CORNELL UNIVERSITY LIBRARY ENGINEERING TN 780.87"""" ""''"^^''y Library .g^toy Of copper. Cornell University Library The original of tiiis book is in tine Cornell University Library. There are no known copyright restrictions in the United States on the use of the text. http://www.archive.org/details/cu31924004606608 METALLUEGY OF COPPER McGraw-Hill DookCompaiiy Electrical World TheEtigineeiTiig andMiiun^ Journal Engineering Record Engineering News Railway Age Gazett€> American Machinist Signal Enginc^er AmericanEngjneer Electric Railway Journal Coal Age M(?tallurgical and Chemical Engineering Power ■MIllMllM ■■Ill I ll iii iiill l^ & BOOKS BY H. O. HOFMAN PROFESSOR OF METALLURGY, MASSACHUSETTS INSTITUTE OF TECHNOLOGY NOW READY GENERAL METALLURGY 909 Pages, 6X9, 836 iUustrations, . . Cloth, $6.00 METALLURGY OF -COPPER 556 Pases, 6X9, 548 illustrations Cloth, S.OO METALLURGY OF LEAD AND THE DESILYERIZATION OF BASE BULLION— Filth Edition—Ninth Impression, 1908 551 Pages, 6X9, 441 Illustrations, qiolh, 6.00 IN PREPARATION- METALLURGY OF LEAD METALLURGY OF MINOR METALS METALLURGY OF GOLD AND SILVER METALLURGY OF COPPER BY H. O. HOFMAN, E. M., Met. E., Ph. D. PROFESSOR OP METALLURGY IN THE MASSACHUSETTS INSTITUTE OF TECHNOLOGY McGRAW-HILL BOOK COMPANY, Inc. 239 WEST 39TH STREET, NEW YORK 6 BOUVERIE STEEET, LONDON, E. C. 1914 Copyright, 1914, by the McGraw-Hill Book Company, Inc. THE. MAPLE. PRESS. TOUK. PA TO THE Pernor; of my Dk. WILHELM HAMPE, LATE PROFESSOR OF CHEMISTRY AT THE ROYAL PRUSSIAN MINING ACADEMY OF CLAUSTHAL. ESTEEMED BY METALLURGISTS FOR HIS CLASSICAL RESEARCHES UPON COPPER PREFACE My aim in preparing this book has been to furnish a treatise on copper which will meet the demands of the metallurgist of to-day. In order to do this, it has been necessary: to present the leading physical and chemical facts about the metal, its alloys, and its compounds which are of metallurgical importance; to record from the older practice that which is of lasting value; and to give the details of the present modes of operating. Though we have several books treating of certain branches of the metallurgy of copper, such as the valuable volumes of Peters, Greenawalt, and others, there does not exist a modern book which covers the entire ground as the present work aims to do. In the study of processes there are given the principles and the practice. The discussion of principles is confined to the essential points; for an ex- tended presentation the reader is referred to my treatise on General Metallurgy. The examples of practice are drawn mainly from the United States. The text and footnotes will show that the technical literature on copper has been covered. The details of practice, not recorded in print, have been obtained through visits to and detailed studies of the leading copper smelteries and refineries of the United States; additional information has been available through correspondence. The tables giving the working data of the leading plants of the United States, Canada, Mexico, Germany, and Australia show the extent of the field which has been covered. In all my visits and correspondence I have met with the greatest cor- diality and liberality. I wish to express here my obligation to the heads and officers of the different smelteries and refineries for the assistance they have rendered me in my task; without this cordial and liberal aid it would not have been possible to prepare this treatise. In working out the detaOs, especially in the large number of calculations, I have been much assisted by the collaboration of my colleague Professor C. R. Hayward. The reading of the page-proof by Professor G. A. Roush detected errata which had escaped my own revision. H. O. HOFMAN. Massachusetts Institute of Technology, Boston, Mass., June, 1914. CONTENTS Page Prepa.ce V CHAPTER I Introduction I. Historical Notice, i; 2. Statistics, 2; 3. Bibliography, 3. CHAPTER II Properties of Copper . . . . . 4 4. Physical Properties, 4; 5. Chemical Properties, g. CHAPTER III Copper or Commerce, Its Impurities and Their Effects 10 6. Grades of Copper, 10; 7. Impurities and their ESfects in General, 12; 8. Oxygen, 13; 9. Lead, !■;; 10. Bismuth, 16; 11. Iron, 16; 12. Manganese, 17; 13. Nickel, 17; 14. Cobalt, 18; 15, Arsenic, 18; 16. Antimony, ig; 17. Sulphur, 21; 18. Selenium and Tellurium, 21; 19. Silver, 22; 20. Lead-silver, 22; 21. Gold, 23; 22. Minor Metals, 23. CHAPTER IV Industrial Alloys 24 23. Industrial Alloys in General, 24; 24. Phosphor Copper, 25; 25. Silicon copper, 26; 26. Brass in General, 28; 27. Regular Brass, 34; 28. Special Brass, 36; 29. Bronze in General, 3g; 30. Regular Bronze, 45; 31. Special Bronze, 46; 32. Alumi- num Bronze, 48; 33. Minor Alloys, 50. CHAPTER V Copper Compounds . . 51 34. Cuprous Oxide, 51; 35. Cupric Oxide, 52; 36. Cupric Carbonate, 52; 37. Copper Silicates, 52; 38. Copper Sulphides, 53; 39. Cupric Sulphate, 55; 40. Cuprous Chloride, 57; 41. Cupric Chloride, 58. CHAPTER VI Copper Ores, Their Metallurgical Treatment 60 42. In General, 60; 43. Sulphide Copper Ores, 60; 44. Oxide Ores, 61; 45. Native Copper Ore, 62; 46. Marketing, 62; 47. Metallurgical Treatment in General, 62. CHAPTER VII Smelting of Copper ... 63 48. Smelting of Copper Ore in General, 63. A. Smelting Sulphide Copper Ore 49. Smelting Sulphide Copper Ore in General, 63 ; 50. Smelting Sulphide Copper in the Blast furnace in General, 64; 51. Roasting and Reduction Process, 64. xi xii CONTENTS Page I. Roasting. 52. Roasting Sulphide Copper Ore, 65; 53. Behavior of Cu-, Fe-, and Mn-sulphides in Powder Form, 65; 54. Behavior of Cu-, Fe-, and Mn-sulpliides in Lump Form (Kernel Roasting), 67; 53. Roasting Apparatus in General, 70; 56. Roasting in Heaps, 71; 57. Roasting in Stalls, 79; 58. Roasting in Shaft Furnaces (Kilns) in General, 82; 59. Roasting Lump Ore in Shaft Furnaces, 82; 60. Comparison of Heaps, Stalls and Kilns, 86; 61. Roasting Fine Ore in Shaft Furnaces, 87; 62. Automatic Furnaces, 87; 63. Shelf-burners, 89; 64. MacDougall Furnace in General, 91; 65. HerreshofE Furnace, 92; 66. O'Brien Furnace, 95; 67. Evans- Klepetko Furnace of Great Falls, 97; 68. Evans- Klepetko Furnace at Other Smelteries, 105; 69. Wedge Furnace, 112; 70. Roasting in Reverberatory Furnaces, 115; 71. Single-hearth Hand Reverbera- tory Furnace, 115; 72. Edwards Furnace, 116; 73. Merton Furnace, 121; 74. Ropp Furnace, 122; 75. Wethey Straight-line Single- hearth Furnace, 123; 76. AUen- O'Hara Furnace, 124; 77. Wethey and Keller Multiple-hearth Furnaces, 126; 78. Brown Horseshoe Furnace, 129; 79. Pearce Turret Furnace, 132; 80. Bruckner Fur- nace, 135; 81. Roasting in Automatic and MuflBe-furnaces, 138; 82. Blastroasting in General, 138; 83. Examples of Up-draft Blastroasting, 139; 84. Examples of Down-draft Blastroasting, 143; 85. Summary of Roasting, 147. II. Smelting in the Blast furnace. 86. The Blast furnace and its Accessory Apparatus in General, 148; 87. Blast- furnace in General, 149; 88. HerresholT Furnace, 149; 89. Blast furnace Building, 152; 90. Great Falls, Anaconda, Cananea, and Mount Lyell Furnaces, 15 2; 91. Hearth, 157; 92. Shaft, 161; 93. Feeding of Charge and Withdrawal of Gases, 165; 94. Fore-hearth or Settler, 169; 95. Disposal of Waste Slag, 173; 96. Disposal of Matte, 17s; 97. Blast furnace Table, ^^(l.• (a) Reducing Smelting. 98. Reducing Smelting in the Blast furnace of Roasted (Raw) Sulphide Ore (Raw Smelting, Matting), 178; 99. Blast furnace Slag in Reducing Smelting, 179; 100, Fuel and Blast, 182; loi. Chemistry of Reducing Smelting, 183; 102. Calculation of Charge, 185; 103. Management and Results, 189. (b) Pyritic Smelting. 104. Pyritic and Partial Pyritic Smelting, 189; 105. Pyritic Smelting Proper, 192; 106. Fuel and Blast, 193; 107. Chemistry of Pyritic Smelting, 194; 108. Manage- ment and Results, 196; 109. Knudsen Process, 196. (c) Partial Pyritic Smelting. no. Partial Pyritic Smelting of Raw Sulphide Ore for Matte, 198; in. Slag, 198; 112. Fuel and Blast, 199; 113. Chemistry, 201; 114. Management and Results, 202; 115. Calculation of Charge, 202; 116. Thermal Balances of Some Partial Pyritic Smelting Operations, 204; 117. General Smelting Operations, 210. (d) Products. 1x8. Productsof the Blast furnace, 212; 119. Matte, 212; 120. Speise, 216; 121. Slag, 217; 122. Gases and Flue-dust, 218; 123. Treatment of Flue-dust, 232; 124. Hearth Accretions (Sows), etc., 235; 125. Results, 235; 126. Production in the Blast furnace of Metallic Copper from Matte, 236. III. Smelting in the Reverberatory Furnace. 127. Smelting in the Reverberatory Furnace in General (Welsh Process), 236; 128. Reverberatory Matting Furnace in General, 237; 129. Reverberatory Furnace in Detail, 238; 130. Furnace of the Colorado Smelting Co., Butte, Montana, 238; 131. Furnace of the Anaconda Copper Mining Co;, Anaconda Montana, 243; 132. Furnace of the Canadian Copper Mining Co., Copper Cliff Ontario, 245; 133. Furnace of the Cananea Consolidated Copper Co., Cananea Sonora, Mexico, 247; 134. Furnace of the Anaconda Copper Mining Co., Great Falls' CONTENTS xiii Page Montana, 233; 135- Other Furnaces, 255; 136. Accessory Apparatus, 255; 137. Dimensions and Worlcing Data of Some Reverberatory Furnaces, 255; 138. Refer- ences to Some Reverberatory Smelteries, 255; 139. Working Bottom, 260; 140. Fuels, 265; 141. Bituminous Coal and Producer Gas, 265; 142. Fuel-dust, 267; 143. Wood and Producer Gas, 268; 144. Oil, 269; 145. Slags, 27s; 146. Chemistry, 278; 147. Calculation of Charge, 279; 148. Heat Balance of Reverberatory Matting Fur- nace at Anaconda, 280;. 149. Management, 280; 150. Accessory Apparatus, Products, Losses, Cost, 286; 151. General Arrangement of MacDougall and Reverberatory Plant, Copper Queen Consolidated Mining Co., 286; 152. Comparison of Blast and Reverberatory Furnace for Matting, 287; 153. Production in the Reverberatory Furnace of Metallic Copper from Matte, 288; 154. Ordinary Process, 288; 155. Extra or Selecting Process, 293; 156. Argo Process, 295. IV. Smelting in the Converter. 157. Converting Copper Matte in General, 298. (a) Converting in Vessel with Acid Lining. 158. The Converter, 300; 159. Upright (Manh&) Converter, 300; 160. Hori- zontal (David-Manh6s) Converter, 305; 161. Upright vs. Horizontal Converters, 308; 162. Ma'nipulation, 309; 163. Lining, 310; 164. Blast, 314; 165. Acces- sories, 314; 166. Arrangement of Plant, 320; 167. Mode of Operating and Chemistry, 323; 168. Eliminationof Impurities, 326; 169. Thermal Balance, 327; 170. Products, Losses and Cost, 331. (b) Converting in Vessel with Basic Lining. 171. Basic Converting in General, 331; 172. Peirce-Smith Converter, 335; 173. Products, Losses, and Cost, 342; 174. Concrete Examples, 343; 175. Anaconda, Basic Converter, 34s; 176. Great Falls Basic Converter, 349. V. The Sulphide Copper Smelting Plant. 177. General Arrangement of Plant, 353; 178. Cost of Plant, 357. B. Smelting Oxide Copper Ores 179. Smelting Oxide Copper Ores, 358; 180. Early Work in Arizona, 359- C. Smelting Native Copper Ore 181. Ore, 361; 182. Process, 362; 183. Reverberatory Furnace, 362; 184. Mode of Operating, 368; 185. Blast-furnace, 370. D. Fire-refining of Impure Copper 186. Introductory, 372; 187. Furnace, 372; 188. Mode of Operating in General, 381; 189. Charging, 381; 190. Melting, 382; 191. Fining (Rabbling), 382; 192. Poling, 384; 193. Examples of Refining, 391; 194. Casting, 395. CHAPTER VIII Leaching of Copper 4°^ A. Leaching Copper Ore 195. Leaching Copper Ore in General, 402; 196. Solvents, 402; 197. Precipitants, 404; 198. Leaching Apparatus and Method, 406; 199. Precipitating Vat and Method, 406; 200. Outline of Leaching Processes for Ore, 406; 201. Leaching Sulphate Ore, 407; 202. Mine Waters, 407; 203. Mill Tailings and Mine Dumps, 413; 204. Leaching Oxide Ore, 414; 205. Laszcynski Process, 414; 206. Carmichael Process, 415; 207. Neill and Van Arsdale Processes, 418; 208. Stadtberge Process, 419; 209. Greenawalt Process, 419; 219, Leaching Atacamite, 420; 211. Leaching Sulphide Ore after xiv CONTENTS Page Conversion into Sulphate by Weathering, 420; 212. Leaching at Rio Tinto, 421; 213. Leaching Ore in Place, 424; 214. Leaching Sulphide Ore after Conversion into Sulphate by Sulphatizing Roasting, 424; 215. Sulphatizing Heap-roasts, 424; 216. Sulphatizing Muffle- roasts, 427; 217. Leaching Sulphide Ore after Conversion into Sulphate by Ferric Sulphate, 430; 218. Siemens-Halske Process, 430; 219. Leach- ing Sulphide Ore after Conversion into Oxide by Roasting, 43 1 ; 220. Leaching Sul- phide Ore after Conversion into Chloride by Ferric Chloride, 432; 221. Doetsch and Froelich Processes, 432; 222. Leaching Sulphide Ore after Conversion into Chloride by Cupric Chloride; . Hoepfner Process, 434; 223. Leaching of Sulphide Ore after Oxidizing and Chlorinating with Ferrous or Calcium Chloride, 434; Hunt and Douglas Processes, 434; 224. Leaching of Sulphide Ore after Chloridizing Roasting, 436; 225. Leaching of Sulphide Ore after an Oxidizing Followed by a Chloridizing Roast, 436; Longmaid-Henderson Process, 436; 226. Crushing and Mixing of Ore and Salt, 438; 227. Chloridizing Roasting and Condensation of Gases, 441; 228. Leaching Chloridized Ore with Water and Tower Liquor, 450; 229. Clarifjdng Copper Liquor, 452; 230. Precipitation of Copper by Iron, 452; 231. Washing and Refining of Cement Copper, 453; 232. Disposition of Residue from Leaching and of Waste Liquor, 453; 233. Precipitation Copper Independently of Silver and Gold, 454; 234. Results and Cost, 455. B. Leaching Copper Matte 235. Leaching Copper Matte in General, 456; 236. Augustin Process, 456; 237. Ziervogel Process, 456; 238. Ziervogel Process at the Gottesbelohnung Works, Mans- feld, 457; 239. Freiberg Vitriolization Process, 463; 240. Hofmann Vitriolization Process, 464. C. Leaching Metallic Copper 241. Leaching Metallic Copper in General, 471; 242. Augustin Process, 472; 243. Vitriolization Process, 472; 244. Examples of Vitriolization, 475. CHAPTER IX Electrolysis of Copper ^82 245. In General, 482; 246. Electrolysis of Ore, 482; 247. Electrolysis of Matte, 482; 248. Electrolysis of Speise, 483; 249. Electrolysis of Metallic Copper in General, 483; 250. Behavior of Individual Impuritie^, 485; 251. Current, 483. A. Multiple System 252. Multiple Systemin General, 491; 253. Electrolyte, 492; 254. Current, 497; 255. Anode, 499; 256. Cathode, 502; 257. Manipulation, 505; 258. Depositing Vat, 506; 259. Corrosion of Anode, 511; 260. Deposition on Cathode, 511; 261. Anode Mud, 513; 262. Treatment of Anode Mud, 515; 263. Recovery of Tellurium and Selenium, 518; 264. Foul Solutions, 520; 265. Cost, 524; 266. Examples of Multiple Process, 525- B. Series System 267. Series System in General, 534; 268. The Baltimore Plant, 536; 269. Electrolyte, 536; 270. Current, 536; 271. Electrodes, 536; 272. Depositing Vat, 537; 273. Corro- sion and Deposition, 537; 274. Clean-up, 537. C. Multiple versus Series System 275. Multiple and Series Systems Compared, 538. Index , , , , , 541 METALLURGY OF COPPER CHAPTER I INTRODUCTION I. Historical Notice.' — Copper is a widely distributed metal, and next to iron, it is the most important.^ In preliistoric times it was used for domestic utensils and for implements of war.' In the archoeological chronology of the stone, bronze, and iron ages, it has been supposed that the use of copper always preceded that of iron; at present it is held that generally iron was converted to use at an earlier period than copper, although in regions in which native copper occurred at the surface it was worked before iron, which had first to be reduced to the metallic state. Thus the races inhabiting this continent* before the Indian were acquainted with copper, as is shown by the utensils found in Western mounds. In opening the Lake Superior copper mines, excavations to the depth of 60 ft. were encountered containing stone hammers and charcoal indicating that fire-setting had been the method of winning native copper. At Ducktown, Tenn.,* are remains of pfehistoric smelting operations. Copper was mined first in Connecticut (1709), later in New Jersey and Pennsylvania, but no work of any importance was carried on until the middle of the last century. In 1845 the whole output of copper was 100 tons, and this came from Vermont, Pennsylvania, Virginia, North Carolina, and Georgia. California became a producer in i860. In 1841 the existence of native copper at Lake Superior became generally known, but copper mining as an industry was not established until 1850. Lake Superior was the leading copper producer up to 1887, when it was outranked by Montana, where copper mining began early in the seventies,^ and smelting about 1880.'' The third great copper ' Beck, L., "Die Geschichte des Eisens," Vieweg, Brunswick, 1891, i, Introduction, p. 17. Agricola — Hoover, "De Re Metallica," London, 1913, p. 402, Eng. Min. J., 1913, xcvi, 359.. "Douglas, Eng. Min. J., 191 2, xcin, 776. » Garland, "Metallograph. Res. Egyptian Metal Antiquities," /. Inst. Met., 1913, x, 329. < Foster, J. W. and Whitney, J. D., "Report on the Geology and Stratigraphy of Lake Superior," Interior Department Land OiEce, Washington, 1850 and 1851. Whitney, J. D., "Metallic Wealth of the United States," Lippincott, Philadelphia, 1854. "Mineral Statistics of Michigan," 1880. Kirchhoff, " Mineral Resources of the United States," 1882, p. 213. Tr.A.I. M. E., 1876- 77, V, i6s (Hewitt); 1890-91, xix, 679 (Douglas); 1906, xxxvii, 288 (Wood, chemical anal- yses); Min. Ind., 1894, iri, 243 (Douglas, chemical analyses); 1895, iv, 269 (Douglas). ' Henrich, Tr. A. I. M. E., 1895, xxv, 175. « Weed, H. V., Professional Paper No. 74, U. S. Geol. Survey, 1912. ' Hofman, Tr. A.LM. E., 1904, xxxiv, 258. I 2 METALLURGY OF COPPER district is in the Southwest, including New Mexico, Arizona,* and Lower Cali- fornia. The existence of copper in this region was known to the Mexicans," but active operations began after 1880 with the building of the Southern and the Atlantic & Pacific Railroads. In recent years Utah, Nevada, California, and some of the Southern states have entered the list of important producers of the country. 2. Statistics.— The world's production' of copper in 1912 is given in Table I, that of the United States'* in Table 2. Table i. — World's Production of Copper (In metric tons) Africa 16,633 Mexico 73,6i7 Argentina 33S Newfoundland 549 Australasia 47,774 Norway 11,156 Austria 4,024 Peru 26,483 Bolivia 4,681 Russia 33.55° Canada 34,2i3 Spain, Portugal 59,876 Chile 39.204 Sweden 1,524 Cuba 4.393 Turkey 508 Germany 24,304 United Kingdom 405 Italy 2,337 United States 563,260 Japan 62,486 Total 1,011,312 Table 2. — United States' Production o* Copper (In pounds) Alaska 2,602,000 Nevada 82,530,608 Arizona 357,952,962 New Mexico 27,488,912 California 31,069,029 Utah 131,673,803 Colorado 7,502,000 Wyoming 1,121,109 Idaho 5,964,542 Southern States 18,592,655 Michigan 231,628,486 Other States 4,396,667 Montana 309,247,735 Total 1,241,762,508 Table i shows that the United States produced about 55 per cent, of world's copper. The leading producing states, as given in Table 2, are Arizona, Mon- tana, Michigan, and Utah, which furnish 83 per cent, of the country's product. There are sold in the American market three grades of copper. The average price of Lake copper in New York for 1912 was 16.56 cents per pound; of electrolytic copper, 16.34 cents; casting copper is usually i to 2 cents per pound less than Lake copper. The costs of producing copper in different parts of the United States have been analyzed by Steele.^ He finds that the average cost for the country lies between 9.6 and 10. o cents per pound. ' Martin, Eng. Min. J., 1913, xcv, 882. 2 Wendt, Tr. A. I. M. E., 1886-87, XV, 25. ' Min. Ind., 1912, xxi, 168. *0p. cit., 1 60. '£»g. Min, J., 1913, xcvi, 251. INTRODUCTION 3 3. Bibliography. — The number of books dealing with copper alone is com- paratively small, and none of them covers the entire ground. Usually the subject is discussed in treatises on non-ferrous metallurgy. Thus, the works of Percy (1867), Kerl (1881), Balling (1885), Schnabel-Louis (1905), Hildebrandt (1906), and Prost (1912) contain valuable discussions of the subject. There is subjoined a list of books dealing exclusively or mainly with copper. Griiner, L., "Metallurgie du Cuivre," Paris, 1884. Howe, H. M., "Copper Smelting," Bull. 26, U. S. Geol. Survey, 1885, now out of print. Peters, E. D., " Modern Copper Smelting," New York, 1895. Lang, H., "Matte Smelting," New York, 1896. Peters, E. D., "Principles of Copper Smelting," New York, 1907. Trochu, P., "Les Pyrites," Paris, 1907. Hixon, H. W., "Notes on Lead and Copper Smelting and Converting," New York, 1908. Venancourt, G. C. de, "Le Water-jacket a Cuivre," Paris, 1910. Peters, E. D., "Practice of Copper Smelting," New York, 1911. Greenawalt, W. E., "The Hydrometallurgy of Copper," New York, 191 2. Ulke, T., " Modern Electrolytic Copper Refining," New York, 1903. The subject of copper alloys is not taken up in these works. There exist many valuable monographs and papers on the various phases of the metallurgy of copper; these are referred to in the text. CHAPTER II PROPERTIES OF COPPERS 4. Physical Properties. — The metal occurs in the native state. The specific gravity of pure copper at 20° C. is 8.89;^ good commercial metal shows lower values owing to porosity,' the presence of CU2O/ and impurities. The specific gravity of molten metal is given as 8.22. The luster of the compact metal is metallic, while precipitated metal is dull. The color of compact metal is a yellowish-red; it ranges from orange-red to rose-color, the shades being governed by the temperature of the cooling-water in which the casting has been quenched. Metal precipitated from solution is a brownish-red; a colloidal solution has a violet to brownish color.^ Copper is transparent in thin films transmitting greenish to bluish light. Fig. I. — Surface of cast copper, 30 diameters. Fig. 2. — Surface of electrode- posited copper, 30 diameters. Fig. 3. — Surface of rolled copper, 30 diameters. Copper crystallizes mainly in isometric forms, and twin crystals are common. Under the microscope^ the surface' of cast copper (Fig. i) is seen to be made up of large primary grains, composed of small secondary grains with definite orienta- tion; that of electrodeposited copper (Fig. 2), of primary grains only, which have > Hampe, Zl. Berg. Hiiiten. Sal. Wesen i. Pr., 1873, xxi, 218; 1874, xxii, 93; 1876, xxiv, 6. 2 Circular 31, Bureau of Standards, 1912, p. 61. ' Stahl, Berg. Hiittenm. Z., 1901, lx, 77. ' Trippel, Eng. Min. /., 1888, xiv, 436. * Rassenfosse, /. Soc. Chem. Ind., 1911, xxx, 1335. ' Baucke, Proc. Internal. Assoc. Testing Materials, 1912, 11, 14. Bassett, Met. Chem. Eng., 1913, xi, 64. ' Campbell, Report Alloys Research Comm., 1904, p. 867; /. Frankl. Inst., 1902, CLrv, 14; Metallurgie, 1907, iv, 828. Addicks, Electrochem. Ind., 1903, i, 582. Huntington, Eng. Min. /., 1905, Lxxx, 1109; Metallurgie, 1906, iii, 40. Abbott, Eng. Min. J., 1909, Lxxxvii, 1040. » Faust, Zt. anorg. Chem., 1912, lxxviii, 201; /. Inst. Met., 1913, rx, 223. Waser-Kuhml, Electrochem. Zt., 1912, xviii, 151, 211. 4 PROPERTIES OF COPPER S no regular orientation to one another. With rolled copper (Fig. 3) , the secondary grains are elongated in the direction in which rolling took place, and this gives the metal the characteristic fibrous structure.' The so-called allotropic copper of Schiitzenberger^ has been found by Benedicks' to consist of ordinary copper containing varying amounts of CujO. The dilatation experiments of Turner and Levy* on hard-drawn and annealed wire give simple curves without any jog whatever. The fracture of cast copper is hackly to granular; that of rolled or drawn copper, fibrous. 60 50 s i 40 I S 30 ' CL. I 20 10 ■^ • > \ \ E lon£ atio 1 % in In< lies \ .•' .^- '■- "~ ._. "" — --- ■^- \ / ^^^ ^v 1 X Jltirr ate Sh3 iBth \ \ \ \ N \ V E asti •Ui lit ,^^ -"~~" -^ 100 200 300 400 500 600 700 800 900 1000 Temperature of Annealins - Degrees Centierade Fig. 4. — Mechanical properties of electrolyte copper as affected by temperature. Copper is soft when pure; with Fe at 4.5 and Ag at 2.7, the scratch-hardness^ of Cu is 3.7. The tensile strength of the cast or hard-drawn metal is 60,000-70,000 lb. per square inch; annealing reduces it to 30,000-40,000 lb.* Shock-tests by Baucke' show that cast copper has a very low resilience, and that the property is improved by forging. A rise in temperature weakens the metal. ^ Fig. 4, ' Robin, "Annealing of Copper," Rev. Mel., 1913, x, 750. ^ Compt. rend., 1887, lxxxvi, 1240, 1397. e,-i907, IV, 5, 33. 'Proc. Roy. Soc, Ser. A., 1907, lxxx, i; Rev. Mil. Exlr., 1908, LV, 655. ' Martens, Mitth. Kgl. techn. Versuchsanst, 1894, xvi, 172; Iron Age, 1894, ltv, 900. "Bennet, "Tensile Strength of Electrolytic Copper," Tr. Am. Electro. Chem. Soc, 1912, XXI, 243. Met. Chem. Eng, i9i2,x, 298. ''Internal. Zt. Metallographie, 1912, in, 195; /. Inst. Met., 1913, ix, 210. ' Grard, Rev. Mit., 1909, vi, 1069. Leidig, Verh. Ver. Beford. Gewerbefl., 1911, xc, 459, 525. Johnson, Met. Chem. Eng., 1911, rx, 399. Huntington, /. Inst. Met., 1912, viii, 127. MuUer, Metall-Erz, 1913, x, 220. METALLURGY OF COPPER by Grard, shows the mechanical changes electrolytic copper undergoes with a rise in temperature. The pure metal is easily rolled into sheets/ hammered into foiP and drawn into fine wire.' The hardness caused by mechanical treatment is removed by annealing* at a temperature ranging from 500° to 700° C. in an atmosphere free from S.* Welding* by ordinary means is possible to small extent; pieces are easily joined by electric welding. Molten copper absorbs'' SO2, H, and CO (disputed by Sie- verts), but no CxH„; upon solidi- fication most of the absorbed gases are given off, at least at atmospheric pressure.* The ab- sorbing power rises with the 1 Powe, Brass World, 1909, I, 183. Copperman, Met. Ind., 1909, vii, 4, 64, 99, 134. 2 Fuller, J., "Art of Copper Smith- ing," Spon, London, 1912. ^ Kiipper, Zt. Verein deutsch. Ing., 1906, L, 1899, 2022; Rev. Mil. Extr., 1907, IV, 722. Pye, /. Inst. Met. 1911, vi, 165; CircularNo. 31, "Copper Wire Tables," Bureau of Standards, 1912. 60000 Con. duottriti 100-2 1 t) 20 30 40 50 80 70 f Percent Reduction in Arela < SSOOO lOO'O / 56000 09.8 ^ 1/ 51000 99-6 V® # 62000 99.4 V A '/ 50000 99-2 ' % 4 "/ 43000 99-0 K -*-i f 40000 98-8 1 44000 08-0 / \ 42000 98-4 ~i Y/ \ \ 40000 98-2 rS 7 ^ "•*■, 38000 98-0 / f ^ ^ 36000 97-8 / \ 34000 97-6 f 33000 97.4 30000 07-2 A.nne£iled Copper Bod of CLSerent Sizes Brawn to No.l2 B,it Sj Fig. s- — Electric conductivity of copper as affected by mechanical treatment. < Howe, Tr. A. I. M. E., 1884-85, xiil, 646. Cummins, Eng. Min. J., 1890, L, 216. Thomas, Iron, 1892, XL, 399. Heyn, /. I. and St. I., 1902, 11, 745. Stahl, Metallurgie, 1908, v, 289. "Johnson, Met. Chem. Eng., 1911, ix, 187. ' Waite, "Leibe Process," Eng. Min. J., 1890, lxdc, 705. McRoberts, "Birmington Process," Iron Age, 1891, XLvm, 1156. ' Carori, Compt. rend., 1866, ucm, 11 29. Hampe, Zt. Berg. Hiitten. Sal. Wesen i. Pr., 1873, xxi, 274; Chem. Z., 1886, xvii, 1692. Stahl, Berg. Iluitenm. Z., 1886, XLV, 414; 1889, XLix, 299; 1893, Lil, 19; 1901, LX, 77. Metallurgie, 1907, iv, 761. Sieverts, Ber. deutsch. chem. Ges., 1910, XLiii, 893; 1912, xLV, 221; Zt. Electrochem., 1910, XVI, 707; Zt. phys. Chem., 1911, lxxvii, 591. Guichard, Compt. rend., 1911, CLIII, 104; /. Inst. Met., 1911, vi, 329. 'SO2: Schenck-Hempelmann, Meiall-Erz, 1913, x, 28. Stubbs, /. Soc. Chem. Ind., 1913, xxxii, 311. PROPERTIES OF COPPER 7 temperature, and is interfered with by CujO, P, As, and Sb. At a red heat copper readily absorbs H.' The melting point of copper is 1083° C; the latent heat of fusion 43.3 cal.; the boiling point 2310° C* In vacuo volatilization^' is noticeable at 700° and decided near the melting point.* When volatilized by heating with the oxy hydrogen blowpipe or the electric arc, it burns with a green flame; the fumes are poisonous.' The specific heat at 170° C. is 0.09244; at 100°, 0.09422; 101 100 I 97 96 95 94 ^ ^ \ ' /' ^ '' ■— = - — — d -^ Te s i "^ """"■ — ■ — . s ^ '^ — == ^ ,__ Pb ^ '^ \' \ ^ ■ —- ■4T ■^ _ca ^ ~ ^nir \ s N • ■— =-- \ ^. V <- \\ V s \ ■v \ > \ \ 5 V. \ \ \\"^ \ V \ \ , V \ \ "13 \ ^' \ \ \ \ \ \ I \ \ \ .01 .02 .03 .04 .05 .06 .07 .08 .09 .10 .11 .12 .13 .14 .15 .16 .17 .18 .19 .20 Percentage of Impurity Fig. 6. — Electric conductivity of copper as affected by impurities. at 300°, 0.09846. The coefficient of linear expansion is 0.000017. The thermal conductivity is 736 when Ag=iooo, or 0.72 g. cal. per degree centigrade for a cube whose side is i cm. The electric conductivity® of i cm. cube at 0° C. = 620,000 reciprocal ohms; ' Heyn, Zt. Ver. deutsch. Ing., 1900, xliv, 508; Metallographist, 1903, vi, 48; Metallurgie, 1906, in, 82. Sieverts, Zt. phys. Chem., 1911, lxxvii, 591; /. Inst. Met., 1911, vi, 342. Heath, "Estimation of Oxygen and Occluded Gases in Copper," /. Ind. Eng. Chem., 1912, IV, 402. ' Greenwood, Eng. Min. J., 1911, xcii, 3. ' Hughes, /. Inst. Met., 1912, vii, 700. * Kahlbaum, Berg. Hiittenm. Z., 1898, lvii, 201; 1902, ixi, 295. 'Hansen, Met. Chem. Eng., 1911, ix, 67. • Wolff, F. A., and Dellinger, J. H., Bureau of Standards, vii, No. i (reprint No. 148), 1911 : "The Electric Conductivity of Commercial Copper." Northrup, Resistivity between 20 and 1450° C, /. Franklin. Inst., 1914, clxxvii, 1. 8 METALLURGY OF COPPER or the resistance of a wire ift. long ando.ooi in. in diameter at o°C.= 9.529 ohms for annealed and 9.71 for hard-drawn wire. The conductivity of cast copper is about 3.5 per cent, lower than that of annealed wire. In smelting works the conductivity is usually given in terms of the Mathiessen Standard. This standard is equal to copper which at 15° C. has a resistance of 1687 reciprocal ohms per cubic centimeter, or i meter-gram of pure soft copper at 0° C.= o. 141 7 2 international ohm. The standard is represented by the figure 100; cathode copper not melted has shown 103.14; mass copper from Lake Superior 102.5; electrolytic wire bar often reaches loi ; Lake copper usually is 99.1 Tests are usually made on annealed wire, No. 12 B. & S. gauge ( = 0.08081 in. in diam- eter). The conductivity as affected by mechanical treatment is shown in the diagram of Addicks^ (Fig. 5). If the conductivity is to be given for hard-drawn 0.0045 0.095 0.0043 0.1 0.0031 0.025 3.4 99i6 99.8 100.0 100.2 100.4 100.fi 100.8 Conductivity Fig. 7. — Electric conductivity of copper as affected by Oxygen, Arsenic and Antimony, severally and combined. wire, it is customary to deduct 2.5 per cent, from the figure obtained with annealed wire. Small amounts of impurity have a decided influence upon the conductivity of copper. Fig. 6 gives some of the experimental results of Addicks.' Arsenic and antimony are the two impurities likely to be found in refined copper which strongly depress the electric conductivity; thus 0.0013 .per cent. As* or 0.0071 per cent. Sb lower it i per cent., while the elements ' Table II, Wolff-Dellinger, op. cit. ^ Electrochem. Ind., 1903, i, 581. ^J. Frankl. Inst., 1905, CLX, 425; Tr. A. I. M. E., 1906, xxxvi, 18. Other data: Keller, Min. Ind., 1898, vn, 243, complete analyses with tensile strengths and electric conductivities. ■■ Hiorns-Lamb, /. Soc. Chem. Ind., 1909, xxvra, 451, curve 1-3 per cent. As. Friedrich, Metallurgie, 1908, v, 533, curve 1-12 per cent. As. Puschin-Dischler, Am. Chem. Soc. Chem. Abstracts, i9€2, vi, 1587; Zt. anorg. Chem., 1913, ixxx, 6s, curve 1-45 per cent. As. PROPERTIES OF COPPER 9 which render copper brittle appear to have little effect upon the electric prop- erties.i Fig. 7 shows the combined effects of As and Sb in Montana electrolytic copper within a range of 0.0034 and 0.0044 per cent. The curve for O in Fig. 4 is abnormal, because CujO reduces the conductivity progressively.^ It finds its probable explanation in the circumstance that introduced as CuaO or CuO into the Cu, oxidizes the slight amount of impurity present in high-grade metal and thus increases the conductivity of the latter. The behavior of a small amount of impurity depends greatly upon the form in which it is present; if in the metallic state, it is liable to form a solid solution with the Cu, and have a greater depressing effect than if present in some other form. Such irregularities are well brought out in Fig. 7. Copper is diamagnetic. 5. Chemical Properties. — At ordinary temperature copper is not attacked by dry air nor by moist air free from CO2; in the presence of this gas it becomes coated with a green basic carbonate.* The chemical theory of corrosion has been in part replaced by the electrolytic theory.* When heated above 185° C. copper begins to oxidize, becomes rose-colored at about 200°, brass-colored at 300°, bluish-green at 350°, and dark above that temperature. At a red heat it becomes coated with a dark scale consisting on the outside of CuO and on the inside of CU2O; the scale is separated from the metal by bending and quenching. According to Heyn^ copper heated for a short time above 500° C. withstands a smaller number of bends than when heated below this temperature because of the CU2O that has been formed. It is clear that overheating^ a cake of copper which causes a superficial oxidation will affect the sheet that is rolled from it. Copper is readily soluble in HNO3, when not too concentrated; in aqua regia; boiling H2SO4 of 66° Be; slowly soluble in hot dilute H2SO4 in the presence of air; in dilute HCl with air; in NH4OH with air; in KCN with or without air ' H2SO3 slowly changes Cu into CuS. 1 Bardwell, Tr. A.I.M. E., 1913, xlvi. ' Walker, Min. Ind., 1898, vii, 248. Hofman-Hayden-Hallowell, Tr. A. I. M. E., 1907, xxxvnr, 178, 183. ' Diegel, Zt. Verein Bef. Gewerbefl., 1899, Lxxviii, 313; 1903, lxxx, 93, 119, 157. Heyn, Mitih. Kgl. Materialpriifungsansialt, igii,xxix, 29. Eastick, Met. Ind., I9i3,xi, 524. * Bengough, /. Inst. Met., 1911, v, 28. Philip, op. cit., 1912, VII, so; 1913, nc, 61. ' Mitth. Kgl. techn. Versuchsanst., 1900, xviii, 327; Zt. Ver. deutsch. Ing., 1902, xxxvi, 1119; Stahl u. Eisen, 1902, xxii, 1234. ' Stahl, Metallurgie, 1908, v, 289; 1912, rx, 418. CHAPTER III COPPER OF COMMERCE, ITS IMPURITIES AND THEIR EFFECTS 6. Grades of Copper.' — In the United States there are marketed three grades of copper: electrolytic, Lake, and casting copper, which are cast in the forms of wire bar, ingot and ingot bar, and cake. Electrolytic and Lake copper contain over 99.8 per cent. Cu; casting copper as little as 98.5 per cent. Cu. According to tests made by W. H. Bassett^ in 1903-04 an average of 511 samples of best electrolytic copper gave on hard-drawn wire 0.003 in. in diameter, tensile strength 65,259 lb. per square inch, and elongation 1.55 per cent, in 8 in.; on annealed wire, 0.06 in. in diameter, in 6 in. 24.8 twists and 13.6 bends; electric conductivity 100.32 per cent. Math. Stand. An average of 55 samples of best Lake copper gave tensile strength 66,141 lb., elongation 1.45 per cent., twists 22.2, bends 12.2, conductivity 99.85 per cent. Thus as regards physical proper- ties electrolytic copper is preferable to Lake copper. If, nevertheless. Lake cop- per' has been sold at i cent per pound more than electrolytic, the reason is to be found in the uniform character of Lake copper and the irregularities in the prop- erties of electrolytic copper. The tests of Bassett show that the progress made in the electrolytic process and in the fire-refining of cathodes has so improved the character of electrolytic that it stands higher to-day than Lake copper. Casting copper is a general term for fire-refined blister copper too low in pre- cious metals to make their recovery profitable, and carrying impurities in too small a quantity to make them objectionable. Its electric conductivity is too low to make it available for electric use, and the amount of impurity too high for making brass that is to be rolled or drawn; it serves therefore for making brass and copper castings. Table 3 gives the forms in which copper was cast in the United States in 1911.* Table 3. — Forms in which Copper was Cast in the U. S. in 1911 Form Pounds Per cent. Wire bars Ingots and ingot bars. Cakes Cathodes Other forms 731,029,349 so 409,786,682 29 143,716,12s 10 135,499,770 9 25,774,328 2 Total 1,445,806,254 1 Ingalls, Eng. Min. J., 191 2, xciii, 887, 939, also selling of copper. Tassin, Met. Ind., 191 2, x, 275, 335, 447- 2 Records, Circuit Court of the U. S., Bigelow vs. Calumet & Hecla Mining Co., October 17, 1907. ' Eng. Min. J., 1908, lxxxvi, 842. * Min. Res. U. S., 1911, i, 311. 10 COPPER OF COMMERCE II Uses or Copper. — Metallic copper is used in the arts for electrical purposes, and for the manufacture of brass, bronze, and other alloys; it is rolled into sheets and tubings, and formed into castings. Table 4 gives the distribution of copper in the metal arts in 1907.^ Table 4. — Uses of Copper in the U. S. in 1907 Use Pounds Per cent. Electric purposes, including wire 228,000,000 156,000,000 34,000,000 127,000,000 41.8 28 6 Brass manufacture Rolling mills, sheet copper 6-3 23 -3 Miscellaneous use, principally castings and alloys Total 645,000,000 100. Specifications. — The standard specifications for copper- wire bars, cakes, slabs, billets, ingots, and ingot bars adopted by. the American Society of Testing Materials, August 21, 19 11,'' are given in the following: 1. (a) Metal Contents. — The copper in all shapes shall have a purity of at least 99.88 per cent, as determined by electrolytic assay, silver being counted as copper. (b) Conductivity. — ^AU wire bars shall have a conductivity of at least 98.5 per cent, (annealed); all ingots and ingot bars shall have a conductivity of at least 97.5 per cent, (annealed), excepting only arsenical copper, which shall have a conductivity of not less than 90 per cent, (annealed). Cakes, slabs, and billets shall come under the ingot classification, except when specified for electrical use at time of purchase, in which case wire-bar classification shall apply. The "Annealed copper standard," or resistance of a meter-gram of standard an- nealed copper at 20° C, shall be considered as 0.15302 international ohm. The per cent, conductivity for purposes of this specification shall be calculated by dividing the resistivity of the annealed copper standard by the resistivity of the sample at 20° C. 2. Wire bars, cakes, slabs, and billets shall be substantially free from shrink holes, cold sets, pits, sloppy edges, concave tops, and similar defects in set or casting. This clause shall not apply to ingots or ingot bars, in which case physical dfefects are of no consequence. 3. Five per cent, variation in weight or i in. variation in any dimension from the refiner's published list or purchaser's specified size shall be considered good delivery; provided, however, that wire bars may vary in length 1 per cent, from the listed or specified length, and cakes 3 per cent, from the listed or specified size in any dimension greater than 8 in. The weight of ingot and ingot-bar copper shall not exceed that specified by more than 10 per cent., but otherwise its variation is not important. The specifications of the leading copper producers of the world are contained in the report made by Guillet^ at the Copenhagen Congress. The congress ' Min. Res. U. S., 1907, i, 639. * Year-book, 1911, p. 127. ^ Rev. Met., 1909, VI, 1245- METALLURGY OF COPPER of New York in 1912' has only laid out a program for future work. Considering that 95 per cent, of United States copper is made up of electrolytic and Lake brands, both more than 99.8 per cent, pure, the choice of electric conductivity as standard for quality is to be expected. This standard, however, is not suited to most European brands which contain less than 99.8 per cent. Cu, but have excellent wearing qualities owing to the presence of As, Sb, Ni, etc.^ 7. Impurities and Their Effects in General.' — Copper of commerce, as stated above, is not pure. It contains CuaO and foreign metals and their oxides which affect the physical and chemical properties, and thereby the availability for use in the arts. Table 6 gives t3^ical chemical analyses of some leading brands of to-day. They show that the copper produced at present is of a higher grade than that of former years, when copper contents of 97-98 per cent, were not uncommon.^ The total impurity in copper is small, but the number of elements composing it is large; and a small percentage of a single element may Table 5. — Chemical Analysis of Refined Copper Element Lake, wire bar' Lake, arsenical, ingot 1 Electrolytic, wire bari Best selected English! Cu + Ag. Cu Ag Pb Bi As Sb Se + Te. Fe Ni Zn S O (by diS.) , Sn 99.900 99.890 o . 0096 (2.8 oz.) 2 . 003 1 O . 0000 0.0062 o . 0000 0.0020 0.0028 o. 0090 0.0000 0.0016 0.0753 99.438s 99.4131 0.02S4 (7.41 02- ) 0.0027 0.0000 0.3183 0.0000 u. d. 0.0056 O.0IS3 0.0000 0.0071 0.2143 99.970 99.967 U.0027 Co. 79 oz.) U.0024 O 0000 0.0006 0.0000 u.OOOO 0.0023 0.0030 0.0000 o. 0026 0.0I9I 99.89s 99.893 0.0020 o. 0072 0.0000 O.OOOI u. 0006 U. 0022 0.0028 O.OOIO . 0000 U.0023 0.0888 99.9548 99.9S3 u.OOlS O.OOIO . 0000 0.0000 o . 0009 0.0026 0.0038 0.0028 0.0000 0.0026 0.03IS 99.9780 99.976 u,0020 0.0056 0.0000 O.OOOI 0.0008 0.0014 o . 0044 U.00I8 0.0000 0.0016 o . 0063 99S5I0 99 530 0.0210 O.I33I . 0000 0.0071 0.0087 o . 0066 o . 0044 u. III2 0.0000 0.0074 0.I70S Conductivity, annealed Conductivity, hard drawn... Difference due to hard draw- ing. Tensile strength, lb. sq in. . Twists in 6 in Elongation per cent Bends, annealed Diam. of wire, in 96.49 93.84 2.6s 67.590 17 103' II 0.080 100.84 97.93 2.91 6s, 000 18 1. 6s* 14 0.080 99.78 96.6s 3.13 67,800 27 I.ISt 14 0.080 100.45 97.64 2.81 66,300 34 l.04t 14 0.080 1.0070 97.93 2.77 66,550 S3 i.oSf 22 0.080 •In 1 W. 8 in. fin 60 in. H. Bassett. ' Guillet, op. cit., 191 2, ix, 1037. ^ Lewis, Met. CP.em. Eng., 1912, x, 540. ' Hampe, loc. cit. Greaves, J. Inst. Met., 1912, vii, 218. Archbutt, op. cit., 1912, vii, 262. , Johnson, op. cit., 1912, viii, 192; 1913, x, 275. Law, op. cit., 1912, VIII, 222. Tassin, Met. Ind., 1912, xviii, 275, 335, 447. Lewis, Met. Chem. Eng., 191 2, x, 540. Baucke, Int. Zt. Metallographie, 1913, in, 195. * Kerl, B., "Metallhiittenkunde," Leipsic, 1881, pp. 189, 200 221. COPPER OF COMMERCE 13 Table $. — Chemical Analysis of Repined Copper {Continued) Element Casting copperfl Casting copper" Lake' Merchant bar, England* Oker, Germany^ Wallaroo, Australia ' Mansfeld, Germanys Cu+Ag Cu Ag Pb Bi As Sb Se+Te . 99.50 0.0s 99-45 99-44 O.OI 99-950 99-879 0-07t O.OI 0.02 0.0s Fe Ni Zn S O Sn Conductivity, annealed 0.06 o.is 0.38 trace U.18 99.930 99.873 O.OS7 trace o . 0006 trace 0.0014 o.ooio trace 0.0022 0.04s 99 935 99.867 0.068 O.OOII O . 0099 trace 0.0063 0.0108 0.0008 o . 0064 0.056 95.7 99 . 904 99.870 0.034 O . 0004 trace 99.397 99.39s 0.072 U.061 0.052 0.135 o. 095 ■ 99 . 648 99.6417 99.6125 O.029i2 U.020 trace U.0007 0.0172 0.0023 0.0027 trace U.0005 o . 0006 0.064 O.OII 0.063 0.064 0.0039 0.2112 0.013 0.068 O.OOI to. I 166 0.0024 0.00752* 101 . 1 Co 0.012 "c IIVO / / mo / r / 1070 ^OM J. / 1050 00 8 IQCusOjl Fig. 8.— Alloy-series Cu-CuaO have a very decided influence upon the properties. The presence of two foreign substances in copper may intensify their respective harmful efifects, or may neutralize them; they may also act independently and interfere with one another. 8. Oxygen.' — Oxygen is insoluble in copper. Nearly all the O is present as CuaO, which is the only copper oxide stable at the melting point of copper. Slade-Farrow^ found that a mixture of Cu and CuoO liquefied at 1195° C. and separated into two layers containing respectively 20 and 95 per cent. CusO. The equilibrium diagram of the Cu-CuzO series of alloys by Heyn' is given in Fig. 8. It has the characteristic V-shaped form of an alloy forming a eutectic mixture. The eutectic contains 3.45 per cent. CU2O and solidifies at 1064° C* In fire-refining copper (§ 186) the metal is saturated with CU2O to form the so-called set-copper containing about 6 per cent. CuzO, while refined copper contains 0.5+ per cent. CU2O, the amount varying with the pitch (ingot-, wire- bar, plate-) to which J Determined. 1 W. H. Bassett. 2 Brass World, 1905, I, 95. ' J. B, Cooper, Private comaunication. '^ Electrical Review, March 3, 1897, p. loi. ^Zt. Berg. Hulten. Sal. Wes., i. Pr., 1873, xxi, 252, 254. « Bay-Plant, Balbach S. & R. Co., Newark, N. J. Great Falls, Mont., "Electrolytic Wire Bar," Burns, Tr. A.I.M. E., 1913, xlvl Collections' of other analyses: Keller, Min. Ind., 1898, vii, 243. HoUard-Bertiaux, Rev. MSt., 1906, m, 205. ' See also Refining Copper, § 186. ^Proc. Roy. Soc. A., 1912, Lxxxvii, 524; /. Inst. Met., 1913, rx, 207; Met. Chem. Eng., 1913, XI, i6s; Zt. Electrochem., 1912, xviii, 817. ^ Mitth. Kgl. Versuchsanst.; 1900, xviii, 315; Metallographist, 1903, vi, 49; Tr. A. I. M. E., igo4, xxxiv, 677. ^ Dejean, Rev. Met., 1906, in, 233; reply by Heyn, p. S43- 14 METALLURGY OF COPPER Fig. 9. m^M Fig. 10. Fig. II. Figs. 9-11. — Three stages in fire-refining of copper the copper has been poled and with the character of the impurities present. Figs. 9-1 1 are photomicrographs^ of cathode copper in three stages of fire-refining. Fig. 9 represents the cathode copper, after melting down, with about 3 per cent CU2O; Fig. 10 set copper with 6.16 per cent. CU2O; and Fig. 11, wire-bar copper with 0.51 per cent. CuzO.^ The amount of CuaO present in copper containing less than 3.45 per cent. CujO can be readily found by measuring on an enlarged photo- micrograph with a planimeter the Cu areas in a given area, deducting them from the total area, which leaves the eutectic area, and calculating in this the percentage of CU2O. With a little practice close valua- tions can be made by examining a polished surface with the microscope, an operation which takes from 6 to 8 minutes.* The method of Hof man- Green- Yerxa has been modified by Huntington-Desch^ to secure greater accuracy, and much sim- plified by BardwelL* The latter projects the image upon Duplex paper so as to cover a circle 15 to 16 in. in diameter, traces the outline with a hard pencil, cuts out the copper areas, weighs them and the residual net-work of eutectic on a chemical balance, and computes the 0. From 4 to 5 determinations are made in one hour, and the results check closely. The CuaO in Cu is not reduced by either As or Sb, but readily so by Sn, Zn, Mg, and Pb.« ' Hofman, Green, Yerxa, Tr. A.I.M.E., 1904, xxxrv, 682; see also Stahl, Metallurgie, 1909, IV, 609. '' Giraud, Rev. Mil., 1905, 11, 297; Eng. Min. J., 1905, Lxxx, 170. " Hofman, Green, Yerxa, Tr. A.I. M. E., 1904, XXXIV, 671, 984. ^ Tr. Faraday Soc, 1908, iv, 51. Tr. A.I. Jlf . £., ,1913, XLVi. " Jolibois-Thomas, Rev. Mil., 1913, x, 1264. COPPER OF COMMERCE IS • The tensile strength of copper begins to be affected by 0.45 per cent. CuaO, but not the malleability; this begins to diminish with 0.9 per cent. CU2O. 9. Lead.— The constitution of the Cu-Pb alloy series has been investigated by Roberts-Austen.i Heycock-Neville,^ Hiorns,» Friedrich-Leroux/ and Giolitti-Marantonio.s In Figs. 12-120 (Friedrich-Leroux), in area Cu 65 Pb, there are formed, upon cooling, crystals of Cu and mother metal; in that of fCV crystals of Pb and mother metal; below the eutectic line and to the right oifx, crystals of Cu and eutectic; and to the left, crystals of Pb and eutectic. Alloys lying between 100 and 65 per cent. Cu form homogeneous solutions above the liquidus, 1084 to about 950° C; as soon as the temperature reaches line Cu-6s, crystals of Cu separate and continue to do this until the point 65 per cent. 1200 1000 800 600 400 Pb 200 / \ 10R4 / / \ 1 -^ .A^ •^ 05!^ / / / 327 c» Eutectic ■Cu.-Pb. C 325 °C. 20 40 60 80 lOa Cui Fig. 12-120. — Alloy-series Cu-Pb. Cu. 0.0 0.06 i< Pb 100.0 99.94ii Cu has been reached. With a further withdrawal of heat the temperature does not fall, but is kept constant by further separations of Cu until the composition of 65 per cent. Cu — 35 per cent. Pb has been changed to that of 10 per cent. Cu — 90 per cent. Pb; only now, after the complete disappearance of the former, does the temperature fall with further separations of Cu until the eutectic point/ (0.06 per cent. Cu, 99.94 per cent. Pb) has been reached with 325° C, when complete solidification takes place. It is thus seen that in Cu, when cooled slowly, there will be found a little Pb, and in Pb a little Cu. If mixtures within the range of 10 and 65 per cent. Cu, or 90 and 35 per cent. Pb, are heated above 1025° C, thoroughly stirred and poured into a chilled mold, an apparently homogeneous alloy will be obtained, which in reality is a conglomerate. Recent investigations by Friedrich-Waehlert' have fixed the critical temperature of ' 4th Report Alloys Research Committee, 1897, 51. ^Philos. Trans. A., 1897, XLn, 189. ' y. Soc. Chem. Ind., J906, xxv, 618. * Metailurpe, 1907, iv, 299. ' Guertler, "Metallographie," i, part i, p. S97. ' Metall-Erz, 1913, x, 578. i6 METALLURGY OF COPPER the saturation-point curve between 65 and 10 per cent Cu at 1025° C. with the critical point at about 35 per cent. Cu. The color of the alloys is a reddish gray. The effect of Pb upon the mechanical properties of Cu depends to a certain extent upon the amount of O present/ as the less the O, the smaller is the Pb permissible, because Pb reduces Cu20.^ Thus 0-free Cu with 0.05 per cent. Pb is red-short, while 0-bearing Cu can stand as much as 0.2 per cent, and be worked cold or hot; Cu can contain as much as 0.675 PbsAsjOs or 1.45 per cent. 2Cu20.PbO and be only just red-short (Hampe). Jolibois-Thomas' have shown that As neutralizes the harmful effect of Pb, in that Pb forms a solid solution withCusAs. Ordinarily it is held that Pb, not to exceed o.i per cent., makes Cu roll better, and that 0.2 per cent, makes it brittle.* Forging tests of Archbutt* showed that 0.2 per cent. Pb did not interfere with working at a red heat. 10. Bismuth. — Freezing-point curves have been drawn by Roland-GosseUn,* Hiorns,' Jeriomin,* and Portevin.' The curve of Portevin resembles that of Jeriomin. This has the V-shaped form of the eutectic with the eutectic point lying at 0.25 per cent. Bi, and the eutectic line extending to the borders of the diagram. As little as 0.02 per cent. Bi, which is mostly present in the metallic state, makes Cu red-short (Hampe). Baucke'^" found that 0.025 PC'^ cent, makes it brittle at a red heat, 0.05 per cent, makes it cold-short, '^ o.i per cent.^^ very brittle (Hampe). Lawrie'^ found that Cu with over 0.0005 per cent. Bi could not be drawn into wire. As'* and Sb'^ counteract to some extent the bad effect of Bi.'^ It is generally accepted that Bi203 is less injurious tJian Bi; Cu20.Bi203 less than Bi203, and that Bi203.xSb206 can be present to the extent of 0.7 per cent, without producing either cold- or hot-shortness. The alloys are coarsely granular and have a strong luster. 11. Iron. — The only freezing-point curve drawn is that of Sahmen.^' It appears to show that Cu and Fe form heterogeneous mixtures excepting at the ' Westmann, Oest. Zt. Berg. HiiUenw., 1903, li, 653. ' Jolibois-Thomas, Rev. Met., 1913, x, 1264. ^Loc. cit. * Lewis, Engineering, 1903, Lxxvi, 753; Eng. Min. J., 1904, lxxvii, 284; Am. Mfr., 1903, Lxxni, 903; Min. Ind., 1903, xii, 127; Met. Chem. Eng., 1912, x, 540. '/. Inst. Met., 1912, VII, 265. ' Bull. Soc. d'Enc, 1896, i, 1310; "Contributions k I'^tude des alliages," 1901, 109. ' J. Soc. Chem. Ind., 1906, xxv, 616; Electrockem. Met. Ind., 1905, in, 396. * Zt. anorg. Chem., 1907, Lv, 412. " Re'v. Met., 1907, iv, 1077. ^"Internal. Zt. Metallographie, 1913, m, 195. " Lawrie, Tr. A. I.M. E., 1909, xl, 604, believes that the figures 0.025 and 0.05 per cent, ought to be reversed. '2 Roberts-Austen, Second Report Alloys Research Committee, 1893, p. 121. "Loc. cit. " Johnson, J. Inst. Met., 1910, viii, 570. " Parravano, Internal. Zt. Metallographie, 1911, 1, 75. " Archbutt's Forging Tests, J. Inst. Met., 1912, vii, 264. " Zt. anorg. Chem., 1908, lvii, 9; Metallurgie, 1908, v, 298; Rev. MU., 1908, v, 366. COPPER OF COMMERCE 17 terminals of the curve, where Cu forms a solid solution with from 2 to 3 per cent. Fe, and Fe the same with a small amount of Cu. The first solution is found in the incomplete diagram of Heycock and Neville' and shown in the photomicrograph of Stead;"" the latter states that Cu with up to 2.73 per cent. Fe shows only a single micrographical constituent, and that Fe with as much as 8 per cent. Cu appears free froni any copper-colored compound. Ruer-Fick' draw the limits of solid solutions at 3 per cent. Fe and 9 per cent. Cu. Pfeiffer,* on the other hand, considers that Fe and Cu form heterogeneous mixtures throughout the whole series of alloys. The evidence for solid solu- tions at the terminals appears convincing. Iron is always likely to be present in Cu; it makes it hard and brittle, but less so than does Pb; the red color of Cu changes gradually to gray with an increase of Fe. 12. Manganese. — Copper and manganese form alloys that are frequently called manganese-bronzes. Their constitution has been investigated by Wolgodin,^ Schemtuny-Urasow-Rykowskow,^ and Sahmen.^ The curves of the last two investigators show a solid solution; the curve of Sahmen has an apparent minimum between 30 and 40 per cent. Mn. The alloys become harder as the percentage of Mn increases. With from o to 80 per cent. Cu, the alloys are gray; beyond this they become yellowish; and with 96 per cent. Cu, reddish. All the alloys are non-magnetic. An addition of from 2 to 3 per cent. Mn to Cu^ increases the tensile strength and the elastic limit, but not materially the hardness. An alloy with 8 per cent. Mn is malleable and ductile; one with 12-15 P^r cent. Mn is brittle; such alloys ought to be free from Pb or Sb. There exist cupro-ferro-manganese alloys prepared by the addition of ferro- manganese to Cu.^, 13. Nickel. — The leading freezing-point curves published are those of Guertler-Tammann,!" Kurnakow-Schemtuny,'' and Tafel.'^ The two metals form solid solutions throughout. The alloys rich in Cu are not attracted by the. magnet; those rich in Ni are. Nickel makes Cu pale red and hard; 0.3 per cent. Ni shows no effect, 2-3 per cent, greatly increases the hardness and raises the tensile strength. The presence of Sb increases the effects of Ni; hence in the ^Philos. Trans. A., 1897, CLXix, 189. '/. Iron Steel Inst., 1901, II, 108. 'Ferrum, 1913, xi, 39. ^ Metallurgie, 1906, lii, 281. ' Rev. Mit., 1907, IV, 25. ^Zt. anorg. Chem., 1908, LVii, 253; Rev. MH., 1908, v, 371. 'Z/. anorg. Chem., 1908, LVii, 201; Rev. Mel., 1908, v, 373. = Lewis, /. Soc. Chem. Ind., 1902, xxi, 842. Guillet, fitude industrielle des alliages, Dunod-Pinat, Paris, 1906, p. 752. Heussler, Verh. Ver. Bef. Gewerbefl., 1903, Lxxxii, 277; Iron Age, June 28, 1904, p. 18. •Parravano, Int. Zl. Metallographie, 1913, iv, 171; Melall-Erz, 1913, x, 503; /. Inst. Met., 1913, DC, 213. " Zl. anorg. Chem., 1907, Lll, 25; Rev. MH., 1908, v, 375. " Zt. anorg. Chem., 1907, Liv, 151; Rev. MH., 1908, v, 377. ^'Metallurgie, 1908, v, 343 ,37s. i8 METALLURGY OF COPPER presence of from 2 to 3 per cent. Ni, the Sb ought to be absent. However, 0.3 Ni+Sb does not affect the malleability in cold-working. According to Stahl,! Mansfeld copper with one-tenth per cent. Ni has a tensile strength of 31,000-47,000 lb. per square inch; an elongation in 8 in. of 39.5-46.0 per cent.; and a reduction of area of 50.5-60.7 per cent. The Cu-Ni-Fe series of alloys has been investigated by Vogel.^ 14. Cobalt.— The Cu-Co alloys are at present of no industrial importance. Freezing-point curves have been traced by Konstantinow^ and Sahmen.* 15. Arsenic. — Passing over the earlier work of Hiorns,^ there exist two freezing-point curves by Friedrich^ and by Bengough-HilL' The revised curve of Friedrich, Fig. 13, shows the following: Cu forms with As a solid solu- tion reaching with 684° C. a maximum in 4 per cent. As at the terminus of the eutectic line; the eutectic with 78.5 per cent. Cu is made up of the solid solu- tion of Cu with 4 per cent. As and the compound CusAs; the summit, 830° C, represents CusAs with 71.8 per cent. Cu. A hidden chemical compound, Cu6As2 (67.9 per cent. Cu), is formed at 710° C. Noth- ing is definitely settled regarding the eutectic line at 604° C, and the trans- formation line at 307° C. Bengough-Hill confirm the existence of the compounds CU3AS2 and CU5AS2, but believe that there exists a series of solid solutions between these compounds. The mechanical properties of Cu are not harmed by 0.5 per cent. As; with 0.8 per cent. As copper can be drawn into the finest wire; i per cent. As begins, to cause red-shortness.^ The amount of O present in Cu has a decided influence upon the permissible quantity of As, as As does not reduce CuaO;^ thus 0.4 per cent. CU2O. x AS2O6 has no effect whatever upon the mechanical properties of Cu, while more than 0.4 per cent, causes cold-shortness. StahP" states that Cu with 0.30-0.35 per cent. As has a tensile strength of 28,000-29,200 lb. per square inch; an elongation of 33-44 per cent, with a reduction of area of 47-62 °c. 1100 900 700 BOO '300 \ ~ \ S s. cu 830 ■* 6& = \ V- 7 :^ 604 ■\ L^ w 2 10 20 30 40 5Q%As Fig. 13. — Alloy-series Cu-As. ^ Op. cit., 1909, VI, 610, 1910, VII, 14; discussions by Heckman, op. cit., 1910, vi, 760. ' Zt. anorg. Chem., 1910, lxvii, i. ' Rev. M6t., 1907, rv, 983; Min. Ind., 1907, xvi, 377. * Zl. anorg. Chem., 1908, Lvii, i; Rev. Met., 1908, v, 364. ^ Electrochem. and Met., 1903-04, in, 648, 734; Ekctrochem. Ind., 1904, 11, 176; Min. Ind., 1903, XII, 124; J. Inst. Met., 1910, m, 54. « Metallurgie, 1905, 11, 484; 1908, v, 529. ' /. Inst. Met., 1910, III, 34. ' See Roberts-Austen, Second Report, Alloys Research Committee, 1893, p. 119. ° Jolibois-Thomas, Rev. Met., 1913, x, 1204. '" Metallurgie, 1909, vi, 611. COPPER OF COMMERCE 19 per cent. The following table of Lewis' shows the influence of As upon the ten- sile strength of Cu; other data are given by Bengough-Hill.^ Lewis^ states that Cu with from i to 1.37 per cent. As rolls very well; that the tensile strength is from 6,000 to 10,000 lb. higher than that of ordinary sheet copper; and that the elongation is not reduced. However, 0.6 per cent. As is generally considered the limit for good copper. Bengough-Hill found that Cu with less than i per cent. As was ruined when annealed in a reducing atmosphere above 650° C. The effects of As upon electric conductivity and absorption of gases has been dis- cussed in § 4. Table 6. — Influence of Arsenic tjpoN Tensile Strength of Copper As, per cent. Tensile strength, lb. per. sq. in. Elongation, per cent. Elastic limit, lb. per. sq. in. 0.00 26,800 2S 14,000 0. 24 33,840 27 -S 20,500 0.53 36,760 295 ig,o20 0.75 36,620 21 17,820 0.94 36,040 2S 18,020 1.37 37,660 28 20,180 1.80 3S,66o 20 ^3,040 16. Antimony. — The constitution of copper-antimony alloys has been in- vestigated by Baikoff* and Hiorns.^ Baikoff's curve is shown in Fig. 14. Start- ing with the Sb-end at 629° C, Sb is seen to form a solid solution a with Cu reaching its maximum with 10 per cent. Cu; B (524° C, 25 per cent. Sb) is the eutectic point of the mixture of solid solution a and chemical compound CujSb, the eutectic line extending to 51 per cent. Cu. This compound, which has a characteristic purple color, is formed at 584° C. by the grayish compound CusSb combining with Sb according to 2 Cu3Sb-)-Sb=3Cu2Sb. The compound CuaSb (61.5 per cent. Cu) solidifies at 681° C; between 681 and 584° C, the solid solution /3 of CusSb and Sb (51-53.5 per cent. Cu) separates, and is transformed at 584° C. in part into Cu2Sb and |3. Between 53.5 and 69 per cent. Cu the solid solution fi separates unchanged. Between 69 and 96 per cent. Cu there ' J. Soc. Chem. Ind., 1901, xx, 254. "/. Inst. Met., 1910, III, 37. Johnson, J. Inst. Met., 1910, rv, 163; Met. Chem. Eng., 1910, viii, 570; J. Inst. Met., 1912, vnt, 192; 1913, X, 275. Lewis, Met. Chem. Eng., 191 2, x, 540. •Greaves, /. Inst. Met., 1912, vii, 218. Archbutt, op. cit., 1912, vii, 262. Law, op. cit., 1912, viii, 222. Baucke, Internat. Zt. Metallorgraphie, 1913, iii, 195. 'Engineering, 1903, lxxvi, 733; /. Soc. Chem. Ind., 1903, xxii, 1351. *Bull. Soc. d'Encour., 1903, i, 626; Rev. Mit. Extr., 1905, n, 433. ' /. Soc. Chem. Ind., 1906, xxv, 616. 20 METALLURGY OF COPPER separates above 630° C. the solid solution of CusSb and 7, and a solid solution of Cu with 2.5 per cent. Sb; below 630° C. the former is transformed into /3 and y solution; below 407° C. the last transformation takes place, leaving on the antimony side of the ordinate 6r per cent. Cu, the mixture of Cu2Sb and CusSb, and on the copper side y and CusSb. The effect of Sb upon the mechanical properties of Cu is similar to that of As. Sb does not reduce CU2O.' The tensile strength is increased by Sb. Thus Hampe^ showed that Cu with 0.26 per cent. Sb gave 73,800 lb., and with 0.529 Fig. 14.— Alloy-series Cu-Sb. per cent., 77,900 lb. per square inch.; the sample with 0.529 per cent. Sb could still be drawn to a fine wire ; i per cent. .Sb caused cold-shortness. The so-called copper-mica (6 CujO.SbaOs+SNiO.SbaOb, gold-colored to yellowish-green scales, formed in refining Cu containing both Sb and Ni, can be present to the extent of 0.726 per cent, and not interfere with malleability, but does affect ductility; ' Jolibois-Thomas, Rev. MU., 1913, x, 1264. 2 Chemiker Z., 1892, xvi, 726, Second Report, Alloys Reseerch Comm., 1893, p. 120. COPPER OF COMMERCE 21 "C. 1 .44 per cent, of the salt makes Cu red-short. It is generally held that Cu should not contain over 0.05 per cent. Sb, as Cu with o.i per cent. Sb has been found to crack at the edges when it is rolled/ and cannot be bent without breaking. Other data are given by Greaves,^ Archbutt,' Johnson/ and Law.^ 17. Sulphur.* — Sulphur is present in Cu as CU2S. The freezing-point curve Cu-CuzS of Heyn-Bauer/ shown in Fig. 15, resembles that of Pb-Cu. Starting with CU2S, its melting-point of 1127° C. is lowered by additions of Cu; when the liquid is cooled and reaches the branch 1127-1102" C, metallic Cu separates with a lowering of temperature until the point at 1102° C. has been reached; a further separation of Cu causes no fall in temperature until the composition has been changed into that of the left terminus of 1102° C, when upon further separation of Cu there is a quick descent of the curve to the eutectic point, 3.8 per cent. CuaS, 1067° C, fol- lowed by a quick rise to the freezing point 1084° C. of Cu. Hampe has shown that Cu with 0.25 per cent. S is still malleable, and that 0.5 per cent. S makes it cold-short, but not red-short. On the other hand, Lewes* found that Cu with o.i per cent. S cracked badly on rolling and bent badly; and that 0.5 per cent. Mn or Al counteracted the bad effect of S. Sperry' found that as little as 0.1 per cent. S caused blowholes; and that the Cu could be forged, but would not stand bending without cracking. 18. Selenium and Telltuium.^" — These two elements are found in pig copper in very small quantities, 0.007 per cent, in Montana copper according to Keller;'^ and are removed by electrolytic refining process to such an extent that they rarely appear in market copper. iiiiU 1 1127 1120 \ 1 ( \ \ / 1 1 \ / I 03^ V 1090 i — -- 1084 1080 — 1 10 67 » 20 40 CO 80 Fig. 15.- iooCu,si -Alloy-series Cu-CuzS. 'Lewis, Engineering, 1903, lxxvi, 753; Am. Mfr., 1903, lxxiii, 903; Met. Chem. Eng., 1912, X, S40. '/. Inst. Met., 1912, VII, 218. 'Op. cit., 1912, VII, 262. ' Op. cit., 1912, VIII, 192. ' Op. cit., 1912, vin, 222. » Stahl, Berg. Huttenm. Z., 1890, xlec, 99, 127. Hinrichsen-Bauer, Metallurgie, 1907, iv, 315, Oest. Zt. Berg. Eitttenw., 1907, iv, 473. ' Metallurgie, 1906, in, 76. 'Engineering, 1903, Lxxvi, 73; Am. Mfr., 1903, lxxiii, 904. 'Brass World, 1913, rx, 91. " Microscopical tests: Heyn-Bauer, Metallurgie, 1906, iii, 84. Hinrichsen-Bauer, op. cit., 1907, iv, 315; Oest. Zt. Berg. HUltenw., 1907, LV, 473. " Min. Ind., 1898, vii, 241. 2 2 METALLURGY OF COPPER Cuprous selenide/ CugSe, melts at 1113° C. and forms with Cu an eutectic containing 2 to 3 per cent. Cu. The constitution of Cu-Te alloys has been investigated by Chikashige'' and Pouchine.^ There exist, according to the former, a gray Cu2Te, a violet Cu4Te3, an eutectic Te+Cu4T3, and several solid solutions; the latter found an additional compound CuTe. The metal was discovered by Egleston* in copper, which upon analysis showed 0.08 per cent. Te, and the copper was red-short. 19. Silver. — The first freezing-point curve, by Heycock-Neville*, determined the general eutectic character of the series of alloys; Friedrich and Leroux^ of carried the work further, and Lepkowski^ completed the curve fixing the extent the solid solutions at the terminals. The curve of the last, with atomic changed ^ 31° \ ~v 1 l'^ ...^ / 7 94 12% Ag 1100 Cu 1100 900 800 700 10 20 30 40 50 60 70 80 90 100 ^cr,* Fig. i6. — Alloy-series Cu-Ag. into weight per cent., is given in Fig. i6. The eutectic point with 28 per cent. Cu lies at 779° C. ; solid solutions are formed at the ends of the eutectic line, Cu holding 2 per cent. Ag and Ag 7 per cent. Cu. A knowledge of the structure is of importance for the correct sampling* of copper ingots that carry precious metal. The electric conductivity and hardness of Cu-Ag alloys have been studied by Kurnakow-Puschin-Senkowski,' and mechanical properties, hot and cold, by Johnson. 1" 20. Lead and Silver. — The investigation of Friedrich andLeroux" has shown that these metals form a ternary eutectic, with Cu 0.5 per cent., Ag 2.0 per cent., ' Friedrich-Leroux, Metallurgie, 1908, v, 356. 2 Zt. anorg. Chem., 1907, Liv, 50; Rev. Mel., 1908, v, 392. ^ Op. cii., 1907, rv, 929. < Tr. A. I. M. E., 1881-82, X, 493. ^ Phil. Trans. A., 1897, CLXxxrx, 25. ° Metallurgie, 1907, iv, 297-7. ' Z<. anorg. Chem., 1908, XLix, 289. s Keller, Tr. A. I. M. E., 1897, xvn, 106; i9ii,xlii, 905; Eng. Min. J., 1912, xcra, 703, 729. Ledoux, School Mines Quart., 1897-98, xix, 366. Wraith, Tr. A. I. M. E., 1910, XLi, 318. Liddell, Eng. Min. J., 1910, xc, 897, 953, 1095; 1911, xcii, 1173. Smoot, op. cit., 1912, xciii, 1213. " Zt. anorg. Chem., 1910, Lxviii, 123. School Min. Quart., 191 2, xxxiii, 405. '"7. Inst. Met., 1910, rv, 163. "■ Metallurgie, 1907, rv, 293. COPPER OF COMMERCE 23 Cu 97.5 per cent., which freezes at from 0.5 to 1° C. below the binary eutectic of Ag-Pb (300° C). 21. Gold. — The first freezing-point curve was drawn by Roberts-Austen and Rose.' It has been supplemented by the work of Kurnakow and Schemtuny,^ they show that the two metals form solid solutions throughout with a low point at 82 per cent. Au, as seen in Fig. 17. This disposes of the supposed existence of definite chemical compounds.' The ternary series Cu-Au-Ag has been studied by Janecke.'' c. iioori Cu 1000 900 800 i^ -^ 1063° A ::::5 ~^ _j^ y/ «H^u Au 10 20 30 40 60 60 70 80 90 lOOjiAu , Fig. 17. — Alloy-series Cu-Au. 22. Minor Metals. — The following freezing-point curves are at present of little metallurgical importance: Cu-Ca/ Cu-Mg," Cu-Cd/ Cu-Tl/ Cu-Pd,^ Cu-Pt/o Cu-Va," Cu-W,i2 Cu-Ti," Cu-Cr," Cu-Cd-Sb.'s ^Proc. Roy. Soc. ,jgoi, lxvii, 105. ' Zt. anorg. Chem., 1907, liv, 159. ' Pearce, Tr. A.I. M. E., 1884-85, xra, 738. * Metallurgie, 1911, viii, 597; /. Inst. Met., 1911, vi, 331. ' Donski, Zt. anorg. Chem., 1908, LVii, 218; Rev. MU., 1908, v, 360. Bensell, Metall-Erz, 1914, xi, 10, 46. ' Ursakow, Chem. Centralblali, 1908, i, 1038; Rev. MU., 1908, v, 371. Sahmen, Zt. anorg. Chem., 1908, LVii, 26. 'Sahmen, op. cit., 1906, XLix, 301; Rev. Met., 1908, v, 362. * Doerinkel, Zt. anorg. Chem., 1906, XLvra, 185; Rev. Met., 1908, v, 395. ' Ruer, Zt. anorg. Chem., igo6, li, 223; Rev. MSt., 1908, v, 386. " Doerinkel, Zt. anorg. Chem., 1907, Liv, 33s; Rev. Met., igo8, v, 388. " Guillet, Rev. MU., 1906, in, i7i;Geme Civil, 1905, xlvii, 147. Norris, /. Franklin Inst., 1911, CLXXI, 561. " Guillet, Rev. MU., 1906, iii, 171; Genie Civil, 1905, xlvii, 147. " Rossi, Eledrochem. Met. Ind., 1908, vi, 257; iQOQ. vn, 88. Bensell, Metall-Erz, 1914, xi, 10,46. " Guillet, Rev. MU., 1906, in, 171; Genie Civil, 1905, xlvii, 147. Hindrichs, Zt. anorg. Chem., 1908, XLix, 414; Electrochem. Met. Ind., 1909, vn, 34. " Schleicher, Internal. Zt. Metallographie, 1912, in, 103. CHAPTER IV INDUSTRIAl ALLOYS 23. Lidustrial Alloys in General.' — Copper forms the basis of a large number of important alloys. As a rule they are more fusible and more fluid than copper, give sounder castings, are harder, less malleable and less corrodible.^ As regards the structure, it may be said that alloys in which copper forms an unsaturated solution with another metal show a high degree of toughness and malleability, while alloys in which copper forms an intermetalHc compound or solid solutions with the compounds of the latter are usually hard and brittle. The solubility of gases in copper alloys' is similar to that in copper.* Thus the solubility of SO2 increases with the rise in temperature and is proportional to the square root of the gas pressure. The mechanical properties of the alloys are greatly affected by a rise in tem- perature. ° The behavior of copper has been illustrated in Fig. 4; examples of alloys are given in Figs. 25, 26, 36, 37, ^S. 1 Japing, E., "Kupfer und Messing," Hartleben, Leipsic, 1883. Guettier, A., "Le fondeur en m^taux," Bernard, Paris, 1890. Wiist, F., "Handbuch der Metallgiesserei," Voigt, Weimar, 1897. Thurston, R. H., "A Treatise on Brasses, Bronzes and other Alloys," Wiley, New York, 1900. Guillet, L., "Les alUages m6talliques," Dunod, Paris, 1906. Brannt, W. T., "The Metallic Alloys," Baird, Philadelphia, 1908. Krupp, A., " Die Legierungen," Hartleben, Leipsic, 1909. Sexton, A. H., "Alloys, non-ferrous," Scientific Publ. Co., Manchester, 1909. Law, E. F., "Alloys and their Industrial Applications," GrifBn & Co., London, 1913. Buchanan, J. F., "Practical Alloying," Penton Publ. Co., Cleveland, Ohio, 1910. Kaiser, E. W., " Zusammensetzung der Gebrauchlichen Metallegirungen," Knapp, Halle, 1911-12. Schott, E. A., "Die Metallgiesserei," Voigt, Leipsic, 1913. Hiorns, A. H., "Mixed Metals and MetaUic Alloys," Macmillan, London, 1913. Buchner, G., "Die Metallfarbung," Krayn, Berlin, 1910. Brown, W. N., "Dipping, Burnishing, Lacquering, etc.," Scott, Greenwood & Sons, London, 1912. Gowland, "History," /. Inst. Met., 1912, vii, 23. 2 Diegel, Verh. Verein. BeVord. Gewerb., 1899, Lxxvni, 313; 1903, lxxxii, 93, 119, 157. Bengough and Bengough- Jones, Reports of Corrosion Committee, J. Inst. Met., 1911, V, 28, 1913, X, 13. 3 Sievert-Bergner, Zt. phys. Chem., 1913, lxxxii, 257; /. Inst. Met., 191,, ix, 2?i. ^See§4. ' Grard, Rev. Met., 1909, rv, 1069. Weidig, Verh. Verein. Beford. Gewerbefl., 1911, xc, 455, 525. Johnson, Met. Chem. Eng., 1911, ix, 399. Bengough, J. Inst. Met., 1912, vii, 123. Huntington, op. cit., 1912, viii, 126. Milller, Metall-Erz, 1913, i, 219. 24 INDUSTRIAL ALLOYS 25 In the preparation of alloys' the pouring temperature shows a decided influ- ence upon the closeness of the grain and thereby upon the strength of the product. The leading copper alloys are those with zinc, tin, aluminum, gold (§ 21), and silver (§ 19); of secondary importance are the alloys with phosphorus, silicon, and manganese, which are reviewed first. The melting-points of the following common industrial alloys have been de- termined by Gillett-Brown.^ Table 6a. — Melting-points of Some Common Industrial Alloys Alloy Composition desired Cu Zn Sn Pb Composition by analysis Cu Zn Sn Pb Melting- point (liquidus) Gun metal Leaded gun metal... Red brass Low-grade red brass. Leaded bronze Bronze with zinc . . . Half yellow, half red Cast yellow brass. . . Naval brass P. ct. 88 85 1 85 82 80 8S 75 67 6i| P. ct. 2 2 S 10 5 20 31 37 P. ct. 10 9i 5 3 10 10 Manganese bronze. . . Cu S6 Zn 41 Sn 0.9 P. ct. 3 5 S 10 P. ct. 85.4 P. ct. 1-9 P. ct. 9-7 Si-5 10.4 31 84,6 7S-0 66.9 61.7 20.0 30,8 36.9 10.4 2.0 1-4 Fe Al 0.4S P. ct. 3-0 S-o 3-0 2-3 Mn 995 980 970 980 945 980 920 895 855 870 24. Phosphor-copper. — The constitution of these alloys is shown by the curve of Heyn and Bauer,' Fig. i per cent. P melting at 707° C; the chemical compound CusP with 14 . 1 per cent. P freezing at 1 100° C. ; one solid solution of Cu with a maximum of 0.175 per cent. P, and another of CU3P with a probable second chemical com- pound CU6P2.i Alloys are prepared in two ways; either by plunging stick-P, held in an inverted cup,^ into Cu, melted in a crucible, and keeping it submerged until it has been taken up, or by causing 8. This curve shows an eutectic with 8.27 c. 1100 Cm lOOO 700 slOS \^ \ s 10 18° s \ / / \ / / \ / 0, s \ / i \ / ^l 707' 8.27 ^ 10 ISBi Fig. 18. — Alloy-series Cu-P. 1 Gillett, Eighth Internal. Congress Appl. Chem., New York, 191 2, 11, 105. 2 Bureau Mines, Techn. Paper 60., 1913. 'Zt. anorg. Chem., 1907, lii, 131; Metatlurgie, 1907, iv, 242, 257; Rev. Met., 1908, v, 377- * Huntington and Desch, Tr. Faraday 5oc, 1908-09, iv, 51. ' Wickhorst, Iron Age, March 25, 1897, p. 2. 26 METALLURGY OF COPPER fused Cu to combine with P-vapor. The apparatus for the second method, shown in Fig. 19/ consists of the crucible A clamped to the funnel B with discharge-opening c. Phosphorus is placed in A and molten copper poured into B. The phosphorus in A is vaporized and forced to pass through the copper as it flows through c. As an alloy containing over 14. i per cent. P gives off P upon heating, alloys with over 14-15 percent. P cannot be produced by fusion. Heyn states that alloys with as much as 20 per cent. P can be produced by mixing Cu-filings and red P in crucibles, connected in series wash-bottle fashion, and heating one at a time to 300-400° C, but not over 700° C, when the vapors from the crucible that is being heated will be condensed by the others. The commercial alloy contains from g to 15 per cent. P ; it is steel gray, so hard that it can be filed only with difficulty, fine-grained and brittle. Small additions of P make Cu hard; Cu with 0.05-0.10 Vie P^'' ^^^^- P 3,nd not over 0.04 per cent. O is still easily , rolled.^ Hiorns' found that Cu with 0.5 per cent. P rolled Fig. 19 . — Appa- -^ ^ ratus tor preparing well giving smooth edges; and Miinker'' maintams that pure phosphor-copper. copper with I per cent. P may be rolled hot or cold, but that the ductility is much reduced by 0.2 per cent. P. The constitutional diagram shows that with over 0.175 P^r cent. P the eutectic, containing hard brittle Cus P, separates. Uses of phosphorus-Copper Alloys. — The principal use of the commercial alloy is in the manufacture of the so-called phosphorbronze; it is added to Cu that is to be rolled, as the metal appears to work more evenly owing to the deoxidation of the CU2O present. It has been suggested for use in the refining of coarse copper in order to assist in the removal of O according to 6CU2O-I-2P = 10 CU-I-2CUO.P2O6. Alloys of Cu and Mn^ have been used for this purpose. StahP shows that the addition of such alloys increases the specific gravity of commercial Cu by reducing the CU2O that is present and by diminishing the absorbing power for gas. 25. Silicon-copper.' — The freezing-point curve of Rudolfi^ (Fig. 20) 1 Hiorns, A. H., "Mixed Metals," Macmillan, New York, 1913, p. 219. ^ Lewis, loc. cit.; Mel. Chem. Eng., 1912, x, 540. ' /. Soc. Chem. Ind., igo6, xxv, 622. '^iMetallurgie, 1912, rx, 185; /. Inst. Met., igi2, vii, 272. 'Rossler, Berg. Euttenm. Z., 1878, xxxvii, 370; Zt. Berg. Hutten. Sal. Wes. i. Pr., 1879, xxvn, 14; Eng. Min. J., 1880, xxrx, 317. Lewitzky, Berg. Huttenm. Z., 1880, xxxrx, 64; Rev. Un. Min., 1879, vi, 24. ^ Berg. Huttenm. Z., 1901, lx, 78. ' Phillips, Metallurgie, 1907, iv, 587, 613; Electrochem. Met. Ind., 1907, v, 468. Baraduc-Muller, Rev. Met., 1910, vii, 711. Frilley, op. cit., 1911, vni, 511. ^Zt. anorg. Chem., 1907, uii, 216; Metallurgie, 1907, iv, 851; Rev. MU., 1908, v, 390. INDUSTRIAL ALLOYS 27 °o ISOO 1200 3 ' uoo £c« 2 1000 I 900 o "^ 800 TO Si replaces for the present the older approximations' of the constitution of copper- silicon alloys, although objections have been made to some of its features.^ Starting at the Cu-end of the curve, it is seen that Cu forms with Si a solid solu- tion reaching 4.5 per cent. Si, next comes a hidden chemical compound Cui9Si4 (8.59 per cent. Si) which forms with the second chemical compound CuaSi (12.9s P^r cent. Si, (melting at 862° C.) the first eutectic (8.3 per cent. Si, freezing- point 829° C); the second eutectic of Cu.iSi-f-Si, with about 18 per cent. Si, freezes at 7 10° C. Photomicrographs have been published by Arnold-Jeff erson' and Albro.* Copper-silicon alloys have been prepared in various ways.^ Electrothermic methods have probably replaced the earlier modes of operating; and in these the electric fusion of a mixture of Cu, sand, and C in a resistance furnace has given place to the simple fusion of Si and Cu, since metallic Si is produced on a large scale and is sold at a reasonable price. The Cowles Electric Smelt- ing and Aluminum Co., Lockport, N. Y., produces pure silicon-cop- per with 20-30 per cent. Si, sold in ingots weighing about 14 lb.; details of the method of working have not been made public. Copper-silicon alloys are brit- tle, and the more so the higher the Si-content. The 20-30 per cent, alloy is easily broken into glassy splinters by a tap with a hammer; a fresh surface is silvery and assumes a reddish tint when exposed to the air. According to Hampe'^ an addition of Si to Cu increases the hardness and at the same time assists in the production of sound castings; 3.472 per cent. Si does not reduce tlie tensile strength and malleability of copper; 6 per cent. 1 DeChalmot, Am. Chem. J., 1897, xix, 118, 871; 1896, xviii, 95; 1898, xx, 437- Lebeau, Sixth Int. Cong. Appl. Chem., 1906, 11, 4". Vigouroux, Compt. rend., 1896, cxxii, 318; 1905, cxli, 890; 1906, cxiii, 87; 1907, cxliv, 1 2 14. ' Bornemann, Metallurgie, 1907, iv, 852. Guertler, Phys. Chem. Centralblatt, 1907, iv, 576. Rudolfi, op. cit., 1908, V, 223. Portevin, Rev. Mit., 1908, v, 391. ' Eng. Min. J., 1896, Lxi, 353. * Electrochem. Met. Ind., 1905, in, 461. ' Maberry, Am. Assoc. Adv. Sc, 1886, xxxiv, 136. Hunt, Tr. A. I. M. E., 1885-86, xiv, 492. Steinhardt, Eng. Min. J., 1899, LXVli, 710. Kroupa, Oest. Zt. Berg. Huttenw., 1903, Li, 285. 'Chem. Z., 1892, xvi, 726; Berg. Huttenm. Z., 1892, li, 321. 1404° ^ ^ ^ " / y — r '-' / \ \ CO / \ ^ V 1 i 10 20 Fig. 30 40 60 CO 70 60 90 100 Si % 20. — Alloy-series Cu-Si. 28 METALLURGY OF COPPER makes it brittle; Cu with 8 per cent. Si can be pulverized; with 11.7 per cent. Si it is as brittle as glass. Rudolfii states that Cu with S per cent. Si is readily drawn into wire. According to Davis'' the addition of o.i per cent. Si to melted Cu increases the fluidity and gives castings that are clean and free from blow-holes, which is due probably' to the reduction of CU2O. The cast alloy 97 Cu and 3 Si has a tensile strength of 55,000 lb. per square inch and from 50 to 60 per cent, ductility; the cast alloy 95 Cu and 5 Si has 75,000 lb. tensile strength and 80 per cent, ductility; over 5 per cent. Si makes Cu brittle. An analysis of Si-Cu spring-wire* gave Cu 97.59, Si 2.31, Fe o.io. 26. Brass (Cu-Zn) in General. — The constitution of brass has been a sub- ject of study since the days of Storer.* The leading freezing-point curves are °c. 1100 1000 -^ ^a+L N \b at 3 800 700 600 h w ^^ >■ ^ \ Y 1 / \ \ V- LiQ\ J> a \ \ /J+J- dz \.] WoK 1 r 7 \ 500- 1 di ',/,!+( :'^ u 63 J, F' 300 « + r dt )■+€ f "»/, ( + 1 1 100 90 10 70 Fig. 21.- 60 30 20 Cui -Alloy-series Cu-Zn brass. those of Roberts-Austen,' Sheperd,^ Sackur,' Tafel,' and Carpenter-Edwards.*" The curves of Sheperd and Tafel resemble one another. Sheperd holds that there are no chemical compounds; Tafel that the compound Cu2Zn3 (Cu 39.33, Zn 60.67, melting-point, 830° C.) is established and that possibly there is a second compound CuZn (Cu 49.3, Zn 50.7)." Carpenter-Edwards have added 1 Loc. cit. " Aluminum World, 1896, in, 241. ' Vickers, Foundry, 1908, xxxii, 1. ' Brass World, 1905, i, 413. ' Mem. Am. Academy, i860, viii, 27. ^Fourth Rep. Alloys Research Comm., 1897, p. 31. '/. phys. chem., 1904, vili, 421; Meiallurgie, 1904, i, 462. * Ber. deutsck. chem. Ges., 1905, xxxviii, 2186. » Meiallurgie, 1908, v, 349, 375, 413 (incl. bibliography pp. 343, 349). " J. Inst. Met., 1911, V, 127, 1912, viii, 51, 59. 1' Guertler, Zt. anorg. Chem., 1906, Li, 429. Hudson, J. Soc. Chem. Ind., 1906, xxv, 503. Bengough-Hudson, op. cit., 1908, xxvii, 43, 654. INDUSTRIAL ALLOYS 29 to the curves of Sheperd and Tafel a transformation point at 470° C, below which the constituent j3 splits into a+7. The diagram of Sheperd-Tafel and Carpenter-Edwards is given in Fig. 21. It shows 6 constituents, the character- istics of which are assembled in Table 7. Table 7. — Characteristics of Components op brass (Sheperd) Component' Color Color of fracture a +7 /3 7 + /3 7 7+« + « E « + 1) Clear yellow to copper red ... Red changing to full yellow . . Reddish yellow with a yellow- ish cast. Light bluish gray Reddish yellow Reddish yellow Yellowish red Silvery Silvery gray to bluish gray . . . Bluish gray Bluish gray, becoming lighter . Zinc color Yellow. Yellowish red. Yellowish red. Yellowish red. Silvery with pinkish tinge. Silvery, very brilliant. Silvery gray, becoming duller. Bluish gray. Zinc color. Zinc color. Alloys consisting of the solid solution a, which has a range of from 100 to 64 per cent. Cu, can show no variety of structure. Figs. 22 and 23 are photo- micrographs of common brass, 66.6 per cent. Cu, cast and annealed. The Common brass, cast. Fig. 23. — Common brass, annealed. dendritic structure of the a-crystals is due to the formation of copper-rich centers surrounded by zinc-rich bo.rders, and to the attack on the borders by etching. The annealed specimen, in which equilibrium has been established between Cu and Zn, shows large polyhedral forms pitted irregularly by etching. Alloys consisting of a-crystals are malleable and ductile. Alloys with a composition lying between 64 and 54 per cent. Cu are likely to be brittle if cooled ' Constituents a and /3 are malleable and ductile; 7, «; « and 7, are increasingly brittle. 3° METALLURGY OF COPPER slowly to below 470° C, because of the presence of the constituent 7; if chilled above 470° they will be tough, as they are made up of the components a+j8. Murray! furnishes photomicrographs of the crystal forms a-17; and Charpy,^ 48 illustrations of different industrial brasses. The leading mechanical proper- ties^ are shown in Fig. 24. The tensile strength is seen to grow with increase of zinc until it reaches a maximum with about 56 per cent. Cu (conglomerate a and /3, chilled above 470° C), and then to fall quickly(appearance of 7 constitu- ent); the elongation reaches its maximum earlier at about 70 per cent. Cu (limiting concentration of a); the compressive strength attains the largest Cu 100 Zn 90 10 80 20 70 30 60 40 50 50 40 60 30 70 20 80 10 90 % 100* Fig. 24. — Mechanical properties of brass at ordinary temperature. figure with 50 per cent. Cu. The total shrinkage^ shows the largest maximum at 40 per cent. Cu; the same is the case with the hardness; and these two phenom- ena coincide with Tafel's chemical compound Cu2Zn3. A table of the mechanical properties of a series of analyzed brasses, cast and annealed, was presented by Guillet and Revillon^ at the London International Congress of 1909. At elevated temperature the mechanical properties show other values than those given in Fig. 24 for ordinary temperature.^ '■J. Instil. Met., 1909, 11, i. ="' Contributions a, FiStude des alliages," Paris, 1901, pp. 1-62. '/. phys. chem., 1913, xvii, i; J. Inst. Met., 1913, ix, 216; see also Bancroft-Lohr-Wilder, VIII, Internal. Congr. Appl. Chem., 191 2, 11, 8. * Turner-Murray, J. Instil, Met., 1909, 11, 98. Wust, Metallurgie, 1909, vi, 709; Iron Age, 1910, lxxxv, 790. Chamberlain, J. Inst. Mel., 1913, x, 193. ^Rev. Mil., 1909, VI, 1 251. * Bengough-Hudson, /. Inst. Met., 1910, rv, 92. Johnson, Met. Chem. Eng., 1911, rx, 399. Bengough, /. Inst. Met., 191 2, vii, 123. Huntington, op. oil., 1912, viii, 126. Guillet, "Wire-drawing," Rev. Met., I9i3,x, 769. INDUSTRIAL ALLOYS 31 The mechanical changes which two brasses, |^ and ^, with a and a+/3 as components, undergo with increasing temperatures are shown in Figs. 25 and 26 by Grard.' Annealed brass has been hardened by mechanical treatment and then tested at temperatures ranging from zero to 900° C. The curves in Fig. 25, show for f^ brass that decided changes in the three mechanical properties given take place between 200 and 300° C, and that the same is the case with lUU 80 ^ GO \ s \ s u N Oh \ V, — — --. ^ \ U]t "^^?"s3;^,, 20 v ; V £ last cLi nit Elo igat on %i^J ■ ^i heB_ .^' ""~~" 100 200 300 400 500 600 700 800 900 1000 Temperature of Annealing Degrees Centigrade Fig. 25. — Mechanical properties of J-J brass at varying temperatures. 100 » g 60 40 20 "x ~~^ \ ^ ncJn ■» . ^ \ \ ,(TS' X"^. %, N ^ .-" > # — - — Uit mat ; Sti cng h \ K s / ^ lash pU_ nit — — ~-1 — — y - '0 100 200 300 400 BOO 600 700 800 900 1000 Temperature of Annealing - Degrees Centigrade Fig. 26. — Mechanical properties of f-J brass at varying temperatures. II brass (Fig. 26) between 200 and 300° C. Additional data are furnished by Guillet.2 The effects of anneattng upon the structure have been studied by Portevin^ and Robin.* When heated in vacuo,^ the Zn is volatilized at a low temperature. ^Rev. Met., igog, vi, 1069; Metallurgie, 1910, vii, 651; Proc. Internal. Congress Testing Materials, New York, 191 2. ^ Rev. M6t., 1913, X, 671. 'Op. cit.,., 1913, X, 677. *0p. oil., 1913, X, 764. ' Turner, J. Inst. Met., 1912, vii, 105. 32 METALLURGY OF COPPER The electric conductivity of brass has been studied by Pushin and Rjaschsky.i Upon heating in air, alloys with over 63 per cent. Cu show irridescent colors. The behavior, with acids, of alloys with < 50 per cent. Zn is similar to that of Cu; alloys with > 50 per cent. Zn are readily dissolved in acids which attack Zn but not Cu. The product of electrolytic corrosion^ of brasses with > 50 per cent. Zn is practically Zn; with < 50 per cent. Zn it has the same composition as the alloy; or a, a+)3, and /3 brasses yield products of the same composition as the brasses; 7 crystals diminish the corrosion; 7+e, e and r; brasses yield as prpduct pure Zn. The fact that most industrial brasses lie within the range of a or the conglomerate a-\-fi, makes them resistant to corrosion.' Brasses are likely to occlude gases* such as CO2, CO, and H.* In the manxifacture^ of brass the purity of the metals,' the apparatus used, and the temperature and time given to fusion and pouring, all have an influence upon the physical properties of the alloy. The in Cu may have a harmful influence' in that it oxidizes Zn and causes infusible salamanders to form in the crucible. The presence of o.oi per cent. = 0.09 per cent. CU2O is harmless; good sheet brass has been rolled with copper containing 0.55 per cent. = 4.91 per cent. CU2O, although the figure is excessive. According to Sparry as little as 0.02 per cent. Sb' or 0.02 per cent. Bi,^" or 0.06 per cent. Te^^ makes common brass (60 Cu, 40 Zn) brittle so that it cannot be rolled without showing cracks; 0.8-0.9 P^' cent. Pb^^ causes no harm, but with i per cent. Pb trouble arises. An addition of 1.5-2 per cent. Pb'' makes (screw or clock) brass sufficiently brittle to cut well with short chips. Similarly As, to the extent of 0.02 per cent.,'* begins to affect the malleability; i per cent. Cd'^ appears to have no harmful influence except that ' Zl. anorg. Chem., 1913, Lxxxn, 50. J. Inst. Mel., 1913, X, 420. 'i Lincoln- Klein-Howe, /. Phys. Chem., 1907, xi, 501. Briihl, /. Inst. Met., 1911, vi, 279. ^ Diegel, Stakl u. Eisen, 1899, xrx, 170, 224. Jones, Met. Ind., 1905, in, 171. Sexton, Eng. Mag., 1905, xxx, 211. Desch-Whyte, /. Inst. Met., 1913, x, 314. * Guillemin-Delachanal, Rev. Met., 1911, viii, i. 'Lewis, Proc. Chem. Soc, 1912, xxviii, 290; /. Inst. Mel., 1913, ix, 217. "Brass Foundries: Foundry, 1902, xx, 142; 1903, xxiii, 169; 1906, xxviu, 131; 1907, xxxi, 176, 28s; Met. Ind., 1908, vi, 341; Iron Age, 1912, lxxxix, 1257; Mel. Ind., 1913, xi, 155. ' Carpenter, /. Inst. Met., 1912, viii, 59. ' Sperry, Tr. A. I. M. E., 1900, xxx, 937. Jolibois-Thomas, Rev. MU., 1913, x, 1264. " Tr. A. I. M. E., 1898, xxvm, 176; Brass World, 1907, in, 297. " Sperry, Tr. A. I. M. E., 1898, xxvm, 427. Carpenter, J. Inst. Met., 1912, viii, 60. '' Tr. A. I. M. E., 1903, XXXIII, 682. '^ Guillet, Rev. Mil., 1906, in, 273. Johnson, /. Inst. Met., 191 2, vii, 201. Carpenter, op. cit., 1912, vni, 63. 1' Sperry, Tr. A. I. M. E., 1897, xxvii, 485. " Sperry, Brass World, 1906, 11, 163. '' Op. cit., 1907, III, 1211. INDUSTRIAL ALLOYS j,z it hardens the alloy. S' makes brass pasty and is thus a cause of dirty castings; such pasty brass showed 0.69 per cent. S. Brass with 0.03 per cent. S rolls as well as common brass; it is not made red-short by S as is the case with Cu. The effects of various impurities upon the constitution of brasses, especially upon the structure of the /3-constituent, have been investigated by Carpenter. ^ Manufacture of Brass. — Brass is produced' by melting together Cu and Zn in a crucible furnace^ which is usually fired with anthracite, coal or coke, sometimes with liquid, but rarely with gaseous fuel.* Electric induction furnaces are being advocated.^ Oil- or gas-fired reverberatory furnaces are employed for melting ingot brass or bundled scrap and borings' after the iron has been re- moved by a magnet or the raw material first purified by washing.* Experi- ments have been made to produce brass from ZnS and Cu.' In a crucible furnace the warmed Cu is charged first, melted under a i-in. charcoal cover and kept just above its freezing-point. Then the necessary Zn, previously warmed, is added in several portions in order to prevent chilling of charge (the heating-up of which would cause much loss in Zn) ; the whole thor- oughly stirred; brought quickly to the right temperature (100-200° C. above the melting-point; overheating causes oxidation, Zn begins to burn) ; skimmed or not; and poured into a suitable mold (sand, cast-iron, or bronze) to furnish an in- got to be sold or a plate to be rolled. In either case the alloy is chilled by spray- ing with water. Brass shrinks about yV in. per foot, hence the cores are made soft."^" Fluxes, ^'^ such as borax, are little used; sometimes salt''' is added as a wash, although it assists the volatilization of Zn, and a deoxidizer" such aso.ox per cent. Mg. An ordinary charge weighs from 50 to 200 lb.; 100 lb. is melted in about two hours; the loss in Zn may reach 6 per cent." and has to be taken into account in making up the mixtures;'* i lb. coke will melt about 2 lb. alloy; I man will operate 4 to 6 furnaces. The manufacture by electro-deposition is confined to plating.'^ ' Sperry, op. cil., 1906, 11, 307. ^ J. Inst. Met., 191 2, vni, 59. ' Stahl u. Eisen, 1913, xxxiii, 522. * Horner, Foundry, 1913, xli, 113, 119. ' iiTrom, " Development of Melting Furnaces," Met. Ind., 1909, vii, 287, 324, 358, 404, 436; 1910, VIII, 80. « Clamer-Hering, Met. Chem. Eng., 1912, x, 702; Foundry, 1912, XL, 483; Brass World, 1912, VIII, ss. Hansen, Met. Chem. Eng., 1912, x, 703. ^ Brass World, 1910, vi, 345; op. cit., 1912, vni, 421. » Wittich, Eng. Min. J., 191 2, xcv, 853- 'Bensel, Melallurgie, 1912, ix, 523. " Chamberlain, Volume changes, /. Inst. Mel., 1913, x, 193. " Krom, Met. Ind., 1910, viii, 203. "Sperry, Brass World, 1912, vili, 307. " West, O in Cu and brass, /. Inst. Met., 1913, x, 371. "Bassett, /. Ind. Eng. Chem., 1912, rv, 164. " Sperry, Mixtures, Brass World, 1912, viii, 41, 83, 121, 167, 204, 239, 28S, 317. Burkey, "Treatment of Brass Scrap," Eng. Min. J., 1913, xcvi, 486. " De Kay Thompson, Met. Chem. Eng., 1912, x, 458. 3 34 METALLURGY OF COPPER Industrial brasses may be classed as regular and special; the former represent the binary alloys, the latter the Cu-Zn alloy with additional metals to furnish special properties. 27. Regular Brass.— There is a great variety in the regular brasses to meet the numerous requirements of the arts. The leading brasses are given in Table 8. The usual range of composition lies between 90 and 35 per cent. Cu; the most important alloys are those containing from 70 to 55 per cent. Cu. Alloys with more than 64 per cent. Cu are composed solely of a-solution, while those between 64 and 55 per cent Cu are made up of a and ^ when quenched above 470°; of a+y when cooled very slowly; a-alloys are rolled or drawnV cold; a-jS (a-y) alloys have to be rolled hot. Fig. 27. — Rolled common brass which Fig. aS.^Rolled common brass which worked well. worked unsatisfactorily. In Table 8 the brasses are divided into four classes: class I, so-called "High Brass" is suited especially for cold-rolling; class II, the standard common metal, can be rolled either cold or hot; class III, so-called "Low Brass, " can be rolled hot only. Annealing^ at 420° C. begins to cause rearrangement of distorted crystals; heating to 600-700° C. removes all internal strains; at 800° C. there is danger of burning the alloy. ^ Fig. 27 gives the structure of rolled common brass that worked well. Fig. 28 that of one which furnished an unsatisfactory plate; the former has a close fine grain, the latter (badly annealed) is made up of large crystals, and is weak. Other photomicrographs are given by Lewis,* Bengough- 1 Grard, iJeti. Mel., 1909, vi, 1069 (London Congress); Metallurgie, igio, vil, 651; Proc. Inst. Assoc. Test. Mat., New York, 1912, 11, 15. Diegel, Verh. Verein. Beford. Gewerbefl., 1906, lxxxv, 177; Metallurgie, 1906, in, 568. Krom, Met. Ind., igio, viii, 8, iii, 157, 342, 375, 459, 499; 1911, rx, 27, 123, 127; 1912, X, 20, 118, 331; i9i3,xi, 18,337. Stilson, Eng. Mag., 1913, XLV, 239. "Moore, "Annealing Furnaces," Met. Ind., 1910, viii, 45. 2 Charpy, "Contributions k I'^tude des alliages," Paris, 1901, p. i. * Engineering, 1903, Lxxvi, 753; Met. Ind., 1903, i, 33; 7. Soc. Chem. Ind., 1903, xxn, 12. INDUSTRIAL ALLOYS 35 -I O O oV .BV o en o *f o ■if o 5^ - 0,5 +3 o •(:; U M £ 3 0,'^ OJ V ■ "g ^ i .a I g a § •S t5 -w o u > M ■S ^ ^ 3 ja J3 XI iC o- o " " ^ • -•S "5 ^ Ifc a .S - 1^ ? O - -S-^-s r-* ^J^ O w a" l-H z Metal Muntz ," Bras 1-1 Munt upon rocess H ° M 71 tu^ c T) a "^ 2 5 '3 r^ CO l-l y and Co :t of Sb, sion or S > M C3 u 3 '"' tQ iJ tH a ? eq ?="w fe H" «0 2 Haml Johns Sperr 36 METALLURGY OF COFFER Hudson/ Carpenter-Edwards^, Bengough.' Class IV, or " WhiteBrass," includes the alloys that cannot be rolled. The so-called " Cast Brass " includes the range of composition occupied by classes II, III, and IV. It usually undergoes no mechanical treatment, hence it need not be so pure as the alloy that is to be rolled or drawn; in fact, the impurities present or purposely added make it run more smoothly, fill the mold more evenly, and permit machining more readily. 28. Special Brass. — The leading special brasses^ are aluminum-, iron-, man- ganese-, tin-, and nickel-brass. (a) Aluminum Brass (Hercules Metal) does not contain over 4 per cent. Al, as the presence of a larger amount renders the alloy difficult to work. The usual range of composition is Cu 67-71 per cent., Zn 31.75-25.50 per cent, Al 1.25-2.50 per cent. "^ Carpenter-Edwards^ have investigated that part of the ternary system Cti-Zn-Al which is richest in Cu. They find that there is no ternary eutectic, that, the larger part of the liquidus surface consists of two areas corresponding to a- and j3-solutions, that the a-alloys undergo no transformation on cooling, but that the j3-alloys are resolved into a +7, and that a thermal change takes place at 7oq° C. They conclude that starting from the Cu-Al-side toward the Zn- side the mechanical properties will change not suddenly, but uniformly and progressively. Guillet,' who studied this series of alloys microscopically, concludes that i lb. Al can replace 3.5 lb. Zn; an alloy with 38 per cent. Zn and 2 per cent. Al, e.g., has the same structure as one with 45 per cent Zn. The alloys are fine- grained, and give good castings which should be cooled slowly. They can be worked at lower temperatures than the corresponding Zn-Cu alloys being more malleable. The mechanical properties of a f^ and a 1^ aluminum brass annealed are shown in Figs. 29 and 30. Their resistance to corrosion is dis- cussed by Rowland.' The addition as a deoxidizer of 0.05 per cent. Al to ordinary brass that is to be cast in sand is helpful in obtaining clean castings; with castings to be made in metal molds the addition of Al is to be avoided.* (b) Iron-brass. — The addition of up to 3 per cent. Fe to brass strengthens and hardens the alloy, increases the malleability when hot, and the resistance to corrosion. The constitution of these alloys has not been studied; the eSect of Fe upon the structure has been investigated by Carpenter.* 17. Instil. Metal, 1909, i, 89, 1910,2 iv, 92; J. Soc. Chem.Ind., 1908, xxvn, i (Muntz Metal). ^J.Inslit. Met., igii.v, 127; 1912, vii, 70; 1912, vm, 51. 'Op. cit., 1912, VII, 123. ' Rosenhain, /. Inst. Met., 1912, vii, 191. "Internal. Zt. Metallographie, 1912, 11, 209; Rev. Mil., 1913, x, 429; /. Inst. Mel., 1912 vra, 322; comment by Guillet, Rev. Met., 1913, x, 463. «i?CT. Mit., jgos, n, iii, 1906; m, 254. '/. phys. chem., i9o8,,xii, i8q. * Sperry, Met. Ind., 1903, i, 35. "/. Inst. Met., 1912, vni, 66. INDUSTRIAL ALLOYS 37 Sterro, Aich and Deltametal, Tobin Bronze, and Duranametal are indus- trial names for this class of alloy. Sterrometal^ has the following range of composition: Cu 60, Zn 38-38.5, Fe 2-1.5; it represents a |^ brass in which part of the Zn has been replaced by Fe; it has a tensile strength of 50,000-70,000 lb. per square inch, with an elongation of from 11 to 39 per cent, in 8 in.'' Sometimes a small amount of Sn is added to improve its quality; such an alloy contains Cu 55-60, Zn 34-44, Fe 2-4, Sn 1-2, and has a tensile strength ranging from 43,000 to 82,000 lb uo 100 90 80 70 60 60 .40 30 20 10 ^N \ \ \ \y A \ Ten Stre Pffth ^ \ 1 -EIoue atlou \ \ N N ►J 35 c S 30 ^ J 25 a "S 20 1 1B| I ra 10 H c 6 H Cm. 60.0 B9.6 69.9 69.6 60.4% Zn.iO.O 40.1 40.3 38.6 35.9% Al. 0.0 0.3 0.8 2.9 4.7% 100 90 80 70 60 60 40 30 20 10 / V / VeI )ngatl on / \ \ \ \ \ } y A / \ / \ \ y< Ten lie Stre igth Cu. - 0.0 69.0 70.0 70.5 70.1% Zn. 30.0" 80.6" 29.1" 26.4 24.7% Al. 0.0 0.4 0.9 3.1 5.2% Fig. 29-30. — Mechanical properties of fS and JJ aluminum brass. ' These figures have been changed slightly from the originals to bring up the totals to 100 per cent. per square inch. An alloy Cu 55.04 per cent., Zn 42.36 per cent., Fe 2.77 per cent., Sn 0.83 per cent., gave tensile strength, cast 40,320 lb., forged 76,160 lb., cold-drawn 40,320 lb.* Aich Metal resembles sterrometal. Hiorns'' gives Cu 58-60 per cent., Zn 36-41 per cent, Fe 0.74-1.74 per cent., Sn 0-1.02 per cent, as the range of composition. Delta Metal contains usually Cu 55 per cent., Zn 41 per cent., Fe 3 per cent., Mn, etc., I per cent. Tetmayer's tests* gave tensile strength, cast 44,000 lb. per square inch and elongation 30-40 per cent, in 7^ in. These alloys are said to resist corrosion better than ordinary brass. In the manufacture, Fe is intro- duced by using iron-bearing Zn, rarely iron-bearing Cu. Tobin Bronze.^— Two analyses gave Cu 59.00, Zn 38.40, Sn 2.16, Fe o.ii ' Guillet, Rev. Ma., 1906, iii, 264. 2 Thurston, "A Treatise on Brasses, Bronzes, etc.," 1900, p. 415- 'Op. cit., p. 368. * "Mixed Metals," Macmillan, N. Y., 1913, p. iS9- ^ Schweizerisches Gewerbeblall, June 8, 1889. • Garrison, /. Frankl. Inst., 1891, cxxxii, 55. 38 METALLURGY OF COFFER Pb 0.31 and Cu 61.20, Zn 37.14, Sn 0.90, Fe 0.18, Pb 0.35; the tensile strength showed 78,500 lb.; the elongation 15 per cent, in 2 in. and 40.5 per cent, in 8 in. The original Tobin bronze* contained Cu 58.22, Zn 39.48, Sn 2.30; it showed a tensile strength cast of 66,000 lb., rolled 79,000 lb., and cold rolled 104,000 lb. per square inch. Durana MetaP contains Cu 64-78 per cent., Zn 29.50, Fe 1.51, Al 1.70, Sn and Sb 2 . 20. The tensile strength is 82,000 lb.; elongation 14 per cent., and elastic limit 70,000 lb. (c) Manganese-brass.' — In the trade these alloys often go by the name of manganese-bronze,* a name that ought to be reserved for Cu, Sn, Mn alloys. Mn appears to harden brass, to increase the tensile strength, and to diminish the elongation; i lb. Mn can take the place of o. 5 lb. Cu.* However, the yel- lowish alloy called Parson's manganese-bronze contains only from a trace to o. ox per cent. Mn as seen by the two following recent analyses:* cast metal, Cu 57.30, Zn 40,44, Sn 1. 01, Fe 0.79, Pb none, Al 0.46, Mn trace; sheet metal, Cu6o. 17, Zn37.47, Sn o. 99, Fe 1.24, Pb trace, Al none, Mn 0.02. Itsmechan- ical properties' cover the following range: tensile strength 81,500-90,500 lb. per square inch; elastic limit 39,300-44,750 lb.; elongation 40-26 per cent.; and reduction of area 47.5-33.0 per cent. The effects of the pouring temperature upon the size of grain and thereby upon the mechanical properties have been studied by Gillett.* The alloy is used for propeller blades, parts of guns, carriages,«automobiles, valve-stems, shafting of motor-boats, etc. The specifi- cations of the U. S. Government Bureau of Steam Engineering, of July i, 1910,' call for Cu 57-60, Zn 37-40, Sn 0.75, Fe 2, Al not >2, Mn not >2 per cent., ultimate strength not > 70,000 lb. per square inch, elongation in 2 in. not < 20 per cent. (d) Tin-brass. — According to Johnson'* Sn is only slightly soluble in a-, but readily so in /3-brass (cast) ; rolling and annealing help the solution in the a-constituent of a 1^ brass. Carpenter'^ found that i per cent. Sn greatly favored the formation of the 7-constituent. These alloys contain Cu 60-62, ' Thurston, Tr. Am. Soc. Civ. Eng., 1881, xi, 1309. ' Knorre, Zt. angew. Chem., 1894, 238. 3 General, Foundry, 1905, xxvi, 116; Mdctdres, Met. Ind., 1909, vii, 173; 1910, x, $; Casting, Mel. Ind., 1903, i, 131; 1910, vni, 410; 1911, ix, 4, 73; Brass World, 1905, i, 153; 1910, VI, 79; Foundry, 1905, xxvi, 87; 1912, xl, 487; Tests, Met. Ind., 1907, vii, 175; Wire, Brass World, 1905, i, 255; Specipications, Met. Ind., 1909, vii, i. ^ Corse-Skillmann, "History" Met. Chem. Eng., 1914, xii, 113. ' Guillet, Rev. Met., 1906, in, 258. ° Met. Ind., 1909, vii, 173. ''Op. cit.,p. 175. * Tr. Am. Instit. Met., 1912, vi, 207. ^■Brass World, 1910, vi, 398. '° Year Book, 191 1, p. 135. •' /. Inst. Met., 1912, VII, 201. '2 Op. cit., 191 2, VIII, 65. INDUSTRIAL ALLOYS 39 Zn 37 . 5-39, Sn i-i . 5 per cent. Guillet' found that i per cent. Sn replaced about 1.5 lb. Zn, but the amount of Sn has to be kept below 4 per cent., as otherwise the alloy becomes brittle. The main advantage of an addition of Sn is an increased resistance to corrosion. In the manufacture of the alloy the Sn is introduced into the stream of Cu-Zn as it flows from the crucible. Capp^ found that the data for elastic limit obtained by the usual methods of testing are unreliable for this class of alloys, as well as for brasses and bronzes in general. The effects of Mn, Si, Cr, Wo and Va on brasses have been summarized and described by Escard,' those of Cr and Va by Carpenter.* As little as 0.04 per cent. Va^ reduces the electric conductivity, but increases the elastic limit, tensile strength, and ductility from 10 to 20 per cent. According to Gin* the Ruebel alloy is prepared by melting together Cu 45-57 parts, Zn 40, and (Va, Cu, Al, Fe) 3-15. The commercial Cu-Va alloy contains 3 per cent. Va, but is difficult to obtain free from Fe and Al. (e) Nickel-brass. — The mechanical properties of brass are improved by an addition of Ni;'' it can replace 1.2 parts of Zn. Fig. 31. — Alloy-series Cu-Sn, bronze. Guillet,* who has studied the effects of Ni upon brass, finds that small additions of Ni make brass easier to work cold, and that additions up to 10 per cent, improve the mechanical properties. 29, Bronze' (Cu and Sn). In General.— The complicated constitution of ' Rev MH., 1906, m, 264. *7. Am. Soc. Mech. Eng., igio.xxxii, sjy.Iron Age, igio,lxxxvi,62&; J. Inst. Met., igio, IV, 310. 'Ginie Civil, 1909, LV, 74, 85; Oest. Zt. Berg. Hiiitenw., 1910, LViii, 201, 215. *J. Inst. Met., 1912, vm, 165. 'Norris, /. Prankl. Inst., 1911, CLXXI, 580. " Melall-Erz, 1913, x, So2- ' Guillet, Compt rend., 1912, clv, 1512; /. Inst. Met., 1913, dc, 213. '^Rev. MH., 1913, X, 1 130. » Thurston, op. cil. 40 METALLURGY OF COPPER bronzes has been studied by Stansfield/ Heycook and Neville,^ Roberts-Austen,^ Sheperd and Bio ugh,* Giolitti and Tovante.^ Following the freezing-point curve of Sheperd and Blough, shown in Fig. 31, bronzes may contain five solid solutions, a, fi, 7, 8, c, and the chemical compound CusSn; the compositions and colors are given in Table 9. Table 9. — Constituents op Bronze Constitu- ent Nature and Composition Color a Solid solution of Cu and Sn with from to 13 per cent. Sn Reddish-yellow to yellow. P Solid solution of Cu and Sn with from 22 to 27 per cent. Sn Yellow. y Solid solution of Cu and Sn with from 27 to 57 per cent. Sn White. s Solid solution of Cu and Sn with from 24 to 33 per cent. Sn White. f Solid solution of Cu and Sn with from 33 to jg per cent. Sn White. CnSn Chemical compound, 61.5 per cent. Cu, 37.5 per cent. Sn White. The liquidus is shown mA,B, C, D, E, F, G, the sohdus iQA,bi,b,Ci,d,d7, e^, ei, e, F,H. The six components and their combinations give bronzes characteristic structures, represented by fields i-xviii. Field i is the region of pure a-crystals; in field 11, a-crystals are stable in contact with the mother-metal; in 111 crystals Fig. 32. — Alloy chilled at 777° C. Fig. 33. — Alloy cooled slowly to 546° C. and then chilled. Figs. 32-33.^Cast bronze with 15,6 per cent. Sn; a-crystals light, /3-crystals dark. a and P form a conglomerate; at 486° C. |8-crystals break down into a and 5, and furnish the stable forms a and S for field iv; in field v, /3-crystals are stable in contact with mother-metal; field vi is the region for pure |3-crystals, as is vn for a conglomerate of P+y, and viii the region for pure 7; the solid solution 3 in field XII (once considered to be Cu4Sn), formed by a transformation in the solid of 7-t-Cu3Sn in field rx, is stable below 600 ° C. and so on. The meaning of the two horizontal lines at 218 and 182° C. in field xvi has not yet been interpreted. ^ Third Rep. Alloys Research Comm., 1895, 269. ^Phil. Trans. A., 1897, clxxxdc, 42. ^Fourth Rep. Alloys Research Comm., 1897, 67; Suppl. by Campbell Fifth Rep., 1901, 1211. * /. phys. chem., 1906, x, 630. ^Gazz. chim. ital., 1908, xxxviii, 2, 209; Rem. Met., 1909, vi, 476. INDUSTRIAL ALLOYS 4.1 Two photomicrographs of Heycock and Neville, Figs. 32-33, show light a- and dark ^-crystals in an alloy with 15.6 per cent. Sn. The alloy represented by Fig. 32, was chilled at 777° C; that by Fig. 33 slowly cooled to 546° C. and then chilled. The range of composition of the bronzes that are of importance in engineering is much smaller than that of brasses; 80 per cent, represents the lowest figure for Cu, or 70 per cent, if bell metal be included. The transformations that are possible in the region 100 to 70 per cent. Cu show that the physical properties of these alloys must be considerably affected by heat treatment. The ultimate strength and elongation^ of bronzes, both cast and annealed, are shown in Figs. 34 and 35. In Fig. 34 the tensile strength is seen 86 80 75 Per Cent Copper Heated to low red, water-quenched. Held one week at S4o° C, water-quenched. Tested as cast. Held one week at 400° C, furnace-cooled. Fig. 34. — Tensile strength of cast bronze. 60 W SO S S20 to increase with an addition of Sn until the maximum is reached with about 80 per cent. Cu; the strength of an alloy with 70 per cent. Cu is very small, falling to about 16,000 lb. per square inch. Heat treatment^ does not affect alloys with from 100 to 86 per cent. Cu, as these are homogeneous, consisting ex- clusively of a-crystals. With alloys containing from 86 to 76 per cent. Cu the case is different. Annealed at 400° C. (curve D) they consist of a — |- S-crystals (Fig. 31) ; while holding at 540° C. and then chUling in water (curve B) has changed a + S into a + j3. The ff alloy with a+5 structure shows a tensile strength of 45,000 lb. per square inch; the same with a-(-/3 structure, one of 67,000 lb. The curve C for cast bronze with from 86 to 76 per cent. Cu Ues between curves B and D, as the cooling was so quick as to allow only part of a-\-5 to change into a+iS, hence such a cast bronze can contain all three constituents a, (3, and 5. An alloy with 70 per cent. Cu may contain a, /?, 7, and 6 crystals, depending upon the ' Sheperd- Upton, /. phys. chem., 1905, ix, 441; MetaUurgie, 1906, in, 29; Rev. Met., 1906, ni, 8. ' Grenet, Rev. MM., 1911, viii, 108; MetaUurgie, 1911, vm, 543. / -^ ~~^ y ys- k \ «* ^-^^ ^<-. 10 100 Fig. 3S.- 96 90 85 Per Cent Copper -DuctUity of cast bronze, water-quenched. 80 76 Aqua = 42 METALLURGY OP COPPER rate of cooling. The constituent 5 makes the alloy brittle; whenever it forms more than 70 per cent, of the alloy the strength decreases rapidly. In Fig. 35 the differences in ductility between cast and annealed bronzes are clearly shown. In cast bronzes with from 100 to 80 per cent. Cu the duc- tility, excepting a slight rise, decreases with an increase of Sn; heating such ^ 30 20 ti ^ teSi p^ 1[1, ^ ~^ V \otv ^ ^v. , ( r^ o-a'i'^ \: A \ \\ Bias LicL mit •n ^. "■- ^^ A — '^ >L \ 100 200 300 400 600 600 Temperature of Annealing:— Desreea Centigrrade Kotoj -OlratatejciflfLo tendjenc^ Lndloated by dotted lines 700 800 Fig. 36. — Mechanical changes of V" bronze at varying temperatures. bronzes to 540° C. and then quenching in water increases the ductility by 5 per cent. The greatest ductility is reached with a bronze of 90 to 88 per cent. Cu, a composition which lies very close to the maximum of Sn in a-crystals. The changes in the leading mechanical properties of cast bronzes, of the corn- when tested between zero and 800° C. are shown positions ^, -V and \\, 900 u 100 200 300 400 500 600 700 Xemperatuce of Annealinsr- Deerees Centrsrade Note; Oharnoterlstlc tendency Indicated by dotted line Fig. 37.— Mechanical changes of V" bronze at varying temperatures. in Figs. 36, 37 and 38.1 The alloys -«/ and -«/ have a as sole constituent, but show a difference in behavior when pulled in the testing machine. In the \\ alloy, consisting when annealed of a+5, the change into a+jS near 500° C. is » Guillet, L., "Trempe, Recuit, Revenu," Dunod, Paris, 1909, p. 572. INDUSTRIAL ALLOYS 43 clearly marked by the mechanical tests. ^ The relation between mechanical properties and heat treatment of drawn bronzes has been studied by Goerens- Dumont,^ and Guillet,' some physical properties by Wyss/ the specific heats by Chappell.* Bronzes occlude little gas;* they resist corrosion, less when rich in Cu than when rich in SnJ GioUiti and Ceccareli,' studying bronzes with up to lo per cent. Sn, found that heat treatment affected the corrosion of the a-solution, that the a+j3 alloy was more quickly attacked than the a 60 40 g Si 20 10 ^. -— _ A A / ^A " — — I. ntittu te S) rengt h y / Ela Stic I limit ^" --- ■->r' ^ ^^ ^ / -- V J % Elongs Ltion J 100 700 800 200 300 400 500 600 Temperature of Annealing -Degrees Centigrade Note: OharaterlstLo tendency Indicated by dotted lines Fig. 38. — Mechanical changes of \\ bronze at varying temperatures. alone, and that the greater the difference in composition between center and edge of a crystal the more rapid the attack. Impurities greatly affect the prop- erties.9 Shrinkage^" is lessened by Zn,ii increased by Co, Al, Si, Fe, and Ni.'^ Tensile strength is considerably lowered by Sb or much Zn;" it is raised by Co, Ni, Mn, and Fe; it is lowered by a rise in temperature.^^ Machining is made ' Portevin, Jfeji. MU., 1903, x, 677. Robin, op. cit., 1913, x, 764. ^ Ferrum, 1912-13, x, 21. ' Rev. Mil., 1913, X, 769. * Ferrum, I9i2-i3,x, 167. ^Op. cU., 1913, X, 271. « Guillemain-Delachanal, Rev. Mit., 1911, viii, i. ' Carpenter-Edwards, Met. Chem. Eng., 1911, ix, 63. ^Gazzetta chim. Ital., xxxrx, 557; -''• If"^- ^^i-> iQ". vi, 333- "Miller, Metallurgie, 191 2, ix, 63. '» Turner-Haughton, /. Inst. Met., igij, vi, 192. " Wust, Metallurgie, 1909, vi, 769; Irak Age, 1910, lxxxv, 790. " Chamberlain, /. Inst. Met., 1913, x, 193. " Guillet-Revillon, Rev. MM., 1910, vii, 429; Metallurgie, 1911, vra, 582. "Johnson, Met. Chem. Eng., 191 1, ix, 399. 44 METALLURGY OF COPPER easier by Sb and Pb and more difficult by Mn and Ni; Pb in excess of 0.15 per cent, afiects the strength; and leady bronzes^ are readily attacked by boiling water and steam. Patina formation is lessened by Zn and Al, intensified by Co, Ni, Sb, Fe, Si, and P. Fe gives the alloy a lighter color. Hardening is discussed by Grenet,^ the wearing qualities by Portevin-Nussbaum.^ In the manufacture, oxidation has to be avoided. Heyn and Bauer* found that CU2O was readily reduced, 2Cu20+Sn = 4Cu+Sn02, the Sn02 separating in large crystals which, insoluble in the alloy, rendered it less fluid. Jolibois- • Thomas^ made similar observations. Large charges are melted in reverberatory furnaces, small ones in crucibles;* in either case oxidation has to be avoided. Even a crucible and a charcoal cover do not absolutely prevent oxidation, as some CU2O is formed. This is most pronounced with alloys containing over 84 per cent. Cu. A powerful reducing agent, such as P or preferably P-Sn, is sometimes used to counteract the oxidation. This alloy, ^ containing about 5 per cent. P, is prepared by charging a graphite crucible with stick P, covering with i in. of charcoal, filHng with granulated (flake) tin (r P: 10 Sn), giving a charcoal cover, putting on a cover and luting it, and bringing gently to a low-red in a pot-furnace. The P being volatilized is taken up by the Sn. When the flame of burning P disappears, the charge is finished, the alloy stirred, skimmed, and poured into small ingot molds set in water. Portevin^ found that in making some castings in a strongly reducing atmosphere, the ingots became porous, and required the addition of an oxidizing agent. Table 10. — Regular AND Special Bronzes Class Name Composition, per cent. Color Fracture Malleability Cu 1 Sn 1 P Si 1 Mn ductility. Malleable bronze.. . Gun metal Bell metal 98-94 92-88 80-75 70-65 2-6 8-12 20-25 30-35 Red to reddish yellow. Reddish yellow to dirty yellow. Yellowish gray to gray. Ash-gray to white. Vesicular. Crystalline to fine- grained. Fine- grained. Conchoidal. hammered. Difficult to roll, hard. Cannot be rolled, hard. Brittle, steely, susceptible of perfect polish. bo « Speculum metal. . . . Phosphor bronze. . . Silicon bronze Manganese bronze.. 90-91 88 9-8 10 9 per cent. Sn show the 3-constituent and the compound CusP which may form a eutectic mixture; Hudson-Law' believe they have found a ternary eutectic. According to the chemical and mechanical investigations of Philip,* phosphor bronzes may be grouped according to their uses in three classes: Heavy castings: Cu 90-92, Sn 7.4-9.7, P 0.3-0.6, tensile strength >34,ooo lb., elongation 20 per cent, in 2 in.; Rod, Sheet, Wire: Cu 91.5-97.5, Sn 8.4-2.25, P 0.1-0.25, when unannealed tensile strength 60,000 lb., elongation 10 per cent, in 2 in., when annealed 40,000 lb. and 40 per cent, in 2 in.; Bearings: Cu 84.5-89.1, Sn 14.5-10.1, P 0.8-1.0 and possibly higher. The specifications of the U. S. Bureau of Steam Engineering of July I, 1910, call for Cu 80-90, Sn 6-8, Zn 2-14, P 0.30, Fe // / y 0.18 ■ / y y A -y /,,, "^ 0.14 X 20 40 60 Temperature 60 Fig. 42.— Conductivity of NaCl-solutions saturated with Cu2Cl2(I, II, III), and of CUSO4-H2SO2 solutions (A, B). 41. Cupric Chloride (CuCla 47.22 per cent. Cu; 63.6 Cu+71 Cl2 = 134.6 CuCl2+Si>4oo cal., in dil. solution 62,500 cal), does not occur as a min- eral. The anhydrous salt is formed by the action of CI upon Cu or CuCl, of HCl upon powdery CUSO4, and of heat and NaCl upon CUSO4. It is a brown to brownish yellow powder, melts at 498° C, is changed at 340° C. with exclu- sion of air into CuCl+Cl, with access of air partly into CuCl+Cl, partly into ' Abegg, "Handbuch der anorganischen Chemie," pp. 419, 556. COPPER COMPOUNDS S9 CuO and CI;' is deliquescent and becomes green. loo gm. H2O dissolve at 0° C— 70.6 g. CuCU; at i7°-7S.6, at 31.5°— 80.8, at 91°— 104 g. CuClj. The hydrous CUCI2.2H2O is formed by dissolving Cu in aq. reg., or CuO in HCl, by the reaction 2NaCl+CuS04 = Na2S04+CuCl2; the salt is light-blue forming a greenish solution when concentrated. The solution has a decomposing effect upon metallic (Fe, Co, Zn, Cd, Pb,^ Ni, Sn, As, Sb, Ag) sulphides, forming metallic chloride and Cu^S ; it is reduced to CuCl by H2SO3, viz. 2CUCI2+H2SO3+H2O = Cu2Cl2+2HCl-|-2H2S04 and by boiling with metallic copper, CuCl2-|-Cu = CU2CI2. The metal is precipitated by Fe, Hg, and Ag, which are converted into chlorides, and by FeO, converted into Fe203 and FeCU; the sulphide by CuS with the separation of S, and by Ag2S with separation of S and formation of AgCl. KOH precipitates Cu(0H)2; H2S separates CU2S and S. > Kothny, Oest. Jahrb., 1910, LVin, 141. 2 Hunt, Tr. A. I. M. E., 1881-82, x, 12. CHAPTER VI COPPER ORES 42. In General. — The minerals fonning copper ores are quite numerous; they are classed as sulphide, oxide, and native, and form the basis of the classi- fication of copper ores.' 43. Sulphide Copper Ores. — The sulphide minerals are: Chalcocite (vitreous copper, copper glance), CU2S, 79.8 per cent. Cu.; covellite, CuS, 66.4 per cent. Cu; bornite (peacock ore), 3Cu2S.Fe2S3 or CuaFeSs, range 50-70 per cent. Cu, formula: 55.5 Cu, 16.4 Fe, 28.1 S; enargite, 3CU2S.AS2S5 or CU3ASS4, 48.3 per cent. Cu, 19. i As, 32.6 S; chalcopyrite, Cu2S.Fe2S3 or CuFeS2, 34.5 per cent. Cu, 30.5 Fe, 35.0 S; tetrahedrite (gray copper, fahlore), 4RS.Sb(As)2S3, R = Cu2, Fe, Zn, Ag2, Hg2, range 15-48 per cent. Cu; 4Cu2S.Sb2S3, 52.1 per cent. Cu, 24.8 Sb, 23.1 S; tennantite, 4CU2S.AS2S3, 57-S P^r cent. Cu, 17.0 As, 25.5 S; chalcantite (blue vitriol), CUSO4+S aq, CuO per cent. 31.8, SO3 32.1, H2O 36.1; Cu 25.4. To this list must be added pyrite and marcasite, FeS2, 53.4 per cent. S, and pyrrhotite, FenSn+i, range FejSe— FeieSiy, chiefly FeuSi2, 38.4 per cent. S., both of which are frequently copper-bearing to the extent perhaps of 5 per cent, through intermingled chalcopyrite, or sometimes tetrahedrite. A characteristic metallurgical difference^ between pyrite and marcasite is that the latter is usually more free-burning and more easily vitriol- ized than the former. The leading sulphide copper deposits in the United States are situated in Montana, Utah, Nevada, California, and the Atlantic Coast beds. The last^ are massive pyrrhitous deposits, with from 2 to 5 per cent. Cu, occurring in strata extending from Newfoundland to Alabama. The ore is generally first roasted in kilns of sulphuric acid plants before it is treated for copper. In Montana,* in and around Butte, occur rich ores in shattered and altered granite. The copper minerals^ are chalcocite, bornite, enargite, and pyrite; covellite, tetrahedrite, and chalcopyrite are subordinate. Up to 1900 1 Weed, W. H., "The Copper Mines of the World," McGraw-Hill Book Co., New York, 1907. 2 Brown, Proc. Am. Phil. Soc, Philadelphia, 1894, xxxin, 225; Zt. prakt. Geol., 1895, iii, 180. Stokes, Butt. 186, U. S. Geol. Surv., 1901. ' Kemp, J. F., "Ore Deposits of the U. S. and Canada," New York, 1900, p. 185. Wilson, W. G., "Pyrites in Canada," Canada Dep't. Mines, Ottawa, 191 2. " Weed, H. S., Profess. Paper, No. 74, U. S. Geol. Surv., 191 2. Sales, Tr. A. I. M. E., 1913, XLVi, 3. ^ Goodale, Tr. A. I. M. E., 1896, xxvi, 599; Goodale-Klipinger, 1913, XLVi. Bard-Gidel, op. cit., 1913, XLVi, 123. 60 COPPER ORES 6 1 chalcocite was the leading copper mineral; since then enargite became more prominent. The ore averaged in iqii^ Cu 3.2, Si02 55, Fe 10 per cent.; 0.0071 oz. Au, 2.20 oz. Ag per ton; there are present 2.3 oz. Te per lb. Cu. The ore is graded as first class, about 26 per cent, of the product (with SiOz 51.2, Fe 13.6, S 17.3, AI2O3 8.1, CaO 0.30 per cent. Ag 2.0 and Au 0.015 oz. per ton)^ which goes straight to the blast-furnace; and second class, the remaining 74 per cent, (with SiOz 58.5, Fe 9.4, S 11.6, AI2O3 11. 7, CaO o.io per cent., Ag 1.26 and Au 0.008 oz. per ton) which goes to con- centrating mills furnishing a coarse concentrate (with SiOa 20.7, Fe 25.9, S 33.8, AI2O3 4.4, CaO 0.3 per cent. ; Ag 3.9 and Au 0.019 oz. per ton) which goes to the blast-furnaces, and a fine concentrate (withSiOz 18. i, Fe 28.5, S 35.9, AI2O3 5.3, CaO 0.3 per cent., Ag 3.0 and Au 0.021 oz. per ton) which is roasted and then smelted in reverberatory furnaces. The ores of Bingham, Utah,' and Northern Nevada* are finely divided chal- cocite and chalcopyrite disseminated through porphyry. Utah ores contain about Cu 2.5, Fe 40, SiOz 25, CaO 4 per cent., 0.015 oz. Au and 0.15 oz. Ag per ton; Nevada ores 1.2-2 per cent. Cu, 0.01-0.02 oz. Au and 0.3-1 oz. Ag per ton. The ore of the Nevada Consolidated Copper Co. in 1911^ assayed Cu 1.80 per cent., Au 0.013, ^.nd Ag 0.079 02. per ton. The ores are concen- trated to a product assaying about 30 per cent. Cu, 25 Fe, 32 Si02. Similar ores occur in Arizona. The leading deposits of California are those of Shasta County,' where the ore consisting of pyrite with chalcopyrite and benide, averaging 3.77 per cent. Cu and $1.99 Ag Au, occurs in a granite porphyry. Partial pyritic smelting is practised in the district. 44. Oxide Ores. — The oxide minerals are: Cuprite (red oxide of copper), CU2O, 88.8 per cent. Cu; tenorite (melaconite, black oxide of copper), CuO, 79.8 per cent. Cu; malachite, CuC03.Cu(OH)2, 57.3 per cent. Cu; azurite, 2CuC03.Cu(OH)2, 55.1 per cent. Cu; chrysocolla, CuSi03-l-2H20, 37.9 per cent. Cu, 34.3 Si02; atacamite, CuCl2.3Cu(OH)2, 59.4 per cent. Cu, 16.6 CI; brochantite, CuS04.3Cu(OH)2, SO3 i7-7 per cent, CuO 70.3 ( = Cu 56.1), H2O 12.0 (Chile). These ores used to occur abundantly in the Southwest,' especially in Southeastern Arizona, in limestone and disseminated through erup- tive rock. A large part of the ore has changed into sulphide and grown less 'Mm. Res., U. S. Geol. Survey, 1911, i, 295. = Goodale, Tr. A. I. M. E., 1913, xlvi. » Boutwell, Pro/. Paper, U. S. Geol. Surv., 38, 1905. «Lawson, Bull. 4, Dep't. Geol. University Cal., p. 284. ''Min. Res.', U. S. Geol. Survey, 911, i, 298. •Diller, Bull. 213, U. S. Geol. Survey, 1903, pp. 123-132, 1904; pp. 169-179. GtaXon, Btdl. 430, 1910, p. 71. ' Wendt, Tr. A.I.M. E., 1886-87, xv, 25. Ransome, Globe district, Prof. Paper, No. 12; U. S. Geol. Surv., 1903. Ransome, Bisbee district, Prof. Paper, No. 21, U. S. Geol. Surv., 1904. Lindgren, Morenci district, Prof. Paper, No. 43, U. S. Geol. Surv., igoj. Lindgren-Graton, Jerome district. Bull. 285, U. S. Geol. Surv. 1906, p. 81. 62 METALLURGY OF COPPER rich so that in 1908 it averaged 4.36 per cent. Cu. The Bisbee ore averaged in 1911: Cu 5.9 per cent., Au 0.0308 oz. and Ag 1.49 oz. per ton;^ the Morenci district 3.1 1 6 per cent. Cu. Sulphide ore of higher grade, with about 16 per cent. Cu, is smelted raw ; that of lower grade, 2 + per cent. Cu, is first concentrated to a product with about 16 per cent. Cu and 30 per cent. SiOa; remaining oxide ore is usually smelted with sulphide. 45. Native Copper occurs in the Upper Peninsula of Michigan'' disseminated through amygdaloid and conglomerate beds of eruptive rocks; the rock assays 0.5-1.5 and averages i.oi per cent. Cu, and is concentrated to 65-85 per cent. Cu and then smelted in reverberatory furnaces. Native copper is very pure' containing 99.92 per cent. Cu with small amounts of Ag and Fe, perhaps some traces of Ni and As. 46. Marketing. — In marketing copper ores* there exist no general standards as is, e.g., the case with Pb and Fe ores, because mine and smeltery usually belong to the same company. Pyritic ores of the Atlantic coast, used for the manu- facture of H2SO4, are rated for the S they contain (not <37 per cent.) in addi- tion to their Cu-contents, and according to size. As regards the latter there are three classes: Lump ore, 8 in. and over; broken, 3 in.-f in.; and smalls, under j in. 47. Metallurgical Treatment in General.* — Copper may be extracted from its ore by pyro-, hydro-, and electro-metallurgical processes. The method chosen will depend upon the character of the copper mineral (sulphide, oxide, native) and the gangue, the copper content of the ore, and the cost of labor, fuel, and material. Smelting is practised with rich and medium-grade ore, because the fuel, the leading expense, increases with the amount of gangue pres- ent which has to be converted into slag. Leaching is in place with low grade ore, the gangue of which is not attacked by the solvent, as the amounts of fuel, solvent, and precipitant required are small and the percentage of extraction high; and with intermediary products, such as impure matte or copper coirtaining precious metal. Electrolytic processes have so far been a failure with ore, and a partial success with matte, and have become the standard method for treating metallic copper containing precious metal. The copper obtained by smelting and leaching is impure and has to be fire-refined; that from electrolytic processes is usually too brittle to be used as such, and has to imdergo a similar fire- refining process. ' Min. Res., U. S. Geol. Survey, 191 i, i, 279. ' Irving, Monograph V., U. S. Geol. SnTv.;Geol. Survey of Mich., vols, v and vi. T. A. Rickard, "The Copper Mines of Lake Superior," McGraw-Hill Book Co., New York, 1905. ' Douglas, Min. Ind., 1894, in, 243. * Barbour, Eng. Min. J., 1911, xcii, 314. ' General Review by Kerl, Berg. Hiittenm. Z., 1892, li, 375; Review of Recent Progress by Croasdale, Pac. Coast Miner, 1903, vn, 471. CHAPTER VII SMELTING OF COPPER 48. Smelting of Copper Ore in General.— The minerals forming copper ores were classed in §42 vmder the heads of sulphide, oxide, and native. The smelting of the ore, which is governed by the character of the copper-bearing mineral, differs accordingly; hence the whole subject is best treated imder the three heads: Smelting Sulphide-, Smelting Oxide-, and Smelting Native-Copper ores. (A) Smelting Sulphide Copper Ore 49. Smelting Sulphide Copper Ore in General. — Metallurgically considered sulphide copper ores consist mainly of CuS, FeS, and gangue, and the aim of smelting is to separate Cu from Fe, S, and gangue. The smelting is based upon the strong affinity of Cu for S^ and its weak affinity for O in comparison with Fe and the other base metals of the ore, as well as upon the fact that CujS and CU2O will react upon one another, giving Cu and SO2. If an ore rich in S is partially roasted and then subjected to a reducing fusion, be the reducing agent C, CO, or S, the gangue will form a slag. Of the metals, Cu will first unite with the S necessary to form the stable CU2S, then the Fe, not taken up by the Si02, will combine with S to form FeS, and subsequently the heavy metals will combine as long as there is S present, in the order of their affinities. The sulphides form a heavy matte which readily separates from the lighter slag, a mixture of silicates of FeO, MnO, CaO, MgO, BaO, AI2O3 .... There exists a variety of roasting apparatus (§ SS)- Smelting is carried on in the blast furnace and the reverberatory furnace; electrothermic methods have not passed the laboratory stage.^ In blast-furnace smelting the two operations of roasting and smelting may be carried on together by the so-called pyritic process, a fusion in a strongly oxidiz- ing atmosphere. There are two ways of bringing forward to metaUic copper the matte produced in ore-smelting. ' Hofmati, "General Metallurgy," 1913, p. 74, "Sulphides." 2 Vattier, Berg. HuUenm. Z., 1903, LXii, 549. Wolkow, Metallurgie, 1910, vu, 99. Ladd, Met. Chem. Eng., 1910, viii, 7. Schilowski, Metallurgie, 1910, vn, 99, 151, 43S; iQ"- vni, 617; Eng. Min. J., 1912, xciv, S°4- Stephan, Metall-Erz., 1912-13, x, 11. Lyon-Keeney, Tr. A.I.M. E., 1913, xlvii; a general review of subject. 63 64 METALLURGY OF COPPER (i) By a series of oxidizing roasts, each followed by a reducing fusion, the aim being to expel the electro-negative components of the matte (S, As, Sb), and to slag the electropositive in the order of their affinities for O, until copper is finally obtained in the metallic state. The roasts are carried on in suitable apparatus, and the fusions sometimes in blast furnaces, but more commonly in reverberatory furnaces. . (2) By converting, i.e., forcing compressed air through molten matte held, at the right temperature, in a suitable vessel whereby oxidation by O and reduc- tion by S go on simultaneously until CU2S and an irony slag is obtained, when, after pouring off the slag, the remaining S is expelled and metallic copper ob- tained. Low-grade matte is frequently brought forward to converting grade by partial pyritic smelting in the blast furnace. Most of the smelteries of the United States are situated far enough away from populous centers or agri- cultural regions to make it possible to allow the sulphurous converter gases to pass off into the open, but in Europe this is not the case; hence converting is the predominant process in the United States for extracting Cu from matte, in Europe it is the exception. The outline shows that several operations are necessary for the extraction of Cu from sulphide ore by smelting. The recovery cannot be accomplished by a dead-roast followed by a single fusion, as it would be difficult to collect all the reduced Cu, especially in case of low-grade ores, and as the recovered Cu would be very impure, containing excessive amounts of Fe, beside much As and Sb. 50. Smelting Sulphide Copper Ore in the Blast Furnace in General. — Three processes have to be distinguished, all of which aim to produce matte and slag. (i) The roasting and reduction process, also called German or Swedish process. The raw ore is subjected to an oxidizing roast, and the partially roasted ore smelted is in the blast furnace with much carbonized fuel or anthracite to furnish the necessary heat and reducing agent. (2) The pyritic process, also called American process. The raw ore is srhelted in the blast furnace without carbonaceous fuel in an oxidizing atmos- phere, the oxidation of Fe and S, and the slag formation, furnishing the neces- sary heat. (3) The partial pyritic process, a modification of (2), in which a lack of heat is made up by charging a small amount of fuel with the ore. The roasting and reduction process, the oldest of the three, is being replaced by the pyritic processes whenever the character of the ore makes this possible. The requirements that an ore has to fulfil to permit pyritic smelting are so strict that smelting without any fuel whatever is the exception, but partial pyritic smelting, in the United States at least, is becoming more and more common. The pyritic processes will be discussed together after the roasting and reduction process. 51. The Roasting and Reduction Process.— The operations to be' considered are roasting of raw ore, smelting of roasted ore in the blast-furnace for matte and slag, roasting of matte, smelting of roasted matte in the blast-furnace for impure, so-called black, copper, and slag. SMELTING OF COPPER 65 I. Roasting 52. Roasting Sulphide Copper Ore.— The- object of roasting is to oxidize S and Fe, and to remove volatile impurities such as As, Sb, and Bi. The degree to which a roast is to be carried depends upon the percentages of S, Cu, and Fe, and the amount of impurity present. An ore rich in Cu will require less roasting than one that is poor for the production of matte with a given copper content. As regards impurities, the larger the amount of S present, and the slower and more prolonged the roast, the greater will be their elimination. Thus Gibb^ found that in a cuprous pyrite with Cu s-SS, As 1.18, Sb 0.035, ^^ °-o" P^r cent., roasted in a heap, there was eliminated As 75.1, Sb 25.4, Bi 27.8 per cent.; Wendt^ roasting a similar ore with Cu 5.15, As 1.30, Sb 1.45, S 32.35, Fe 29.36 per cent., in a kiln,^ the roasted ore having lost about 20 per cent, in weight, gave the following- losses. As 97, Sb 86, S 72 per cent. In roasting a concentrate of chalcopyrite and pyrite with some bornite and chalcocite in a reverberatory furnace, Gibb* found the expulsions to be As 61.2, Sb 18.8, Bi 11. 3 per cent. In special cases, the roast may be purposely carried so far that in the subsequent reducing fusion there will not be enough S present to cover all the Cu, with the result that there will be formed some metallic copper which will carry down with it impurities, such as As, Sb, Bi, etc., to be refined by special processes, and a matte correspondingly cleaned. The general discussion of the behavior of metallic sulphides in an oxidizing roast, both in powder and lump form, is given elsewhere.* 53. Behavior of Cu-, Fe-, and Mn-sulphides in Powder Form, (i) CuaS, Chalcocite. — The changes chalcocite imdergoes in an oxidizing roast are generally stated to be as follows: Cu2S+30 = Cu20+S02, Cu20+S02+0 = CuO+SOa, CuO-f-S03?^CuS04. Aubell's^ laboratory experiments with pre- pared finely divided CU2S show that roasting starts at 200° C with the reaction 2Cu2S-|-s02=2CuO+2CuS04, and continues up to 330°; above this tempera- ture the reaction Cu2S-|-02=2CuO-|-S02 begins; up to 550° more than half of the sulphide-S is converted into svilphate-S. The SO3, formed by the dissocia- tion of CuSO 4, acts oxidizingly, Cu2S+3S03 = Cu20+4S02 and Cu20-1-S03 = 2 CUO-I-SO2. As long as the roasting ore contains CU2S, there will be formed CU2O, so that after all the S has been expelled, the roasted ore may retain as much as 30 per cent. CU2O, which has to be converted into CuO by air at a tem- perature below 1069° C. Chalcocite^ ignites in air in the range of 430 and 697'' C. according to the size of grain, o.i->o.2 mm.; it does not decrepitate. That CU2S shows a ^Tr.A. I. M. E., 1903, XXXIII, 654. ' 0/1. (c//., 1890-91, xix, 100. ' Gmehling, Oest. Zt. Berg. Hiitknw., i890,xxxviii, 272, details and drawings. * Loc. cil. > Hofman, "General Metallurgy,'' 1913, p. 403. 'Oest. Jahrb., 1910, LViii, 131. ' Friedrich, Melallurgie, 1909, vi, 1691. S 66 METALLURGY OF COPPER tendency to sinter while roasting, as stated by Plattner/ cannot be due to the fusion of CU2S, as its melting point lies at 1127-1130° C; the sintering he did observe must have been due to some other cause, perhaps the formation of an oxysulphide. In a laboratory sulphatizing roast carried on between 420 and 440° C, Warlimont^ succeeded in rendering 52.5 per cent, of the Cu water- soluble. (2) CuS, CovELLiTE. — As CuS gives up one molecule of S when brought to a bright red with exclusion of air, and is converted into CuaS, its behavior in roasting should be similar to that of CuaS. (3) FeS, Pyrrhotite, rEiiSi2. — The changes FeS undergoes may be expressed according to Plattner's outline by FeS+30 = FeO+S02, 3FeO+0=: FesOi, and S02+0+catalyzer=S03; 2Fe304+S03=3Fe203+S02 and FeO+ S03 = FeS04; 2FeS04+heat=Fe2S06+S02 and Fe2SO6+heat = Fe2O3+S03. According to Kothny,' heating FeS04 in a current of CO2 at 280° C. causes it to be converted into Fe203.S03+S02 up to 530° C, when the basic ferric sul- phate is dissociated into FeaOs and SO3. Heating in a current of air gives rise to the reaction 4FeS04+02=2(Fe203.2S03) within a temperature range of 150 and 380° C. Between 380 and 530° some SO2 is set free; above 530° decomposition again sets in. The oxidation of FeS may progress more directly than shown above, as seen by the equations 4FeS+ 762= 2Fe203-|-4S02 and FeS -l-202 = FeS04. In roasting there has to be considered also the oxidation of FeS by 3SO3 to FeO and 4SO2. Pure FeS ignites in air at from 325° (o.i mm.- grain) to 472° C (>o.2-mm. grain), pyrrhotite at from 430 to 590°. The former does not decrepitate; the latter does somewhat.* Kothny^ found that finely divided prepared FeS begins to oxidize to FeS04 at 170° C, continues to do this up to 430°, when mainly Fe203 and SO2 are formed accompanied by some FeS04. In fact the presence of FeS04 is noticeable up to 600° C. In a laboratory sulphatizing roast, Warlimont^ rendered in 5 hr. 31.8 per cent, of the Fe water-soluble. (4) CU2S. FE2S3, Chalcopyrite. — The general behavior is similar to that of pyrrhotite except that beside Fe^Oj, and FeS04 there is formed CuiO„ and CUSO4; the formation of water-soluble CuS04is greatly assisted by the presence of Fe^S. Thus Warlimont' succeeded in rendering 97.7 per cent. Cu water- soluble with a mixture of 1CU2S: loFeS. Similar extractions were obtained by Wedge. ^ Chalcopyrite decrepitates upon heating. (s) FES2, Pyrite. — If heated to 700° C. with exclusion of air the reaction FeS2-|-heat = FeS-|-S takes place, the product resembling the magnetic ' "Rostprocesse," p. 79. - Metallurgie, 1909, vi, 132. ^ Oest. Jahrb., 1910, Lvm, 112; Metallurgie, 1911, vin, 389. ■• Friedrich, he. cit. ' Loe. cit. " Loc. cit. ' Loc. cit. s Eighth Internal. Congress of Appl. Chem., 1912, m, 151; Tr.A.I. M. E., i9i2,XLiv, 818, in large-scale work with his roasting kiln (§§ 6g, 214). SMELTLXG OF COPPER 67 sulphide;' however, the expulsion of S begins at 200°. ^ According to Friedrich,' pyrite from Elba heated in air gives ofi SO2 at 405° C. and glows at 533° C; the more free-burning mineral from Rio Tinto begins to roast at from 260 to 275° C.'' Kothny's^ experiments show that 250° C. is the lowest roasting temperature. Pyrite from some localities decrepitates readily; from othefs it does not; the latter brings a higher price. FeS2 is more readily roasted than FeS, as the expul- sion of S by distillation makes the mineral porous, and as the larger amount of SO3 formed exerts its strongly oxidizing influence. Assuming that no S is distilled, the complete oxidation of FeS2, as formulated by Waring,'' is expressed by 4FeS2+ii02=2Fe203+8S02; the formation of Fe304 with a short supply of air, by FeS2-f02 = FeS+S02andFeS-|-ioFe203 = 7Fe304-)-S02. Theexperi- ments of Kothny^ have shown that between 250 and 290° C. the oxidation of FeS2 takes place according to FeS2-f302 = FeS04-|-S02, and between 290 and 500° C. according to 4FeS2-|-ii02 = 2Fe203+8S02 when the FeS2 kindles. Oxidation to FeS04 is, however, noticeable up to 600° C. He also showed that the reactions FeS2+sFe203 = iiFeO-f2S02 and FeS2+i6Fe203=iiFe304 -t-2S02 do occur, beginning at 380° C. and requiring that for i molecule FeS2 there be present 16 mol. Fe203. (6) MnS, Alabandite. — This compoimd, which melts at 1162° C.,* is con- verted by roasting into MnS04 and Mn304. The MnS04 begins to give off SO3 at 699° C; the dissociation^ into Mn304, SO3, SO2, and O is energetic at 790°. Friedrich'" found that alabandite with 2.02 and 1.98 per cent. Fe, when o.i mm. fine, ignited at 355° C; when coarser than 0.2 mm., at 700° C; the mineral did not decrepitate. 54. Behavior of Fe-, Cu-, and Mn-sulphides in Lump Fonn. Kernel Roasting. — In roasting a sulphide copper ore or matte in lump form, the oxida- tion will start at the surface of a lump, when this has been brought to the kind- ling temperature, and then penetrate toward the center at a speed governed by the heat evolved, which causes the lump to swell, become porous, and crack, and thus furnish channels for the travel of the air. In time the surface will become covered with a rind or shell of more or less porous Fe304 and Fe203, through which finally the air cannot penetrate sufficiently to complete the oxidation. This is effected by SO3, viz., S03+Met. S = S02-f-Met.O; but part of the SO2 1 Valentine, Tr. A. I. M. E., 1889-90, xviii, 78. Geodel, /. fur Gasbeleuchtung, 1905, XLVm, 400. Friedrich, Stahl u. Eisen, 1911, xxxi, 2040. Barth, Metallurgie, 191 2, ix, 204 (600° C). ' Kothny, loc. cit. Barth, loc. cit., 350° C. ' Loc. cit. * Chalon, Rev. Un. Min., 1902, LVii, 201. ' Loc. cit. ' Min. Mag., 1905, xii, 196. ' Loc. cit. * Fay, Proc. Am. Soc. Test. Mat., 1908, vm, 92. ' Hofman-Wanjukow, Tr. A. I. M. E., 191 2, XLiii, 548. " Metallurgie, 1910, vii, 329. 68 METALLURGY OF COPPER is decomposed, 3802 = 5+2803, setting free S-vapor, which, traveling toward the cooler center of a lump, is likely to cause the formation of a kernel of sulphide. If the ore is low in Cu, the S in a roasted lump may have been almost com- pletely expelled, and the Cu converted into CuO, only a small part retaining the transitional forms of CujO and CU8O4. If the ore is rich in Cu, the roasted lump may contain some metallic Cu formed by 2Cu20+Cu2S = 6Cu+S02 and by 3Cu20+FeS = 6Cu+FeO+802. If the ore contains some Ag2S, the silver in the roasted lump will be present as Ag2S04 or Ag, the latter having been formed by Ag28+20 = 2Ag+S02 or by Ag2804+Ag28 = 4Ag+2S02,^ and not by decom- position of Ag2S04, which takes place at 925° C.^ A special form of roasting copper ore in lump form is Kernel Roasting,^ whereby the small quantity of Cu in a low-grade pyritic ore is concentrated in the center of the lump as a kernel of sulphide, which thus becomes rich, while the surrounding shell or rind, converted into a mixture of Fe304 and Fe203, is correspondingly impoverished, retaining small quantities of copper as CUSO4 and CujO. Edward^ has noted an enrichment of the rind in Ag; a similar phenomenon was observed by Plattner^ with matte whether lead-bearingor not. Figs. 43-45 show three characteristic stages in heap- or stall-roasted cop|)er- bearing pyritic ore of suitable size and character. Liirzer^ shows a very high degree of concentration in the following analyses: Crude ore, Cu 1.60, Fe 43.50, S 50.25, Si02 5.00 per cent. Pure kernel, Cu 41.64, Fe 28.76, S 29.28, Si02 0.08 per cent. Rind next to kernel, Cu 3.31, 8 0.92, Si02 2.85, CuO 1.58, Fe o.io, Fe203 85.70, SO3 2.50, ignition loss 3.04 per cent. Concentration' of ore with 3-4 per cent. Cu into kernels with 15-20 per cent. Cu and rinds with 2 per cent. Cu was common at Fahlun. At Agordo,' ore with 2 per cent. Cu gave 13 per cent, kernels with 6 per cent. Cu, and 87 per cent, rinds with 1.28 per cent. Cu (calculated). SchnabeP gives from his work in the Caucasus a concentration of Cu, from ore with 7-10 per cent. Cu, into kernels with 35-40 per cent. Cu, the rinds assaying 3 to 4 per cent. Cu, of which 2| to 3 parts were CU8O4 and J to i part CuO. This concentration of Cu toward the center has been attributed by Plattner'" to CU2S becoming liquefied and being drawn toward the center. Now CU2S melts at 1127-1130° C, a temperature not reached in a heap- or stall-roast. 8chertelii thinks it a process of adhesion and not one of fusion. Considering that the kernel at the end of the roast is a ' Sackur, Ber. deuisch. chem. Ges., 1908, XLi, 3356. ' Hofman-Wanjukow, loc. cit. ' Peters, Min. Res. U. S. Geol. Surv., 1882, p. 287. ^ Eng. Min. J., 1895, lk, 411. ^ " Rostprocesse," pp. 183, 205. ° Turner's Jahrb., 1853, in, 339. ' Bredberg, Erdmann's J. Technisch-Oekonomische Chem., 1829, iv, 300. ' Egleston, Sch. Min. Quart., 1887-88, ix, 124. ' Schnabel-Louis, Handbook of Metallurgy, 1905, i, 39. '" "Rostprocesse," p. 195. " Dingl Polyl. J., 1872, ccvi, 284. SMELTING OF COPPER 69 a. Crust of ferric oxide. h. Layer of enriched sulphide (chalcopyrite). (.. Unaltered center. FiG. 43. — Roasted for a short time. -■■r^^ o. Crust of ferric oxide. 6. Layer of enriched sulphide (bornite). c. Layer of enriched sulphide (bornite). d. Layer of enriched sulphide (chalcopyrite) . e. Unaltered center. ■-^■v-^^-h V', \ ' — ■!■ ndiiliiiiifL in ^11 Fig. 44. — Roasted for a longer time. a. Crust of ferric oxide usu- ally much cracked. b. Kernel of enriched sul- phide. ^^'i&\g1^^ '^ N. Fig. 45. — Final stage of roast. Figs. 43-45. — Three stages in kernel-roast of pyritic copper ore (Plattner). 70 METALLURGY OF COPPER mixture of CuaS and FeS, it may be that instead of CujS, the eutectic (see matt* §119) of the two components, which melts below 950° C, travels inwarc Another theory is that of Poole, ^ who states that S, which thickens when heate for some time to above its melting point, forms a very thin liquid when finel divided, CuaS or FeS is stirred into it, and suggests that this thin liquid travel toward the center. The latest theory is one by Friedrich,^ who bases it on th experimental fact observed by Schenck-Hempelmann* that heating a mixtiu: of CUSO4 and CuaS to a temperature lying between 300 and 400° C. causes th formation of a viscous brownish fluid. He believes that the essential point the fusion of CuaS at a very low temperature, is satisfactorily explained. Presupposing the existence of a plastic or fluid substance, there remain to be [given an explanation of the cause for the inward travel from th bottom upward and from the side inward. Howe^ suggests capillary at traction which causes the plastic part to adhere to the solid; McRoss^ an( Austin' suggest magnetic attraction on account of the plastic mass becominj not only magnetizable, but magnetic; Knapp'' and Roberts-Austen' sugges dififusion which might be active without any fusion whatever. There is m reason why the three forces should not supplement one another. The sugges tions of H.G.Z.' of an outward movement of FeS, and of Edwards^" of an in ward movement of CuO do not appear reasonable. A successful kernel-roast''^ requires a copper-bearing pyrite of say 4 in. ii diameter that does not decrepitate upon heating and is nearly free from gangue and a slow roast at a regulated low temperature. It is carried on in heaps anc stalls. The oxidation of a lump begins at the surface, the heat generated sub limes some of the S which protects CU2S from oxidation; the CuzS alone with some FeS travels inward while the excess FeS is oxidized by the action o air or SO3. In ordinary heap- or stall-roasts, the limiting size for lump-ore is i 3-in. ring, as with larger pieces there is a tendency to the formation of kernels a condition which is usually avoided as much as possible. 55. Roasting Apparatus in General. — The roasting of copper ores is carrie( on in heaps, stalls, shaft-, reverberatory-, and mufifle-furnaces, and in blast roasting apparatus. The roaster gases from heaps, stalls, and reverberator furnaces contain under i per cent. SO2 and are contaminated with fuel gases hence they cannot be economically utilized. ^^ Those from shaft- and'muffli 1 Tr. A. I. M. E., 1906, XXXVI, 403. ^ Metall-Erz, 1914, xi, g. 'Op.cit., 1913, X, 293. ' Eng. Min. J., 189s, Lrx, 104, 267, 364. " Op. cit., pp. 195, 339. * Min. and Meth., 1911, n, 119. ' Eng. Min. J., 1895, Lix, 339. * "Introduction to the Study of Metallurgy," 1910, 55. " Eng. Min. J., 1895, nx, 147. "0/>. cit., p. 411. ■' Howe, op. cit., 189s, LIX, 104-267. AyT„*.„n —. )J „ 00 SMELTING OF COPPER 71 furnaces contain over 4 per cent. SO2' and are free from fuel-gases; tfley can be converted into H2SO3, H2SO4 or SO3. The SOa-content from intermittent blast- roasting apparatus varies considerably; usually the gases go to waste; in some instances several pots are run in series in such a manner as to furnish a roaster- gas of uniform grade with over 4 per cent. SO2. From the continuous blast roaster of von Schlippenbach^ the sulphurous gas is utilized in the manufacture of H2SO4. The choice of furnace is governed by the chemical composition (mainly percentage of S and Fe), physical character (coarse or fine), and the value of the ore (percentage of Cu, Ag, Au), or the time that can be given to the roasting. It is further influenced by the degree of desulphurization that is demanded, by the practicability of recovering the sulphurous gases or the necessity of rendering them harmless, and lastly by the price of labor, fuel, and material. 56. Roasting in Heaps.' — Since the advent of partial pyritic smelting, this method of roasting has lost much of its former importance. It is suited for coarse ore with an admixture of only a small amount of fines; and consists, Figs. 46-48, in piling the ore to the form of a truncated pyramid on to a bed of wood on suitable ground, and igniting the fuel, which heats the superincumbent ore and starts the roasting. If the ore contains sufficient S to keep up com- bustion, the process of roasting will proceed of its own accord; if not, the lack has to be made up by mixing in fuel (coke or coal fines, refuse wood, etc.). Ores containing less than 15 per cent. S require intermixing of fuel. (i) The Roast-yard. — The location of the roast-yard has to be so chosen that the heaps are protected from strong winds, and that the prevailing wind carries the gases away from the works. Ore-heaps are rarely covered by sheds as is often the case with the more valuable matte-heaps. In order to have a cheap roast it is essential that there be as Uttle handling as possible of raw and roasted ore; special provision has to be made for this. One of the simplest arrangements of roast-yard is that formerly in operation at the Vershire copper mines, near Corinth, Vt., shown in Figs. 49-51. The ore is brought in side-idump cars on a trestle at an elevation of from 10 to 12 ft. across the places where the heaps are to be erected; the trestle carries T-rails and has a slight grade, from f to i per cent. Parallel with the upper track and about 4 ft. below the level of the yard is a second track over which the roasted ore is run to the feed floor of the blast-furnace, the tops of the cars being on a level with the floor of the yard. Similar permanent plants are those 1 Hofman, op. cit., p. 881. ^Hofman, Min. Ind., 1910, XLi, 761. Kroupa, Oest. Zi. Berg. Hiiltenw., 191 2, xl, 539. » Peters, Min. Res., U. S. Geol. Surv., 1882, 283; 1883-84, 283; "Modern Copper Smelting," New York, 1895, p. 104. Glenn, Eng. Min. J., 1883, xxxvi, 392. Wendt, Sch. Min. Quart., 1885-86, vii, 154, 281, 301. 72 METALLURGY OF COPPER Section on CDEFGH Section on AB \Y Figs. 46-48. — Roast-heap. Section on XY SMELTING of' COPPER 73 k given by Peters/ and the more complicated arrangement of the Tyee Copper Co., Ladysmith, B. C, with Kiddie movable bridges.^ At Keswick, Cal.,' the ore was delivered in 10- to 20-ton cars to bunkers of 150 tons capacity closed with grizzUes (3-in. slots); beneath was another set of screens with i- to f -in. holes, thus furnishing coarse, medium, and fine ore, which dropped into separate cars of 2500 lb. capacity. The cars, with i8-in. gauge, were run on i6-lb. rails, which were spiked to 4X4-in. ties laid across 6X8-in. stringers, 16 ft. long, connecting the bents 10 to 12 ft. above the roast-yard. Tramway to Smelters Roast Yaxi (26 PUes) Vershire Copper Mines ,X^. ,XtV. .XIV. ,xtx. .^^ Figs. 49-51- The bents of a trestle were made of round poles, 4-6 in. in diameter at the small ends. When a heap was erected, the rails, stringers, and ties were removed to be used in another heap, while the poles were left in place, the cost of extract- ing them being greater than their value. At the works of the Canadian Copper Co., Copper Cliff, Ont., the roast-yard has a capacity of 100,000 tons. In building a heap, the ore is wheeled from a flat car and spread, the base of the heap being 1-2 ft. distant from the track. The roasted ore is loaded from either side of the heap by means of a steam- shovel, having a bucket of 2.5 tons capacity, on to so-ton ore cars; it takes about S min. to load a car. When the ore is badly sintered, 40-per cent, dynamite is used to loosen the material. The roast-heaps of the Tennessee copper were arranged similarly to those at Copper Cliff, only they were smaller. ' "Copper Smelting," p. iii. ' Rep. to Minister of Mines, Brit. Columbia, 1902, p. 243. Brewer, Min. Sc. Press, 1903, ixxxvii, 7; Eng. Mag., 1904-05, xxviii, 348. Jacobs, Eng. Min. J., 1904, lxxviii, 748. ' Neilson, op. cit., 1899, lxviii, 457. Scale feet 74 METALLURGY OF COPPER At the works of the Mond Nickel Co., Victoria Mines, Ont., the roast-yard is situated half way between the mine and the smeltery, which are two miles apart. The ore is crushed at the mine and passed over a f-in. screen; it is then trans- ported by a Bleichert elevated tramway to the roast-yard, and the roasted ore by the same means to the blast-furnaces. The train handles in lo hr. 400 tons of ore in buckets holding 800-900 lb.; there are 180 cars on the two lines, which travel at a speed of 3.5 miles per hour. The tension station is at the roast-yard. Here the buckets are unloaded into small ore-cars and the contents distributed on to the heaps over temporary trestles. The roasted ore is shoveled into 2-ton cars; these are hauled by horses to the base of incline, raised by a gasoline hoist, and discharged into bins, the contents of which are drawn into the Bleichert buckets and carried to the feed floor of the blast-furnaces. At the new plant of the company at Coniston, Ont., the ore is discharged from bottom side-dump cars into ditches running longitudinally between the roast-heaps, which are placed end to end. Grab-bucket cranes on standard-gauge tracks pick up the ore from the ditches and drop it on the beds. The same cranes excavate and load the roasted ore into so-ton steel dump cars, which are hauled to the smelter bins. The grab-buckets are designed for digging up ore that is more or less sintered. (2) The area of roast-yard necessary for furnishing daily a given amount of ore is very large. A heap 40X24 ft. and 6 ft. high, holds 240 tons ore and burns 70 days; adding 10 days for building and removing makes 80 days. Such a heap then furnishes per day 3 tons of roasted ore; 35 heaps give 105 or in round figures 100 tons ore per day. Allowing 10 ft. at the ends and 6 ft. at the sides of a heap for working, gives an area of 60X36 = 2160 sq. ft. for a heap, or 75, 600 sq. ft. for 35 heaps or 100 tons roasted ore per day. (3) The ground on which heaps are to be built ought to be dry and hard, similar to a macadamized road. A ditch dug at the upper end prevents water from entering the yard which slopes either toward the lower end or the two sides. Drainage may be assisted by underground drain pipes. At Ducktown, Tenn.,' 3 T .4 per cent, of the C u in a heap was lost by defective drainage, being leached by frequent heavy rains. If the ground is soft, it becomes mixed with the roasted ore; 2 it may be hardened by removing the surface with a scraper, filling the excavated space with rock or coarse slag, and the interstices between the latter with gravel, concentrator-tailing, or granulated slag, and covering the new sur- face with loam and rolling it down. The finished yard should be two or more inches higher than the surroundings. (4) Crushing and Sizing of Ore.— The most suitable size of ore for roasting is from 1.75 to 2 in. if it contains under 25 per cent. S, and 3 in. if it contains more sulphur; ore larger than 3 in. is likely to form kernels (§54). These general figures will undergo slight changes with the character of an ore as dictated by practical experience. • Wendt, Sch. Min. Quart., 1885-86, vii, 173. 2 Op. cit., p. 180. SMELTING OF COPPER 75 The ore is crushed by machinery (rock breakers) or by hand (spalling). The former works cheaply, but makes many fines; the latter permits sorting out barren rock and may thus compete with the forioer in small plants. However, crushed ore can be conveyed by picking belts and barren rock removed. The crushed ore is sized into three classes: coarse, i to 3 in. ; medium (ragging) i in. to 3-mesh; fine, under 3-mesh. The sizing is done by grizzlies, trommels, or shak- ing screens; if hand labor is employed, by forking out the coarse, and separating the imdersize by shoveling on to an inclined screen. The relative amounts of the sizes obtained in crushing vary greatly; thus, Peters"^ gives coarse 55, ragging 25, fines 25 per cent., and Glenn^ coarse 82, ragging 7, fines 11 per cent, as aver- age figures for the product. (s) The Heap (Figs. 49-51). — The most important dimension of a heap is its height which varies with the percentage of S. An ore with 15 per cent. S can stand a height of 8 ft. above the bed of wood; one with over 35 per cent. S only about 5 ft. ; an average height is about 6 ft. The length and width have little influence upon the result of the roast; large heaps which burn a long time are, however, more advantageous than small ones, as both furnish about the same amount of imperfectly roasted ore, which has to be retreated. Examples are given in Table 19. (6) Building or Heap. — The rectangle, 24X40 ft. in Fig. 50, is staked out, and fines, a, are spread over the ground to the depth of 4, 6, and even 8 in. They prevent roasting ore from adhering to the ground, are in part roasted and caked, so that they can be reroasted in a neighboring heap with lump ore. On to the fines is placed a bed of partly seasoned cordwood 3-5 in. thick and 4 ft. long. The best quality goes to form the border, which is 4 ft. wide and 8-14 in. (av. 9 in.) high; inferior uneven wood serves to fill the center. Work is begun by making the border b, which ends 6 in. from the staked lines. Sticks are laid down on the ends and sides of the proposed heap, placed end to end and parallel with the ends and sides of the heap, the side-borders butting against the end-borders. Flues (not shown), 6-8 in. wide, communicating with central chimneys, c, are left open in the sides every 8 or 10 ft., to be filled with kindling; the chimneys, about 8 in. sq., are boards (nailed together) of sufficient length to reach above the finished heap. On to the first layer of cord wood and at right angles to it is placed a second layer, d, of best-seasoned wood so as to reach to the staked lines and thus extend 6 in. over the bottom layer. Not coming in contact with the floor, this second layer greatly facilitates firing. The wood is piled as closely as possible; any open spaces are filled with small dry sticks. Upon the sticks, d, forming the second layer, are placed crosswise three to four sticks, e, to form a support for the ragging and fines, and thus prevent them from rolling down while the heap is being built. The space, /, between the border is now filled with sticks of varying size. They are laid on the fines end to end and parallel with the sides, but overlapping one another (Fig. 51), so as to leave air-spaces open. The wooden chimneys, c, are now put in place on top of the ' Op. cit., p. 90. « Op. cit., p. 353. 76 METALLURGY OF COPPER H < O Canadian Copper Co., Ont. M X V o,^ "0 2 O X " ^ .- [S ? 2 o Mond Nickel Co., Ont. 1^ 00 M o Tyee Copper Co., B. C. X o O u ^ o ^ ^ 3 to CJ CO o o o S (J 3 CO X o X ;? 1 ^ ^ °l CO 5 Ore Knob, N. C. 00 X o o X " M O ■^ o X ° ° <^ ro t^ ^ CD N Tt fn o o u u 8 M Ov ro _,. fO (^ ^ fr> o o X ■t X " On to vO O op ■^ Point Shirley, Mass. I 1 00 in lo d O ■SS X s X ■4- § Si o - 'o J3 1 - o *-< § ° Jo g 4 U - I-I C/3 ■4-J u 1 u I-I t3 a CJ a ^ C 1 '1 c in 1 C .c i 0. t 0. "o c 0. C u X > 0. 1- c "c c c c S a IS S S o "_ a •^ o CJ ^"5 :.. „ I-S m n O " 2 +J s ° w •c ^ ■o (U o > ■n Ph c >-< M l-t H CO l«l M Jl fo w ■ oo o £f ^ >—> ^ O ^■ t" >■ „ " !> .- iC "' S " •' M ^ to CO 00 M • 00 5 t^ H I .S o 1 jJ 00 a .- SMELTING OF COPPER -j-j kindling-flues, the boards facing two opposite flues being made about 8 in. shorter than the others in order to insure the necessary communication. Fig. 49 shows that the central and main part of the heap, g, is made up of coarse ore. This receives a cover of ragging, h, and the latter a thin layer of fines, i. The coarse ore is dumped from the overhead cars on to the finished bed of wood or at first perhaps on a wooden platform to break the shock; some of the ore is piled around the chimneys to steady them, and the rest evenly distributed over the bed. In order to do away with part of the hand labor necessary for spreading the ore delivered at the center, a movable table, similar to the one shown in Fig. 52, is placed on the main track and connection made by two heavy rails held together by clamps and supported at one end by a bent. This permits dumping the ore on different parts of the heap. The firpt 3 ft. of coarse ore is piled carefully, and the sides are kept smooth and at as steep an angle as possible (42 deg.) ; the lower edge of the truncated pyramid of coarse ore reaches to within 2 to 2.5 ft. of the upper layer of cord wood. Top and sides of coarse ore are now covered with about i f t. of ragging, which grows a Uttle thicker toward the bottom. The fines, about 10 per cent, of the weight of the ore, are arranged in small piles, about 3 ft. from the heap, so as to be ready for spreading later on the heap; or, one-third of the total is spread evenly over the heap, leaving uncovered the top and a border about 20 in. wide along the bottom. (7) Firing. — ^The firing of a heap is started early in the morning on a bright day. If there are no draft-flues, the firing is begun at the ends, otherwise the kindling in the flues is ignited. The bed of wood will be burning fully in from 4 to 5 hr. after starting, and the ore starting to roast. When the burning of the ore has progressed about i ft., a thin layer of fines is spread over the surface with a shovel and patted down. This smothering is continued as the fire creeps up, leaving visible a border of ignited ore. When the heap is well started, dense yellow fumes arise, the surface becomes damp (sweats), the heap settles, and fissures appear which are filled with fines. About the third day, the fire will have reached the top, when a workman ascends the heap and covers it with a layer of fines. The sides of the heap will show sublimed S, and As^Sj, if the py- rite was arsenical; if the S is melted the temperature is too high; if the ore is cemented together by the S, the crusts formed have to be broken. The tem- perature on the sides and top must be kept even; it is correct when the hand can just bear touching the cover. If too hot, the thickness of the fines is in- creased; if too cool, it is decreased in order to draw the fire in the direction of the cool place. An average thickness of fines is 4 in. on the top and 3 in. on the sides. After 10 days, a heap requires little more attention than a daily inspection. In some European works" treating pyritic ores rich in S, sublimed S is collected on top of the heap during the first period by making in the cover 25 or more spherical depressions, 14 in. in diameter and 7 in. deep, lining them with raw or roasted fines, and enclosing the top of the heap with boards to protect the S from the prevailing air-currents. Nevertheless much S is burnt oS; that which 78 METALLURGY OF COPPER remains, about i per cent, of the S of the ore, is ladled into wooden molds, refined, and sold. (8) Opening (Stripping, Turning) of Heap.— This begins when roasting has ceased, and the heap has cooled sufficiently to allow transferring the ore to the feed floor of the blast-furnace. First, the heap is stripped, that is, unroasted fines and ragging are removed and transferred to a neighboring heap that is building; the rest then is pulled down, well roasted ore being kept separate from that which is imperfectly roasted or fused (heap-matte) , the last two going to another heap that is being built. Heap-matte often has to be blasted. (9) The products are: (a) well-roasted ore which is porous, reddish-brown (Fe203) to brownish-black (Fe304), light, more or less friable; has an earthy fracture; and retains 4-7 per cent. S, the S-content rising and falUng with the percentage of Cu, as most of the S is sulphide-S in combination with Cu; (6), imperfectly roasted ore, mostly fines and some ragging; (c) sintered and fused ore, a gangue-skeleton near the top, from which matted sulphide has eliquated and collected on the bottom. The proportions of these three products vary, but a:6:c = 9o:7.s:2.5 are not uncommon. (10) The cost' is given as ranging from 20 to 80 cents per ton of ore. This great difference is caused mainly by the handling of raw and roasted ore as shown clearly in the costs of Keswick, Cal.,^ given in Table 20, where labor was $1.85 for 10 hr., wood $3.00 per cord, and the daily capacity from 500 to 800 tons. Table 20 . — Cost of Heap-roast Cents per Ton Building heap and roasting Indirect' loading Direct' loading Discharging heap Manual labor Steam shovel |-yard capacity Labor .... . I S • 00 6.25 2.5° Labor 20.50 4-5° 2.50 0.20 0. 16 0.04 0. 10 OS4 0.02 0.20 Stores Stores 2.50 0.12 Total 17.90 931 23.00 4.92 Grand total 40.90 14- 23 ' 1 The differences in arrangement for building, 17.90 115. 9.31 cents per ton, and for discharging by hand or machinery, 23.00 vs. 4.92 cents, are made evident. With the Tyee Copper Co. at Lady smith, B.C., with labor at $1.00 and fore- ' Church, Eng. Min. J., 1893, lvi, 666. Peters, "Modern Copper Smelting,'' 1895, p. 132. 2 Neilson, Eng. Min. J., 1899, Lxvin, 458. ' Indirect loading is the method described on page 75. Direct loading means that the ore screened at the mine was broughtin loo-ton cars on a strong trestle running over the yard and discharged direct without any intermediate process, thus saving bunkers, grizzlies, tram cars, and small trestle. When one series of heaps had been built, the bents were transferred to another. SMELTING OF COPPER 79 man at $3.00, the cost per ton was: Laying wood, 0.007 cents; tramming from receiving lines to heap, 3.590 cents; shoveling from heaps to 2-ton cars {\ of heap to be broken up with wedges), 15.490 cents; cordwood at $2.00, 0.700 cents; total 19.787 cents; tramming by horse 1500-2000 ft. from yard to blast-furnace, 5 cents per ton. Examples in addition to those given in Table 19 are: V-method of Peters and MacArthur' at Sudbury, Ont., given up as being of little advantage; roasting bituminous shale at Mansfeld, Germany;^ roasting at Rio Tinto;' kernel roasting at Agordo;^ and at Ducktown, Tenn.^ The two main advantages^ of the process are cheapness of plant and process, with lump ore as a product. The main disadvantages, slow, imperfect, intermittent roast depending upon the state of weather; loss of ore by dusting, tramping under foot and leaching; locking up of large amounts of ore in an extended roast-yard; loss of S; killing of vegetation; and exclusion of fines in excess of say 10 per cent. The last is remedied in part by briquetting (Tyee Copper Co.) or moistening with FeS04+aq. and allowing to harden (Agordo). 57. Roasting in Stalls.'' — A stall is an oblong space surrounded on three sides by permanent walls; the fourth side, the front, when the stall is to be filled, is closed, wholly or only in part, by brick set dry or by an iron plate; the front wall is removed again when the stall is to be emptied. Frequently a number of stalls are built side by side against a main wall forming a single row as, e.g., at Keswick, Cal.* Another arrangement is to have two rows of stalls, back to back with a main flue between them, Figs. 52-53- It is more compact, requires less brick and ironing (if not built of slag brick), retains the heat better, and makes it con- venient to carry ofi the gases. The top of a stall is either open as in Fig. 53, or closed by a brick arch" or an iron plate. Ore-stalls are usually open, while matte-stalls are closed. Closing the stall gives a better utilization of heat, and insures withdrawal of gases through flues in the back or pipe in the roof. Figs. 52-53 represent the stalls of the Butte Reduction Works, Butte, Mont., with some modifications. They are very similar to those formerly used at the ^Tr.A. I. M. E., 1889-90, xvra, 284. ' Jungfer, Berg. HiUlenm. Z., 1887, XLVi, 471. Egleston, School Min. Quart., 1890-91, xii, 85. Wagner-Primrose, Eng. Min. J., 1907, Lxxxiv, 673. ' Peters, Min. Ind., 1893, 11, 26. Chalon, Rev. Un. Min., 1902, Lvn, 201. < Church, Eng. Min. J., 1872, xiv, 131. Egleston, School Min. Quart., 1887-88, ix, 124, 256, with cross references. Ernst-Monaco, Berg. HUllenm. Z., i8gi, l, 26. ' Wendt, School Min. Quart., 1885-86, vii, 218. " Henrich, Tr. A.I.M.E., 1895, xxv, 224. 'Peters, Mm. Res., U. S. Geol. Surv., 1882, p. 290; 1883-84, p. 389; "Modern Copper Smelting,'' 189s, p. 40. Henrich, Tr. A. I. M. E., 1895, xxv, 229, 232. ' Keller, Min. Sc. Press, 1896, Lxxin, 497. ' Wendt, School Min. Quart., 1885-86, vn, 306. 8o METALLURGY OF COPPER Parrot Smelter, Butte.^ A stall, a, is 8 ft. deep, 6 ft. 6 in. wide, 6 ft. high, and holds about 20 tons average Butte ore. The larger the stall, the more diflScult the regulation of the air-supply and thereby of the temperature; hence small stalls are the rule. Large stalls have I Figs. 52-53.— Roast-stalls of Butte Oil Reduction Works (with additions). been failures as, e.g., the older stall of the Parrot Works,^ 40 ft. long by 9 ft. high, and the later one at Keswick, Cal.,^ 14 ft. long by 7 ft. wide by 6 ft. deep, holding 35 tons. 1 Peters, Min. Res., U. S. Geol. Surv., 1883-84, p. 389; "Modern Copper Smelting," 1895, pp. 148-150. 2 Hofman, Tr. A. I. M. E., 1904, xxxiv, 261. = Keller, Min. Sc. Press, 1896, Lxxin, 497. Neilson, Eng. Min. J., 1899, Lxvin, 457. SMELTING OF COPPER 8i The walls, h, are usually of slag brick, as these are cheap and little attacked by the roasting ore; they are made very thick, 2 ft. 6 in. to 2 ft. 9 in., in order to require no ironing. The back wall has three openings: c, 6 by 8 in., for the pas- sage of the gases into the main flue d, 24 by 48 in., leading into the stack e, 3.5 ft. square by 75 ft. high. The flue is provided with a damper (not shown). With this arrangement, characteristic of the older stalls, air enters only at the front. In order to allow admission of air at the sides and the bottom, modern stalls have flues, /, in the side-walls, and usually a single one, g, in the floor which has a grated intake. The air-supply can be reduced or shut ofi by placing bricks in the flues. As a carefully regulated admission of air is essential for good work, the roasting in modern stalls is more satisfactory than in the older type. Before a stall is charged, the walls are usually plastered with clay to prevent adhesion of roasting ore. The manner of placing the fuel-bed varies. A central air-passage from front to back with two cross-ways is built with lump ore, filled with kindling, and the whole covered with rotten wood (sound cord- wood often gives too much heat), or the cross-ways are omitted as with h in Fig. 52. Upon the bed of wood is charged coarse ore, i, then coarse and ragging, placing here and there sticks against side and back walls to prevent chiUing, and light wood toward the front, as this is being built of coarse ore backed by brick laid dry and in the form of an inverted arch, k; then comes ragging alone fol- lowed by a 3- or 4-in. layer of chips of wood, bark, shavings, etc. , which is covered with from i to 1.5 tons of fines. At the Butte Reduction Works the front of the stall is closed down to 2 ft. from the bottom with a plate of sheet iron, m, braced by cross-bars to prevent bulging and supported by wooden blocks, n, small sticks and kindling being placed between the latter. When the swelling of the ore during the roast has caused the charge in the stall to rise to its maximum and it begins again to shrink, the front plate gradually gUdes down to the floor. A stall is usually fired at night, as the draft is better than during the day and carries away most of the offensive smoke that arises at first in considerable amounts. After the first day, the lower part of the front brick wall is warm; after the second the middle part, and after the fourth day the upper part. The roast is finished after nine days, when the front wall is taken down, the stall entered and its contents removed. Butte sulphide copper ore loses about 15 per cent, in weight and rises as much as 12 in. during the roast; if instead the surface sinks, there is furnished proof that the heat was too great or the admission of air insufficient, both of which would cause fusion. The stall in Figs. 52-53 holds 20 tons ore, requires 8-9 days for roasting, or 10 days between charges. Thus a stall treats 2 tons per day and, assuming 15 per cent, loss in weight by roasting, there will have to be roasted about 120 tons of raw ore to ftirnish 100 tons of roasted ore, and this demands 60 stalls. The cost of roasting lump ore in seven stalls, each with a capacity of 25 tons, as estimated by Henrich,* is i roast-yard foreman $2.25; 3 men preparing seven empty stalls and laying wood-foundation, at 50 cents per stall, $3.50; 8 men ' Tr. A. I. M. E., 189s, XXV, 232. 6 82 METALLURGY OF COFFER filling seven stalls, building front walls, ready for firing, at $1.50 per stall, $10.50; 14 men emptying seven stalls, at $2.50 per stall, $17.50; 4 cords of wood at $1.55, $6.20; 200 hard-burnt brick, at $5.00 per M., $1.00; tools, oil, repairs, clay, etc., $7.55; unforeseen expenses, $5.00; total for 175 tons, $53.50, or 30.57 cents per ton. The advantages of stall- over heap-roasting are: A more uniform distribu- tion and better utilization of heat, hence a smaller amount of wood; a smaller loss of ore by scattering and leaching ; a quicker roast, requiring a smaller locking- up of ore; easy disposal of gases. The main disadvantages are: Cost of plant; greater cost of labor; close attention to process on account of danger of insuffi- cient roast or of fused charge. At Keswick, Cal.,' stalls were replaced by heaps. In general, stalls will be used only with small amounts of coarse sulphide ore. 58. Roasting in Shaft Furnaces (Kilns) in General.^ — The furnaces are shaft-like structures of varying heights in which the ore rich in S is roasted with- out the use of carbonaceous fuel, the oxidation of S and Fe furnishing the neces- sary heat. The process is continuous, raw ore being charged periodically at the top and roasted ore drawn at the bottom; the gases containing over 4 per cent. vol. SO2 and being free from carbonaceous matter are suited for the manufac- ture of SO3, H2SO4, or H2SO3. The furnaces are usually classed as lump-ore and fine-ore furnaces. Lump ore ranges in size from J (perhaps i) to 3 in., a piece larger than 3 (perhaps 3.5) in. not being satisfactorily desulphurized in the center; the actual size within the range is governed by the more or less free- burning character of the ore. It may be necessary to carry the sizing farther and separate the coarse into three classes, 0.25-1.00, 1.00-2.00, 2.00-3.00 in. in order to obtain the best results. Fine ore, smaller than 0.25 in., is roasted separately. A lump ore cannot stand more than 10 per cent, fines; a larger amount blocks up the air-passages, which results in imperfect roasting and in clinkering. 59. Roasting Lump Ore in Shaft Furnaces. — The ore, in which pyrite is the usual sulphide, may not contain less than 25 per cent. S, if the S-content is to be reduced to below 4 per cent.; the lower limit in the eastern United States is 37 per cent. Usually the pyrite ore charged contains over 40 per cent. S. The ore, further, shoiild not be rich in Pb (10 per cent, causes clinkering), in Cu (< 8 usually < 5 per cent.), as chalcopyrite, the common form of copper mineral, decrepitates upon heating, causes clinkering, and furnishes gas low in SO2. The best material is coarse (< 8 per cent, fines), hard, non-decrepitating pyrite with < 6 per cent, gangue, and < 5 per cent. Cu. Thus' Spanish, Portuguese, some 1 Neilson, Eng. Min. J., iSgg, Lxvni, 458. 2 Jurisch, K. W., "Handbuch der Schefelsaeurefabrikation," Enke, Stuttgart, 1893. Lunge, G., "Sulphuric Acid and Alkali," Gurney and Jackson, London; Van Nostrand, New York, 1913, i, pp. 415-501. Wilson, W. G., "Pyrites in Canada," Canada Dep't. Mines, Mines Branch, Ottawa, 1912, pp. 94-132, "Roasting of Pyrites." Wyatt, Eng. Min. J., 1887, XLiv, 165. Falding, Min. Ind., 1898, vii, 665. ' Falding, loc. cii. SMELTING OF COPPER 83 Norwegian and Newfoundland ores contain 48 per cent, available^ S and 3-4 per cent. Cu; pyrite from New England,^ northeastern Canada, New York, and Virginia, 35 per cent, available S and < 3.5 per cent. Cu. Pyrrhotite in lump form does not roast satisfactorily in kilns. ; } .,n..>.J ^ ^ Section on Line EF Section on Line CD Fig. 54. — Oker pyrite-kiln. Furnace. — The furnace for burning pyrite is essentially a rectangular brick chamber 5-6 ft. sq. and 4-6 ft. high, with doors at the front for charging, stirring, and discharging, and an opening in the roof for carrying off the gases. In the chamber there are horizontal grate-bars reaching from front to back which carry the ore. The depth of the charge is regu- lated by the size and S-content of the ore, as is the draft, and with this the rate of roasting. (i) The Oker Kiln.' — A block of eight kilns built in two rows of four, back to 'back, is shown in Figs. 54'~SS- ^ is the roasting chamber 5 ft.- 9 in. wide, 4 ft. 9 in deep, and 3 ft. 3 in. high; B grate; C center wall; D flue for roaster gas, covered with cast-iron plates, on which rests a lead pan for concentrating dilute H2SO4; E damper; F chamber for iron or lead vessel charged with NaNOs and H2SO4 to generate N:rO!,-f umes, which pass through vertical circular, uptake D; G cinder pit with car to receive roasted ore through grate by rocking; the ends of the pit are closed by folding or sliding doors. The movable grates 1 This is the S remaining after the amount held back by constituents such as Zn, Cu, Pb, Ca, Mg in the form of sulphide or sulphate has been deducted from the total. ^Min. Res., V. S. Geol. Surv., 1885, p. 501. 3 Brauning, Zt. Berg. Hulten. Salin. Wesen., i Pr., 1877, xxv, plate rv. For other drawmgs see Treatises of Lunge and Jurisch, and paper by Falding, Min. Ind., 1908, vii, 666. Fig. 55. — Oker pyrite-kiln. 84 METALLURGY OF COPPER (Figs. 56-57) are square wrought-iron bars made circular in the bearing places so that they can be rocked by means of a T-handle key. By setting the bars on edge, the open spaces between them are smallest; when flat on the sides the spaces are largest; when alternately on edge and on side, the size of the open spaces is between the above extremes. Ore is fed from the front through the upper door, which otherwise is closed by a cast-iron plate, the joint being made air-tight by a clay lute. The central door, through which access is had to the grate bars, is closed by a hinged cast-iron door. The lower opening has a sUding door for regulating the admission of air. The ore-bed is only i ft. 5 f in. high. The eight kilns roast in 24 hr. 3.5 tons of pyrite 1.25 in. in diameter with 40 per cent. S, or 35 lb. of ore per square foot of grate area, the S being reduced to 4 per cent. ; with ore assaying 6-9 per cent. Cu, the roasted ore retains 5-6 per cent. S. Figs. 56-57. — Shaking-grate of lump-ore roasting-kiln. The leading operations are drying and heating, charging and discharging, loosening up, and regulating the air-supply. With the damper in the gas- flue closed, the furnace is dried by having a low fire on the grate and allow- ing the smoke to pass into the building through the charging doors. When dry, the roasting chamber is filled with burnt ore or with pebbles to about 3 in. below the charging door, and a wood or coal fire started upon the bed and kept going for about 24 hr. ; this will bring the upper layer of the bed as well as the roof to a dark red, whereupon, pyrite is charged, which will kindle and burn. When burning freely, the charging door is closed and the damper in the gas-flue opened. A kiln is usually fed every 12 hr.; fines are spread mainly along the sides, and some are scattered over the surface of the coarse ore. A charge weighs from 700 to 1000 lb. Roasted ore is discharged before feeding raw ore. Sometimes it precedes the feeding as much as 2 hr. in order that the ore-bed, loosened and freed from dust by the rocking of the grates, may roast freely and be thus better prepared for receiving the cold charge. The grate may need shaking between chargings in order to remove fines in the crumbled roasted ore. Any packing or caking of roasting ore is corrected by inserting a slice bar or poker. A furnace is in good SMELTING OF COPPER 8S working order when the charging door is too hot to permit touching with the hand, and when the air-inlet door is just warm. The success of roasting depends upon the air-supply (an excess gives dilute gas, a lack causes clinkering), the depth of bed (an excess gives too high a temperature, a lack insufficient desul- phurization), the percentage of fines (an excess chokes the draft and causes clinkering), and the rate of discharge (an excess gives imperfectly roasted ore, a lack reduction of capacity). The temperature of the escaping gas should not exceed 420° C. Table 21^ gives the distribution of Cu-Fe-S-compounds in 5 samples of kiln-burnt pyrite. Table 21. -Distribution op Cu-Fe-S Compounds in Kiln-burnt Pyrite, Per cent. CUS04 CuO CujS CuFeSa FeS2 Fe2.(S04)3 sss 0.89 i-SS 0.69 2-3S 0-59 1.67 0. 76 3 76 I .og 1.56 0.5S 3 -43 1 .26 2-54 0.00 4.32 2.2s 0.5s 0.46 1.27 0.72 7-34 2 . 29 0.82 0.2s o-Si 0-35 5-9° 0.68 Dimensions and work of some coarse-ore kilns are given in Table 22.^ Table 22. — Dimensions and Work of Coarse-ore Roasting Kilns Ore, source and kind Harz Mts. . Low grade. General. . . . Ireland. . . . Portugal. .. Spain Spain Spain Sulphur, per cent. 44 41 39 42 49 48 48 49 Grate area, sq. ft. 28.21 24.97 24.76 22. 72 22.72 10. 75 25.00 21 .64 Ore-bed, depth, in. 18. s 26,4 26.4 36.2 31S 21 . 7 26.4 24.0 Ore roasted, lb. in 24 hr. 96s 89s 1008 89s 89s 397 787 992 per. sq. ft. grate area 34 35 40 39 39 37 31 45 A block of over twenty furnaces is in charge of two men; with a looo-lb. charge of ore per kiln in 24 hr., this gives 0.4 day labor per ton of ore. Usually, how- ever, the blocks are smaller, as is the weight of charge; the labor necessary is correspondingly increased. (2) The Freiberg Kiln, Saxony.— This kiln, shown in Figs. 58-59,' is an oblong shaft furnace, 8 ft. high by 4 ft. 6 in. wide by 7 ft. 6 in. long; it has on the floor a saddle, g, of cast-iron plates, which deflects the roasted ore toward the discharge-openings /, through which the air necessary for roasting enters. Raw ore is delivered to the receiving hopper a, and discharged through chutes h, 1 Kothny, Oest. Jahrb., 1910, LVlii, 107; Metallurgie, 1911, viii, 390- ' Jurisch, op. cil., p. 73. ' Lunge, op. cil., p. 421. 86 METALLURGY OF COPPER into the furnace c; doors d, serve for spreading charge, and e as working doors. Five furnaces form a block. The material roasted is either pyritic ore with < 40 per cent. S, containing blende and arsenopyrite, or lead and copper matte with 22 + per cent. S. About 2,500 lb. pyritic ore is roasted in a kiln in 24 hr. The S-content is reduced to 4-5 per cent. The original furnaces of Oker have been reconstructed on the lines of the Freiberg kilns. There are two types: one for medium-S ore having a bed 6 ft. high, the other for low-S ore with a bed 9 ft. high. Ore consisting of ZnS 28, FeS2 25, CuFeS2 15, PbS 11 per cent, has its S-content reduced to ia-12 per cent. Figs. 58-59. — Freiberg coarse-ore kiln. Matte with 30-50 per cent. Cu clinkers in the kiln if roasted alone; it is satis- factorily rough-roasted when mixed with 25 per cent, in weight of similar matte' that has passed once through the kiln. The furnaces at Kedabeg, Caucasus,^ are 7 ft. 2 f in . by 3 ft. 11 J in. and 11 ft. i| in. high. A double furnace treats in 24 hr. 9 J tons of pyritic ore with > 5 per cent. Cu and >o.5 in. in size, reducing the S-content to 8-10 per cent. 60. Comparison of Heaps, Stalls, and Kilns. — A comparison of the three roasting apparatus shows that kUns have the advantage in that the process is continuous, independent of the weather, requires less time, gives a better elimination of S, and needs no fuel; the S can be recovered; there is practically no mechanical loss, nor any leaching-loss whatever. The disadvantages are the great size and cost of plant per unit of daily product, and the necessity of skilled ' Koelle, Tr. Inst. Min. Met., 1904-05, xiv, 508. Golowatschew, GWcA Aiif., i9I3,xlix, 733, transl. by Hahn, Eng. Min. J., igis.xcvi, 15. SMELTING OF COFFER 87 labor. In general, kilns are used only where there is profit in or necessity of not allowing the sulphurous gas to escape into the air. 61. Roasting Fine Ore in Shaft Furnaces. — Fine ore is produced in mining (mine-fines), in breaking coarse ore for kilns (smalls), and in ore-dressing plants (concentrates). In sulphuric-acid works the mineral commonly roasted is pyrite; in metallur- gical plants other metallic sulphides often take its place. Attempts to make sul- phuric acid from the gases obtained in roasting pyrrhotite were so unsatisfactory that the mineral was condemned as a raw material until Sjostedt^ showed in the sulphurous acid plant at Sault Ste. Marie, Ont., that with pyrrhotite assaying 24.46 per cent. S a gas averaging 7.91 per cent. vol. SO2 was obtained. The S in the roasted ore was brought down to 1.81 per cent., provided the loss of heat by radiation was reduced to a minimum by combining 4 furnaces of the MacDougall type in a single block, and by enclosing this with a bad conductor of heat. The furnaces used for roasting fine ore are oblong or circular shaft-like struc- tures into which the ore, fed intermittently or continuously at the throat, meets during its descent horizontal or inclined shelves, or prisms, which retard its down- ward progress sufficiently to effect the desired degree of desulphurization. The furnaces, provided with chambers to settle out dust, do satisfactory work with ores that run high enough in S (about 28 per cent.) to sustain combustion and do not become sticky while roasting. At first, fines were pugged with clay, 10 per cent, or more, and made into balls which, when air-dry, were roasted to the extent of perhaps 16 per cent, with coarse ore. Very fine dust pugged with water alone, molded and exposed to warm air, vitriolized superficially, hardened, and thus furnished a material that could be treated together with coarse ore. Untreated fines have been roasted in connection with coarse-ore furnaces, the gases from the latter serving to burn the fines (Olivier Ferret, older Haasenclever-Helbig furnaces^). These devices and furnaces which were formerly the transition between coarse- and fine-ore furnaces have become obsolete. The fine-ore furnaces proper may be classed (i) as Automatic, in which the ore glides downward through a shaft; and (2) as Shelf-burners, in which the ore is moved from shelf to shelf either by hand or mechanically. 62. Automatic Furnaces. The Gerstenhofer Furnace. — This is shown in Figs. 60-61. It is the only automatic furnace which is still in operation in a few plants. It consists of a shaft 17 ft. high, 2 ft. 7 J in. wide, and 2 ft. 3 in. deep, which contains from 15 to 17 horizontal tiers of staggered triangular fire-brick bars, 6 or 7 bars to a tier. The ore, fed continuously by fluted rolls (not shown), drops onto the upper bars, accumulates until it reaches its angle of repose of about 33 degrees, then glides on to the bars below, and so on until it arrives at the bottom. The fresh ore flows downward and displaces the particles previously lying on the bars and takes their places. The air necessary for oxidation enters near the bottom and meets the descending ore; the furnace- ' /. Canad. Min. Inst., 1904, vii, 480. "Lunge, op. cii., First Edition, 1879, pp. 193 and 200. 88 METALLURGY OF COPPER gases pass off on either side of the feed through flues leading into dust chambers. At the front of the furnace are openings, usually closed by cast-iron doors, to allow watching the process and inserting of tools to remove clinkers and clean the passages. The roasted ore is removed from the bottom of the shaft either by shoveling or by a screw conveyor. In starting, a temporary grate is put in place, and a fire maintained on it until the shaft has been brought to a bright red. Firing is continued for from 6 to 7 hr. until the fourth tier of bars from the bottom begins to receive ore, when the grate is removed, and the damper of each of the flues leading into the Charge Hole Charge Hole Figs. 6o-6i. — Gerstenhofer kiln. dust chambers drawn out. The admission of air has to be well regulated, for an excess drives the heat up, and makes the furnace too hot; a lack draws it down, and cools the furnace. The furnace works with ore of i mm. size containing from 25 to 36 per cent. S ; it fails with less than 25 per cent. S. It treats in 24 hr. from 2 to 5 tons of concentrate, reducing the S-content to from 8 to 10 per cent. The roast is therefore imperfect; much flue dust is formed, and repairs are likely to be costly. The furnace was first used at Freiberg, Saxony, in about 1866, but has been abandoned there; it was in operation in Swansea, 1 and is still in opera- tion in Kedabeg, Caucasus.^ At these works it has twice the width of the stand- ard furnace; treats in 24 hr. 12.5 tons of pyritic ore, < 0.25 in., with over 5 per cent. Cu, reducing the S-content to 6 per cent. The latest report' of this plant • Vivian, Eng. Min. J., 1881, xxxi, 231. ' Koelle, Tr. Instil. Min. Met., 1904-05, xiv, 510. « Golowatschef, GlUck Auf., 1913, xLrx, 732; transl. by Hahn, Eng. Min. J., 1913, xcvi, 15. SMELTING OF COPPER 89 states that there are 10 furnaces. A furnace is 7 ft. 4 in. by i ft. 9 in. and 24 ft. 6 in. high; has four feed openings, 14 tiers of fire-brick bars (10 bars 6.9 in. wide to a tier) and two oil-fired fire-places at the front. A furnace treats in 24 hr. II tons of ore, consumes 572 lb. oil, reduces the S-content to 3 per cent., and makes 5 per cent, flue-dust in which the Cu is present mainly as CUSO4. 63. Shelf-bumers.— (i) The Maletra Furnace.— Of the hand-raked shelf-burners that of Maletra constructed in 1867 was the first. A double fur- nace designed by Niedenfuhr is shown in Fig. 62.1 The roasting-chamber, 9 ft. long by 3 ft. II in. wide by 6 ft. lof in. high, has seven horizontal shelves, made of 3|-in. rebated fire-brick tiles, which extend over the entire width of the — K — H Fig. 62. — Maletra shelf-burner (dimensions in millimeters). furnace, but are 9 in. shorter than the length of the furnace and are staggered so as to leave open spaces alternately at front and back. The distance between the upper two shelves is greater than that between the lower five; above the top shelf it is 7 in.; between shelves i and 2, 6 in.; the other shelves are 5 in. apart. The ore, from dust to pea-size, is charged through a hopper in the roof, spread over the top shelf with a rabble or rake from the opening at the front, usually closed by a sliding door, and pushed after from 4 to 8 hr. through the slot on to the next lower shelf, when a fresh charge is dropped from the feed-hopper and so on. Air, admitted over the roasted ore in the pit, travels in a direction opposite to that of the ore; the gases deposit the suspended dust in part on the bottom of the flue connecting a set of kilns, in part in special dust chambers. The ore remains 4X7 = 28 or 8X7 = 56 hr. in the furnace. The roasted ore is moved from the chamber beneath the bottom shelf. In rabbling or in transferring ore as 'Lunge, op. cit., i,' p. 468; other drawing by Falding, Min. Ind., 1908, vii, 667. 90 METALLURGY OF COPPER little false air as possible ought to enter a furnace, as it dilutes the gas. Irregu- larities are avoided by having a number of furnaces in a block. A new furnace is brought to a red heat by burning fuel on the shelves for 4 or 5 days. Furnaces are built in single or in double rows forming blocks with from 8 to 16 kilns. A furnace roasts in 24 hr. about 2200 lb. so-per cent, pyrite, or 3200 lb. when of lower grade, or on the average from 6^ to 7^ lb. per square foot hearth area, reducing the S to i per cent, and lower. Pyrite ought to contain not less than 38 per cent. S. .Table 23 gives some data for a six-shelf furnace' charged every 4 hr., and for a seven-shelf furnace^ charged every 8 hr. Table 23. — Results of the Maletra Furnace Raw S, per Tem- After Raw S, per After Raw S, per After pyrite cent. 49-65 perature, °C. hours pyrite cent. 50 hours pyrite cent. 50 nours Shelf I . . . . 37-90 680 4 Shelf I... 32.27 8 Shelf I -■ 32.81 8 Shelf 2.... 30-13 75° 8 Shelf 2... 21 41 16 Shelf 2 -. 17 55 16 Shelf 3.... 10.10 720 12 Shelf 3... 12 77 24 Shelf 3 . . II 09 24 Shelf 4-. ■ • 1 . 26 650 16 Shelf 4.- - 6 39 32 Shelf 4 S 05 32 Shelf s.... 1,08 380 20 Shelf S-- ■ 4 08 40 Shelf 5 - 3 42 40 Shelf 6.... 0-S9 310 24 Shelf 6... 2 35 48 Shelf 6 2 S6 48 Shelf 7... 2 27 56 Shelf 7 I 96 56 Under normal conditions no blue flame is visible when the ore is transferred from shelf i to shelf 2; shelf 2 is bright-red, shelf 3 a shade darker and so on; the last shelf is dark. Just as in the Gerstenhofer furnace, a draft that is too strong causes the heat to creep up and cool the lower shelves, the reverse holds good with a draft that is too weak; in either case the ore on the hottest shelf begins to sinter. From four to five furnaces are given in charge of one man on each shift. (2) The Spence Furnace.^ — This furnace, shown in longitudinal vertical section in Fig. 63, is a shelf-burner in which the ore is mechanically fed from a hopper in the roof and then pushed at intervals from shelf to shelf by reciprocat- ing rakes until it is discharged at the bottom into a car. The rakes of the several hearths are provided with trippers so that they will act upon the ore only in one direction; the rods are connected outside of the furnace to a frame which is moved to and fro at certain intervals by steam or hydrauUc power (1-2 h.p.). The heads of the rakes when not in motion rest at the ends of the hearths on the parts of the shelves lying beyond the discharging slots. Thus a rake will travel over the hearth in i| min., come to a stop, rest for 5 min., and then make the return trip. 1 Krutwig-Dernoncourt, Rev. Univ. Min., 1898, XLi, 288. 2 Crowder, /. Soc. Chem. Ind., 1891, x, 298. ' Marten, Min. Res., U. S. Geol. Surv., 1883-84, p. 892. Adams, Tr. A.I. M. E., 1884-85, xra, 345. Wendt, School Min. Quart., 1885-86, vii, 322. SMELTING OF COPPER 91 The hearth of a furnace is 5 ft. wide and varies in length; a hearth 5 ft. by 16 ft. 8 in. inside has a roasting area of only 5 by 10 = 50 sq. ft. The furnaces are usually built in pairs with a common middle-wall. A double furnace will treat in 24 hr. from 3 to 4 tons pyrite with 46 per cent. S, and reduce this to about 2 per cent., or about 70 lb. in 24 hr. per square foot of hearth area. The original furnace has undergone several changes, notably at the hands of Adams and Hammond. The latter, ' among other important alterations, has provided the furnace with a fire-place which may be used only for warming, or also continuously and thus change the kiln into a reverberatory furnace if the flame travels over the ore, or into a mufBe furnace if it passes through flues Fig. 63. — Spence automatic desulphurizing furnace. placed between the hearths. In the United States the Spence furnace has been widely used in sulphuric acid plants, but has had to give way to the Mac Dougall type of kiln. 64. The MacDougall Furnace in General.' — This furnace is a vertical cyl- inder with superimposed horizontal hearths and central rotating shaft with radial stirring arms provided with teeth set at a proper angle. The ore, fed mechanic- ally at the top, is turned over by the rabble arms, moved on one hearth from the periphery toward the center, where it drops through a slot on to the next following hearth to be moved in the opposite direction that it may drop through slots near the periphery on to the third hearth and continue to travel until it is finally discharged from the bottom into a receiver. During the fall of the ore from hearth to hearth a large part of the S and Fe in the ore is oxidized and the required heat generated. The air necessary for oxidation enters through doors situated either wholly 'Peters, "Modern Copper Smelting," 1895, p. 221. ^ Lunge, "Sulphuric Acid and Alkali," 1913, i', 474- Benker-Hartmann, Zl. atigeu: Chcin., 1906, xix, 1125, 1188. Pierron, Rev. C/iim. Ind., 1907, xviii, 8. Wilson, G. W., "Pyrites in Canada," Canada Dep't. of Mines, Mines Branch, Ottawa, 1912, pp. 101-125. 92 METALLURGY OF COPPER on the bottom hearth or in part on one or more of the upper hearths, and travels in a direction opposite to the ore. This type of furnace, being automa- tic, does uniform work at a low cost, and permits full control of air and tem- perature which means a good roast; on the other hand it makes a considerable amount of flue-dust. The original furnace was in operation in Liverpool, England, about 1870, but was abandoned mainly on account of mechanical difficulties. In this country it was taken up again by Herreshoff in 1896, and later by others. The leading types at present are those of Herreshoff, O'Brien, Evans-Klepetko, with modifi- cations, and Wedge. 65. The Herreshoff Fximace.' — The original form of this furnace, shown in Figs. 64-65, has four or five hearths, and a centra air-cooled hollow shaft with Figs. 64-65. — Herreshoff furnace, original form. pockets to receive the hollow rabble-arms. The shaft ends in a 30-ft. stack. An arm. Fig. 66, has on the upper side of the inner end a projection, a, which fits into a corresponding cavity, b, in the pocket c, and being thus hinged is held in place by its own weight. The ore is fed mechanically near the center of the top hearth; the teeth form- ing part of the rabbUng arms are placed at an angle in order that, traveling over the hearth, they may not only turn over the ore, but move it gradually toward the periphery, whence it drops through slots on to the second hearth. The 1 The Nichols Copper Co., New York, N. Y. Herreshoff, Min. Ind., 1897, vi, 235; igoa, XI, 205. Gilchrist, /. Soc. Chem. Ind., 1899, xviii, 459. Hofman, Tr. A. I. M. E., 1904, xxxrv, 277. SMELTING OF COPPER 93 teeth of the pair of rabble-arms of the second hearth point in the opposite direction enclosing an angle of the same magnitude as those of the first hearth ; they therefore move the ore from the periphery to the slot near the center, and soon. At the works of the Montana Ore Purchasing Co., Butte, Mont., there were in operation in 1899, 70 furnaces, 10 ft. 10 in. in diameter and 11 ft. 6 in. high, made of J-in. steel casing, lined with a full course of red brick. A few furnaces with two full courses of brick, hooped with iron bands, worked more satisfactorily in the cold winter than the rest. A furnace had five arched brick hearths 4! in. thick and 11-12 in. apart, the pitch of the arch being 4I in.; the central shaft was 14 in. in diameter, and each arm had seven or eight teeth. The top hearth acted as a dryer for the wet concentrates fed, the second started the roast, on the third the ore was roasting freely with many sparks visible, on the fourth there was no spark- ing, on the fifth the ore appeared dark. With the shaft making 50 r.p.hr., a furnace roasted in 24 hr. Fig. 66.— Rabble-arm of Herreshoff furnace, original form, from S to 6 tons of wet concentrates with about 35 per cent. S, or 30 lb. per square foot of hearth area, reducing the S-content to 6 per cent, and making 6.4 per cent, flue-dust; the cost of roasting was $0.40 per ton. In eastern sulphuric acid plants the furnace treats in 24 hr., with the shaft making 30 r.p.hr., from 3 to 4 tons of Virginia pyrite with 44 per cent. S. or Tharsis pyrite with 49 per cent. S, or 19.6 lb. per square foot hearth area, re- ducing the S to 2J-3I per cent. Herreshoff patented^ a device for reducing the amount of flue-dust formed in the descent of the charge. Some temperature measurements have been published by Clevenger.^ The improved Herreshoff Furnace, 20 ft. inner diameter and six roasting and one drying hearths, installed at Clifton and Douglas, Ariz., is shown in vertical section^ in Fig. 67. The f-in. boiler-iron shell is lined with 8-in. fire- brick. The hearths have a spring of arch of 10.5 in.; they are built of 6- and 8-in. fire-brick, and furnish thus a support for the loose bottom on which the ore travels. At the periphery of an odd-numbered hearth are 54 drop-holes 18 by 14 in. An even-nvunbered hearth has a central discharge on an 8-ft. circle; the central shaft is 42 in. in diameter. The top hearth has four doors, the other hearths have six each. The central hollow cast-iron shaft is made in three sections connected by 'U. S. Patent, No. 729,170, May 26, 1903. ' Met. Chem. Eng., 1913, xi, 448. 'Min. Sc. Press, 1913, cvii, 683; Eng. Min. J., 1914, xcvii, 262. 94 METALLURGY OF COPPER tongue and groove joints; bolt-heads and nuts are protected by pockets filled with refractory material. The hollow rabble-arms are similarly connected with the shaft. Inside the 42-in. shaft is an i8-in. steel tube to which are bolted pipes (not shown) reaching into the rabble-arms. Air is forced into the tube at the bottom (5500 cu. ft. per minute at 1.75 oz. pressure), it passes outward Fig. 67. — HerreshofE furnace, improved form. in the pipes to the ends of the arms, returns between pipes and arms, and is either delivered as warmed air on to the bottom hearth or passes off at the top. The manner of attaching the rabble-blades to an arm is shown in Fig. 68. In some furnaces a hollow rabble-arm has a vertical partition extending to near the end, so that the coohng-air passes outward in one compartment and returns in the other. SMELTING OF COPPER 95 Fig. 68. — Rabble-arm and blades of im- proved Herreshoff furnace. The ore is discharged from the feed-hopper on to the top hearth mechanically by a passing arm which makes 60 r.p.hr. The space between the top of the fur- nace and the central shaft is sealed with sand; in the furnace calcined material replaces the sand. Part of the dust carried away by the gases is collected by a hopper which discharges on to one of the lower hearths. The same arrange- ment is found at Hayden, Ariz. It is expected that a furnace wUl rough-roast in 24 hr. about 80 tons of copper-bearing concentrate. At the works of the General Chemical Co., Edgewater, N. J./ a 16 -ft. fur- nace roasts in 24 hr. 5-6 tons Spanish pyrite reducing the S-content from 46 to 0.5 per cent. At this plant it takes from five to six days to exchange a broken rabble-arm, as the furnace has to be cooled down slowly and suffi- ciently to allow a man to enter and make the repair. In another Eastern sulphuric-acid plant the cooling air is introduced at the top of the central shaft, admitted on to the hearths through the rabble- arms, and the sulphurous gas drawn off near the bottom. The claim made is that by this means the desulphurization is carried farther than when the air travels upward, and that the amount of flue-dust formed is greatly diminished. 68. The O'Brien Furnace.^ — Fig. 69 gives a side elevation of a six-hearth furnace with mechanical screw-feed placed on a drying floor 6 ft. above the top of the furnace; the same feeding apparatus for dry ore placed on the furnace is indicated by dotted lines. To the right are seen the single gas-outlet and six doors; there are twelve in all, two for each hearth. The leading characteristics of the furnace are the air-cooling and the attachment of the rabble-arms, both shown in Figs. 70-72. The hollow built-up, cast-iron driving-shaft, tapering from top to bottom, has two vertical f-in. ribs making three compartments, a central c, for cold-air inlet, and two laterals, d, for hot-air outlet. It further has for each hearth two sockets, e, to receive the conical ends of the rabbling arms, which are locked in place by rib b, and socket a. On the lower inner side of socket e, is the opening /, for admission of cold air (natural or forced draft) from the central compartment c. The air travels outward in the rabbling arm under the horizontal partition m, returns above the latter, and leaves the arm through port g, into the hot-air compartment d. Putting in and taking out the arms horizontally permits diminishing the distance between hearths and having therefore a larger number of hearths or a greater hearth area for a given height than ordinarily; thus an O'Brien kiln with six hearths is only 13 ft. 2j in. high, outside. In Fig. 69 the central shaft, running on ball-bearings 'Illustration, Eng. Min. J., 1914, xcvii, 262. ' American Coke and Gas Construction Co., Camden, N. J. Falding, Min. Ind., 1900, ix,623. Editor, Irqn Trade Rev., April 20, 1905, p. 56. 96 METALLURGY OF COPPER held in adjustable supports, is seen to be driven from below by means of a flexi- ble plate, n, of large diameter, with steel rack driven by means of pinion o- In case any obstruction hinders the circuit of a rabble-arm, the plate jumps out of mesh, and the furnace comes to a standstill. Opposite pinion o, is a second pinion, p, through the shaft of which the screw-feed receives its motion by means of a sprocket wheel and chain. In Floof: Plate Fig. 69. — O'Brien furnace. The rated capacity of this six-hearth furnace is 7000 lb. 4S-per cent, free-burn- ing pyrite; in practice the amount treated in 24 hr. has shown a range of from 5000 lb. so-per cent, to 7500 lb. 40-per cent, pyrite when fairly free-burning. The following furnaces have become of some importance in the roasting of pyrite for the manufacture of H2SO4. The KaufEmann,^ Herkules,^ Heinze- 1 Erzrostgesellschaft, Cologne, Germany. ' Dr. L. Lutjens, Chemist and Engineer, Hanover, Germany. SMELTING OF COPPER 97 Freeland,' Merton, Hhe Sjostedt' already quoted, and the Renwick^ with a down- ward path of the gas current. 67. The Evans-Klepetko Furnace of Great Falls.*— The leading differences between this furnace and the two preceding are the size, the water-cooling of the central shaft and rabble-arms, and the form and attachment of the rabble- teeth. Figs. 73-76 give the details of the furnace, and Fig. 77 the general arrangement of building with two rows of furnaces at Great Falls, Mont. Figs. 70-72. — Rabble-arm of O'Brien furnace. Thefurnace, Fig. 73, a-d, is a vertical cylinder of f-in. boiler iron, 18 ft. 3^ in. high by 15 ft. 10 in. in diameter, lined with a full course of red brick; it stands on columns, a, 12 ft. high, to allow the roasted ore to be collected in hoppers, h, and thence discharged into cars. The kiln is divided into six roasting chambers or floors by brick arches, with a 9-in. spring, that are 3 ft. apart and have for the passage of ore and gases annular openings, c, on the odd-numbered, and pe- ripheral ones, c', six to a floor, on the even-numbered hearths. The outer half of ' Wedge Mechanical Furnace Co., iij Chestnut St., Philadelphia, Pa. ' Merton Furnace Co., 62 London Wall, London, E. C. '/. Can. Min. Inst., 1904, vii, 480; Pyrite Engineering Co., Carthage, N. Y. ^Mel. Chem. Eng., 1911, ix, 156. ' Allis-Chalmers Co., Milwaukee, Wis. Croasdale, Pac. Coast Miner, 1903, vii, 471. Sorensen, /. Canad. Min. Inst., 1903, vi, 306; Canad. Min. Rev., 1903, xxii, 87. Hofman, Tr. A. I. M. E., 1904, xxxrv, .277; Eng. Min. J., 1903, lxxvi, 122. Austin, Tr. A. I. M. E., 1906, xxxvn, 462. Moore, Eng. Min. J., 1910, lxxxdc, 1021. At Tooele Smelter, Min. World, 1910, xxiil, 944. Corwin-Rodgers, Tr. A. I. M. E., 1913, xlvi, 383. 7 98 METALLURGY OF COPPER Fig. 73&. Plan Fig. 73a!. Horizontal Section Figs. 730-1^.— Evans-Klepetko furnace Oi Great Falls, Mont. SMELTING OF COPPER 99 a hearth is 9 in. thick, the inner only 6 in., in order to reduce the weight at the center, to prevent contact with rabble-teeth when the hearth rises upon heating, and to furnish an offset for holding in place and horizontal the loose working- bottom (crushed limestone) which is 4 in. thick at the center and 9 in. at the periphery. In the center is the hollow, built-up, cast-iron driving-shaft, d, 9 in. inner diameter, to which are bolted opposite, hollow, cast-iron rabble-arms, e, two to each hearth, the arms of one hearth being set at 90 deg. to that of the other.i I Copper gaskets are used to make the joints water-tight. The shaft extends through the bottom and top of the furnace, is carried by foot-step bear- ing /, and driven from below at the rate of 38-60 r.p.hr. by bevel gear g; it is supported at the top in a journal-box, h, and carries a gear-wheel i (Fig. 75), which drives the mechanical feeder. The cooling-water from tank j (Fig. 77) is received in 6-in. pipe k, (Fig. 72), extending upward vertically about 8 ft. sAdjuatlnfeX Fig. 75.// -C^ Nut \\ / / p ^ Fig. 76, Rabble-arm and teeth Figs. 74-76. — Evans-Klepetko furnace of Great Falls, Mont. above the central shaft; it is then reduced to 3 in., extends downward toward the bottom of the driving-shaft, and has branches I, leading almost to the ends of the rabble-arms. The cooling-water thus flows through a 3-in. stationary pipe toward the bottom of the revolving shaft, and toward the ends of the stirring-arms through i-in. pipes. In its upward passage between shaft and pipe it takes up the return-water from the arms and discharges at the top through 2 or 4 spouts m (Fig. 72) into a stationary launder, n. About 20 gal. cooling-water are required per minute with the overflow-water at 80° C. A furnace has two 33-ton feed-hoppers, o, covered with a grating to keep out lumps; they are 10 ft. apart, 4 ft. 6 in. in diameter for 9 ft. 6 in., and then conical for 7 ft. 6 in.; at the bottom they are 3 ft. 6 in. by i ft. 2 in. (the cross-section of the opening in the roof of the furnace) . The weight of the ore-column pressing downward prevents any hanging of wet concentrate. The contents of the hop- pers are discharged continuously by two bars, /» (Figs. 74-75), passing to and fro under the mouth with a throw of 10 in.; the rods of the feeders are joined to a 1 Klepetko has patented (U. S. Pat. No. 744,359, Nov. 17, 1903) the use of three arms on each hearth; two arms are simply to stir the ore, the third is to move it. lOO METALLURGY OF COPPER cross-piece, q, which is pivoted in the center and connected by a rod to one end of a link, z, while the other is attached to the pitman, s, of a crank and spurwheel, the latter being geared to the central shaft of the furnace. The ore is spread on the top hearth to a thickness of 3-3.5 in. The tops of the rabble-arms are protected by cast-iron caps against rapid wear at the places where the ore-column comes down through the feed-openings. The two arms of a hearth have 7 and 8 cast-iron teeth, t, 8 in. long by 6 in. wide by f in. thick; Fig. 77. — General arrangement of building with Evans-Klepetko furnaces at Great Falls, Mont. the lower 3 in. which come in contact with the ore are chilled. The teeth (Fig. 76) have grooved collars, u, fitting over ribs, v, in the arms, and are slipped over the latter. As the teeth wear off, the ore builds on the loose working-bed; when they are to be replaced, a plow is first slipped over the rabble-arm and moved a little toward the center after every circuit in order to break up and bring to the surface any crust that had been formed. The teeth on the top hearth last one month, those on the bottom hearth seven months, the difference being due to wear caused by decrepitation; on the hottest hearths, the fourth and fifth, the teeth are attacked by the ore, thus, at Anaconda, pyrrhotite and chal- copyrite, hardly existing in the ore, formed on the teeth. SMELTING OF COPPER lOI Gases. — The gases,' having the composition SO2 2.25, SO3 0.53, 18.45, N 78.77 per cent, vol., and a temperature of about 315° C, pass off through two gas-flues, w, 2 ft. in diameter and 12 ft. apart, ending in a main, x, 6-7 ft. in diam- eter for three furnaces; this is provided with discharge-hoppers for removing flue- dust, which amounts to 4-6 per cent, of the weight of ore treated. The sheet- iron uptakes have been replaced by brickwork at Garfield,^ as the iron was quickly corroded. Furnace Battery. — From six to eight furnaces form a battery. They are placed 18 ft. apart from center to center in one direction, and 21 ft. 3 in. in the other. Three or four furnaces receive their motion from one main shaft, with which they are connected by friction clutches; 10 h.p. is required by six furnaces. 1st Hearth Znd Hearth 8rd Hearth 4th Hearth Bth Hearth 6th Hearth 100 KS 600 800 1000 C Fig. 78. — Changes of temperatures and sulphur-contents in two Evans-Klepetko furnaces. The cost' of a furnace is given as being from $5 to $6 per square foot of hearth area. The necessary draft for four furnaces is supplied by a lined steel stack, 200 ft. high, 20 ft. inner and 23 ft. 6 in. outer diameter. Starting Furnace. — In starting a furnace, crushed limestone is fed to form the working bottom. A fire of dry' long-flame wood is started from the three side-doors of the third and fifth hearths, is kept going for two days with an old, and for three to four days with a new furnace, when the walls will have been ' Austin, op. cit., p. 468. ' Moore, op. cit., p. 1023. ' Editor, Eng. Min. J., 1906 lxxxii, 226. I02 METALLURGY OF COPPER brought to a dark red. Concentrate averaging perhaps 25 per cent. H2O is now- fed; roasting begins on the second or third hearth, depending upon the strength of the draft (normal depression 0.3 in. water) and the number of cir- cuits of the arms (normal i rev. in 60 sec). Usually a blue flame is seen near the periphery of the second hearth; sparks become visible on the third; they cease on the fourth; the highest temperature is reached on the fifth. Changes in S-content and in temperature of two separate cases recorded by Austin and Croasdale are given in Fig. 78. Temperature. — ^The temperature can be reduced either by opening doors on the third hearth and thus checking the draft, or by feeding more low-S mate- rial (fourth-class ore screenings, slimes), than normally; it is increased by acceler- ating the machinery and by feeding more coarse ore or high-S material than normally. Usually, however, the speed of the arms and the rate of feed, and with it the tonnage and the character of the roast, are kept constant so as not to interfere with the subsequent smelting in the reverberatory furnace. , The ore-charge passes through the furnace in about 90 minutes, the necessary air being admitted through the three doors on the bottom hearth. While the ore drops from hearth to hearth, the upward gas-current carries with it dust which striking the roof adheres to it in part and forms crusts. In the same man- ner the ports become incrusted. The different parts of the furnace where the ore is found to stick are sometimes lined with cast-iron plates to facilitate removing the crusts; these form a welcome blast-furnace ore, especially as they run lower in Si02 than the calcine, e.g., 6 vs. 30 per cent. Si02. The dust, from 4 to 6 per cent, of the weight of the ore depending upon the size treated, is richer in S than the ore. Capacity. — A furnace treats, in 24 hr., 40 tons of Montana concentrate, (with S 35 per cent., Cu 10 per cent.) or 84 lb. per square foot hearth area, reducing the S-content to 7 per cent. ; a greater reduction causes the formation of Fe304, and this makes trouble in the subsequent smelting. Improvements. — Since 1894 some changes have been made at Great Falls in the details of construction and in the mode of operating which have resulted in a simplification of the rabble-arms, a strengthening of the hearth, an increase in tonnage without materially raising the S-content of the calcine, and a decrease in the amount of flue-dust formed.^ At present the rabble-arms in the first hearth have no water-cooling whatever, as this has been found to be unnecessary; further, the i-in. branch-pipes {I, Fig. 72) delivering coohng-water from the vertical 3-in. supply-pipe to the ends of the other rabble-arms have been removed, as incrustations of lime or oxide of iron, and suspended mud are likely to choke them. Cement Hearths. — The hearths since 191 2 have been built of cement con- crete instead of brickwork. The experience with this change has been satis- factory (May, 1913), as the hearths show no cracks, no wear, and incrustations are readily removed from the smooth surface. A concrete hearth. Figs. 79-81, made of i part Portland cement, 2 parts tailings sand, and 4 parts crushed slag, ' Corwin and Rodgers, Tr. A. I. M. E., 1913, XLVi, 383. SMELTING OF COPPER 103 is reinforced by two pair of concentric iron rings, each tied by 24 radiating o.2S-in. iron rods; one pair is placed 1.5 in. below the top, the other 1.5 in. above the bottom of the concrete; the inner ring is a 0.5-in. iron rod, the outer a f -in. rod. No.S Hearth N0.4 Hearth Figs. 79-81.— Concrete hearth in Evans-Klepetko furnace of Great Falls, Mont. Increased Draft. — The first experiments made in 1906 to raise the ton- nage of a furnace by increasing the draft were successful. Strengthening the draft without feeding more ore caused the furnace first to grow cool and then to go out. By raising the draft from a water-depression at the furnace of 0.139 to 0.93 in., the tonnage could be increased from 36.1 to 45.7 tons without any change in the percentage of S; the S-content of the calcine became too high I04 METALLURGY OF COPPER Figs. 82-83. — Center drop-holes, 14 and 18 inches wide, in Evans-Klepetko furnace of Great Falls, Mont. when the feed was increased to 50 tons per day. This increased tonnage caused the center drop-hole in the third hearth to become more incrusted than usually; the hole, 14 in. wide, Figs. 82-83, of one furnace incrusted more quickly than that of another furnace where it was 20 in. wide; in fact 7 per cent, more time was lost with it in barring, etc., than with the larger hole. This shows that either the distance of 14 in. must be increased to 20 in., or additional drop-holes provided. Enlarging the central hole caused the S-content in the calcine to become higher than desirable; the alternative is the present practice; the third hearth. Fig 86, has a 14-in. central drop-hole and two extra drop-holes, 6X7 in., close to the edge of the central hole. The inner blades of the rabble-arm in the third hearth, one to an arm, are set to push the ore away from the center so that most of the ore is dropped at intervals through the two extra drop- holes instead of being showered continuously through the cen- ter-hole. Speed of Revolution. — A third factor in the roasting power of a furnace is the speed of revolution of the central shaft. A high speed keeps the ore-bed shallow and counteracts the banking of ore in front of the rabble-arms; that speed is best which exposes a new surface to the air as soon as the rapid oxidation of an old one begins to cease. Thus the old speed of i rev. in 53 sec. has been abandoned; four furnaces have i rev. in 45 sec, and the others 18 have i rev. in- 38 sec, with the result that the amount of charge treated in 24 hr. has been increased from 47.2 to 88.8 tons. With an increased tonnage there is to be expected a rise in temperature. If this becomes excessive, it must be decreased by reducing the fuel value of the charge; this can be accomplished either by charging screenings from first-class ore (< 5 in.), or limestone required as flux, or both. Samples of roasting ore taken in igii gave upon analysis: Screenings, Cu 7.10, Si02 46.9, FeO 16.6, AI2O3 9.7, CaO 0.1, S 17.4 per cent.; and fine concentrate, Cu 7.45, Si02 24.0, FeO 32.8, AI2O3 6.1, CaO o.i, S 33.4. On account of their low heat- value the screenings cannot be roasted by themselves without the use of extra- neous fuel (coal or oil on the fifth hearth) even when air is blown into the fourth and fifth hearths from the sides or through radial perforated arms to hasten oxidation; in addition the drop-hole on the second hearth has to be narrowed to cause satisfactory ignition on this hearth. As limestone remains unchanged in its passage through the furnace, it acts only as a diluent and not as a heat-absorbent; hence it is advantageously charged at the outer edge on the sixth hearth, where it becomes mixed with and warmed by the hot calcine. Its place is advantageously taken by screenings from first class ore. In this way the tonnage was increased in February 1913 to 105.08 SMELTING OF COPPER 105 tons charge (97.09 concentrate and 7.99 screening) with 11.6-13.1 per cent. S in the resulting calcine. Experience has shown that while the furnace can readily roast a 100- ton charge in 24 hr. with the central shaft making i rev. in 38 sec, the machinery cannot stand the strain. At present (record of April, 1913), the regular tonnage is 77.1 tons copper-bearing material with a reduction of the S-content in the calcine to 8.9 per cent. .Flue-dust. — The tests made to reduce the percentage of flue-dust were begim in 19 1 1 by remodeUng a furnace according to the device of Repath and Marcy^ which provides separate ports for the passage of the ore and the gas. They showed that while the amount of flue-dust formed was reduced from 17.6 to lo.o per cent., there took place with it a reduction in tonnage, 76.85 to 45.02 tons dry concentrate.^ Crouse Equipment. — In 1912 tests were begun with the Crouse equip- ment, shown in Figs. 84-88, which carries the ore from the first hearth to the second, and from the second to the third through extra drop-holes, while the gas travels upward along its usual path. The extra drop-hole, 7X8 in., is situated about 4 in. from the center drop-hole; the two inner blades of the rab- ble-arm, one to an arm, push the ore away from the center toward the extra hole. The rabble-arms on the first, second, and third hearths are placed so as to be in line one above the other. The rabble-arm of the second hearth carries two iron boxes, open top and bottom, which extend from near the floor to close to the roof; the arrangement on the third hearth is similar. When a blade of an arm in the first hearth pushes the ore toward the extra drop-hole, the receiving box on the arm in the second hearth passes below the hole and protects the drop- ping ore from the gas-current. The same occurs when the ore from the second hearth drops through the side-holes on to the third hearth. The extra, 6X7 in. , drop-holes on the third hearth permit using a 14-in. center-hole. Fig. 82, without incurring the trouble of heavy incrustation. Spark Catcher. — The vertical section shows a spark-catcher on the fourth hearth. It consists of horizontal plates, which, extending in front of and in back of the ends of the arms, catch the sparks that are formed where the rabble-arm pushes the ore through the side drop-holes, and thus prevent the building of incrustations on the roofs. The results of a five-day test showed that the regular furnace treated in 24 hr. 61.78 tons dry concentrate, reduced the S-content to 10.3 per cent. = 81.3 per cent, "elimination, and made 17 per cent, flue-dust; the corresponding figures for the furnace with the Crouse equipment were 63.079 tons, 9.6 per cent. S, 81.2 per cent, elimination, and 10.7 per cent, flue-dust. Great Falls furnaces are being provided with the equipment. 68. Evans-Klepetko Furnace at other Smelteries. — This furnace was first built at Great Falls and has undergone modifications in various ways at copper smelteries to meet local conditions and to make the handling of ore and calcine ' Drawing in paper of Corwin and Rodgers. ' The paper by Corwin and Rodgers gives a method of determining the amoimt of flue- dust formed, without stopping the furnace. io6 METALLURGY OF COPPER Hearth No. 1 Hearths Nos. 2 and 4 * . The center drop hole is now used as a gas Usually one or more of these side drop passageway only, and extra drop hole for holes are covered to damp the air current. concentrates. Discharge Hoppers Hearth No. 3 Hearth No. 6 The two extra drop holes through which The discharge drop holes are 8 in. wide the calcines fall to the hearth below, localize and 28 in. long: on outer arc. and reduce incrustations • Extra Drop Hole 2nd Hearth" BeceivLng Box 3rd Hearth Becelvlng Box '[jijijijijijijij r^ i Figs. 84-88.— Grouse equipment for dust -prevention in Evans-KIepetko furnace, SMELTING OF COPPER 107 more and more mechanical. A few of the more important examples are given to illustrate the changes. (i) The Washoe Smelter, Anaconda, Mont. ^ — The plant has 64 furnaces contained in a steel building, 96 by 142 ft. The furnaces are placed in 16 rows ; j izzz : ^=~~=P: ^ m ^ m j^ Figs. 89-92. — Kelly-Thomson rabble-arm of Evans-Klepetko furnace at Anaconda, Mont. each row of foiur receives its power from one main shaft; a steel structure with three working platforms embraces two rows of furnaces. The furnaces, placed 18 ft. apart in one direction and 21 ft. 3 in. in the other, are carried by steel frames 13 ft. 2 in. high. A furnace is 18 ft. 5 in. high outside and 16 ft. 6j in. in diameter inside and 18 ft. outside, and has six hearths. The first, third, and fifth floors each have a center drop-hole, 16 in. wide, and three doors; the second and fourth floors have six side drop- holes, 18 in. wide, and the sixth hearth 2 side drop-holes, 14 in. wide, and four doors. The water- cooled shaft makes i rev. in 60 sec. The rabble used is the Kelly-Thomson^ shown in Figs. 89-92. The i-in. pipe which formerly delivered the cool^ ing- water to the end of the arm, shown in Fig. 93, is omitted. In the ordinary arm, Fig. 93, the joint A becomes incrusted so that it is hard to remove the blade; further, leakages are not infrequent at this place. In the improved form, the joint of arm and blade is better protected against the calcbe and is coated with tar; then the top of the arm is provided with ears which carry flue-dust and thus counteract the cooling-eflfect of the circu- lating water. The ore-charge is made up of 96-97 per cent, concentrate, 4-3 per cent, screening of first-class ore ( < | in.), some MacDougall flue-dust (analysis below), ' Tr. A. I. M. E., 1904, XXXIV, 258 (Hofman); jgofi.xxxvii, 431 (Austin). Min. Ind., 1903, xii, 98; 1906, xv, 254 (Austin). Peters "Practice of Copper Smelting," 1911, p. 78. 'Eng. Min. J., 1911, xci, 455. Fig. 93. — Original rabble- arm in Evans-Klepetko fur- nace at Anaconda, Mont. io8 METALLURGY OF COPPER and 3-5 per cent, limestone. The charge dumped from cars into the 33-ton feed-hoppers, passes in about 3^ hr. through the furnace. The furnace treats in 24 hr. about 40 tons charge assaying Cu 7-8 per cent, and S 3 2-33 per cent., reduces the latter to about 8 per cent.; has a draft of 0.5-0.6 in. water; requires 1.7 h. p. and 20 gal. water per minute with the overflow at 80° C. The loss in weight of the raw ore is about 18 per cent. A screen analysis of the calcine' gave, >io-mesh, 10 per cent.; 10- to 30-mesh, 25; 30- to 80-mesh, 50; <8o- mesh, 35 per cent.; and a chemical analysis of a month's average sample, Si02 27.7, FeO 39.0, CaO 2.4, S 8.0, Cu 8.66 per cent. A similar analysis of flue- dust showed Si02 29.0, FeO 21.8, CaO 0.9, S 14.3, Cu 7.73 per cent. The 64 furnaces are attended in 24 hr. by 76 men; the cost of roasting is about 30 cents per ton with labor at $3.00 for 8 hr. 60 Angle \ Cored Hole in Lue for Lifting out Rakea Section on AA 1i lb. Rail Section on'B B 2J4 Cored Holes 6-Nozzles per Arm Hole for Rail Lobs 3 Thick Net Area Approz. 35 Sq. In, Ohipping Surface in Main Shaft Section showing Joint between Sections of Center Shaft Figs. 94- -Shaft and rabble-arms of MacDougall furnace Steptoe Valley smeltery. (2) The Steptoe Valley Smelter, McGill, Nev.^ — There are 16 fur- naces, which are 18 ft. in diameter and have six hearths. The floor of the first hearth is built of i-in. cast-iron plates strengthened by i-in. ribs, which hastens the drying of the concentrate containing about 10 per cent, water. The center shaft in air-cooled; air under a pressure of 2 in. enters- at the bottom of the shaft and leaves through the arms. The construction of shaft and rabble- arms is shown in Figs. 94-98. The parts constituting the shaft are connected by bayonet couplings, one of which is shown in Figs. 94 and 95. The shaft, 16 in. in diameter, has oval openings for the rabble-arms, which have the form shown in Fig. 98; each of the two openings is provided with two 3-in. lugs, which carry a 7S-lb. rail serving as support for the pair of arms of the hearth. The arms (Fig. 96), are slipped over the rail and wedged to its ends by wrought ' Offerhaus, Eng. Min. J., 1908, Lxxxv, 1238. 2 Sorensen, Eng. Min. J., 1913, xcv, 1273. SMELTING OF COPPER 109 iron keys and at the same time pressed against the flanges on the shaft. In this manner they are firmly connected with the shaft and can be readily removed when necessary. The cast-iron rabble-blades (Fig. 96 and 98), are provided with ears or trunnions by means of which, they are suspended in hook-shaped lugs of the rabble-arm; a blade has a cored hole through which it can be lifted off the lugs. The lack of S in the ore makes extraneous firing necessary. An oil-burner with a steam atomizer is used on the third hearth to ignite the ore. Roasting takes place mainly on the fourth and fifth hearths; often a flame is seen when the ore drops from the fifth to the sixth hearth. The shaft makes i rev. in from 38 to 55 sec. The concentrate, which makes up the whole charge, assays Cu 10.6, Si02 32.1, Fe 22.0, AI2O3 5.0, S 24.7 per cent.; the calcine Cu 12.5, SiOz 31.9, Fe 25.7, AI2O3 5.8, S 11.7 per cent. In 24 hr. a furnace treats 87 tons of dry concentrate ( = 90 tons wet), makes 2 per cent, flue-dust, consumes 27.4."lb. oil (14° Be and < i pet cent. H2O). The depression at the outlet-flue is only 0.07 in. water and the temperature 260° C. Limestone is not now fed with the ore nor on to the last hearth, as no advantage accrued from doing this. The 16 furnaces of the plant require in 24 hr. three foremen and 25.5 men excluding delivery of ore. (3) The Garfield Smelter, Garfield, Utah.^ — There are in operation 14 furnaces with shells 19 ft. 6 in. in diameter and 16 with shells 18 ft. in diameter. The internal diameters are 18 ft. and 16 ft. 6 in. The central shaft and arms are air-cooled; each furnace is supplied with about 2000 cu. ft. air per minute at a pressure of 4 oz. The charge is delivered to the furnace by larry cars. It consists of concentrate varying in size from |-in. down to 200-mesh. The chem- ical analysis of an average charge is Cu 10, Si02 24, Fe 21, CaO 6, AI2O3 3, S 27 per cent. Additions of limestone are made on the fifth hearth. No extrane- ous fuel is used except in starting. An i8-ft. furnace puts through in 24 hr. 60 tons charge; a 19 ft. 6 in. furnace 70 tons. The calcine contains Cu 12 and S 10 per cent. There is made about 2 per cent, flue-dust. An average analysis of flue-dust shows: Cu 5.2, Si02 30.3, Fe 6.6, S 3.5, SO3 30.5, AI2O3 5.8, Pb 0.5, Zn 1.3, CaO 2.8, MgO 0.8, As 0.3. The draft in the roaster flue ranges between 0.25 and 0.30 in. water. The power required is 3 h.p. (4) The Tooele Smelter, Tooele, Utah.^ — There are 32 furnaces arranged in four rows. They are 16 ft. inner diameter and have six hearths; on the third hearthisanauxiliaryfire-placeforstarting the furnace when 1-2 per cent, fuel is mixed with the ore. The water-cooled central shaft takes 7500 gal. water in 24 hr. and makes 1 2 rev. in 60 sec. The charge, deHvered by belt to hopper, is 5-in. and smaller; it consists of pyrite with chalcopyrite, bornite, etc., in a siliceous gangue and contains as much as 10 per cent, sulphide copper concen- trate. An average analysis shows Cu 3.56, Si02 26.44, Fe 30-21, S 19.99, ' Private Notes, 1912; Private Communication, i9r3. 2 Thompson-Sicka, Tr. A. I. M. E., 1913, xlvi. Austin, Mill. Ind., 191 1, xx, 235. Private Notes, 1912; Private Communication, 1913. no METALLURGY OF COPPER Ca0 4.36 per cent., Auo.ii and Ag 1.5 oz. per ton. On the fifth hearth about 10 per cent, siliceous copper ore (SiOa 70-80 per cent., S none) is added. The draft in the discharge- flue is 0.2 in. water. A furnace roasts in 24 hr. 45 tons of charge; the calcine assays Cu 3.2 and S 7.4 percent.; there is made 4.5 per cent, flue-dust with Cu 3.45, SiOa 26.0, Fe 26.0, S 13.5, CaO 6.0 per cent., Au 0.16 and Ag 1.85 oz. per ton. The power including line and countershafting and motor drives is 2 h.p. The labor for four furnaces per shaft is one furnaceman and one laborer. (s) Copper Queen Smeltery, Douglas, Ariz.' — The new reverberatory- furnace division is laid out with 16 roasting furnaces placed in two rows. There are in operation nine air-cooled furnaces, 16 ft. inner diameter and six hearths. A furnace has three hoppers, two for ore and one for flux (limestone); an oil- burner is used in starting. The shaft makes i rev. in 60 sec. The concentrate to be roasted is assembled in 20 concrete bins, 16 ft. long by 15 ft. wide by 16 ft. deep, laid out in two divisions separated by the space taken up by the elevator; each division has two rows of five bins placed end to end, between which run surface-cars bringing in the concentrate. The bins are V-shaped, and closed at the bottom with 8- by 12-in. boards; they are discharged on to belt-conveyors, one for each row of five, the bottom boards being removed one after the other and replaced as soon as there is enough room. The belt-conveyors discharge into a pit from which the elevator delivers to bins above the roasting furnaces. The three furnace hoppers, holding together about 75 tons of material, are fed during the day from the bins alone by belts and trippers. The chemical analysis of the ore shows Cu 13.9, Si02 12.6, Fe 31.5, AI2O3 3.6, S 34.0 per cent. Limestone is fed from the central hopper through an exter- nal pipe on to the fifth hearth; thus the reverberatory-furnace charge is made up in the roasting furnace. The draft is 0.5 in. water. A furnace treats in 24 hr. 50 tons of charge; the calcine contains Cu 15.8 and S 11. 8 per cent.; there is made 4.5 per cent, flue-dust with Cu 13.6, Si02 27.5, Fe 19.0, CaO i.i, AI2O3 6.6, and S 15.8 per cent. The power required varies from 2 to 5 h.p. according to the condition of the furnace; the labor for eight furnaces per shift is one furnaceman and two helpers (one Mexican); the charging crew is separate. (6) Hayden Smelter, Hayden, Ariz.^ — There are in operation eight fur- naces, arranged in a single row, 18 ft. and 21 ft. 9 in. in diameter; four furnaces have 5 hearths, and four 6 hearths; the center shaft is air-cooled. The charge consists of concentrates, which are delivered by belt-conveyors and trippers. A screen analysis of the concentrate shows: > 20 mesh 11.89 per cent.; 20-40 mesh 14.89; 40-80 mesh 17.21; 80-100 mesh, 11.82; iod-120 mesh, 14.95, < 120 mesh, 29.64 per cent. A partial chemical analysis gave Cu 18.75, S 23.2, Fe 21.6, Si02 26.2 per cent. From 2 to 3 gal. California oil are burnt per ton of charge. A furnace puts through in 24 hr. from 90 to lo'o tons of charge, which has a depth of from 6 to 8 in. The calcine assays Cu 19-2 1 per cent, and S 14-15 1 Private Notes, July, 191 2 and Communication, Aug., 1913. 'i Private Notes, 1912; Private Communication, 1913. SMELTING OF COPPER ><-,-$;S/->i 112 METALLURGY OF COPPER per cent. The flue-dust collected in the hoppers of the roaster building amounts to about 0.75 per cent, of the charge; a partial analysis gave Cu 9.9, Si02 52.8, Fe 7.8, S 6.3 per cent. The furnaces are run with a draft of 0.2 in. water. The eight furnaces require in 24 hr. six men to supply ore* and nine men for attendance. (7) Cananea Smelter, Cananea, Mexico.^ — There are in operation 10 fur- naces (Figs. 99-1 11) arranged in two rows; a furnace has an inner diameter of 16 ft. 6 in. and six hearths. The center shaft is water-cooled with from 25 to 28 gal. water per minute. It makes 47 r.p.hr. The charge is made up by bedding and reclaiming as are the blast-furnace charges (§177). The charge consists of concentrate (Cu 9.10, SiOz 19.6, AI2O3 4.5, Fe 26.5, CaO 1.3, S 29.0 per cent.) and siliceous flux. A furnace treats in 24 hr. 40 tons (dry) charge and burns 0.026 bbl. of oil per ton charge; the calcine assays Cu 10.83 ^^'^ S 9.9 per cent.; there is made 6.54 per cent, flue-dust averaging Cu 9.01, Si02 37.9, AI2O3 7.7, Fe 13.8, CaO 1.5, S 12.7 per cent. The draft is 0.45 in. water. The 10 furnaces are run by two 20 h.p. A. C. motors. 69. The Wedge Fiimace.^ — The leading features which distinguish this furnace, shown in Fig. 112, from the preceding Evans-Klepetko and its modifica- tions are: a central accessible air-cooled driving shaft, 4 ft. in diameter, which without bottom-step runs on roller bearings; severally cooled rabble-arms which are removable through the central shaft; a large diameter (20 ft.) made feasible by the manner of support of the rabble-arms; a mechanically stirred drying-hearth for wet ore, which forms the top of the furnace. The furnace, 21 ft. 7 in. in diameter outside and 27 ft. 5 in. high, has seven roasting-hearths and one drying-hearth. It is jauilt of a |-in. steel shell lined with a full course of red brick; the shell stands on columns 5 ft. 8| in. high of structural steel to allow for automatic discharge of roasted ore into cars. The roofs are arched with a spring of 13 in., the floors are level, the hearths are 16 in. apart (2 ft. 9I in. centers) and are built of special brick. Each hearth has the usual two rabble-arms, which make i rev. in 30 sec. The central driving shaft (53) is air-cooled, 4 ft. in diameter, of |-in. steel plate, to which are riveted cast- • iron furnace-arm holders (47); at the top it carries the dryer-arms (50). The furnace-arms (48) are provided with cast-steel breech blocks by means of which a man inside the shaft can loosen or fasten the holders. An exchange of arm including water connections is made inside of 30 min. The dryer-arms (50) are provided with adjustable plows (see below). The shaft (53) is supported by a master-gear (26) 12 ft. 3 in. in diameter, which is centered by a center pin (29) and provided with a cast-steel race-way (28) running on rollers (32) supported by pedestals (34). Power is derived from driving pulley (i) making 100 r.p.m., bevel gears (5 and 6), vertical intermediate shaft (16), worm and wormwheel (15), lower horizontal shaft (18), bevel pinion (23) (provided with safety-pin (25)), and master-gear (26). On top of the center shaft is the steel water-pan (42) supplying severally the cooling-water to the rabble-arms through lead pipes 'Private Notes, 1912; Private Communication, 1913. 2 The Wedge Mechanical Furnace Co., 115 Chestnut Street, Philadelphia, Pa. SMELTING OF COPPER 113 (54) provided with regulating-valves; the outlet pipes of the arms (55), as well as the overflow pipe from water-supply pan, end in a 4-in. cement water-receiving pan in the foundation, which has a 6-in. outlet. The hollow cast-iron rabble-arms ':i'i m^ Hed Brick Wrm r,re Brick Fig. 112. — Wedge furnace. (48) each have a central partition to enable the cooling-water to circulate; each furnace requires 32 gal. water per minute. The rabble-arms are so placed that they are not touched by the ore dropping from hearth to hearth. They are provided with lifting blade-holders (45) and 8 114 METALLURGY OF COPPER blades (49) in order to prevent breakage of arms in case the blades strike any obstruction; a blade-holder is readily removed without stripping the arm, and the blade without disturbing the holder. The dryer-arm (50) is also provided with plow-holders (51); the plows are so secured to the holders that they can be lowered as they wear off; they do not drag on the hearth. The ore is fed by an elevator or belt to a lo-ton hopper (not shown) at the edge of the top of the furnace. The hopper has a rectangular discharge through which the ore drops on to the dryer-hearth, whence the plows of the dryer-arm move it gradually toward the center. There is a cast-iron feed-box (41) with removable feed- blades by means of which the amount fed to the furnace is regulated. The ore forms an air-tight seal and prevents the escape of any gas. Any large pieces not intended for the furnace cannot get into it, and are readily removed when noticed by the attendant. Drying the ore before feeding into the furnace has greatly reduced the formation of accretions, and keeps all moisture out of the dust-flues. The furnace has a rectangular gas-flue (not shown) leading from the top hearth to the main flue. The bottom of the gas-flue is inclined in order to allow the flue-dust to glide downward to the cleaning doors in the side of the main flue. From 8 to 10 sq. ft. flue-area is required for a furnace; the power necessary is 2 h.p. ; two men on a shift tend four furnaces. This seven-hearth furnace, 20 ft. inside diameter, with a hearth area of 2381 sq. ft., treats in 24 hr. 100 tons sulphide copper ore with 35 per cent. S, reducing the S to 7 per cent., or 84 lb. per square foot hearth area. In Eastern sulphuric-acid works the capacity is from 16 to 20 tons pyrite with 50 per cent. S, roasting down to below 2 per cent. S, or 15 lb. per square foot hearth area. The temperatures of a furnace, treating 100 tons sulphide copper ore with Table 24.- -Temperatures in Wedge Furnace Hearth No. 1 I 1 2 3 1 4 5 1 6 1 7 Degrees C 1 660 1 700 66s 1 620 370 200 1 <26o 29 per cent. S and 15 per cent. Cu, are given in Table 24. The temperature of the second hearth is the highest. In a sulphuric acid plant treating pyrite, the highest temperature (880° C.) is usually attained on the third hearth, while that of the seventh hearth falls to 150° C. The cost of roasting in an Eastern acid plant with four furnaces treating 60 tons pyrite, in 24 hr.,was in one instance $0.18 per ton, i.e., labor $0.12, repairs 0.03, sundries 0.03. At the new smeltery of the United Verde Co., Clarkdale, Ariz.,' six Wedge furnaces, 21 ft. 6 in. in diameter, with air-cooled rabble-arms have been installed; the air enters the arms under a pressure of 2 oz. The ore is brought from the storage-bins by belt-conveyors to the furnaces. ' Eng. Min. J., 1913, xcvi, 289. SMELTING OF COPPER IIS 70. Roasting in Reverberatory Furnaces.— This form of furnace is suited for all sulpliide copper ores that have been crushed to tf or perhaps \ in. ; matte will have to be crushed more finely, say to tV in. As the usual aim is to rough-roast only, reducing the S-content to about 7 per cent., the imperfectly roasted coarse and the overroasted fine particles will furnish a satisfactory average. The far- ther the expulsion of S is to be carried, the lower will have to be the limiting size of the ore, the more uniform the grain, and the slower the operation. 71. The Single-hearth Hand Reverberatory Furnace.— This form of furnace is httle used at present for rough-roasting sulphide copper ore that is to be smelted. Horizontal and vertical sections of one form are given in Figs. 113 and 1 14.1 A furnace with a hearth 60 by 16 ft. will roast in 24 hr. as much as 16 tons of free-burning pyritic ore, reducing the S-content from about 35 to 7 per cent.; this is equal to roasting 33 lb. ore per square foot hearth area. Longitudinal Section ffh£#l ^^ m Figs. 113-114. — Single-hearth hand reverberatory roasting furnace. In operating, 2 tons of raw ore are charged and their equivalent of roasted ore drawn every 3 hr. There are required two men in an 8-hr. shift and from i to 2 tons bituminous coal depending upon the quality, or 0.4 days labor and 0.094 ton coal per ton of raw ore. With ore that is not free-burning, the roasting capacity is smaller, perhaps 12 tons in 24 hr., the 2-ton charge being shifted every 4 hr. The depth of ore-bed varies with the character of the ore from 2.5 to 4 in. Thus, e.g., at the Parrot works,^ a furnace 60 by 16 ft. treated in 24 hr. 10 tons concentrate, assaying from 16 tb 40 per cent. Cu and about 35 per cent. S, in charges of 3600 lb., reducing the S-content to 7 per cent.; it consumed 2.25 cords of wood, and required four men. At the Orford Copper Co.'s plant,' a furnace 60 by 12 ft. treated in 24 hr. •Peters, "Modern Copper Smelting,'' 1895, p. 175. 2 Peters, Min. Res. U. S. Geol. Surv., 1883-84, p. 391. ' Howe, " Copper Smelting," p. 104. ii6 METALLURGY OF COPPER 12 tons matte, assaying from 40 to 50 per cent. Cu and 27 per cent. S, in 4000-Ib. charges, reducing the S-content to 5 per cent.; it consumed 2 tons of coal and required eight men. The thickness of bed was 7 in. The amount of flue-dust formed in a hand reverberatory furnace is from i to 2 per cent. ; the total cost of roasting in the western states is from $1.50 to $2.00 per ton. The two-hearth furnace raked by hand is used only in very exceptional cases, when there is not room enough for one with a single hearth. While the hand reverberatory roasting furnace fulfils all the conditions necessary to obtain a satisfactory roast, the quantity of its product is small, as it is not possible to renew by hand-work the surface of the roasting ore with sufl&cient frequency to hasten the process in the required degree. The small capacity makes the cost high. Mechanical furnaces have therefore replaced most hand-furnaces for rough roasting. 72. The Edwards Furnace.' — This furnace was designed in Australia for dead-roasting gold ores previous to cyaniding. Its work has been so satisfactory that its field has been extended to rough-roasting sulphide and arsenical ores. Figs. 1 1 5-1 2 1, from drawings of the Steams-Roger Mfg. Co., Denver, Colo., represent the Simplex Stationary Furnace. It is a single-hearth, straight-line, mechanical reverberatory furnace with a roasting hearth, a, 76 ft. ij in. long by 6 ft. 6 in. wide=495 sq. ft.; it has an inclination from feed to discharge of 0.5 in. per foot, is heated from two fire-places, one, b, at the discharge-end, the other, c, at one side, 30 ft. 2 in. distant from a. On both sides are doors to furnish access to the rabbles and for admission of air. The ore, fed at d, is moved down the inclined hearth to the fire-place, b, where it is discharged through port e, into an underground passage, e', containing an ore-car. The rabbling is accomplished by 21 revolving horizontal arms making from i.j to 2 r.p.m.; the arm nearest the feed, the feed-scraper/, is without teeth; the other arms each have four plow- shaped teeth, g. The arms, fastened to vertical spindles h, and h', which pass through the roof, are held in position by iron frames i supported by the side- walls. Spindles h', are water-cooled. All the spindles are rotated by bevel gears, j, meshing with bevel pinions, f, on a single line-shaft, k, extending the length of the furnace. This carries at the center a spur gear, m, meshing with pinion, n, which is rotated through bevel gears, 0, shaft, and pulley, p. The rotation of the spindles is such that each moves in the direction opposite to its immediate neighbor, causing the rabbles, which throw the ore outward, to move it in a zigzag course over the hearth, and to cover a path of from two to three times the length of the hearth. The circle of rotation of the outer end of one rabble-arm intersects that of its two neighbors; the only dead spaces are the triangles between the paths of the rabbles and the side-walls. The furnace construction requires 35,000 red brick and 3800 fire-brick. Operating requires 2 h.p. and one man per shift. The uniform rate of desulphurization of a sul- phide copper ore attained at the Yampa smeltery, Bingham Canyon, Utah, ' Simpson, Tr. Inst. Min. Met., 1903-04, xm, 27; Eng. Min. J, 1903, Lxxvi, 849. Warwick, Min. Mag., 1905, xii, I96. Metallurgist, Min. Sc. Press, 1910, ci, 5°. SMELTING OF COPPER 117 e * 1 4. 1 «tt; !■ 1 fep a J —t. I ii8 METALLURGY OF COPPER in a duplex fxirnace, as given by Waring^ and plotted by Austin,^ is shown in Fig. 12 2. Beside the furnace described, there are three other types in the market, the Duplex Stationary and the Tilting which is made either single or duplex. A DUPLEX STATIONARY FURNACE is shown in Vertical cross-section in Fig. 123, in which the parts similar to those in Figs. 115-121 have received the same designations. The hearth is 13 ft. wide and has two rows of rabbles; the latter are so distanced that there is an intersection of circles of rotation of the arms in each row, and no dead spaces exist along the center line of the furnace. The main details of the drawing are similar to those of Figs. 115-121. wi 0)( Flue mi 1 £■20* --^ Fire-box 11 DOOES 15 Fig. 122. — Desulphurization in Edwards simplex furnace. The power for the rotation of the rabbles is derived from pulley p, the shaft of which carries at the opposite end a bevel pinion, q, meshing with bevel gear r; the shaft of the latter carries a spur pinion, s, meshing with spur gear m, and this drives through spur pinion t, the second spur gear m'. With a fixed furnace the inclination, usually i in. per ft., is unchangeable; with the Tilting Furnace the angle can be changed with the character of the ore. A TILTING FURNACE consists of two Upright steel girders joined at the bottom and ends by a steel plate, and supported at the center, 2 ft. above the floor, by trunnions placed midway between the ends, while at one end there is a jack- screw for raising and lowering in order to give the hearth a slope suited to the required progress of the ore through the furnace. The iron work is connected in such a way that a furnace is readily dismantled. The results obtained at the Yampa plant are given in Table 24. Table 24. — Work of the Edwards Furnace at the Yampa Plant Simplex Duplex Hearth 6XS7ft. \ in. per tt. 2 r.p.m. I h.p. 25 tons 11X92 ft. f in. per ft. I r.p.m. 4 h.p. Inclination Power . . S in raw ore 26-39 PS"^ cent. 2 . 2-8 . 6 per cent. 4 . 3 per cent. vol. S in roasted ore SO2 in flue gases ^ Loc. cil. ' Min. Ind., 1906, xv, 253. SMELTING OF COPPER 119 T3 H S M ■«1J< a" I20 METALLURGY OF COPPER Warwick^ states that the capacity would be 15 per cent, greater if the draft were sufficiently strong. The furnaces have been replaced by MacDougall furnaces. The works of the Consolidated Arizona Smelting Co., Humboldt, Ariz.,^ have 3 duplex furnaces, 12 ft. by 91 ft. 10 in. with a total inclination of 36 in. The furnace is fired with 2 oil-burners, the coal fire-place 8 ft. by 2 ft. 4 in. serving as a combustion chamber; the furnace has 21 rabble-arms on a side, each with 6 teeth; 9 spindles are water-cooled; the 4 arms near the discharge make 4 r.p.m., the others 2. The copper ores treated are: a heavy dense pyrite crushed to 0.75 in. (Insol. 8, Fe 38, S 40 per cent.), a friable schistose pyrite (Insol. 50, Fe 15, S 18 per cent.), and a fine concentrate (Insol. 45, Fe 18, S 18 per cent.). One furnace treats in 24 hr. 50-60 tons mixture with 27 per cent. S, reducing this to 4-5 per cent., and makes 2-3 per cent, flue-dust; it consumes 250 gal. California crude oil per shift. One man and helper in an 8-hr. shift tend three furnaces; power is furnished by a 40-h.p. alternating-current motor; 50 gal. cup grease per month are used with the three furnaces and the transmission lines. EabWe Number 2726 25 24 23 22 21 20 19 18 17 16 15 14.13 12 11 10 9 3 7 6 5 1 5 2 1 19 1900 ^ 1800 1700 1600 1600 1400 1300 17 16 15 ^ 14 g 13 1: 7 6 5 4 3 2 1 / ™,.' ^ Y • ■*/ t / S \ / k. \ ^! 4. 1200 \ 1100 1000 900 800 700 COO ^ \ .- '— ili ■^ \ — '■ . y \ '' ' 50 / ^>»^ / \ ^ / "*- ■-, ^. / <-' 1 / ■^ ^ / 200 100 J = "~ _ _ Fig. 124. — Changes of temperature and sulphur-content in Edwards duplex furnace At the Goldfield Consolidated MilP a furnace 112 by 13 ft. with two rows of 27 rabbles making 2.25 and two finishing rabbles making 4.8 r.p.m. roast dead in 24 hr. 40 tons concentrate (SiOz 51.60, Fe 19.90, S 18.93, AI2O3 2.00, CaO 0.20, MgO o.io, Sb 0.08, Se and Te 0.46, Cu 0.50) containing 10 per cent, water. The elimination of S in the furnace and the temperatures observed at the rabbles are shown in Fig. 124. The S begins to burn at the seventh rabble, and burns > Loc. CU. ' Draper, F. W., Private Communication, November, 1910. ' Hutchinson, Min. Sc. Press, 1913, cvi, 171, 204. SMELTING OF COPPER 121 freely at the tenth; the loss in weight is 17 per cent. A screen analysis of the roasted ore gave >ioo-mesh— 12.5 per cent, >iso— 15.5, > 200—25.5, < 200— 46.0 per cent.; a chemical analysis showed: SiOs 54.60, FeaOs 32.20, S (sulphide) 0.15, S (sulphate) 0.75, AUO3 3.00, CaO 0.20, MgO 0.13, Cu 0.60, Se and Te 0.19, Sb 0.07 per cent. The furnace consumes per ton concentrate, 9 gal. California crude oil, and makes 1.5 per cent, flue-dust. 73. The^ Merton Furnace^ shown in front elevation and vertical longitudi- nal section in Figs. 125-126, is similar to the simplex stationary Edwards fur ^H^=^ Side Elevation Figs. 125-126. — Merton furnace. nace as regards the general mechanical features. The leading differences are: that it has three super-imposed level hearths, B1-B3, communicating by separate gas-passages, I, and drop-holes, D; that there is a separate finishing-hearth, E; that the spindles, F, 4 ft. 3 in. apart from center to center and making from 1.5 to 3 r.p.m., pass through the three hearths and are supported underneath the lowest hearth by foot-steps; that there are four rabble-arms to a hearth clamped to the ' Power, Eng. Min. J., 1903, Lxxvi, 775. Simpson, Tr. Inst. Min. Met., 1903-04, xiu, 30. Merton, Met. Chem. Eng., 1912, x, 432 (new design). U. S. Patent No. 1022961, April 9, 191 2, 122 METALLURGY OF COPPER spindles; and that the spindle and rabble-arm of the finishing-hearth are water- cooled. The furnace is 32 ft. 9 in. long over all. A roasting-hearth is 23 ft. 9 in. long, 8 ft. wide, I ft. 4I in. high at the center and 9 in. at the sides; it has four doors on either side situated opposite the spindles; the finishing side has only one door. The furnace requires 2 h.p. The following capacities for 24 hr. are claimed: Kalgoorlie sulpho-tellurides 18-25 tons, blende-galena (withZn 33, Pb 19, S 20 per cent.) reducing the S-content to 1.5 per cent., 8-10 tons; pyrite, 6-15 tons, depending upon the degree of desulphurization; sulphide copper ore up to ro tons; arsenopyrite, 5-8 tons; and galena 8-20 tons. 74. The Ropp Furnace.* — The furnace represented in Figs. 1 27-131, is a single-hearth, straight-line mechanical furnace, usually 150 by 14 ft., with four exterior fire-places, h. *Along the center of the hearth, a, runs a channel, c, 1.5 in. wide, through which extend vertically at equal distances heavy trapezoidal steel-plate arms, d, attached to four-wheel trucks, e, each carrying a rake of the width of the hearth with teeth set at an angle of 45 deg. Table 25. — The Ropp Straight-line Furnace Length of hearth, feet 105 100 150 Width of hearth, feet 11 14 Hearth area (deducting slot), sq. ft 1,120 1,367 Number of fireplaces 3 3 Length of single grate, feet 4J 5 5 Width of single grate, feet 4 4 4 Grate area, sq. ft 18 20 20 Ratio hearth to grate area 64 i 70 : i 105 : i Number of carriages 4 4 to 6 6 Number of blades on carriage 26 Carriage makes circuit in (minutes) 3 J Ore stirred every (seconds) 52 53 53 Horse-power required 4 to 5 s 6 to 8 Depth of charge near fluebridge f l^''^^' !"'^"= ' ' ' ^ ^° ^4 ^ *" ^i ^ t° =* \ Matte, inches... i| ij ij Time ore remains in furnace (hours) 6 to 8 6 to 8 6 to 8 I Pyrite 26 to 42 35 to 40 45 to 7a Tons in 24 hours Matte 26 to 34 32 to 45 45 to 60 i Dry gold ore 80 to 98 Pyrite.., o. i to 8 o.ito8 o. i to 8 Matte... I to 8 t to 8 i to 8 Dry gold 0.14 to 0.16 0.14 to 0.16 0.14 to 0.16 ore Number of men in 12-hour shift i to J i to ^ 1 to ^ Tons of coal per ton of ore ito ^ij t'o to J^ /i to t'j Gallons lubricating oil per ton of ore 15 14 j , ' Built by Harron, Rickard and McCone, San Francisco, Cal. ^ Character, composition and size of ore : Pyrite: SiO^ 18-30, Fe 20-30, Cu 0-6, Pb 0-20, S 20-30, Zn o-io, per cent.; 30-mesh 14 2,050 4 32 32 3i SiV Per cent, sulphur in pulverulent roasted product,'' to Matte: Fe 20-40, Cu 0-50, Pb 0-18, S 18-30, Zn 0-4, per cent.; 20-mesh to | in. Dry gold ore; S 1-2 per cent., 30-mesh and finer. Missing Page SMELTIXG OF COPPER 123 There are six rakes working in pairs. They enter the hearth at h, where ore is introduced by the mechanical feeder, i, carry it slowly to the discharge, ;, where it falls into trucks, k. The steel traction rope, which does not become hotter than the hand can bear, lasts about 1.5 years, and the rakes are thoroughly cooled on their return trips. The furnace has working doors on the sides, which permit access to the hearth and admit air. The heat is regulated from the different fire-places; these have perforated fire-bridges to insure complete combustion with air preheated in the side-walls. The products of combustion pass off through the smoke flue, w. The furnace is built in two sections to allow for expansion and contraction; the sections are joined by bars of channel iron, which overlap and glide one within the other; they are tied independently, and the length of a longitudinal tie-rod does not exceed 50 ft. Table 25 gives some working details. 75. The Wethey Straight-line Single-hearth Furnace.— This is shown in plan, side- and end-elevations in Figs. 132-134. It covers a floor-space of 115 ft. 3^ in. by 16 ft. 5 in.; has a roasting-hearth 100 by 10 ft.; is fired from two pairs of exterior fire-places, A A', the two forming a pair being placed opposite one an- other; the first pair, ^, is 10 ft. 8 in. away from the feed-end; the second, A', 22 ft. 8 in. distant from the first. The flame travels with the ore, enters a down- take, B, returns underneath the hearth, and passes near the feed into an under- ground flue, C, ending in a dust chamber connected with the stack. The moving and raking of the ore is accomplished by six rabble-arms, D, each with 20 exchangeable teeth, E. The arms, which form part of the carriages, travel on one set of outside rails, F, while the ore is being rabbled, and return on another, G, placed above the furnace. The continuous slot on either side is closed by tripping doors, H, through which enters most of the air necessary for oxidation. The carriages are connected and set in motion by two endless steel-link chains, 7, passing over sprocket-wheels, J, at the ends of the furnace. The chains are kept taut by 2000-lb. weights attached to gliding journal-boxes of the upper sprocket-wheels; and the latter are prevented from sliding back by ratchet and pawl, L. The shaft, M, of the lower sprocket-wheel opposite the feed-end is connected by gearing, N, with the power furnished through belt and pulley, 0. The concentrate arriving in a mine car is emptied into two feed-hoppers, P, whence it passes through a curved chute-like feeder, Q, driven from a sprocket- wheel, into the path of the descending teeth of a rabble-arm to be carried into the furnace. This has two balanced inlet-doors, R; at the opposite end are simi- lar outlet-doors, 5. The roasted ore is discharged into suitable ore-cars, T. The roof of the furnace is supported by /-beam skewbacks, U, which are sus- pended by steel straps, V, from cross-beams, 11', and braced by struts, Y, firmly connected with the vertical /-beams, Z; the latter are tied by rods, a, and spaced by distance pieces, b, which in their turn are let into the lower furnace-wall. The iron frame carries the supports, c, of the upper and lower tracks for the rab- ble-carriages. The furnace with its 1000 sq. ft. hearth area puts through in 24 hr. 70 tons sulphide copper ore, ranging in size from J-in. to loo-mesh, and having the following composition, SiOj 38, Fe 31, Cu 7.7, Zn 6.0, S 6.8; or 140 lb. per 124 METALLURGY OF COPPER square foot hearth area; the depth of the ore-bed near the feed is about 4 in. A rake makes the circuit in 13.5 min. The ore, remaining in the furnace about 4 hr., is stirred every 2 min.; the roasted ore retains 6.8 per cent. S. The furnace requires 15 h.p., and consumes 7000 lb. coal in 24 hr.; one man attends two furnaces in an 8-hr. shift. The cost of roasting in Butte was 34 cents per ton of ore. H. F. Brown^ has constructed a similar furnace. 76. The Allen-O'Hara Furnace.^ — O'Hara' was the first to construct a straight-line mechanical roasting furnace with two superimposed hearths. The difficulties encountered at first by wear of chains, rabbles, and hearth have been overcome by Allen and by Brown. The Allen-O'Hara furnace, shown in Figs. 135-138, has two hearths, B and D, 90 by 9 ft., and 5 pair of exterior fire-places, Q, two on the upper and three on the lower hearth, three being in use at one time. The grates are 5 ft. by 2 ft. 4 in. Each hearth has a pair of rails, L' , 3 ft. 5 in. apart, to carry six trucks, K, supporting the rabble-arms, each with 14 teeth of steel plate, to which are riveted white-iron shoes. Extra rabbles prevent the accumulation of ore on the rails. The trucks are connected by a weldless-link chain driven by sprocket-wheel, F, through shaft G, and kept taut by weight /, acting upon the gliding box T, of sheave H. Carriages, M, placed midway between the rabble-arms, prevent the sagging of the chain and permit exchanging it in parts when damaged. The backs of the carriages carry swinging double-bladed scrapers which prevent the ore forming a ridge along the center-line of the hearth not covered by the teeth of the rabble-arms. The sprocket-wheel, H, is placed at a distance of 23 ft. from the end of the furnace to allow the chain and rabbles to cool. Beneath wheel F, is a grating, P, for repairing of chain and arms. The ends of the furnace are closed by hori- zontal turnstile doors, N, which are moved by the carriages. The ore fed me- chanically at ^ , is moved over the upper hearth, drops through port C, at the opposite end on to the lower hearth where it travels in the opposite direction and is discharged into the brick hopper, E. The air necessary for oxidation enters through dampers in the side-doors, R'. At Butte, Mont., from 45 to 50 tons concentrate, ranging in size from i| in. to slime and averaging SiOa 23, Cu 14.6, Fe 23, Zn 2, S 37.4 per cent, were roasted in 24 hr., or 60 lb. per square foot hearth area. The ore lies about 3 in. deep near the feed, is stirred every 37 sec, with the carriages making the circuit in 3.75 min. It remains about 6 hr. in the furnace, and is discharged with a S-content of 8 per cent. ; 100 lb. moist concentrate give 72.7 lb. roasted ore and i lb. flue-dust. A partial . • Hofman, Min. Ind., 1897, vi, 450. Strauss, Min. Mag., 191 1, v, 61, Casapalca smelter, Peru. " Hofman, Min. Ind., 1893, 11, 432. Wythe, Eng. Min. J., 1893, LV, 460. Kroupa, Oest. Jahrb., 1894, XLii, 328. Peters, Min. Ind., 1894, in, 203. Hofman, Tr. A.I. M. E., 1904, xxxiv, 272. ' Kustel, G., " Roasting Gold and Silver Ores," A. J. Leary, San Francisco, Cal., 1880, p. 84. SMELTING OF COPPER I2S •S > K 6 o 126 METALLURGY OF COPPER analysis of an average sample of roasted ore gave, Cu 10.8, Si02 31.6, Fe 38.0, S 7.6 per cent., and one of flue-dust, Cu 10.8, Si02 34-4, Fe 36.5, S 8.2 per cent. The furnace requires 3.64 h.p.; 100 lb. coal are consumed per ton of ore; the labor needed in an 8-hr. shift is J car-man who brings the raw ore, J fire-man, i wheeler who removes the roasted ore, and | repair-man during the day-shift only. The rabble-arms last about i year, the teeth 6 weeks, the flange-wheels 6 months, the chain i year. The cost of roasting at Butte in 1904 was $0.78 per ton. The Brown-O'Hara furnace is referred to below. 77. Wethey and Keller Multiple-hearth Furnaces. — The Wethey Mul- tiple Hearth Straight-line Furnace, shown in vertical cross-section in Fig. 139, is a double furnace with four superimposed hearths, 50 by 5 ft., which have slots alternately at opposite ends in order that the ore fed near one end of the top hearth, when moved to the other by mechanically driven rakes, may drop on the second hearth and so on until it reaches the discharge on the bottom hearth. The furnace is built in an iron framework as is the single-hearth furnace given in Figs. 132-134. The framework consists of vertical channel-beams, a, joined by riveted horizontal castings (not shown) which form the supports for horizontal I-beams, b, carrying the three upper hearths. Channel-beams a, are braced by tie-rods c. The brick bottoms of the upper hearths are laid on steel plates, d, which rest on 3-in. I-beams, e. Along the inner sides of each hearth run longitudinally two 6-in. I-beams, /and g; the lower ones, /, rest on channel-beams, h, and carry the tracks as well as the angle-irons against which are built the inner side-walls; the upper beams, g, are suspended from channel -beams, h, by castings, h, and form the inner skewback of the roof; the outer skewback, i, is backed by vertical channel- beams, a. The three upper hearths are supported independently of one another. The ore is raked and moved by rabble-arm sj, which are attached to two end- less link-belt chains passing over two pairs of sprocket-wheels (not shown). The 2-in. slots in which the rabble-arms travel are closed by tripping doors. The fire-boxes are placed on the levels of the second and fourth hearths, on which the fire-gases enter the furnaces to travel in a direction opposite to that of the ore. The double furnace with eight hearths, 50 by 5 ft., has 2000 sq. ft. hearth area; there are four fire-places with grates 38 by 33 in., which gives a grate area of about 80 sq. ft., or a ratio of hearth to grate area of 25 : i. There are eight rabble- arms to a furnace, each with 14 teeth; four arms make the circuit of two floors in 185 sec, hence the ore is stirred once in 46 sec. The concentrate treated at Butte ranged in size from 12- to 4-mesh, and assayed Si02 10, Fe 35, Cu 10, Zn 5, S 40 per cent. The depth of charge near the feed is about 4 in.; the ore remained in the furnace from 8 to 10 hr., and was dis- charged with about 8 per cent. S. The furnace put through in 24 hr. 45 tons concentrate, consumed no lb. slack coal and o.i gal. black lubricating oil per ton of ore, required i man per shift, and made 0.5 per cent, flue-dust. SMELTIXG OF COPPER 127 128 METALLURGY OF COPPER The four-hearth furnaces built after 1894 were single. They combined some of the features of the single-hearth furnace shown in Figs. 132-134 and the four- hearth furnace just described. The hearths were 65 by 10 ft.; the rabble- arms protruded on both sides and were supported by exterior carriages. The fire- boxes were raised so that the flame entered the top hearth to dry and kindle the ore, the fire-gases then traveled downward with the roasting ore. The furnace with a roasting hearth of 2600 sq. ft. treated in 24 hr. 90 tons of ore with 35 per cent. S or 70 lb. per square foot hearth area, reduced the S-content to 5-6 per cent., consimied 80 lb. slack coal per ton of ore, and required one attendant on a shift. The Keller Automatic Roaster.' — This is an improved form of Spence furnace (p. 90) -which was in operation at the Parrot works, Butte, under the name of Keller-Cole-Gaylord for roasting sulphide copper ore, and under that of Keller Automatic Roaster at the Germania lead works. Salt Lake City, for roasting a lead-bearing pyritic concentrate. Fig. 140. — Keller automatic roasting furnace. It consists of two five-hearth kilns with the machinery for rabbUng placed between. Fig. 140 gives a longitudinal vertical section through one of the kilns. It shows five arched hearths, 21 ft. 5 in. long by 6 ft. 3 in. wide, which are covered with siliceous taihngs or ashes to form a level working bottom. At one end on the ground floor is situated the fire-place, from which the flame rises to be deflected by the roof into the top roasting-hearth to ignite the ore. The roof serves also as a dryer. The ore fed mechanically through the hopper, a, is moved over the hearths and discharged into a car in the passage, q. The sulphurous gases travel in a direction opposite to that of the ore, and meet the products of combustion from the fire-place on the top hearth; both then travel 1 Peters, "Modern Copper Smelting,'' 1895, p. 214. Hofman, "Lead," 1898, p. 191. SMELTING OF COPPER 129 through flues 6 and c and an underground passage d to the stack. They thus envelop the roasting part of the furnace by hot gases and thereby diminish the loss of heat by radiation. The rabbling apparatus, which required from 2 to 3 h.p., is complicated and probably has become obsolete as have most rabbling- devices which move to and fro over a hearth. The furnace treated in 24 hr. 45 tons of Butte concentrate, reducing the S- content from 40 to 7 per cent., consumed 1.25 tons of slack coal, and required I man per shift. At Salt Lake City it roasted in 24 hr. 35 tons of wet pyritic concentrate containing some Cu, 8 per cent. Pb and 30 per cent. S, reduced the the S-content to about 3 per cent. ; and consumed 2 tons of slack coal. 78. The Brown Horseshoe Furnace.' — The late H. F. Brown constructed three types of roasting furnaces, all of which have two patented improvements, Fig. 142, a recessed chamber on either side of the hearth for the reception of the mechanism of the rabble-carriages, and a continuous slot for the passage of the rabble-arms from the chambers to the hearth. His three types of furnaces are the Horseshoe, discussed below; the Elliptical,^ in operation at one time at the Golden Reward works, Deadwood, S. D., for roasting silver-gold ores; and the Straight-line,' which is similar to the Wethey furnace shown in Figs. 132-134. In the Horseshoe furnace, represented by Figs. 141-148, the roasting hearth. Fig. 141, occupies five-sixths of the annular space, the remaining one- sixth forms an open span, I, for the feed and discharge of ore, and for the cooling of the rabbles. The hearth, m, 135 by 8 ft., has a roasting surface of 1080 sq. ft. which is heated from three exterior fire-places with grates 3 by 5 ft., giving a ratio of hearth to grate area of 24 : i. The first fire-box is placed about 60 ft. from the feed-end of the furnace; the other two are 35 and 40 ft. apart. The flame from a fire-box enters the furnace through the roof, q, Fig. 143; it spreads uniformly under the latter and does not strike the ore. Ore and flame travel in opposite directions. The stack, /, carries off the gases. In the center of the furnace, Fig. 141, is an engine, a, varying from 5 to 8 h.p. (only two are required) which drives the mechanical feeder, b, and the endless cable, c, running over sheaves, d. Opposite each sheave is a door, e, which gives access for oiling and repairing, and admits air for cooling the bearings and for roasting the ore. The doors, /, on the outer side of the furnace are far enough apart to furnish access to the hearth when the outer wall, g, between the doors and the tiling, h, has been taken out. The rabble-carriage, Figs. 147-148, consists of an L-shaped frame with one flat-tread wheel, x, running in the outer, and two double-flanged wheels, y, in the inner passages, Figs. 142, 143. The front inner wheel. Fig. 148, has the cable- grip, z. The tooth or point, t, Fig. 147, is plow-shaped, of cast iron, and is keyed to the cast-steel rabble-arm, u, Fig. 146, which is attached to the stirrer- arm, v, reaching across the hearth. Arms, m, are lowered as the plows wear off; ' Allis-Chalmers Co., Milwaukee, Wis. ' Brown, Eng. Min. J., 1896, LXii, 80. Rothwell, Min. Ind., 1896, v, 270. ' Hofman, Min. Ind., 1897, vi, 459. 9 130 METALLURGY OF COPPER when a crust forms on the working hearth, they are lowered one at a time to cut underneath. The arms are placed about 12 in. apart and are mismatche'Q so as to cover the whole surface of the ore. One carriage, Fig. 147, has side-stirrers, w, to prevent the accumulation of ore. There are two carriages to a furnace which are alternately attached to and released from the continuously traveling cable, c. The ore is delivered at b, Fig. 141, by an automatic feed, which dis- charges with every passage of a car a weighed amount of ore on to the hearth outside of the swinging door, j. The ore then enteres the furnace, travels over the hearth in from 12 to 14 hr., leaves it by the other door, k, and is discharged into wheelbarrows or trucks placed at I. A heavy sulphide ore ignites at a dis- tance of about 1 5 ft. from the flue-end or in about i hr. after it has been fed ; an ore running low in S will be brought to the kindling temperature in from 2 to 2J hr. One traveling stirring-carriage (not shown), after leaving the furnace at k, strikes the other carriage, m, at rest in the open space, h, and pushes it along until its grip-lever comes in contact with a fixed stop and is forced down, when it grips the cable which draws the carriage into the furnace. At the same time the heated carriage is automatically released, comes to a stop in the place just occu- pied by carriage, m, and cools in the open space for i to 3 min. In this manner the two carriages are used alternately at intervals of i to 3 min. and do not become sufficiently heated to be much afifected by their passages through the furnace. Fig. 142 gives some details of construction. The heavy cast-iron frames, e and /, of the inner and outer doors, carry 9-in. channel-beams, n, which are backed by buckstays, 0, secured in their turn by tie-rods, p. Thus the roof is supported independently of the brickwork between the doors, which can betaken out to furnish access to the interior. The sheaves, d, are supported in the doors outside of the inner recess-chamber and thus remain cool. The tiUng, h, is set in the floor and is easily removed when occasion calls for it. The tiling, s, forms part of the roof, and special care must be had to secure it firmly. The inner passage shows a T-rail for the inner double-flanged wheels of the stirring carriage; in the outer passage is a smooth rail for the outer flat-tread wheel giving room for expansion and contraction. From i to 2 per cent, flue-dust is ordinarily made. The furnace, as seen by Table 26, has been used for roasting different kinds of ore. The advantages claimed for the recessed chambers as regards the pro- tection of the stirring mechanism have been much exaggerated. The leading disadvantage ^ lies in the fact that the carriage in being moved over the hearth is pulled on the inner side only. The result is that the ore on the outer side is not as well stirred as that on the inner because the plows pass over obstructions that are forming on the working bottom instead of plowing through them; and, secondly, in consequence of this the carriage frequently jumps the track and creates havoc in the furnace. 1 Testimony of W. H. Smyth in case H. F. Brown vs. Metallic Extraction Co. U. S. Circuit Court of Appeals, Eighth Circuit, pp. 453, 457, 459, 461. Missing Page SMELTING OF COPPER 131 ■T3 .-« 1 I2IS Tf u- 1 PC 10 [ 1^ 1 M > 00 1 M 5 M 00 cn in ■M H (U (3 c e g a '- 9 1) g H 2 .; S < Oi 10 H H 6 HI u [^ Tj -^ in U ■> ^ 1/ ^ I5 ;3 =3 c « 1/ > P) H \^ as %1 f ■> ■^ h 1/ T 00 1^ ^1 Is Ml ■i vc <: 12 ^ 2 C al .^ C/] n (1) Td 3 "3 M « G 1/ •i 1/ 1 li- 1 ■> w ■i C) H 1^ ^ 2 -i CM cs c h ly 7 'C vC Ph +j 1§ a. ^2 ■> " ■) in M 1 vd s 5 in (L) 1 oc c < -> If < 0. 1 & a h ^T +J rt 3 1 u ■) ^ IS 1 c t- ^ c ■i I' 1 1 1 u 1 ^ 00 cs C M tn C 1- ■^ t. c C vC c tI (N ^ f T \- ^ ■5 1- cs i a ^ > ^ > \ a C U C u Q L. -t- ,r X. X. T I U 01 a •g % '% p ^ * , OJ OJ u g 3 a in tn "7/ c/ j3 U <6 d C 3 3 2 01 \ c a c c L rj (U > PL, ^ '^ u CA u; Cfi c/ 5 .0 3 > R o u a c8 ■3 g 43 Ph 3 ** aj . Pi '3 • a 132 METALLURGY OF COPPER 79. The Pearce Turret Furnace. ' — This is the second circular or turret- shaped mechanical reverberatory roasting furnace, with a space left open for the feed and discharge of ore, which has been used to a considerable extent in the roasting of sulphide copper ore. It is built with one, two, and six hearths. The single-hearth furnace is represented in plane and vertical section in Figs. 149-150. These show a circular hearth, 6 ft. § in. wide (area 505 sq. ft.),^ heated from three exterior fire-places, k, the fire-gases traveling in a direc- tion opposite to that, of the ore, and passing off through the down-take, c. The outer wall has a number of doors. Fig. 152, usually closed, which furnish access to the hearth; between each pair of doors and 2.5 in. above the level of the hearth, Figs. 149 and 152, is a horizontal flue, e, 8 in. wide, through which air, warmed in its passage, can be admitted at the surface of the roasting ore, provided the draft in the furnace is sufficiently strong. The inner wall is divided by the rabble-arm slot,/, into two parts. The lower one, a', is erected in the usual way; the upper one is suspended by stirrups, g, in angle-irons, V, from radial I-beams, h, and cross-beams, i, and is braced by radial struts, c". The I-beams are supported by the outer wall and the central column, /, and the cross-beams by the I-beams. The stirring mechanism turns over and moves the ore, fed automatically through the roof, and discharges it on the side opposite the feed. It consists of a pair of horizontal pipe-arms, n, which are placed diagonally in a hollow hub, 0, revolving around the stationary hollow central cast-iron column, j, and are supported at r by rollers traveling on track z. A rabble-arm is made up of two pieces of 5-in. gas-pipe joined at r; the outer part reaching into the furnace is provided with the necessary rabble-blades, m. The blades, of J-in. steel, are graduated in length and direction from the inner to the outer circle of the hearth so that the ore on the outer circle, which has to travel the greater distance, shall remain in line with that of the inner circle. The continuous slot, /, in the inner wall is closed by a 12-in. band of sheet-steel, which is carried by the two rabble- arms and two reserve-arms, and is pressed against the slot by weighted bell- cranks, u, attached to a circular angle-iron, z, also carried by the arms. The four arms are braced by rods, q. The cast-iron column, j, is connected through f^ pipe, b', with a blower. The air passing through b' and j, enters hub, 0, * acting as an air-box, and thence into the rabble-arms keeping these cool,i and is slightly warmed and delivered (between the blades) into the furnace through small pipes which point downward, and through horizontal holes in 1 the arms situated between blades and pipes. I The rabble-arms receive their motion through spur-wheels, s, centered by \ rollers, t, which are in gear with pinions, c', on pinion-shafts, v; the latter are driven by bevel-gearing, w, and main shaft, x. The arms travel at an average! speed of 75 ft. per min., taking the circumference of the center of the hearth as the line of measurement. This makes 53 sec. the time of a complete circuit with a 6-ft. hearth, and causes the ore to be rabbled once in 26.5 sec. • Stearns-Roger Mfg. Co., Benver, Colo. 2 Furnaces with hearths 7 ft. wide (area 609 sq. ft.) and 8 ft. wide (area 788 sq. ft.) have been built. Missing Page Missing Page SMELTING OF COPPER 133 The lower I-beams, y, Fig. 151, serve as inner braces of the furnaces; they hold the boxes of the vertical pinion shafts, and carry the circular track, z, which bears the weight of the rabble-arms. The ore is fed mechanically from the hopper, d', by a device which is con- trolled through shaft, e', connected by pulleys, /'and /", and belting, g', with the main driving shaft, x. Thus the rate of feed is regulated by the speed of the rabble-arms. A rabble-arm, after leaving the discharge-opening, travels over the open space and arrives at the uncovered part of the hearth which receives the ore from the feeder. It spreads out this ore and enters the covered hearth by the swing-door, i', Fig. 152. As at first the ore is only dried and brought to the kindling temperature, any air admitted at this period would retard the process. It is therefore automatically controlled by a butterfly-valve, 7, in the arm, which closes as it strikes the first stop, k', and opens again when it reaches the second stop, m'. The arm leaves the covered hearth by the second swing-door, i", Fig. 152, carrying with it some roasted ore which drops through the discharge- hopper n' into a truck, the fumes being carried off by the sheet-iron stack 0'. The ore remains on the hearth about 6 hr. The rabble-teeth last from 4 to 6 weeks with pyritic ore assaying about 40 per cent. S; the outer part of a rabble-arm which is exposed to heat and fumes about 12 months. The furnace requires 2 h.p. Temperature-measurements have been published by Clevenger.^ The cost of roasting in Butte is about $1.00 per ton. Some work done by this furnace is given in Table 27. The Two-hearth Furnace. — The "double-deck" furnace, with its two superimposed hearths separated by an air-space of 2 ft., has a height of 16 ft. 6 in. against 11 ft. of the single-deck, and twice the hearth area. Each hearth has the two usual rabble-arms. A horizontal section through the upper hearth is shown in Fig. 152 and a vertical section through the entire furnace in Fig. 151. The details are nearly the same as those of the single-hearth furnace; hence the different parts common to Figs. 149, 150, 151, 152, are designated with the same letters. The lower part of the inner wall of the upper hearth rests on I-beams, h, while the upper parts of both hearths are suspended and held in place as with the single-hearth. furnace. Underneath the lower hearth is a dust chamber through which passes the driving-shaft of the lower rabble-arms; it is enclosed by a s-in. pipe for protection against sulphurous gases. The upper driving-shaft, directly connected with the main pulley, transmits the power to the lower shaft by means of bevel-gearing and a vertical shaft. The furnace requires 5 h.p. It has two upper and two lower fire-places. Fig. 151 shows a step-grate, k, which is used with non-caking coal or lignite; there is also a curtain arch, m", over which the flame travels for a short distance, and is thus prevented from overheating the roasting ore. The ore is fed on to the upper hearth as in the single-hearth furnace; it travels in the direction of the arrow over the upper hearth, drops through slot, ' Afet. Chcm. Eiig., 1913, xi, 449- 134 METALLURGY OF COPPER < a i { Ph 1 c G i &4 c a * 0^ 1 « o o o 00 o oo o oo o 00 o O M o o (*3 £ -3 3 S S u " o VO o o in 1 o O < M U ft 8 S, =• 00 00 00 - „ 5 o 1^ s-S s " S s r. M - r- H H g 1 -2 -; ■§ r -S ■^ p. o o 5 o o o 5 o 00 to 3 s a 3 3 z n n 1 o 2: 1 00 1 o 00 8 ct < § .S J „ £ 5 * -S s 2 o & lO O i s XI tin W g h O g t S In a E bo • S I 1 1 Ph E o O c 1 C ■ 1/ t- c c D. c c 3 r- c C ■J 3 O c N o* 0) oo" o I cu . . o fp 1 ^ O « N en o i C N oo" id o o W ■) (S cr c p- ro CO ci! d 'Z > p. 4 > p. 3 0.5 in., 3.3 per cent.; < 0.5 and > lo-mesh, 9.6; 10- to 40-mesh, 30.9; 40 to 80-mesh, 25.5; 80- to 120-mesh, 14.0; 120- to 200-mesh 17.8; < 200 mesh, 8.9 per cent. In starting, the grate was covered with ashes to a depth of 0.75 in., a fire was kindled at the center with the blast at 2 oz. pressure, then saw-dust was fed until about 6 in. deep at the center and tapering toward the periphery, and lastly 2 tons of warm (80° C.) flue-dust were charged to be followed by a 3-in. layer of concentrate. After 30 min. blowing, the concentrate was well kindled, the blast was turned on fully giving a pressure of about 18 oz., and the rest of the concentrate fed in as fast as the heat crept up. The time of a blow was 20 hr.; a charge furnished 93 per cent, coarse material with Si02 17.6, Fe 34.5, AI2O3 5.8, CaO 0.8, MgO 0.6, S 9.4, Cu 21.4 per cent. Metallurgically this work was satisfactory, as was that carried out later in a Dwight-Lloyd straight-line machine; but blast-roasting with blast-furnace smelting has had to yield to roasting in MacDougall furnaces and smelting in oil-fired reverberatories. Neill records^ the blast-roasting experiments he carried out at Mine La Motte, Mo., in 1883 with nickel-bearing pyrlte, and his later work in Bingham canon in 1894 or 1895 with sulphide copper ore from the Commercial mine; he illus- trates the plant he built for the Yampa smelter, Utah, in 1903. The first experiments with blast-roastir^g are probably those recorded by Roswag,^ which were made in Spain. Collins* gives some facts about blast-roasting in a Spanish plant. The ore treated is a concentrate with from 15 to 18 per cent. Cu, and a mine fine with from 8 to 15 per cent. Cu. The mixture averages 15 per cent. S and receives from s to 6 per cent, fine iron ore; if it is sufficiently coarse, containing much 0.25-in. material, it is blown direct; if a large portion is finer than 8-mesh, about 3 per cent, slag is added to the ore. Fig. 156 gives a vertical section of a Collins blast-roasting pot. The pot. A, is of mild steel and is enclosed by an open jacket filled with water to prevent overheating; the bottom is closed by a cast-steel wind-box, E, and a conical cast-iron grate, T, having 123 conical holes, f and f in. in diameter. The pot is supported by trunnions, C, on uprights, D, forming part of the carriage, B, ' Hofman, loc. cil. 2 Tr. A. I. M. E., 1910, XLi, 915. , ' Fr^my "Encyclop^die Chimique" Dunod, Paris, 1884, section I, vol. v, p. 243, and sec- tion II; Eng. Min. J., 1902, xcii, 750. * Min. Mag., 19O9, i, S 2. SMELTING OF COPPER 141 and is turned by worm and hand-wheel. The wind-box, E, is connected by a 6-in. alum-soaked canvas hose with an underground main. The gases are con- ducted away through a hood connected by a 3-ft. pipe with a dust chamber built of brick soaked in tar. A pot holds from 3400 to 3600 lb. charge. In starting, glowing coals are introduced, a slight blast turned on, some fresh coal added, and about 100 lb. mixture charged. In about 10 min., when this primer has begun to glow, the pot is disconnected, moved to a loading platform, charged and returned to its place, when a blast with from 7 to 8 oz. pressure is turned on; this is then increased to 12-13 oz. A coarse charge requires from 6 5'9~ Fig. 156. — Collins blast-roasting pot. to 7 hr. blowing, a fine charge from 9 to 10 hr. The sintered product, ordinarily from 90 to 95 per cent, of the whole, assays: Insol. 56.2, FeO 12.6, MnO 0.2, CaO 6.S, MgO 1.2, AI2O3 7.1, Cu 7.7, S 2.0, SO3 1.4, total S 2.6. At the Wallaroo Works, South Australia, ' blast-roasting in spherical pots, provided with hoods, mounted on trunnions and tipped by worm and hand- Svheel, is in successful operation with ore, low-, and high-grade matte. The pot is 8 ft. 8 in. in diameter, 4 ft. 6 in. deep, has steep sides (14° from vertical), and holds from 8 to 9 tons of charge; the grate, perforated by f-in. holes, is 10 in. above the bottom; the blast-pipe is 8 in. in diameter. The success depends to a large extent upon keeping down the temperature by the addition of abun- dant water; thus, ore requires 6-9, ordinary matte 4-6, rich matte 3-4 per cent. ' Cloud, Tr. Insl. Min. Met., 1906-07, xvi, 311. Austin, Min. Ind., 1907, xvi, 335. RadclifiE, Min. Set., 1909, lix, 67. 142 METALLURGY OF COPPER water. The tendency to reach an excessively high temperature is due to the combustion of Fe, causing premature fusion and thereby stoppage of roast. An ore-charge — made up of \ coarse (i-ii in. in diameter) and f fine material- assays Si02 15-35, S 15-25, Cu 10 per cent.; it is blown in from 8 to 12 hr., coarse working more quickly than fine. In operating the grate is protected by roasted pieces, 3-4 in. in size; the fire is started with wood and saw-dust and a slight blast. When the fire has extended from the center to the periphery, the pot is half filled with charge, the latter being piled up on the side where the blast tends to rise quickly. The blast is now increased, and 0.5-in. holes are punched in the surface. In i hr. the fire becomes visible around the center, when full blast (1000 cu. ft. per minute at 13 oz. pressure) is given. The blown cake, weighing about 6 tons, is dumped, lifted by a crane and dropped on to cast-iron cones. The cake retains 5-6 per cent. S. The converter-gases con- tain SO2 5.4, O 8 per cent. vol. For blowing ordinary matte, this is crushed to §-f in., mixed with 15-25 per cent, siliceous ore, wetted, and charged into the pot gradually so as to cover the fire, as this appears on the surface until the pot is filled. The blast-pressure is 6-8 oz.; the issuing gas contains SO2 5.5, O 12 per cent. vol. Rich matte (white metal) is crushed to f -in. size, wetted, and blown without additions in pots, 5 ft. 3 in. in diameter at top, 11. 5 in. at bottom and 3 ft. 6.5 in. deep, holding 2-ton charges. When the fire of wood and saw-dust has been started, lumps of roasted matte are introduced and brought to a red heat; the pot is half filled with matte and blown with a pressure of 6 oz., then completely filled. After the fire has appeared on the surface, the charge is blown with a pressure of 8 to 10 oz., a blow lasts 8-10 hr. The dumped cake, which consists of a mixture of sulphide, oxide, and metal, is broken by an iron ball weighing 1600 lb. and dropping 20 ft. The converter-gases average SO2 11. i, O 5.6 per cent. vol. As to the~ elimination of impurities in treating white metal, blast-roasting gives 1.87 per cent. Ni and 1.39 per cent. Bi; when the Direct or Nichols- James process (§ 153) shows 1.5 per cent. Ni and 1.12. per cent. Bi; and roast-smelting (§ 153) I per cent. Ni and i per cent. Bi. At Blagodatny^ matte is blast-roasted in a conical tilting pot, 3.28 and 4.01 ft. in diameter and 4.687 ft. in height, with central perforated air-delivery pipe extending to near the top, in charges of 3600 lb. (5.66 matte : i siliceous concen- trator tailing) in from 30 to 35 hr.; the product retains S 12-13 per cent. At the Casapalca smeltery, Peru,^ ore with Si02 25, Fe 15.6, AI2O33.8, CaO 9, Zn II. I, S 24.4, Pb 6, Cu 3.5 per cent., and Ag 44 oz. per ton, is crushed to pass a 3-in. screen, mixed with siliceous material and lime, moistened with 5 per cent, water, and blown in a Savelsberg pot. This is 8 ft. 8 in. at the mouth, where it is 2 in. thick, and tapers to ij in. at the bottom; the height from grate to rim is 3 ft. 6 in., that from rim to center of trunnion 2 ft. ( = 2 in. below center of gravity of charged pot). The inside of the pot has four rings, | in. wide, to counteract ' Metall-Erz., 1913, x, 612. 2 Strauss, Min. Mag., 1911, v, 59. SMELTING OF COPPER 143 creeping up of blast. The grate, i| in. thick, is arched to a radius of 11 ft.; it consists of a center and two side-pieces, stands 14 in. above the bottom, and rests upon shoulders, i in. wide, bolted to the outside of the pot. The air-holes, I in. in diameter at inlet and f in. at outlet, are 2\ in. apart. The blast-pipe is 6 in. in diameter; the conical hood has one feed and seven working doors; it enters a telescope stack 30 in. in diameter, and is raised and lowered by a hand-wheel. The pot is tilted by a worm gear, the carriage is held by a wire rope on its slightly inclined track. A full charge weighs 9 tons. The fire is started with shavings and wood to a depth of 3-4 in., blast is turned on, 100 lb. coke breeze ( < | in.) are charged, a layer of limestone is given, and 1200 lb. ore are shoveled in in three charges, each being spread with a rake; the blast is correspondingly increased. The rest of the charge is dropped through a chute from an overhead hopper, in lots of 800 lb. and more at a time; the sur- face of the charge is kept open by punching. When the full charge has been given, the blast shows a pressure of 5.5 oz., and rises to 9 and 10 oz. toward the end of the ii-hr. blow. The desulphurization reaches 75 per cent., the loss in weight is 10 per cent.; little flue-dust is made. A sample of sinter-cake gave Si02 28.6, Fe 15.8, AI2O3 3.9, CaO 10.6, Zn 10.8, Cu 4.1, S 6.1, Pb 5.4 per cent, Ag 48 oz. per ton. 1^ stuck Feed Kegulating Gate Water Tr~^ Fio. 157. — Dwight-Lloyd straight-line roasting machine. 84. Examples of Down-draft Blast-roasting.— The leading apparatus is the Dwight-Lloyd straight-line roasting machine,^ shown in Fig. 157. It consists of a frame of structural steel supporting a feeding-hopper, an igniting furnace, a suction-box, and a pair of endless-track circuits to accommodate a train of small truck-like elements called pallets which, in combination,^ form practically an endless conveyor, with the continuity broken at one place in the 'Hofman, Tr. A. I. M. E., igiOjXLi, 759, "General Metallurgy," 1913, p. 43°- 144 METALLURGY OF COPPER circuit. Each pallet is provided with four wheels, which engage with the tracks or guides at all parts of the circuit, except when the pallet is passing over the suction-box, and then the pallet slides on its planed bottom over the planed top of the suction-box, thus making an air-tight joint. A pair of cast-steel sprocket- wheels, turning inside of concentric guide-rails, lift the train of pallets from the lower to the upper track by engaging their teeth with the roller-wheels, and launch each pallet in a horizontal path under the feed-hopper and igniting-fur- nace, and over the suction-box. In a train of pallets in action, all the joints are kept closed, and air-tight, by the pallet being pushed from behind. At the begin- ning and the end of the track formed by the planed top of the suction-box, there is a planed "dead-plate" over which the pallets must glide; it serves to prevent any leakage of air. After a pallet passes over the suction-box and ter- minal dead-plate, its wheels engage the ends of the circular discharge-guides. These are adjusted with the view of raising the pallet about 0.5 in. vertically and thiis automatically prying up the cake of sinter and freeing it from the grate- slots. A "breaking-roller" prevents the prying action from extending too far back, and tends to form a line of fracture. This roller, however, is not essential in all cases. On reaching the curve of the guides, the pallets one by one drop into the guides, each strikes the pallet which has preceded it and, at the same time, discharges its load of sinter-cake, and shakes free the slots of the grates. The force of the blow can be regulated by the gap left in the train of pallets at this point.' The weight of the train keeps the pallets fed down to the lower teeth of the sprocket-wheels. The igniter frequently used with this machine is a small coal-burning fur- nace built of tiles, having a grate-area of 10 by 30 in. and burning 500 lb. of coal in 24 hr. The flame after passing over the fire-bridge is deflected downward upon the ore by a brick curtain that can be raised and lowered, and then is drawn upward by the natural draft of a small stack or bleeder. The suction-box on top is 12 ft. 6 in. long and 30 in. wide, and gives for the grates an effective hearth-area of 31.25 sq. ft.; this is the true measure of the capacity of the machine. The pallets are each 30 in. wide by 18 in. long and weigh with grates 550 lb. The power delivered to the machine has its speed-factor reduced by passing through a train of gear-wheels, the last of which engage the internal gear-teeth cast in the large sprocket-wheels, and actuate the train of pallets. The complete cycle of operations is as follows : A pallet, being pushed onward tangentially from the top of the sprocket-wheels, passes under the feed-hopper, where it takes its load in the form of a continuous even layer of charge, say 4 in. thick, passes next under the ignition-furnace, where the top surface is kindled, and at the same time comes within the influence of the downward-moving cur- rents of air, induced by the suction-draft; these carry the sintering action pro- gressively downward until it reaches the grates. The roast-sintering operation is complete, the cake is discharged by dropping into the discharge-guides, and the pallet crowds its way back to the sprocket-wheels, is slowly raised to the upper tracks, and begins a new cycle. SMELTING OF COPPER US A straight-line machine of the size described with effective area of 31.25 sq. ft. weighs, without accessories, approximately 16 tons. The general arrangement of a Dwight-Lloyd sintering plant has gradually Fig. 157^ —Belt Conveyor ^^^ ^ Scale Vi«=l Foot Section through Elevation of Mixing Bins bj^^ ^^d Feeders Fig. 157C. Scale Vi6 = l Foot Figs. 157 A, B, C. — Plant of Dwight-Lloyd straight-line roasting machine. taken a standard form of which a diagrammatic sketch is given in Fig. 157 A, B, and C.^ The ores to be blast-roasted arrive on a belt-conveyor, Fig. 157 ' Drawings of plant of Ohio and Colorado Smelting & Refining Co., Salida, Colo., Mef Chem. Eng., 1912, x, 87. 146 METALLURGY OF COPPER A and B, on the top floor of the building containing the mixing bins and are dis- charged by means of a tripper into the cylindrical hopper-bottom bins. The content of each of these bins is discharged in the desired amount by its traveling belt through a regulating gate on to a main belt conveyor which dehvers on to an inclined belt conveyor raising the unmixed charge components to the feed- hopper on the top of the roaster-building, Fig. 157 C. This holds the charge, now mixed somewhat, but insufficiently to fiurnish a uniform product. This is obtained in the mixer, which receives its material from the hopper through an automatic feeder and gate, and discharges the uniform material through a chute into, the feed-hopper of the machine proper. By this arrangement the handling and mixing of the ores as well as the blast-roasting have become entirely me- chanical, and require only attendance for overseeing. In blast-roasting sulphide copper ores at Cerro de Pasco, Peru (14,000 ft. elevation), Lloyd' found that the ignition flame had to be hotter than at lower altitudes, that charges could be worked with as high a S-content as 25 per cent., and that the process proceeded a little more slowly. He also states that the flowers of sulphvu- and dust from pyritic ore, which collect in the fan and have to be removed at intervals, show no tendency to self-ignition or to forming explosive mixtures. At Trail, B. C.,^ producer gas is used for igniting the 60-mesh concentrate which contains Cu i, Fe 3, Si02 40, AI2O3 15, CaO 1.5, S 15.5 per cent, and Au I oz. per ton. From 30 to 35 tons are treated by a machine in 24 hr. with a reduction of the S-content to 1.0-1.5 per cent. According to Jacobs' the herring-bone grate of the standard machine has been replaced by one with straight slots. A mechanical cleaner^ has been devised which eliminates the man required to remove from the grates adhering particles of blast-roasted material. At Plant A, a 42- by 264-in. machine, with pallets moving at a speed of from 12 to 24 in. per minute, treats a mixture of siliceous sulphide ore (Cu 6-10, Fe 15, Si02 55, CaO 2, S 10 per cent.), sulphide concentrate (Cu 12, Fe 24, Si02 30, S 25 per cent.), and pyrite cinder (Cu 2.5, Fe 54, Si02 8, S 2.5 per cent.), all passing through a \-m. screen and 10 per cent, through a 40-mesh sieve, at the rate of 90-110 tons in 24 hr.; crude oil is used as igniter. At Plant B, a 42- by 264-in. machine, with pallets moving at a speed of from 20 to 36 in. per minute, treats flue-dust (Si02 22-30, Fe 25-29, AI2O3 10-17, Cu 6.5-8.5, CaO 1.5, total S 5-16, sulphate -S 1-5.5, high in AS2O3) of which 26 per cent, passes through a loo-mesh sieve and 90 per cent, of the loo-mesh material through a 200-mesh sieve, at the rate of 100-120 tons in 24 hr., pro- ducing a sinter, usually all coarse, with 1-2 per cent. S. The former work at the plant of the Tennessee Copper Co. is recorded by Smith. ^ The treatment of flue-dust at Mason Valley is discussed in § 122. ' Min. So. Press, 1913, vii, 908. ^ J. Buchanan, Sept., 1913. ' Canad. Min. J., 1913, xxxrv, 518. * Eng. Min. J., 1913, xcvi, 789. ' Min. World, 19 10, xxxni, 460. SMELTING OF COPPER 147 85. Summary of Roasting.' — The apparatus discussed in §55-84 represent the leading types that have been or are in use at various industrial plants. Quite a number would probably not be built again for new plants. Thus, heaps will be used only, if the ore is not suited for pyritic or partial pyritic smelting as is the case with Ontario nickel copper ores; coarse ore-kilns will be confined to roasting in connection with the utilization or rendering harmless of sulphurous gases. Of the fine-ore kilns, those of the MacDougall type have proved on the whole to be more satisfactory than the others. Mechanical reverberatory furnaces will also be replaced in time by MacDougall furnaces run with auxiliary fire-places. Blast-roasting is confined to localities in which large modern reverberatory matting furnaces are not practicable. As regards matte, blast-roasting promises to be more satisfactory than the customary reverberatory-roasting for subsequent blast-furnace smelting. Fine-ore mechanical roasting furnaces make much flue-dust, especially those having super-imposed hearth, such as the MacDougall. As most silver-bearing ores usually contain some arsenic, the gases from furnaces roasting them will be charged with AS2O3. Thus the dust-free gases from a MacDougall planf contained at standard conditions: SO2 2.545, SO3 0.275, CO2 0.1136, H20-vapor 2.784, AS2O3 0.073, 14.02, N 81.18 per cent, volume. Dust as well as vapor can be collected as long as the velocity and temperature of the gas-current are sufliciently reduced, and the time necessary be given for settling. A current-velocity of 6 ft. per sec. permits the collection of practically all suspended particles; for the complete condensation of vapor and of AS2O3 the temperature of the gases must be reduced to 143-144° C A screen-analysis of real flue-dust from Great Falls, Mont.,^ showed that it was finer than 0.5 mm. ; considering that at this plant 55 per cent, of the material charged into a MacDougall furnace is smaller than 0.5 mm., this furnace will produce a large part of the total made at the works. The method ordinarily employed in the copper smelteries of the United States to clean roaster-gases is to have them travel through a dust chamber and then through long flues before they enter the stack. The flue-system of the MacDougall furnaces at Cananea has been shown in Figs. 99, 107, and 108. At Anaconda the roaster gases travel 3017 ft. at a decreasing velocity. The research of Dunn^ has shown that all the dust and most of the fmne are settled. At Great Falls, after an exhaustive study of the subject of condensation by Goodale,* the Roesing wire-system was introduced, which has proved itself to be very efficient. The question of condensation of flue-dust in copper-plants will be discussed in §122. 1 See also Dwight, School Min. Quart., 1911, xxxm, i; Eng. Min. J., 1911, xcn, 1267. ' Dunn, Tr. A.I. M. E., 1913, XLVi. ' Elton, op. cit., 1913, XLVI. ' Goodale, op. cit., 1913, xlvi. » Tr. A. I. M. E., 1913, XLVI. • Tr. A. I. M. E., 1913, XLVI. 148 METALLURGY OF COPPER II. Smelting in the Blast-furnace 86. The Blast-furnace and Its Accessory Apparatus in General. '— The blast-furnaces in operation in the United Statfes in copper smelting resemble one another so much that they are approaching standard forms, whether a re- ducing or a pyritic fusion is carried on; in fact they are the outcome of fur- naces developed at Great Falls, Mont.^ They differ greatly from those in operation 25 or 30 y^rs ago. The earlier furnaces were copies of European models.* At first they were built of stone and brick, were square, and had a single tuyere-pipe at the back; later they were made slightly oblong and had two or three tuyeres at the back. When the greatest length with this arrange- ment of air-supply had been reached, tuyeres were added at the sides, and the width of the furnace was increased. In order to protect the walls of the old furnaces, blown from the back, against fusion and corrosion, the fuel used to be charged toward the front and ore with flux toward the back; smelting thus took place in the center of the furnace. With the advent of tuyeres on four sides, the brick walls were protected by having water-cooled tuyeres project beyond them into the fiu'nace, when the blast passing through the nozzles would strike the carbonized fuel several inches away from the wall and cause the hottest zone to prevail nearer the center than the wall. The transition from water-cooled tuyere to water-cooled smelting-zone and later to water-cooled furnace was gradual. At present most copper blast-furnaces for smelting sulphide ore are water- cooled throughout, as almost every furnace-man works for some pyritic effect in order to oxidize part of the S and Fe in the charge. He accomplishes this in part by forcing into the furnace a large volume of air, or by having a low charge- column, or by both means. The result is that usually the heat creeps up and the top becomes hot. With the upper part of the shaft built of brick, the wall- corrosions or -accretions would become unmanageable, hence the water-cooled shaft, and sometimes even water-cooled parts above the feed-floor. In a strictly reducing fusion, in which the smelting zone reaches only a short distance above the tuyeres, the upper part of the shaft is of brick, as this material abstracts less heat than a water-cooled jacket. Water-jackets at present are nearly always of soft steel, as their large sizes preclude the use of the cheaper cast iron. Mount Lyell, Tasmania, forming the leading exception. The furnaces are all oblong with tuyeres on the sides, as with a given limiting distance between tuyeres the length can be adapted to the desired capacity. The vertical section of an oblong furnace shows that the ends are usually vertical; and that the sides either taper uniformly from throat to bottom, or only the 'Mathewson, Eng. Min. J., 1911, xci, 1057. 2 Church, Tr. A. I. M. E., 1913, XLVi, 423. ' General Treatises on Metallurgy of Balling, Kerl, Schnabel-Louis, and others. Wendt, School Min. Quart., 1885-86, vii, 174, 181, 304, 314 (Alleghanies) . Egleston, Tr.A. I. M. E., 1881-82, x, 25 (Ore Knob), School Min. Quart., 1885-86, vii, 360 (Point Shirley, Boston). Tables in Metallurgie, 1905, 11, 417; 1907, iv, 104. SMELTING OF COPPER 149 lower sides enclosing the smelting zone taper, while the upper are vertical. The amount of bosh thus given is governed by the reducing effect to which the charge is subjected; the greater the angle of bosh, the stronger the reduction. All furnaces have a detached external crucible, and this is either fixed or movable. The disadvantages of loss of heat and thereby of imperfect separa- tion of matte from slag, characteristic for external crucibles of small furnaces, have been overcome in large furnaces by the large stream of molten material which often keeps the fore-hearth so hot as to necessitate water-cooling in order to prevent matte from breaking through the lining. The fore-hearth has grown with the size of the furnace, and sufficiently so to become a reservoir of matte for the converter. Little need be said from a general point of view about the acces- sory apparatus, such as slag-pots, matte-pots, etc. 87. The Blast-furnace in General.— The Great Falls, the Anaconda, the Cananea and the Mount Lyell furnaces, which are the outcome of the furnace developed by F. Klepetko at Great Falls,i showing the leading points of a mod- ern copper blast-furnace, are discussed in § 90. Figs. 158-159. — Herreshofif circular water-jacket blast-furnace with detached fore-hearth. t 88. The Herreshoff Furnace. — The prototype of this furnace is the circular furnace designed by J. B. F. Herreshoff,^ for the Nichols Copper Co., Brooklyn, N. Y., and shown in Figs. 158-159. The shaft. A, consists of a double shell of boiler-plate, 10 ft. high, 4 ft. inner diameter at tuyeres and 5 ft. 6 in. at throat, with a 2-in. water-space, water-inlet and outlet pipes, H and /, and hand-holes, not shown, for removal of sediment. At the bottom the shells are connected by rivets passing through a wrought- ' Church, Tr. A. I. M. E., 1913, XLVi, 423. 'McDowell, Tr. A. I. M. E., 1884-85, xra, 124. ISO METALLURGY OF COPPER iron ring 4 in. wide; at the top the inner shell is turned outward 4 in. to overlap the outer, similarly bent, and the two laps are riveted together. The bottom, C, is a shallow cast-iron basin bolted to the ring connecting the inner and outer shells of the jacket. It is supported by four hollow cast-iron columns, 2 ft. 6 in. high. The working-bottom is made up of a mixture of crushed fire-brick and raw clay rammed down, to be level with the discharge, and to rise slightly toward the opposite side. The blast is supplied by a No. 5 Roots blower to the air-box, D, 20X4 in., which extends around the furnace except for one-sixth of the circumference. The blast passes through four 2-in. tuyeres into the furnace; opposite each tuyere is an opening in the outer wall of the air-box, closed by a removable cover with peephole, for inspecting and poking. The detached fore-hearth is a rectangular box on wheels made up of ribbed cast-iron plates bolted together. The bottom is lined with slag-wool and fire- brick, the sides with fire-brick alone; the cover consists of a number of iron- clamped fire-brick arches. At the discharge-end near the top is the slag-spout, at one side near the bottom the matte-tap. The water-jacketed inlet for the slag-matte, 7X6 in., is placed 12 in. above the brick bottom and below the slag- outlet in order that the melted mass may trap the blast. A close contact between fore-hearth and furnace is obtained by flattening the latter at the discharge. A tight joint is formed by slag chilling around the two water-cooled surfaces which have been coated and luted with clay. During the run, the slag- matte runs continuously from the furnace into the fore-hearth where the two separate. The slag overflows continuously into a slag-pot; the matte is tapped periodically in such quantities that the slag in the fore-hearth will not sink beneath the inlet and allow the blast to issue forth. If the fore-hearth should become too cool, enough matte is tapped to allow this to take place and thus heat the hearth. Should this not remedy the evil, or should the hearth be- come choked, the water-connections are severed, and a clean heated hearth put in its place. This furnace treated in a straight reducing fusion in 24 hr. about 60 tons of charge (52 per cent, roasted pyrite, 5 raw pyrite, 3 sand) using 20 per cent, gas coke. A similar furnace,' elliptical in cross-section, 3 ft. 7 in. X6 ft. 4 in. at tuyeres, 4 ft. 7 in.X7 ft. 4 in. at throat, 10 ft. high, 13 tuyeres 2 in. in diam., 26 in. above bottom-plate, outlet 9X7 in., with circular fore-hearth, water- jacketed on sides, lined with one course of brick, put through in 24 hr. 95 tons of charge (77 per cent, roasted pyrite, 13 raw pyrite, 5 sand) with 18 per cent, gas coke and 12 oz. blast. Its former work with copper-nickel ores at Copper Cliff, Ont., is given by Levat.^ The present furnace of the Nichols Copper Co., is noted in Table 28. The orford fore-hearth, designed by H. M. Howe in 1879, which sepa- rates and continuously discharges both matte and slag, may be noticed in con- nection with the Herreshofi furnace. One form is given in Figs. 160-161,' ^ Wendt, loc. cit. "Ann. Min., 1892, i, 170. 'Peters, "Modern Copper Smelting,'' 1901, p. 294. SMELTING OF COPPER 151 which represent, a movable oblong box of ribbed cast-iron plates, fast- ened together with hooks and eyes and tied with iron rods; the top is open; there is no inlet in the end, nor a matte-tap in the side. A vertical partition wall of fire-brick, 9 in. thick, divides the fore-hearth into two unequal compartments the areas of which are as 2:5; commmiication between the two is established through a slot, 3X8 in., near the bottom of the wall. In starting, the fore-hearth, dried and heated with the slot closed by a clay plug, is rolled in front of the furnace with the larger receiving compartment facing the spout over which the slag- matte flows. This enters the larger compartment and fills it, matte settles and 'collects on the bottom while the slag overflows into a waste-slag pot. When enough matte has collected in the compartment to appear at the slag-over- flow, the plug in the communicating slot is pierced by a bent bar, whereupon matte alone will flow into the smaller compartment. Both compartments will now be slowly filled, the smaller one with matte, the larger one with matte over- lain by slag; slag will overflow continu- ously from the larger compartment and matte from the smaller, the matte- spout being laid 2 in. lower than the slag-spout. The top is kept covered with burning wood and charcoal to prevent chilling. A furnace must produce at least 50 tons slag-matte in 24 hr. to allow using the fore-hearth. Matte with under 30 per cent. Cu. gradually eats away the partition- wall; matte with 50 per cent. Cu and over is likely to choke the slot. Figs. i6o-i6i.—Orford original fore-hearth. I* ll l'\ ■f Figs. 162-163. — Orford improved fore-hearth. An improved form of this fore-hearth is shown in Figs. 162-163. This again is an oblong box of cast-iron plates lined with fire-brick, having a cross- partition with a 3X4-in. slot. The distance from bottom to slag- and matte- spouts is the same; the difference in level is obtained by placing the box on an inclined plane. IS2 METALLURGY OF COPPER The overflo.w-slag from both forms of fore-hearth has to pass one or more settling-boxes or pots to give entrained matte the time necessary for com- plete separation. The fore-hearth is better adapted for matte-concentration than for ore-smelting, as with much matte there is formed a slag usually too rich in Cu to be discarded. However, the Hall Mining and Smelting Co.,' the Tyee Copper Co.,^ and the Copper Queen Mining Co., at one time had adapted the principle of continuous matte-discharge to their ore-furnaces. 89. Blast-furnace Buildings, — Every blast-furnace building has at least three floors, the feed-floor, the furnace-floor, and the slag or matte-floor. They are shown clearly in Fig. 164, representing the furnaces of the Shannon Copper Co. The details of the feed-floor vary with the manner of delivering the charges to the furnace; the distance between the feed- and furnace-floors is governed by Fig. 164. — Blast-furnaces Shannon Copper Co. the height of the furnace; the slag- or matte-floor has to be a sufficient distance (10 ft. more or less) below the furnace-floor to admit waste-slag cars holding from S to 15 tons of slag, and the matte-cars or matte-receiving ladles with a capacity of say 10 tons. If the slag is granulated, special provision has to be made to carry away the granulated material. 90. Great Falls, Anaconda, Cananea and Mount LyeU Blast-furnaces.— The Great Falls or Klepetko, Figs. 165-167,=' the New Anaconda or Mathewson, 1 Harris, Eng. Min. J., 1906, Lxxxi, 178. 2 Jacobs, op. cit., 1909, lxxxvii, 1232. 3 Hofman, Tr. A. I. M. E., 1904, xxxiv, 283. Church, op. cit., 1913, xlvi, 423. SMELTING OF COPPER IS3 Figs. 165-167. — Great Falls or Klepetko blast-furnace. 154 METALLURGY OF COPPER -E- ^._ Missing Page SMELTING OF COPPER T-SS iS6 METALLURGY OF COPPER xaqmt3T[0 !1ppq Qtl^j SMELTING OF COPPER 157 Figs. 168-170,1 the Cananea or Shelby, Figs. 171-172,* and the MoimtLyell or Sticht, Figs. 173-174,' furnaces represent the modern forms of blast-furnaces. They have this in common: they are oblong, have vertical ends, sloping or boshed sides with the necessary tuyeres, a shallow crucible which discharges slag-matte mixture continuously over a raised spout trapping the blast (first used by R. H. Sticht), a large fore-hearth for separating and collecting matte, to be tapped periodically from slag overflowing continuously into a slag car or a granulating device. Only the leading features will be briefly reviewed; the details and those of some other important furnaces are assembled in Table 28. 91. The Hearth. — This is sometimes built up solid from the concrete founda- tion, Figs. 165-167; in most plants it is erected on ribbed cast-iron plates sup- ported by iron posts and jack-screws, Figs. 171-174;^ the latter are sometimes carried by a steel truck. With the new Anaconda furnaces. Figs. 168-170, both arrangements aire found, the center division being carried by jack-screws, the end-divisions built up solid. The masonry of the hearth is encased by heavy ribbed cast-iron plates firmly bolted together. The jacked cast-iron bed-plate supporting the bottom of the hearth usually also carries the jackets; in this case the water-cooled side-walls as well as the air-cooled bottom are made thinner than with the hearth erected upon concrete, e.g., side 22 in. and bottom 18 in. thick vs. 14 and 9 in. Sometimes the bottom-plates of the hearth contain pipe-coils for water-cooling. The refractory material used for lining used to be exclusively fire-brick. The corrosive action of hot matte not too high in Cu has been in many cases the cause of replacing fire-brick by silica-brick or chrome brick. ° An analysis of the chrome brick used at Garfield gave: Cr203, SiOa 6.9, FeO 14.0, CaO 0.5, MgO 17.0 AI2O3 11.6. The brick, usually laid both end- and side-wise, are stepped down from near the level of the tuyeres to the bottom and give the crucible a trough-like shape. The trapped spout^ for the continuous flow of slag-matte is situated at one side, Figs. 165-167, or one end. Fig. 174; sometimes there are spouts at both ends. For emptying the furnace, there is a tap-hole, sometimes water-jacketed, at the lowest point of the crucible or in the spout. ^ With the new Anaconda furnace. Figs. 168-170, having a tuyere-section 56X612 in., the bottom of the hearth slopes from the center toward the ends, where are situated crucibles from the deepest points of which the slag-matte flows out at one side over two 1 Mathewson, Eng. Min. J., 1906, lxxxi, 370. Austin, Tr. A. I. M. E., 1907, xxxvii, 442. Offerhaus, Eng. Min. J., 1909, Lxxxvm, 243. * Shelby, Eng. Min. J., 1908, ixxxv, 841. ' Sticht, Melallurgie, 1906, iii, 766; Private Communication, 1911. * See also Mather, Tr. A.I.M. E., 1903, xxxiii, 675. 'Lang, Eng. Min. J., 1897, lxiii, 89. Packard, op. ciL, 1897, LXiil, 159. Glenn, Tr. A. I. M. E., 1901, xxxi, 374. "Poupin, Eng. Min. J., 1912, xciv, 785. ' Church, Tr. A.I.M. E., 1913, xlvi, 436. iS8 METALLURGY OF COPPER Tapping Breast. Oast Iron. 2 Ken. Figs. 177-179. — Cast-iron tymp of Cananea blast-furnace. Holes Lug- 1 Iron Bar 2 Oil Figs. 180-182.— Wrought-iron water-cooled spout of Cananea blast furnace. SMELTING OF COPPER 159 spouts. Each crucible has a tap-hole to empty the furnace. The latest fur- nace, 56X1044 in., has three continuous discharges. Many materials have been tried in the construction of the water-cooled tymp and discharge-spout. The tymp is made usually either of cast-iron or of copper,' sometunes of fire-clay ,2 which requires frequent renewal. The cast-iron tymp of the Cananea furnace is shown in Figs. 177-179; the copper tymp of the Great Falls furnace in Figs. 165-167. Longitudinal Section Through Breast Jacket, Spout, and Nose Pieces mmmmmmmm ^j Water.Jacketeii ?• Upper Nose ^ Piece ;5J Solid Lower Nose Piece Discharge End of Spout Figs. 183-185. — New blast-furnace-spout of Great Falls. A spout' of solid cast-iron with water-cooled nozzle used to be common; wrought-iron pipe surrounded by cast-iron has proved unsatisfactory; a water- cooled wrought-iron spout, as the one at Cananea, shown in Figs. 180-182, especially if provided with a cast-copper removable nozzle, lasts a long time. ' Hixon, Eng. llin. J., 1904, Lxxviii, 992. - United Verde Copper Co., Vail, Eng. Min. J., 1913, xcvi, 341. ' Hixon, Eng. Min. J., 1905, Lxxx, 673. i6o METALLURGY OF COPPER The spout, which usually rests with the lower end upon the breast-opening and .with the upper upon the fore-hearth, is fastened by tie-rods to the breast-jacket. Ordinarily the floor of the crucible lies below the breast-opening and forms a well, which necessitates the tap-holes, mentioned above, to empty the crucible. At Great Falls,' the breast-opening is now placed at the lowest point of the cruci- ble, and the trough of the spout, shown in Figs. 183-185, pitches downward from the breast-opening instead of inclining upward; thus the contents of the /f¥=^- f'42 Id 15^'G.P.Outlet Ij/'o.P.Tap. % Patch Bolts Used. Through-out IJ^'pitch * — O.L.Patoh_Bolte:;2': *1" ++ Inlet 2 G.P.Tap f 2_ 2 GJ.Iap . Other side only—; ^'K ?+=* =* =i=^ J_l-/ia •i2 l*~4!rH -3 1- 'f-ei^ -2 Gas Pipe Tap <4=^ (T'T O— I'^'^S.P.Tap W'^G.P.Tap^gjT j; ^■t 1— J^-t ^^-H \'J ,, . T"n-lK"G.P.Tap IM" G.P.Tap? 5 , ,,. Figs. 186-188. — Blast furnace-spout of Copper Queen smeltery. 4^^ =t=fc5 crucible drain into the spout which has a tap-hole opposite the breast per mitting the tapping the fluid contents of the furnace into the fore-hearth. At the Copper Queen blast-furnaces the bottom pf the crucible is flush with the breast-opening, and the wrought-iron water-cooled spout, shown in Figs. 186-188, with the bottom parallel to the sides is placed horizontally. A dam of refractory material, which has to be frequently patched or replaced, is rammed in the spout to trap the blast. If the furnace contents are to be drained, the dam is removed. In some cases, especially with corrosive Ni-Cu matte, it has been found neces- sary to discard metal altogether and substitute for it magnesite or chrome ' Church, Tr. A. I. M. E., 1913, xlvi, 436. SMELTING OF COPPER i6i brick. 1 Such a spout is shown in Fig. 189, in which A is magnesite or chrome brick, B water-cooled copper nozzle, C cast-iron plate, D water-cooled tymp. At the works of the Canadian Copper Co., Copper Cliff, Ont., even the brick have had to be cooled by a water-coil placed next to casting C. Fig. 189. — Blast-furnace-spout lined with magnesite- or chrome-brick. 92. The Shaft. — The width of the oblong shaft at the tuyere-level shows a range of from 42 to 56 in., the length from 150 to 612 and even 1044 in. with the latest Mathewson furnaces at Anaconda. The advantages of increasing the length of a furnace are, saving of end-jackets, diminution of loss of heat by radiation and hence saving of fuel, and increase of regularity in operation and of smelting power. The investigations of Roberts* at Great Falls give numerical data for the fact that the saving in radiating surface by lengthening a furnace and thereby diminishing for a given area the surface occupied by the end-jackets, takes place at a rate which decreases as the furnace grows in length. The corollary is that the heat units carried away by the cooling-water of the jackets decrease at the same rate. This is the reason why Great Falls has adhered to 15 ft. as a standard length of the tuyere-section. However, at Ana- conda, with a length of 87 ft., the saving in fuel has been from 10 to 15 per cent.; in addition there has been a reduction of floor area of 50 per cent., and of labor of 25 per cent. The tendency today is toward long furnaces. The increase in area from tuyere to throat of furnace i : 1.30-1.60, is accom- plished either by a bosh or by gradual enlargement. The Great Falls, Figs. 165-167, and Anaconda, Figs. 168-170, furnaces have a bosh of i in. per ft. for a distance of 7 ft. 5 in. ; the rest of the shaft is vertical. The Cananea, Figs. 171^174, and most other furnaces show a gradual enlargement. The Mount ' Reader, Mines and Minerals, 1911-12, xxxii, 55. ' Tr. A. I. M. E., 1913, xLvi, 445. i6i METALLURGY OF COPPER Lyell furnace, Figs. 175-176, has vertical sides for a distance of 18 in., followed by a bosh 48 in. high with a total deflection of 4-S in-; the rest of the shaft is vertical. The arrangement at Granby, B. C.,' is similar. At Keswick,^ Cal.,'' cast-iron lower jackets similar to the Mount Lyell, and steel upper jackets were used. % z 1 X 5 Plate Figs, igo-igi. — Cananea steel water-jackets, end. The working-height or smelting column is usually from 10 to 14 ft.; the level of the top of the charge is adapted to the coarseness of the mixture. Usually the sides of the furnace are built of two tiers of fire-box steel-plate jackets. The upper jackets either rest directly upon the lower. Figs. 165-167, or they are suspended by hangers from I-beams (mantle-frame) which carry the 'Lathe, /. Can. Min. Inst., 1910, xni. 2 Keller, Min. Ind., 1897, vi, 232. SMELTFXG OF COPPER 163 structure above the feed-floor, Figs. 171-174. Upper and lower jackets are sometunes separated by a course of brick. The jackets are braced by longi- ^ X 4 z 6 Plate Figs. 192-193. — Cananea steel water-jackets, side. tudinal I-beams bound by tie-rods. The space between the tops of the upper jackets and the feed-floor is usually covered by cast-iron mantle-plates which 164 METALLURGY OF COPPER receive the impact of the charges as they are fed into the furnace. The water- jacketsi ^j-e now nearly always flanged steel plates. The inner or fire-plate is made heavy, i-i in. thick, to prevent buckling or warping,^ the outer | in. thick; the former receives its support by distance-pieces, riveted to the outside plate, and offers a smooth strong surface to the descending charge. Stay-bolts on the inside plate, which used to burn off or be knocked off in barring or be attacked by corrosive material settling at ""f U 2'lii— I Overflow ¥ i 11 !! 1 44 +f4+-li I I II I I I 1 I U I ill III II! Ill li ! V h I I < I 1 1 I II 1 1 1 1 III! : I !^ I 11 li ]i II I " 1 1 1 1 1 1 II I 'I I I II I 111! !l 1! ii I l ! -ti ,, i I I I ig >T! ! [nl,^p-i -lc . «(., i9ii,xxxii, 55. Cerro de Pasco: Strauss, Min. Sc. Press, 1908, xcvii, 637; Min. World, 1910, xxxii, 709; Lloyd, Tr. Mel. Inst. Min. Met., 1909-10, i, 11. Copper Queen Cons. Min. Co.: Editor, Eng. Min. I., 1905, lxxx, 197; Woodbridge, op. cit., 1906, Lxxxii, 242, 298 (blast-furnace, ore-bedding) ; Brinsmade, Mines and Minerals, 1907, xxvn, 273; Milton, op. cit., 1909, xxx, 148; Lee, Eng. Min. J., 1910, xc, s°4 (Dust); Rose, Gliick Auf, 191 1, xlvii, 107. Douglas Smeltery: Barbour, Eng. Min. J., 1908, lxxxv, 303; Tucker, op. cit., lxxxvi, 413. Ducktown Sulphur, Copper & Iron Co.: Alabaster- Wintle, Tr. Inst. Min. and Met., 1905-06, XV, 274; Freeland, Eng. Min. J., 1903, lxxv, 664. Garfield Smeltery: Beason, Eng. Min. J., 1906, lxxxi, 509; Ingalls, op. cit., 1907, lxxxiv, 576; Brinsmade, Mines and Minerals, 1908, xxviii, 305; Kroupa, Oest. Jahrb., 1908, lvi, 213, Granby Cons. Min., Sm. & Power Co.: Hodges, J. Can. Min. Inst., 1908, xi, 408; Sacket, Mines and Minerals, 1910, xxx, 524; Lathe, /. Can. Min. Inst., 1910, xiii, 275; Jacobs, Met. Chem. Eng., 191 1, DC, 406; 191 2, x, 113; Avery, £»g. Min. J., igi2,xciii, g35;Lee, Met. Chem. Eng., 1912, X, 147. Great Cobar Smelting Works: Correspondent, Eng. Min. I., 1908, lxxxv, 950; Austin, Min. Ind., 1911, xx, 225. Greenwood Copper Smelting Works: McAllister, Eng. Min. I., 1911, xci, ion; Bell, Tr. Can. Min. Inst., 1913, xvr, 152. General: Christensen, Eng. Min. I., 1908, lxxxvi, 847; Min. World, 1909, xxx, 381; 1910, XXXIII, 489. Horseshoe Mg. Co.: Fulton-Knutzen, Tr. A.I.M. E., 1905, xxxv, 326. International S. & R. Co.: Palmer, Min. World, 1910, xxxii, 419; Mines and Minerals, 1911, XXXI, 321; Mines and Methods, 1909-10, I, 149; Repath-McGregor, Met. and Chem. Eng., igii, IX, is; Thomson-Sicka, Tr. A. I. M. E., 1913, XLVi. Kyshtim Smelter: Carlyle, Eng. Min. I., 1912, xcin, i23i;Lange, Metall-Erz., I9i3,x, 108. Mammoth Copper Min. Co.: Campbell, Min. Sc. Press, 1908, xcvi, 30; Martin, op. cit., 1908, xxDC, 309; 1909, XXXI, 311; Min. World, 1908, xxix, 310; Haskell, Mines and Methods, 12 178 METALLURGY OF COPPER throat to hearth area is about as 1.3 : i, showing that with a working height of about 12 ft. the sides are very steep, or that most furnaces aim to have very little reducing action in the shaft. This is shown similarly by the very small amount of bosh of the jackets. The water-jackets in nearly all cases extend down to the bottom of the crucible. Most of the tuyeres are 4 and 5 in. in diameter; the tuyere-ratio shows a considerable variety, which seems to prove that there exists still a diversity of opinion upon this point; its expla- nation lies in part in the character of the ore treated. (a) Reducing Smelting 98. Reducing Smelting in the Blast-furnace of Roasted (Raw) Sulphide Ore for Matte. — A reducing fusion in the blast-furnace is a process in which enough carbonized fuel is added to the ore-charge to furnish the reduction and the heat necessary for the operation. It is intended that the blast shall oxidize only the C and no S. Any elimination of S as SO2 taking place during the descent of the charge in the furnace may be due to oxidation by the ascending gas cur- rent, but is probably caused by the action in the charge of oxide upon sulphide. Roasted sulphide copper ore contains oxides, sulphates, and undeifcomposed sulphides of Cu and Fe, subordinately also of Zn, Pb, Mn, perhaps some As- and Sb-compounds, and the gangue. In the reducing fusion, Cu, Fe, and S form a matte which takes up the precious metals and part of the Zn, Pb, As, igo8, XXXVIII, 392; Rice, Eng. Min. J., 1911, xci, 614; Tupper, Min. Eng. World, 1912, xxxvi, 337- Mason Valley Smelter: Read, Min. Sc. Press, 1912, cv, 267. Mount Lyell Min. and Ry. Co.: Sticht, Min. Ind., 1907, xvi, 428; Melalhirgie, 1906,111, 563, SQij 638, 664, 686, 709, 760, 788 (drawing of blast-furnace, Min. Ind., 1907, xvi, 350). Rio Tinto, Baron, Min. World, 1909, xxxi, 681. Shannon Copper Co.: Corresp. Min. Sc. Press, 1902, lxxxiv, ioi. Tennessee Copper Co.: Heywood, Eng. Min. J., 1904, lxxvii, 231; Alabaster- Wintle, Tr. Inst. Min. Met., 1905-06, xv, 269; Channing, Eng. Min. J., 1905, lxxdc, 1195; lxxx, 6; Min. Sc. Press, 1908, xcvi, 97; Freeland-Renwick, Eng. Min. J., 1910, lxxxix, 116; Guess, op. cit., 1910, xc, 866; Morgan, Min. Sc. Press, 1910, ci, 677; Falding-Channing, Eng. Min. J., 1910, xc, 555; Emmons, op. cit., 1911, xci, 15; Tr. A. I. M. E., 1910, XLi, 723; Nelson, Mines and Methods, 1912, in, 407; Min. Sc, 1912, lxv, 149; Offerhaus, Metall-Erz., 1913, x, 863. Teziutlan Smeltery: Corresp. Eng. Min. J., 1909, lxxxviii, 655; igio, xc, 169. ' ^ Trail Smeltery: Turnbull, Mines and Minerals, 1910, xxxi, 121; Buchanan, Tr. Can. Min. Inst., 1913, xvi, 156. Tyee Copper Co.: Maynard, Eng. Min. J., 1909, lxxxviii, 905; Jacobs, 0^. cit., 1072; Phelps, Min. Sc. Press, 1907, xcv, 782; "B. C. Report Minister of Mines," 1902, 243. United States Metals Refining Co.: Vail, Eng. Min. J., 1913, xcv, 1031; xcvi, 553. United Verde Copper Co.: Vail, £»g. ifire. 7., igij^ XCVI, 287, 341. Wallaroo Smelter: Cloud, Tr. Inst. Min. Met., 1906, xvi, 55, 100 Washoe Plant: Hofman, Tr. A. I. M. E., 1904, xxxiv, 258; Austin, op. cit., 1906, xxxvn, 431: Correspond., M^nes and Minerals, 1907, xxviii, 131, 248; Offerhaus, Eng. Min. J 1908, LXXXV, I189, 1234; IXXXVI, 747; 1909, LXXXVIII, 243. ' Yampa Smeltery: Palmer, Min. Sc Press, 1909, xcK, 225; Christensen, Min. WorU 1909, XXX, 621. SMELTING OF COPPER 179 and Sb; the gangue with the necessary fluxes form the slag, consisting of Si02, FeO and CaO, some AI2O3' and other bases. The formation temperatures with the effects of different bases have been discussed elsewhere.^ As it usually takes many tons of ore, or slag, to make i ton of matte, the ratio of concentra- tion being 10+ :i, the composition and character of the slag to be formed is one of the first considerations in making up the charge. foCu \ \ <\ R V \ • V ^^r— n fi - \ \ r ^... ^<,„ "~~-~^^ "■~--~^ ■ ---__ -^igcbo ^~~~~^^^ -SS^c^o — ■— __^ 1 - — --___^H 1.B 2.0 3.0 Fig. 225. — Solubility of CuaS (30-per cent, matte) in ferro-calcic silicates of different degrees of silica tion. 99. Blast-fumace Slag in Reducing Smeltmg.— The slags formed in the redudng smelting of roasted sulphide copper ore show a great variety in silicate- ' Bellinger, Eng. Min. J., 1912, xciv, 321; Min. Sc. Press, 1912, cv, 114; Met. Chem. Eng., I9I2,X, 693. 'Hofman, " General Metallurgy," 1913, p. 454-463. i8o METALLURGY OF COPPER degree and composition. The former ranges from sub- to bi-silicate, but usu- ally is near a singuJo- and sesqui-silicate. With ores rich in Fe, the percentage of SiOa covering a range of from 28 to 40 per cent., FeO is generally high (50 per cent.) and CaO low (10 per cent.) ; the reverse is the case with ores contain- ing little Fe; AI2O3 is rarely high (15 per cent.), more frequently low (4 per cent.) than medium (8 per cent.). The main requirements that the slag has to fulfiP are that it shall form at a low temperature, require little superheating to be fluid, have a specific gravity not too high to allow a satisfactory settling and sepa- ration of matte, and be cheap, i.e., not require much flux. There is little danger Per Cent 0.6 K \ \ 0.4 Y Per Cent ' Cu 1.6 / n ^ 1.4 \ A ^1.2 3 /i / 0.8 « ^^Ss.-^ 1 \ y4 n X 1 i 0.6 J^ f .s 0.4 "a ■-n ^^. 1 \ M ^ ^ 0.0 0.0 H 2/4 54 Beplacement of CaO in Curves a, 5 and c Beplacement of FeO in Curve d, V4 Fig. 226.— Effect of replacement of FeO and CaO by AI2O3, MgO and ZnO in a sesqui- ferro-calcic silicate upon the solubility of CujS (30-per cent, matte). of Cu being scorified as long as there is enough FeS present to sulphurize any metallic or siUcate of copper that may have entered the furnace or have been formed in the downward passage of the charge. Wanjukow'' has investigated in the laboratory the solubility of CU2S of a 30-per cent, matte in ferro-calcic silicates. He finds that the solubility falls with the degree of silication as shown in Fig. 224; and that it decreases with the ' Mostowitsch, Metallurgie, 1912, K, 559. ' Op cit., 191 2, DC, I, 48. SMELTING OF COPPER i8i replacement of FeO by CaO, Fig. 225. The results of the effects of replacement of the constituents FeO and CaO by the bases AI2O3 MgO and ZnO in two sesqui-silicate slags are assembled in Fig. 226. Here curves a, b, and c represent the solubility of CU2S (30-per cent, matte) in the sesqui-silicate SiOz 42.42, FeO 21.58, CaO 36.00; and curve, d, that in the sesqui-silicate SiOa 39.78, FeO 48.22, CaO 12.00. In general, slags which contain a metal having much affinity for S will carry more Cu than those which have little. In the following list by Wanjukow, Cu stands at the head, Al at the bottom: Cu, Ni, Co, Fe, Mn, Zn, Ca, Mg, Al. The effects of varying percentages of Cu in matte upon the Cu-content of slags is taken up in § iig. The compositions of slags formed in a reducing smelting may be the same as those made in pyritic smelting (§ 106) or in partial pyritic smelting (§ no, in) as the latter must fulfil requirements similar to the former as far as specific grav- ity and fluidity are concerned. On account of the great latitude in composition, and of the fact that smelteries usually treat ores from a single district, typical slags, such as have been developed in lead-smelting, have not been devised, although these will work in the reducing fusion of a copper blast-furnace as they do in that of the lead blast-furnace. When the leading ore was a cupriferous pyrite or pyrrhotite with little gangue, the slags made consisted mainly of Si02 and FeO, totaling over 90 per cent., the remainder being small amounts of AI2O3 and earthy bases. The investigations of Hofman^ have shown that with pure ferrous silicates the for- mation temperatures decrease as the silicate-degree rises, but experience has proved that the reverse is the case with fluidity; the singulo-sihcate forms at a higher temperature and is more fluid than the bi-silicate, and thie 3 : 4 or the 2:3 siUcates lie between the two. The curves of Hofman also show that the replace- ments of FeO by CaO lower the formation tfemperatures to a certain point, be- yond which they rise again; also that low-SiOs ferrous slags can endure more CaO before they reach the minimum than high-Si02 slags. Experience has shown that additions of CaO up to certain amounts increase the fluidity. These statements give the reasons for the preference of the i : i and 2 : 3 ferrous sili- cates over those that are either more basic or more acid. With slags more basic there is danger of hearth accretions, with slags more acid, there is either a small tonnage or a high coke-consumption to give the slags the required fluidity. Thus 2FeO.Si02, with Si02 29.20 and FeO 70.80 per cent., reduced to a total of 95, gives Si02 27.74 and FeO 67.26, leaving 5 per cent, for other oxides; in the same manner, 3Fe0.2Si02, with Si02 35.70 and FeO 64.30 per cent, gives Si02 33.915, FeO 61.085, RO 5.000 per cent.; and the 4R0.3Si02, with Si02 38.46 and FeO 61.54 per cent., gives Si02 36-S37> FeO 58.463, RO 5.000 per cent. With these ferrous slags, 28 per cent. Si02 is about as low as one dares to go, s:^ per cent. SiOz is a better figure; 38 and 39 per cent. Si02 is rather high. There is an old rule which it is safe to follow in starting: to make SiOz about 33 per cent., to figure the iron as Fe = Si02, and to have about > "General Metallurgy," 1913, p. 4SS; Tr. A. I. M. E., 1899, xxi, 682. i82 METALLURGY OF COPPER 10 per cent. CaO. This will give a total of 95.5 per cent, and allow 4.5 per cent, for other oxides. Slags made in some of the few remaining copper blast-furnaces in which a strictly-reducing fusion is carried on are given in Table 28. 100. Fuel and Blast. — The fuel ordinarily used is coke; a common ratio is 6 charge : i coke, which is equal to 14 per cent, coke; this figure sometimes falls to 13 and again rises to 17 per cent. ' An overheating of a ferruginous slag by an excessive amount of coke is likely to cause reduction of iron to the metallic state. The furnace runs best if it has just the right amount of fuel; any lack will cause the forming of long noses at the tuyeres and a corresponding reduc- tion of tonnage. Charcoal, which used to be the universal blast-furnace fuel, has been given up in practically all non-ferrous blast-furnaces.^ Where one is forced to use it, the amount required may be one-third larger than the necessary coke. In a few instances green wood^ sawed into 2-ft. lengths has been successfully used to replace as much as \ of the coke, i lb. coke being equal to from 2.6 to 3.0 lb. wood. Experiments with oil as blast-furnace fuel have been carried on by Hamilton' Kiddie,* Waters,^ and Lang. ^ The blast in a reducing fusion is hardly ever preheated; in some instances,' as with the Giroux hot-blast, Figs. 197-198,' part of the heat of the tunnel-head gases is utilized for this purpose or the Kiddie hot-blast system.^ The blast-pressure will vary greatly with the width of the furnace, the diam- eter and number of tuyeres, the amount of fines in the charge and the percentage of iron. Formerly a pressure of 12 oz. per square inch was common; with the increase of distance between tuyeres this figure has grown materially (see Table 28). Lloyd'" gives as his experience of smelting at Cerro de Pasco, Peru, altitude 14,000, that a blast-furnace behaves about the same way as at sea-level, except that its smelting power is smaller; that with slag-composition the same holds good; that any pyritic effect (§ 104) is lower; that on account of the diminished smelting power radiation losses are to be avoided (no jacketing of crucible walls), the tuyeres should be of larger diameter and the distance (or width of fur- nace) smaller, and the coke of good quality. 1 Modern exception: C. S., Eng. Min. J., 1911, xci, no. 2 Trans. A. I. M. E., 1891, xx, 54s (Lang); Eng. Min. J., 1902, Lxxiv, 646 (Collins); 1906, Lxxxii, 700 (Mitchell); 837 (Bromly); 1013 (Bretherton); 1910, lxxxix, 774 (Bretherton). ^ Eng. Min. J., 1911, xci, 224. * Op. cit., 1911, xcii, 434 (Jacobs). ^ Op. cit., 1912, xciii, 877. ^ Min. Sc. Press, 1913, cvi, 248. ' Bretherton, Eng. Min. J., 1899, lxviii, 604, 698; 1900, lxix, 614; ixx, 760; Min. Sc. Press, 1900. Lxxxi, 572, 1912, civ, 243. 8 Traylor Engineering Co., Eng. Min. J., 1906, lxxxii, 698; Min. Sc. Press, 1906, xcill, 792. Vail, Eng. Min. J., 1913, xcvi, 341 (United Verde Copper Co.). "Jacobs, Eng. Min. J., 1906, lxxxii, 598. 1" Tr. Met. Inst. Min. Met., 1909-10, i, n, SMELTING OF COPPER 183 Sacio' concludes from his study of smelting at high altitudes, that the ca- pacity of blowers, air-conduits, and tuyeres has to be increased; that effort has to be made to diminish the loss by radiation; and that the use of hot blast is desirable. loi. Chemistry of Reducing Smelting.— The details of the chemistry of the reducing fusion of roasted sulphide copper ore have been studied little. In. general, the processes to be considered are reduction, sulphurization, decompo- sition, and slagging; oxidation is confined practically to the burning of the fuel although some sulphide may be attacked by free 0. The principal reducing agents are C and CO. (i) Reduction may be expressed by 2CuO-|-C = 2Cu-f CO2, CuO+CO = CU-I-CO2, 2CuS04-|-3C = Cu2S+S02-f-3C02; Fe,0„+yC = Fe.-l-yCO begin- ning at 400° C and Fe^O;c+!,+yCO = xFeO-fyC02 beginning at 200° C; CaS04+ 4 (or 2) C = CaS-f4C0 (or 2CO2) and CaS04+4CO = CaS+4C02 beginning at 700° C; similarly BaS from BaS04 by C beginning at 600° C, and by CO at 650° C. (2) Sulphurization by 2Cu-)-FeS<=±Cu2S+Fe, Cu20+FeS = Cu2S+FeO, Cu4Si04-|-2FeS = 2Cu2S-|-Fe2Si04, Cu20+Ca(Ba)S = Cu2S+Ca(Ba)0. (3) Decomposition of MSO4 by MS04+heat = MO-|-S03(S02+0); of (Alk. earth)S04 by Ca(Ba)S04+Si02 = Ca(Ba)Si03-|-S02-fO beginmng at about 1000° C. (4) Slagging has been referred to on page 179. (5) Matting and slagging by 4CuO+3FeS-fSi02+2C = 2Cu2S,FeS+Fe2Si04 -I-2CO. Before considering the changes in the ascending gas-current, it is necessary to picture the conditions of the charge extending from tuyere to throat. In the smelting zone coke will prevail over melting refractory parts of the charge requiring the high temperature of this region to become liquefied; matte and eutectiferous constituents of the reduced original charge have been melted above and have run down below the tuyere-level. Higher up, the relative amounts of fuel and charge will be approximately the same as when fed at the throat. The temperature at the tuyeres of about 1200° C. will decrease toward the top of the charge from which the gases leave at a temperature of 250° C. or higher. The blast entering through the tuyeres strikes coke at a temperature of 1200° C. According to Ernst^ the C burns to CO, but the large volume of air enter- ing oxidizes the CO in part to CO2, so that at the tuyere-level the gases are a mixtureiof N, CO2, and CO. As they ascend in the furnace they arrive quickly at the region of lower temperature; the reducing power of C, burning now to CO2, and that of CO, also burning to CO2, increase as long as this power is not weakened by the increasing presence of CO2. On the whole the percentage of CO2 in the gas-mixture becomes larger as this rises in the furnace, and will strongly prevail over CO when it leaves the top of the charge. The older gas ' School Min. Quart., 1913, xxiv, 344; Mel. Chem. Eng., 1913, xi, 499. 'Hofman, "General Metallurgy," 1913, p. 294. 1 84 METALLURGY OF COPPER analyses of Bunsen/ Kersten,^ Schubin,' and Heine^ appear to substantiate this ; modern analyses are wanting. The gases from the blast-furnaces at Mans- feld, Germany (Table 29), form an exception, as the furnaces are run more in the manner of producing pig-iron than oi matte; in fact, part of the gases is used to preheat the blast, the rest in gas engines. Table 29. — Ttjnnel-head Gases from Mansfeld' Name of smeltery Krug Koch Eckhardt Kupfer-hammer CO2 CO N (diff.) 9-3 18.8 71.9 II. I 20.6 68,3 13.6 IS-2 71.2 IS. 6 12.8 71.7 Considering the changes in the descending ore-charge, at first any H2O present will be expelled, then CuO will be reduced and sulphurized, CUSO4 decomposed or reduced, CuSi04 converted into CU2O and CU2S lower down in the furnace; incidentally reverberatory furnace reactions Cu2S-|-Cu20 = 4Cu-|-S02 may occur and leave Cu to be sulphurized. In the upper part of the furnace porous FeaOa will be reduced both by C and CO to FeO, and this lower down will either combine with Si02 or be reduced by C to Fe and then sulphiurized; FeSOi will be decomposed near the throat without being reduced. Reduction and sul- phurization progress as the charge sinks in the furnace. Slag-formation begins only a short distance above the tuyke-level. The Cu2S-FeS mixture of lowest melting-point fuses between 850 and 900° C. and runs downward, is to a small extent oxidized by the O of the blast, gives up some of its Fe to form slag, and collects below the tuyke-level; the rest of the sulphide melts at a slightly higher temperature and follows the first. The melting of the last sulphide is coin- cident with the lowest slag-formation. In this the eutectic mixture will form, soften, and fuse first; flow downward; and gradually dissolve the less fusible parts. Every charge-component passing through the hot tuyere-region filled to a great extent with incandescent coke is melted; below it, take place the separa- tion of matte- and slag-particles, the adjustments of matte-components to form the desired matte, and of slag components to form the desired slag. Matte trickling through fused slag sulphurizes slagged copper and carries down with it the CusS formed, thus cleaning the slag. Any Zn in the charge is either volatilized, or enters the slag both as ZnO and ZnS, or the matte, and causes imperfect separation of matte and slag. Precious metals go with the matte, as ^ Poggend. Ann., 1840, l, 81, 637. 'Berg. Hiittenm. Z., 1844, in, 137. ' Op. cit., 1846, V, 569. ^Bergwerksjreund 1843, v, 208, vi, 513, 1844, vii, 547; cited by C. F. Rammelsberg in his Lehrbuch der Chemischen Metallurgie," Luderitz, Berlin, 1865 p 308 5"Der Kupferschieferbergbau und der Huttenbetrieb,"' Mansfeld'sche Kupferschiefer- bauende Gewerkschaft, Eisleben, 1889, p. 81. SMELTING OF COPPER i8s does most of the Pb; As and Sb are volatilized and enter the matte if present in small quantities; with considerable amounts a speise may be formed.' 102. Calculation of Charge.^— Of the different methods of calculating charges the one based upon the production of a slag of a given silicate-degree is chosen here, as it serves for the large range of composition of the slags that are made in a reducing as well as a pyritic fusion. Another method based upon a slag of definite percentage composition is given in § 115. The slag is to be a singulo-siUcate; the factors necessary for the computation of bases and Si02 to form silicates are given in Table 30.' Table 30. — Computation op Bases and Silica Required to Form Silicates I lb. base requires lb. SiOj to form a Name of base I lb. SiOz requires lb. base to form a Singulo-silicate Sesqui-silicate Bi-silicate Singulo-silicate Sesqui-silicate Bi-silicate 0.S3S 0.196 0.750 0.873 0.416 0.422 0.803 0.294 1. 125 1. 310 0.62s 0.633 1 .070 0.392 1.500 1-747 0.883 0.84s CaO BaO MgO Al,Oa FeO MnO 1.86 S-io 1-33 1. 14 2.40 2.36 1.24 3 -40 0.88 0. 76 1.60 I-S7 0-93 2-55 0.66 0-57 1 . 20 1. 18 The analyses of ores, fluxes, and fuel are assembled in Table 31, which gives the summary of the calculation. The ore for which the charge is to be calculated is a low-grade basic roasted sulphide copper ore; there are available a high- grade siliceous oxide copper ore to furnish the Si02 that is necessary, and a roasted gold-bearing pyrite which is to be used to combine with the excess- sulphur of the roasted ore. The weight of the charge is to be 1000 lb. and the amount of coke used 14 per cent. (i) S1O2 Available in 100 Lb. Siliceous Copper Ore. 15 lb. FeOXo.4i6 = 6.2 lb. SiOa required 8 lb. CaOXo. 535 = 4.3 lb. Si02 required 3 lb. MgO Xo. 750 = 2. 3 lb. Si02 required 9 lb. Al203Xo.873 = 7.9 lb. Si02 required Total, 20. 7 lb. Si02 required 43.0 lb. Si02 present Remain, 22.3 lb. Si02 available for fluxing purposes. ' G. C. McMurtry, "Smelting antimonial concentrates," Tr. Int. Min. Mel., 1913, xxn, so; Min. Eng. World, 1913, xxxviii, 9. ^Furman, School Min. Quart., 1896, xviii,i. Barbour, Min. Sc. Press, 1909, xcix, 664. Mostowitsch, Metallurgie, 1912, ix, 559. 'Hofman, "General Metallurgy,'' 1913, p. 435. i86 METALLURGY OF COPPER 3 ^ d NO NO CI CO i »o d c^ 00 CO It o O q < $ d 00 00 o -o o M M i' o. . y ' Eng. Min. J., 1903, lxxv, 664. SMELTING OF COPPER 191 nel (converting-slit). It consists of porous friable gangue carrying the charge; its position, however, is not stationary nor its form fixed. As the furnace forms its own pyritic bosh, the question of bosh in the jackets is not of paramount importance, but rather the temperature of their cooling- water; thus at Kyshtim, Siberia, the temperature of the lower jackets is kept cool in order to obtain a pyritic bosh of the desired thickness. The tuyeres in the furnace are mostly dark, light being rarely discernible when they are punched; a bar has been driven in from one side and withdrawn from the other with the naked hand. This proves that the tuyeres are bridged, that the melted charge passes downward in the spaces between them, and that the hot fusion-zone lies above. The blast, warmed and finely divided in its passage through the Fig. 227. — Vertical section through pyritic blast-furnace. Fig. 228. — Horizontal section through pyritic blast-furnace. porous boshes, must be delivered upward into the fusion zone against the descending streamlets of FeS and the quartz fragments of the charge, and exerts there its powerful oxidizing effect. Sticht^ smelting a pyritic ore at Mount Lyell with 1.25 per cent, coke, notes a similar porous bosh which consists of quartz fragments and slag, and is free from FcaOa or Fe304, CaO, and particles of matte. Partial pyritic smelting standing between true pyritic and reducing smelting will show some of the phenomena of both processes. The first experiments in pyritic smelting are those of John Hollway' with pyrite from Rio Tinto. In 1889 L. Austin^ did some work in Toston, Mont.; ' Metallurgie, 1906, iii, 115. »"A NewAppUcationof Bessemer's Process of Rapid Oxidation, by which Sulphides are Utilized for Fuel," Soc. of Arts, April 12, 1879. » Tr. A. I. M. E., 1887-88, XVI, 257. 192 METALLURGY OF COPPER in 1891 R. Sticht made the first successful runs in this country at Boulder Val- ley, Mont.; in 1893 the Bi-Metallic Co. ran a pyritic plant at Leadville, Colo.' The process lay dormant for a while until it received a new impetus in 1895, when Sticht introduced it at Mount Lyell, Tasmania. Since then it has grown in importance. At present entire pyritic smelting is carried on at the Mount Lyell, Ducktown, Isabella, Keswick, Mammoth, Rio Tinto, Kosaka, Sulit- jelma, and probably other smelteries. Partial pyritic smelting has become very common in this country. 105, Pyritic Smelting Proper, (i) The Slag. — From what has been said regarding pyritic smelting, it is clear that a high formation temperature of the pyritic slag is of paramount importance for the process, as the charge has to stand unmelted to enable the blast to attack the FeS, and form FeO and SO2, when the FeO must combine instantaneously with SiOa, as in converting matte; the fer- rous silicate formed will then dissolve gradually the remaining refractory slag- forming constituents of the charge and form with them the final slag. Referring toTable 32, which gives the formation temperature of some ferrous silicates, it is seen that omitting the impossible 4RO:Si02, the ferrous singulo- "Silicate 2RO:Si02 has the highest formation temperature (1270° C); thesiH- cate 3RO:Si02, comes next (1140° C); then follow the sesqui-silicate, 4RO. 3Si02, with 1120° C; and the bi-silicate, RO: Si02, with 1110° C. The silicate- degree is then a function of the temperature, and the formation temperature falls as the silicate degree rises. The aim must be therefore to form a ferrous singulo-sihcate, and the three factors, FeS, O, and Si02 have to be so balanced as to make this possible. If there is too much FeS, which is equivalent to a lack of O, there is too little oxidation; and this means lack of heat in the combustion zone, accompanied by the passage of undecomposed FeS through the blast followed by collection below the tuyere-level. If for a given amount of FeS and O, there is an excess of Si02, this will accumulate in the shaft (silica-sow) and block the smelting; if there is a lack, some Fe will be peroxidized to Fe304 or Table 32. — Formation Temperatxire of Some Ferrous Silicates^ Formula of silicate Chemical composition Formation temperature SiOj per cent. FeO per cent. Deg. C. 4RO 3RO 2RO Si02 17. 20 21.70 29.20 35-70 38.46 45-45 82.80 78-30 70-80 64.30 61-54 54-55 SiOj SiOa 1270 1 140 ' 1120 3R0 4R0 RO SiOz sSiOj SiOj ' Doolittle-Jarvis, Tr. A. I. M. E., 1910, xu, 709. 2 Hofman, Tr. A. I. M. E., 1899, xxix, 682. SMELTING OF COPPER 193 Fe203 which, taken up by the slag, decreases its fusibility aftid may stop the fur- nace. As a slag consisting solely of SiOz and FeO does not separate well from the matte, it has to be lightened by some earthy base. The percentage of CaO, including its equivalents of MgO, BaO, AI2O3, should not be less than 10 per cent.; the usual range is 12 and 16 per cent, (see Table 28). The percentage of AI2O3 in any case must be low; perhaps 7 per cent, is the limit, usually it does not exceed 5 per cent. The slag of Mount Lyell,i SiOz 32.47, FeO 52.15, CaO 4.77, AI2O3 7.22, BaO 0.90, S 0.88, Cu 0.39 per cent., Ag 0.189 oz. per ton, is a singulo- silicate. Other slags are given in Table 28. 106. Fuel and Blast. 2— The roles that fuel and blast play in p5Titic smelting have been discussed in § 104. There remain the calculation of the blast required and the consideration of hot blast. Quantity or Air. — The calculation of the cubic feet of air per minute re- quired must be based upon the amount of FeS that is to be oxidized, and this is best referred to the square foot of hearth area. Assume that the smelting power of the furnace is 6 tons of charge per square foot hearth area in 24 hr., equal to 500 lb. per hour; that the charge contains 50 per cent. FeS2; and that this loses 4 of its S by sublimation. The 500-lb. charge corresponds to 250 lb. FeS2 (46.7 Fe, 53.3 per cent. S) or 116. 8 lb. Fe and 133.2 lb. S. With f of the S volatilized, this is changed to 116.8 lb. Fe and 76.1 lb. S. Let 90 per cent, of this be oxidized and 10 per cent, go to the formation of matte; the blast has to be supplied for 105.1 lb. Fe and 68.5 lb. S. NowFe:0 = 56 : 16 = 105.1 :«;, x =3o.olb.O; and S:02 = 32 132 = 68.5 :y, ^ = 68.5; hence there are required 30.0-f 68.5 =98.5 lb. Oper hour with 98.5X3.33 or 328.3 lb. N=426.81b. air=426.8Xi3.o67 (for 15° C.) or 5577 cu. ft. air per hour = 92.9 cu. ft. air per minute. Assuming an efficiency of 90 per cent., gives 103 cu. ft. as the amount of air required. Wright* found at Keswick, Cal., that his furnace, with a charge containing 50 per cent. FeS2 and treating 11 tons of charge per square foot hearth area in 24 hr., did good work when it received per minute per square foot hearth area 365 cu. ft. air; this would oxidize 7 lb. FeS2 per minute, and the furnace gases would contain 12 per cent. vol. SO2. Preheating Air Blast. — Heating the blast has been the subject of much discussion. The idea has been prevalent that hot blast would raise the degree of concentration of the matte. Sticht has shown that not only does it not effect this, but that the result is just the reverse, because hot-blast causing a more siliceous slag to be formed is equivalent to a lower degree of oxidation. The main reason for the higher concentration of matte with cold blast is that, in order to produce the same heat more cold air is required in a given time than hot air; the larger volume of air rises in the shaft to the point at which FeS begins to melt, and this trickling down through a longer column of hot gangue matter is exposed to oxidation for a longer period of time. At Mount Lyell the degree of concentration with hot blast was 7 into i ; it is at present with cold blast from 18 ' Min. Ind., 1907, xvi, 435. * Walter, Eng. Min. J., 1913, xcm, 797. ^ Eng. Min. J., 1905, Lxxrx, 957. 13 194 METALLURGY OF COPPER to 20 into I, furnishing a matte with 40 per cent. Cu, and could be made greater by additional blast if this was desirable. It is believed that not a single bla^- furnace doing true pyritic smelting is supplied with hot blast. 107. Chemistry of Pyritic Smelting. — 'In discussing the chemical reactions that take place in the blast-furnace it is convenient to consider separately the ascending gas-current and the descending ore-charge. The blast of atmospheric temperature upon entering the furnace at the tuyere-level comes in contact with the porous boshes and bridges, is warmed and being further heated in traveling upward through melted descending charge strikes, a short distance above, fused red hot FeS trickling downward through siliceous gangue material; it oxidizes the FeS to FeO and SO2 giving up its entire O; simultaneously the FeO forms slag while the gas-mixture of N and SO2 with a temperature of 1200° C. rises in the shaft and preheats the descending charge; cooled by this to 700° C. it becomes charged with S-vapor (FeS2-|-7oo° C. = FeS-|-S), the amount of Mount Ly ell averaging 1.7 per cent, at a distance of 4 or 5 ft. below the top of the charge; the N-|-S02-(- S-vapor rises in the furnace to the top of the charge (250° C.) where the S ignites and burns. If the temperature of the stock-line should sink below 250° C, the volatilization and ignition points of S, some S would be deposited and clog the passage of the gases. During the ascent of the gases the small amount of coke charged, <3 per cent., is consumed by the reaction S024-C = S-t-C02-f-27,94o cal., which assists in warming the charge. Table 33 gives analyses of Mount Lyell gases when the furnace is running normally, and Table 34 similar data when running "wild," when FeS is not being oxidized. The fig- ures^ in Table 33 show a very small amount of free O, which varies very little in Table 33. — Normal Gas from Pyritic Furnace, Mount Lyell No. of samples Sample, distance be- neath throat, feet SO2 1 CO2 CO 5 2-2,5 6.64 S.08 0.16 0.84 4 2-2. s 7 95 3 -075 none 1.50 4 2-S 8 92s 5-45 none 0.70 13 2-3 5 7 88 5-93 0.02 0.35 5 3-35 6 12 7.86 none 0.66 14 6 8 44 4-5° none 0.4 A' 6 9 475 3 •70 none none j^ 6 10 60 4 '.40 0. 2 none 5' 6-7 7 90 3.56 none 0,88 a distance of 2 to 7 ft. below the surface of the charge, and practically no CO owing to the reaction S02-t-C0 = S-f CO2. In other words, the O entering at the tuyeres is almost wholly consumed by the oxidation of the FeS, and the small amount remaining as free is rendered innocuous by the great dilution with ' Objections by Guess, Eng. Min. J., 19:2, xciv, 925. " From the preceding 14. SMELTING OF COPPER 19s indifferent gas. As soon as the regular process is disturbed, the composition of tunnel-head gas changes (Table 34). The air rushes up through the charge effecting only little oxidation; the percentage of in the gases is high, and that of SO2 low. Table 34. — Abnormal Gas from Pysitic Furnace, Mount Lyell No. of Sample, distance be- SO2 CO2 CO samples neath throat, feet I 2 22.40 4. 20 none 3-4° I 2-5 2.40 0. 20 0.2 14.60 S 2.S-2-7S S.26 1.78 0.07 11.29 s 3 3.02 3.16 none 11.08 I 3 1 . 20 1 .10 none 16.70 7 3-3 S 2-59 4-73 0.03 6.91 Following the descending ore charge, it is convenient to distinguish three zones : (i) The Zone of Preparatory Heating. — The top of the charge is at 250° C, S-vapor is burning over it, and air is rushing in through the feed-doors at the rate perhaps of three times the volume of the ascending gas-current. The charge when introduced gives up quickly its hygroscopic water and more slowly that which is chemically combined; between 250 and 700° C. it absorbs heat from the gas-current and some of its coke is oxidized by the SO2 of the gases; at 700° C. pyrite begins to give off i mol. of S. With these changes the charge passes downward, at Mount Lyell for a distance of 7 or 8 ft., whereupon it reaches the oxidizing zone (2). (2) The Zone oe Oxidizing Smelting or Focus. — This at Mount Lyell extends downward to 2 or 3 ft. above the tuyeres. The altered charge, now FeS + gangue -f Umestone, reaches the region of 880° C. where CaCOs is dis- sociated; lower down the FeS begins to fuse, trickles down over the pieces of SiOj and CaO in separate droplets or in assembled rivulets, and is met by the of the blast; the FeS is oxidized and simultaneously combines with Si02 to Fe2Si04; this hot ferrous singulo-silicate traveling over siliceous gangue mat- ter and CaO dissolves these in its downward course and slags them so that, arrived at the level of the tuyeres, there is little of them left with exception of the silica-boshes and -bridges. (3) The Zone or Adjustment. — The melted slag-matte which descends through the more or less open spaces between the bridged tuyeres collects below these. Here the different silicate mixtures form a uniform mixture, and the matte tends to separate to some extent from the slag, although it is not given sufficient time and space to accomplish this satisfactorily, as both matte and slag leave the furnace together over the blast-trapping spout. The final separation takes place in the external fore-hearth. 196 METALLURGY OF COPPER In the oxidation of FeS there is always left enough unaltered FeS to re- sulphurize any Cu that may have become oxidized in its downward course. 108. Management and Results. — -The management of a pyritic furnace requires considerable care, as even a slight irregularity is likely to disturb the normal working. Thus e.g., the matte-fall is likely to be irregular owing to the slight changes in the ore, in the moisture of the air, or in the mechanical condi- tion of the furnace, which hinder or favor the O from doing its proper work in the focus. At the same time the character of the slag formed may not change materially although the amount will be decreased or increased. The analyses of gases from normal work, Table ^2>j show only a trace of free O and as much as 12 per cent. vol. SO2; those from abnormal work, Table 34, much free O; the furnace makes a large amount of low-grade matte and a small amount of acid slag; simultaneously the focus begins to cool. Correcting the evil by addition of coke or coke- and slag-charges wUl heat up again the lower part of the furnaces and cause the production of much low-grade matte. Regular pyritic smelting can then be started again, just as in blowing in a pyritic furnace; where the start is made with a reducing fusion. When a fur nace gets out of order, it is usually cheaper to blow it down and start fresh in an interval of 24 hr. instead of trying to doctor the patient. Thus Sticht's campaigns last about four weeks. ■■ The operations of blowing-in, etc., which are the same as in partial pyritic smelting, are. discussed in § 117. The elimination of As, Sb, Bi, and Pb in a pyritic furnace,^ is much greater than in a reducing fusion. Sticht' gives 70 per cent, as the direct efficiency of his work at Mount Lyell; 35 per cent, of the heat generated is absorbed by the chemical work of smelting and the fusion of the solid products; 35 per cent, by dissociations preceding or accompanying the chemical reaction; the balance of 30 per cent, is lost by radiation. An average of 1 1 years' work (including early experimental work, converting, and resmelting intermediary products), treating since 1907 ore with Cu 2.25 per cent, and producing matte with Cu 44.3 per cent., gave Sticht a yield of Cu 85.72, Ag 92.57, Au 102.28 per cent. 109. The Knudsen Process.''— This process, a modification of true pyritic smelting, consists in smelting intermittently with a small addition of coke and raw pyrite in a special tilting converter for matte and slag. At Sulitjelma, where it is in operation, the ores are of two kinds, siliceous (CU6.5, Fe 32.8, S 36.8, Si02 16.5, AI2O3 9.3 per cent.) giving a fluid, and schistose (Cu 5.1, Fe 26.3, S 23.9, Si02 28.8, AI2O3 12.2 per cent.) giving a sticky slag; they are mixed and furnish a waste slag with Si02 28-29, FeO 56-59, AI2O3 10-12, S 1.5, Cu ' Peters, "Principles," p. 305. 2 Lang, Eng. Min. J., 1904; Lxxvm, 461. ' Min. Ind., 1907, xvi, 435. * Schiffner, "V Internal. Kongress fur Angewandte Chemie," Berlin, 1904, n, p. 112. Knudsen, Eng. Min. J., 1904, lxxvii, 757; Min. Ind., 1903, xii, 119; 1908, xvii, 315; Oest. Zt. Berg. Huttenw., 1909, lvii, 426; 1912, lx, 568; Eng. Min. J., 1909, Lxxxvn, 1080. New furnace, U. S. Patent No. 1075214 Oct. 9, 1913. Dyck, Metallurgie, 1907, iv, 416. Hofman-Mostowitsch, Tr. A. I. M. E., 1908, xxxrx, 652. SMELTING OF COPPER 197 0.3-0.6 per cent. The ore ought not to be too fine. Charges have been run successfully with one-third coarse, one-third fine, and one-third dust. The converter shown in Figs. 229-230, consists of three parts, a cylindrical body suspended by trunnions which are connected with the air-blast and the rotating mechanism; a conical hood; and a bottom contracted toward the blast- inlet, which is close to the bottom. The converter is lined with 10 in. of magne- site brick and has an inner diameter of 6 ft.; there are 18 tuyeres || in. in diam- eter. In operating, the converter, at a yellow heat from the preceding charge and retaining a small amount of matte, receives a bed of coke or coal of 0.5 to 1.5 per cent, of the weight of the charge, which is sufficiently deep to extend a little above the tuyeres; next the blast is started to bring the fuel to a red heat. ^!flfe ^^^&!p!m!mm!mzm!z:^7^^^,r E^z my- Figs. 229-230.- — Knudsen converter. then 10 tons of charge are run in from a hopper, and the blast-pressure is raised to 5 lb. The pyrite is decomposed; the FeS melts and trickling down is oxidized, and at once combines with the Si02 of the charge. After 10 min. or after the smelting has been well started, the tuyeres being kept bright by punching, and the melted charge reaches the tuyere-level, the blast-pressure is raised to 10-15 3.nd even to 22 lb. per square inch. In from 1.5 to 2 hours the charge is smelted, whereupon concentration of matte to 40-50 per cent. Cu begins; this lasts about 2 hr., making the total time required 3.5 to 4.5 hr. for the treatment of 10 tons of charge. The content of the converter is discharged into an open-hearth furnace of 15 tons capacity, lined with magnesite brick, to settle the matte which takes from 1.5 to 2 hr., whereupon waste slag with Cu < 0.5 per cent, is drawn from the surface of the bath. The open-hearth furnace serves also as a reservoir for matte to be blown in the Manhes con- verter for blister copper with Cu 99.5 per cent. The original furnace had a capacity of 210 cu. ft. and treated 10- ton charges; the latest furnace has a volume of 35o'cu. ft. and takes 20 tons of charge. The 198 METALLURGY OF COPPER work of the larger furnace is more regular than that of the smaller. The charge ought to contain S 25 per cent, in order to do good work. The magnesite lining is good for about 180 charges; as many as 459 charges have been smelted on one lining, which costs about $5000.00. It has been observed that slags when basic (irony) attack the lining more quickly than when acid. As MgFe204 does not form at a temperature as high as 1500° C./ the corrosion must be due to the eutectic CaOiFe2 O3 which readily fuses at 1 200° C. With a furnace of 350 cu. ft. capacity treating charges of from 20 to 25 tons there are required during the first 1.5 hr. 40 h.p.; during the next i-i-S hr., 80-100 h.p.; during the last hour, 150-160 and even 200 h.p. to furnish the necessary blast. The temperature of the gases measured in the flue t,^, ft. from the mouth of the converter ranged from 600 to 780° C; they contained from 2.5 to 11. 5, average 6.5, per cent. vol. SO2 and SO3; with gas showing 11. 5 per cent. vol. SO4 the per cent. vol. of SO2 was 9-9.5, and of SO3 2-2.5. At Sulitjelma, with two 20-ton furnaces, the cost of treatment is $0.91 per ton of ore, excluding air-compression and re- pairs; adding these two items gives a total of $1.25. Attention may be called to the tremendous power-consumption of the process, which makes it suited only for localities where water-power is extremely cheap, as, e.g., in Norway ($4.00 per horse-power year). As a matter of record may be mentioned the various attempts of Garretson^ to smelt sulphide copper ore and convert the matte formed in a single furnace. Similar experiments were made at Great Falls, Mont., in 1895^ with the so-called Charles Allen process of bessemerizing in the blast-furnace crucible, which also were unsuccessful. c. Partial Pyritic Smelting no. Partial Pyritic Smelting of Raw Sulphide Ore for Matte.— The essential requirements for entire pyritic smelting were, massive pyritic ore containing free Si02 and little AI2O3 and other bases. Such ores are not of frequent occur- rence; most sulphide copper ores contain pyritic material disseminated through a gangue which is likely to run high in AI2O3. They are treated raw by partial pyritic smelting, a process in which the lack of heat from the oxidation of in- sufficient FeS is supplied by the use of carbonaceous fuel, and, in addition, sometimes by preheating the blast. III. The Slag. — In pure pyritic smelting, the slags made are singulo silicates high in FeO and low in CaO and AI2O3; in the partial process they run high in SiOa, low in FeO, high in CaO, and are likely to contain considerable amounts of AI2O3. Examples of slag compositions are shown in Table 28. Temperatures of slags as flowing from the blast-furnace measured by Clevenger* cover a range of 1123-1261° C. The part AI2O3 plays in these 1 Hofman-Mostowitsch, loc. cit. 'Min. Ind., igo2, XI, 206; Eng. Min. J., 1908, Lxxxv, 776, 1063; 1909, Lxxxvm, 1266. ^ Church, Tr. A. I. M. E., 1913, xLVi, 426. ' Met. Chem. Eng., 1913, xi, 447. SMELTING OF COPPER 199 slags is often of great importance.^ In general, Vogt has found^ that in slags with less than Si02 43 per cent, and moderate amounts of AI2O3 (ratio, 3 CaR( = Fe.Mg.Mn) : AI2 > i), the AI2O3 wUl act as a base. Lloyd^ found in partial pyritic smelting that when the Si02-content of the slag exceeded 44 per cent., AI2O3 began to act as an acid and make the slag bad. Hofman^ has shown that if in the singulo-silicate with SiOz 32.10, FeO 35.90, CaO 32.00 per cent, and a formation temperature of 1x50° 'C, the Si02 is replaced by AI2O3, the formation temperature rises, that if the CaO is similarly replaced, the formation temperature falls slightly. If therefore both Si02 and CaO are replaced by AI2O3, the formation temperature is likely to remain constant. This corresponds in part to the practice usual with a high-CaO slag, of lowering the percentage of Si02 with an increase in that of AI2O3. Some furnace-men assume that AI2O3 acts always as an acid, add the Si02 and AI2O3, and figure their charge to form a bi-silicate slag; others neglect the presence of AI2O3, and figure their slag as being made up of a mixture of a bi-silicate of lime, CaO SiOz, and a singulo-silicate of iron aFeO. Si02. Table 35, by C. S. Witherell, gives the amounts of FeO and Si02 necessary for slags with from 10 to 25 per cent. CaO, for a range of the total of the constituents of from 90 to 100 per cent. Magnetite (Fe304) as a flux° is undesirable in partial pyritic smelting, as the reducing power of the ascending gas-current is too weak to complete the reduction to FeO by solid C; hence Fe304 is likely to enter the slag and reduce its fusibility and fluidity, as well as to be taken up by the matte. The presence of Fe203 in some slags is explained by Wright'* as due to the reaction FeS-f- ioFe203 = 7Fe304-fS02. According to Hofman' the percentage of Fe304 in a slag is governed by the percentage of Si02 to which it is indirectly proportional.* 112. Fuel and Blast. — In §104 the line of separation between entire and partial pyritic smelting was provisionally drawn at 3 per rent. coke. In Table 28 the amount of coke used in partial pyritic smelting is seen to be about 8 per cent, and upward of the weight of the charge. The less the weight of coke required to furnish the heat necessary for smelting, the larger will be the amount of O available for FeS in a given volume of blast. The percentage of coke required can be diminished by heating the blast, « with the result that the pyritic effect i£»g. Min. J., 1908, Lxxxvi, 107, 177 (Heberlein) ; 264, 270 (Shelby); 483 (Bretherton); 730 (Koch); iiii (Hooper); 1909, Lxxxvii, 222 (Beardsley). ' Smith, Eng. Min. J., 1910, xc, 1261. ' Peters, Min. Ind., 1909, xviii, 245. * Tr. A. I. M. E., 1899, XDC, 717; "General Metallurgy," 1913, p. 461. '■Eng. Min. J., 1907, Lxxxin, 817 (Wells); 1909, txxxvii, 962 (Bennetts); ucxxvlii, 367 (Rizo); 742 (Shelby). » Eng. Min. J., 1913, xcvi, 825. ''Min. Ind., 1913, xxn, under "Lead." 8 See § 119, Fe304 in Matte; and § 176, Great Falls basic converter. ''Eng. Min. J., 1902, ixxiii, 525 (Grabill); 1906, Lxxxn, 598 (Kiddie), 698 (Giroux); 1907, Lxxxii, 692. Min. Sc, 1908, lvii, 46 (Parry). Min. J., 1908, Lxxxiii, 113 (Moore). Ekckochem. Met. Ind., 1906, iv, 420 (Giroux). Tr. A.I.M. E., 1904, xxxiv, 422 (Bretherton) See Hofman, "General Metallurgy," §321 and following. 200 METALLURGY OF COPPER 8 H o CO o o to t*3 lO ro lO fO *d f*3 00 ro ro ro ro 00 ro 00 PO s di PO o i in d O o o O fO »o H o lO 00 ■^ PO 00 PO Oi PO 00 PO o PO to PO 4 ro 00 i 00 ro ro ro to ro -d ro ro H ro NO PO 00 PO lO 00 ro o o> PO Oi ro o. d. PO ro O O 4 lO M to M O lO »o 00 ro I- 00 lO ro di PO »o OO PO o PO NO ro 4 PO pj PO o 00 M 4 ro 4 ro M ro ro M ro "d ro o PO d ro o PO ro o PO w OO ro to 00 PO o d 0\ 00 O »o H ro d 00 ro PO ro PO PO 00 PO 'it PO d> 00 o o c3 H M M H 1 lO o ro ro ro ro fO -Ti- ro 00 Tl- PO PO lO ro 00 to PO ■d ro -o PO PO PO o 00 fO O % ^ r* 't to lO 5 O ro M i 00 ro ro PO IN 4 PO OO PO M PO a o 00 o CO 00 O CO 6 o l-l fO ro 00 ro ro ro ro 00 ro ro ro ro to ro PO M ■d PO lO >d ro o ro PO O • 3 d 00 ro ro 00 lO PO PO PO 00 PO H ro o 00 O >o to PI 00 O CO 6. d d ro ro 00 H ro ro ro ro ro ro ro rj- PO O »o ro lO lO PO Oi »o PO PO 00 PO I i ■* i OO 00 ro ro ro 00 ro ro ro H PO O O PO lO 00 r- to M 4 M s 6 en OO d> ro r- d d ro ro ro H ro M fO ro O ro ro lO ro PO o PO ■<*■ PO PO PO ro 00 to ro xd ro O t^ 1 ro 00 00 ro ro lO fO O fO PO PO o d ro OO ■N t- >o C4 00 00 o CO OO Oi ■ d ro d ro O M ro lO ro o ro ro ro PO ro 00 PO ro Tl- PO i> 4 PO to PO to PO O 3 00 d 00 ro -^l- o ro ro O ro ro PO d ro OO CI I> »o PO 4 00 PI o 00 "(5 6 00 00 o> lO d ro o 6 ro fo 00 PO 00 PO ro ro PO PO 4 PO PO o to ro ^ a ^ d 00 ro O o s H ro ro H ro p< o PO OO OO »o PO 4 PI O d o o o M O o o H o »o O -d O o 00 O di H o d O o o fO o N o to SMELTING OF COPPER 201 in the furnace will be increased. Heating the blast to 200 or 300° C. will make a marked difference in the degree of concentration.^ Experiments in sub- stituting oil for coke have been made at Van Anda, B. C, by Kiddie^ and at Tucson, Ariz., by Waters.^ 113. Chemistry. — Detailed investigations into the chemistry of partial pyritic smelting are lacking. The blast entering through the tuyeres will find the smelt- ing zone higher and more narrowed than in a reducing fusion, and lower as well as less narrowed than in pure pyritic work. It will oxidize mainly coke and in a less degree some of the FeS that is trickling downward. Above the smelting zone the ascending gas current will consist of N,C02,CO,S02 and some free O; the oxidizing power of the O will be greatly weakened by the presence of CO2 andSOa. The O will act upon FeS; whether its oxidizing power is strong enough to have a converting effect as in the smelting zone or only a roasting effect, will be determined by the temperature and by the volume of the other gases in the current. Thus Wright^ found that combustion of S ceased in an atmos- phere containing 12 per cent. vol. SO2. The gas-current will be charged with S-vapor at 700° C, and this will burn at the surface of the charge. It may be noted that in most partial pyritic furnaces the volume of blast forced into the furnace is so large that the top of the charge is usually at a red heat. In other words, in order to obtain any considerable pyritic effect in the presence of coke, it is essential to have a large volume of air that unconsumed O may reach the region above the smelting zone and have there some oxidizing effect upon FeS. The overlying bed of charge is not sufficiently deep to take up most of the heat, which causes the top to become heated. The atmosphere in the furnace will have hardly any reducing power whatever. This is shown by the two gas analyses of Hermann^ which showed CO2 (from fuel) 8.3, CO2 (from limestone) 2.6, SO2 2.5, CO 2.15, O 8.00 per cent, vol.; and CO2 (from fuel) 14.1, CO2 (from limestone) 3.1, SO2 3.5, CO 3.2 per cent, vol., O n.d. Fur- nace gases from the Copper Queen smeltery contained 10 per cent. vol. O. D.unn* gives the following analyses: SO2 1.274, SO3 0.086, CO2 6.493, H2O 3.490, AS2O3 0.0091, O 10.18, N 78.13 per cent. vol. In the descending ore-charge the changes are probably the following: There will be first a loss of hygroscopic and chemically combined H2O; then the charge will become permeated by S-vapor, which has a sulphurizing effect; farther down O, CO2, and SO2 will act upon coke, the O alone upon FeS, which may be converted into Fe2Si04 or only into Fe^O^ to combine later on with SiOj. As the charge reaches the smelting zone proper, fusion will take place in the same way as in the regular reducing smelting. 114. Management and Results.— The characteristics of the furnace are a ■ Peters, Min. Ind., 1908, xvil, 293. ^Eng. Min. J., 1911, xcii, 434. ' Op. cit, 1913, xcvi, 203. * "Pyrite Smelting," p. 227. ' West. Chem. and Met., 1905, i, 145. « Tr. A. I. M. E., 1913, XLVi. 202 METALLURGY OF COPPER hot top and a cool tuyere-region on account of the large volume of blast and the small amount of coke used to obtain a pyritic effect. The tuyferes, therefore, have a tendency to become dark and hard, with the result that they have to be punched more or less continuously in order to get the air into the furnace; in fact, many plants have a special punching crew, which often has to use an air- hammer drill in its work. Results of operations are given in Table 28. 115. Calculation of Charge.' — The calculation of a charge with the aim of forming a slag of a certain degree of silication has been given in § 102. In many smelteries it has become the custom to run the blast-furnace on a slag containing definite percentages of SiOa, FeO, CaO, and other bases, as is usually the case with the typical slags made in the lead blast-furnace. As the calculation of such a charge differs from that given in § 102, it is carried through by the method with simultaneous equations, which is the most accurate. A charge is to be calculated which is made up of ores, fluxes, and fuel given in Table 36. Table 36. — Ores, Fluxes, and Coke for Blast-furnace Charge Charge-component Si02 Fe(Mn)0 Ca(Mg, Ba)0 AI2O3 S Cu Ash First-class ore Coarse concentrate. . Limestone 5° 20 3 17 33 I I SO 8 S 17 30 6 10 Coke Coke-ash 50 25 3 20 The slag desired is: SiOa 41, FeO 22, CaO 23 (AI2O3 8) per cent. ; the matte shall contain Cu 45 per cent.; the coke used shall be 8 per cent, of the weight of the charge, which is to weigh 1000 lb. A summary of the calculation is given in Table 37. (i) Materials Required to Produce 100 lb. Slag. — ^Let x = \h. first-class ore, y=\h. coarse concentrate, 0=lb. limestone; then 0.08 (x+y-f 2)= necessary lb. of coke which carry 0.1X0.08 {x-\-y-\-z) lb. coke-ash. (2) Iron Required for Matte. — The matte is to assay Cu 45 per cent.; it will contain Fe 27.8 per cent, which corresponds to FeO 35.7 per cent. The first-class ore contains Cu 6 per cent, which requires FeO 4.8 per cent, for matte and leaves FeO 17—4.8 = 12.2 per cent, to be slagged. The concentrate contains Cu 10 per cent, which requires FeO 7.9 per cent, for matte and leaves FeO 33 — 7-9 = 25.1 per cent, to be slagged. 1 Furman, School Min. Quart., 1896, xviii, i. Barbour, Min. Sc. Press, 1909, xcix, 664. Mostowitsch, Metallurgie, 191 2, ix, 559. SMELTING OF COPPER 203 3 u '53 si 00 10 • CO - 10 fe s ru g vO H c/: 4-1 1 00 -^ 10 M W r^ CO CO 0) ri Ah Oj r^ H ro • 00 ■ w i < 4-) f 1" '^. « M ■ M % M • CO 1 00 10 • 'i? m 6 .1 Tt- 00 " CO CO CO fO M M 10 H 8 U f^l " H fO 10 • « • II CO CI i !> ■^ CO \o 10 w M ~0 lo <0 'O Ov CO ■* CO H H S a M fO ,1* CS (S en 1. CD I 00 M d 2 So t^ u 2 204 METALLURGY OF COPPER (3) Ore, Flux, and Coke-ash Required to Furnish S1O2, FeO, and CaO for 100 LB. Slag. The Si02 in 100 lb. slag = o.5oa;+o. 203/ +0.032+0.50X0.008 (x+y+z). The FeO in 100 lb. slag = o.i22X+o.2Siy+o.2sXo.oo8 (x+y+z). The CaO in 100 lb. slag^o.oix+o.oiy +0.502+0.03X0.008 (x+y+z). The three simultaneous equations are 0.504a; +o.204y +0.034Z =41 (1) 0.124X +0.2533/ +0.002Z =22 (2) o . OI024.T+0 . oio243'+o . 50024Z =23 (3) Solved, they give: x = 54.o lb. first class ore ^ = 60.3 lb. coarse concentrate 2=43.7 lb. limestone Total= 158.0 lb. ore and flux for 100 lb. slag. (4) Reference to looo-lb. Charge. — To find the amounts of each of the charge-components required in a looo-lb. charge, each has to be multiplied by a factor w; i58»i= 1000, w = 6. 329. This gives: 54.0X6.329 = 342 lb. first class ore 60.3X6.329 = 382 lb. coarse concentrate 43.7X6.329 = 276 lb. limestone Total = 1000 lb. charge (5) Proof of Calculation. — From Table 37, 1000 lb. charge contains 58.7 lb. Cu which corresponds to 130.5 lb. 45-per cent, matte; this amount of matte requires 46.6 lb. FeO; deducting 46.6 from 186.1 FeO present, leaves 139.5 lb. to be slagged. There are present in 1000 lb. charge 259.7 lb. Si02; the factor for reducing this figure to 41 is 259.7 ^=4^> ^ = 0.158. Multiplying the totals of Si02, FeO, CaO, and AI2O3 of Table 37 entering the slag, by this factor gives: Si02 41, FeO 22, CaO 23, AI2O3 7.6, the desired ratio of the slag-components. (6) Pyritic Effect. — The 130.5 lb. matte produced from 1000 lb. charge contain 130.5X27.2=35.5 lb. S. There are present 172.7 lb. S, hence 172.7 — 35.5 = 137.2 lb. = 79.4 per cent, have to be burned off. 116. Thermal Balance Sheets of Some Partial Pyritic Smelting Operations. — Details of two thermal balance sheets are given, representing sulphide-ore treatment and matte-concentration as carried out by the Ducktown Sulphur, Copper, and Iron Co., Isabella, Tenn., the data having been furnished by Mr. W. F. Lamareaux. There is added for sake of comparison a condensed balance sheet of the partial pyritic smelting of ore by the Washoe plant of the Anaconda Copper Mining Co., furnished by Mr. E. P. Mathewson. These heat balances furnish an insight into the manner in which the heat has been generated, in which it has been utilized by chemical action and fusion, and in which it has been lost by radiation, convection, escaping gases, etc. SMELTING OF COPPER 205 Before the thermal balance sheet can be cast, it is essential to prepare a theo- retical balance sheet of the materials. I. Partial Pyritic Ore Smelting. — The basis of calculation chosen is that of 1000 kg. charge with 50 kg. coke. The ultimate analysis of the ore is, Cu 2.55, Fe 26.8, S 17.27, SiOj 28.38, CaO 8.11, MgO 3.83, AI2O, 3.39, Zn 2.93, 0,C02, etc., 6.74 per cent. The rational analysis calculated from the ultimate and the known character of the mineral constituents shows that the charge is composed of: chalcopyrite (CuFeSa) 7.4, sphalerite (ZnS) 4.4, pyrrhotite (FeySs) 33.5, biotite (AI Fe)2Si40i6) 12.2, actinolite [Ca(MgFe)3Si40i2] 17. i, cafcite (CaCOs) 10.4, quartz (Si02) 14. i, Undetermined 0.9, Total 100 per cent. The matte produced in the smelting had the following composition: Cu 16.0, Fe 49.8, S 24.9, SiOa 0.8, CaO 0.3, Insol. 2.1, Undetermined 6.1 per cent. In the theoretical balance-sheet of materials given in Table 38, it has been assumed, (i) that all the Cu has entered the matte, and that any Cu found in the slag is present as a matte pellet; (2) that the weight of the matte may be cal- culated from its analysis and the weight of the Cu in the charge, only Cu, Fe, Zn, and S being assigned to matte, the rest to slag; (3) that of theZn not present in the matte, one-half has entered the slag as ZnS, the other has been carried off as ZnO in the gases. Of the materials entering the furnace and placed on the debit-side, there remain to be determined the O, or air necessary for oxidation of constituents, and the accompanying moisture. The O required by the charge is: (i) CuFeSz. — ^The whole is assumed to enter the matte unchanged. (2) Fe to FeO.— Of the 202.0 kg. Fe furnished by 335.0 kg. FejSa, 57.1 kg. enter the matte and 144.9 the slag; all thp Fe of the coke-ash, i.e., i.i kg., enters the slag, or 144.9+1.1 = 146 kg. Fe enter the slag as FeO. These require (Fe:0 = s6: 16 = 146 :x) 4i-7 kg. O. (3) S to SO2.— Of the 14.5 kg. S furnished by 44 kg. ZnS, 6.3 kg. are oxidized; of the 133 kg. S furnished by 335 kg. FevSs, 120.7 kg. are oxidized; all the S in the coke, 0.8 kg., is oxidized; or 6.3 -fi 20.7 -f 0.8=127.8 kg. S are oxidized. These require (S:02 = 32 :32 =127.8 :3') .... 127.8 kg. O. (4) C to C02.^— All of the 42.0 kg. C of the coke are oxidized; they require therefore (0:02 = 12:32=42.0:2) 1 12.0 kg. O. (5) Zn to ZnO.— Of the 29.5 kg. Zn furnished by 44 kg. ZnS, 13.0 kg. are oxidized. They require (Zn :0 = 65 : 16 = 13.0: w) 3.2 kg. O. (6) The total O theoretically required is therefore 41.7-I-127.8-I-112.0 +3.2 ■ 284.7 kg. O. and the accompanying N, 953, corresponding to 756.4 cu. m. (7) The volume of gases is 127.8 kg. S + 127.8 kg. 0=255.6 kg. 802 = 88.7 cu. m.;42.o kg. C+ii2.okg. 0=154 kg. CO2 from coke, and 45.8 kg. CO2 from limestone, or 199.8 kg. C02= loi cu. m. This gives as total volume 756.4 N+ 88.7 SO2+101.0 002 = 946.1 cu. m. But the waste gas carries 8 vol. excess O which corresponds to 40 vol. excess air. The above 946.1 cu. m. form therefore only 60 per cent, of the true volume, which is 1577 cu. m. ato° C. and76omm. Hg. ' The gas contains free O, hence no CO can form. 2o6 Table 38.- METALLURGY OF COPPER -Theoretical Balance Sheet of Materials of One Ore Charge of iooo kg. Debit Credit Mineral Per cent. Wt., kg. Constit- uent Per cent. Wt., kg. Matte, kg. Slag, kg. Gas, kg. CuFeSz 7-4 74. u Cu 34-6 25.6 2S.6 Fe 30. 5 22.6 22.6 S 34-9 25.8 25.8 ZnS 4-4 44. u Zn 67 29.5 3.4 13.1 13. u S 330 14s 1.7 6.5 6.3 FerSs 33. 5 335.0 Fe 60.4 202.0 57.1 144.9 S 39-6 133.0 12.3 120.7 (AlFe,)2SiiOn 12.2 122.0 Ore Charge, AhOa 27.8 33-9 33.9 1000 kg. SiOi 39-4 48.1 48.1 FeO 32.8 40.0 40.0 Ca(MgFe)jSi40i2 17. 1 171. CaO 13-4 22.9 22.9 MgO 22.4 38.2 38.2 FeO 9.0 15.4 15.4 Si02 S5-2 94-5 945 CaCOa 10.4 104. u CaO S6.0 S8 2 S8.2 CO2 44" 45.8 45.8 Si02 14. JL 141.0 141. . X 0.9 9.U 9.0 Fe 2.3 I . I I.I S 1.6 u,8 8 Coke, Si02 8.4 4.2 4.2 SO kg. AI2O. 3.6 1.8 1.8 c 83.9 42.0 42.0 X 0.2 O.I Blast 23. u 478.0 41.4 436.6 (dry) 2079 kg. N 77.0 1601.U 1601.0 Water, 31.7 kg. In charge 10. In blast 21.7 21.7 Totals., 3160.7 148.5 714.^ 2298.0 1 1 1 1 SMELTING OF COPPER 207 (8) Volume of Blast.— The volume of O necessary to form CO2 and SO2 is 189.7 cu. m., that to form FeO and ZnO, 31.0 cu. m.; the accompanying N and excess air give 1387.3 cu. m.; hence the volume of blast at 0° C. and 760 mm. Hg is 1608.0 cu. m. = 2079 kg. (9) Moisture in Charge and Blast. — ^The charge has i per cent or ro kg. H2O, the blast 13.5 g. per cubic meter, or 0.1608X13.5 = 21.7 kg., all of which passes off with the gases. In casting the thermal balance for a charge of 1000 kg. ore and flux, and 50 kg. coke, given in Table 39, the in-coming heat is placed in the debit column, the out-going in that of the credit. The details of the calculation are as follows: {A) Burning C to CO2, 8100 Cal. per kg. C: 42X8100 = 340,200 Cal. {B) Burning S to SO2, 2164 Cal. per kg. S: (6.3 kg. S from ZnS) + (i2o.7 kg. S from Fe7S8) + (o.8 kg. S from coke) = 127.8 kg. 127.8X2164=276,560 Cal. Table 39. — Thermal Balance for One Ore Charge or 1000 kg. Debit Credit Kilogram calories Per cent, of total Kilogram calories Per cent, of total (A) Burning C to CO2 340,200 276,560 169,968 16,96s 31,816 18,554 5,250 9,744 39-2 31.9 195 2,0 3.6 2. I 0.6 (I) Reduction of CaCOa (J) Reduction of FeS 46,991 62,104 8,600 33.413 232,11s 354,633 131,201 5.4 (B) Burning S to SO2 7.1 (K) Reduction of ZnS (D) Burning Zn to ZnO (L) Heat in matte 3.9 (E) Formation of aFeO.SiOj (F) Formation of CaO.SiOz (G) Sensible heat in charge (H) Sensible heat in blast (M) Heat in slag 26 7 (0) Radiation and conduction (by difference). ISi 869,057 100. 869,057 100. (C) Burning Fe to FeO, 1173 Cal. per kg. Fe: The 144.9 kg. Fe from Fe7S8 entering the slag are oxidized to FeO. 144.9X1173 = 169,968 Cal. (Z)) Burning Zn to ZnO, 1305 Cal. per kg. Zn: The 13.0 kg. Zn not entering matte and slag as ZnS are oxidized to ZnO. 13.0X1305 = 16,965 Cal. (£) Formation of 2FeO.Si02, 154 Cal. per kg. FeO: The 33.9 kg. AI2O3 in (AlFe)2Si40i6 are assumed to enter the slag as Al203.2Si02. This silicate contains 39.9 kg. Si02, leaving 48.1-35.9 = 8.2 kg. Si02 which are assumed to be combined with FeO as 2FeO.Si02 The 8.2 kg. Si02 require 19.7 kg. FeO, leaving 40.0-19.7 = 20.3 kg. 2o8 METALLURGY OF COFFER FeO uncombined with SiOj. The 144.9 kg. Fe from (C) correspond to 186.3 kg. FeO; the total FeO =20.3 + 186.3 = 206.6 kg. =154X206.6 = 31,816 Cal. (F) Formation of CaO.Si02, 318.8 Cal. per kg. CaO. The 104.0 kg. CaCOs furnish 58.2 kg. CaO. 58.2X318.8 = 18,554 Cal. (G) Sensible Heat in charge at 20° C, spec. heat = o.2S app. 1050X0.25X20=5250 Cal. (H) Sensible Heat in Blast at 20° C, spec, heat = 0.303 app., volume of blast =1,608 cu. m. 1068X0.303X20 = 9744 Cal. (/) Dissociation of CaCOa, 1026 Cal. per kg. CO2. 45.8X1026 = 4,6991 Cal. (/) Dissociation of FejSs, 428.6 Cal. per kg. Fe. The Fe entering the slag from FeySs has to be set free before it is oxidized. 144.9X428.6 = 62,104 Cal. (K) Dissociation of ZnS, 661.5 Cal. per kg. Zn. The Zn entering the gas from ZnS has to be set free before it is oxidized. 13.0X661.5 = 8600 Cal. (L) Heat in Matte, 225 Cal. per kg. app. 148.5X225=33,413 Cal. (M) Heat in Slag, 325 Cal. per kg. app. 714.2X325 = 232,115 Cal. (iV) Heat in Gas, temperature 600° C. Gas analysis=S02 5.4, CO2 6.3, 8.0, N 80.3 per cent. vol. Volume of gas, at o°C and 760 mm. Hg, 1577, cu. m. divides as follows: S02= 1577X0.054 = 85.2 cu. m.; C02= 1577X0.063 =99.4 cu. m.; N and 0=1577X0.883 = 1392.4 cu. m. The mean spec, heats between zero and 600° C. are SO2, 0.54; CO2. 0.502; O and N, 0.3192, hence the total heat in the gases, (85.2X0.54+99.4X0.502+1392.4X0.3192) X6oo = 324,2i7 Cal. There has to be added the heat contained in 31.7 kg. H2O. and 13.0 kg. ZnO. The heat of evaporation of H2O at 0° C. = 606.5 Cal. per kg., hence 31.7X606.5 = 19,226 Cal.; the mean spec, heat of the gas between zero and 600 °C. = 0.531, hence 31.7X0.531X600=10,098 Cal., or the total heat in the water vapor= 19,226+10,098 = 29,324 Cal. The mean spec, beat of ZnO between zero and 600° C. = o.i4, hence 13.0X0.14X600=1,092 Cal. The total heat in the gas is therefore 324,217 + 29,324+1,092 = 354,633 Cal. II. Partial Pyritic Matte Concentration. — ^The mode of procedure in- calculating a thermal balance was exactly the same as the one followed in the first case with ore-smelting. The calculation is based upon the concentra- tion of 1000 kg. matte (Cu 16.0, Fe 49.8, Zn 2.1, S 24.9, Undet. 7.2 per cent.) with 293 kg. quartz and 50 kg. coke to converter-matte containing Cu Si.i) SMELTING OF COPPER 209 Fe 22.2, S 23.3, SiOz 0.21, Zn 0.8, Undet. 2.5^ per cent. Table 40 gives the theoretical balance sheet of materials of one charge of 1000 kg. matte, and Table 41 the thermal balance sheet. Table 40.— Theoretical Balance Sheet of Materials op Ore Charge of iooo Kg. Debit Credit Per cent. Weight, kg. Matte, kg. Slag, kg. Gas, kg. Matte, IOOO kg. Cu :. 16.0 49.8 2. 1 24,9 7.2 100. 83.9 ?-3 1.6 8.4 3-8 77-0 23. u 160.0 498.0 21.0 249.0 72.0 293.0 41.9 I. 2 0.8 4-2 1.9 2213.0 661 .0 10. 30.0 160.0 69s 2.S 73.6 Fe 428. s 9.2 4.6 72.0 293.0 Zn S 9-3 170.8 X Quartz, 293 kg.: SiOa Coke, so kg.: C 41.9 Fe 1.2 S 8 SiOj 4.2 X 1.9 2213.0 538.2 Blast, 2874 kg. : N.... 122.8 Moisture, 40 kg. : In matte 4257.0 305-6 935-5 3015-9 Table 41. — Thermal Balance for One Charge of iooo kg- Matte. Debit Credit Item Kilogram calories Per cent, of total Item Kilogram calories Per cent. of total Burning C to CO2. , Burning S to SOa. . Burning Fe to FeO, Burning Zn to ZnO Formation of slag. . Heat in charge Heat in blast 339.390 371.342 504,038 12.137 44.075 6,71s 17.416 1. 29s. 113 26.2 28.8 38.9 0.9 3.4 o-S J-3 100. Dissociation. FeS Dissociation, ZnS Heat in matte Heat in slag Heat in gas Radiation, conduction (diff.) 183. 6SS 6. 152 68,760 304,038 471. 41S 261,090 14 2IO METALLURGY OF COPPER III. Partial Pysitic Ore Smelting at Washoe Smeltery, Anaconda Copper Mining Co. — ^The thermal balance sheet given inTable42 is one figured to a basis of looo kg. charge from the oflftcial sheet funished by Mr. E. P. Mathewson. Table 42. — Thermal Balance Sheet for One Ore Charge of iooo kg. Anaconda, Mont. Credit Iten Kilogram calories Per cent. of total Kilogram calories Per cent.of total Burning coke Burning S Burning Fe Burning Zn Formation of slag. . Brought in by blast 569,948 138,952 80,461 8,88r 77,158 8,599 64. 5 lS-7 9-1 1 .0 8.7 883,999 Dissociation, CaCOs Dissociation, PeS Dissociation, MgCOs.. Heat in gases Heat in l!ue dust Heat in matte Heat in slag Heat in cooling water. . Loss by radiation, (diff.) 121,710 7,086 3.804 459,953 13.449 30,311 200,773 14.576 32,337 883,999 13.8 0.8 U.4 52.0 IS 3.4 22.7 1-7 3.7 lOQ.O One striking feature of the two theoretical balance sheets of materials, given in Tables 38 and 40, is that the weight of the gases produced greatly exceeds that of the charge fed. Of the three thermal balances, the data of Ducktown, Tables 39 and 41, show that more heat is derived here from oxidation of Fe and S, than at Ana- conda, Table 42; also that their losses of heat by radiation and conduction is very much larger. The three thermal tables show, that the largest part of the heat generated in the furnace is carried off by the gases, and that the slag follows next in order. 117. General Smelting Operations.— These are the blowing-in, the regular work on the feed- and fiu-nace-floors, and the blowing-out. In starting, it is advisable to make up a charge which will run easily and furnish a matte with Cu 30-35 per cent., as this runs hot and heats up the crucible and the fore-hearth. When the furnace runs well on a slag that is easily formed and on a high percentage of coke, the change is made to" the kind of charge it is the intention to run, be the process reducing, pyritic, or partial pyritic smelting. The following gives in detail two modern examples of blow- ing-in furnaces in partial pyritic smelting. At the smeltery of the Tennessee Copper Co., Geo. A. Guess^ used to proceed as follows: A new crucible is warmed for 24 hr. by burning wood. When warm, it is cleaned out, and light wood (scrap boards, broken lumber), charged to reach to the tuyeres. With an old furnace the bottom is only 6-8 'Private Communication, August, 191 2. SMELTING OF COPPER 211 in. below the tuyeres, hence much less wood is required than when the cruci- ble is new. The wood is ignited from end to end with oil-soaked waste; no blast is used; the necessary air enters through the tuyere openings. When the wood burns freely, more is charged to reach well above the tuyeres; care is taken that the wood lie flat, i.e., that there shall be no pieces of cord wood pointing upward. If the wood is dry, no blast is needed, but if wet, some air is turned on which is made to enter through alternate tuyeres. As soon as the bed of wood is burning freely all over, coke is charged, about 40 lb. per square foot of hearth area. Some blast is now necessary; as soon as the coke appears to be well ignited all over, the plugged tuyeres are uncovered, and air with a pressure of about 2 oz. turned on. The breast of the furnace has not been closed; an open breast assists in blowing out ashes, etc. When the coke is red on top, the furnace receives its charges in quick suc- cession. The blowing-in charges, six in number, differ from the normal, in that they are made less siliceous and carry 50 per cent, more coke. As soon as the first charge has been fed, the blast-pressure is raised to about 8 oz. Bits of charcoal, coke, etc., are blown with the flame out of the breast, and the bottom is cleaned and heated. This continues until fluid slag with some matte has trapped the blast, which happens in from 30 to 40 min. after the first blowing- in charge has been given. Matte and slag now overflow into the preheated settler; the blast is gradually increased until, in about 10 hr. after the first blowing-in charge, full blast has been put on. It will be noticed that no slag is used in blowing-in an ore-furnace; with a matte-concentrating furnace some slag is used at first, although it is not abso- lutely necessary. At the Garfield smeltery A. E. Wells^ used the following procedure: The bottom of the crucible is lined with about 12 in. of silica well tamped down. In order to dry it, a slow wood fire is kept going for 18 hr., and then a brisk fire 6 hr. A new settler is dried and warmed with a wood fire for 48 hr. When crucible and settler are warm, the furnace is filled with scrap wood to reach about 2 ft. above the tuyeres, the wood is kindled and the furnace filled; the wood burns with natural draft while the blowing-in charges are dropped. On top of the wood are spread 6000 lb. coke followed by nine blowing-in charges of 6000 lb., consisting of sulphide ore (>f in.) 2000 lb., converter slag (fist-size) 3000 lb., limestone 1000 lb., coke 600 lb. The blowing-in charges make a matte with Cu 25 per cent, and a slag with Si02 36, FeO 45, CaO 14 per cent. As soon as the blowing-in charges have been fed, blast is put on (15,000 cu. ft. per minute, engine displacement), which gives at the tuyeres a pressure of 20-25 oz. A flame is allowed to escape for a few minutes through the connecting-hole in the breast-block to blow out half-burnt wood and coke which might obstruct the passage of the slag. When slag begins to collect around the connection-hole, this is loosely plugged with a clayey brasque, and a 2-in. bar pushed through the latter. In about 15 min. the slag has risen sufficiently to trap the blast, the bar is pulled out, and the slag. 'Private Communication, August, 1912. 212 METALLURGY OF COPPER matte allowed to overflow into the settl er. The bio wing-in ch arges are gradually replaced by ore-charges; one of the former is followed by two of the latter; and the furnace is run for 2 hr. on this mixture. The charge-column is now raised to its normal height, and the blast is increased to 21,000 cu. ft. air per minute, which raises the pressure to about 35 oz. The regular work of the furnace is given in Table 28. The operations such as feeding of charge ( § 93) and tapping of matte and handUng of slag (§94) have already been discussed. The work on the feed-floor and furnace-floor has been indicated in §§93 and 94. In blowing-out, slag- charges are substituted for ore-charges until most of the ore-charge has been smel- ted. Charging is stopped, the charge slowly sinks, the volume of blast admitted is lowered. When the charge has sunk to about the lower tier of jackets, the blast is stopped, the tuyere valves are closed, the furnace is tapped clean, the breast-jacket is removed, and the material remaining in the furnace raked out. When the flow of slag ceases, the contents of the fore-hearth are tapped. Products 118. Products of the Blast-furnace. — The regular products are matte, speise, flue-dust, and gases; the irregular products, wall accretions, hearth accre- tions, furnace-drawings, and refuse. 119. Matte.' — ^In table 43 are given analyses of mattes with increasing con- tents of Cu selected from a collection published by Keller.^ The leading con- stituents of this intermediary product are Cu, Fe, and S; the other compo- %FeS 100 90 80 70 60 50 40 %Fe 10 20 30 40 50 60 30 70 20 10 %FeS 80 90 100 qoFe Fig. 231. — Alloy series FeS-Fe. nents, such as Ni, Co, Zn, Pb, Bi, Sb, As, Se, Te, Ag, Au, etc., are of minor import. There will be considered the relations of Cu-Fe, CU-CU2S, Fe-FeS and Cu2S-FeS. The mefals Cu and Fe (p. 16) form solid solutions within a range of 96.5 per cent Fe+3-S per cent. Cu and 97 per cent. Cu-l-3 per cent. Fe. ' In England the word "Regulus" is used synonymously. 2 Min. Ind., 1900, ix, 243; see also Channing in Rickard's "Pyrite Smelting," p. 263. SMELTING OF COFFER Table 43.— Analyses of Copper Mattes 213 Source Cu per cent. S per cent. Pe per cent. FejOi per cent. Ni per » cent. Co per cent. Zn per cent. Pb per cent. Elizabeth M. Co., Vt Parrot, reverb, furn 21.36 36. IS 49.02 49.17 49.34 55 00 53.73 54-89 57.83 60.76 61 .42 61.52 22.95 23.88 22.78 19.45 23.96 23.17 23.36 22.47 23.2s 22,52 41.03 24.97 23.86 22.79 22.44 13.8s 19.49 20.25 15.28 11.43 1450 13.68 10.44 8. SI 0.0020 0. 1984 0.24 0.0226 Le Roi Mine, B. C 0.0383 0.0436 0.0332 0.0222 0.09 0.77 Ducktown, Tenn B. & M. Co., reverb, furn . . . Jerome, Aria 2.58 1.24 0,26 0.34 i:.09 2.41 Silver City, N. M O.03S4 0,0498 0.0341 0.0240 o.ooso 0.007^ 0.0034 Copper Queen, blast-furn. . . . 0.1178 Mountain Copper Co., Cal. . . Anaconda, reverb, furn B. & M. Co., blast-furn 1.13 0.5900 U.0370 Santa Rosalia, Mexico O.064S 0.4140 0.2333 Table 43. — Analyses of Copper Matt-es-^{C ontinued) Source Elizabeth M. Co., Vt Parrot, reverb, furn Le Roi Mine, B. C Ducktown, Tenn B. & M. Co., reverb, furn . . , . Jerome, Ariz Silver City. N. M Copper Queen, blast-furn Mountain Copper Co., Cal. .\ . Anaconda, reverb, furn B. & M. Co., blast-furn Santa Rosalia, Mexico : . Hi per cent. 0.0337 u. 0174 0.0 o . 0044 0.0014 u. 0420 o . 0049 o . 0008 Sb per cent. 0.0 0.0348 0.0206 0. lOIO 0.2693 0.0032 0.0232 0.0143 U.0790 0.1330 0.0032 As per cent. o . 004 I 0.0434 0.0 0.0480 0.0914 0.0 0.0171 U.0130 0.0450 o. 1280 0.0013 Te per cent. 0.0063 Se per cent. Ag oz. per ton 0.0 0.0021 .0474 I 0.1172 Trace .0088 I 0.0113 0.0060 .0112 I 0.0038 0.0042 .0 I 0.0 26.0 5-9 14.6 127.0 I. .i 6.0 13-4 60.4 18.2 Au oz. per ton 10.73 0.04 o.os 2.28 Trace o. 10 o.sr U.30 o.os Trace The saturation point of CU2S for Cu (p. 21) is 15 per cent.; and the eutectic CU-CU2S contains 3.8 per cent. CU2S+96.2 per cent. Cu. The freezing-point curve for FeS-Fe has been investigated by Tammann- Treitschke^ and Friedrich.^ According to the diagram of Friedrich, Fig. 231, the two components form an eutectic mixture of 15 per cent. Fe and 85 per cent FeS solidifying at 983° C; Fe can hold in solid solution less than 3 per cent FeS and FeS less than i per cent. Fe. The constitution of Cu2S-FeS has been traced through freezing-point curves by Rontgen/ Hofman-Caypless-Harrington/ Baykoff-Troutneff,^ and Borne- mann-Schreyer,^ and through other methods by Miinster,^ Bolles,^ Gibb-Philp^ ^Zi. anorg, Chem., 1906, xldc, 320; Metallurgie, 1907, iv, 54. ^Metallurgies 1910, vn, 257. ^ Metallurgie, 1906, iii, 479. * Tr. A. I. M. E.J 1908, xxxvin, 424. ^ Rev. MM., 1909. VI, 518. ^Metallurgie, 1909, vi, 619. ''Berg. Huttenm. Z., 1877, xxxvi, 195, 210, 219. « Tr. A, I. M. E., 1905, XXXV, 666. *• Op, cit.f 1906, XXXVI, 665. 214 METALLURGY OF COPPER FeS Eutectic 15 Fe+85 Fe. CuiS Saturation Point 'usS for Cu ^-Eutectic 3.8 C«+ 96.2 Cu CmjS Fig. 232. — Triaxial diagram of copper-iron matte. and Fulton- Goodner.i ^jj eutectic mixture found by some is denied by others, although it can be seen clearly in copper matte; similar disagreements exist as regards chemical compounds and solid solutions. The question requires further investigation. In plotting the accepted relations between Cu, Fe, and S in a tri-axial dia- gram, as was done by Baikoff-Troutneflf,^ but substituting the data of Friedrich for those of Tammann-Treitschke, Fig. 232 is obtained. This gives four fields: Field I represents the areas of S-Fe and S-Cu compounds which are dissociated above their melting-points; field II the stable FeS-Cu2S mixtures forming pure matte; field III the region in which there is in the liquid state a stratification of components belonging to II and IV; field IV a matte with large amounts of solid solution of Cu and Fe. The connecting lines are drawn straight, because there are as yet no data to show their accu- rate positions between the end- points. The smallness of field II shows how likely ordinary mattes are to contain metallics in which either Fe or Cu prevails. The presence of Fe304 to an extent of 10 per cent, in low-grade, and especially in pyritic matte,' which causes much trouble in the settler, is due to Fe304 in the charge, or to imperfect reduction of Fe203; it may be caused also by oxidation of the Fe in the matte, but not by that of FeS, as in normal pyritic operation this is presupposed to be oxidized to FeO and directly combined with SiOj. In abnormal pyritic work, however, it may be formed, if, e.g., there is a lack of SiOj.* As Fe304, with a specific gravity of 5.0 to 5.2, forms more readily with a low- than a high-grade matte, and as the specific gravity of matte increases with the Cu-content (Cu 13.62 per cent., spec. gr. 4.8; Cu 43.00, spec. gr. 5.18; Cu 60.22, spec. gr. 5.42; Cu 80.00, spec. gr. 5.55) Fe304 will enter low-grade matte and float on high-grade matte, and thence will be taken up in part by slag and carry Au into it. Thus at Blagodatny, Ural,« with a matte of SiOa 1.2, CujS 17.2, FeS 61.7, PbS 6.7, Fe304 12.4, As, Sb.Bi 0.6 per cent, Au 6.23 oz. and Ag 66.4 oz. per ton, the slag (SiOa 45-2, FeO 28.5, CaO 22.2, N.D. 4.1) assayed 1 Op. cit., 1908, xxxrx, 584. '^ Loc. cit. ^ Keller, Eng. Min. J., 1895, lx, 465. Larison, op. cit., 1909, lxxxvii, iiqs- Rizo-Patron, op. cit., 1909, lxxxviii, 367. Shelby, loc. cit., p. 742. Keller, Min. Ind., 1900, rx, 243. Gibb-Philp, Tr. A. I. M. E., 1906, xxxvi, 671, 1907, xxxvra, 913. Keller, op. cit., 1906, xxxvi, 837. ^See page 199. ' Private Communication by F. W. Draper, Nov., 1908, SMELTING OF COPPER 215 Au 0.03 oz. and Ag 0.59 oz. per ton., while in the absence of Fe304 the slag ran Au 0.003-0.013 oz. and Ag 0.50-0.75 oz. per ton. The silver content in the slag did not appear to be affected by Fe304; it varied directly with the assay of the matte whether Fe304 was present or not. Finely divided Cu/ so-called "moss-copper," is of frequent occurrence in matte assaying from about 30 to about 60 per cent. Cu. Fulton-Goodner^ noticed it in lo-per cent. Cu-matte. It has its origin in the insolubility of Cu in CU2S in the solid state (§ 21). Fulton-Goodner call attention to the fact that the Cu separates from solid matte when this is relatively cool, but still too hot to be held in the hand. They attribute the separation at this low temperature tentatively to the dimorphic point of CU2S, which occurs at 103° C. The crack- ing vertically of conical or hemispherical cakes of Cu-matte' upon cooling, if the Cu-content is much below 50 per cent., and horizontally if over 50 per cent., may be due in part to the separation of Cu (see also Ni-Cu matte). ^ The large needles of Cu found in some mattes, not to be confounded with moss-copper, are due to the reaction of CU2S upon CuO or Cu20.^ Matte is an excellent carrier of precious metals.® The leading reasons for this are that CuaS and Ag2S form solid solutions,' the curve showing a maximum depression at 677° C; that CU2S readily dissolves Au; that Cu easily alloys both with Ag (p. 22) and Au (p. 23); that Fe is a strong solvent for Au;* and that the same is the case' for AU2S3 in the presence of AgjS. Little if any solvent action has been noticed with FeS for either Ag2S''' or Au2S3;^^ and Fe has little affinity for Ag.^** The equilibrium diagram for Ag2S-FeS by Schoen^' shows an eutectic with 11 per cent. FeS freezing at 600° C, and a transformation at 175° C. characteristic for Ag2S. The presence of PbS and ZnS in matte will not materially assist the collection of Ag2S, provided CU2S is present, as both form eutectiferous alloys'^ in which the eutectic line extends to near the ordinates. > Plattner, Berg. HiiUenm. Z., 1855, xiv, 143. Hampe, op. cit., 1893, lii, 448. Palmer, Min. Sc. Press, 1906, xciii, 604. Gibb-Philp, Tr. A. I. M. E., 1906, xxxvi, 677. Larison, Min. World, 1907, xxvil, 550. 'Tr.A. I. M. E., 1908, xxxDC, 617. ' Bellinger, Min. Ind., 1894, in, 229. ■* Browne, School Min. Quart., 1894-9S, xvi, 297. ' Miinster, Berg. HiiUenm. Z., 1877, xxxvi, 220. « Bolles, Tr. A. I. M. E., 1905, xxxv, 666. Fulton-Goodner, op. cit., 1908, xxxrx, 584. ' Friedrich, Metallurgie, 1907, rv, 671. « Tr. A. I. M. E., 1886-87, XV, 767 (Spilsbury); 1889-90, xvm, 454, 457 (Pearce); 1900 XXX, 769 (Carpenter); 1905, xxxv, 666 (Bolles); Zt. angew. Chemie, 1907, Liii, 291; Rev. Met., 1908, V, 188; Metallurgie, 1907, rv, 469 (Isaac-Tammann's solid solution curve). ' Muir, Eng. Min. J., 1872, xiv, 56. Pearce, loc. cit. >» C. J. B. Karsten, "System der Metallurgie," Reimer, Berlin, 1832, v, 525. " Spilsbury, loc. cit. " Spilsbury, Pierce, Bolles, loc. cit. "Metallurgie, 1911, vin, 737. " Friedrich, Metallurgie, 1907, iv, 671, and 1908, v, 114. 2l6 METALLURGY OF COPPER In a copper matte CU2S is therefore the leading carrier of precious metals. The question is, how much Cu must be present to effect a complete collection. The consensus of metallurgists seems to be' that 0.5 per cent. Cu is sxifficient, provided the degree of concentration is not too great. Carpenter^ states that with little matte, this should contain 10 per cent. Cu, and that with much matte 2-3 per cent. Cu would be ample. Lang' calls attention to the formation temperature of the slag produced in smelting; if this be high, the matte will contain some metallic Fe, and this is a good collector for Au. 120. Speise — This is not often formed in the smelting of sulphide copper ores, as As and Sb are usually present in small amounts, and as most of the AS2S3 and Sb2S3 is readily eliminated, either as oxide in the roasting which pre- cedes a reducing fusion, or as sxilphide in the pyritic smelting which treats raw ore. Sometimes speise is purposely produced in the treatment of ores contain- ing Cu,Ni,As, and S in order to collect the Ni in the speise and the Cu in the matte; the speise, however, locks up considerable amounts of Cu. Some analyses of speise are given in Table 44. Table 44. — Analyses of Copper Speise Locality Cu Pb Fe Ni Co Zn Sb As Ag Au Bi Schmoellnitz Neusohl^ Brixlegg*. . . Brixlegg^. . . Pretoria" England'. . . 12.99 41.18 51-73 25-85 52-50 7.78 0.09 0.69 ,35 - 20 16.68 0.25 2.94 12.63 35-41 x-6s 22. 17 3.60 10. 20 40 0.09 0.13 60 09 0.04 0. 24 1. II 1.82 3-31 6-57 60.00 10.79 3-34 13-50 38.00 41.82 7.42 6. 10 2-75 11.04 2.00 18.56 0.36 0.03 o. o. I -59 0.02 0.06 175 08s 2.60 2.04 2.60 1.38 4-13 2.06 1.72 1.26 1.63 The treatment of speise consisted usually in a series of oxidizing roasts followed by reducing fusions, by means of which Ni with its great affinity for As is more and more concentrated, forming a nickel speise. More recently roast- smelting in the reverberatory furnace has become the favored method at Frei- berg, ^ Oker,9 and Brixlegg.!" The latest proposition is that of Guillemain," 1 Rickard, " Pyrite Smelting," p. 134. 2 Op. cit., p. 34. ^ Op. cit., p. 37. * Balling, C. A. M., "Metallhuttenkunde," Springer, Berlin, 1885, p. 192; Min. Eng. World, 1913, XXXVIII, 9. ^ Kroupa, Oest. Zt. Berg. Hiittenw., 1906, Lw, 73, 84. " Bettel, Eng. Min. J., 1891, Lii, 74. '.McMurtry, Tr. Inst. Min. Met., I9i3,xxii, 50. * Hubner, Gliick Auf., 1905, XLi, 6. Hofman, Min. Ind., 1905, xrv, 414. ° Huhn, Gluck Auf., 1905, xli, 1143; Min. Mag., 1906, xiii, 312; Min. Ind., 190s, xiv, 414 (Hofman); 1906, xv, 286 (Austin). " Kroupa, Oest. Zt. Berg. Hiittenw., 1906, liv, 73, 84. Austin, Min. Ind., 1906, xv, 286. » Metallurgie, 1910, vii, 595; Eng. Min. J., 1911, xci, 50; Min. Ind., 1910, xix, 450. SMELTING OF COPPER 217 in Slag 1.2 to blast-roast speise in a Huntington-Heberlein pot. The experimental results have been most satisfactory. Converting speise with an addition of about 50 per cent, copper matte in a basic converter has been successful, while convert- ing speise alone, has not. The electro-negative component of speise usually is As. In the analyses of Table 44, Sb prevaUs over As. There is an inter- esting record by Betteli of the collection of Cu and Ag in an antimonial speise, by smelting in a reverberatory furnace; the analysis is given in Table 44; another record is that of McMurtry.^ 121. Slag.— Compositions and Cu-contents of some blast-furnace slags are given in Table 28. About twenty years ago, foul slag from a matte concentrating blast- furnace^ of the Orford Copper Co., was partly freed from Cu by running it from the fore-hearth direct into one end of the ore- furnace a slight distance above the level of the tuyeres. This worked satisfactorily, as long as the ore-furnace ran smoothly, without any obstructions forming to check the inflow of the slag. With ordinary care this could be avoided, but when accidents did occur, there was no end of trouble; hence, the method has been abandoned. The Cu-losses^ are caused by imperfect settling of matte (due to lack of time and temperature, insufficient difference in specific gravity, gas-flotation, mushi- ness of matte, viscosity of slag), by solution of metal, oxide or sulphide in slag, and by scorification of copper (silicate, perhaps ferrite). The second cause was once thought to be so insignificant that it could be neglected. The experiments of Wanjukow^ prove that this factor has to be considered. The curves, Fig. 233, representing sesqui- and bi-silicates with 12 and 36 per cent. CaO, show how. Fig. 2ii.- -Relation of copper-content in slag to that in matte. ' Eng. Min. J., iSgt, Lll, 74. ' Tr. Inst. Min. Met., 1913, xxii, 50; Min. Eng. World, 1913, xxxvni, 9. ' Eustis, W. E. C, Private Communication, April, 1894. ' Heywood, Eng. Min. J., 1904, Lxxvii, 395. Wright, Tr. A. I. M. E., 1909, XL, 492; 1910, xli, 316. Channing, op. cit., 1910, xli, 885; Min. Sc. Press, 1909, xcrx, 668. Vallely, Eng. Min. J., 1905, lxxdc, 1223; Ann. chim. analytique, 1905, x, 193. Heberlein, Eng. Min. J., 1910, lxxxdc, 617. Gabrill, op. cit., 1910, ixxxrx, 776. Schertel, L., Thesis, Freiberg, igio. ' Metallurgie, 19x2, ix, 148. 2l8 METALLURGY OF COPPER with as high a ratio of matte to slag as 2: s, the Cu-content of the slag increases with that of the matte. Wright finds that the Cu-content of slag increases with that of the matte produced, as shown in Fig. 234. Heywood, Fig. 235, states that slags rich in iron and mangar nese carry more Cu than when rich in SiOa. A relation be- tween the percentage of SiOa and the Cu-content has been noted in some cases; thus acid slags are to contain 0.5 per cent, and basic slags i per cent, of the Cu-content of the matte. The present knowledge of the dififerent factors is still too in- complete to permit application of laws of physical chemistry for drawing general conclusions which might assist in explaining satisfactorily individual cases. The Ag- content of slags appears to run parallel with that of Cu; that of Au shows no recognized regularity. Waste blast-furnace slag with 40- to 4S-per cent, copper matte contains from 0.2-0.5 per cent. Cu. 0.4U 1 = 0.36 m a g0.30 ^ 0.25 1 "-^ J _^ / / .-' y 10 15 20 25 30 35 40 45 Feri3&ntage of Copj)er lu Matte Fig. 234. — Relation of copper-content in slag and matte. 30 S5 40 45 50 Per cent Copper in Matte 65 Fig. 235.— Relation of copper-content in slag to SiOj- and Fe(Mn)0-content. 122. Gases and Flue-dust.— The average temperature of the waste gases at the open throat of a copper blast-furnace is low, in a reducing fusion (150° C.),in true pyritic smelting(25o°C.). It is high in partial pyritic smelting (over 300° C). SMELTING OF COPPER 219 Their velocity is also high on account of the small difference between tuyere- and throat-areas, an average figure being about 750 ft. per minute. * Though the temperature of the gases may drop to 100° C. when a new charge has been fed, it will rise in partial pyritic smelting to 600° C. by the time the next one is intro- duced, and at this temperature the velocity may rise to iioo ft. per minute. In the downcomer, which carries away also the air that enters by the feed-doors, the velocity varies from 1000 to 1500 ft. per minute. The composition of the gases has been given on pp. 184, 194, 201, 225. The temperature will have to be reduced to 300° C. and the velocity to 200 ft. per minute in order that the fine dust may fall out.^ ' Studies carried on within the last six years upon the deposition of dust and fume in flues and chambers have been the cause of many advances made in the recovery of values otherwise lost. The installations of Cananea, Copper Queen, Anaconda, Great Falls, Mammoth, and Balaklala may serve as examples for the discussion of the subject. At Cananea' the blast-furnaces treat much fine material, 50 per cent, being smaller than |-in. The gas from a blast-furnace passes through a goose-neck into a balloon-flue common to eight blast-furnaces; from this the combined gases travel through two cross-over flues into a large dust cham- ber. Of the total dust collected, 55 per cent, was recovered in the balloon- flue, 2 per cent, in the cross-over flue, 28 per cent, in the first half of the chamber (the major portion lying beneath the cross-over flue), and the remain- ing 15 per cent, in the second half. A screen-analysis of the dust collected in the chamber showed that the bulk of the part which settled in the first half was of 60- and 80-mesh, and that the rest was 150- and 120-mesh and finer. At the Copper Queen works the experiments carried on by G. B. Lee* led to the conclusion that for that plant the flues and chambers need not exceed a length of 125 ft. as long as the velocity of the gases is not greater than 150 ft. per minute, even with dust of which 90 per cent, will pass a 200-mesh screen. 5 ' Haas, Eng. Min. J., 1910, xc, 814. ' Kiddie, Tr. A. I. M. E., 1909, xl, 900. ' Shelby, Eng. Min. J., 1908, lxxxv, 204. * Eng. Min. J., 1910, xc, 504. tMin. Sc. Press, 1913, cvii, 929. Blast Furnace Plant Fig. 236. — Plan of flues at Washoe smeltery, Anaconda. 220 METALLURGY OF COPPER At Anaconda' there is in operation for the collection of dust and fume a system of long flues through which the gases travel with diminishing tempera- tures and velocities. The plant has three blast-furnaces 56 in. by 51 ft. and one 56 in. by 15 ft., 56 Evans-Klepetko MacDougall furnaces 16 ft. inner diameter, seven re- verberatory furnaces 19 by 102-115 ft., and ten horizontal basic converters 8 ft. in diameter and 12 ft. 6 in. long. The general plan of the condensation is shown in Fig. 236. The blast- furnace-, roaster- and converter-departments each have a dust chamber for collecting the coarse dust; in the reverberatory department the two Sterling boilers back of a furnace serve as catchers of coarse dust. Having been freed from coarse material, the dust and fume of the blast-furnaces travel through a flue of brick and steel construction (653 ft. long, 20 ft. wide, and 15 ft. high) to the main flue (1234 ft. long, 60 ft. wide, and 20 ft. high) of I-beam and brick-arch construction, and enter the twin flue (995 ft. long, 120 ft. wide, and 20 ft. high). Fig. 237. — Twin condensation-flue at Washoe smeltery, Anaconda. shown in cross-section in Fig. 237, leading to the stack (300 ft. high and 30 ft. inner diameter). The gases thus travel 4182 ft. The roaster-gases travel 3017 ft.; the converter-flue is 7 by 7 ft., and the gases travel 3662 ft.; the reverbera- tory gases travel 3782 ft. Near the foot of the 60-ft. main flue is the arsenic plant,^ which treats the dust recovered in the 120-ft. flue. The velocity of the gases in one of the double flues is 12 ft., and in the stack 16-17 ft- per second. There are recovered in the flue of the double main 0.114 lb. dust per 1000 cu. ft. gas, and 1,450,000,000 cu. ft. gas pass through it per day. The dust that passes off into the open is finer than 200-mesh; an analysis gave: SiOa 4.19, AI2O3 3.14, FezOs 2.78, CaO 0.50, MgO 0.69, total S 10.05, total SO3 23.25, free SO3 3.60, Cu 0.84, Zn 9.50, AS2O3 34.34, Pb 10.80, SbjOs 1.17, BijOs I. IS, NazO 0.49, K2O 0.42, C 0.8, Se 0.28, Te o per cent., and Au 0.005 and 1 McDougal, Canad. Min. Rev., 1905, xxiv, 26. Austin, Tr. A.I. M. E., 1906, xxxvii, 478. Dunn, op. cii., 1913, xlvi. 2 Elton, Tr. A. I. M. £., 1913, xivi. SMELTING OF COPPER 221 Ag 4.9 oz. per ton. An average sample of the dust collected in the 120-ft. main flue gave Cu 0.986 and AS2O3 34.2 per cent., and Au 0.0058 and Ag 4.44 oz. per ton. At Great Falls' the Roesing wire system^ for retarding the gas current by increasing the resistance through suspended wires is in successful operation. The general arrangement of plant and new system of condensation is shown in Fig. 238, a cross-section of the dust chamber in Fig. 239, and partial longitudinal section in Fig. 240. The plant has 24 Evans-Klepetko-MacDougall furnaces (16 ft. inner diameter), three gas-fired reverberatory smelting furnaces (42 ft. 6 in. by 15 ft. 9 in.) with regenerative chambers, five blast-furnaces (54 by 160 in.), seven upright converters' (7 ft. diameter and 14 ft. 7 in. high). Elaborate working experiments were carried on with various dust-arresting devices before the present system of condensation was introduced. The results of these tests are represented graphically in Fig. 241. Curves 32 and 34, representing gases passing through an open or ordinary flue, show that under the conditions of the tests only 30-40 per cent, of the dust was collected. With narrow plates sus- pended in such a way that the gas-current strikes the edge (Freudenberg plates)* matters are improved, curve 33, as the collection of dust is over 40 per cent. By suspending in the flue long narrow plates in such a way that the gas-current strikes the faces, a great deal more dust is precipitated as seen in curve 36 (35-in. bafile-plates),. and in the entrance- and exit-parts of curve 35 (6j-in. baffle-plates). The wider plates reducing the cross-sectional area 50 per cent, were more effective as dust catchers than the narrower which reduce it 25 per cent., but both strongly interfere with the draft. The difference between open- and open- and baffled-flue is shown strikingly in curve 35. The action of suspended wires is seen in curve 39 ("wire baffles"). The wire-baffles do not obstruct the draft as do plate-baffles and at first do not collect as much dust, but make up for this later on, causing 84 per cent, of the dust to separate, which is more than the other arresting devices. Curve 38 brings out the difference in settling power of an open flue and one provided with wires. In curve 37 is represented the effect sudden reductions and enlargements of area of flue at certain distances have upon the settling of dust. In the tests, the flue, 304 ft. long and 4 ft. by 4 ft. 6 in. = 16 sq. ft. area, was partly closed, 100 and 104 ft. from the ends, by two partitions each having in the center an opening i ft. 6j in. in diameter= 1.8 sq. ft. area. This arresting device is effective as a collector of dust and is cheap to build. With a reduction of area of from 18 to 1.8 sq. ft. the interference with the draft was too great to work satisfactorily, but the data show that the method is promising. In the new flue-system. Fig. 238, there have to be considered the dust- ' Herrick, Mines and Minerals, 1909, xxx, 257. Goodale, Tr. A. I. M. E., 1909, xl, 891. Goodale-Klepinger, op. cit., 1913, XLVI. 'Hofman, "General Metallurgy," 1913, p. 846. ' Since the basic converter has replaced the acid, the sizes of vessels have been increased and the number decreased; see § 176. *Ho£man, "General Metallurgy," 1913, p. 845. 222 METALLURGY OF COFFER — rfi SMELTING OF COPPER 223 s o a T3 d 224 METALLURGY OF COPPER chamber, the connecting flue leading to the stack, the flues leading to the dust- chamber, and the results obtained.^ The brick dust-chamber, 357 ft. iiyV in. long by 176 ft. wide by 27 ft. high, is supported. Figs. 239 and 240, by columns, 11 ft. 9 in. high, so as to leave room for the bottom-discharge of dust through sheet-steel hoppers (over 1000 in num- ber, arranged in 22 lines) into cars running on tracks of 3 -ft. gauge. The cham- ber is divided longitudinally by a partition wall so that by means of dampers the gas-current can be made to travel through either half. At the inlet of the cham- Total Dust is Gases 60 100 150 200 250 SCO Length of Experimental Flue in Feet Fig. 241. — Relative efficiency of dust-arresting devices, Great Falls. ber a space is left clear to afiord an unobstructed passage for the gases that they may distribute themselves over the fuH width of the chamber. In order to cool the gases, 22 air-admission pipes, Fig. 240, enter from top and bottom. The wires are suspended in two divisions. From the entrance of the chamber to a distance of 150 ft., and back from the exit also for 150 ft., the space is fully occupied by wires; the intervening space of 47 ft. is left free from wires. The purpose of this arrangement was to collect dust in the entrance- and fume in the exit-division. The wired part of the chamber holds about 1,200,000 steel wires spaced 2.3 in. center to center; for a distance of 51 ft. from the inlet the whes are No. 8 W. & M. gauge and 16 ft. long ; the rest of the wired chamber has No. 10 wires 20 ft. long. For the support of the wires, steel-wu-e netting, Figs. ' Some minor changes not shown in the drawings have been made since the new flue-system went into operation June 12, igog, but these are passed over. SMELTING OF COPPER 225 242 and 243, of i|-in. mesh is bolted to the I-beams of the roof; the baffle-wires are bent at one end to the form of a shepherd's hook and suspended from alternate intersecting points; they are thus staggered, which aids in arresting dust. For the shaking of the wires to dislodge adhering dust, angle-iron frames, Fig. 240, 10 ft. wide extending from the side-^walls to near the partition wall, are suspended by hangers about 10 ft. long. A frame has a wire netting with open- ings 4 by 7 in., is shaken for 30 min. at intervals 60-90 days by a connecting- rod extending through the flue-wall and attached to a bell-crank lever actuated by an eccentric with a stroke of 9.5 in. and 60 strokes per minute; the eccentrics on each side of the chamber are carried by a line-shaft operated by an electric motor. The flue connecting the dust-chamber with the stack, Fig. 238, is (section EE) 48 ft. wide by 21 ft. high. At its term- inus it is divided into two branches (sec- tion FF). The stack is 506 ft. high and 50 ft. in diameter at the top; it is circular inside, octagonal outside for 46 ft. from base, and circular for the remainder of the height (drawings in reference). The gases from the converters. Fig. 238, pass through the converter-flue which joins the flue from the MacDougall roasters (section CC) as this enters the cross-take flue (section DD). The gases from the reverberatory furnaces pass into a collecting flue before they either enter the dUst-chamber, or are by-passed around the dust-chamber to the flue leading to the stack. The gases from the blast-furnaces (section BB) travel through their main to the cross-take flue. The temperature of the gases (excluding reverberatory-furnace gases) at the entrance of the dust-chamber is 163° C. and at the exit 135° C; near the chimney at the branching of the connecting flue (including the reverberatory-furnace gases), the temperature is 163" C. Table 45 gives temperature- and draft-readings in the leading divisions of the flue-system; Table 46 the velocities, volumes, average temperatures, and weights of gases; and Table 47 the amounts of gas produced per furnace and per ton of charge. Practically all the dust and most of the metallic fume are recovered. The escaping gases contain free H2SO4 22.23, Si02 2.30, Cu 0.70, (FeAOaOs 7-o8, S 6.67, SbaOs 1.47, BizOs 0.81, PbO 0.49, CaP 0.18, ZnO 3.31, O (calculated for S) 10.23 PP^ cent. In Fig. 244 is shown the relative deposition of dust through the chamber; the superimposed fine full-drawn lines represent the outline of the dust-chamber, IS Figs. 242-243. — Method of hanging wires in dust-chamber at Great Falls.- 226 METALLURGY OF COPPER Table 45. — Temperature- and Draft-readings Locations of readings in Fig. 238 Elevation Temperature in flue, F.° Temperature of atmos- phere, F.° Draft-readings, in. water Impact tube Static tube (A) Blast-furnace (B) MacDougall (D) Cross-take 3,338 3,361 3,413 3,413 3,413 3,57° 391 419 358 346 312 310 80 80 80 70 70 70 1 . 12 0.94 0.88 0.94 2. 20 1.80 1.16 0.98 0.97 (D) Cross-take 1. 01 (L) Connecting flue (U) Near chimney 2.26 1.84 Table 46. — Velocities, Volumes, Average Temperatures, and Weights of Gases 1 Loca- tions in Fig. 338 Date, 191 1 No. and kind of furnaces Average temper- ature, po Clear area of fiue, sq. ft. Vel. in ft. per sec. Vol. at observed temperature, cu. ft. per min. Pounds gas per min. Pounds gas per furnace per min. A B C G Di.... Apr. 6-8 Apr. 6-8 Apr. 6-8 Mch. 21-24 •• Apr. 6-8 1 Mch. 21-24.. Mch. 21-24.. Apr. 10 Apr. 10 4B. F.... 10 MacD... 5 Conv .. . 2 Rev 4B. F.... 10 MacD... 5 Conv... . 4B. F.... 6 MacD . . 5 Conv. . . 4 B. F.... 6 MacD... 5 Conv. . . 2 Rev . . . . 4B. F.... 10 MacD . . 5 Conv. . . do 345 352 303 496 331 286 311 290 322 401.6 169.0 78.5 152. S 636 977 977 977 636 17.26 17. SO 50.37 43.34 21.80 15.51 21.03 15.54 22.74 415,960 253.500 213.630 396,700 859,300 901,400 1,234,900 910,950 867,760 18, 510 10,980 9,840 14.530 38,190 42,280 s6,iio 42.270 38,700 4.063 1,100 1.970 7,270 E, F Ej Di. . Table 47. — Amc UNT OF Gas pee Furnace and per Ton of Charge Kind of Observed temperature correspond- ing to given volume, F.° Rate per min. Rate per 24 hr. Aver- Per ton of charge furnace Cu. ft. Lb. Cuft. Lb. charged Cu. ft. Lb. 345 352 303 496 103,990 25,350 42,730 198,350 4.630 1,100 1,970 7,270 149,745,600 36,504,000 61,531,200 285,624,000 6,667,200 1,584,000 2,836,800 10,468,800 391.61 70.7 32.02 188.5 384,900 516,300 1,922,900 1,515,200 MacDougall .... Converter. ..... Reverberatory . . 22,400 88,600 55,500 1 Includes flux but does not include fuel, 8 Tons copper produced per converter day. SMELTING OF COPPER 227 20 26 Hopper Number Fig. 241. — Relative deposition of dust through dust-chamber, Great Falls. Table 48.— Quantity and Analyses op Flue-dust Tons dust Cu per cent. Ag oz. per ton Au oz. per ton Insol. SiOj FeO AhO, CaO Name of flue Total for 41 months Aver- age per month S Blast-furnace MacDougall furnace. Up- take and cross-take Main dust-chamber. . Connecting fluei 87,020 18,741 17,360 64,048 4,000 2,122 457 423 1,562 98 8.08 10.22 12.59 8,61 3.09 2.7 3.6 4-1 3.3 3.1 0.019 0.023 0.026 0.020 0.012 34-7 37.6 34-8 33.3 12.6 26.8 29.0 26.0 23.6 8.5 33.0 27.1 27.2 14.5 5-4 7.1 7.5 7.2 8.0 4.0 1.7 0.3 0.7 0.7 0.1 16. 1 21.3 19. 5 II. 8 10.6 Total igi,i6g 4,662 » Weight estimated. Average analysis is from sample taken in June, 1912, at different points from dust-chamber to chimey. the dotted lines the areas filled with wires. The quantities and analyses of dust collected in a period of 41 months are shown in Table 48. The distribution of the material is shown in Table 49. Table 4g. — Percentage Distribution of Material in Flue System Weight Copper SiOj Blast-furnace flue MacDougall-furnace flue Cross- take flue Main dust-chamber Connecting flue Stack discharge Totals 39 8 7 28 I 13 41.4 II-3 12.9 32-5 0.7 1.2 47- II. 8, 30. o, I 228 METALLURGY OF COPPER The Mammoth Smeltery, Kennett, Cal.' — The method for the recovery of flue-dust is by filtration. The plant has five blast-furnaces, 50 by 180 in., in which partial pyritic smelting is practised. During the summer and autumn two furnaces are in blast, during the winter four. The cooling system is not large enough to handle in the warm season more than the 250,000 cu. ft. gas per minute at 0° C. that come from two blast-furnaces and the converters at a temperature of 275° C. The matte with Cu 20-30 per cent, is converted in two stands with acid-lined vessels 96X150 in.^ The general arrangement of the condensation plant is given in Fig. 245. The gases from the blast-furnaces and converters travel through two long brick flues, in which most of the dust is settled, to the base of an old stack (marked 18X200 Fig. 245.— Condensation plant of Mammoth smeltery, Kenneth, Cal. ft. in Fig. 245) and thence through four steel pipes, 8 ft. in diameter, to a steel collecting chamber, 42 ft. wide by 15 ft. high, which contracts toward the exit and ends in a steel flue, 1 5 by 1 5 ft., terminating in the fan house. Small amounts of ZnO and ZnS04 are collected in the flues and chamber. In order to cool the gas to 100° C. and less, air is admitted to the fan discharge chamber (later it will be blown into the distributing chamber in front of the baghouse). In the fan- ' Campbell, Min. Sc. Press, 1908, xcvi, 30. Martin, Min. Eng. World, 1908, xxix, 309. Rice, Eng. Min. J., 1911, xci, 614. ■Nevins, Min. Sc. Press, 1913, cvi, 374. Martin, Mines and Minerals, 1913, xxxin 323. ''Kervin, Eng. Min. J., 1914, xcvii, 713. SMELTING OF COPPER 229 house are two Sirocco fans 11 ft. in diameter making 219 r.p.m., each driven by a 400-h.p. General Electric induction motor. Below the fans there is a vacuum of 1 .7 ; above, a pressure of 3 in. water. At the entrance of the two 8-ft. delivery- pipes into the fan-discharge chamber, 20 by 15 ft. and 160 ft. long, is a machine for feeding zinc oxide^ to neutralize SO3 that has remained uncombined. Most of the SO3 in the blast-furnace gas has formed ZxixSOi with the ZnO formed in smelting the charge containing about 4 per cent. Zn. The gases from the converters are similarly treated with ZnO and powdery Ca(0H)2 before they join the blast- furnace gases. From the chamber the gases travel through 39 steel cooling-pipes, 4 ft. in diameter and 200 ft. long, supported by iron rings held in wooden stands. The pipes discharge into the baghouse distributing chamber; each is provided with a butterfly valve so that when several are closed the small amount of dust collected in the others may by swept out by the stronger gas-current passing through them. From the chamber, 13 ft. 6 in. by 14 ft. and 210 ft. long, 20 short pipes, 3 ft. 6 in. in diameter, provided with flap-valves, deliver the gas into the baghouse which is 210 by 63 ft., has 3000 bags, equivalent to 25,000 sq. ft. filtering surface per ton of fume per day.^ The baghouse has the usual thimble- floor; beneath it are 20 transverse hoppers for collecting the fume to be dis- charged through circular gates into cars (a continuous discharge is to be installed) , dampened, and briquetted. The filter chamber, 40 ft. high, is divided into five bays, each with 600 bags 34 ft. long and 18 in. in diameter, some of wool, others of cotton. The bags, 3-10 in. apart, are placed in rows 21-28 in. center to cen- ter. Each bay has a monitor tower, 20 ft. square and 1 2 5 f t. high, which draws in enough air to dilute the SOa-content of the filtered gas to 0.75 per cent. vol. The dust adhering to the bags is loosened by means of Benedickt shakers.' The gas-pressure in the baghouse is 1.7 in. water. Men entering the filter chamber carry Draeger fire-fighting helmets. The crew of the plant for 24 hr. is made up of i foreman and 12 men, 4 on an 8-hr. shift. The cost of installation is $800 for 1000 cu. ft. of gas per minute. With two blast-furnaces there are filtered per minute 250,000 cu. ft. gas; these furnish in 24 hr. 10 tons of fume, which consists largely of ZnO and ZnS04, runs high in Ag, carries some Au and Pb, and Cu 0.5 per cent. An anal- ysis quoted by Nevius^ shows: Cu 1.04, Insol. 7.8, Fe 6.2, CaO 1.8, Pb 7.0, As 4.3, ZnO 4.8, ZnS04 47.2 per cent., Au 0.03 and Ag 4.08 oz. per ton. Balaklala Smeltery, Coram, Cal. — Here the Cottrell system of electric condensation^ was in operation before the plant was shut down. In this process a high-potential direct current jiunping through an air-space from needle-points of one pole to the plate of the other causes particles of suspended dust and vapor to travel toward the plate at a speed which is proportional to the charges and to the potential gradient between point and plate. ' Sprague process: Hofman, " General Metallurgy," 1913, p. 849. Sprague, Eng. Min. J., igio, Lxxxix, 520. ' Sprague, he. oil. 'Hofman, "General Metallurgy," 1913, p. 855. *Loc. cU. 'Hofman, 0/1. ci/., 1913, p. 859; Cottrell, J. Ind. Eng. Chem., 1911, ni, 542. 23° METALLURGY OF COPPER The smeltery treated 700-1000 tons sulphide ore with Cu 2.5-3.0 per cent., S 30 per cent., and more or less Zn; 90 per cent, was smelted in blast-furnaces; 10 per cent, was roasted in MacDougall furnaces and smelted in oil-fired rever- beratory furnaces; the matte was converted in two stands. The gas from these departments entered a common flue, 18 by 20 ft.; it varied with operating con- ditions from 250,000 to 500,000 cu. ft. per minute.; the linear velocity in the flue varied correspondingly from 10 to 20 ft. per second; the temperature at the inlet-flue ranged from 100-150° C. A plan and section of the condensation plant, costing $100,000, are shown in Figs. 246 and 247. The gases from the smelting and roasting furnaces are col- lected in the inlet-flue. Fig. 246, pass through nine precipitation chambers con- nected in parallel, are assembled in the outlet-flue, and delivered by two fans into ggggggg Outlet Flue ■ FtowJ 1" f ,1 s a \nrw dj s Rectifier Bldg 36'x 98 u^ ^ Rectifiers 1 h & a a □ a a| Fig. 246. — Cottrell electric condensation plant at Balaklala smeltery, Coram, Cal. the chimney; the fans suck in enough air to reduce the S02-content of the cleaned gas to 0.75 per cent, or less.i In the rectifier an alternating current of 2300 volts is transformed up to 25,000-30,000 volts, rectified into an intermit- tent direct current, and distributed to the precipitating chambers. In the cross- section of a precipitating chamber, Fig. 247, the twenty-four vertical double lines represent the collecting electrodes (6 in. wide, 10 ft. high, of No. 10 sheet iron) carried by bars direct connected with the frame of the chamber and thus grounded. The vertical dotted lines represent the discharge electrodes con- sisting of two iron-wire strands between which is twisted the discharge material— asbestos and mixed preparations; they are spanned by springs between busbars carried by insulators placed outside and enclosed in auxiliary chambers through which passes an air-current to prevent con- ductive dust or fume from settling. ^ The cam, shown on the left side, and ' Later, chambers i to 9 were connected up in three series receiving current with increasing potential. 2 The wires were replaced by split iron rods holding "micanite" formed to have saw-teeth, as the wires became clogged with zinc oxide. SMELTING OF COPPER 231 the shaker-rod, extending across the chamber, intended to remove precipitate, were found to be unnecessary, as the electrodes could be easily shaken by hand from the top; this had to be done every 6 or 8 hr. The dust collected in the fume-hopper and was removed by a conveyor. The average power-consumption was 120 kw.; the plant required one fore- man, one man in the rectifier house, and two in the precipitation house. With all parte under good control the saving of suspended matter was over 90 per cent., the average was nearer 75 per cent. The plant was closed before all the details had been perfected. Fig. 247. — Cottrell electric condensation plant at Balaklala Smeltery, Coram, Cal. Garfield Smeltery, Garfield, Utah.' — At these works a plant of one hundred and fifty s-in. pipe-electrodes, similar to those discussed in §173, each 10 ft. long, operated under 20,000 volts, was installed for experimental purposes near the chimney, which took gases from blast-furnaces and Mac- Dougall calciners. The apparatus treated from 4000 to 9000 cu. ft., blast-furnace gas per min.; 8 hours were given to a test; the treated gas was sampled con- tinuously in order to determine the amount of clearance. Tests lasting several weeks gave the following average results: With velocities of from 5 'Private communication, W. H. Howard, December, 1913. 232 METALLURGY OF COPPER to 7.5 ft. per sec. in the electrodes and a temperature of 90° C, 95 per cent, clearance was obtained, all the dust being collected; the remaining 5 per cent, consisted mainly of moisture and a small amount of acid. With velocities of from 3 to 3.5 ft. per sec. and a temperature of 85" C, perfect clearance was obtained. With roaster gas having a velocity in the electrodes reaching 7.5 ft. per sec. and a temperature of 85° C, practically complete clearance was obtained. Observations on the poisonous eSects of copper when melting in an electric furnace have been published by Hansen.^ _ BO DLa. Dust Bio ^ Figs. 248-250.— Mixer for flue-dust and converter slag, Copper Queen smeltery, Douglas, Ariz. 123. Treatment of Flue-dust.— The treatment of flue-dust varies with the character of the plant. > If reverberatory-smelting is carried on, the flue-dust is usually added to the reverberatory-charge. At Morenci, Ariz.,^ it is charged into an oil-fired so-ft. reverberatory furnace which serves as a settler for blast-furnace slag. Incorporating the dust in converter-slag was in successful operation at the Copper Queen smeltery, Douglas, Ariz., up to the time of the erection of the ' Met. Chem. Eng., 1911, ix, 67. ' Douglas, Eng. Min. J., 1907, Lxxxni, 198. SMELTING OF COPPER 233 reverberatory division for smelting roasted concentrate.* Figs. 248-250 show Eide- and end-elevations and plan of the apparatus. The leading parts are a du£t-bin with bottom-discharge, a tilting converter-ladle emptying into an in- clined trough, and an inclined conical drum built of ribbed cast-iron sections, i.S in. thick, running on friction rollers. The discharges of dust-bin and con- verter-ladle meet at the upper end of the drum; the slag readily takes up the dust in its passage through the drum forming balls varying in size from 0.5 to 6 in. in diameter. At Cananea^ the dust is incorporated in converter slag by feeding the dust from a hopper with spout into the bowl of a slag-car while converter slag is being poured from the same level through a trough. The dust is drawn down into the slag and becomes well mixed. When the bowl is filled, the slag is poured down an incline over which it rolls as a viscous mass and balls up, form- ing balls 0.5-12 in. in diameter which are sprayed with water to be cooled and rendered porous. A third method of compacting is that of agglomeration. Heberlein' patented a process for agglomerating iron or manganese ore, blue billy, flue-dust, etc., which consists in adding carbonaceous fuel to the charge and blowing it in an Huntington-Heberlein pot as in the ordinary operation of blast-roasting. This process is in operation* at the works of the Zenith Furnace Co., Duluth, Minn. Two 100-ton D wight-Lloyd straight-line sintering machines treat the flue-dust of the smeltery of the Mason Valley Mines Co., Thompson, Nev.^ A screen- analysis of an average sample of dust gave: On 20-mesh, i per cent.; 20- to 40- mesh, 12.3; 4o-to8o-mesh, 29.1; 80-to loo-mesh, 22.7; 100- to 150-mesh, 13.5; 150- to 20o-mesh, 10.8; through 200-mesh, 16 per cent. The finest dust gave upon sifting: On 20-mesh, nil; 20- to 40-mesh, nil; 40- to 8o-mesh, o.i; So- to loo-mesh, 2.9; 100- to 150-mesh, 0.5; 150- to 200-mesh, 3.5; through 200- mesh, 93.0 per cent. This is mixed with the coarser dust. A chemical analysis of average dust showed: Cu 3.5, Fe 18.5, SiOa 36.5, CaO 12.0, S 4 per cent. As the fuel value of the material is too low to permit treating the dust alone, there is mixed in 25 per cent, of fine sulphide material, with Cu 4.0, Fe 18, Si02 37, CaO 12, AI2O3 10, and S 12 per cent., and about 3 per cent. coke. The coke is omitted when the S-content of the mixture is 7-8 per cent. ; considerable attention has to be given to regulating the amount of water that is added to the charge. The grates of the machine are of malleable iron. The charge is made up as follows: The dust runs from a hopper-bottom flue into a drag-conveyor which delivers to the boot of an elevator emptying into a bin having a feed-belt beneath the discharge. Fine ore and coke are trammed to bins alongside which are similarly equipped. From the three bins forming a ' Editor, Eng. Min. J., 1913, xcvi, 627. ' De Kalb, Min. Sc. Press, 19 10, ci, 9. 'Oest. Zt. Berg. HUttenw., 1908, XLVi, S5S; l>Iin. Ind., 1908, xvii, 605. * Eng. Min. J., 1913, xcvi, 394. " Private Communication of J. Labarthe and A. J. McNab, August, 1913. 234 METALLURGY OF COPPER group any desired amount can be delivered to a common conveyor-belt by adjust- ing the speeds of the three belts, or the sizes of the openings, or both. The com- mon conveyor drops the mixture into' the boot of an elevator which discharges on to the conveyor-belt in the sintering plant which feeds the two machines These treat in 24 hr. 100 tons of mixture each and require 40 gal. of naphtha. The labor for each machine is one furnaceman, one helper, and one man moving cars. With two machines, one man with motor moves cars and takes away sinter. Total Dust in Gases -With Briquettes EO leo 160 250 250 Length of Experimental Flue in_ Feet 350 Fig. 251.— Cumulative dust-curves of blast-furnace-charges with and without briquettes, and of MacDougall calcimers. Briquetting plue-dust is a method of compacting fiue^dust which was in more general use a few years ago than at present. At Anaconda! the Chambers Bros. (Philadelphia, Pa.) No. 7 augur end-cut soft-mud brick machines^ are in operation. One machine makes in 24 hr. 840 1 Austin, Tr. A. I. M. E., 1906, xxxvii, 460. ^ Hofman, "General Metallurgy," 1913, p. 643. SMELTING OF COPPER , 235 tons of briquettes, weighing 5-10 lb., with 9 men on a shift. The mixture is made up of \ screenings from first-class ore, \ table concentrate, and \ concen- trator slime; to this is added 5 per cent, washed coke from the ash-pit droppings of the reverberatory smelting furnaces. The handling of material is mechanical throughout; the briquettes are stored in bins, from which they are drawn, im- perfectly dried, into the charging cars of the blast-furnaces. The investigations of Goodale and Klepinger^ into the condensation of flue- dust at Great Falls, Mont., have shown that blast-furnace charges containing briquettes make a great deal more flue-dust than those that do not; also that it is more difficult to collect the values in blast-furnace dust than in the dust from the MacDougall roasting furnaces. This is brought out clearly in Fig. 251. 124. Hearth Accretions (Sows), Etc. — Hearth accretions are of less common occurrence in treating sulphide ores, since the internal crucible has been aban- doned. An accretion consists of Fe-Cu alloy mixed with Fe304, matte, slag, perhaps some speise and other metallic compounds. The formation of a sow in treating roasted ore is probably caused by an excessive reducing effect upon the oxidized iron in the charge. In partial pyritic smelting, especially when a slag of high formation temperature is formed, causing FeS to split in part into Fe and S, the formation of a sow has probably to be attributed to the separation of Fe from matte; Fulton^ believes the separation to be due in part to the reac- tions 2FeS-|-Fe304=sFe-|-2S02 and FeS-f-2FeO = 3Fe-|-S02, but this has still to be proved. The other irregular products, such as wall accretions, furnace drawings, and refuse, need no further discussion. 125. Results. — The yield of metal in smelting sulphide copper ore carrying precious metal, but not contaminated with Pb and Zn, is high, as the only losses involved are those by dusting and slagging. The recovery of Cu is therefore well up in the nineties, say gy-f- per cent., that of Ag 98-^ per cent., and of Au loo-f per cent. The cost of smelting in the U. S. varies within wide ranges; the lowest is probably $0.50 per ton of charge, the highest $3.00. Beardsley^ estimates the cost at Mount Lyell to be $2.36; at Copper Hill (Tennessee Copper Co.), $1.24; in Mexico, $2.03; at Granby, B. C, the same as at Copper Hill. Austin^ gives for the Tennessee Copper Co., $0.96, for the Granby Cons. M. S. & P. Co., $1.20. The report of the Tennessee Copper Co. for 1911^ gives the cost as $0.89 per ton charge; that of Balaklala, $3.00 per ton;^ that of Cananea, $2,571 per ton of copper-bearing material.'^ These figures are low when com- pared with the sworn data for Butte in 1901-02* which were with the Montana 1 Tr. A. I. M. E., 1913, xLVi. ^ Eng. Min. J., 1904, ucxviii, 333. ' Eng. Min. J., 1906, txxxii, 3971. * Min. Sc. Press, 1911, cii, 178. ''Eng. Min. J., 1912, xcra, 1035. 'Eng. Min. J., 1912, xcm, 937. ' Eng. Min. J., 1912, xcrv, 114. 'Eng. Min. J.,igos, lxxv, 708. 236 METALLURGY OF COPPER Ore-Purchasing Co., $5.96 and with the Butte and Boston Cons. M. Co., $4.84 per ton of ore. The cost in Montana today is about $2.00 per ton charge. As regards the smelting power of a furnace, that obtained with a reducing fusion is by far greater than with pyritic treatment.^ The rating of a furnace,^ i.e., the amount of charge smelted in 24 hr. per square foot of hearth area, ranges from 4 to 9 tons, the larger figure referring to roasted ore. However, a pyritic charge ought to be coarse; many plants have to deal with ore finer than desirable, and fine ore reduces tonnage. Further, the character of the slag that is being made has an important influence upon the amount of charge that can be put through. Thus it is not possible to say oS- hand what ought to be the amount to be treated in a given time per square foot of hearth area, either in a reducing or a pyritic fusion. 126. Production in the Blast-furnace of Metallic Copper from Matte. — Formerly matte was brought forward to metallic copper by roasting and then smelting in the blast-furnace. This mode of procedure has become practically obsolete in the U. S., while it is still practised in other countries. Low-grade matte is enriched to converting-grade in the blast-furnace by pyritic smelting. The practice of the Tennessee Copper Co. is given in Table 28. It was found there that a 44X i8o-in. furnace put through more matte than ore, and that the reverse was the case with a s6Xi8o-in. furnace; also that the 44Xi8o-in. furnace gave a gas richer in SO2 and SO3 than the 56Xi8o-in. furnace. With the matte' is mixed flue-dust. The matte, held in a car of 105 cu. ft. capacity is poured on to a sloping yard 80 ft. long which is divided into beds 18 ft. wide. III. Smelting in the Reveeberatory Furnace 127. Smelting in the Reverberatory Furnace in General (Welsh Process).*— The characteristics of matting sulphide copper in the reverberatory furnace are, that fine ore, usually rough-roasted by a separate operation, is smelted on a silica-hearth for Cu-Fe matte, with from 33 to 45 per cent. Cu, and an acid slag, with 36-1- per cent. Si02. The matte is brought forward to metallic Cu either by several steps in reverberatory furnaces or by a single operation in a converter. 1 Douglas, "Power-plant of Copper Queen Smeltery," Tr. Inst. Min. Met., 1913 ; Eng. Min. J., 1913, xcv, 757; Min. Eng. World, 1913, xxxvm, 669. 2£»g. Min. J., 1903, Lxxv, 442 (Van Liew), 472 (Channing), 513 (NeiU, Heywood) 624, 661 (Metcalfe); Table 28. ' Guess, Eng. Min. J., igio, xc, 866. ^Le Play, "Description des Precedes MfitaUurgiques employes dans le pays des Galles pour la Fabrication du Cuivre," etc., Ann. Min., 1848, xm, 3, 389, 557; transl. into German by C. Hartmann, 1851, sold by Craz and Gerlach, Freiberg, Saxony. Levy, "Note sur la M6taUurgie du Cuivre par la M6thode Galloise," Rev. Un. Min., 1884, XVI, 286-339; Berg. Huttenm. Z., 1885, XLiv, 396, 493, 469, 485, 497, 507. Moore, Eng. Min. J., 1910, Lxxxix, 1021, 1063. Mathewson, Eighth Internal. Congr. Appl. Chem., 1912, m, p. 113; Tr. A. I M E IQ12 XLIV, 781. ' ■ ■ • • '^., •i , Laist, Eighth Internal. Congr. Appl. Chem., 1912, in, 97; Tr. A.I.M. E., 1913, xLiv, 806. SMELTING OF COPPER 237 The slag goes to waste. In the reverberatory furnace S is the leading reducing agent, the carbonaceous fuel burnt serves only to furnish the heat necessary for the chemical reactions to take place between ores and fluxes. The reverberatory-furnace charge is best made up of fine ores, hence the method of smelting is used mostly for concentrates; coarse ores, rich enough to pay for direct-smelting, usually go to the blast-furnace (comparison, see § 152) The ores are rough-roasted in fine-ore furnaces and are charged, if possible hot (400-500° C), into the smelting furnace. Reverberatory smelting of ore for matte has been developed in the United States first by Pearce in Colorado, and later by Allen, Keller, Klepetko, Mathewson, and others in Montana, so that it occupies to-day a position quite different from the early Welsh or European Continental practice. It will be therefore discussed as an independent process. 1800 1848 1878 1882 1887 ISbT 1893 1894 1900 «^^^, 7-^^! 1904 190B 1906 19U Figs. 252-265. — Evolution of reverberatory matting furnaces. 128. The Reverberatory Matting Furnace in General. — The sketches given in Figs. 252-265 represent the leading stages in the development' of the non-regenerative matting furnace. An example of the regenerative type is given in detail in § 134. The figure for the year 1848 resembles the early form described by Le Play. This has a large deep fire-place, an oval concave hearth contracted slightly near the fire-bridge, very much so near the flue; there are a working-door on one side, a matte-tap on the other, and a skimming door at the ' Editor, Min. Sc. Press, 1910, ci, 69. Mathewson, loc, cU. 238 METALLURGY OF COPPER end, above which is an inclined flue leading the gases into a well-drawing stack. The hearth slopes from fire- and flue-bridges toward the center and from the back toward the front, the deepest point, at which is situated the tap-hole. As this furnace treated in 24 hr. only 8.6 tons of charge (see Table 50), it was essential that the capacity be increased, especially in the United States, if it was to compete with the blast-furnace. This was done by R. Pearce, first at Black Hawk,i and later at Argo,^ Colo. The original oval form was retained from 1878 until about 1891, when one side was slightly straightened in order to furnish room for two working doors. The cast-iron rule of requiring an oval plan having been broken, the oval sides were straightened more and more and thus a gain in hearth area secured until the standard of 1900, with a hearth 20 by 50 ft., was reached. Then E. P. Mathewson increased the -length of the coal-fired furnace to 102 and even 115 ft. 10 in. and witn it the mode of operating. Oil-fired furnaces reached in 1911 a length of 120 ft. 10 in.; in 1913 one of 130 ft., which is probably the greatest permissible length. The lead- ing facts of the development of the coal-fired reverberatory furnace have been brought together in Table 50. Table 50. — Types in Development of Non-regenerative Coal-fired Matting Furnaces Reverberatory Hearth E to .S g a": ^^ ^ Ratio Coal, tons "■■3 Locality, date Xii C 'CI Hearth: grate Grate: chimney Charee, lb. 24 hi- : I sq. ft. hearth H " s St s a Wales, 1848 1 14X101 120 20 4.66 8.6 6 :i 4.3 :i 144 II. 2 0.56 0.77 Colorado and Montana, 1891. 24X14 265 28. s 9 28 9.3:1 3.17 :i 211 10 0.36 ^,80 Anaconda, Montana, 50X20 1900. 886 53.3 30.68 105 16.6:1 i.73 :i 236 3S 0.66 3.03 Anaconda, Montana, 112X19 1906. 1 1,98s 112 ? 300 17.7:1 ? 302 62,5 0.55 4.8 129. The Reverberatory Furnace in Detail.— Modern reverberatory fur- naces, Figs. 252-265, show much similarity in their general form. A few typical forms have been selected for the discussion of the construction; the modes of firing with lump coal, fuel dust, oil, and producer gas; and the management; they are the furnaces of the Colorado Smelting Co., Butte, Mont., 1903; of the Anaconda Copper Mining Co., Anaconda, Mont., 1908; the Canadian Copper Co., Copper Cliff, Ont., 191 2; the Cananea Consolidated Copper Co., Cananea, Sonora, Mex., 1912; and the Anaconda Copper Mining Co., Great Falls, Mont., 1912. 130. Furnace of the Colorado Smelting Co., Butte, Mont, 1903.— This is shown in Figs. 266-271. It has a hearth 49 ft. 6 in. by 20 ft. 2 in. which has air- flues underneath to cool the bottom and to furnish preheated air to a box on the ' Egleston, Tr. A.I. M. E., 1875-76, iv, 276. ' Pearce, Tr. A. I. M. E., 1889-90, xviii, 55. SMELTING OF COPPER 239 roof above the fire-bridge, whence the air enters the furnace and assists the com- bustion of the fire-gases. The furnace, dismantled several years ago, served to produce high-grade matte which was tapped at intervals into a series of commu- nicating cast-iron molds and shipped to Argo, Colo., for further treatment. The ground-plan, Fig. 266, gives the rock and brick foundation-walls with the ash- pit floor of the fire-place at one end and the base of the stack at the other. Cold air is forced in at this level; it enters through flue a, passes around partition- wall h, returns by flue c, and travels along vault d. Following it in Figs. 269 and 267, it rises 2 5 ft. in two corners, e, of the stack (square on the outside, circular on the inside), to descend in the others, e', when it is divided by tongue/. Fig. 267, into branches g and h; branch g furnishes the air for flues gig2g3, while branch h supplies it to flues Ai A2 h^. Arriving at the fire-bridge end of the furnace, the Fig. 266. — Reverberatory matting furnace Colorado Smelting Co., ground plan. heated air rises in corresponding vertical flues, enters the air-box i, Fig. 269, and descends through ports, j, situated in the roof above the fire-bridge wall. In Fig. 267 there are indicated branch-inlets K and K' from flues a and c which communicate with longitudinal channels I and m; they serve to control the tem- perature of the air passing through gi_ 3 and Ai_3,and are regulated by sliding dampers V and m'. Turning to Figs. 267 and 268, the inner side-walls, n, of the hearth are of silica-brick backed by fire-brick, 0, which in their tturn are replaced by red brick, p, wherever possible. The roof, always made of silica- brick 15-18 in. long, has five charging-ports, the inner sides of the fire-place are also of silica-brick. The fire-bridge air-space is enclosed by two heavy cast- iron plates, the bridge-plate q and the E-shaped casting r; beneath the skimming- door is placed the front-plate t; both serve to distribute the loagitudinal thrust evenly upon the buckstays. In smaller furnaces the bridge-plate, extending beyond the fire-bridge to the 240 METALLURGY OF COPPER O U o U o M SMELTING OF COPPER 241 16 242 METALLURGY OF COPPER Figs. 269-271.— Reverberatory matting furnace, Colorado Smelting Co., vertical sections. SMELTING OF COPPER 243 points s and Ji (Fig. 267) and forming parts of the end-wall, is called conker- plate;* the buckstays are placed against conker- and front-plates. The putting-in of the working bottom is taken up in § 139. The tap-hole v is placed on the side near the fire-bridge between the first and second doors. With a grate 10 ft. long, it is necessary to have a fire-door on either side to per- mit successful grating. The fire-place receives its coal through an opening in the roof fed from a hopper; it is run with undergrate blast. The fire-gases leave the furnace through a short inclined flue, w, Fig. 269, ending in a small dust- chamber, X, and then pass into the stack. Flue w is covered with tiles clamped with heavy channels. The work of the furnace is given in Table 51. The furnace of the Anaconda Co. of 1903, at Anaconda,* had the same general form as that of the Colorado Smelting Co., only the waste heat was utilized' by passing the gases through a 300-h.p. Stirling boiler.* The draft shovdd not be less than 1.7 in. water. Clark* published recently a drawing of a modern small-size reverberatory matting furnace. Nearly all reverberatory smelting plants have followed the example set by Anaconda of using Stirling waste-heat boilers. Sorensen' compared the Stirling with the Babcock & Wilcox boiler at the works of the Nevada Consolidated smeltery, where they were in operation, and came to the conclusion that the cost of cleaning in the Stirling is much less than with the Babcock & Wilcox, and that the evaporative power of the Babcock & Wilcox greatly exceeds that of the Stirling; in fact, i lb. of oil burned in the reverberatory matting furnace evapo- rates only 3.315 lb. water in the Stirling as against 7.91 lb. in the Babcock & Wilcox, and this more than compensates for the extra cost and the time lost in cleaning and repairing. 131. Ftxmace of the Anaconda Copper Mining Co., Anaconda, Mont., 1908.' — ^This furnace designed by E. P. Mathewson is shown in Figs. 272-274. The hearth is iii ft. 8 in. long; its sides are parallel for a distance of 74 ft.; the enlargement from the fire-bridge (length 16 ft.) to normal width of 19 ft. is abrupt, the contraction from 19 ft. toward the flue-end to 7 ft. is gradual, taking up 32 ft. of the length of the furnace. A right degree of contraction is one of the essentials for obtaining a temperature sufficiently high to keep the slag fluid at the flue-end. The bottom is built up solid of slag, 24 in. deep, poured hot, and covered by 1 2 in. of silica brickbats. The earlier practice of having a concrete foundation has been abandoned, as ' Barbour, Conker-plate details, Eng. Min. J., 1912, xciv, S4i- ' Drawing, Min. Ind., 1902, xi, 202; Work, Table 41. ' Hofman, Tr. A. I. M. E., 1904, xxxrv, 295. * Waste-heat boiler of the Colusa-Parrot M. & S. Co., Butte, Mont., Eng. Record, 1907, LVI, II. ' Min. Sc. Press, 1910, c, 579. 'Min. Sc. Press, 1913, cvii, 573. 'Austin, Tr. A. I. M. E., 1906, xxxvii, 468. Ofiferhaus, Eng. Min. J ., 1908, lxxxv, 1234; Detail Drawing. Mathewson, Tr. A. I. M. E., 1912, XLrv, 781. 244 METALLURGY OF COPPER on penetration of the heat, the concrete has tendency to rise and affects unfavor- ably the silica bottom. The omission of underground flues, shown in Fig. 266-271 and formerly deemed necessary, is now characteristic for most large reverberatory furnaces which treat ore and store considerable amounts of matte going direct to con- verters; the principle followed is to make the bottom strong and to keep it- hot; air cooling still prevails with matting and copper-refining furnaces which are tapped at short intervals when enough matte has been accumulated, or the cop- per is ready for molding. The parallel sides of the furnace have ten working doors, 8X15 in., the sills of which are 18 in. above the skimming plate at the flue-end. The roof, made of iSX6X3-in. silica-brick, has near the fire-bridge only four feed-openings for ore, as the mode of charging difiers from that of the older furnaces, and four Figs. 272-274. — Reverberatory matting furnace and boilers, Anaconda, Montana. ports, 12 in. sq., for coal. There are twenty jXs-in. holes in the roof above the fire-bridge for the admission of air to assist in the combustion of the fire- gases. An expansion-space of \ in. to running foot is provided for by the intro- duction of cardboard. In the front wall of the fire-box are two circular openings, 12 in. in diameter, closed by wheel-doors, which serve to ascertain the height of the bed of coal on the grate. The furnace is ironed by 8-in. i8-lb. I-beams tied only at the tops by 2-in. rods, the bottoms being secured in the foundation. The waste heat of the gases is utilized by two 300-h.p. Stirling boilers placed in series; the gases leave ,the hearth at a temperature of 1100° C, enter the tubes at 950° C. and issue from them at 330° C; the draft ranges from 1.7 to 2.0 in. water. The results attained are given in Table 41; the mode of operating is discussed in § 149. The coal-fired furnaces at Garfield^ and especially the oil-fired furnaces have economizers between the boilers and the stack. ' Eng. Min. J., 1911, xci, 752. SMELTING OF COPPER 245 132. Furnace of the Canadian Copper Mining Co., Copper Cliff, Ont.,i 1912. —The firing end of this furnace is shown in horizontal and vertical longitudinal Figs. 275-276. — Reverberatory matting furnace, Canadian Copper Co., firing-end. sections in Figs. 275 and 276, the flue-end in Fig. 277, and some detail of the side wall in Fig. 278. The leading points wherein the furnace differs from others ' Browne, Tr. Canad. Inst. Min. Eng., 19 12, xv, 114; Private Notes, 1912. 246 METALLURGY OF COPPER are, that the working hearth is of magnesite brick; that the inner sides are of the same material excepting at the slag-level, where it is replaced by chrome brick; and that it is fired with fuel-dust. The hearth is 112 by 19 ft., has parallel sides for a distance of 83 ft. 3 in., which then contract in a distance of 28 ft. 9 in. to 7 ft. 9 in. The inner height at the firing end is 6 ft. ; there are nine doors on a side distributed over the par- allel sides; the side walls are 27 in. thick, of which 18 in. are of silica-brick sup- porting the roof; inside of the silica-brick is a 9-in. course of magnesite brick with chrome brick at the slag level, Fig. 278. The roof of silica-brick is 20 in. thick for the first 35 ft. near the burners, and 15 in. for the remainder of the way. The hearth of magnesite brick slopes from the ends to the lowest point, which is 11 Btept^B^ a 11^ T IS „ . O'O' , \ i lOBt«I«i@!IJi.e'l« \ -+ Fig. 277.— Reverberatory matting furnace, Canadian Copper Co., flue-end. 28 ft. 8i in. distant from the burners; here is placed one matte tap 12 in. above the floor, and 5 ft. 7 in. nearer the burners a second, flush with the bottom. On either side of the furnace near the matte taps is an in-pour for converter slag arriving in a tunnel, through which also the matte is hauled to the converters; flue-dust, mixed with green-ore fines, is charged through two hoppers in the roof which discharge along the center line of the furnace. The waste slag is removed from both sides about 11 ft. from the point at which the side walls begin to converge. The foundation for the working hearth is of slag, 2 ft. deep, poured into place between the outer walls and shaped to the form of an inverted arch with a spring of 1 2 in. On this are placed a 2i-in. course of fire-brick, and then a 9-in. course of magnesite brick laid with a mortar of ground magnesite and Unseed oil; an expansion space of i in. to the foot is allowed for by using wooden strips, i in thick, every six courses, and having them staggered. The firing is taken up in § 142. The throat through which the gases leave the furnace is about 27 feet square. They pass, Fig. 277, into a cross-over flue, 6 by 9 ft. and covered by cramps, lead- SMELTING OF COPPER 247 ing into the main dust flues (15 by 19 by 177 ft.) of the two reverberatories, which is connected with the stack 200 ft. high, 17 ft. 2 in. in diameter at the bot- tom and IS ft. 4 in. at the top. Very little dust is collected in the chamber. There are charged hourly about 10 tons of flue-dust with green-ore fines and 16 tons of converter slag; the coal is shut off when dust is fed, and this is allowed to drop from the feed-hopper until it ceases to float off on the slag. The flame is about 100 ft. long, the temperature at the fire-end is 1550° C, at the skimming- SiaenesLte Olirome Silica Tire Brick Fig. 278. — Reverberatory matting furnace, Canadian Copper Co., bricking of side-wall. door 980° and in the- stack 500° C. Fine ore is used as fettling. Two furnaces treat in 24 hr. about 400 tons of converter slag and 125 tons of flue-dust. Data are given in Table 51. 133' Furnace of the Cananea Consolidated Copper Co.,i Cananea, Sonora, Mexico. — This furnace, fired with oil, is shown in Figs. 279-287. General data ' Ricketts, Tr. Inst. Min. Met., 1909-10, XDC, 147; Eng. Min. J., 1910, Lxxxrx, 315, 404 (Heywood), 619 (Collins), 827 (Grabill), 959 '(Gormly), 1021, 1063 (Moore); 1911, xcii, 693. Mathewson, Tr, A. I, M. E., 1912, xliv, 781. 248 METALLURGY OF COPPER SMELTING OF COPPER 249 2 50 METALLURGY OF COPPER No.2 k 48'0- m M ii ] 16'. EIM66 mm'mTm Figs. 288-301. — General data of oil-fired reverberatory matting furnace, Cananea. Area, Sq. Ft. by 19 ft. Two reverberatory furnaces, Hearth, K. K Seven, burner-holes, 6 by lo in 3 Four peep-holes, lo-in. dfameter 2 Throat of furnace No. i, 5 ft. wide, 2 ft. 8 in. high 14 Throat of furnace No. 2, 7 ft. wide, i ft. 9 in. high 14 Furnaces Nos. i and 2, throats 28 Cross-over flue No. i, 5 ft. wide, 5 ft. high, SO ft. long 28 Cross-over flue No. 2, 7 ft. wide, 6 ft. high, 27 ft. long 44 Flue to boiler at 1 1 60 Flue to boiler at 12 40 Flue from boilers at 13 76 Flue from boilers at 14 120 Boiler No. i, area entering, 34 sq. ft.; leaving. Boiler No. 2, area entering, 34 sq. ft. Boiler No. 3, area entering, 34 sq. ft. Boiler No. 4, area entering, 38 sq. ft. Boiler No. S, area entering, 38 sq. ft. Boiler No. 6, area entering, 30 sq. ft. Boiler No. 7, area entering, 30 sq. ft. Boiler No. 8, area entering, 30 sq. ft. Boilers, Total area entering, 288 sq. ing 151 Boilers Nos. i, 2, 3, Aultmann & Taylor, 250 h.-p., each. General Data, Area, Sq. Ft. Boilers Nos. 4 and 5. Stirling, 250 h.-p., each. Boilers Nos. 6, 7. 8, Stirling, 300-h.-p., each. Economizer No. i, area entering, 83 sq. ft.; leaving. tR leaving. 18 leaving . 20 leaving. 20 leaving . IQ leaving. 10 leaving . ig ft.; leav- leaving ■■ ■ 45 Economizer No. 2, area entering, 83 sq. ft.; leaving _ ■ ■ • 4S Economizer No. 3, area entering, 83 sq. ft.; leaving ■ • ■ 45 Economizer No. 4; area entering, 83 sq. ft.; I eaving 45 Underground flue ^^5 Stack, 12 ft. 6 in., 171 ft. i 1/2 in. high, base. 123 Smallest area for passing gases, 14 sq. ft.; largest area '^3 California crude oil used, S. G., 0.966. Weight per barrel of 42 gal., 339 lb. Heat-content per lb., 18,700 B.t.u. Economizers, Green Economizer Co., 288 tubes. . Seven oil-burners per furnace. 0-43 lb. of steam required to atomize i lb. of oil. . Total travel of gases in furnace No. i. ■ ■ -690 ^J* Total travel of gases in furnace No. 2. . ■ -730 "• Draft. Temperature of gases at 9 and 10, 2,300"- 2,400" F 0.25 Temperature of gases at, 14, 500° F O-60 Temperature of gases at 18, 350° F 0.9O SMELTING OF COPPER 251 a o 252 METALLURGY OF COPPER SMELTING OF COPPER 253 are assembled with Figs. 288-301. The hearth, 100 by 19 ft., is contracted com- paratively rapidly at the flue-end. At the opposite end are the openings for four oil-burners. The six ore-hoppers originally used have recently been reduced to three. The inner parts of the hearth are of silica-brick. The fire-gases pass 56 * Kallg Figs. 304-305. — Gas-fired reverberatory matting furnace, Great Falls. in parallel through two 300-h.p. Stirling boilers; there is no by-pass. Crediting the furnace with 0.464 bbl. oil for steam raised in the waste-heat boilers, the total oil consumed per ton or ore of 1.121 bbl., given in Table 51, is reduced to 0.657 bbl. 254 METALLURGY OF COPPER 134, Fumaceof theAnaconda Copper MiningCo.,GreatFalls,Mont.' — This furnace fired with producer gas and provided with regenerative chambers, is shown in Figs. 302-305, selected from the ten published by Mathewson.'' It has a hearth 42 ft. 6 in. long by 1 5 ft. 9 in. wide, which is rectangular. Fig. 302 ; this gives it a larger hearth-area than if the sides are tapered at the ends, but is accompanied by the drawback that the sides tend to bulge inwardly instead of outwardly as is the case with oval side-lines. The furnace is built in five sec- tions in order to allow for expansion amounting to 14 in. over all. The sections are separate in the roof; in the side walls the necessary spaces are left open be- tween the bricks. The roof, 9 and 12 in. thick, is of silica-brick; the sides, 18 in. thick, are of fire-brick. The upper tie-rods. Fig. 303, 1.5 in. in diameter, pass i ft. above the roof, as this rises as much as 9 in. when hot; the lower tie-rods pass through open flues, 2 ft. 6 in. by 3 ft. 8 in., and are thus kept cool. The checker work. Fig. 303, at either end of the furnace has a special construction to permit cleaning and thus to reduce the slagging of the checker-brick by ore-dust. The fire-brick flues in the checkers are horizontal, 1 8 in. high by 7 in. wide and 11 to 2 1 ft. long, according to position. The 9-in. brick forming the tops and bottoms of the flues are laid 4.5 in. apart in order to leave open spaces for the dust to fall through and collect in the pit at the bottom, 4 to 5 ft. deep, whence it is removed periodically. The air-chambers, Fig. 303, are 8 ft. 6 in. by 14 ft. 7I in. and 13 ft. 9 in. high, and have each, including the exposed end-walls, a heating surface of 4900 sq. ft.; the gas chambers are 7 ft. by 14 ft. 7J in. and 13 ft. 5 in. high, each with a heating surface of 4200 sq. ft. The air. Fig. 304, enters the hearth through a single port, 15 ft. 9 in. by 2 ft. next to the roof; the gas through four ports, 20.25 by 30 in., beneath. The products of combustion from the fur- nace descend in a chamber. Fig. 302, at one end of the checkers, pass through these and ascend at the opposite end. The life of the checkers is about three months. It can be increased by allowing the gases to travel during the charging- period through a by-pass direct into the main flue and thus prevent the dust from settling in the checker-flues. In the roof of the furnace are seven openings through which the charge, mainly hot calcines, is dropped into the furnace from corresponding hoppers. The end-hoppers are first emptied in order that the charge may flow toward the center, then follow the central hopper and its two neighbors, and lastly the two remaining, the flow being regulated by gates in order to insure an even distribution. In the middle of one of the sides. Figs. 302 and 305, is a spout for tapping slag, which is granulated, and matte, which is collected in ladles and transferred to the converting department. Close to this spout is a tap-hole, Figs. 302 and 304, serving to remove the total liquid content of the hearth. In Figs. 303 and 305 are seen three ports leading into the air- flue. Fig. 302 shows the positions, of the gas-flue bringing the gas (CO2 9.4, CO 16.8, CH4 2.9, H 13.3, N 57.6 per cent, vol.) from six producers gasifying together in 24 hr. from 60 to 75 tons of bituminous coal (H2O 4-6, Vol. H-C 1 Hofman, Tr. A. I. M. E., 1904, xxxrv, 258. Mathewson, op. cit., 1912, XLiv, 781. ^Loc cit. SMELTING OF COPPER 255 21-27, F.C. 48-53. Ash 19-33 per cent.); of the air-flue, connected withablower; and of both the gas- and air-reversing valves. The work done with this furnace is given in Tables 41-42. Practical experience has shown that the surfaces of the regenerators are much too small. In the new smeltery ^.t Great Falls which is to replace the present one, each furnace will have its own gas-producer, fed and cleaned me- chanically; the heat of fire-gases will be utilized in recuperators. The predecessors of the present stationary furnaces were tilting furnaces erected in 1890 and 1892, which were similar to the Campbell open-hearth steel furnace. They were abandoned on account of their small capacities (hearth 16 by 13 ft.) and because of the fact that it was impossible, in pouring slag, to control the flow in such a manner as to prevent matte from flowing off with the slag. 135. Other Furnaces. — The leading facts about other furnaces are given in Tables 51 and 52. Drawings of other modern furnaces have been published by Mathewson.* The furnace at the Peyton Chemical Works, San Francisco,'' is an oil-fired furnace with regenerators for preheating the air, an arrangement which has much to commend it. In recent years a few special furnaces of passing importance have been put on the market, such as the Fink smelter,^ the Cotton furnace,* the Brown furnace,^ and the Dawson furnace.^ Examples of the older reverberatory furnaces have been tabulated by Howe.' 136. Accessory Apparatus. — The accessory apparatus for the collection and disposal of slag (§95) and matte (§96) are the same as with the blast-furnace. The recovery of flue-dust (§122) is also similar; its disposal (§123) differs in that it can be readily worked with the ore-charge. In fact, the reverberatory furnace often serves to work with its ore-charges the flue-dust produced by roasting and blast-furnaces. 137. Dimensions and Working Data of Reverberatory Furnaces. — The dimensions and work of the above reverberatory furnaces and some others are assembled in Tables 51 and 52. Table 51 had been prepared by the author when Table 52 was published by E. P. Mathewson. Both are given, as they treat the subject from different aspects; they supplement one another, even if there is some duplication. ' Loc cit. *Leas, Eng. Min. J., igo8 lxxxvi, 898; Pacific Foundry Co., San Francisco. ' Carter, Technical Work Mag., 1909, xi, 194. Traphagen, West. Chem. and Met., 1909, v, 172. Fink, London Min. J., 1909, Lxxxv, 237; Electrochem. and Met. Ind., 1909, vii, 287; Mines and Methods, 1912, iv, 63; Min. Eng. World, I9i2,xxxvii, 795. Neill, Min. Sc. Press, 1909, xcviii, 300. Pulsifer, Min. Eng. World, 1913, xxxvin, 952. * Editor, West Chem. and Met., 1909, v, 325. ^ Mines and Methods, 1911, II, 167. 'Min. World, I9ii,xxxiv, 691. ' Bulletin No. 26, U. S. Geol. Surv., Washington, 1885, p. 40. 256 METALLURGY OF COPPER Table 51. — Reverber Item Coal-fired Wales, 1848 Butte & Boston, 1903 Anaconda, 1903 Colorado S. Co,, 1903 Butte Red. W'ks.. 1903 Boston & Montana, 1903 Length of hearth Length "a" of bridge-section.. . . Length "b" of middle-section. . , Length "c" flue-section Width "d" of hearth at bridge. Width *'e" of hearth at middle. Width "f " of hearth at flue Hearth area, sq. ft Hearth thickness 13' Length of grate — Burners, No Width of grate — Burners, diam., in. Depth of grate below top of bridge at bridge — Air, cu. ft. per min. Depth of grate below top of bridge at opposite end — Air, pressure, lb. Grate area, sq. ft Ratio, hearth to grate area Height of roof above bridge Height of roof above hearth at bridge, Height of roof above hearth at flue. , Width of bridge Size of flue at vulcatory or verb Size of flue leading to chimney., Chimney, inside diameter Chimney height. . Charge, tons Charge, time of melting, hr.. Charge, tons in 24 hr Charge, tons per sq. ft. of hearth in 24 hr. Ratio of concentration Fuel, bituminous coal, manner of firing, oil. Per cent, of ash , Percent, fixed carbon... Oil deg. B6 Tons charge per ton coal., Tons charge per bbl. oil.. , Labor in 8 hr. shift 7' 4i" 8' II" 2' 3J" 120 18" 4' 8" 4' 3i" 3' n" 3'S" 20 6 :l I' 73" 3 I' Ai" 2' 6" IS"X? 22" X? 2' 2}" 46' 7J" 1.43 8.6 0.071 3:1 Bit. & anth. Direct 2 & 7 68 & 76 0.77:1 SO' 6' 28' 16' 4 842 10' S' Si" I' 10" t' 2" 53.9 IS. 6:1 2' 7" 4' 8" 2' li" 3' 0" 6'X3o" 30"X48" S'6" 70' 25 112. 5 0. 129 5.7:1 Direct 8.9 51.9 so' 6' 32' 6' 886 24" S' 4" 2' 9i" 2' 6i" S3. 3 16.6:1 2' 6J" 4' 4" 2' 10" 3' si" 6'X30" 30"X30" 6' 3" IS 3h lOS 0.118 4.7:1 Direct iS.u 44 -S 49' 6" 4' 6" 32' 13' 10' 9" 20' 2" 4' 878 19" 7' S'6" 2' 4" I' 4" S3.0 l6.s:i 2' 10" 4' 4" 3' o" 3' o" S'X30" 28"X30" 6' o" 70' 24 6i 90 O. 102 7.56:1 Direct 14.85 44.5 SO' 6' 32' 4 878 10' 9" 5' 53.75 16.3:1 2' 8J" 4' 8" 2' 2" 3' o" 6'X30" 30"X30" 6' I" 75' 3j" 18 90 0. 102 S:i Direct 5.0 SS." Matte, Cu, per cent Matte, sp. gr Slag, SiOa, per cent Slag, Fe(Mn)0, per cent. Slag, AUOa per cent Slag, ZnO, per cent Slag, Ca(Mg)0, per cent. Slag, Cu, per cent 33.7 4.56 60.5 28. s 2.9 Other ox- ides 1 . 4 3.05:1 2 & 2 53-8 36.8 SI. 9 8.4 48.3 4.8 42.8 47.31 7.5-80 4:1 2 &'2J' SO 2.81:1 i'l'&ii' 50 34. s 43.0 8.5 38. 6 51.4 i-S Slag, Ag., oz per ton. Slags, sp.gr 1. 1 0.75 0.4s 1 .2 0.40 0.30 3. 58 0.60 I .0 3.3 0.40 U.40 42"dia. at each end 42' 6" 42' 6" 42' 6" 42' 6" IS' 9" IS' 9" IS' 9" 688 28" Producer gas regen. chamber 24" 7' 0" 4' ll" 35 6 150 0.218 3i-l Prod, gas 17.0-30 45." 2:1 50 4-77 41.9 42.7 10.9 I.I 0.58 3.54 SMELTING OF COPPER 257 ATORY Furnaces Dust-fired Oil-fired Mont Ore Purchasing Co., 1903 Anaconda, 1908 Garfield, 1911, coal. Canadian Copper Co., 1912 Cananea, 1911 Garfield, 1913 Hayden, 1912 Copper Queen, 1913 Steptoe Valley, 1912 ]- 112' 4" 6' 71' 4" 16' 7' 2' 3" 2'0" 112 18:1 5' a" 7' 6' 3' 31" 4' I" S'X3'2" 6'X3'8" Main stack of plant do 15 every 80 min. Continuous 300 0.150 13.4:1 Direct II 56 112' 29' S" 47' 5" 35' 2" 19' 19' 7 1959.5 24" 18' 6' 5" 2' 2" l' 10" 112. 7S 17.3:1 ^;?|" 7'X3'-3" 8'XI2' 30' 0" 300 l8± Continuous 250-390 0. 163 Direct 6 48 112' 0" 83' 3" 83' 3" 28' 9" 19' 0" 19' 0" 7' 9" 1970 9" magnesite S 6" 100' 112' 29' 5" 47' 5" 35 2" 19' 0" 19' 0" 7' i960 24" 5 li" 1000 10 112' 3" 8'o" 73' 3" 37' 11" 16' ' 19' 0" 7' 1967 24" 7 (3 in use) H" 92' 7' 10' si" so' si" 31' 8" 17' 3" 19' 0" 8' 1576 24 5 120' loj" 94' oi" 94' oj" 26' 10" 18' 11" 18' 11" 8' io"> 2125 I..'. 8' 0" 14' 6" 82' 9" 17' 3" 19' 19' 8' 1688.7 30" 7 (burners) air 2" oil i" 22" 8' 6' 3'o" 22" quartz 12" silica brick 7 A-i A" 400-600 15 10 13-14 12-14 10 48.0 4' 7' l'\" Throat 7'- 9"X3'6" 2 furnaces I5'XI9' 17' 2" 200' 16 slag 10 flue-dust Continuous 525 0.27 6! •d W rt W B > w f^ 3 ClJtj OS 6? 6? 00 2 °o H 00 A oqo %o62 • »o xxg- :i?v„o ! 0>« f*3 M "0 I/) Qj aj (I) (u *v CO Co Co ra nj ;foooo o iS,, Odiia, xxg 0°'?". o 32 wJ - c o u O « O " H 0) OJ QJ Ol /^ o o o:i„ "-■ o^^:^;^o X ■J^v ^ . CM C^ "^"-""xx^? C»vO -H l§; 6 OR •^-a I ■ M t 3 . (^ ■ ot; o • '-' 000 p, O O O _r3 POOO o 10 N r*^^o g* bO i-i O " Wi''^ «. "-I t^ rtji tj ft a 7! O j> O -|j o fi j^ (D >>> S O dj s > oS3 til 13 bD "o.S ^-1 °— -'■SjJ •-■o h u a) 3 m rt c S C CO .^ ffl 2 rt «^ o K 3 "OOfoOmh M tfi +J -4J- m c c cfl rt .a o o f« t'„'*s ^ S S S o oj3 ftg aBoi&.s? a •j--^ O O "U •- PQHHM P _ -> s cJ h/l frt rr! o u u y (-. Cfl O •i-iij aj +^13 cij bOcd*53 fl *r] O I-l • =: L^ t-< ^ 6-^ 5 >^ " C g rt •s-g-s •S^S' ■•s-s.s|ggr ■» a s B c« .§ 2 as Q OQQ EB-Sf in M W Q. rt ri ™ '7 bo bO bo^ ^^ ii V- n S :J 3 n rt ctl cS ., += OJ qj u 00 Qj sees > 0) V C O o o 2 S o o h iH iH R « *"*" C a> ^^^ o u Q) 0) . " O 0,2 CI <-■ t- tJ s QJ ID ™ SMELTING OF COPPER 259 89 0. O 3 < T3 a O f=fi§ ft" o I. O 4* 10 »fl>o ^ (no f- r- Oi f) t-eo r- Oi I-" O Oi OiO OOOOOOOO NfOOOOOO ui u^ N 1000 -ftoo >ooo>oio>oato PTN^MloiOf*) tJ-(*5ihi-00^ •O ■* w O O O O noo -^OO OOfO ■^t-'^OOOO TlT^Oii'iOiMO 00 lO-O wJOiOiO\ uiMiyiOO"-"© OMONirt'^i') NfOroOoo ino O t^ m r* Oi Oi (^ 00 M 00 00 O HI O "-< O O M 0000000 0) a),»rt»^oo i-H y y o o o ^' 22000 ; O O O O O O vOMr^'*)OOt*3 OOOOOOO oOminOhiOi^ Oi-O Oi N fO (*) N ooooooo *-* *■ ■* M (o Tf -fr Tj- (N o fl ^ ^ "I rt C 0) C cil u «) -^ 13 ft^riJ o 34^1) P.«?ii iH^'SO <5C4)U*'KO ri+j°0 artiHd+>"o CO C ft) I- ^ o O f iiU 1 S t-sO t^O 9 woo o>« 10 +j w m M •d " POOO pq ^2^:^" ja rt ,; r* 000 r- >-i ft M M V CU c XI •d tC -^-lo in c -t Tf "!j in 't tl4 S C) 000 > <*3 N fo f*3 fn ni a 0^ 1) 3 n d rO 000 t-O >o r-oo N (h & 1 V u >** ^ rt : aj -gS^s^^S ■ ofl4-Uay?iJ ■ o 0, 2 72 METALLURGY OF COPPER hearth 100X19 ft., the furnace lasts six to seven months. For the same reason the distance between roof and floor has to be greater with the oil- than with the coal-fired fiurnace, viz., 6-75 vs. 5 ft. In Fig. 307 is represented a temperature-record taken at the vulcatory or verb of an oil-fired reverberatory furnace of the Steptoe Valley smeltery. About Btreet K11 Botes to clean Kozzle by Steam Figs. 308-309. — Shelby oil-burner, original. every hour there are dropped about 14 tons of charge near the firing-end of the furnace, which causes the temperature to fall; skimming and claying have simi- lar effects. The burners used in matting reverberatories have various forms, each smeltery having developed a type that suits best its own purposes. Two may serve as examples. Oil Fig. 310. — Shelby oil-burner, modified. The Shelby burner of the Cananea Smeltery,^ is shown in Figs. 308-310. It is an atomizing burner of the chamber type using steam, mounted on a Moran ball-joint. The oil enters, at 40-lb. pressure, an annular chamber which has an orifice ^ in. in diameter. There it meets steam of 125 lb. pressure which issues through an annular opening -5-^ in. wide (|| in. outer and | in. inner diameter) at right angles to the flow of the oil. The flow of steam is regulated by a needle- ' Eng. Min. J., 1910, Lxxxrx, 31. SMELTING OF COPPER 273 valve with rod held by a screw-thread and moved by a hand- wheel. In Fig. 310 a simpler form of nozzle is shown which throws the flame farther into the furnace; the steam-inlet has been changed from the original ring form to a cylin- drical port of -jV in- bore; atomized oil and steam pass through a pipe -^ in. inner and f in. outer diameter. A burner handles 50-60 bbl. oil in 24 hr., -4"/io ^ Figs. 311-312.— Sorensen oil-burner, high-pressure air. requires 0.3-0.4 lb. steam per lb. of oil excluding the 2^ per cent, which it consumes. The Sorensen burner oe the Stepioe Valley smeltery is also an atomiz- ing burner and uses high-pressure air. It is shown in Figs. 31 1-3 12. A |-in. oil-pipe with conical nozzle is held in position by three adjusting screws in a il-in. air-pipe provided with a flaring nozzle. The general arrangement of the 274 METALLURGY OF COPPER SMELTING OF COPPER 27s supsended pipes at the Steptoe furnaces is shown in Fig. 3 13. Above the burner- ports the bricks are laid so as to leave open vertical slits through which air is to enter in order to furnish additional oxygen, but they very soon become more or less slagged over. The seven pipes of the furnace have burnt in 24 hr. 421 bbl. oil, or 60 bbl. per burner, at an air-pressure of 15 lb. 145. Slag. — The slags produced in reverberatory furnaces are nearly always more acid than those in blast-furnaces; they may also differ considerably in constitution. In the blast-furnace all the SiOa must be combined with bases to form a fluid slag; in the reverberatory furnace this is not necessary, as the slag can be and is drawn off, and need not be run off. Slags low in Cu are made, consisting of a moderately fusible silicate-mixtiu-e as matrix, holding in suspen- sion particles of quartz, which readily separate from the matte. The Welsh slag (Table 51) given by Le Play as containing SiOj 60.5 per cent, is such a porphy- ritic slag, 30 per cent, is combined and 30.5 per cent, free SiOj. Such slags used to be made more frequently than at present, when the desire for increased smelting power is attained more by high temperatures with completely fluxed SiOa than by a saving of fuel accompanied by a porphyritic slag. All re- verberatory furnace slags must be acid in order to avoid rapid corrosion of the siliceous furnace-lining^ and quick cooling; they must be moderately fusible so as not to require too much fuel or time in forming; they must not be too fluid, as this hinders skimming, nor too viscid, as this prevents matte from settling; finally, the specific gravity ought not to be too high, as it would hinder the separation of matte. All slags, then, run high in Si02, usually not above Si02 45 per cent., rarely below SiOa 36 per cent.; the percentage of CaO used to be lower than is the case to-day. Examples are given in Tables 51 and 52. Fulton* has investigated the constitutions, melting-points, and fluidities of some reverberatory slags made in matting copper ores. The work embodies three series of experiments with three basal slags of the compositions given in Table 56. The slags forming the bases of series I and II run high in AI2O3; Table 56. — Compositions of Basal Revekbeeatoey Slags. — (Fulton) Number Composition Ratio of of series SiOa FeO AI2O3 CaO SiOz : FeO : AI2O3 I Per cent. 42.6s 44-50 44-30 Per cent. 36.12 38.20 43-1 Per cent. 11.30 11.8s 4-8s Per cent. 6.20 5-IO 6.80 377:319: 100 376 : 323 : 100 914 : 888 : 100 11 Ill those of series I-III contain increasing amounts of FeO. In each series the percentage of CaO has been increased and the silicate degree correspondingly decreased. Table 57 gives the numerical results of the experiments in series I, and Fig. 314 the plot; Table 58 stands for series II and Fig. 315 for the plot; Table 59 for series III and Fig. 316 for the plot. ' Tlie magnesite lining of Copper Cliff forms tlie exception. ' Tr. A. I. M. E., 1912, XLiv, 751. 276 METALLURGY OF COPPER Table 57.— Compositions and Melting-points of Series I of Fulton's Reverbera- TORY Slags Composition Silicate, degree Melting- Number SiOa FeO CaO AI2O5 Total point, deg. C. Per cent. Per cent. Per cent. Per cent. Per cent. I 42.65 36.12 6. 20 11.30 96-37 1 .40 979 2 40.66 34-33 10.80 10.70 96.49 1.28 lOII 3 38.80 32.60 14 -9S 10.02 96-37 1. 19 1060 4 36.90 31.10 18.52 9-75 96.27 1.09 108 1 S 35-40 29.65 22.12 9-3S 96.52 1.02 "37 6 33-92 28.48 25-14 9.00 96-54 0.96 1151 1,225 , 1.175 1 1.150 6 1.126 S 1.100 s ^•"''^ ,^ r*' / / / / Q / / 1,000 975 _ ,/«' ^ r^ 6.20 10.80 14.95 18.52 22.1225.14 Per cent Lime 1.40 1.30 1.20 1.10 1.00 0.90 Silicate Degree Fig. 314. — Melting-point curve, first series, Table 57. Table 58. — Compositions and Melting-points of Series II of Fulton's Reverbera- TORY Slags Number Composition SiOa FeO CaO AI2O3 Total Silicate degree Melting- point, deg. C. Per cent. Per cent. Per cent. Per cent. Per cent. I 44-5° 38.20 5 -10 11.85 99-65 i-Si 1050 2 44.00 37-80 6. 14 11-73 99 67 1.48 1025 3 43-5° 37-30 7-iS 11.60 99 55 1-45 988 4 43.20 37.00 8.16 II. 51 99 87 1.42 982 5 42.25 36.20 10.04 11.23 99 72 1.38 1010 6 40.50 34.65 13.60 10.75 99 5° 1.28 J0S4 7 38.90 33-25 16.80 10.40 99 3.'? 1.20 1091 8 37-55 32.10 19.90 10.05 99 60 I -13 1103 9 36.20 30-807 22.56 9-67 99 30 1.06 1107 SMELTING OF COPPER 277 Silicate Desrrce Fig. 315.— Melting-point and fluidity curves, second series, Table 58. Table 59.— Compositions and Melting-points of Series III or Fulton's Rzverbera- TORY Slags Composition Silicate Melting- Number Si02 FeO CaO AI2O3 Total degree point, deg. C. Per cent. Per cent. Per cent. Per cent. Per cent. I 44-30 43-10 6.80 4-8S 99-05 1.66 1000 2 42.6s 4i-.';i 10.29 4.69 99.14 54 lOIO 3 41 . XI 40.01 13 -S3 4-Si 99.16 44 1034 4 38-34 37.22 I9-3S 4.21 99.12 27 1097 S 35-92 34-97 24-43 3-94 99.26 13 1167 6 33-79 32.89 28.92 3-71 99-31 03 1207 7 31.90 310S 32.90 3-50 99-35 934 1186 1.225 1,200 U75 .§ 1,.150 1 1,125 J 1,100 1 1,076 1 1.060 n 1.025 1,000 976 / ^ i:^ N xi / /' \ \ \ V^.*J f- K ^ / V \ P '' / V , > V .<^> / ,^°' t> y -.^ ^i^ r V N ^ i^ v^ 6.80 1.70 10.29 13.53 19.35 24.43 28.92 32.90 Per cent Lime 1.60 1.50 1.40 1.30 1.20 1.10 1.00 0.90 Silicate Degree Fig. 316. — Melting-point and fluidity curves, third series, Table 59. 278 METALLURGY OF COPPER In Figs. 315 and 316 the fluidity curves represent the temperatures at which the slags held in platinum crucibles were sufficiently fluid to permit easy stirring with a heavy platinum rod. Curves in Figs. 315 and 316 are V-shaped; in Fig. 314 the eutectic point was missed. The lowest melting-points of the three curves, Figs. 3 14-3 16, lie between silications i. 4-1 .65, depending upon the character of the several bases. Fig. 314 shows that in a highly aluminous slag an increase of CaO alone causes a rise in the melting-point; this confirms the experience in practice of having to raise both the SiOa- and CaO-content for the lowering of the melting-point. The curve in Fig. 315, which covers a higher range of SiOa-content (44.50-36.20 per cent.) than the curve in Fig. 314 (42.65-33.92 per cent.), shows that an addition of CaO lowers the melting-point until a minimum has been reached as was to be expected. The curve in Fig. 3 16, beginning with a lower percentage of Si02 than in Fig. 315 (44.30 vs. 44.50), reaches the lowest melting-point earlier. 146. Chemistry.' — It has beenstatedin § 127 that in thecopper reverberatory furnace S was the reducing agent. The main reactions as regards Cu are reduc- tion by means of S, and sulphurization by means of FeS; and as regards the gangue, slagging in a partially oxidizing atmosphere. Reduction of Cu may be expressed by Cu2S-|-2CuO = 4Cu-|-S02; CU2S-I- 2Cu20 = 6Cu-|-S02; Cu2S-F3CuO = 3Cu-+-Cu20-|-S02; Cu2S-|-6CuO = 4Cu20+ SO2. The first two reactions^ begin at about 500° C, and all SO2 is set free at 1000°. Sulphurization may be formulated by 2Cu-f-FeS<=*Cu2S-f Fe; CU2O+ FeS = Cu2S-|-FeO; 6CuO+4FeS = 3Cu2S-|-4FeO-|-S02; Cu2Si03+FeS = Cu2S -HFeSiOa; Cu6Si30i2-|-4FeS=3Cu2S-|-Fe4Si30io-|-S02. The recent investigations of Juschkewitsch* having established the for- mation of the compound (Cu2S)2.FeS in a pyrometallurgical process, the reactions between CU2S, CuzO and Fe, FeS assume the following forms: (Cu2S)2.FeS-t-2Cu = 3Cu2S-|-Fe; 2Cu2O-|-3FeS = (Cu2S)2.FeS-i-2Fe0; and i2CuO-|-ioFeS = 3(Cu2S)2.FeS-f7FeO-|-S02-j-30. The decomposition of CuO by Fe according to Friedrich^ begins at about Bases 740° C. Reduction of Iron Oxide may be expressed by 3Fe203-|-FeS = 7FeO -fS02; Fe203-f-Fe=3FeO; Fe203-hCO=2FeO-f CO2; Fe203-f.heat+Si02= 2FeO-hO-fSi02; and slagging by FeO-fa;Si02 = FeSi,02x+i, and by Other -|->'Si02 = 0.-B. Si„02„+i. As the hot flame impinges upon a newly fed charge, the most fusible parts on the surface will melt first and passing downward will leave new parts exposed to the action of the flame. This goes on until the whole is fused. "Blanket Slag," or slag which forms on the charge and consists of partly fused material, ' Laist, Tr. A. I. M. E., 1912, xliv, 806 ^ Doeltz, 'Metallargie, igoy, iv, 421. ^ Metallurgie, 1912, ix, 543. * Stahl u. Eisen, igii, xxxi, 2041, SMELTING OF COPPER 279 gaveFultoni SiOz 52.08, FeO 27.96, AI2O., 9.00, CaO 5.90, MgOo.23,K(Na)20 i.oo per cent, and Si02 38.68, FeO 38.10, AI2O3 8.48, CaO 9.92, MgO 0.79, (NaK)20 1.00 per cent. This would not form if the charge components were well mixed, and they usually are not.^ In the meantime bubbles of SO2 rise from the pasty parts of the surface; they increase in number and stir up the unmelted and partly melted parts of the charge, which slowly decrease in quan- tity; matte will collect on the bottom and slag float on the top. The FeS held in solution by the slag sulphurizes any slagged Cu, so that the Cu finally remaining in the slag is present more in the form of a suspended matte-pellet or dissolved matte than of a copper silicate. The Zn and Pb of the charge are partly vola- tilized, partly matted and scorified; As and Sb are volatilized, matted, and slagged; precious metals are collected in the matte. The records of Gibb' show that in smelting roasted sulphide ore with Cu 10.60, As 0.102, Sb 0.025, Bi o.oio per cent., there was eliminated. As 62,8 per cent, Sb 57.6, Bi 20.2; As and Sb entered the slag, while Bi was expelled mainly by volatilization. Comparing these data with those obtained by a reducing fusion of the same ore in the blast-furnace (p 189), it is seen that the reverberatory removed more As and Sb than the blast-furnace, while the reverse was the case with Bi. 147. Calculations of Charge.*— The basis of calculation is the same as with the blast-furnace (§ 102, 115), i.e., the kind and quantity of matte and slag that are to be produced. The Cu is calculated as entering the matte a CU2S accom- panied by a given amount of FeS, which depends upon the grade of matte desired. It may be assumed that 75 per cent, of the Pb, 50 per cent, of the Zn, and 25 per cent, of the Mn will be found in the matte so that this contains perhaps only 95 per cent. CU2S, x FeS; the remaining constituents of the ore will enter the slag. Some flux, usually limestone, may have to be added to the charge to obtain the right degree of fusibility and fluidity. As the ore charged is fine, there is some loss (2 -|- per cent.) by dusting; further, in smelting, there is a vari- able loss of S, which, according to Vivian,^ in the old Swansea furnaces amounted to 13 per cent., and which, according to more recent observations by Peters' causes the matte to run about 8 per cent, higher in Cu than expected. Again, all the SiOa may not be scorified, part being carried out by the slag mechanically suspended, especially parts of the charge which are larger than ^ in. Lastly, the results obtained are governed to a considerable extent by the handling of the furnace. These variables are the reason why preliminary calculations give results which agree only approximately with the figures obtained in actual work. When the necessary factors have been obtained by experience, theory and practice will be in better harmony. Some preliminary factors for the first calculations can be obtained by making crucible fusions to find the quantities and grades of 'Loc. cit. ' See Demond, loc. cit. 'Tr. A. I. M. E., 1903, xxxiir, 657. * Walker, Eng. Min. J., 1906, LXXXi, 852. ° Eng. Min. J., 1881, xxxi, 249. •Modern Copper Smelting," 1895, p. 83. 28o METALLURGY OF COPPER matte collected in the form of buttons, and to form an idea of the fluidity of the slag. 148. Heat Balance of Reverberatory Matting Furnace at Anaconda.— The data of the subjoined heat-balance, kindly furnished by Mr. W. Wraith, are the result of an investigation, lasting loi hours, carried on in 1905 with furnace No. 2. This has a hearth 102 by 19 ft. (area 1807 sq. ft.,) and a grate 7 by 16 ft. (area 112 sq. ft.). There were smelted in 24 hr. 272 tons of material with a ratio of 4.37 tons of material to i ton of coal. The assay of the charge gave SiOa 26.1, FeO 40.2, CaO 2.9, Cu 9.0, S 8.1 per cent. Summary or Heat Distribution Debit B.t.u. Per cent, of total debited heat 5,520,200,000 293,500,000 94.92 Heat in calcine and flue-dust, above atmospheric temperature. 5.08 Credit 5,815,700,000 100.00 950,400,000 184,700,000 671,900,000 9,020,000 45,330,000 1,908,380,000 766,800,000 16. ?4 1 3.x8r'-5^ 11-55 0.16 Sensible hieat in droppings from grate above atmosplieric temperature. TTeat in steam from boilers 0.78 32.81 13.18 4,536,530,000 78.00 There have been accounted 78.0 per cent, of the heat with which the furnace is debited; the remaining 22 per cent, have to be sought in the heat which passed into the ground, and the heat due to slag-formation. A striking feature is the fact that only 19.52 per cent, of the heat is found in the slag and matte, while the two waste-heat boilers saved 32.81 per cent, devel- oping 567 h.p. The calcines which entered the furnace with a temperature of 950° F. form 5.08 per cent, of the heat with which the furnace is debited. Con- sidering that only 19.52 per cent, of the total heat appear in the slag and matte produced, the evidence is furnished that 5.08 : 19.52X100 = 26 per cent, more coal and time would be required if the calcine had been at atmospheric tempera- ture.'- The case actually would be more unfavorable if the manner in which a cold charge balls up in the furnace is considered. 149. Management.^ — Formerly all reverberatory ore-smelting furnaces 'See Demond, loc. cit. ''Lloyd, Eng. Min. J., 1904, Lxxvii, 716. Peters, Metallurgie, 1905, n, pp. 35, 63. Offerhaus, Eng. Min. J., 1908, lxxxv, 118 Moore, op. cit., 1910, Lxxxrx, 1021, 1063. 1234. SMELTING OF COPPER 281 received their entire charges at stated periods ranging from 4 to 8 hr. ; a charge was smelted, the slag drawn, and the matte tapped completely or only in part. Since the advent of the loo-ft. furnaces at Anaconda, this periodical charging of large quantities has been changed into feeding relatively small amounts at inter- vals of 20 to 60 min. as is common in still shorter periods with the blast-furnace. The second method, which is a radical departure from the first, is becoming the general practice in ore-smelting as fast as large furnaces are replacing the smaller* a. Feeding Large Charges at Long Intervals. — Presupposing the hearth to have been emptied, the charge is dropped from the hoppers in the roof. Roasted concentrates (calcines) are charged hot, whenever this is possible, and the fluxes added while charging calcines. If the charge is to be made up in the furnace, an undesirable procedure, ^ siliceous ore will be dropped first, the roasted ore next, and lastly foul slag. In a non-regenerative furnace the bulk of the charge is collected near or at the fire-bridge; in regenerative furnaces near or at the ends. If the desired distribution of charge cannot be effected through the hoppers, the charge will have to be spread by hand. After charging, the doors are closed and luted, and the fire is urged. A strong steady heat is essen- tial for quick smelting. The flame at first will soften readily-fusible components of the charge; these will sink and expose others to the high temperature of the furnace; matte will collect on the bottom and carry slag and unmelted parts of the charge. At intervals these will be rabbled from the skimming door, the bottom will be scraped to detach adhering lumps and to float them to the surface that they may be Hquefied. If the bottom feels rough and gritty instead of smooth and soapy, firing has to be continued, as the fusion is not completed. With a zinky charge it is essential to have the furnace especially hot toward the end of the fusion in order that mushy zinky matte which on account of its low specific gravity floats on top of the normal matte may separate completely from the slag. When a charge has been smelted, the slag will be rabbled to set free particles of matte which may have remained entangled. In order to free slag from scorified Cu, pyrite has been sprinkled over it; in a smiliar manner CuO has been introduced to have it react with FeS dissolved in the slag, then the matte formed in its passage downward through the slag reduces the silver-con- tent; at Argo, Colo., 75 per cent, of the Ag of the slag has been thus recovered. When fusion is completed and matte well separated from the slag, the skimming door is opened, perhaps also a working door, for a short time, to cool and stiffen the slag. The slag is skimmed into an overflow slag pot to collect any particles of matte that have been accidentally drawn out; the overflowing clean slag is collected in waste-slag pots and hauled to the dump or granulated. In former times, with smaller furnaces than used in recent years, a series of basin-shaped molds was formed in a sand bed beneath the skimming doors; the slag was drawn into the mold beneath the door and allowed to overflow into the neighbor- ing molds after breaking down the sand ridges separating the molds in order to form communicating channels. In the mold beneath the door was collected any matte drawn out with the slag; the "plate slag" floating on top of the matte 'See Demond, loc, cit. 282 METALLURGY OF COPPER was resmelted, perhaps also the cake on either side, as they contained pellets of matte. When the slag has been removed, the matte-tap is opened and the out- flowing matte collected in sand molds built along the side, or in cast-iron molds having overflow lips. Sometimes the matte is granulated by allowing a heavy jet of water to impinge upon it when it runs in a thin stream from the spout. The granules are collected in a deep basin from which the combined granulating arid cooling water flows through shallow settling tanks to collect floating par- particles. When the furnace has been emptied any damage to the hearth and the fettling is repaired before the next charge is introduced. Most furnaces were tapped dry after every second or third charge; at Argo, Colo., the matte-fall was so small that twenty-four and more charges were smelted before the matte was removed for further treatment. This holding of a bath of matte in the furnace is accompanied by so many advantages that it has become the common practice even in cases when the matte is to be shipped cold instead of going in the liquid state direct to the converter. With converting plants, the reverberatory furnace forms a reservoir for matte, forming a bed 3-14 in. deep, as does the settler of the blast-furnace (p. 169). As long as the furnace construction is strong, the bottom built up solid, the connection between sand- bottom and side wall made secure, there is no danger of matte getting beneath the sand-bottom and causing trouble. Such a mass of loosened furnace-bottom, "floater," gave upon analysis' SiOa 84.62, FeO 8.53, AI2O3 2.13, CaO 0.90, MgO trace, K(Na)20 i .00 per cent. The advantages of having a bath of matte on the hearth are: It prevents corrosion of the bottom, as basic constituents of the charge do not come in contact with it; it reduces the repair of the fettling from say once a day to once a month, because the slag line can be held at about the same level; it diminishes the repair of the furnace, as the temperature can be kept more even; it does away with the labor required for spreading the charge, or greatly reduces it; it diminishes the loss of heat, as the side doors need to be opened very little; it increases the smelting power by the reduced necessity of cooling the furnace and by the more rapid fusion of the charge when floating upon hot matte than when resting upon a partly cooled bottom. With all these improvements in the mode of operating, the dropping of large charges at long intervals has the drawback that the furnace, say 50 by 20 ft., is cooled considerably when it receives from 15 to 20 tons of charge every 4 or S hr. ; much time is taken up to bring the furnace up to the required temperature, and less time is given to actual smelting. Fig. 317 gives a chart showing the variations of temperature in smelting in an Anaconda ore-furnace of 1903. When the furnace had reached its maximum heat of 2750° F. at 10.40, and the new charge had been dropped and leveled, the temperature fell to 2150° F. at 11.00; it now took from 11.00 to 2.15 to gradually raise the temperature which reached 2850° F. when the fusion was completed. The introduction of the new charge caused the temperature to fall 550° F., whereupon the heating-up had to be begun again. The depressions in the curve also bring out clearly the falls of temperature caused by stoking (indicated by o) and rabbling. 1 Fulton, Tr. A. I. M. E., 1912, xliv, 751 SMELTING OF COPPER 283 If the time wasted in heating to fusion temperature can be saved, the smelting power is greatly increased. This is attained by feeding small charges frequently. b. Feeding Small Charges at Short Intervals. — This was first developed by E. P. Mathewson at Anaconda. ^ The furnace, Figs. 272-274, has two ore- hoppers discharging into four openings in the roof near the fire-bridge, and four coal-hoppers with corresponding openings in the roof above the fire-box. B ef ore dropping a charge, a side door near the ore-hoppers is opened; and a rabble inserted to ascertain whether there remains any unfused material ("floaters"). This is moved toward the fire-bridge^ and fused, or a pipe pushed underneath it a O "i. I S I 9 o Si! u 5 8 e^ « 2900 ^.2800 ^•2700 „2600 H2E00 ■§2400 §2300 a2200 12100 H2000 1900 Rabbled Gratins Rabbled Skimming C Skimming D Charge Droi Charge Lev< Removing S Grating Gratins Rabbled Rabbled Rabbled Gratins Charge Droi Charge Leve _ __ C\ __ t.-^ f^ 1 A^.i^^.i... ...^ .:s.7!,._Xii_vi,!, v^r . : ...siA.Ji \^..n^4----^^—--\r-\ k •^i \_L_^[ _L__w. .:__:: \y— < ^ Fig. 9.20 9.40 10.00 10.20 10.40 11.00 11.20 U.40 12.00 12.20 12.40 LOO L20 1.40 2.00 2.20 2.40 Time Koto: The Time of stoking is Indicated thus: o 317. — Variations in temperature of Anaconda reverberatory matting furnace, fed with large charges. and air forced through so that conversion of matte furnishes the heat and the base necessary to flux the siliceous floater. Now lo tons of charge are dropped from the first and 5 tons from the second hopper. The charge falls upon a bath of not less than 3 or 4 in. of matte covered by about 8 in. of slag, spreads more or less, and moves downward toward the flue, at first in two streams. In this way IS tons of charge are dropped into the furnace every 80 min. In the same man- ner 3000 lb. of coal are dropped every 40 min. into the fire-place through four op- enings, the level of the coal being ascertained by feeling with a rod through open- ings in the side. The temperature in the furnace is pretty uniform as shown in Fig. 318, the highest being 1550° C; the temperature of the slag as it leaves the furnace is 1 1 20° C. ; that of the gas 1060° C. ;* depressions in temperature caused by feeding of charges are not shown; they resemble those given in Fig. 307. •Austin, Tr. A.I.M. E., 1906, xxxvii, 470. Offerhaus, Eng. Min. J., 1908, lxxxv, 1191, detail records. ' Editor, Eng. Min. J., 1911, xci, 1243. 'Other measurements of furnace- and slag-temperatures: Clevenger, Met. Chem. Eng., I9I3.XI,447. 284 METALLURGY OF COPPER Slag is skimmed every 4 hr. ; between skimming and dropping of last charge there is left an interval of 40 min.; the skimming door is removed, the sand-dam cut open, and the slag allowed to run out into a settling- trough 7 by 24 ft. and 18 in. deep, when the overflowing waste slag is granulated. The flow of slag is regu- lated by a rabble (blade 5X9 or 6X16 in., handle 12 ft.); from 40 to 60 tons of slag are thus drawn in from 15 to 20 min. When the furnace is filled with matte, i.e., with about 200 tons, part is tapped as needed by the converting department. The grating is done every 4 hr. and takes from 35 to 45 min.; the clinkers on the grate are broken up with heavy bars and removed; those on the bridge, likely to be very hard, are attacked at the same time. A good deal of fine coal falls through the grate spaces (grate, 2 in. square, space 6 in.), especially in grat- ing. Clinkers and ashes are removed by a stream of water, and pass over a Fig. 318. — Temperature-curve of Anaconda reverberatory matting furnace, fed with small charges. grizzlie, the oversize going to waste and the undersize to jigs which recover about 10 per cent, coal and coke to be incorporated in the flue-dust briquetting mixture (§122) of the blast-furnace. Claying, or repairing the fettling, at Anaconda is confined mainly to the prox- imity of the fire-bridge. It has to be done once a month, the necessity for it being indicated by the conker-plate becoming red hot. When claying is contem- plated, the furnace is gotten ready by allowing the matte to rise to the level of the skimming plate, skimming the slag clean, tapping the 200 tons of matte into twenty ladles holding from 7 to 10 tons each; this takes several hours. Two or three doors are opened on each side near the fire-bridge, and about 20 tons of self-baking sand (SiOa 95 per cent.) thrown in during the operation. At Cananeai the device of Gmahlin- Shelby, Fig. 319, is in successful operation in the looXig-ft. oil-fired furnace (Figs. 279-287). In the roof , extending the length of the furnace, are openings, 5 by 5 in. and 18 in. apart, which are covered by fire-brick made air-tight with fine ore. Above the furnace is a traveling hopper for distributing the fettling material, fine siliceous ore which has been ' Ricketts, Tr. Inst. Min. Met., 1909-10, xK, 160; Min. World, 1909, xxxi, 1116: Eng. Min. /., I9IO, LXXXIX, 317, SMELTING OF COPPER 28s moistened. Every day some of the fettling is dropped through the holes in places where it is needed; it trickles down over the side walls, which are built with a slight batter, and forms a bank to be tamped down gently. Thus every day from 10 to 15 tons of siliceous ore are fed and smelted; the level of the slag line is kept in repair; and the furnace work need not be stopped for claying. At the Garfield plant and other works this method of claying has also proved very satisfactory. In fact, the temporary shutting down of a furnace causes the Feed Tracks Those Hoppers are at one cad of the Building [Hopper 5'6'i 6[ I tor I Fettling Matoridl ^Hopper t^ distribute Fettling Material alongside of Furnace Holes 5x5 Fig. 319. — Gmahlin-Shelby fettling apparatus at Cananea. brick in the roof to become loose and not form tight joints again when smelting is started; hence air leaks into the furnace. It has been proved that with this continuous claying the smelting power of the furnace is increased and the fuel consumption decreased. Automatic records of the different operations (charging, skimming, tapping, etc.) as carried out one after the other may be kept by recorders of the Bristol type.' ' Jager, Eng. Min. J., 1909, lxxxvii, 1240. 286 METALLURGY OF COPPER 150. Accessory Apparatus, Products, Losses, Cost. — The accessory appara- tus such as ladles for slag and matte are the same as with the blast-furnace (§95.96). The products, matte, slag, and flue-dust are similar (§ 118 and foil.). At Argo, Colo.,1 the ore-slag from the reverberatory furnace, with hearth 42 by 19 ft., was skimmed or tapped into a launder delivering into a smaller reverberatory furnace, with hearth 20 by 14 ft., and a small amount of pyrite added to clean the ore-slag. The ore-furnace treated a 12-ton charge in 3.5 hr., and the ore-slag remained the same length of time in the cleaning furnace. The losses of Cu in reverberatory slags are usually greater than those in the slags of blast-furnaces; they range from 0.3 to 0.5 per cent. Cu. By using the blast-furnace settler with the reverberatory furnace, as is the case with some furnaces at Great Falls, Mont.,^ the slag-loss in the reverberatory has been reduced to the figure of that of the blast-furnace. At Anaconda with the Math- ewson reverberatory furnace the slag-loss' is only 0.30-0.35 per cent. Cu owing to the long path that the slag has to travel before it is tapped. The cost of smelting has been greatly reduced in recent years. The official statements of 1903* made the cost at Butte, Mont., $3.40-3.60 per ton of ore; with the saving of some of the waste heat this figure is reduced to about $2.50; with the new Anaconda furnace the cost is about $1.50 per ton of ore. Cananea* smelts for $1.40-1.77 according to the character of the charge;* Utah Con- solidated' for $1 .40. An analysis of the cost of smelting at the Bingham works, now closed, is made by Barbour.* Moore' gives the costs of total treatment of i ton of ore from ore to blister copper as shown in Table 60. Table 60.- -CosT OP Smelting in Reverberatory Furnace and or Converting Resulting Matte Smeltery Fuel Power Labor Sup- plies and re- pairs Sala- ries Lab- ora- tory Gen- eral ex- penses Sam- pling Roast- ing Con- vert- ing Flux and fet- tling Total $0.91 0.91 0.02 0.02 0.23 0.37 0.36 0.13 0.08 0.04 0.17 0. 22 0.68 0.50 0.08 0.03 $2.53 2.6s Garfield. . . . 0,02 0. 19 0, 22 151. General Arrangement of MacDougall and Reverberatory Plant, Cop- per Queen Consolidated Mining Co. — A plan of the new roasting and rever- " H. N. Pearce, Tr. A. I. M. E., 1906, xxxvi, 891. " Eng. Min. J., 1906, ucxxi, 92. ^ Eng. Min. J., 1908, Lxxxv, 1237. ' Eng. Min. J., 1903, Lxxv, 708. ' Ricketts, loc. cit.; Eng. Min. J., 1911, xcii, 694. ' Eng. Min. J., 1911, xci, 1252. ' Austin, Min. Sc. Press, 1911, cii, 178. 'Eng. Min. J. 1911. xci, i2'S2. '0/>. cit. 1910, Lxxxrx, 1065. ' SMELTING OF COPPER 287 beratory smelting divisions of the Copper Queen Consolidated Mining Co., at Douglas, Ariz., is shown in Fig. 320. The plant went into operation in 191 2. The plan shows six MacDougall roasting furnaces, 18 ft. in diameter (see page no) arranged in two rows. The gases pass with a velocity of 3 ft. per second into the elevated dust-chamber, built of hollow brick on a steel frame; these chambers are provided with Roesing wires (§122) where they drop their dust, to be returned to the roasting-furnace charges, before they enter the central stack. This also furnishes the draft for the two oil-fired reverberatory smelting furnaces in operation. The roaster charge is so made up that the calcine can go direct without further additions to the reverberatory furnaces, where it is delivered on the ele- vated calcine tracks. The reverberatory-furnace gases, after passing through boilers and economizers, are collected in a dust-chamber and then enter the cen- tral stack. Beneath the throats of the furnaces is a tunnel in which travel the "^A of 8m«tter P[ r^2fl-0 Tunnel xBoilere nh^Economizers ^ g C.L .ofDuBt •^ ChBmber ' Eco Domi z ere j. ^ ^ V^ ^ O.L. of Router BPf ._j^ 0.l» olKoMterai i«:i_,_|-40p_0 lunber T — r^ /^^ iS O.L. Chunber 000 -Ui'O^ Fig. 320. — Plan of new roasting and reverberatory-smelting departments at the Copper Queen Works. waste-slag pots; at the firing-end is the matte-ladle pit, the ladles of which are handled by the same cranes that manipxilate the ladles of the ten blast-furnaces (Table 28). 152. Comparison of Blast- and Reverberatory-Fumace for Matting.— The older discussions upon the relative merits of these two apparatus for smelting sulphide copper ores^ have had to be revised to some extent to meet new condi- tions.'' Many points have to be cjjnsidered to arrive for a given locality at a 'Vivian, Eng. Min. J., 1881, xxxi, 248. Peters, Min. Resources U. S., U. S. Geol. Surv., 1882, p. 270. Howe, Bull. 26, U. S. Geol. Surv., 1885, p. 99. Lang, Eng. Min. J., 1890, l, s7o. 'Min. Sc. Press, 1906, xcn, 136 (Editor), 138 (Neill), 197 (Mathewson), 215 (Bretherton) 379 (Hixon), 296 (Austin), 342 (Subscriber); 1907, xciv, 114 (Charles), 815 (Neill). 288 METALLURGY OF COPPER conclusion which is approximately correct; actual practice may be necessary to give a definitive answer. Some of the points are the following: The blast-fur- nace requires coarse ore; can make slags of high or low silicate degree, but requires generally much flux to obtain a slag of desired fluidity, and the large quantity of slag formed carries away much Cu although its assay- value is low; it eliminates impurities effectively, especially in pyritic work; uses a small amount of expensive coke; requires little labor but much power and cooling water; has small repairs; makes much dust; is cheap to build for a given ton- nage; requires little space; locks up comparatively little capital. The rever- beratory furnace, on the other hand, requires an extendfed roasting plant; is especially suited for fine ore (including the flue-dust of the blast-furnace) ; must make slags high in Si02 and is therefore restricted to siliceous charges using little flux, and the higher slag-assays for Cu may be balanced by the small amount of slag made; it consumes enormous amounts of cheap coal (replaced sometimes by oil) ; the labor cost is high, the power consumption is low, especially if there are waste-heat boilers; the original cost, the floor space, and repair will always remain large, as will the capital locked up in the bath of matte and the sand- bottom. The metal locked up in the matte of a reverberatory furnace and a blast-furnace is about as lo : i. With most large plants both reverberatory and blast-furnaces are in success- ful operation, the former treating fine ore and flue-dust, and furnishing through its waste-heat boilers part of the power of the plant, the latter treating coarse ore. 153. Production in the Reverberatory Furnace of Metallic Copper from Matte.; — Bringing the copper forward in the reverberatory furnace, from the first matte to the metallic state was the method employed in most American, English, and many European works; it has been superseded by converting, wher- ever this is permissible, but still holds its own in many places, at least for the production of metallic copper from matte containing Cu 70-80 per cent. There were and probably still are in operation various ways of obtaining metallic copper in the reverberatory furnace from matte containing the old stand- ard of Cu 33 per cent, (so-called coarse metal), characterized by a dark brown surface, a fracture bluish-black when hot and yellowish-brown to bronze when cold, granular to vesicular, and uneven. Two ways may serve as examples, the Ordinary Process with six steps, and the Extra or Selecting Process with seven steps; to these will be added the Argo Process, an adaptation of the selecting process to special ends. 154. The Ordinary Process.— The aim of this mode of operating is to bring forward the matte as quickly as possible. The result is that only a compara- tively small proportion of impurity is eliminated and copper of ordinary grade produced. Hence it is suited mainly for ores which contain little As, Sb, Zn, Sn, Ni, Co, etc. The following tree gives an outline of the process: SMELTING OF COFFER 289 The Ordinary Process Coarse Metal (Matte I, Cu 33 per cent.) Roast I Smelt White Metal (Matte n, Cu 75 per cent.) Roast-Smelt Metal Slag (SiOa 30 ± per cent., 'Cu > i.o per cent.) Return to ore-furnace Blister Copper (Cu 98+ per cent.) I Refine ^ Roaster Slag (Si02 33+ per cent., Cu 2o± per cent.) Add to ore-charge or smelt alone for in- ferior copper. Refined Copper (Cu gg-t- per cent.) Refinery Slag (SiOj 40 ± per cent, Cu 45 ± per cent.) Add to white metal charge or smelt alone. (i) Roasting Coarse Metal. — The aim is to oxidize Fe that it may be slagged, S that it may be volatilized, and perhaps As and Sb that they may be eliminated. The roasting is carried far enough to oxidize all the S that is not needed in the subsequent smelting to combine with the Cu and some Fe to form white metal. If oxide copper ore is available for the smelting, the roasting can be cut short, as CuO will act upon undecx>mposed S-ide and drive off some S as SO2. Coarse- and fine-ore roasting kilns (§ 58 et seq.) are used occasionally when sulphurous gases are to be converted into H2SO4; the reverberatory furnace ( § 70 et seq.) is the common apparatus. Gibb^ found that in roasting matte with Cu 33.40, As 0.185, Sb 0.060, Bi 0.017 per cent, there was eliminated As 35.4 per cent., Sb none, Bi 17.6 per cent. The expulsion of the impurities in matte is less marked than in ore on account of the smaller amount of S present. (2) Smelting Roasted Coarse Metal for White Metal. — The aim of the operation is to carry all the Cu into the matte and most of the Fe into the slag. Pure CuaS (Cu 79.8 per cent.) cannot be made without forcing an excessive amount of Cu into the slag, hence white metal is allowed to retain from 4 to 8 per cent. Fe. The Si02 necessary for the slag comes in part from the sand adher- ing to the coarse metal (if tapped into sand molds), in part from the fettling of the furnace or from siliceous oxide ore; occasionally roaster- and refinery-slags are added to the charge, a practice to be avoided if feasible, as impurities are carried back into the process. The ' 'metal furnace' ' has the same general form as the charge- fed ore-furnace. Table 61 gives a few old examples. As, in the United States at least, the first matte produced in the reverberatory furnace contains Cu from 40 to 50 per cent., and is brought forward by other means, 'Tr. A. I. M. E., 1903, xxxra, 658. 19 290 METALLURGY OF COPPER Table 61. — Metal Furnaces Locality Hearth Grate Charge, tons Coal, lb. per charge Charge, time of smelt- ing, hr. Reference Mansfeld Oker is'ii'Xg' 10" 14' i" Xg' 10" i6'Xi2' 4'X4' 3'ii"X3'3" 4'X4' 6.5 3 2 325° 330° 6 8 12 Leuschner, Zl. Berg. Hmten.Sd- inen W. i. Pr., 1869, xvn, 135. Brauning, op. oil. 1877, XXV, 133. Levy, Rev. Univ. Min., 1884, XVI, 286. Cwm Avon this process in its old form has little practical interest at present. In order to be economical, it would have to be vsrorked in larger units. The chemical reactions taking place in smelting are similar to those in ore-smelting, excepting that the sulphurizing action by FeS must be more prominent than the reducing effect of CU2S upon CU2O or CuO, as there is very little ebullition, consequently little SO2. Thus Le Playi found that upon heating a charge of 3520 lb., first matte low in Cu trickled through the charge and collected on the hearth; in from 3 to 4 hr. three distinct layers could be distinguished; fluid matte on the bottom, next well- fused slag, and lastly partly fused charge. The last resolved itself into well- fused slag and matte. During the fusion the matte grew richer in Cu and the slag poorer as seen in Table 62. Table 62. Changes of Copper-content in Matte and Slag Time after charging Matte, Cu, per cent. Slag, Cu, per cent. 3 hr. 10 min 54 52 65 74 4 hr. min 9 5 hr. min 9 5 35 5 hr. 50 min In charging the furnace, the coarse parts are first introduced and then cov- ered with the fine, as these are more difficult of fusion and adhere more readily to the bottom than the coarse. About i hr. before the fusion is completed, the bottom is scraped to bring adhering parts to the surface. The slag is skimmed and the matte tapped, or both are tapped together into a series of overflow slag- pots. White metal has a dark bluish to grayish color, is compact, brittle, more or less crystalline and free from moss copper. An analysis of high-grade white metal gave Le Play^ Cu 77.4, Fe 0.7, Ni.Co.Mn trace, Sn.As o.i, S 21.0, Insol. 0.3 per cent.; one of lower grade, Cu 64.8, Fe 0.9, Ni.Co.Mn 0.5, Sn.As 0.7, S 22.6, Insol. 1.8 per cent.; an average of fourteen samples, Cu 73.2 per cent. The slag contained, SiOz 31,0, FeO 56.0, AI2O3 6.9, CuO 3.5, MgO 0.6, ' Op. cit., p. 423. 'Op.'cit., p. 419. SMELTING OF COPPER 291 Other Oxides 0.3, Entangled Matte 1.67 per cent. In smelting roasted coarse metal with Cu 31.04, As o.iii, Sb 0.062, Bi 0.013 per cent., Gibb' found the elimination to be As 58.5, Sb 59.0 and Bi 35.7 per cent. 3. Roast-smelting of White Metal for Blister Copper.— In this proc- ess, called in English technical literature "Roasting," the two operations of roasting and smelting are carried on in the same furnace, the smelting following immediately after roasting. The aim is to oxidize all the Fe and part of the Cu and S by melting cakes of white metal slowly in a strongly oxidizing atmosphere so that when the temperature is raised to liquefy the mushy bath of sulphide and oxide, the two will react upon one another and form Cu and SO2, while the FeOj, will combine with Si02 and form a slag. The blister copper retains some impurities to be removed by subsequent fire-refining, ^he slag formed is very rich in Cu, and is retreated. The "blister-furnace" resembles in its general form the metal furnace; special attention, however, is given for air to have free access to the charge dur- ing the first stage of the process. The fire-bridge is made hollow and has air- ports extending into the hearth, so that air enters through the bridge, and through the side doors; sometimes openings are provided near the ends of the fire-bridge for the admission of air; air may be blown through pipes inserted into these side openings. The furnace has to be strongly built, as the hearth holds a heavy charge. Table 63 gives a few examples of older furnaces.^ Table 63. — Blister-furnaces Locality Hearth * Fire-box Charge, tons Time, hr. Coal, lb. per ton charge Reference Wales 4.08 S 6.3 25 24 24 24 1 190 1600 Le Play, loc. cit. Levy, loc. cit. Howe, loc. cit. Cwm Avon Caldera U. S i6'Xi2' i6'6"Xii'6" 4'X4' S'8"XS'3" The- process consists of five operations: (a) Cakes of white metal are charged with paddles through the side doors, piled up almost to the roof in such a manner as to leave open channels between the cakes, and passages for air and gases along the sides and between matte and roof; a space of about 4 ft. is left uncovered between fire-bridge and charge. (6) The doors are closed, and the fire is urged for about J hr. until the cakes just begin to melt; then the doors are more or less opened, and the temperature of the furnace is held at this or a slightly more elevated temperature in order to fuse the matte slowly. The parts of matte fusing first will trickle down over the still unfused parts, and both will be oxidized in the strongly oxidizing atmos- phere. In about half the time given to the whole treatment, the charge will have been converted into a half-fused mushy mass that has been hardened somewhat ^Tr. A. I. M. E., 1903, xxxiii, 660. ' Drawings of gas-fired blister-furnace with regenerative chambers are given in Petera "Modern Copper Smelting," 189s, p. S09- 292 METALLURGY OF COPPER near the doors by the inrushing air; the reaction between CuS and CU2O will have been started causing Cu to be set free, the latter collecting on the bottom, the SO2 passing ofE with a hissing noise. (c) The air-ports and side doors are closed and the fire is urged to liquefy the charge completely; the liberations of Cu and SO2 continue, oxidized Fe rises to the surface, enters the slag, and is skimmed. There remains in the furnace a bath of Cu covered with a thin layer of matte. (d) Air-ports and fire-door are opened to cool the metal and oxidize the matte. With much matte, rabbling and skimming become necessary to oxidize the sul- phide and remove the oxide formed. Cooling is continued until the Cu begins to harden, in order to finish the action of CujS upon CujO, and to assist the expul- sion of SO2 held in solution by the fluid copper. The SO2 in passing off through the half-fluid Cu forms crater-like excrescences and causes the metal to swell to from two to three times its original volume; the spongy mass becomes more or less oxidized by being exposed to air at this temperature.' (e) The temperature is now raised again with th* air-ports in the fire-bridge, kept open, and the Cu is completely liquefied; the slag floating on top is skimmed the blister copper tapped and collected in sand or cast-iron molds. The runners connecting the cakes of Cu in the molds are broken while hot. Blister copper receives its name from the cavities in the body and the excres- cences on the surface caused by dissolved SO2 being given off while the metal is solidifying. Its surface is dark from oxidation, the fracture brick to dark red and granular to fibrous; the blisters vary in color from brown to yellow. Table 64 gives the chemical composition of blister copper. Table 64.— -Blister Copper Locality Cu Fe Ni, Co, Mn Zn Sn, Sb Sn, As S Reference Kaafjord Swansea 99.2-99.4 98.4 97. S 98.0 98-5 . i-o . 2 u.7 0.7 0.8 . 2-0 . 3 0-3 0-0.02 0.4 . i-o . 1 2 0. 2 0. 2 0-3 0. 1 Kerl, "Metall- niittenkunde," p. 215. Le Play, op. cit., p. 486. Swansea 1 .0 0.7 Swansea i Percy, op. cil., Swansea 1 p. 362. 6S- The composition of roaster-slag (average Cu 20 per cent.) is given in Table Table 65. — Roaster-slag Locality SiOj FeO AI2O3 CaO MgO CazO Cu Ni, Co, Mn SnO ZnO S Reference Wales Kaafjord. . Wales 47-5 36.0 45-0 28.0 7.0 28.0 3-0 6.0 tr. 2.7 tr. 0.8 16.9 43-2 25.0 2,0 0.9 4-9 0-3 0.6 2.0 3-2 2 Le Play, p. 487. Kerl, p. 215. Percy, p. 363- ' Vivian, Eng. Min. J., 1881, xxxi, 250. SMELTING OF COPPER 293 The elimination of impurities by volatilization as shown by Gibb' is As 18, Sb 20.5, Bi 73.4 per cent. His comparison of roast-smelting and converting.^ is given in § 168. The research of Keller' has shown that for every one part of Cu there may be slagged 1.95 Pb, 0.80 Bi, 2.02 Sb, 0.38 As, and 0.98 Se-Te. The losses in precious metal are said to be very low. This is given in Europe as one of the reasons for the preference given to roast-smelting over converting for bringing forward matte with Cu 80 per cent, to the metallic state, as in the latter process (§ 170), the blister-forming stage is known to be accompanied by considerable losses. 4. The Direct (Reactor, Nichols- James) Process.^ — This is a modifica- tion of roast-smelting in which coarsely crushed raw and roasted white metal are mixed, the correct proportions (usually 1:2) being ascertained by crucible experiment, and melted down in the reverberatory furnace where the same products will be formed as in roast-smelting. This process is in operation at Mansfeld, Germany, where according to Stahl,^ in treating 111,757 tons of matte and oxides of copper, with Cu 74.46 per cent, and Ag 65 oz. per ton, the loss by volatilization in Cu was only 1.303 per cent, and in Ag 2.5091 per cent. In smelting sulphide copper ore by the ordinary process as outlined, includ- ing the refining of blister copper, the total elimination of impurities was found by Keller" to be Pb 99, Se-Te, 60, Bi 54, Sb 50 and As 2 1 per cent. 5. Refining oe Blister Copper. — This will be taken up in § 186. 155. The Extra or Selecting Process.' — The aim of the process is to produce from sulphide ore the highest grade — "Best Selected" or "B. S." — copper. This is accomplished by two means: (i) By bringing forward the copper more slowly than in the ordinary process, whereby impurities, such as As and Sb, remaining longer in the state of sulphide, are more easily eliminated, and, (2) By forming "bottoms," i.e., roast-smelting in such a manner as to pro- duce some metallic Cu which carries down with it most of the remaining impuri- ' Tr. A. I. M. E., 1903, xxni, 662. ^Tr. A. I. M. E., 1904, xxxrv, 957. ' Tr. A.I. M. E., 1898, xxvm, 127, 816; Min. Ind., 1898, vii, 245, 1900, ix, 240. * Vautin, Tr. Inst. Min. Met., 1893-94, 11, 76. Peters, Min. Ind., 1893, II, 269, "Modern Copper Smelting," 1895, p. 519. Terrill, Eng. Min. J., 1898, ixvi, 665. Keller, Min. Ind., 1898, vii, 244, 245; Tr. A. I. M. E., 1898, xxviii, 821. Kroupa, Berg. Hiittenm. Z., 1899, LVin, 483; Oest. Zt. Berg. Hiittenw., 1899, XLVii, 241, 39S- Thofem-Seine, Eng. Min. J., 1902, lxxiv, 340. Glenn, op. cit., 1902, Lxxrv, 381. Wagner-Primrose, op. cit., 1907, lxxxiv, 671. Stahl, Metallurgie, 1908, v, 353. Styri, op. cit., 19I2, DC, 426, 449. ' Metallurgie, 1908, v, 353. '' Tr. A. I. M. E., 1898, xxvni, 145 ' Gibb, 3d Report, Alloys Research Committee, 1895, pp. 254-286. Peters, Eng. Min. J., 1895, XLix, 512. Keller, Min. Ind., 1900, rx, 512. 294 METALLURGY OF COPPER ties and thereby purifies the remaining rich matte (Cu, 80 per cent.) called "regule." Bottoms are refined in the same way as is blister copper, giving an inferior brand of copper, or are worked up in the wet way. The matte is treated as is the white metal in the ordinary process and furnishes a very pure ^copper. If impure ores are treated by this process, there will result "G. O. B." (good ordinary brand) or " G. M. B." (good merchant brand) copper which is the "Standard" brand' of the London Metal Exchange. The accompanying tree gives an outline of the operations. The Best Selecting Process Coarse Metal (Matte i, Cu 33 per cent.) I Roast Smelt Blue Metal (Matte 11, Cu s 5 per cent.) Metal-slag I I Roast— Smelt To ore-charge Bottoms (Metal, Cu go* per cent.), Regule (Matte in, Cu 80 per cent.). Roaster— slag I 1 I Work up dry or wet Roast-Smelt To Blue-metal charge I ^ I Blister Copper (Cu 99 ± per cent.) Roaster-slag Refine Smelt for second-class copper, or return to Blue-metal charge, or to Ore-charge. Best Selected Copper (Cu 99 -|- per cent.) Refining Slag Return to Blue-metal, or to Ore-charge. In roasting coarse metal for the subsequent production of blue metal, either the elimination of S is not carried so far as would be necessary in smelting for white metal, or, as is more common, no change is made in the roasting, but unroasted rich sulphide ore is added to the smelting charge. Roast-smelting of coarse for blue metal has become obsolete. The smelting furnace for producing blister copper has the form and size of the white-metal furnace. A furnace used in the ordinary process may not be employed for the selecting process, as the work- ing bottom may take up some impurity and become "poisoned"; for the same reason a furnace for extra process may not be used occasionally for an operation in the ordinary process. The management of the furnace is the same as in smelt- ing for white metal. ' Eng. Min. J., 1909, Lxxxvn, 375. SMELTING OF COPPER 295 "Blue Metal" has a purplish-blue color; the fracture when cold is purplish- red, uneven, and has a submetallic luster; the matte is brittle and contains moss copper. An analysis by Le Play^ gives Cu 56.7, Fe 16.3, Ni 1.6, Mn trace, Sn 1.2, As trace, S 23.0, Insol. 0.5 per cent. The blue-metal slag has no particular characteristics. Le Play's analysis^ gives Si02 36.0, CujO 0.7, FeO S44, AI2O3 0.8, MgO 0.2, CaO 1.2, Other Oxides 2.5, Matte 4.2 ( = Cu 2.4, Fe 0.8, S i.o) per cent. Matte lying between coarse metal (Cu 33 per cent.) and blue metal (Cu 55 per cent.) has been called "red metal." In roast-smelting for bottom and regule the aim during the roasting period is to drive oil enough S so that in the fusion all the Fe may enter the slag and enough metallic Cu be produced to carry down with it all the remaining impuri- ties in order that resulting 80-per cent, matte (regule, spongy regulus) shall be as pure as it can be made. The investigations of Gibb' have shown that the amount of copper separated as bottom should not exceed 20 per cent, of the Cu-content of the charge, the usual amount being 14 per cent., as a larger per- centage not only diminishes the direct yield in Cu, but may also impair its purity. Table 66 gives some of the results of Gibb's work. Table 66. — Elimination of Impurities from Matte by Bottoms Impurity Per cent, of Cu-content of matte sepa- rating as bottom Per cent, of impurity of matte entering the bottom Impurity Per cent, of Cu-content of matte sepa- rating as bottom Per cent, of im- purity of matte entering the bottom Sn, Sb. Sb. Sb. Sb. Bi. Bi.. Bi., As. As. 20.6 8.2 47 -S 54-5 8.2 16.0 47-S 8.2 16.0 93-4 21 .0 80.8 93-7 92.6'' II. I 43-1° 47.6 2I-S 30.6 As Ni Ni-t-As. . . . Ni-i-As. . . . Ni-f As, Ag. 25.2 8.0 193 28.3 Au Au Ag Ag, Pb, Sb, Au. 8.2 14.4 19.0 3S(+Fe) 60. 2 8.2« 47.9 Ni-t-77.7 As ? -f 83.9 As Ag diminished by Ni-t-As. 4I.S 100. o 42.9' 74.5 Ag, 83. s Pb, 97.0 Sb, 92.8 Au The furnace, the mode of operating, and the chemical reactions taking place are similar to those in roast-smelting white metal. Analyses of products (bot- tom, regule and slag), are given in Table 67. 156. The Argo Process.* — The characteristics of this process are that the " Op. cil., p. 439. 2 Qp_ cit., p. 440. '^°'^- "'• ' Decrease. ' Practical maximum. ' Practically no concentration. ' Maximum. 'Egleston, Tr. A. I. M. E., 1875-76, rv, 276 (Black Hawk). Pearce, R., op. cit., 1889-90, xviii, SS (Argo). Peters, Eng. Min. J., 1890, l, 189. Ulke, Min. Ind., 1893, 11, 290. 296 METALLURGY OF COPPER o Pi a o 1 Pi d(id>^^^ , ft d^^. r J- J- -; H § t J- J- -; a CO d •ei, < s. M d V Ag, oz. per ton CO a 1 T U *^ 1/1 ^5 .. M 4J ^ i> d l£ U 0. g ^0 " c S c t/3 a g CO t^ " rO H Bi, per cent. d c S fl 10 M d 00 d I (O CO CO NiCo, per cent. iJ -W -0 W 3 C 3 n' U "• " As, per cent. 00 •+ >t 1 iJ ■^ „• "^ ■" "00 "" -'ii CO cr> M ir> 00 ^O 1 M 00 rO ^ -^ W H. V. Pearce, Tr. A. I. M. E., 1908, xxxix, 722. 2 Eng. Min. J., 1909, Lxxxvu, 787. 'Hofman, "General Metallurgy," 1913, p. 74. « Berg. Huttenm. Z., 1882, xli, 152; Eng. Min. J., 1882, xxxm, 261. ' Douglas, Tr. Inst. Min. and Met., 1899-1900, xiii, 2. Jannetaz, P. "Les Convertisseurs pour Cuivre," Baudry, Paris, 1902. Kroupa, Oest. Zt. Berg. Hiittenw., 1903, Li, 695, 715. Mayr, F. "Das Bessemern von Kupfersteinen," Craz und Gerlach, Freiberg, 1906. Sticht, R. C. "Progress in Rapid Oxidation Processes Applied to Copper Smelting," Australian Assoc. Adv. Sciences, Jan., 1907. SMELTING OF COPPER 299 air in thin streams is forced through Cu-Fe matte held in a refractory vessel at 1150-1200° C; Fe is oxidized to FeO and combines with Si02 forming a slag; S forms SO2 and passes off; and Cu is set free to be cast into suitable forms; the oxidation of Fe and S, and the union of FeO and Si02 furnish the necessary heat. The first attempt at enriching matte by a pneumatic process was that of A. Rath in 1866' who at Ducktown, Tenn., forced air through matter to oxidize Fe and S, and continued the process until 1875. In 1867 Semenikow of the Bogoslowsk mines, Ural Mts., proposed making blister copper in a converter; the working tests were carried out by Jossa and Laletin who published their results in the Russian Mining Magazine of May, 1870.^ They succeeded in bringing forward coarse metal (Cu 31 per cent.) to white metal Cu 72-80 per cent.), but failed to produce blister copper. Converting lay practically dormant imtil Manhes and David in 1880 at Eguilles, France, succeeded in obtaining blister copper.* In 1883-84 their proc- ess was introduced at the works of the Parrot Silver & Copper Co. of Butte, Mont.* The original "mode of procedure, of blowing in two separate stages, melting matte with Cu 35 per cent, and blowing to 80 per cent., followed by casting and remelting the white metal and blowing to blister copper, was retained until 1885* when A. J. Schumacher laid the foundation of the modern practice of starting with matte of 40-50 per cent. Cu and blowing to blister copper in two consecutive stages without any remelting of white metal. The last improvement was the working of direct matte instead of cupola matte, the matte being tapped from the blast-furnace settler or the reverberatory hearth either into a ladle and poured into the converter, or made to flow direct into it (now abandoned). This method was planned in 1890-91 by C. O. Parsons for Great Falls, Mont., and carried out there in 1892 by F. Klepetko." Modern American (acid) practice then represents the combined improve- ments of Schumacher and Parsons upon the original Manhes-David process.' So far the converter had always been lined with siliceous material which fur- nishes the Si02 necessary to slag the FeO. About 1888 Claude Vautin experi- mented with a basic lining at Cobar, Australia, but gave up the attempt. In 1890 Keller' made unsuccessful attempts at the Parrott smeltery in Butte at Hixon, H. W., "Notes onLead and Copper Smelting," McGraw-Hill Book Co., New York, 1908. Peters, E. D., "Principles of Copper Smelting," "Practice of Copper Smelting," Mc- Graw-Hill Book Co., 1907 and 1911. ' U. S. Patent No. 57376, Aug. 21, 1866; Eng. Min. J., 1879, xxvii, 260, 1883, xxxv, 250. ^Berg. HiiUenm. Z., 1871, xxx, 7, 17, 57- ' Griiner, Bull. soc. d' Encouragement, 1882, rx, 439; Ann. Mines, 1883, in, 429; Bull. Soc. Ind. Min., 1885, xrv, 607. * Repath, Min. Sc. Press, 1902, ixxxv, 144. ' Hofman, Tr. A. I. M. E., 1904, xxxiv, 261. ' Hofman, loc. cit. ' See also Hass: "Development of Converter Practice," Min. Sc. Press, 1913, cvii, 653. 'Peters, "Modern Copper Smelting," 1895, p. 510. Mathewson, Tr. A. I. M. E., 1913, xlvi, 469. Keller, Tr. A. I. M. E., 1913, xlvi, 474. 300 METALLURGY OF COPPER converting matte in a vessel lined with magnesite. Others did the same at the old Anaconda and the Boston and Montana works. Later Westinghouse' experimented first at Pittsburg, Pa., and later at Ely, Vt., with a basic lining for pyritic smelting; Baggaley^ worked in 1903 along similar lines and with convert- ing at the Pittsmont smelter, Butte, Mont., but did not succeed entirely with his ideas. The work at the plant of the U. S. Smelting Co. at Midvale, Utah, met with a similar result. The first successful converting in a vessel lined with basic, or rather neutral material, is that of Pearce and Smith in 1909, at Baltimore.' The pneumatic treatment of matte in a converter with a neutral lining, the Si02 necessary being furnished by the addition of warmed acid ore, has so many advantages over the original acid process that it has largely replaced the latter. In a discussion of converting, it is convenient, for the present at least, to keep separate the acid and basic converters. With each there will have to be considered (i) the apparatus, i.e., the converter with its manipulation and lining, the blast, and the arrangement of plant; (2) the chemistry and mode of oper- ating, the elimination of impurities, and the thermal features; and (3) the prod- ucts and their disposal, the loss, and the cost. a. Converting in Vessel with Acid Lining 158. The Converter.^ — Matte converters have this in common, that they are side blown and not bottom blown as is the case with all large steel converters. They are usually classed as Upright and Horizontal. 159. The Upright (Manhes) Converter.* — In his first attempt at convertbg, Manhes used a pear-shaped bottom-blown upright vessel with a capacity of 440 lb. matte. At the beginning of a blow everything went smoothly; toward the end the slag thickened from having been overblown and was ejected in-part; metallic copper solidified, having been cooled by air passing through it, and choked the vertical tuyere openings in the bottom. He therefore placed the tuyeres in the side a short distance above the bottom lining, and thereby furnished a space beneath them in which the metallic copper formed would be out of reach of the blast, could settle, and be poured ofi later with the slag.* . ' Metallurgie, 1904, i, 346. ^ Heywood, Eng. Min. J., 1906, Lxxxi, 574; Min. Sc. Press, 1906, xcn, 281. Baggaley, Bull. 83, A. I. M. E., Nov., 1913, p. 2677. ' U. S. Patents, Nos. 942346 and 942661, Dec. 7, 1909; Nos. 942973 and 943280, Dec. 14, 1909; Vail, Eng. Min. J., 1910, ixxxrx, 563; Editor, op. cit., 1914, xcvii, 720. * Christensen, Min. World, 1910, xxxm, 1036. ^ Wheeler-Krejci, Tr. A.I. M. E., 1913, XLVI. « Experiments at Great Falls, Mont. (Hofman, Tr. A. I. M. E., 1904, xxxiv, 304; Wheeler- Krejci, loc. cit.) with an upright converter (13 ft. high and 9 ft. in diam., an initial charge of SO tons of 50-per-cent. matte) were satisfactory as far as the bringing forward to blister copper was concerned which remained suf&ciently flmd to permit pouring; they were not followed up, as the life of the bottom was too short. The difference in the two capes is due to the small amount of charge treated by Manhes which was chilled, and to the fact that his slag was not skimmed and hence was overblown, causing some FeO to be changed into Fe804, infusible at converter temperature. SMELTING OF COPPER 301 The leading data of the original converter of Eguilles are given in Table 68. The original Parrot converter of 1894I was a copy of that of Eguilles; its general form and the details of construction were changed to meet the new conditions. Dimensions and working results are given in Table 68. The old Anaconda converter^ and the plant are described by Hixon.' The Stallmann converter of 1890, square in plan and curved in such a way as to reduce the blowing-out of molten material, was in operation at Anaconda for four years, but was then replaced by the circular type. It is still doing good ser- vice at Mount Lyell, Tasmania, where it was introduced about 1897.* The two leading upright converters to-day are those at Aguas Calientes, Mexico,^ shown in front elevation in Fig. 321, and at Great Falls, Mont., built in 1892. Figs. 322-324 give the details of the converter of 1904. TrMdi ■afi2V.2ii*4 SECTION A-B Fig. 321. — Upright converter of Aguas Calientes. To the left of Fig. 321 are seen the horizontal rack and the pinion by means of which the converter is rotated in a vertical plane. The details of the con- verter of 191 1 are given in Table 68. The Great Falls converter "of 1904, which is similar to the Aguas Calientes type consists of an upright cylindrical boiler-iron shell. A, with refractory lining, h, supported from a cast-iron trunnion-ring, c, by a pair of trunnions, d, in such a way as to permit swinging in a vertical plane for the reception of matte and 'Peters, "Modern Copper Smelting," 1895, p. 529. ' Stickney, Eng. Min. J., 1893, LV, 370; 392, 417; Min. Ind., 1892, i, 151. ' Notes on Copper and Lead Smelting," 1908, p. 95. 'Peters, "Modern Copper Smelting," 1895, p. 530- Faws, Tr. Inst. Min. and Met., 1895-96, iv, 279. Sticht, Presidental Address, p. 33. 'Hamilton, Bull. 83, A. I. M. E., Nov., 1913, p. 2672. 302 METALLURGY OF COPPER Table 68. — Acid Upright Manhfes, Eguilles, 1880 Parrot Copper and Silver Min. Co, Boston & Montana, Great Falls Aguas Calientes Cylinder, height outside. . . . Cylinder, diameter outside. 7' 6" 4' 8" Bottom, height outside. . Throat, diameter Shell cylinder, thickness. Shell head, thickness. Lining, character Quartz & clay Lining, thickness at bottom Lining, thickness at tuyeres Lining, thickness opposite tuyeres . . . Lining, thickness at hood, tuyere side Lining, thickness at hood, opposite tuy. &re side. Tuyeres, number of Tuyeres, diameter Tuyeres, height above bottom lining. Charge, first, lb Charge, last before repairing, lb , Blow, duration, min Blows, number in 24 hr. . . Blast, pressure, lb. per sq. Grade of matte blown, per cent. Cu. . Charges, number per lining Tons copper per lining. Men, per shift. Number of stands. Number of shells. . sr si" o.s" & 0.6" 6" 25-30 16 12-20 33 7-8 8' 6" 5' 2' 5" Quartz & clay 18" 18" 18" 16" \" 6" 2,S00 9,000 80 14 45 9 8' 9" tront-back 9'9,"width trunnion 8' 9" 7' 8" 3' 4}" i" i" Ore 33" 39'' 25" 27" 34" 12 ' l" 2I" 6,500 11,500 170 7 16 40-SO 5 Converting 2, accessory 25 12' front-back 10,' width trunnion 10' 7' 8" 3' 4J" J" i" Ore 30" 55" 37" 38" IS l" Si" 8,500 18,500 170 7 16 40-50 6 36 Converting 2, accessory 2i 16' i' li" 7' 8" 3' 4i" J" i" Ore 30" 65" 16" 37" 38" 15 l" 5J" 8,500 18, 500 170 7 16 40-50 6 40 Converting 2, accessory 2i 3' i" i" Siliceous ore & clay 23" 31}" 3ii" 5,000 -11,000 33,000 -66,000 90-12S 4 I3i-I3i 38-42 3 135 SMELTING OF COPPER 303 Converters Horizontal Manhte- David, Jeres Lanteira Copper Queen Balaklala Con. Copper Co. Tennessee Shannon Copper Co. 1 Copper Co. Granby Con. Min. Smelt. Power Co. Mammoth Copper Min. Co. British Columbia Copper Co. Anaconda Copper Min. Co. 4' 3" 4' 2" 8' S'8" 8' 4" 10' 5" 7' 10' 6" 7' 6" 10' 6" 7' 3 ' 8' 10' 6" 7' 12' 6" 8' l' 10" Quartz & clay 12 (?) 12 (?) 12 (?) S (?) 5 (?) \ 20-40 16 (?) 7J 20-25 16 (?) A" »" Quartz & clay 4" 4" Si 4' J" plate li" cast Brick, I clay: s-7 quartz; lin^ ing 2 clay: 3 quartz 2' 3' 3" IS"-I8" IS"-l8" 16 in. use 14 3" 3"-4" 5,000 10,000 40-so 7-9 lo-is 2S 2-3 Converting 9, lining 7 3' 7'' i" i" Fire-brick and silice - ous copper ore 4" brick 17" ore S" brick 20i" ore 4" brick 19" ore 22" ore 20" ore 10 li" 6" 3-4 charges of 5,000 lb. to white metal and this to blis- ter copper 400 3J-4 10 33-35 3-4 4-27 4 IS 3' 6" 1" 1—3" cast 2 siliceous ore: i fine concen- trates 9" 24" 18" 32" 36" 14 li" 18,000 -20,000 20,000 -24,000 200-250 6-7 4S 3 IS Convertmg 7, relining s 4 }" i" Siliceous gold-cop- per ore 24" 24" IS" 18" 14" 14 li" 6" 8,000 16,000 9i 40 3 7-5 3 10 3' 10" H" I & li" cast Siliceous ore 2ir 26" 26" 16 1" A\" 9,000 20,000 4S-SS 13 & 18 13 20 & 37 2 &3.7 13.47 & 27 2' 10" I" J" 7 siliceous gold ore (80 % SiOj) : I clay 18" 27" 24" 17" 14 li" 10,000 -14,000 3' 9" i" i" Siliceous ore, 2d class ore, cone. slime 25" 31" 31" 31" 31" 16 i" 6" 12,000 17,000 IS 120 & 13s 6 8-10 40 & 50 3, 40% Cu 4, 50% Cu 5, 60% Cu 12, 40% Cu 16, 50% Cu 20, 60% Cu 6 & 6 13s 9 16 44 5 J with sil- iceous ore, 4i with 2d class ore 18.8 with siliceous ore, 15. 63 with 2d class 342 12-14 36 3°4 METALLURGY OF COFFER the discharge of slag and blister copper on one side; on the opposite is the air- box, e, receiving the blast through pipe,/, and delivering it to the interior through tuyere openings, g, traversing the lining. The shell, A, is made up of four parts: the upper, a, forming the hood or head, carries at the lower end a cast-iron collar which serves for bolting it to the trunnion-ring that encloses the middle part, a', of the shell; the lower part, a", is similarly connected above to the trunnion-ring -2'5«^> Ijinr Ty igr Sgr ^ff!-- |<--2'4Ji-'y ( lb O « O 01 ,iiii/ijiii/iti//iis//i^!- ^^'mmmmwmmi Fig. 349. — Kelley slag-casting machine. The circular casting-machine, in operation at Great Falls, Mont., ^ is shown in plan in Fig. 350. It consists of an annular turntable, A, 70 ft. in diameter, carrying 160 tilting molds, B, represented in section in Fig. 351. The table is driven like a pulley by a wire operated by two motors, C and D. The slag is brought from the converters in a ladle by an overhead electric crane, emptied into the s-ton pouring-ladle E, which is similar to the one discussed in Figs. 353-354, and then cast through a curved spout into white-washed molds. When the filled molds arrive at the dumping pocket F, they are tilted auto- matically and discharge on to a steep grizzlie to break the cakes. The broken slag is hosed and collected in a bucket which is raised on the inclined skipway G and delivers the slag to a bin whence it is drawn into blast-furnace charging cars. ' Eng. Min. J., 1908, Lxxxvi, 610. 2 Wheeler-Krejci, Tr. A. I. M. E., 1913, xlvi, 508. SMELTING OF COPPER 317 The slag-skulls are collected in the center of the table, broken, and shoveled into the skip. The blister copper from the converter is poured direct into molds at small plants. In order to prevent spattering, the Bennetts Pouring Spoon, ' Fig. 352, is frequently used. The inside dimensions of a mold holding 200 lb. of copper are: length at top 24 in., at bottom 20 in., width at top 9I in., at bot- tom si in., depth si in. The car carrying the molds^ is constructed of I- beams or channels, and has roller-bearing wheels and axles. The rails should Fig. 350. — Great Falls converter-slag-casting machine. not be less than so-lb.; Shelby' advocates the use of cast-iron rails of ±-section, 8-in. height, 6-in. base, s-ft. length, laid in concrete. In large plants the blister copper is poured from the converter into a ladle, thus leaving the converter free to be recharged with matte. The copper is molded with a casting machine. The machine* in operation at Great Falls, Mont., since 1903, is shown in ^'S^- 3S3~3S4' This is a modified Walker machine (§ 194) with an electrically ' Allis-Chalmers Co., Milwaukee, Wis. ' Glasser, op. cit.; Min. Mag., 1904, x, 135. ^Eng. Min. J., 1907, lxxxiv, 499. 'Klepinger, Eng. Min. J., 1908, Lxxxv, 903. Wheeler-Krejci, Tr. A. I. M. E., 1913, xlvi, 497. 3i8 METALLURGY OF COPPER driven annular frame, 24 ft. in diameter, carrying 24 anode molds, or a corre- spondingly greater number of cake, bar, or ingot molds. The automatic tilting, the inverting of the mold over the sweep (bosh), and the reversing character- istic of the Walker machines are not shown in detail. The main novelty of the Fig. 351. — ^Great Falls converter-slag-casting machine. apparatus lies in the pouring mechanism. The ladle containing the copper is placed by an overhead electric crane in the pouring apparatus which consists of an hydraulically-operated plunger carrying a frame with V-shaped pockets Fig. 352. — Bennetts pouring spoon. for receiving the trunnions of the ladle, to be held in place by two wedges or keyes. The ladle-frame, pivoted on top of the plunger, is also suspended by links, so that upon raising the plunger and tilting the ladle forward, the ladle- spout remains a uniform distance from the mold. While the ladle passes from SMELTING OF COPPER 319 its original position, when full, to its final position, when empty, the spout moves across the mold from front to back, and thus changes continuously the point at which the hot blister copper strikes the mold, and thereby lengthens the life of the latter. The anodes are raised from the bosh by an endless con- veyor which returns overhead. Over the discharge-end of the bosh are pivoted two trolley-beams, 38 ft. 6 in. long, running in a circular track at their outer ends. They are provided with air-hoists for picking up anodes and delivering 320 METALLURGY OF COPPER them within the space controlled by the beams. This is large enough to store all the copper produced between shipping periods. The machine handles 150 tons of metal in 12 hr. with only three men. Other casting machines are dis- cussed in §194. __^s. ^^^ Fig. 354. — Klepinger copper-casting machine, Great Falls, Mont. 166. Arrangement of Plant.^-In the early plants, cold matte was melted in a cupola, tapped and run direct into the converter. The use of Cupola- Matte has been abandoned by most modern plants. At Mount Lyell matte is still remelted, as the grade obtained by the pure pyritic smelting varies too much in its copper-content to permit the use of matte direct from the settler of the blast-furnace, as is customary in plants carrying on partial pyritic smelting. Typical examples of early cupola-matte plants are the Parrott;* and the Old Anaconda.^ The converting plant of Mount Lyell is described by Fawns' and Sticht.* One of the early cupolas described by Stickney^ is shown in Fig. 355. It has a circular water-jacket shaft; a detached movable crucible lined with converter material to furnish a cavity holding a full converter-charge; a tuyere- region of lining material about 18 in. thick; a launder, 12 in. in diameter and lined, having a fall of 12.5 per cent, or more to insure a free flow of matte. It is suspended from a swinging crane by a chain-tackle. Matte in lump form is charged with 12 per cent, coke and some flux for the coke-ash. The small amount of slag (Cu 4-5 per cent.) formed is drawn from the crucible filled with matte when this is to be tapped. The furnace melts 30 tons of matte in 24 hr. with a moderate heat and requires two men on each shift. In some Direct-Matte plants conditions do occur in which the converter •Peters, "Modern Copper Smelting," 1895, P- S4o. ^ Hixon, op. cit., igo8, p. 95. ' Tr. Inst. Min. Met., 1895-96, iv, 284. ^ Min. Ind., 1907, xvi, 437. ' Op. cit., 1892, 1, 152. SMELTING OF COPPER 321 has to melt temporarily its own matte. Thus at Greenwood, B. C./ the pro- cedure with a barrel converter, 126X84 in., having at its disposal 4500 cu. ft. air at 10 lb. pressure, was as follows: a wood fire was started, then a charge of 1500 lb. coke given and a light blast used; 3 tons of cold matte was charged and full blast turned on. The matte fused quickly, more matte was added, the slag skimmed; again matte was charged; and so on until the converter was filled, when it was blown to blister copper. The time required from charging cold matte to obtaining blister copper was 3^ hr., 15 tons of 40- to 50-per-cent. matte were thus converted with one stand in a shift; 45-per cent, matte worked well; Fig. 3SS- — Matte cupola. SS-per-cent. matte required careful handling and occasional additions of coal to prevent freezing. If a converter cannot be run continuously, say for only 8 hours out of 24, it may be kept hot the remaining 16 hours by filling in the hot ashes from a reverberatory furnace. In Direct-Matte plants the converters are always placed in a straight line. If the size of the works is not too great, the smelting and converting departments will be under a single roof, otherwise they are in separate buildings as, e.g., at Anaconda. The plan and vertical cross-section of the smelting and converting departments of the Great Cobar smelting works^ are shown in Figs. 356-357. The plant is built on level ground. There are three blast-furnaces, 56X240 in., arranged end to end, with settlers 18 ft. in diameter and 4 ft. deep, and served by three Connersville pressure blowers, each of 300 cu. ft. capacity per revolution. The slag overflows into 25-ton ladles mounted on two 4-wheel trucks; the matte 'Jacobs, Eng. Min. J., 1906, lxxxii, 440. *Corresp. Eng. Min. J., igo8, lxxxv, 952. 31 322 METALLURGY OF COPPER '81 Z aSO CoiiTBrtcn tt] ft] aaa Capper Cuting Funutoe VuBt Chamber or CoDTorter OosoE m -72 X 100 Stack Fig. 3s6.— Blast-furnace and converter departments, Great Cobar, Australia. Fig. 3S7.-Blast-furnace and converter departments, Great Cobar, AustraUa. SMELTING OF COPPER 323 is tapped into lo-ton ladles handled by a 40-ton electric traveling crane with 50-ft. span, and poured into the horizontal converters, 84X126 in. There are two stands for blowing, three for drying, and two for relining; back of the last are the edge-roller mixing mills. Air is supplied to the converters by a duplex blowing-engine of 10,000 cu. ft. air per minute at 15 lb. terminal pressure. The blister copper is discharged into ladles and then poured into a 15-ton casting furnace, where it is refined and cast into anodes. 167. Mode of Operating and Chemistry.' — The matte entering a converter usually assays about Cu 45 per cent. This grade is satisfactory for obtauiing a lean slag in the ore-smelting furnace, a high temperature in the converter, and a reasonable length of life for its lining. Taking a matte with Cu 30, Fe 40, S 30 per cent., for i part Cu about 1.7 parts FeO will have to be slagged; a matte with Cu 51, Fe 23, S 26 per cent., will give for i part Cu only 0.6 parts FeO, i.e., the life of the lining will be more than doubled. Going higher than 50 per cent, is not advisable for mechanical reasons^ and for the insuflSciency of heat evolved to carry the process through to blister copper. However, under exceptional conditions, matte with 66 to 6g per cent. Cu has been successfully converted.' Supposing a charge has been blown and the copper poured, the converter will be turned up suflSciently to have the mouth in the right position to receive a new charge of liquid matte. This will be poured in from a ladle or run in through a launder, full blast will be turned on, and the converter righted so that the mouth will deliver the gases into the hood. The blast penetrating the matte starts the process. (i) The Slag-forming Stage. — ^The main chemical reactions taking place are: FeS-F30 = FeO+SO!!; 2FeO-f Si02 = Fe2Si04; CuS + 30 = Cu20+S02; CujO-|-FeS = Cu2S4-FeO. The Fe is oxidized to FeO and simultaneously combines with the SiOa of the lining. As the temperature of the matte-bath is about 1200° C, the ferrous singulo-silicate is produced, the formation temperature of which lies at 1270° C. The matte upon entering the converter is at a lower temperature, averaging 900° C. at Anaconda.* Iron continues to be oxidized, any CuzS that may be oxidized is again sulphurized by the FeS that is present. At the beginning of a blow, white fumes of SO2, ZnO, PbO, etc., pass off; the flame issuing from the converter* is tinged rose, becomes greenish, and then turns pale blue.^ The tuyeres need punching at intervals to insure a free passage of air; a small jumping flame indicates the necessity of it. 'Parsons, Eng. Min. J.,'i8g7, Lxm, 481. Moore, op. cit., 1910, xc, 460. ' Schroder, Met. Chem. Eng., 1910, viil, 590. 'Longbottom, op. cit., 1910, vill, 426. * Austin and Offerhaus, loc. cit. 'Levy, Tr. Inst. Min. Mit. 1910-11, xx, 117; Eng. Min. J., 1910, xc, 1207 (flame colors at Anaconda). ' In converting Cu-Ni matte the flame has an entirely different aspect. In fact, the presence or absence of some metals changes the character and succession of colors. 324 METALLURGY OF COPPER The duration of the slagging stage is about 60 min. for a charge of 7 to 9 tons of 4S-per cent, matte. All the iron having been slagged, the converter is turned down, the blast shut off, the larger part of the slag poured and the rest skimmed into a ladle. The approach of white metal toward the rim of the mouth of the converter is ascertained by cutting the stream of slag with a rabble at the place where the white metal ought to appear first; it looks like sizzling grease under the smooth surface of the slag, and boils on the rabble, while the slag appears rather to adhere to it. The slag contains about 2.4 per cent. Cu; an average of one year's runs at Great Falls gave Lloyd SiOj 30, FeO 55, AI2O3 9, CaO i, Cu 2 per cent. Full blast is now turned on again, the converter righted, and with it is started (2) The Blister-forming Stage. — The chemical reactions are: CU2S-I-3O = Cu20-)-S02; Cu2S+2Cu20 = 6Cu+S02; 2Cu20+Si02 = Cu4Si04 to a mod- erate degree. The white metal produced in the first stage is oxidized, and the CU2O formed reacts upon undecomposed CU2S as in the Welsh reverberatory- furnace process. Incidentally some CU2O is scorified. The flame, pale blue at the start, becomes bluish-white, rose-colored to reddish-brown, and disappears. The indication of a finished . charge is that the sparks thrown from the nose of the converter against the back of the hood cease to be dull, having become brilliant, and no longer adhere to the hood but rebound from it. The appear- ance of the Cu on the rod inserted into and quickly withdrawn from the con- verter clearly indicates the state of the metal. A sample poured on the floor may be underblown and contain matte; or overblown to reddish blister,r to expel all the gas, when it will show pimples; or blown only to gas- finish, when SO2 will be given ofi' freely upon solidifying; or to a partial set, when the surface of the sample will show a brown to blackish surface. The tuyeres need almost continual punching in order to furnish the air a free passage. The time required with an initial charge of 7 to 9 tons of 45-per-cent. matte is from 60 to 90 min. When the charge is finished, the converter is turned down, the blast shut off, and the blister copper poured into ingot- or anode-molds, or into a lined ladle to be transferred to a casting-machine or a reverberatory re- fining furnace. A wooden pole is usually inserted through the nose of the converter to hold back the slag; this slag assays at Anaconda^ Si02 22.7, Cu 21.43, FeO 39.4, S 0.04 per cent. The white metal produced by the slag-forming stage is usually too hot for a satisfactory blow for blister copper; it is therefore cooled by adding "dope," i.e., cold white metal, rich slags, old lining, cleanings from hood, dust chamber, floors, etc. If it is too cool, some coal will be charged; injecting warmed crude oil through the tuyferes^ at a pressure of 20 lb. has been a successful remedy at the works of the Mountain Copper Co. The use of air slightly enriched in ^Offerhaus, loc. cil. "Rountree, Eng. Min. J., 1907, lxxxiv, 639. SMELTING OF COPPER 325 has shown a striking effect in laboratory experiments.^ The flame be- coming reddish instead of being reddish-brown is an indication of the con- verter being too cool; the temperature is too high if the flame becomes yellow, and if at the same time there is encountered an excessive resistance in the punching of the tuyeres. The operation of "Doubling," i.e., working two matte charges consecutively in one vessel before blowing for blister copper, is carried on for three reasons: If the temperature of the matte was too low, the slag-forming stage is unduly prolonged, and the white metal not entirely freed from FeS, so that the charge is in danger of freezing during the blister stage; hence another matte- charge is given which will prevent freezing. If the cavity of the converter has become large, it may be necessary to have two white-metal charges to obtain sufficient submersion of the tuyeres; hence the white metal of one converter is often poured into that of another. If low-grade matte is being converted, the charge may have to be doubled to obtain the necessary white metal. Foaming of the slag during the first stage of blowing is the result of not pouring off slag at the correct time; white metal becomes oxidized, and the CU2O formed reacts upon CU2S; this sometimes takes place with suflScient rapidity to cause explosions. Skimming the slag frequently remedies the difficulty. According to McKenzie,^ foaming before the matte reaches the white-metal stage is due to the surface of the matte sinking below the tuyere level when the air will blow direct into the slag. Foaming during the blister-forming stage, when little slag is made, occurs only rarely; the remedy is to charge some coal in order to reduce CuaO. Overblowing is an occasional accident, which is accompanied by the cooling of the copper. The ordinary treatment of such a case' consists in pouring into the converter small amounts of matte, from 10 to 15 lb., at a time, righting the converter after every addition until the enormous volume of SO2 suddenly set free has had time to escape; continuing with additions until SO2 ceases to be evolved; and finally adding enough matte to make up for the heat lost by the endothermic reaction of CU2S upon CuzO and by radiation. This final addition s about equal to the sum of the several additions necessary to reduce the CusO. In order to increase the life of the lining, feeding siliceous ore has been at- tempted in different ways. Blowing it in through the tuyere has always proved a failure;* the probable reason is that the Si02 being cold does not com- bine with the FeO but floats to the surface and becomes entangled with the slag. Replacing sand by dried concentrates^ did not prove sufficiently successful to adopt the procedure in regular practice. It will be shown later that with a large 'Brandt, Metallurgie, 1905, n, 311, 331. Hesse, op. ut. igo6, m, 287, 375. ' Eng. Min. J., 1910, xc, 750. McKenzie, op. cit., igio, xc, 1147. • Wheeler-Krejci, Tr. A.I.M. E., 1913, XLVi, 503. ' Wheeler-Krejci, loc. cit. 326 METALLURGY OF COPPER body of matte, as in the basic converter, in which the Si02 can be brought to the right temperature, blowing in through the tuyeres gives a different result. A method of introducing Si02 or siliceous ore into the acid converter, first practised at Great Falls, is to charge it after pouring the copper, spread it evenly, pour in some matte, allow this to stand for 2 or 3 min. in order to cement the ore to the lining,, add the rest of the matte-charge, start the blast, and right the converter. The siliceous ore is brought gradually to the conversion temperature and slowly slagged. Thus at Anaconda 1000 lb. of ore was shot into the converter from a scoop (boat), and as many as 5000 lb. at Great Falls, as long as the acid converters were in operation. 168. Elimination of Impurities. — ^This important subject has been studied by Keller,! Douglas,^ Van Liew,' Gibb,^ Mathewson,* and Browne.^ Keller gives the data assembled in Table 70 as the average elimination of impurity in converting. Table 70.- —Elimination of Impurities in Converting Element 1 S |Fe 1 Zn 1 Co 1 Ni 1 Pb 1 Bi 1 Sb 1 As JTel Se Per cent, elimination 1 99 1 99 1 99 1 99 1 37 1 96 1 97 1 71 i 81 40 47 The curve of Mathewson, Fig. 358, gives the speed of elimination of Fe, S, and As in converting in the acid barrel converter, 8 ft.Xi2 ft. 6 in., at Ana- conda. For plotting, the percentage of Cu in the samples was chosen as the basis. The curves show that at the beginning of the blow during the first 10 min., S is more quickly oxidized than Fe, but that later the reverse is the case, until at 90 min. 98 per cent, of the Fe has entered the slag, when only 75 per cent, of the S has been changed into SO2. The oxidation of Fe progresses slowly and evenly; tjiat of S at a uniform rate up to no min., when CU2S is oxidized and CU2O reacts upon undecomposed CU2S causing SO2 to be emitted freely. The expulsion of As is very rapid during the first 10 min., and then progresses at a uniform slow rate. Gibb^ points out that as regards the degree of elimination of impurities, such as As, Sb, and Bi, there is not much difference between converting and roast-smelting, but of course as regards speed there is no com- parison. Losses in precious metals are taken up in § 770. As to the composition of converter gases, Sticht* has found that during the slag-forming stage practically all the O of the blast is utilized. The percent- age of SO2 in the gases rises, during the first 10 min., quickly from i to about 10 per cent, volume, and remains constant during the first stage; the free O present ranges from a trace to perhaps i per cent, volume. During the blister-forming ' Tr. A. I. M. E., 1898, xxviii, 146, 816; 1900, xxx, 310; Min. Ind., 1900, ix, 237. ^ Tr. A. I. M. E., 1899, XXIX, 543. ' Op. cit., 1904, xxxiv, 418, curves replotted by Gibb, p. 960. * Op. cit., 1903, xxxiii, 664. ''Op. cit., 1907, xxxviii, 154. * Op. cit., 1910, XLi, 296. ^ Tr. A. I. M. E., 1903, XXXIII, 664. * Presidential Address, p. 44. SMELTING OF COPPER 327 stage the content of SO2 rises quickly to about 18 per cent, volume, while free appears to the extent of 0.5 per cent, and reaches 2 or 3 per cent, at the end of the blow. Hence the air-efficiency in the converter is high. 100.- 95. 90.- 85. 80. 1Z. 70.- 65. 60.- II 66. \ 50.- i «•" 40. 35.- 30.- 25.- 20.. 16. ■ ^10.- 5. 0. - — , ,-' ^ ^' .V w?i ^ / ^ y s 1 4 ^ X \ I \ s ^ \ ■\ s^ \, '<" to \' s .^ 1 \ ;; \ s. 1 s \ N ._ ... -ir. en,v \ ^~ -- ■\ ..^ \ _ "^•n. ~- -0.45 -0.40 I -0.36 S -0.30 •? y # / / / / / A f ■i/ / V ? .o'V- y i fi f V' >' / / .--; Fe l< {Fe 0)ii.£ iOi / i/ f &■ ^^y 4^y y A V, k A y / A^' V ■ ./ ( m ■v/ •ify rtoO^« mic : teacti( ms // y .\^i A^ —^ ^. Hea tof F jrmati on of AssC B, Sb sOj,. fnO, mtKi •.uiOl iSiO, 1 3 2 3 Ti me in 4 Mimit ea 5 6 % Fig. 359.— Graphical representation of thermal changes in converting. In Figs. 360-361 are given two curves tracing the Calories evolved in blowing two acid charges at Anaconda in 1910. The results were obtained by a procedure similar to the one just given. Both curves show a quick rise of temperature at the beginning of the blow, a gradual fall during the second period, SMELTING OF COPPER 331 and a sudden rise during the third. While the curve rises to the first and highest peak, siliceous ore may be added to the charge to keep down the tem- perature; when it reaches its lowest depression white metal may be added to furnish additional source of heat and prevent the temperature from becoming too low. ^ 040. OSOii '•^ tN 518 Oat |i81,«9 C.L V 1 770, .03 Cn / i [399 IllCa . 3K,DC I 0.1, / \ ] / (» MOO L i s IG.^Cl Y Oil / 033,1 3Cftl 815,8 10.1. / Ti \ / 'P si \ \ ii,,'.L &> --^ .520. .21 C 1. , "(3 07,403 .\ u '«0,4b0C.L \ 105 0,1. \ / \ r"'"y«» 317 1 lu. 917 1. N / 10 20 30 40 60 60 70 SUlolouf Oa Ihould be add«l Time o£ Blew In Mlautea 80 00 White Metal Benip Bhould be added 1 40 50 00 70 80 00 100 110 120 130 Time of Blow in Minutes Figs. 360-361. — Heats evolved in blowing acid converter-charges at Anaconda. 170. Products, Losses, and Cost. — The leading products of converting copper matte are blister copper, slag, flue-dust, and flue accretions; incidentally there have to be considered the slop and splash of the converters, and the shells and sculls of the converters and the ladles. Copper has a copper color; the surface of the ingot varies with the point to which the blow has been carried (p. 324). Table 74. — Analyses of Converter Copper Element Impure Boston & Mon- tana, normal Boston & Mon- tana, over- blown Ana- conda Mont. Ore Pur- chasing Co. Copper Queen El Paso Pb Per cent.8 0.055 0.057 o.is u. 108 U.027 0.112 98.841 Per cent. . 0069 0.0029 0.0546 0.0156 0.0034 Per cent. 0.0016 o.oors U.0I9S 0.0072 0.0033 Per cent. U.0103 0.0040 0.0130 o.02ir 0.0072 Per cent. 0.0517 o.oosi 0.0533 0.0231 0.0078 Per cent. 0.084 Per cent. 0. 103 Per cent. 0.07 Per cent. 0.16 Per cent. 0. 16 Per cent. 0.41 Bi Sb U.009 0.012 U.043 0.244 98.64 U.159 0.04 0.027 0. 196 12 . 245 0.014 or6 045 0.403 U.I3S 0.039 0.059 0.067 trace 0.008 As 0.06 SeTe S i .■; 0. 26 98.8 0. 26 98.6 0.30 Cu 99.9166* 99.9669* 99.9944* 99.8590^ 98.30 Pe. . . . 0.184 Ni 1 Zn.. . 0.004 Insol Ag 0.641 O.OOI r II. I.lB III. 8» 83.7' 0.63' "-43' I08.li 0.37' 6 184.45 Au ,j.84S Rel ...^....,...^....| 2 1 2 3 1 3 1 6 6 6 ' Keller, Tr. A. I. M. E., 1897, xxvii, 108. ' Douglas, op. cil., 1899, xxex, 543. ' Ounces per ton. ' Keller, op. cil., 1898, xxvill, 150, 157. ■* Includes Ag and Au. * Private Notes, 191 2. 332 METALLURGY OF COFFER Table 75. — Rare Metals in Blister Copper' (In 100 Tons Blister Copper) Locality I Au, oz. Ag, oz. Pt, oz. Pd, oz. Se, lb. Te, lb. Bi, lb. Ni, lb. Garfield Steptoe Omaha Mountain Tacoma Aguas Calientes. Cerro ile Pasco. . Mount Lyell 298.0 169.0 360.0 I,4l8.u 2,187.0 482.0 170.0 464.5 3.480 S50 23,090 10,990 8,710 67,300 9,900 7,20s .342 .016 .82s .320 .710 .416 .319 .624 1. 183 4.402 6.486 0.607 3327 0. 226 0.589 1.374 56.0 no. I 26.6 36.0 42.0 170.0 13-7 42.0 5.54 none 67.1 3-3 none none none none 6.1 0.33 18.6 27.3 5.7 4.0 13. S 4.3 40. 64. 944. 32.0 166.0 The converter slag, forming with 45-per-cent. matte from 60 to 70 percent of the weight of the matte, usually approaches in composition the singuld- silicate. As it contains from 2 to 3 per cent. Cu and is more basic than the ore- slag, it is returned to the ore-smelting charge. In some cases it is disposed of by means of a slag-casting machine (§ 165) and goes as lump-ore to the blast-furnace charge; in others it has been found to be more profitable to pour it into the blast- furnace settler; again it has been poured into the reverberatory furnace in order to render more fluid the siliceous charge. The skimming-slag, which usually runs very much higher in Cu (over 10 per cent.) than the converting slag, has to go back to the ore-smelting fur- nace. Analyses of converter slag are given in Table 76. Table 76. — Analyses op Converter Slag Component Mont. Ore Purchasing Co. Anaconda Great Falls Copper Queen Cananea SiOs 29.97 58.40 35.70 55.83 0.22 i.76 30.0 55.2 29.8 57.8 33.46 63.48 32.57 65.35 26 8 FeO 58.3 MnO AUOs 3.5S 2. 12 2.36 I .09 9-5 l.o 10. 6 2 CaO 1.5 ZnO 0.86 2.14 0.51 3.18 0;I9 2.25 Cu 2.00 i-^ 4-47 0.96 2.44 Ni Pb 1.245 O.OOI 0.087 0.021 u.OOO O.2S0 2 Bi Sb r As Si, Te S 1.03 3 Re£ 4 4 S 5 Flue -DUST, of which less than i per cent, of the weight of the matte is formed, may run high or low in Cu depending upon where it settled. Material of Ana- conda' collected near the hood averages Cu 71.23 and S 12.5 per cent.; a lower- 'Eilers, A. Tr. A.I. M. E., 1913, xlvii. 2 Keller, Tr. A.I.M. E., 1898, xxviii, 153. 'Private Notes. ^Hofman, Tr. A. I. M. E., 1904, xxxrv, 307. 'Douglas, op. cii., 1899, xxix, 543. "DeKalb, Min. Sc. Press, 1910, ci, g. 'Offerhaus, Eng. Min. J., 1908, lxxxvi, 751. SMELTING OF COPPER 333 grade material collected farther away at Great Falls^ assayed Cu 63.4, SiOz 3.8, FeO 12.8, S 16.0 per cent., Ag-Au 53.8 oz.; and Cu 37.8, Si02 1.5, FeO 12.8, S 14.3, As 4.8 per cent. The following two more complete analyses of dust from the Montana Ore Purchasing Co., taken near and far from the converter, are by Keller:* Cu 23.32, Fe 3.30, Zn 2.85, Pb 18.81, Bi 0.42, Sb 0.935, As 1.805, SeTe 0.0026, S 1.34, SO4 3446, H2O 10.87 per cent., Ag 18.2 oz; and Cu 2.36, Fe 2.80, Zn 5.53, Pb 32.24, Bi 0.48, Sb 2.186, As 5.694, SeTe 0.0054, S 0.0, SO4 36.15 per cent., Ag 6.4 oz. An analysis of the gas passing through a con- verter flue gave Dunn' SO2 2.845, SO3 0.0515, CO2 0.2084, H2O 1.061, AS2O3 0.00073, 12.04, N 83.64 per cent, volume. The loss in Cu and Ag-Au* has been variously estimated; so far no accurate determinations have been made excepting perhaps those of Mount Lyell, where the loss of Cu is 2.02 per cent., of which 1.17 per cent, is incurred in converting and 0.85 in retreating intermediary products; and that of Ag 1.18 per cent, of which 0.68 per cent, is due to converting. It is generally held that with a matte free from or low in volatile metals such as Pb and Zn, and with suitable dust- chambers, the loss of Cu in converting does not exceed 1.5 per cent.; the loss of Ag is about 2.5 per cent., and the loss in Au is nil. The presence of lead greatly increases the loss as shown by the recent figures of Moore,^ and Semple,* but here also the loss in Ag rises and falls with that of the Cu in the fume. With the present possibility of filtering converter gases, this disputed question ought to be near its decision.^ There is full agreement in the opinion that the loss during the slag-forming stage is negligible, and that it becomes apparent only during the blister-forming stage. The danger of drawing conclusions from laboratory experiments upon this question of loss in converting is evidenced by Giinther* who found, in treating, in a small barrel converter, matte from Mansfeld with Zn 3.7-7.4 per cent., losses which exceeded the entire loss of the complete treat- ment from raw ore to blister copper. The cost of converting 40-per-cent. matte in the Central West is given by Austin' as $4.83 per ton of matte. According to Moore,'" the cost per ton of blister copper varies from $4.53 to $9.76 as seen in Table 78. The Tennessee Copper Co. converts^' in a basic vessel at a cost of 0.709 cents per pound fine copper; the Mammoth Copper Mining Co.'* at a cost of 0.75 cents starting with matte of 15-20 per cent. Cu. 'Hofman, Tr. A. I. M. E., 1904, xxxiv, 30. ^Tr. A. I. M. E., 1898, XXVIII, 154. 'Tr. A. I. M. E., 1913, XLVi. * Sticht, Presidential Address. Summary. '£««. Min. J., 1910, xc, 263; Tr. A. I. M. E., 1913, xlvi. 'Eng. Min. J., 191 1, xci, 508- ' Ricketts, Loss at Cananea, Eng. Min. J.. 1911, XLi, 1246. ' Metallurgie, 1905, 11, S39- 'Min. Sc. Press, 1911, oil, 178. " Eng. Min. J., 1910, xc, 464. " Eng. Min. J., 1912, xcni, 1034. "Op. cit., 1912, xciv, 982. 334 METALLURGY OF COPPER Table 78. — Cost of Converting per Ton of Blister Copper Plant Matte, Cu per cent. Cost, $ Matte, Cu per cent. Cost, $ Matte, Cu per cent. Cost, $ Matte, Cu per cent. Cost, $ A B 38.2 42.8 41.9 43 S 40.2 8.40 8.50 9.00 453 4-47 370 8.SS 41. 1 7-73 41. 1 5-76' C D 42.7 4.92 41.6 5-3° E^ b. Converting in Vessel with Basic Lining 171. Basic Converting in General. — The necessity for frequent renewal of the lining in the ordinary acid converter has been the cause of experiments with a lining which is not attacked chemically by the process. The best lining so far has been one of magnesite brick; and this has given the new apparatus the name of Basic Converter, although the basic character of the lining has nothing to do with the process, because it remains unchanged. The SiOz necessary for slag- ging the FeO is supplied by the siliceous ore charged. The early working tests with a really basic process by Keller' and Westinghouse (p. 300), as well as the laboratory experiments by Schreyer^ and Styri^ have proved the futility of trying to work without Si02. Pyritic smelting in a converter lined with magnesite by Knudsen (p. 196) and the partially successful work of Baggaley' have already been described. The use of an inactive lining in the form of magnesite brick in a barrel-shaped vessel for converting copper matte, with the addition of SiOa to slag the FeO formed, has been carried to a successful end by W. H. Peirce and E. A. C, Smith at the works of the Baltimore Copper Smelting and Rolling Co., and the result is the "Peirce-Smith Basic Converter," which is the outcome of work extending over several years.'' The process has been so eminently success- ful that it is replacing acid converting in most smelteries and will continue to do this unless it meets special conditions, such as occur in localities where a high smelting charge can be obtained for siliceous ore better suited for a lining than for charging into the converter, which make it more profitable to use an acid lining. The main advantages of the basic over the acid converter are:^ the decreased cost of lining (one basic lining for 2500 tons of Cu vs. one acid for 10 tons); the greater air-efficiency (75 vs. 60 per cent.), on account of the use of a metal tuyere- I913, XLVI, 474. ' Basic converter. ^ Lead-copper matte. 3 P. 299, and Tr. A.I.M. E., * Metallurgie, 1909, vi, 190. ' Op. cit., 1912, IX, 426, 449. 8 P. 300, Tr. A.I. M. E., 1913, XLVI, 480. ' Editor, Eng. Min. J., 1911, xci, 944. * Moore, Eng. Min. J., 1910, xxxK, 319; Mathewson, Tr. A. I. M. E:, 1913, xlvi, 473. SMELTING OF COPPER 335 pipe; the use of larger vessels (40 tons Cu vs. 20 tons), accompanied by economies in labor, power (one-half), and repairs; the use of low-grade matte and of siliceous ore with Si02 35+ per cent.; the small amount of intermediary product to be retreated (per ton Cu: 1.6 tons of slag with 1.5 per cent. Cu vs. 2.4 tons with 6 per cent.), and the consequent large direct output of blister copper (95 vs. 70); the formation of basic slag to serve as flux in the blast-furnace; the neatness and cleanliness of plant; and the decrease in danger from accident and dust. The disadvantages are: blowing out of fines; time required for repairing and lining; care in manipulation; continual punching of tuyeres. Shortly after the Peirce-Smith converter had proved to be a success at Garfield (early in 1910), the Anaconda Company lined its horizontal converters with magnesite and proved that success was not confined to the details of the Peirce-Smith vessel; the same was done at Great Falls (first attempt 1897, second 1901, chrome brick 1906, use of magnesite 1911) with the upright con- verter. In fact, the work with this upright 12-ft. vessel lined with magnesite brick has been so advantageous that in new basic installations' such as Copper Queen, Cananea, Calumet & Arizona, Arizona Copper Co., Mason Valley, and others, the upright converter is preferred to the horizontal. The Great Falls converters, after having been in operation for two years, had their original lin- ing, when put out of commission to make way for the 2o-ft._ converters now in operation. Basic converting will be discussed in connection with the three forms of con- verters which have become typical for the present practice: the Peirce-Smith, the Anaconda, and the Great Falls. 172. The Peirce-Smith Converter.^ — The converter manipulated with electric power is shown in Figs. 362-365; the hydraulic manipulation is represented by Fig. 366. The converter is a horizontal cylindrical shell, a, of f-in. steel plate; it usually is 26 ft. long, 10 ft. in diameter, and holds 30 tons of copper; the shell is open at the top for 4 or 5 ft. to allow for expansion of the lining, and is tied by rods, h. Some converters recently installed are 36 ft. long (Steptoe, Copper Cliff), 10 ft. in diameter, and hold from 50 to 60 tons of copper. The arch of the lining rises slowly but steadily while the converter is running, and in time breaks down. The pair rests on and revolves with three cast-steel I-beams, c, riveted to it, on three pairs of carrying rollers, dd'. The base of the front roller, d, of each pair es fixed, while that at the back carries on trunnionp two smaller rollers, d', in order that the weight of the converter, over 175 tons when filled, may be evenly distributed. The flat heads, e, also of f-in. steel, are made adjustable in order that they may yield to the longitudinal thrust of the expanding lining. This is accomplished by a riveted flange which extends telescopically into the cylindrical shell. A head is held in place by I-beams ' Editor, Eng. Min. J ., 1913, xcv,335. 'Vail, Eng. Min. J., 1910, Lxxxrx, 563. Moore, op. cit., 1910, txxxEX, 1317; A. I. M. E. Bttil. 83, Nov., 1913. Neel, Eng. Min. J ., 1911, xci, 707, 964. Hamilton, ^4. /. M. E., Bull. 83, Nov. 1913. Editor, Eng. Min. J ., 1914, xcvii. 720. 336 METALLURGY OF COPPER uo;4D|ncliu»ui 3jj433j3 SMELTING OF COPPER 337- which are spaced over the outer surface and connected with the nearest cast- steel I-beam friction-ring by tie-bolts provided with nuts. A bolt passes through the ring and a piece of flat-iron bearing on the flanges of the cast beam. Loosening the nuts, when the furnace is being heated, furnishes the necessary expansion space. The strain on a bolt is indicated by the bending of the flat-iron. On the blowing-side a short distance beneath the median horizontal plane are 32 cast-iron tuyeres, 2 in. square with i.2S-in. holes, extending through the lining and pitching downward, which are provided with improved Dyblie valves and severally connected by a flexible hose, h, to the wrought-iron blast-pipe, i. Fig. 366.^Hydraulic manipulation of Peirce-Smith converter. This is bolted to the converter shell and connected by a cast-iron pipe, 7, and a swing-joint, k, with the stationary blast-main leading to the blowing-engine. Sometimes tuyeres are omitted opposite the discharge spout in order to di- minish the loss by splashing. Along the tuyere line the shell contains movable plates to allow repairing the i8-in. tuyere-belt from the outside without cooling the converter. ■■ At some plants the cast-iron tuyere has been replaced by a wrought-iron pipe, which works equally well and makes replacement more convenient. In a converter 26 ft. in length the number of tuyeres may not fall below 20, as the heat generated becomes insuflacient; with the throat of the converter placed at the center instead of at the side (see below), the number of tuyeres has been increased to 37, with the result that converting proceeded 27 per cent, more quickly than with 32 tuyeres. ' Detail, see Eng. Min. J., 1913, xcv, 322. This arrangement is of more importance when lead-bearing matte (Cu 40, Pb 15-18) is treated than with regular copper matte. 22 338 METALLURGY OF COPPER On the pouring side, i.e., the breast, is an opening with spout, I, for the re- moval of slag and blister copper. It is closed with magnesite brick and clay. For pouring off slag, an opening, about 3 in. in diameter, is made in the upper or clay part of the breast; for casting the copper, the tap-hole in the lower part is opened. In pouring the copper, the magnesite breast acts as a skimmer by holding back the slag which remains in the vessel. Some vessels have separate openings for matte and copper. In one end of the converter is an opening, not shown, for the insertion of an oil-burner to heat up the furnace, either at the start or in case the copper cools too much on account of delay or some other cause. On the top and at one end of the vessel is the throat, m, which receives matte and siliceous ore, and delivers the gases to the flue leading to the dust-chamber. The sheet- or cast-iron throat is lined with magnesite if the flame tends to burn out the lining; with clay if there is a tendency to form accretions, as is especially the case with Pb-Cu matte. ^ As stated above, the throat has been transferred to the center at some plants. This has been found advantageous .for the even spreading of the siliceous ore; dried concentrates move more easily than does ore. The manipulation is effected by two wire ropes, n n' , of which one end is attached to the converter, the other to a block, 0', sliding in guides. The blocks are firmly connected with a threaded hollow shaft, p, and this is rotated by a threaded gear which in the casing, q, engages a pinion direct-connected with the shaft of the dynamo, r. The hydraulic manipulation is shown in Fig. 366. Here the blocks, in which the wire ropes are anchored, are direct-connected with the piston-rod, s, of the hydraulic cylinder, t. The working lining is made of 9-in. magnesite brick excepting at the tuyere- belt, where i8-in. magnesite slabs extend 8 in. below and 24 in. above the tuyeres. The bottom is protected with fire-brick, d, 13 in. thick along the center, tapering to about 4 in. on the sides and reaching 8 in. below the tuyere level. In a 26-ft. vessel the dimensions of the oval cavity are: length 20 ft., width 7 ft. The magnesite brick are laid in dry magnesite powder except in the tuyere-belt where a mixture of magnesite powder and linseed oil is used. In order to furnish space for expansion, j-in pine liners are placed after every eighth course of g-in. brick. In many works, the tuyere-region is packed 8 in. above and below the tuyeres with a ground mixture of calcined magnesite and asbestos with 5 per cent, silicate of soda, first used at Anaconda by Mathewson. This mixture appears to work most satisfactorily, but requires careful warming in order to prevent spalling. A lining of chrome brick or some other non-corrosive material has been suggested to replace magnesite. The 26-ft. converter in working order weighs from 125 to 135 tons; with the charge, this weight is increased to 165-175 tons. The blast-pressure ordinarily does not exceed 11 lb. per square inch, and the volume ranges from 90,000 to 100,000 cu. ft. of air at sea-level per ton of blister copper starting with 40-per cent, matte. The air-efficiency is about 75 per cent, and reaches 90 per-cent. 'Howard, L. O. "Collar-puller for Converters," Min. Sc. Press, 1913, cvi, 733. SMELTING OF COPPER 339 Accessories are similar to those of the acid converter and the arrangement of plant shows little difference. In starting, the converter is brought to the required temperature by a wood fire (18 hr.), followed by an oil-burner (8 hr.) introduced at one end, or by heating with coke (6 hr.). When hot, an initial bath of matte, 30 to 40 tons with a 26-ft. vessel, is poured in through the throat and followed by from 6000 to 7000 lb. dried siliceous ore shot in, from a boat or a rectangular box of sheet iron con- tracted at one end to a pouring spout and suspended from a crane. Blast, of about s lb. pressure, is turned on for 2 or 3 min., and the converter turned over until the tuyeres are well covered by the matte. The ore spreads out over the surface of the matte. When evenly distributed, the vessel is turned over an additional amount, and the volume of air increased to from 6000 to 8000 cu. ft. per min.; this gives a pressure of 10 to 12 lb. with the tuyeres immersed 24 in. below the surface of the bath. Blowing "on matte" 30 to 45, and sometimes 60 min., will cause the FeO and the gangue of the siliceous ore to form a thin slag which can be poured off readily. The fluidity of the slag is tested with an iron rod. The slag is poured off, and from 8 to 10 tons of liquid matte poured in from one ladle, or perhaps 13 tons from two ladles. The vessel is brought to its central position, i.e., one in which the tuyeres are not immersed and the slag is not flowing, and a tubful of siliceous ore (5000 to 6000 lb.) added. After this, the vessel is again turned over on the tuyeres, and converting continued for about 30 min. This cycle of operations is repeated with decreasing charges a sufficient number of times (usually for 7 hr.) to obtain the requisite amount of white metal, about 60 tons, to furnish 42 to 48 tons of blister copper; the white- metal bath will reach within 2 to 4 in. of the tuyere-level. It is essential that in the several pourings of slag, only little slag be left floating on the matte, as this retards oxidation. The critical stage of the entire converting lies in slagging off the last iron before the white-metal stage; an excess of Si02 makes the slag sticky and a lack causes Fe304 to be formed. The white metal is now blown in the usual way in about 4 hr. to blister copper. At the end of the blow a small amount of rich slag is formed, which is taken care of by the next charge. Thus blowing on matte for about 7 hr. and on copper for about 4 hr. gives one pour of blister copper. The grade of matte usually blown is the same as in the acid converter, viz., Cu 45 per cent. Matte of lower grade, say Cu 30 per cent., is worked in the same manner; only it takes a longer time to fill the converter with white metal. The Ore-charge. — The siliceous ore charged may vary greatly in com- position and size; the conditions to be fulfilled are that there shall be an excess- SiOa, and that the other components shall not cause the slag to be sticky. A content of SiOa 35 per cent, forms a low figure; however, the Utah Copper Co. concentrates with SiOa 28, Fe 12, AI2O3 1, Cu 26, are used regularly as flux; the amount that can be used is limited, as the material tends to fill the converter. A few partial analyses of siliceous ores charged are given in Table 79. 340 METALLURGY OF COPPER Table 79. — Analyses of Siliceous Ores Charged with Basic Converter Element El Paso El Paso Cananea Garfield Steptoe Cu 6,8 none 78.8 6.6 none 76.3 70.1 6.6 none trace I. 2 4-32 Pb Si02 82.0 3-4 76.0 7.0 66.0 Fe 51 none 1 .1 4.82 Mn CaO 1 .0 0.69 MgO AI2O3 8.0 9-95 Zn none 3-4 none trace s 30 |in. 3.83 Size Ag oz. per ton . . . Au oz. per ton , . . 2.4 0.05 As to size, the ore is usually not coarser than | in.; ore crushed through a 4-in. ring makes sufficient fines to allow the whole to be used without further comminuting. Very fine concentrates are blown into the vessel through one or more tuyeres by means of the Day apparatus, which is shown in Fig. 373. The flux, of course, has to be approximately dry, i.e., it may not retain more than 4 or 5 per cent. H2O. The Mac Dougall furnace (§ 64) heated from a small coal fire or an oil-burner serves this purpose effectively ; a rotary dryer 3 ft. in diameter, 30 ft. long, will handle 40 tons of material in 10 hr., with two men, at a cost of $0.10 per ton. The tuyeres have to be punched consecutively during the entire blow. Two and sometimes three men per shift are necessary for this work. The tuyere- pipe has a removable cap which contains an opening for this purpose. During stops the cap is taken off and the entire face of the tuyere cleaned. Two sets of punching bars are used, each of f-in. steel about 4 ft. long; the ordinary tool has a punching-head i in. in diameter and a hammer-head 2i-in. ; the cleaning tool, used when the converter has been turned down to be emptied, has a punch- ing-head i| in. in diameter, and a 2j-in. hammer-head. The slag, as stated above, has to be thin in order to pour well. The Si02- content may range from 20 to 30 per cent.; with clean copper matte it varies from 25 to 27 per cent. ; with leady matte it is as low as 23 per cent. If the SiOj goes below 20 per cent., Fe304 forms and incrusts the bottom. The additions of flux have to be carefully regulated. At Garfield the maximum capacity of vessel of 65 tons blister copper was reduced in the early stages of converting to 25 tons by Fe304 building along the bottom. It was once held that, if the SiOr content was higher than 30 per cent., it attacked the magnesite linmg. This belief has been proved to be erroneous by Mathewson^ who by the analyses given in Table 80 shows that there is no connection between SiOa- and MgO- contents of a slag. ^Tr. A. I. M. E., 1913, xLvi, 473. SMELTING OF COPPER Table 8o. — Content of MgO and S1O2 in Slags of Basic Converter 341 Cu SiOj FeO CaO MgO Per cent. Per cent. Per cent. Per cent. Per cent. 2.2 37.8 46.3 I . I 0.6 2.2 37-2 46.6 I.O 0.6 380 3S-I 48.3 I.O 0.6 2.80 34-S 49-3 1 .0 0.6 3.20 33-5 SO. 3 I.O o-S . 380 31.6 51-4 1-4 0-5 i.8d 30,8 S2-S 1-3 0.8 2,20 28.9 56.0 I.O o-S 2.40 28.0 S6.S o.g o-S 3.00 25. 8 S6.6 I.O o-S 4.60 24.6 S6.2 1 .0 0.7 4.60 22.4 S8.3 1-3 0.7 4.20 23-4 S8.i I . I 0.4 In blowing, the fluidity of the slag is tested by inserting a bent rod through the slag-pouring opening and quickly withdrawing. The rod ought to pass readily through the slag, and the coating on it when withdrawn ought to be smooth without showing any nobs. Another method of testing the fluidity is to pour some slag from the converter or a ladle onto a shovel and observe the flow. Foaming of slag occurs during the blister stage; it is attributed by Moore' to the charge being too cool; others attribute it to an excess of Si02 in the charge; LiddelP found in the early stages of the process at Baltimore, when the converter was heated in part as is a reverberatory furnace, that CO was evolved. The temperature of the converter is lower than the temperature of the acid vessel; the slag formed is a subsilicate which has a lower formation temperature than the singulo-silicate; large quantities of ore are charged which cool the bath; and the solution of the constituents of the ore absorbs heat. A temperature exceeding 1150° C. retards the conversion, attacks the lining, and makes sticky tuyeres; running the temperature too low prevents FeO from combining with SiOj, causes the formation of Fe304, and makes the punching of tuyeres very hard. Temperatures of the different stages of a blow have been taken by Clevenger.' The progress of a blow may be followed with clean matte, by watching the flame. It is brown at first, becoming greenish as the Cu-content increases; the greenish flame becomes paler showing pale blue streaks, and is about two-thirds pale blue when the matte has reached the white-metal stage. With leady or dirty mattes the color of the flame gives no indications of the progress made. Examination of the fracture of matte from cooled ladle-samples gives a good idea of the Cu-content. The approach to the white-metal stage is recognized by the behavior of the white metal on the rabble. In blowing for copper the flame becomes at first reddish-brown and later bronze. '£Mg. Min. J., 1910, LxxxDC, 1319. '£nj. Min. J., 1910, xc, 104. 'Met. Chem. Eng., 1913, xi, 448. 342 METALLURGY OF COPPER The end of a blow is determined by the appearance of the copper on a rod inserted into the bath and quickly withdrawn (finishing on the rod). The metal is rose color, full of pinholes and free from dark patches of undecomposed matte. The time required for a blow with 40 tons of 45-per-cent. matte to furnish 45 tons blister copper is about 20 hr.; or a converter produces about 50 tons of blister copper per day from 40-per-cent. matte. A campaign lasts about 100 charges of 40 tons of 4S-per-cent. matte with a production of about 3000 tons of blister copper. It then becomes necessary to repair the lining surrounding the tuyeres for a width of about 18 in. This takes six days, i.e., three days for cooling sufficiently to allow a man to enter the vessel (quicker cooling causes the brick to spall), two days for repairing, and one day for heating up. Smaller repairs at the tuyere-belt, of a section at a time, which can be made from the outside and do not require cooling the vessel, take about two days. Repairs cost about $0.25 per ton of blister copper. The rest of the lining lasts much longer than that of the tuyere-belt; thus the roof, which becomes coated with slag, lasts from two to three times as long, and the lining below the metal line is not attacked by the process; it may have to be replaced only on account of the accumulation of Fe304, which has to be broken out. Additions of spiegeleisen may assist in eating out the crust. No facts have been recorded so far as to the elimination of impurities excepting that the expulsion of As goes on as long as there is S present in the charge. The rate and amount of expulsion ought to be about the same as in the acid process. 173. Products, Losses, and Cost. — The main products are blister copper, slag, and flue-dust. The blister copper is practically the same as in the acid process; analyses have been given in Table 74. It is sometimes poured from the converter into a barrel-shaped oil-heated receiver, resting on a car, and delivered from this to a casting machine. Typical examples are El Paso and Hayden. The slag analyses given in Table 81 show the general character of the product. Table 81. — Analyses of Converter Slag Locality Si02 Pe 1 Mn CaO AI2O3 Cu S Zn Ag, oz. Reference Perth Amboy Garfield 26.6 27-S 23.4 25.0 30-31 22.8 23.6 43-2 47-4 45. 1 44.0 46-47 49-4 48.0 1.6 1.73 3-1 l.l 2.18 2.25 Moore, Eng. Min. J., ipio, 0.9 r.2 1 .0 1.5 LXXXEX, 1319. Neel, oi». cit., 191 1, xci, 964. El Paso l.o I.O 3.6 3.0 2.7 2.7 Private Notes. Steptoe 0.3 0.3 4-7 4.8 1 .0 I. 5 Steptoe Private Notes. Flue-dust. — The amount of flue-dust formed varies of course greatly with the character and size of the siliceous ore charged; it is about 3 per cent. It has been found that with the throat of the converter at the center instead of at the side, the distribution of siliceous ore is more even and thereby the loss in flue-dust about one-half. SMELTING OF COPPER 343 Table 82. — Analyses of Flue-dust from Basic Converter Localily SiOj Fe 1 Mn CaO AUO3 Cu 1 S Zn Pb |Ag, Au, oz. Reference El Paso Cananea 16.3 6.7 8.6! trace 15. o| 1.6 I . I 2,0 45-4 10.9 4.8 12.0 49.56 Private Notes. At the Garfield smeltery an experimental Cottrell electric condensation plant' is in successful operation treating the fume and dust from one Peirce- Smith converter. The collecting-electrodes are vertical wrought-iron pipes, 10 ft. long and 5 in. in diameter, standing on a floor; in the centers are suspended discharge wire-electrodes. The condensation-flue contains 608 pipes and receives through natural draft 40,000-50,000 cu. ft. gas per min. An A. C. of 2300 volts is stepped up by a transformer to 23,000-30,000 volts, and changed by rectifiers, run by three-phase synchronous 2 20- volt motors, to a D. C. of the same voltage. The results have shown that 97.25 per cent, of the Pb in the gases is recovered; the collected material containing Pb 40, Cu 3, Insol, 3, Fe 4, As 7, Zn 2, S 9 per cent; Ag 6.0 and Au 0.05 oz. per ton. An installa- tion for the treatment of the gases of the converting department is being erected. Copper Loss. — The loss of Cu in the basic process is small, as the Cu-con- tent of the slag is < 2 per cent.; in fact 95 per cent, of the Cu in the matte is recovered as blister copper vs. 65 per cent, in the acid converter. The labor on a shift consists of one skimmer and two helpers for punching tuyeres. The Cost.— The cost per ton of copper is inversely proportional to the Cu-content of the matte, e.g., $6.00 with 30-per-cent. vs. $4.00 with 4S-per- cent. matte; but considering that the Cu lost in the ore blast-furnace slag with 30-per-cent. matte is perhaps 0.25 per cent, and with 45-per-cent. matte 0.50 per cent., the balance may lie in favor of 30-per-cent. matte. The cost in Utah is about $4.00 per ton of blister copper, including casting, sampling, and loading; with acid work it was about $9.00. 174. Concrete Examples. — Three concrete examples may serve to further illustrate the work of the Peirce-Smith basic converter. (i) Garfield.— The cylinder is 24 ft. long and 10 ft. in diameter outside dimensions, the central throat 3 ft. 4 in.; the shell of the cylinder is of \-m. steel plate, that of the hood of f-in. plate lined with i-in. cast-iron plates; the lining at the bottom is 20-24 in. thick; at the tuyere belt 18-20 in.; opposite the tuyeres 12 in.; at the hood on the tuyere-side 12 in. and opposite 9 in. There are 37 tuyeres i| in. in diameter placed 18 in. above the bottom. The first charge is 70,000-80,000 lb. matte and 8000 siliceous ore, subsequent charges are 12,000-15,000 lb. matte and 2000-8000 lb. siliceous ore. A single blow on matte lasts 20-30 min., one on white metal 4.5-5 hr.; 'Cottrell, Tr. A. I. M. E., 1912, XLm, 512, 755. Hofman, "General Metallurgy," 1913, p. 858. Howard, W. H., Private Communication, Dec., 1913. 344 METALLURGY OF COPPER the total blowing-time of a charge is 6 hr. on matte, s hr. on copper, or a total of n hr. ; the operating time of a charge is about 16.5 hr. The blast-pressure is 10 lb.; the grade of matte, Cu 40 per cent.; the character of the siliceous ore: SiOa 76, Fe 7, Cu 1.2, S 3 per cent., its size 2I in. The tons of Cu produced per lining average 4200, the maximum has been 6000 tons. The labor required is one skimmer and two punchers per shift. (2) Steptoe.— There are two Peirce-Smith converters, one 32 by 10 ft., the other 24 by 10 ft. outside dimensions. The central throats are 42 and 38 in. in diameter; the lining is 13-18 in. thick; there are 46 and 30 tuyeres 1.25 in. in diameter; the blast-pressure is 14 lb., the matte used assays Cu 34.7, Si02 0.3, Fe 32.2, AI2O3 0.7, S 26 per cent.; the siliceous ore analyzes Cu 4.32, Si02 66;o, Fe 4.82, CaO 0.69, AI2O3 9.9S, S 3.83 per cent.; 0.53 ton is used per ton of Cu. The life of a lining is good for 2000 tons of Cu. There are produced 34 tons of Cu per charge, 1.145 tons of slag (analysis given in Table 81) per ton Cu. The recovery of Cu in form of blister is 80 per cent. (3) Copper Cliff, Ont.' — The Canadian Copper Co. converts its furnace matte with 23-29 per cent. Ni-Cu into refined matte with 80 per cent. Ni-Cu which is poured into cast-iron beds, 30 ft. long by 6 ft. wide made up of sections, when it is broken and shipped to be treated further. The works had ten acid barrel converters with shells 126 by 84 in., which were replaced in 1911 by five Peirce-Smith basic converters. A shell is 37 ft. 2 in. long and 10 ft. in diameter outside dimensions, has one central throat, and two lateral pouring openings. The vessel is manipulated by an oil- instead of a water-cylinder. There are 44 tuyeres, 1.25 in. in diameter and 7 in. apart; beneath the throat there are no tuy- eres. The bottom lining is 15 in. thick, the roof 12 in., the tuyere-belt 24 in.; the inside length is 33 ft. 3 in. In starting, a charge of 60 tons of furnace matte is given and followed by 10 tons of quartzite, the blast is turned on for 1/2-3/4 hr., the slag poured, and at the same time 5-6 tons of furnace matte introduced to be followed by 3 tons of quartzite, when the slag has been removed; the vessel is blown for 40 The cycle of operations is repeated until 300-400 tons of furnace matte mm. have been introduced, and 70-80 tons of converter matte produced, which takes 30-50 hr. blowing-time. A comparison of acid and basic work is given in Tables 83 and 84. Table 83. — Work op Acid Bakrel Converters with Canadian Copper Co., 1911 Furnace matte, Ni-Cu, per cent. One ton converter matte requires Month Quartzite, tons Clay, tons Hour.-min. blowing on one slieU Furnace-matte, tons Tulv . ... 29.00 28.65 25-75 24-35 1 . 10 1 .40 1 .46 1,67 0.52 0.47 0.52 0.77 2-01 i-S6 1-46 2-04 ■4-45 4.66. 4.90 Sept Oct ' Browne, Tr. Canad. Min. Inst., 1912, xv, 115 and Private Notes, 1912. SMELTING OF COPPER 345 Table 84. — Work or Peikce-Smith Basic Converter with Canadian Copper Co., 1911 Furnace matte, Ni-Cu, per cent. One ton converter matte requires Month Quartzite, tons Rock, tons Hour.-min. blowing on one shell Furnace-matte, tons Tulv 29.00 28.6s 2S-7S 24.3s 23.70 0.77 0.71 0.86 0.97 0.86 O.S7 0.50 0.43 0.83 0.77 0-23 0-26 0-17 0-24 0-2S 3.16 2.8; Auff Sept 3-3S 3 -95 3.72 Oct A striking feature is the larger amount of furnace-matte the acid converter requires than the basic, the difference being due to the excessive slopping and spilling with the acid vessel. 175. The Anaconda Basic Converter. — Since 1910 the Washoe works at Anaconda have had in operation as basic converters their former acid vessels (§ 160) lined with magnesite brick. In Figs. 367-369 are given the form and lining of the vessels in use. Tig. 867 U^lKIui! UhdUIMo tud Ijodlum Sll^ta ■nund Tajwia Tlan of Sections A to K Figs. 367-369. — ^Anaconda basic converter. The leading change in form is that the tuyere-pipes are placed tangentially with the lower circle, and that the diameter has been increased from f in. to li in. The shell and tuyere-belt are lined with magnesite mixtures which carry the magnesite brick. The shell-mixture consists of crushed magnesite brick, |-| in., slaked lime, and enough water to make a plastic mass. This is 346 METALLURGY OF COPPER rammed to furnish a firm but elastic support for the brick. The brick used are the standard 9-in. arch, side- and end- wedge forms; the mortar is the undersize (^ ^ /^^ ^Tn \^^ ^^\ fff / V^/^ 11 F 1 ^^'*- 11 ^ \m ^''" r /I --20'0"Outside-Dia.-of Shell- Figs. 379-381.— Great Falls 20-foot basic converter. of a large mouth are, that the gases escape freely, that it is easy to charge liquid matte and solid ore, that there is less tendency to form crusts than with a small one, and that the crusts when formed are easHy removed. If the diameter of the moudi exceeds the dimension advocated, there is an excessive loss of heat, which reduces the amount of material that can be melted by the heat of the charge. The linings and cavities of the two converters are shown clearly in the figures; the lining is backed by a mixture of ground magnesite and sodium silicate; 352 METALLURGY OF COPPER the tuyere-belt is packed with a grout of a similar composition. The position, number and size of tuyeres and their material have been the subject of many experiments. The conclusions reached are, that there should be at least 5 in. from the lowest point of tuyeres to that of the bottom and that there should be enough copper at the finish of a blow to cover the tuyeres; the 12-ft. vessel has 12.5 in., the 20-ft. 11.5 in. The results obtained with a greater and smaller number of tuyeres of different sizes have shown that 12- and 20-ft. converters should have tuyeres 2.25 in. inner diameter, and that the former does its best work with 26 and the latter with 62 pipes. As to the material, tuyere-pipes in the shape of cost blocks of copper worked well while blowing for white metal; they melted while blowing for blister copper. Heavy copper tubes are used at present. They fit into cast-steel blocks attached on the outside of the con- verter sheet; the space between the pipes is packed with a mixture of calcined magnesite and sodium silicate. The amount of air blown into a 12-ft. vessel is about 11,000 cu. ft. per minute. ; the 20-ft. vessel takes as much as 2 2,000 cu. ft. while blowing for blister, and about 18,000 while blowing for white metal. With these amounts the air efficiency, determined by analysis of gases taken from converter, ranges from 83.3 to 100 per cent., the lower figure coming from the end of the blister-stage, the higher from the slag-forming stage. A study of the mechanical effects of the blast upon the molten matte has shown: (i) that the air penetrates the charge only for a short distance and then rises to the surface in larger or smaller bubbles depending upon the sizes of the tuyere-pipes, and (2) that the surface of the charge travels upward away from the tuyeres toward the front, descends there toward the bottom, and returns across the bottom toward the tuyeres.' In starting a newly lined converter, e.g., the 12-ft., the lining is dried and warmed with a wood fire for 36-48 hr.; the ashes are removed, and two ladles or 16 tons of matte (Cu 35 per cent.) poured in; the vessel is turned over, blown 2-3 min., and turned down to see the effect the blow has had on the brick lining. This operation is continued with additions of fresh liquid matte and with extensions of the duration of the blows until the cracks between the bricks have been filled and the surfaces so coated that the individual bricks cannot be recog- nized; the remaining matte is poured off, and the converter left undisturbed for a few hours to harden the matte in the crevices and the coating on the outside of the brick. The vessel is now ready for normal work. In this there are charged two ladles or 16 tons of liquid matte, then 3700 lb. siliceous ore (Cu 3.5, Si02 57-8, FeO 13.0, AI2O3 n.o CaO o.i, S 12. i per cent.); the blast is started, the vessel turned over and blown for 30 to 40 min. with a pressure of 14 lb.; the slag formed is poured off. There is now poured in one ladle or 8 tons of matte and shot in 1800 to 3700 lb. of ore depending upon the grade of the matte, followed by 1.0-2.5 tons of converter cleanings and cold matte, all of which reduce'the temperature of the bath; the vessel is again blown and skimmed. The opera- ^ See also Haas, Nov., 1913, Btdl. A. I. M. E. SMELTING OF COPPER 353 tions are repeated until five or six ladles of liquid furnace matte have been intro- duced, and the converter matte has been brought forward to near the white- metal stage (Cu 70-75 per cent., recognized by its behavior on the rabble); the slag formed is poured ofi as much as possible and the rest skimmed; "dope" are fed in the form of scrap copper, white metal, and cleanings which reduce the temperature of the bath, which is now blown to a finish in 5-6 min. for every ton of copper in the vessel (12-14 tons of Cu or 18 tons of white metal). The tuyeres are punched more or less continually, less so during the slag- ging, than the blister-stage. The progress in the process is usually judged by the flame; when in doubt, the vessel is turned down and the matte examined, or the copper is tested on the rod. The actual blowing- time is 20 min. per ton of copper produced; a charge with matte of 38.9 per cent. Cu takes 8.5 tons of ore (Fe 10, Cu 3.5 per cent.), or 0.7 tons of ore per ton of Cu; the charge produces 12-14 tons of Cu. The converter copper is poured into ladles and transferred to the copper- casting machine (Figs. 353-354); the converter-slag (Si02 19.7, FeO 60.7, AlzOs 6.3, CaO 1.3, MgO 0.25, Cu 1.6-2.0, S 2.0 per cent) is also poured into a ladle and handled in the slag-casting machine (Figs. 350-351). A new departure has been made by Wheeler and Krejci' with the 20-ft. basic converter at Great Falls, Mont., in that they coat the magnesite lining with a mixture of fused oxidized iron (magnetite ?) and slag which is formed by blowing matte at a low temperature either alone or in the presence of a small amount of Si02. A newly-lined, converter is brought slowly to a bright red, charged with liquid matte of a low grade, say 35 per cent. Cu, blown for 10-15 niin., whereupon cold matte is added to reduce the temperature. These operations are repeated until the matte in the vessel has been brought forward to white metal, which is then poured. The brick of the empty vessel will be found to have been coated with what may be briefly called magnetite. A fresh charge of liquid matte is given and a little less siliceous ore added than in regular work; the charge is blown in the usual way, and is followed by other similar charges until the coating has become so thick that all joints of the brickwork have disappeared. The thickness is regulated by the temperature of the converter and the percentage of SiOz of the slag. V. The Sulphide Copper Smelting Plant 177. General Arrangement of Plant. — The general characteristics of modern smelteries treating sulphide copper ores are: the large scale of operations demanding mechanical handling of materials in the different departments, and the combination, wherever feasible, of blast- and reverberatory furnace smelting of ore followed by converting of matte into blister copper. With most blast-furnaces the coarse raw ore is smelted by the partial pyritic process; ' Bull. A. I. M. E., Feb., 1914. Discussion, Eng. Min. J., 1914, xcvii, 431 (James), 530 (Krejci), 628 (Howard), 724 (Williams), 821 (Merton, Krejci). 23 3S4 METALLURGY OF COPPER it is the exception that coarse ore undergoes a preliminary oxidizing roast. The fine ore or concentrate to be smelted in the reverberatory furnace is usually first rough-roasted in a mechanical furnace. The flow of materials, Fig. 382, and the plan, Fig. 383, of the Washoe smelting at Anaconda show in general the sequence of operations and the locations of the several divisions. At this plant' first-class ore, which is coarse and' contains 10 per cent. Cu and over, goes direct to the blast-furnaces (^§90); second-class ore, with perhaps 3 per cent. Cu, passes through the concentrating plant and goes in the form of concentrate, with about the same tenor in Cu as first-class ore, to the MacDoug- all roasting furnaces (§64). The roasted concentrate ("calcine") is smelted in Seoond Class Ore Otusber t^ IlollB^ u^m IfnrefiDcd Cflpper _^r^f=p^ To Rcanaiy Fig. 382. — Flow of materials through the Anaconda smeltery. reverberatory furnaces (§131). The matte from the blast- and reverberatory furnaces is hauled by compressed-air locomotion on the same level in ladles to the converters (§175), whence the blister copper is poured into ladles and transferred by means of traveUng electric cranes to reverberatory casting furnaces, where it is fire-refined, and shipped in the form of anode-copper to be refined electrolytically. The plan, Fig. 383, of the works shows the location of the different departments. The sampler, the concentrator, the roaster, the blast-furnace, the reverberatory furnace, and the converter-divisions, are in separate buildings. Communication between them is effected by thirteen compressed-air locomo- tives working under a pressure of from 800 to 900 lb. They handle daily about 13,000 tons of materials,"which represents a daily capacity of 7000 tons of ore and a production of 500,000 lb. blister copper. The new works of the International Smelting and Refining Co. at Tooele'' ' Hofman, Tr. A.I.M. E., I904,xxxiv, 251. Austin, op. cit., 1906, xxxvii, 431. Correspondent, Mines and Minerals, 1907, xxvill 131, 248. "A Brief Description of the Washoe Smelter" by members of the Anaconda Copper Mining Co.'s StaflF, Anaconda, Mont., 1907. 2 Palmer, Min. World, 1910, Xxxil, 419. Repath and McGregor, Met. Chem. Eng., 191 1, rx,i5. Thomson-Sicka, Tr. A. I. M. E., 1913, xlvi. Missing Page SMELTING OF COPPER ZSS are built on the same general plan as the Washoe plant excepting that they smelt exclusively concentrates in reverberatory furnaces, and that the matte is tapped direct into the converters. This last feature has been changed, as it was found impossible to judge of the amount of matte a converter received. The matte is therefore weighed as is the case in other works. The smeltery of the British Columbia Copper Co. at Greenwood, B. C. which smelts in the blast-furnace only and converts the resulting matte, has Fig. 383. — General plan of Anaconda smeltery. been described with drawings by McAlUster.^ It resembles in general arrange- ment the plant of the Canadian Copper Co.^ An early plan of the smeltery at Great Falls, Mont., is given by Higgens' and a later illustration is that of Fig. 238. The ore-treatment is similar to that of the Anaconda works. The Flow-sheet (191 2) of the Cananea Consolidated Copper Co.'s smelting, shown in Fig. 384, gives in detail, with diagrammatic sketches of apparatus, the complications of the work in a modern plant treating sulphide ore by blast- and reverberatory-furnace smelting followed by converting. It shows ' Eng. Min. J., 1911, xci, ion. ^Eng. Min. J., 1903, lxxvi, ioo8; Browne, Canad, Min. J., 1907, xxviii, 305, ' Eng. Min. J., igog, lxxxvii, 156. 3S6 METALLURGY OF COPPER how the raw materials enter each division and gives the paths of the inter- mediary and finished products. The operations of the several divisions have been discussed in the preceding pages. The large number of roads indi- cated in the 'figure, at first bewildering, becomes simple and clear when followed step by step. A general review is therefore unnecessary. The feature that is unique is the Dwight-Messiter Ore-bedding System' recently introduced also at the smeltery of the Calumet & Arizona Mg. Co.^ and the Tennessee Copper Co.' by means of which ores for the blast-furnaces are sampled, bedded, reclaimed, and delivered mechanically to the blast-furnace storage bins. I^a J^^^ ■^ .^^^^ -•iswiS^P^' Fig. 385. — Overhead skeleton structure of Dwight Messiter Ore-bedding System, Cananea. In the flow-sheet of Cananea, the narrow- and standard-gauge railroad brings all blast-furnace materials, such as coarse ore, coarse concentrates, fluxes, and secondaries (intermediary products) and delivers them over scales to nineteen bins. From the hoppers of these the materials are supplied by traveling shaker-feeds to the conveyor belt. No i, which discharges onto a grizzlie whence the oversize goes to Farrel crushers, and the undersize as well as the discharge of the crushers to conveyor belt, No. 2, and the Bishop-Vezin sampler. The sample goes by the 20-in. sample conveyor to the sampling mill, while the rejects return to the main stream by bucket elevator and conveyor No. 3. Trippers at the junctions of conveyor, No. 3, and belts 4^, 4^, 4*^, deliver the materials to these belts, which pass over the center lines of the three storage beds and deposit the material in 1 Woodbridge, Eng. Min. J., 1906, Lxxxii, 624. Messiter, Min. Sc. Press, 1907, xcv, 528; 1909, xcviii, 361. Herrick, Mines and Minerals, 1909-10, xxx,6s; Pamphlet, Robins Conveying Belt Co., New York. " Editor, Eng. Min. J., 1912, xciii, 682. ' Wierum, Eng. Min. J., 1913, xcvi, 435. SMELTING OF COPPER 357 long narrow strips. A bed holds 8500 tons of ore. A high-speed traveling tripper moves constantly to and fro over a bed upon an overhead skeleton structure, Fig. 385, and deposits the materials uniformly so that all portions of any cross-section of the bed will have the same composition. On the floor and along the sides of each bed are rails upon which travels the reclaiming machine, Fig. 38s. at a speed of f in. per minute. The machine is a bridge-shaped struc- ture spanning the width of the bed and reaching beyond it on one side to cover the conveyor belt placed there in a trench, Fig. 385, and Fig. 384 belts 5^, 5^, 5^. In the front of the bridge is suspended a scraper conveyor which operates in a trough consisting of a bottom and a back plate. The trough is placed at an angle to the axis of the conveyor and carries at the front plow-shaped extensions. In front of the machine is mounted an adjustable triangular harrow having a reciprocal motion to dislodge the ore, that it may run down the face of the bed, fall on the bottom plate of the conveyor, be carried along to the discharge end, and fall through a short chute onto the belt conveyor in the trench. The belts 5^, 5^, s'^j'in the trenches deliver the ore-mixture to conveyor No. 6, whence the material travels over belts No. 7 and No. 8 to be delivered by movable trippers into the eight 80-ton steel storage bins of the blast-furnaces, Figs. 171-174. It takes 4 days of 10 hr. each to deliver the 8500 tons of ore forming a bed; the labor required is i ore-bin man, 2 helpers, 4 men on belts, i crusher-man, and 5 to 6 laborers. A bed is reclaimed at the rate of 150 tons per hour by i man on machine, i helper, \ oiler and i tripper. The cost of handling from unloading to delivery into the blast-furnace bins, including labor, power, repair, and belt removal, is 9 cents per hour. The handling of ores at the Copper Queen Smeltery, Douglas, Ariz.,"^ form another instance of the application of machinery to bedding and re- claiming. In the yard there are five ore pits, 40 ft. wide and 11 ft. deep; four are 825 ft. long, the fifth 1000 ft. Ore from the Copper Queen mine is received in 14-car trainloads. The cars are loaded at the mine by means of belts and trippers;^ the contents are therefore of uniform composition and can be placed in the pits one lot above the other in layers and form a uniform mixture. Other ores are spread over the mixture in order to obtain a self-fluxing charge. The bed in a pit is reclaimed by means of a No.4Thew steam-shovel (Lorain, 0.) which has a capacity of 100 tons per hour. The shovel is set to dig to the level of the track, it loads a train of 20 blast-furnace charging cars, each with a measured amount of about 2700 lb. mixture. The train is hauled under 20 coke-hoppers, where each car receives a weighed amount of fuel, and then to the feed-floor on either side of a row of 10 blast-furnaces (Table 82) where there are two tracks, one for full, the other for empty cars. 178. Cost of Plant. — Most of the earlier copper smelteries were connected with little mines and had small capacities. The ores treated were comparatively ' Woodbridge, Eng. Min. J., 1906, Lxxxii, 242. Douglas, Tr. Inst. Min. Met., 1912-13, xxii, 532. ' Milton, Mines and Minerals, 1909, xxx, 148. 358 METALLURGY OF COPPER rich in Cu, so that the cost of treatment could be high and still leave a satisfactory profit. The Cu-content of the ores mined has grown smaller, the mines have had to be worked on a larger scale, the smelteries have had to work larger quanti- ties of ore with a diminishing copper-content. In other words, the capacities of modern smelteries have become very large. This calls for large expenditures of capital for building a plant and an equally large capital for operating. The figures given in Table 85 are reUable estimates furnished by the best of authority. Table 85. — Capital of Establishment and op Working for Sulphide Copper Smelteries Product in 24 hr. Necessary capital: Annual product: Character of plant Of establish- ment 1 Of working (additional) 2 Amount and kind Value A. Copper blast-furnaces, partial- pyritic smelting, 1000 tons of ore per 24 hr. 100 tons, 4S-per cent, copper matte. $1,250,000 $1,000,000 36,500 tons matte $3,285,000 B. MacDougall roasting and rever- beratory smelting, 1000 tons ore in 34 hr. 100 tons, 4S-per cent, copper matte. $1,250,000 $1,000,000 36,500 tons matte $3,285,000 C. Converting 45-per cent, copper matte, 200 tons per 24 hr. 90 tons blister copper. $250,000 $2,000,000 32,850 blister Cu tons, $7,570,000 As regards the depreciation of a copper smeltery, Mathewson' estimates that, providing repairs are well kept up, the life of a plant is 20 years, i.e., that 5 per cent, of the original cost of plant should be written off for depreciation He also states that if a plant is dismantled, only 5.49 per cent, of the original cost will be recovered as salvage. B. Smelting Oxide Copper Ores 179. Smelting Oxide Copper Ore in General.^ — Oxide copper ores rich in copper used to be smelted in Arizona and New Mexico in water-jacket blast- furnaces, having internal crucibles, for black copper (96 per cent. Cu), and waste slag, with from 1.5 to 2.5 per cent. Cu.^ The industry started about 1881, 1 Capital of establishment: If the three operations A, B, C are combined in one plant a reduction of $250,000 can be made from the above totals. All buildings are supposed to be fire-proof throughout and of best modern construction. The prices include all accessories necessary for the operation of such plants as A and B; plant C would be merely an addition to A and B, or A-|-B. ' Capital of working: If all three operations are combined in one plant, $2,000,000 is suflScient for all working capital. The value of copper in form of blister is figured at 10 cents a pound. No values of silver or gold are included in the above estimates. ' Eng. Min. J., 1906, Lxxxii, 888. ■■ Douglas, Min. Res., U. S. Geol. Surv., 1882, p. 261; 1883-84, p. 397; Tr. A.I.M. £., 1909, XL, 422. Wendt, op. cit., 1886-87, xv, 25. Austin, Min. Sc. Press, 1908, xcvi, 196. ' Channing, Tr. A. I. M. E., 1910, XLi, 883. SMELTING OF COPPER 359 when the cost of coke was too high to permit charging enough fuel to make clean slags. Smelting the old slag dumps later on with sulphide ore showed' that in some instances the copper-content exceeded 2.5 per cent, viz., Globe 3.5, Morenci 4.5 per cent. At present rich oxide ore is mixed with sulphide material and smelted for matte, which is converted; poor oxide ore has been leached (§ 204). Attempts have been made^ to reduce to the metallic state the copper from a mixture of rich oxide ore and fuel by passing it through an inclined revolving cylinder heated internally from a fire-place at the lower end, and then separating the globules of copper by mechanical concentration. Such work can give only a small yield, and may be justified in regions where other processes are not feasible. 180. Early Work in Arizona. — Some of the leading facts of the former Arizona practice are given in Tables 86-89. The blast-furnaces were water- jacketed throughout; most of them were cir- cular in cross-section, only a few rectangular. In order to prevent the black copper from chilling, all furnaces had internal crucibles from which slag was tapped at intervals into slag-pots and black copper into pig- molds holding about 250 lb. The Arizona copper furnace is shown in Fig. 386 and a pig-mold in Fig. 387. The furnace, 36 in. in diameter at tuyeres, 54 in. at throat and 6 ft. working height, has the form of an inverted cone resting on a cast- iron bed-plate supported by four hollow cast- FiG. 386. — Arizona copper blast-furnace. Fig. 387. — Copper mold. iron columns. It is water-jacketed from the throat to below the wind-box; the water-space is 9 in. wide at the bottom and 4.5 in. at the top. The outer shell extends downward below the jacket and forms the side wall of the crucible. The water-inlet pipe, 2.5 in. in diameter, is usually at the center, and not both near bottom and top as seen in the figure. The feed- ' Douglas, Tr. A. I. M. E., 1909, xl, 422. 'Experiments of Caspari-FIegel, Metall-Erz, 1913, x, 253. 36o METALLURGY OF COPPER water, under a pressure of 8 or lo ft., strikes a deflecting plate so as not to im- pinge upon the inner shell; the overflow-pipe, 2.75 in. in diameter, is tapped into the top of the jacket; a 35-in. furnace takes 950 gal. water per hour, a 42-in. 1 200 gal., a 48-in. 1500 gal. The crucible is 24 in. deep; its bottom is formed by two cast-iron plates hinged to the bed-plate; the lining reaching to the tuyeres is brasque, or sand and burnt clay, or quartz and slag. The removable air-box, 10 in. above the slag-tap, has six bronze tuyere-openings; opposite each there is on the outer side a peep- and poking-hole closed by a cap provided with a mica shield. The feed-door is in the hood ending in the chimney, the blast-pipe ends at the wind-box, the metal-tap is 14 in below the slag-tap. Walker and Murphy^ surrounded the crucible with an air-jacket to keep the walls cool and to warm the blast. This plan has not been, adopted by others for obvious reasons. The cost of smelting a ton of ore in 1892-94 at the works of the Old Dominion Copper Co. at Globe, Ariz., with two 36-in. blast-furnaces is given by Austin^ as $8.96 per ton yielding 260 lb. black copper Table 86. — Analyses of Arizona Oxide Copper Ore Locality Cu Si02 FeO MnO CaO Reference IjOnefellow 38.80 21 .67 17.17 21.9s II. 17 15.17 II. 15 17.25 26.80 48.90 67.00 35.3 10.40 13.76 12.09 8.88 28.7 Longfellow 7.43 7.49 \ Wendt, Tr.A.I.M. ' E., 1886-87, XV, 25 Old Dominion 22. 2 Austin, Min. Sc. Press, 1908, xcvi, 196. Table 87. — Arizona Blastpcrnaces for Oxide Copper Ore Tuyere section, in. Throat section, in. .2? oe • •S -^ is Tuyeres P. *^ H CO " Charge la "a! 3 I . U N ft C s 3 ■ CD P. Smeltery 6 2; E .2 Q a 4-J c aj is u a S Copper Queen . . Old Dominion . . 36 diam. 36 diam. 42 diam. 33X66 36X90 54 diam. 6.0 6 6 6 14 10 3 and 5 3 4 and 5 ii.7S 3. SO JO 14 47 56 SO 85. 12.9 17.3 II. 8 17. 01 10 13.2 10 IS. I 8 1S.2. 12 1.27 1 .11 Wendt Douglas Peters 54 diam. 54X87 48X90 6.S 10.5 0.0 8 6 II 12 14 24 Detroit Howe ' Walker, Eng. Min. J., 1893, LVi, 619. 2 Min. Sc. Press, 1908, xcvi, 196. SMELTING OF COPPER 361 Table 88. — Analyses or Arizona Blast-furnace Slags Smeltery Copper Queen . . Copper Queen . . Detroit Detroit Prince Old Dominion. . . United Verde. . . Bisbee SiOj 24.67 30.06 34-34 29. SO 27. 16 27 ■ 23 35-79 28.0 PeO 44-85 53-36 32.27 37-08 34-62 SI -30 37-89 29.0 MnO 0.39 II. 10 8.05 1. 13 0.49 • 65 CaO 10.92 10.13 9.02 17.42 5.14 12.98 9. MgO i:.30 7-44 3. SI 2-54 0-75 .MM, I Cu I CuO I S I Reference 15.57 11.64 14.07 14.70 5-22 8.29 27.0 2 . ro O.T5 .82! .64 o. 18 1.32 3-76 2.59 Wendt, Tr.A.I. M. E., 1886-87, Howe, "Copper Smelting," p. 78 Table 89. — Analyses of Arizona Black Copper Smeltery Cu s Pe As Sb Bi i Ins. Reference P ince 95.00 98.91 98.27 98.24 97.52 0.44 0.64 0.60 0-53 0.69 4-23 0. 12 0.73 0.80 0.Q7 0.51 0.065 0.060 0.007 0. 180 Wendt, op. cit. 0.057 0.039 U.0S4 0.052 0.008 0.019 U.021 0.014 O.OIO trace 0.006 trace Old Dominion Old Dominion Old -Dominion Austin, Min. Sc. Press, 1908, xcvi. 196. C. Smelting Native Copper Ore' 181. The Ore. — The discussion is confined to the smelting of Lake Superior ore (§ 4S). This is passed through ore-dressing works, the products of which are treated, with some mass copper coming direct from the mines, in five smelting plants, viz., the Calumet & Hecla Mining Co., Hubbel, Mich., and Buffalo, N. Y.; the Quincy Mining Co., Ripley, Mich.; the Lake Superior Smelting Co., Dollar Bay, Mich.; and the Michigan Smelting Co., Houghton, Mich. The product of the ore-dressing works, called "mineral," contains three grades of material: barrel work, headings, and grades. The smelting works treat: (i) Mass.— Pure copper with adhering rock, sorted at the mine, varying in weight from a few pounds to several tons; Cu 70+ per cent. (2) Barrel Work. — Pure material, too large to go into mill, sorted at the mine and the mill (as the rock is fed to the stamps), consisting of pieces orange size and smaller; Cu 65-75 per cent. (3) Headings. — Pieces of clean copper, egg size, taken from the stamp- mortar; Cu 65-85 per cent. (4) Grades. — Concentrates varying from slime to walnut size; Cu. 20-70 per cent. The average of the four classes of materials is about 70 per cent. Cu. •Egleston, Tr. A. I. M. E., 1880-81, ix, 678. Douglas, Min. Res., U. S. Geol. Surv., 1882, 259. Cooper, Proc. Lake Sup. Min. Inst., 1901, vii, 44. Rickard, Eng. Min. J., 1904, Lxxviii, 984 (also "The Copper Mines of Lake Superior," McGraw-Hill Book Co., New York, 1903, pp. 142-151). White, Eng. Min. J., 1905, LXXix, 842. Austin, op. cit., 1906, lxxxi, 83. Conant, School Min. Quart., 1911, xxxii, 285. 362 METALLURGY OF COPPER The proportion which each of these four classes of materials forms, varies with the character of the deposit, whether conglomerate or amygdaloid, and with the practice of the concentration plant. In Table 90 are given two examples from the Michigan Smelting Co. for 1906 and 191 1. Data from some of the other works, which are not for publication, are similar.^ Table 90. — Proportion and Copper-content or Smelting Material of the Michigan Smelting Co. 1936 igii Per cent., total Per cent., Cu Per cent., total Per cent., Cu Mass ii-S 6.2 15.1 67.2 68 68 90 68.72 5 S 30 60 70 Barrel work. . 65 8S Grades . 50 Total and average. . . . 100. 71.82 100 62.25 182. Process. — The smelting process is essentially an oxidizing fusion in a reverberatory furnace, with or without the addition of some limestone as flux, in which the gangue is scorified and the slag skimmed as fast as it is formed. The oxidizing smelting is followed by a refining of the copper, either in the same or in another furnace into which the metal has been tapped. The refined copper is ladled or cast into commercial forms or into anodes. The reverberatory furnace ore-slag (Cu 12-15, Si02 40, FeO 16, CaO 11, MgO 5, AI2O3 13 per cent.) is crushed and smelted with the necessary flux (limestone) in a blast-furnace in which the anthracite {\ of the whole fuel) charged is intended to serve mainly as a reducing, and the coke (f of the whole) mainly as the heat-producing agent. The refining slags, Cu 30-35 per cent., sometimes are added to the ore-smelting charges. The "cupola copper" of the blast-furnace is tapped into molds to form blocks weighing 200-400 lb., and refined. The refining is usually carried on in a special furnace, as the cupola copper is impure (Cu 94+ per cent.). The waste-slag (Cu 0.75, SiOj 40, FeO 20, CaO 16, AI2O3 14 per cent.) is run through brick-lined forehearths either into slag-pots, or is granulated, and goes to the dump. With the exception of slime concentrates, all the material received by a smeltery goes straight into the reverberatory smelting furnace; the slimes are agglomerated for blast-furnace treatment, either by fusion in a reverberatory furnace, or by briquetting. 183. The Reverberatory Fvimace. — The reverberatory smelting and refinmg furnaces have the general form of the English reverberatory smelting furnace; the main difference is that the hearth slopes from the fire-bridge to the flue, beneath which the refined copper is taken out either by ladling or by allowing it to run into a casting-ladle. In furnaces treating mass copper, part of the ' Parmelee, "Quincy Smeltery," Met. Chem. Eng., 1913, XI, 122. SMELTING OF COPPER 363 roof can be, or formerly could be raised by a crane and swung to one side to permit charging of the mass. As this weakens the roof, and as there are now simple means of cutting up large pieces of copper, the hole in the roof has become the exception. Large pieces of mass copper, which are of less frequent occurrence than formerly, are often put aside until the roof of a furnace has to be renewed, when they are lowered onto the uncovered hearth. The fire-bridge of a reverberatory furnace as well as the roof above it are provided with ports to admit the necessary air to the hearth. The reverberatory furnace used thirty years ago had a capacity of from 7 to 9 tons of copper; the hearths were 7X12 ft. and 8X14 ft., and the respective fire-boxes 36X48 in. and 42X42 in. The capacities were later increased to 20 tons per day with a hearth 11X14 ft. and a fire-box 4X4 ft.; in recent years furnaces of larger capacities have been erected. A further improvement is the use of waste-heat boilers for utilizing the heat in the waste gases. Experi- ments with mechanical stokers have not given results which show clearly that they are more advantageous than hand-firing. The dimensions of some reverberatory furnaces and the work done in them at Lake Superior are given in Table 91. Tabl"e 91. — ^Lake Supekior Reverberatory Furnaces Michigan Smelting Co. Calumet & Hecla Mining Co. Length of hearth, ft. in Width of hearth at bridge, ft. in Width of hearth at middle, ft. in Width of hearth at flue, ft. in Hearth area, sq. ft Hearth depth, in Hearth thickness, in Hearth material Hearth support Grate, length, ft. in Grate, width, ft. in Grate, depth below top of bridge at bridge, ft. in Grate, depth below top of bridge at opposite end, ft. in. Grate, area, sq. ft Ratio, hearth to grate area Roof, height above bridge, ft. in Roof, height above hearth at bridge, ft. in Roof, height above hearth at flue, ft. in Bridge, width, ft. in Vulcatory (flue leading out of roof), ft. in Flue leading to chimney, ft. in Chimney, inside diameter, ft. in ." . Chimney, height, ft Waste-heat boiler, number, kind, h.p 23 3 7' o" 14' /)" 0' 11" 260 40" 27" Brick and fused sand Concrete 4' 3" 6' 2" 3' 10" 3' 10" 26 10; I I' 8" S'6" o' 6" 4' 0" 2' 4"X4' 2' 4"X4' ro' for all furnaces 250' One 250-h.p. Stirling Hubbell, Mich. 20' 6" 6' o" 13' o" 3' 0" 202 9.5" 12" Sand Fire-brick arch 4' 0" 6' 0" 2' 9" 2' 9" 24 8.41:1 i' 7" 2' 6" 1' o" 3' o" 45 sq.' (feet) 3'X3' 90' None Buffalo, N. Y. Melting 21' 0" 8' 4" 13' 8" 4' "" 23s At side 7" 14" Sand Iron plate and arch- brick 4' 10" 8' o" 3' 3" 3' 4" 38.66 6.08:1 t' ioJ" 3' S" 2' t" 3' s" 3' 3"Xl'6" l' 4"Xl' 6" 3' 83' iSO-h.p. Cook Buffalo, N. Y. Melting and reflning 21' 2" 8' 6" IS' 6" S' o" 244 13 Sand Brick 4' 6" 7' 6" 4' 2" 4' 3" 33.7s 7.23:1 t' 9" 3' 0" 3' 3" 3' S" 3' 3"Xl'6" l' 4"Xl' 6" 3' 83' 364 METALLURGY OF COPPER Table 91. — Lake Superior Reverberatory Furnaces. — Continued Michigan Smelting Co. Calumet & Hecla Mining Co. Hubbell, Mich. Buffalo, N. Y. Buffalo, N. Y. Molten cop- per, and cupola blocks 22 I 8 6 2 Mineral, mass barrel work Mineral and cathodes 2 Charge, time of melting and raking, hr Charge, time of fining, hr 16 2 2.5 Charge, time of casting, hr Charge, time of cleaning up and recharging, hr . . 4 I 150 0.6 IS 9 54 Hardwood 15 2000 14 Cake, ingot 60 Walker machine 46 17 12 10 7 45 Blast-furnace Charge, tons per sq. ft. of hearth in 24 hr 7-5 59. 5 Charcoal, kind, lb. per charge 3500 lb. Hard wood 680 lb. Cake, ingot, anode SO=fc Hand Ingot, cake. bar, anode go Bull-ladle Slac SiOs 21 .3 Fe(MnO)0 AhOa 8.4 CaO Cu 43.4 15 beratory, mostly to blast-furnace Two furnaces, shown in Figs. 388-391 and 393-39S, may serve as examples of the old and modern types. The furnace, Figs. 388-391, is the one formerly used by Park Brothers & Co., Pittsburgh, Pa.'^ The drawings with legend require little discussion. The sand-bottom, i, is burnt in on the inverted arch, i' ; the crown of the latter rests upon the arch, j, built up from the foundations; the space between the arches is filled with slag and mill tailings. The furnace is run with undergrate blast delivered to the central underground flue; the blast strikes in the ashpit a hinged deflecting-plate which forces part of the under-wind to pass through the upturned ports in the hollow bridge-wall. This furnace, 12 ft. 10 in. long by 9 ft. 9 in. wide, used to receive a charge of 7 tons of mineral to be worked in 24 hr. The furnaces in operation at present are larger, holding from 25 to 35 tons of copper; further, the former sand-bottom has been replaced in a few instances by two courses of silica-brick laid in the form of an inverted arch, with about 1 Egleston, Tr. A. I. M. E., 1880-81, ix, 678. Johns, Eng. Min. J., 1912, xciii, 1183, history. SMELTING OF COPPER 36s J_L JJ L Vertical Longitudinal Section on Line A-A Fig. 391 '//////////^///////^^M^ Horizontal Section on Line B-B In. 12 6 1 2 3 4 5 Scale 10 15 Ft. Plra pUM 6 iia pit h Lftbontoij of the fninaoe n Ohlmnej 53 feet bleb z Sand bottom O Fluo to n C Brldti W.11 , i' Mortoa Iriolt «rob Q M.r.blo roof d Iljbt ilr ponaco 1.6 i S In. oonttoUod bj iC->e e j Awi ■apportliij! tbo heulh S Cut boo I bor. boldlni ftmooo tosotter e B,Ug. «;„ ^Obt^tagLr « Al.p»»s..l»ai.g fdf" •»?!»« •" / Air puiaso oootiollod bj tb« vnlu 17 / WotkluB door " Ba:k«U [or Dm ff Boof nlvo m Elro door Figs. 388 to 391. — Reverberatory furnace for Lake Superior copper ore. 366 METALLURGY OF COPPER 1 2 in. spring, on cast-iron plates resting upon brick piers. The melting furnace of the Michigan smeltery, shown in Fig. 392,' is of this type. ^ > Fig. 392. — Cross-section of melting furnace, 16 by 35 ft., Michigan smeltery. Fig. 393,— Reverboratory furnace, Calumet & Hecla Mining Co., Buffalo, N. Y. With the Calumet & Hecla furnace of Buffalo, N. Y., Figs. 393-395, only the central part of the hearth is supported by cast-iron plates, which cover an under- 1 En^. Mifi, /., 1906, Lxxxi, 83. SMELTING OF COPPER 367 368 METALLURGY OF COPPER ground chamber running along the median line, while the rest is carried by a concrete foundation. The hearth, 21 ft. by 13 ft. 8 in., has a silica-brick floor on which is the burnt-in sand-bottom. The sides are silica-brick backed by two 4|-in. courses of fire-brick; the roof is of silica-brick; an inclined flue leads the gases to a circular brick stack. Of special interest are the tie-rods which bind .together the sides. They pass through the side and foundation walls in inclined steel pipes and are connected in the longitudinal chamber by short horizontal rods; the rods, some distance frorn the hotter part of the furnace, are air-cooled, as is the case with the longitudinal tie-rods taking up the end-thrust Fig. 39S. — Reverberatory furnace Calumet & Hecla Mining Co., Buffalo, N. Y. of the furnace, by air passing through the pipes. The fire-box and roof have the usual air-ports. The 4-ft. bars of the rocking grate run parallel with the length of the furnace; the outer wall of the fire-box rests on a cast-iron plate supported by cast-iron columns, thus leaving open the ashpit below. 184. Mode of Operating. — The mode of operating, which includes the six operations of charging, melting and skimming, fining or rabbling, poling or refining, ladling or casting, and repairing is similar to that of refining black or blister copper to be taken up in detail in §186. Only a brief outline will be given here. At most of the works the entire process is carried on in the same reverberatory furnace,, and the time given to a charge is about 24 hr. : charging, melting and skimming, 14-16 hr.; fining (rabbling) 1-2 hr.; poling (refining) 2 hr.; ladling (casting) and repairing, 4-5 hr. The casting is intended to be begun with the arrival of the day shift. SMELTING OF COPPER 369 At the Michigan smeltery' melting and skimming are carried on in two melt- ing furnaces with shallow hearths, 35X16 ft., each of which treats in 24 hr. from 60 to 80 tons of charge, according to the character of the material As fast as the slag is formed and has accumulated in quantities of about 5000 lb., it is skimmed into a set of five molds resting on a car. When the full charge for a furnace has been smelted, the slag is skimmed off, and the metal tapped into an adjoining refining furnace, placed 7 ft. below, with deep hearth, 23 ft. 3 in. by 14 ft. 4 in. As this process does not require 24 hr. to finish the molten copper, it is charged with from 10 to 15 tons of mass or barrel work which is melted down in time for the furnace to receive the liquid copper from its melting furnace. At the Buffalo works of the Calumet & Hecla Mining Co., a similar proceeding is in operation with No. i mineral which assays Cu 70 per cent, and passes for the greater part through a 40-mesh sieve. The melting furnace is fired and charged with mineral to which 1.5 per cent, limestone has been added as flux. When the first charge is melted, more charge is given, and the slag skimmed when necessary. This is continued until the furnace is filled with copper, and then its content tapped into a refining furnace, where the metal is brought to tough-pitch and cast with a Walker machine (§ 194) into ingots and ingot-bars. The question of keeping melting and refining separate is governed mainly by the size of the charge to be treated. With charges of 20 to 35 tons there is no reason for keeping apart the two operations, as the depth of the hearth is not so great as to prevent the smelting from being efficient. With a charge of 150 tons, where the hearth must be about 3 ft. deep, the smelting efficiency would be very low, as the mineral lying on the bottom would not be exposed to a greater heat than the temperature of the molten copper floating above, and the formation temperature of the slag is higher than that of the molten copper. Both causes make it extremely difficult for the mineral to reach the hot zone of the furnace. In ordinary practice the coarser concentrate to be charged into a furnace is drawn from draining storage-bins, either into i- or 2-ton cars and brought to the hoppers in the roof of the furnace, or on to a conveyor which delivers into the hoppers. In starting a furnace, the working-bottom of the hearth is covered with concentrate, then follows coarse copper, such as mass, or barrel work, headers fed through the side doors; the rest of the charge is dropped through the hopper in the roof. The charge may receive as much as 8 per cent, limestone as flux; frequently it is covered with rich refinery slag. The doors are closed and luted, and the fire is urged. The charge melts down gradually in about 16 hr.; slag is skimmed as fast as it is formed. During the melting most of the Fe present is scorified. The fining (rabbUng) was formerly done by hand (flapping) with a rabble (head 4X6 in., handle 8 ft. long). At present it is accomplished ' White, Eng. Min. J., 1905, Lxxrx, 842. Austin, op. cit., 1906, Lxxxi, 83. 24 370 METALLURGY OF COPPER better and more quickly by compressed air of 16-20 lb. pressure passing through two or four |-in. pipes thrust downward into the metal through the side doors. The slags rising to the surface are skimmed, and oxidation is continued until the stage of set copper (Cu with about 6 per cent. CuaO) is reached. In is impor- tant that the dip sample (button) show a smooth concave surface, and the frac- ture a single bubble under the depression. The CU2O is now reduced by poling with poles of hard (poplar) wood 18-24 ft. long and 6-8 in. in diameter at the butt end, thrust into the copper through the front door. When the poling is well advanced, the surface of all or only of part of the metal is covered with char- coal to prevent oxidation. Poling is continued until the hemispherical button- sample, 2 in. in diameter and f in. thick, shows a level, smooth or wrinkled surface, and the fracture a rose color and a silky luster. From 8 to 12 poles are used up in the 15-2 hr. it takes to pole a charge. The copper is now cast by dipping with small hand-ladles, or with large swinging (bull-) ladles, or by a casting machine (endless chain, or Walker) through a trolley- or dipping-ladle, into ingots (20 lb.), bars (60^90 lb.), round or square cakes (loo-iooo-j- lb.), anodes (250+ lb.). Electric conductivity tests are made on ingot, bar, and cake copper. The furnace is fettled after every cast; it has to be repaired about every eight weeks; the roof lasts about eight months, the fire-box about four months. Slime concentrate which is too fine, and usually too impure, to go into the reverberatory furnace with coarse ore, is charged into a special reverberatory furnace with cupola copper, refined and cast into anodes, or it is agglomerated in a reverberatory furnace, or briquetted. The agglomeration is carried on on a bath of cupola copper. The fine material is piled on the bath from the hopper in the roof and melted down grad- ually. The sides of the conical heap become glazed, the slag formed trickles down and is skimmed as fast as it is formed. When the cone of fines has dimin- ished somewhat in size, new material is fed into the hopper by means of a belt elevator. Thus the process of agglomerating is continuous. The slag contains about 17 per cent. Cu and goes to the blast-furnace; part of the accumulated copper, forming about 25 per cent, of the ore, is tapped at intervals, and then worked with cupola copper. In briquetting, lime is used as a bond. The White briquetting machine' furnishes the well-known small cylindrical briquettes. They are piled on iron cars, placed in a cyUnder, and steamed as is the case in the manufacture of hme- sand brick. ^ The hardness of the briquettes is sufficient to make them suited for blast-furnace work. 185. Blast furnace. — Some data of the blast-furnace work of the Michigan and the Calumet and Hecla smelteries are given in Table 92. 1 Hofman, "General Metallurgy," 1913, p. 638. 2 Eckel, E. C, " Cements, Limes and Plasters," Wiley, New York, igos, pp. 130-147. SMELTING OF COPPER 371 Table. 92 — Blast-furnaces for Smelting Native Copper Ores and By-products Horisontal section at throat Area at throat, sq. ft Horizontal section at tuy feres Area at tuyferes, sq. ft Height, tuy feres to throat Height, tuyferes to top of crucible. Water-jackets, height Bosh, height Bosh, amount inches in ft Crucible, depth Forehearth, fixed or movable Forehearth, shell dimensions Tuyferes, number Tuyferes, diameter Tuyfere-ratio Charge (ore, + flux) weight, lb Charge, tons in 24 hr Charge, tons per sq. ft. hearth area in 24 hr Cu, per cent, of charge Coke, per cent, of charge Coke, per cent, ash Anthracite, per cent, of charge Anthracite, per cent, ash Blast, cu. ft. per min Blast, pressure oz Cooling water for jackets, gal. per hr Granulating water, gal. per ton-slag Men, number in 24 hr., a (regular crew) + b (accessory labor). Slag, SiOj Fe(Mn)0 Ca(Mg)0 AhOa. . . : Cu Specific gravity Michigan Smelting Co. 4' 9"Xi2' S7 3'4"Xl2' 39.9 7' 6" 0' 7" 6' loi" 6' loj" 9-S" in 6' lo.s' I' g}" Fixed 4'X6' 18 4" 5.92:1 6000 200 S 20 6 II 6i 14 6000 4 40 17 20 IS 0.9 3-4 Calumet & Hecia Hubbel Buffalo 3'6"Xi2' 40 7' 3" o' 7 . S" 10' I J" S'X7'6" 37 S 2'9"X7' 2" 19.8s 8'o" I' 3" 5' 2" 2' 2" Fixed Two 3X6' and 3' 6" deep 24 753 : I I' o" Two movable 3'XS' and 4' deep 18 4.S" 4.58:1 2780 160 16.6 7 6000 18 9000 30.8 24-3 31.7 12. 1 0.64 3.362 The blast-furnaces, formerly circular or elliptical in horizontal section, are all rectangular and have internal crucibles. The charge consists of reverbera- tory slag, briquettes, siliceous or ferruginous flux, limestone, and a mixture of anthracite and coke. The reverberatory slag- from amygdaloid mineral is acid (Si02 40, FeO 20 per cent.); that from conglomerate basic (Si02 25, FeO 45 per cent.). If the slag belongs to different mining companies, it has to be smelted separately, as the irregularity of the copper-content makes sampling impracti- cable. The normal blast pressure of about 16 oz. has been reduced in some instances to 4 oz. to insure a better reduction through slower smelting, 240 tons of charge vs. 70 tons. The black copper, with 95 -f- per cent. Cu and some S from the fuel, is tapped at fixed intervals; the slag overflows continuously through a trapped spout either into a single large brick forehearth, or into two smaller ones placed in series, from which settled copper is tapped at long inter- vals. The waste slag with 0.75 per cent. Cu is collected in pots or granulated. The cost of smelting in 1906 at the works of the Lake Superior Mining Co., treating 41,176 tons of mineral, was $7,293 per ton of ore.^ ■Austin, Min. Sc. Press, 1909, xcvin, 592. 372 METALLURGY OF COPPER D. Fire-refining or Impure Copper^ 186. Introductory. — In smelting copper ore in the blast- or reverb eratory- furnace, and copper matte in the converter, there are obtained black copper, blister copper, and coarse or converter copper, all of which contain impurities which make the metal unfit for industrial use. These impurities may be Fe, Pb, Zn, Sn, Co, Ni, As, Sb, Bi, S, Se and Te; further Ag, Au, CU2O and gases. The aim of refining is to remove the impurities as much as possible, and to pro- duce a copper of a required purity having the necessary mechanical properties and electric conductivity. The process of fire-refining consists in an oxidizing fusion in order to volatilize some metals and to oxidize and scorify others which have a greater affinity for than has Cu, followed by a reducing fusion in which most of the CU2O formed and held in solution by the copper is reduced to Cu. A small amount of CU2O (0.5 + per cent.) is always left in the Cu in order to in- sure the presence of any remaining impurity in the state of oxide in which it is less harmful than if in that of metal, when it is likely to form a solid solution. 187. Furnace. — Formerly the process was carried out in hearth-furnaces,'' from which the copper was obtained in the form of round discs as so-called "rosette-copper." In recent years the electric furnace has been suggested' for refining, but it has not yet come into practical use. The only furnace that need be considered is the reverberatory furnace. The general- statements re- garding this furnace for the smelting of native copper ore (§ 183) hold good for the refining of copper. The main difference is that of size of furnace, which is caused by the form and character of the material to be treated, and with it that of the manner of charging and discharging. A refining furnace may have to work cakes of low-grade black copper and high-grade blister copper, sheets of pure cathode copper, or liquid converter copper. If the copper be low grade and require a prolonged refining operation,* the furnace will have to be small, probably not exceeding 20 tons capacity; if the copper be nearly pure, the capacity may be very much increased, reaching 100, 150, and even 250 ' Stetefeldt, Berg. Huttenm. Z., 1863, xxii, 185, 205, 2ig. Hampe, Zt. Berg. Hiitlen. Salin. Wesen i. P., 1873, xxi, 218; 1874, xxii, 93. Egleston, Tr. A. I. M. E., i88or8i, rx, 678. Stahl, "Dissertation," 1886; Metallurgie, 1907, iv, 761. Keller, Min. Ind., 1898, vii, 245; Tr. A. I. M. E., 1898, xxvm, 137; rgoo, xxx, 310. Gibb, op. cit., 1903, xxxiii, 661. Hofman-Green-Yerxa, op. cit., 1904, xxxiv, 671. Hofman-Hayden-Hallowell, op. cit., 1907, xxxvin, 171. Wanjukow-Schmidt, Metallurgie, 1909, vi, 749. Johnson, Met. Ckem. Eng., 1911, ix, 396. Emrich, Tr. A. I. M. E., 1912, XLm, 446. Peters, E. D., "Practice of Copper Smeltery," McGraw-Hill Book Co., New York, 1911, PP- 531-547- ^ Percy, J., "Metallurgy," Murray, London, 1861, pp. 399, 406. Schnabel, C- Louis, H., "Handbook of Metallurgy, Macmillan, New York, 1905, i, 248. ' Rauschenplatt, Metallurgie, igro, vii, 151, 435 (Borchers). ^ Platten, /. Soc. Chem. Ind., 1906, xxv, 449. SMELTING OF COPPER 373 tons.' The reduction in fuel-consumption by this increase in capacity is shown by the curves of Keller^ and Peters' given in Fig. 396. The curve of Peters shows that the saving in fuel is small when the charge is greater than 50 tons. The bottom of the furnace and the manner of its support, as well as the con- struction of the sides of the furnace, have received much consideration with the increase of the weight of the charges. The bottom used to be built up exclu- sively of sand burnt in, as is still the common practice with matting furnaces (§139). In a few instances silica-brick have replaced the sand-bottom, as has been done with some matting furnaces and in the furnace of the Michigan smelter. Fig. 392. With some furnaces the cooling vault, given up in matting furnaces, has been retained; with others the hearth has been built on cast-iron plates supported by brick pillars. The reason for air-cooUng the bottom* is 400 ■aSg'aoo t^. 1 i \ \ /> ^ ^ V '■s > ■« ■ — ~ T 100,000 200,000 300,000 Pounds of Copper Refined per Charge 400,000 Fig. 396. — Relation between coal consumption and size of refining charge. that in melting large charges of metallic copper there is danger of the hearth becoming too hot and breaking out; hence the additional cooling with increase of weight of charge. With furnaces holding 20oand 250 tons metaP there has arisen the difficulty of making the bottom sufficiently strong so that it will not come up. This has been overcome by a forced circulation of air through a series of 2j- or 3-in. pipes, running crosswise underneath the brickwork carrying the sand- bottom, and joined at the ends to two longitudinal pipes, one of which is con- nected with a fan, while the other is extended into the ashpit of the fire-place. With a 225-ton furnace 100 pipes receive 5000 cu. ft. air per minute. This forced ventilation with a foundation built up solid is believed to be preferable to natural ventilation with a hearth resting upon cast-iron plates supported by brick pillars. 'Addicks, Eng. Min. J., 1907, lxxxiii, 1002. Prosser, op. cit., 1907, Lxxxrv, 171. Walker, Min. Ind., 1910, xix, 221. ' Min. Ind., 1898, vn, 250. 'Practice of Copper Smelting, 191 1, p. 567. *Eng. Min. J., 1913, xcv, 576 (Herreshoff), 816 (Cloud), 864 (Nason). 'Walker, Min. Ind., 1910, xix, 221. 374 METALLURGY OF COPPER The inner sides of a furnace have been recently constructed in part of mag- nesite brick instead of silica-brick throughout as before, as they appear to stand better the corrosive effect of the charge. The magnesite brick reach from below the silica bottom to a short distance above the highest level of the metal bath. The roof is always made of silica-brick. The contact planes between mag- nesite and silica-brick are formed by a course or two of neutral chrome brick. The latest innovation is that of Addicks and Brower^ who have in operation at Chrome, N. J., a refining furnace with bottom of magnesite brick, and sides and roof of chrome brick. The hearth of a refining furnace has to be seasoned with copper as has that of a matting furnace with matte (§ 139). The bottom is covered with a few inches of copper; this is melted slowly and gradually absorbed. The tempera- ture is slightly lowered, and a second portion of metal charged, melted, and absorbed. The operations are repeated until all the copper that will be absorbed has been taken up. Some of the Cu will be oxidized to CU2O and slagged by the SiOa; the slag formed acting as a cementing material strengthens the bottom. A sand-bottom absorbs more copper than does one built of silica-brick; a furnace resting upon plates locks up less copper than one supported by a sohd or even a vaulted bottom. It is essential that the copper used for seasoning be of a high grade so as not to "poison" the metal to be refined. An interesting case of such poisoning was that of Baltimore where Te was the harmful metal.^ A small fxurnace will be charged by hand and the copper cast with hand- ladles, a large furnace requires mechanical charging and molding by means of a casting machine. The air necessary for fining may be admitted through ports in the fire- bridge, or blown in through tuyere-pipes placed on either side of the fire-bridge. The fuel may be bituminous coal, oil residuum, or producer gas. The waste heat of the furnace may be recovered in part by recuperation or regenera- tion; in recent years the gases of most furnaces have been conducted through waste-heat boilers. All operations are usually so conducted that it takes 24 hr. to work a charge. The furnaces of the Anaconda Copper Mining Co. at Great Falls, Mont., Figs. 397-399, the Balbach S. & R. Co., Newark, N. J., Figs. 400-404, and the United States Metals Refining Co., Figs. 405-410, may serve as examples of modern refining furnaces. Their leading dimensions and then: work as well as similar data of other furnaces are assembled in Table 93. Descriptions of some refining plants given in the technical literature are those of the First Raritan Works,' United States Metals Refining Co.,* Rio Tinto works, ^ and the Second Raritan Works.* ^Eng. Min. J., 1914, xcvii, 421. " Egleston, Tr. A. I. M. E., 1882, x, 493. ' Addicks, Min. Ind., 1900, rx, 261. * Addicks, Min. Ind., 1906, xv, 301; Eng. Min. J., 1907, Lxxxra, looi. ' Walker, Eng. Min. J., 1907, lxxxiv, hi. * Easterbrooks, Electrochem, Met. Ind., 1908, vi, 245. SMELTING OF COPPER 375 376 METALLURGY OF COPPER SMELTING OF COPPER 377 The furnace of the Anaconda Copper Mining Co., Great Falls, Mont., shown in Figs. 397-399, serves for the refining of cathodes.' Its capacity is 35 tons of cathodes or 62 tons of fluid converter metal. The hearth, 23 ft. 6^ in.XiS ft., has the pear-shaped form characteristic of the original Welsh furnace. It has the form of an inverted arch and rests on the center line of a segment arch covering a central cooling chamber. The dish-shaped sand-bot- tom slopes 9 in. from fire-bridge to flue, where is situated the depression necessary to receive the suspended bull-ladle, 14-19 in. in diameter., for casting ingots, cakes, or wire-bars. There is a single working door on either side, and a rab- bling and ladlmg door beneath the inclmed flue; the roof, of silica-brick, is horizontal for half the length of the hearth to allow for free development of flame, and then only pitches downward toward the flue. Air for oxidation is admitted through ports in the roof and the fire-bridge. The grate-area is large when compared with that of the hearth area, viz., i :3.7; the cast-iron grates are provided with mechanical shakers; the coal is fed through two doors placed on either side. The furnace of the Balbach Smelting & Refining Co., Newark, N. J., shown in Figs, 400-404, also serves for the refining of cathodes. Its capacity is 80 tons of copper. The hearth, 23 ft. 3 in. X 13 ft. 10 in., has straight sides and curved ends; it is carried by heavy ribbed cast-iron plates supported by brick pillars; the open space is cooled by forced draft. The sand-bottom slopes from fire- and flue-bridges toward the tap-hole placed in the side toward the flue-end. There are two working doors on the side opposite the tap-hole, and a rabbling door at the flue-end. The roof of silica-brick, has an expansion joint i\ in. wide. The grate area is relatively small, presupposing the use of a good grade of coal; the grate bars are wrought iron; there is a single fire-door. The furnace of the United States Metals Refining Co., Chrome, N. J., shown in Figs. 405-410, also a cathode furnace, has a capacity of 250 toris of copper. The hearth, 33 ft. 8 in.Xi4 ft. 10 in., has straight sides; the fire- bridge end is not curved; the flue-end is contracted as usual. The hearth con- sists of a full course of fire-brick on which is burnt in a mixture of sand and rolling-mill scale to serve as working-bottom; the brick rest on cast-iron plates which are supported by red-brick pillars. The hearth slopes from fire- and flue- bridges to the tap-hole placed in the side toward the flue-end. The sides are built of silica- and magnesite brick, the two are separated by a course of neutral chrome brick. There are three charging doors in the side opposite the tap- hole, and the usual rabbling and skimming door at the flue-end. The long horizontal part of the silica-brick roof is separated by one expansion joint, i| in. wide, from the short pitching part leading to the flue, and by another, also \\ in. wide, from the roof of the fire-box. The wrought-iron grate bars, ij in. square, have a 3-in. pitch; there are two fire-doors on one side. Peters^ describes a gas-fired refining^ furnace with regenerators having a 'Hofman, Tr. A. I. M. E., 1904, xxxrv, 313. ' "Modern Copper Smelting," 1895, p. 573. 378 METALLURGY OF COPPER Section C-B Fig. 405. — Copper refining furnace, United States Metals Refining Co., Chrome, N. J. Section G-H Fig. 406. — Copper refining furnace, United States Metals Refining Co., Chrome, N. J. SMELTING OF COPPER 379 ;0:,?:v»--^;5i--i>;-?i;-^*^>/:.<'-i.:o-^-".:tf--d-S 5 M d 0* O c a co%i- Oi t-t ^■5 •O d 3 +3- ■ op -:> So d : rid a ^ a d ^ dCO d p< O O O CO o w to'" J! (u=3'" Slf H H " '^ O O O " . 0} -fi-i o.-S'S +J -M +» t N tH (U U EU »G *J +J t3 *" - -p 0) aTrS (u P. qJ o' ?I rt^ rt o rt ti u ih'" 1-. '-' tH d 00 o ocKi «■" ho 4 -a ~TI PSciJ P=i m o ^ J E 0. J3 SMELTING OF COPPER ,389 I I I I PI ^t1 M .1 + iS-d S O O •ON Psa ■o a C OO C O > U <; ^ w ft n 11 s ft ^ S ■a fti3 t fto +3 ft aj o ^ "! ;»-«3 N o o S ij a o O E«>8 ■S 2 CI ■^ ("J ho "3 " • ^0 ?ft 2 3 ft 03 ;3 .fei-3 « 3'" S "■a ftg •a o S V bo**- 4.5 ° CO : I ■S o ti 13 •" •V-if fo 01 ft r 3 ■ ■ ^^^ O O «■;_ 02.1 tOcC JO rt 3 O" •^ -3 ?! e B u n1 A ■^ U C ;5 N t"- '^ o y y tn S g u 3 C 1 ,02 4J O Sa ma Kill "si 0,^0 art «S6|g C-3U rt ov M^ B" ! fc es T^^ ^1 " So ■:: ^■■ > u d 8te" 13 bO II QO d fi" t- ^ -^■g 33 *^ ■ M t; iH 1/3 u rt rt tn » ' a it o o 00 O a Sod a)w ri 3 u fe<00 g. SMELTING OF COPPER 391 193. Examples of Refining.— Three graphic representations of the changes taking place in refining will furnish details of the chemical and physical changes outlined in the preceding discussion. Wanjukow' studied by means of chemical analyses the changes that took place in refinmg a charge of 4450 lb. of impure black copper CU94.55, Ag 0.0021, Pb 0.0123, Fe 3.0373, Co 0.8944, Ni 0.4080, P 0.0105, As 0.1257, Sb 0.0020, VIII IX X XI xu Fig. 411. — Elimination of impurities in refining black copper (Wanjukow). 0-09S3. S 0.8678, SO2 0.0006, Insol. 0.0004; total 100 • 0064 per cent, in a reverb- eratory furnace provided with two tuyeres and heated with producer gas gener- ated from twigs, brush-wood and pine-needles. The copper was brought to tough-pitch by two consecutive refinings. The results are drawn in Fig. 411. The abscissa gives the time and the different stages of the process. Thus, gas and air are turned on at 2.22 o'clock, changing is begun at 2.23 and finished at ' Melallurgie, 1909, vi, 749, 792. 392 METALLURGY OF COPPER 3.30. At 4.30 the charge is melted; at 4.40 the blast is turned on; boiling begins at 5.50 and reaches its maximum at 6.25-6.30. At 7.00 boiling ceases and pohng is begun, which is a dense-poling. At 7.38, the blast and other access of air are shut off, whereby dense-poling changes into tough-poling. At 8.15 the copper is at tough-pitch. The second oxidation is started by turning on the blast; at 8.30 the second dense-poling begins, lasts until 8.50, when the blast is shut off and thereby again dense-poling changed into tough-poling. This continues until 9.30, when P is added. At 9.40 ladling the refined copper is started. The ordinate represents percentages on the basis of 100 parts of Cu (which changes the analysis above to Cu 100, Ag 0.0022, Pb 0.0130, Fe 3.2124, Co 0.9455 Ni 0.4312, P o.oiii, As 0.1328, Sb 0.0215, O 0.1009, S 0.9173) and gives the relative eliminations of impurities. A different scale is employed for certain impurities as explained in the legend in Fig. 411. The lines covering the period of charging to complete fusion, 2.23 to 5.00 o'clock, are dotted, as their positions have not been determined; they indicate the probable rates of oxidation. The refined copper gave upon analysis the figures shown in Table 96. Table 96, —Analysis of Refined Copper Cu Ag 1 Pb 1 Fe 1 Co 1 Ni 1 P | As- | -Sb | | S302 | S total | Total 99.66 100.00 0.0016 0.0016 O.OOIO O.OOIO 0.0080 0.0080 0.0243 0.0244 0. 1091 0.I09S 0.0012 0.0012 0.0402 . 0403 0.0017 0.0017 0,00486 0.00487 tr. 0.1093 0. 1093 99-9051 The curves show the following: Fe is oxidized during the melting-period and quickly scorified as soon as the copper has become liquefied, falling to 0.0080 per 100 parts Cu at the beginning of the boiling period; from then on the ehmination progresses more slowly. Co. — The rise in the Co-cilrve during the melting period would indicate that Co was not oxidized; the contrary is true, but CoO is unstable at a red heat. The rapid drop of the curve after fusion shows that Co is scorified nearly as fast as is Fe; at the beginning of the boiling, the elimination is weaker; at the strongest boiling, it reaches 0.024 per 100 Cu and then practically ceases. Ni. — This metal is difficult to slag; the refined metal retains as much as o.ii per 100 Cu. At first, oxidation and scorification progress very slowly; after most of the Fe and Co have been slagged (before the boiling-stage) the elimina- tion progresses more rapidly, then slows up when boiling is in full progress, and finally ceases. The first refining operation has taken out all the Ni that can be removed, so repeating the process does little good. 5. — Some S is oxidized in melting the charge; the elimination then proceeds more quickly but still very slowly until most of the Fe has been scorified, when with the lowering of the temperature from 1109 to 1091° C. the boiling period sets in, and SO2 is set free rapidly bubbUng up through the copper (copper rain), and the S-content drops from 0.7189 to 0.0911 per 100 Cu; from now on the evolution of gas goes on more slowly reaching 0.0379 in stage IV, 0.0256 in stage V (dense-poling), and 0.0097 in stage VI, when the S-content changes little. SMELTING OF COPPER 393 0. — The metal takes up O during .the melting; the 0-content rises quickly during the fining period and especially so after the Fe and part of the Co and Ni have been scorified; boiling assists the formation of CU2O. With the beginning of the poling, the percentage of O falls quickly and regularly; it increases with the second oxidation and decreases agaifi with the second poling. Ph. — This metal is scorified during both oxidizing periods, more quickly in the first, when there is more Pb present, than in the second; little Pb is driven off after this. Ag. — A small amount of Ag enters the slag with the Pb. Sh. — This metal, difiicult to eliminate, appears to be slagged to a greater extent during the melting than the finmg period; the boiling of the copper is favorable to oxidation. As. — The behavior of As is similar to that of Sb; a large part is oxidized in melting, fining favors elimination, poling has no effect. The second refining stage appears favorable to the scorification of both Sb and As. The second example of refining impure copper is that of Stahl^ who treated at Mansfeld, Germany, in the usual course of work two lo-ton charges of blister copper in about 21 hr. at temperatures ranging from 1200 to 1450° C, and took samples for chemical analysis at the end of each stage of the process. The re- sults of one of his tests are given in Table 97 and plotted in Fig. 412. Table 97. — Stahl, Elimination of Impdrities in Refining Blister Copper Blister copper After melting After fining After dense- poling .\fter tough- poling Per Per Per Per Per Per Per Per Per. Per cent. 100 Cu cent. 100 Cu cent. 100 Cu cent. 100 Cu cent. 100 Cu Cu 98 . 140 100 . 000 98.950 100.000 98.550 100. oco 99.060 100.000 99 ■ 300 100.000 Ag O.OII O.OII O.OIO O.OIO O.OIO O.OIO 0.009 0.009 0.009 0.009 Pb . . . , 1.060 1.080 0-330 0.330 0. 160 0. 160 0. 190 0. 190 0. 170 0.017 Ni 0.610 0.620 0.350 0.350 0.350 0.350 0.360 0.360 0.360 0.360 As 0.058 0.059 0.058 0.059 0.05s 0.056 0.057 0.057 0.054 0.054 0.121 0.123 0.302 0.306 0.875 0.888 0.324 0.329 0. 107 0. 108 Total. . 100 . 000 100.000 100.000 100.000 100 . 000 In refining impure Cu the elimination of impurities does not proceed in a fixed order or at a given rate, but is governed rather by the amount of impurity and the form in which it is present, by the lining of the reverberatory furnace (acid or basic), the size of the charge, the admission of air, the temperature and other details of the mode of operating. The work of Wanjukow and Stahl shows, whatever may be the variations as to detail, in general: (i) that Zn, Fe, Co, Sn are removed completely at the beginning of the finmg period, and S at end of dense-poling; (2) that the elimination of Ni, Pb, As, and Sb continues through the entire process and is imperfect; and, (3) that Ag and Bi are re- ^ MelMurgie, 1912, ix, 362, 377. 394 METALLURGY OF COPPER moved .only to a very small extent, the former mainly by volatilization, the latter by scorification. Hofman-Hayden-HallowelP followed the chemical and physical changes that took place in cathode copper as it passed through a loo-ton refining furnace to be cast into wire-bar. The results' are given in Fig. 413. Sample No. i was taken after melting and skimming, No. 2 after fining for 6 hr. with compressed air when the stage of set copper was reached; samples Nos. 3 to II at is-min. intervals during tough-poling; No. 11 represents tough- pitch copper; Nos. 12 and 13 overpoled copper. 1.3 - 1.2 1.1 ■s 1.0 \ ^N 0.9 |0.8 . -<*d S 0-' - "n ^-0.6 ~r-^. . A. \ Aa As -P -^^ \ « — °~~°~>.o ^ 0.5 "^^-..^ \ § "--~-^i^\ «0.4 ■ "^g^ — Ni Ni 0.3 - • — -.^^^ 0.2 . Pb 0.1 Ag Ag '~~ T^*^^ Time, Hours 0.0 1 > t 1 — 1 Uo i 1 oJ. 2 3 4 6 6 7 Melting: 9 10 11 12 13 14 15 16 17 18 I Fining: 19| 20 21 Fig. 412. — Elimination of impurities from blister copper (Stahl). On the whole, the chemical changes up to the tough-pitch stage are what would be expected; the physical properties are in harmony with the rise and fall of the Cu-content. The percentage of FeO, usually lower than that of S, is higher; it was reduced to 0.086 per cent, by 6 hr. fining and reached the minimum of 0.022 per cent, orjy after poling for 15 min., the probable reason being that some Fe was taken up from the iron pipe conveying compressed air into the copper, and removed only after the pipe had been withdrawn. Practically no S was driven off; the 0.030 per cent. S, high for electrolytic copper, remained about constant. The specific-gravity and electric-conductivity curves show the same general trend as does that of the Cu-content. The tensile strength shows a gradual decrease as the poling progresses, and the elongation a corre- sponding increase. Before taking samples Nos. 1 2 and 13, the bulk of the copper in the furnace was cast, and the remaining small amount overpoled until it threw a worm. No regularity can be expected from these two samples. * Tr. A. I. M. E., 1907, XXXVIII, 171. SMELTING OF COPPER 39S Clevenger' gives the following temperatures as averages of a number of refin- ing furnaces: Charge melted and ready to rabble 1 141° C. ; after 25 min. rabbling 8.900 ^2 8.600 I 2 8.400 OQ O 8.100 Befining Charee No.l Number of Sample B 6 7 8 7 8 Number of Sample FtG. 413. — Changes in refining cathode copper. 1103°; after 75 min. 1103°; at end of rabbling 1103°; after 20 min. poling 1107°; before ladling 1125°, after ladling 20 liiin. 1121° C. ' 194. Casting. — Whenthecopper has "come to nature "or arrived at the tough- ' Met. Chem. Eng., 1913, xi, 448. 30 METALLURGY OF COPPER pitch stage, it is ready to be cast into marketable forms. These are ingots, ingot-bars, wire-bars and cakes. The ordinary three-heel ingot is shown in Fig. 414-415; it is usually 3X3X9 in. and weighs about 17 lb. The weight ranges from 10 to 22 lb., e.g., Raritan 10-16 lb.; Phelps-Dodge & Co. and American Smelting & Refining Co. 18-20; Calumet & Hecla 20 lb.; ingots in a ship- ment vary 2 and 3 lb. in weight. For convenience in shipping the copper is cast sometimes in the form of ingot-bar, Fig. 416, 3X3X27 in., which is readily broken at the two deep webs into ingots of ordinary size. Beside the three- heel form, there are molded one-heel, four-heel ingots, etc.^ F;g.419 •i +i Ifl 00 SNa •s -g^ 3 -o a Q) 111 I W ft ft ^ g, M (*3 \0 « (^ -tt M 3 tM % H 'y^ ■" E? "^ W N N N % h-- S 0000 c3. d _r fl 10 m 1/) 10 N CH CS N H H M W - +i aJ g a -* o node rrent siency, cent. < gifi s •C d " d Catho curre efficien per ce w Oi Oi a ulphuric acid reduced, er cent. in \ m ft P. Copper de- posited, grams « w fO H M >-■ M . . ength experi ment hours Tj- N "^ -tt Tf 1 J Mean po- tential, volts -0 ■^ 00 fO Oi (N (N M 01 M Mean urrent rength, amp. m m >o >o ifl " -s '^■§g i> M sO Ol s a s H^a a sJ S g E ^ S ■* HH in N a kJ po -^ ■* m to h J3 p. n 0. ■0 ■C d" S li;-'' Si 00000 •n a •a " tj 0. 6 y B fe fe ft 10 m m 10 m N N « N IN t^ t^ t* m t^ s. =■ d 1 B " 10 lo m in m el- < , H N ro ■^ w Q ft a ^ g 6 4i8 METALLURGY OF COFFER show that of the appearing at the anode, from 60 to 65 per cent, can be util- ized in the oxidation of SO2; that a cathode efficiency of 95 per cent, can be obtained without the use of a diaphragm from a ferruginous copper sulphate solution, with the recovery of a high quaUty of copper. Experiments carried out with pure CUSO4 showed that the cathode efficiency remained about the same, while the anode efficiency was reduced 5 per cent. Table 104. — Experiments showing Effect of SO2 No. of experi- ment Composition of electrolyte Copper deposited, grams Length of experiment, hours Mean cur- rent strength, amps. Mean potential, volts Current efficiency, per cent. Cost, Copper, per cent. Ferrous iron, per cent. Ferric iron, per cent. cents per lb. copper deposited I 2 3 2.83 2.60 2.54 I. 25 1. 01 1. 16 0.12 0.36 0. 21 91 36 75 24 24 15 S S 5 2.00 2.00 1.90 70 26 90 2.393 6.054 1.720 To the original tables of Reinartz have been added costs, cents per pound Cu deposited as figured by Austin. Thus in Table 102, experiment No. i, the average current was 5 amp. at 3.21 volt; 5X3-21 = 16.05 watt; divided by 1000 gives 0.01605 kw. This deposited 126 g. =4.44 oz. Cu in 24 hr., hence required 0.01605X24 = 0.3852 kw.-hr., which is equivalent to 1.3872 kw., hr. for I lb. Cu. Taking the kilowatt-hour at 2 cents makes the cost of de- positing I lb. Cu 2.774 cents as given in the table. 207. The Neill and Van Arsdale Processes.' — In this process SO2 generated by roasting metallic sulphide in a kiln is forced through a series of tanks hold- ing a watery pulp of oxide ore in order to form and hold in solution CuSOs; this salt is unstable, and changes when in contact with CuO into (CU2SO3-I- CuS03)-l-CuS04. Upon heating the solution, from 66 to 75 per cent, of the Cu is precipitated as CU2SO3-I-CUSO3 with the formation of some CUSO4. The remaining Cu has to be recovered by other means, e.g., by precipitation with Fe. The process at Caconino, Ariz., was not successful for mechanical reasons, but tests at the works of the Montana Ore Purchasing Co., Butte, with ore roasted in MacDougall furnaces, gave a total yield of go per cent. Cu with a direct extraction of 64 per cent, as CU2SO3.CUSO3. In the Van Arsdale process, SO2 is used as a reducing and precipitating agent. Oxide ore is leached with H2SO4; the solution is saturated with SO2 when the reaction 3CuS04+3S02+4H20 = Cu2S03.CuS03-|-4H2S04 takes place. The solution placed in a closed vessel and heated to 100° C. liberates gas which generates a pressure of 30 lb. causing the reaction Cu2S03.CuS03-|-4H2S04 = 1 Neill, Circular 1901, Min. Rep., 1902, XLVi, 395, 464; Eng. Min. J., 1908, Lxxxv, 556; Min. Metk., 1910, I, 389. Jennings, /. Can. Min. Inst., 1901, rv, 123; Eng. Min. J., 1901, txxi, 400; 1908, Lxxxv, 152, 821; Lxxxvi, 418 (Westly-Sorensen process). Van Arsdale, Eng. Min. J., 1903, Lxxv, 853 (Van Arsdale process). Gin, Zt. Electrochemie, 1903, rx, 857. Burfeind Min. Rep., 1902, XLVi, 206, 464. Zwaluwenburg, Min. Metk, igio, I, 376. Austin, op. cit., 1911, 11, 241. LEACHING OF COPPER 419 CU+2CUSO4+2H2SO4+2SO2+2H2O to take place; 40-50 per cent, of the Cu falls out as metal from an original solution with 10 per cent. Cu. Releasing the pressure causes most of the SO2 to pass off, so that the regenerated solvent can be used again. According to Gin' the temperature has to be 180° C. and the pressure 142 lb. The process has not been tested on a large scale. 208. Stadtberge Process. Leaching with HCl.^ — With the change of the ore from oxide to sulphide, the process has been replaced at Stadtberge by the Doetsch process (§ 221). Oxide ore with about 2 per cent. Cu, crushed to i in., used to be leached for 3 days with HCl of 12-13° Be in vats holding 75 tons (depth of charge 3.3 ft.) until the acid was neutralized and thereby enriched to 19-20° Be. From the liquor, the Ag with part of the Cu was precipitated with Fe, and, later on, the remainder of the Cu. The leached ore, moist with HCl, was piled in heaps and allowed to weather for 12-15 weeks; decomposition was assisted by wetting at intervals with mother liquor from the precipitating vats. In this way 75 per cent, of the Cu was recovered. The twice-leached ore was removed to the dump to undergo further alteration by weathering. The drainage from the dump was collected and treated with Fe. An additional 17 or 18 per cent. Cu was thus recovered, making the entire yield of Cu 92-93 per cent. Other examples of leaching with HCl are that of Rochlitz,^ and the tests of Stahl.* 209. The Greenawalt Process.' — Oxide or roasted sulphide copper ore is leached with HCl dissolved in brine; the CUCI2 formed is reduced to CU2CI2 by SOa-gas, viz., 2CuCl2+S02+2H20 = Cu2Cl2+2HCl+H2S04, and some HCl formed by H2S04+2NaCl = 2HCl-|-Na2S04; the solution of CU2CI2 in HCl and NaCl is electrolyzed with Acheson graphite electrodes and at the same time SO2 pumped into the vat, whereby the CI set free is converted into HCl, the reactions being Cu2Cl2+current = Cu2+Cl2 and Cl2-|-S02+2H20 = 2HCl-|-H2S04, also H2S04+2NaCl = 2HCl+Na2S04. It is thus seen that the raw materials con- sumed are salt (| lb. NaCl for i lb. Cu), and SO2. Impurities, such as Bi, As, and Sb, are to be precipitated by H2S from decopperized liquor. Iron going into solution as FeCl2 is precipitated as Fe203 by CuO according to the Hunt- Douglas reaction FeCl2-|-3CuO = Fe203+CuCl2-|-Cu2Cl2. For a complete elim- ination of base metal, NaCl is electrolyzed, giving CI and NaOH (2 kw.-hr. furnishing i lb. CI and 1.5 lb. NaOH), the CI is conducted with SO2 into brine: 2Cl-|-S02-|-2H20-|-2NaCl = 4HCl-l-Na2S04, and the NaOH serves to precipi- 'V Internat. Congress Angewandte Chemie, 1903, 11, 116. ^Miszke, Oest Zt. Berg. Huitenw., 1871, xix, 108. Gerhardt, Zt. Verein. deutsck. Ing., 1872, xvi, 305. Francke, Metallurgie, 1910, vii, 484. Mengler, op. cit., igii, viii, 176. 'Meyer, Berg. Hiittenm. Z., 1862, xxi, 173, 201. ' Op. cit., 1894, LHi, 65. " Greenawalt, Eng. Min. J., 1910, xc, 1064; Austin, Min. Melh., 1911, n, 339; Greenawalt, Hydrometallurgy of Copper," 1912, p. 349. 420 METALLURGY OF COPPER tate the base metals. Accumulation of Na2S04 is said to have no deleterious effect. At the experimental plant in Denver, Colo., with a variation in current density of 6 . 2 to 66 amp. per square foot cathode area, i lb. Cu was deposited per kilowatt- hour. On a basis of $50 per kilowatt year and i lb. Cu per kilowatt- hour the cost of electro-deposition of i lb. Cu is estimated to be 5.8 cents. The cost of a 100-ton plant for 5-per cent. Cu ore is given as $120,000 to which, with sulphide ore, $30,000 should be added for a roasting plant. The cost of treatment per ton oxide ore is estimated at $1.96, and $2.71 per ton sulphide ore, if $0.75 is the cost of roasting. 210. Leaching Atacamite. — This chloride mineral cannot be smelted direct, as about 50 per cent, of the Cu is lost by volatilization, hence wet processes have to be used. According to Argandofia,' heating the ore in a cast-iron re- tort to 230° C. and conducting over it steam at 2 lb. pressure decomposes - the mineral: CuCl2-f-H20 = CuO-t-2HCl. The CuO is smelted and low-grade ore leached with the condensed HCl. Evans^ proposes solution with HCl and precipitation with H2S. Thompson-McGovern^ discuss the leaching and electro-deposition with an anode of fused magnetite, in Chile, of bronchantite, CuS04.3Cu(OH)2, asso- ciated with the NaCl and NaNOs in a sUiceous decomposed granite.* 211. Leaching Sulphide Ore after Conversion into Sulphate by Weathering. Pyritic ore which is readily disintegrated by atmospheric agencies, such as free-burning pyrite and especially marcasite, and which carries disseminated copper mineral readily sulphatized, such as chalcocite, is suited for weathering; covellite is slowly sulphatized, and chalcopyrite hardly at all, unless it is first converted into chalcocite,* which is a very slow process. But even under the most favorable conditions the process is lengthy, requiring years for a satis- factory extraction. The pyritic material while disintegrated in the process is only slightly altered and can be utilized for the production of SO2 in the manufacture of H2SO4. The leading chemical reactions taking place in weathering have been variously formulated." The simplest is that of Jones: FeS2-|-70-hH20 = FeS04-hH2S04; 2FeS04-FH2S04+0 = Fe2.(S04)3+H20; Fe2.(S04)3+Cu2S = 2FeS04 + CUSO4 + CuS; Fe2.(S04)3 + CuS + H2O -|- 3O = 2FeS04 -|- CuS04-f- H2SO4. The H2SO4 set free would further decompose CU2S and CuS. Winchell suggests that inasmuch as CU2S is more readily oxidized than FeSj, its solution may precede that of FeS2 and hasten the oxidation of the latter. The sulpha- - Min. J., 1906, LXXK, 854; .Ereg. Min. J., 1906, LXXXii, 205. " Min. /., 1908, Lxxxiv, 453. ^ Eng. Min. J., I9i3,xcv, 171. * See also: Editor, Min. Sc. Press, 1913, cvi, 933. ' Finlayson, Tr. Inst. Min. Met., 1910-11, xx, 70. * Emmens, Eng. Min. J., 1892, Lrv, 582. Probert, Econ. Geol., 1903, Lxxvi, 958. , Stokes, oj). cit., 1907,. il, 14, 290 (Winchell). Jones, Tr. A.I.M.E., 1905, xxxv, 3. LEACHING OF COPPER 421 tization of CuSi is accompanied by the conversion of sulphides of Ni, Pb and Zn into sulphates and the oxidation and solution of As, Sb, and Bi. From the solution Fe will precipitate Cu first and later on part of the As, Sb, and Bi. Table 105 gives some data by Gibbs^ upon the elimination of these three elements. Table 105. — Elimination of As.Sb and Bi by Weathering Pyrite Precipitate Total eli- Metal % actual % relative, Cu=ioo % actual % relative Cu=ioo mination % Cu 2.00 0.40 0.030 0.016 100. 20.0 1-5 0.8 62.0 35 0.08 0.06 100. 564 0. 129 0.097 As Sb Bi 71.8 91.4 87.9 212. Leaching at Rio Tinto.— The leading example is that of Rio Tinto, Spain,^ where weathering has superseded all other processes, especially since heap-roasting has been forbidden since 1888. At San Domingo, Portugal, the law went into effect in 1878. The ore is a massive pyrite (marcasite) and an impregnated schist with from i to 3, average 2.5 per cent. Cu, and 45-48 per cent. S. The analyses^ given in Table 106 show the general character of the ores exported and treated locally. Other analyses are published by Vogt.* Table 106. — Ore or Rio Tinto, Spain Ores Massive pyrite Schistose pyrite Export ore ■Local treatment ore Local treatment ore Fe 43-1 495 2.5-3° °-3 0.8 0.4 0-5 0.6 1 . 2 42.8 48.3 1.5-2.0 1-3 1.8 0.7 - 0.8 0.7 1. 1 S 0-15 I. 0-1.25 Cu Pb Zn 0.05 As HiO 3 -IS 93-6 SiOj and various metals. . Total 99-4 99 S 98.3 ' Tr. A. I. M. E., 1903, xxxiii, 667. . 'Deumifi, Bull. Soc. Ind. Min., 1887, i, 843; Launay, Ann. Min., 1889, xvi, 491; Berg. Batknm. Z., iSgo, xlix, 229, with cross ref.; Eng. Min. J., 1890, l, 741; Brown, /. Soc. Chem Ind., 1894, xill, 472; Dingl. Polyt. J., 1894, ccxciv, 48; Berg. Hiittenm. Z., 1895, Liv, 8: Courtney, Proc. Inst. Civ. Eng., 1896, cxxv, 136; Adcock, Min. Ind., igoo, rx, 235; Chalon ^«i. Vn. Min., 1902, lvii, 205; Jones, Tr. A. I. M. E., 1905, xxxv, 3; Truchot, Sixth Inlernat. Congress Appl. Chem., 1906, 11, 170; Bull. Soc, Ind. Min., 1908, ix, 68; Min. Ind., 1906, xv, 288, his treatise, p. 163; Probert, Min. Sc. Press, 1908, xcvi, 27; Corresp., Eng. Min. J. 1910, LXXXIX, 748. 'Eng. Mini J., 1907. lxxix, 371. *Zt. prakt. Geologie, 1894, n, 44; Zt. Berg. Hiitten. Salin. Wesen. i. P., 1898, XLVi, 225. 422 METALLURGY OF COPPER The ore is crushed to pass a 3-in. ring and sorted into coarse and fine in the ratio of 4 :s with 1.25 and 2.25 per cent. Cu. The site for a weathering heap is a slightly sloping ground. On this, stone flues are erected, each 12 in. in diameter, to serve as air-inlets and draining-channels; on the junctions of the flues are erected, as the heap is being formed, chimneys 80 ft. apart, also of rough stone. The heap is now formed of alternate layers of coarse and fine ore, the start being made with side-dump cars on the upper part of the site. The top layer consists of fines in order to assist in the distribution of water. The heap, which has a horizontal surface, is 33-40 ft. high and holds 100,000 tons of ore. On the surface are formed squares separated by ridges to regulate the flow of water admitted through gutters. While the heap is being formed, H2O is ad- mitted to dissolve existing CUSO4 and to furnish the ore the moisture necessary for oxidation. The temperature in the chimneys may rise to 82 and 90° C. and even to the kindling temperature of the ore; it is, however, not allowed to exceed 82° C. ; at some mines it is held at 30-32° C, and the average temperature is 45- 60° C. If the heat rises too quickly in places, the chimneys are closed, which not only checks such rise, but causes the oxidation to spread more evenly through the heap. The combined warmth and oxidation cause hardly — perceptible fissures to form in the pyritic ore into which solvents penetrate, dissolve the Cu, and more or less disintegrate the disulphide. When oxidation has proceeded sufELciently, H2O charged with liquor from the cementation tanks is run on at the rate of 50 cu. m. (13,210 gal.) per hour until the soluble Cu has been ex- tracted. The water is turned off, the heap allowed to drain, and the sulphide mineral to oxidize, whereupon leaching is repeated. After about one year, the surface is dressed, i.e., the locations of the squares and ridges are inter- changed, and the gutters correspondingly shifted; further, the edge of the heap, having become hardened by crystallized copper salt, is dug into and formed into terraces in order that the salts may be readily extracted. It takes from six to seven years with massive ore, and from three to four years with schistose ore to reduce the Cu-content to 0.25-0.30 per cent., which is as far as tlie extraction can be carried with a profit. There is a loss of IS to 30 per cent, of weight of the ore in the process. The exhausted ore (washed sulphur ore) is removed and shipped to sulphuric acid plants. Upon screening through a 2-mesh sieve to prepare for the coarse- and fine-ore roasting kilns, as much as 50 per cent, will pass through the sieve, showing how strongly it becomes disintegrated by weathering. The solution with from 0.015 to 0.5 and even 0.6 per cent. Cu is of a reddish- green color; it contains Fe2.(S04)3, FeS04, CUSO4, H2SO4, besides Bi2.(S04)3, Sb2.(S04)3, Ag2S04, and Fe2.(As04)2. In order to reduce the Fe-consuming Fe2. (804)3 to FeS04,the liquor from the heap is run over freshly mined pyrite- fines when the reaction 7Fe2.(S04)3+FeS2+8H20 = i5FeS044-8H2S04 takes place. The "filter-bed" is laid in a reservoir formed by a masonry dam across a small ravine having a slope of about 5°; its surface is topped with the finest material and divided into so-ft. squares. The liquor remains in contact with LEACHING OF COPPER 423 the filter until its color has changed to a clear blue, when 90 per cent, of the Fe2.(S04)3 has been reduced to FeS04. It contains per cubic meter (35.31 cu ft.) 4 kg. (8.8 lb.) Cu, I (2.2 lb.) Fe^O,, 20 (44 lb.) FeO, 10 (22 lb.) H2SO4 and 0.3 (0.66 lb.) As. The precipitation (cementation) with pig iron is carried on in a series of flumes placed along the slope of a hill so that the liquor flows to and fro until after passing through a mile of flume it is discharged. and collected in part to be used again as solvent. Each main flume is made of 2, 3, or 4 smaller flumes separated from one another by walls. Each flume is 320 ft. long, 5.5 ft. wide and 2.25 ft. deep, constructed of 9X3 m. planks held together by wooden frames set in cement, the joints of the planks being rope-calked and pamted with asphalt. In order to permit cutting out of circuit, each end of a flume is pro- vided with a door with holes, closed by plugs, to drain off the liquor before remov- ing the cement Cu. The fall of the flumes increases from inlet to outlet, i. e., 0.5 per cent, for the first third, i per cent, for the second, and 2 per cent, for the last third, in order to diminish the increased Fe-consumption by the partly decopperized liquor. The bottom of a flume is loosely covered with boards, and the pig iron piled in grids at right angles to one another until the flume is filled, I ft. of flume holding i ton of pig iron. The liquor from the filter bed passes into a reservoir and thence runs into the flumes at the rate of 300 cu. m. (io,S9S cu. ft. or 79,260 gal.) per hour. In summer it reaches a temperature of 38° C, the higher the temperature the more rapid the precipitation. The precipitate near the head of the flume is the purest (Cu 93-94, As < 0.3 per cent.) and adheres to the pig u-on; lower down it is still red but becomes granular (Cu 75-90, As 0.5 per cent.); the remainder is more or less black (Cu 50, As S per cent.) and contains most of the graphite. Some flumes are cleaned daily; the liquor is run into settling tanks to recover suspended Cu, the pig iron is removed and piled on the dividing wall, adhering Cu knocked off, and the iron then returned to the flume. The cement Cu goes to a cleaning and concentrat- ing plant. Here red precipitate (Cu 70 per cent.) is screened, washed with a hose, compressed into cylinders, dried, and shipped. Black precipitate is briquetted and sun-dried, and so hardens and is ready to go to the blast-fur- naces. The liquor with Cu 15-20 g. per cubic meter goes to waste, as the Fe-consumption is too high to pay for the recovery of the Cu. The extrac- tion of Cu reaches 95 per cent. Table 107 gives analyses of in- and out-going liquors. Table 107. — Head- and Tail-liquors of Rio Tinto Liquor Per cent. Spec. gr. Cu FeO FeaOs H2SO4 Tot. solids Head 0.271S o.ooig 1.3908 I . 7202 0.0610 0.4874 0.4129 7.0872 6.9662 1.05818 Tail 1.05718 The consumption of iron under favorable circumstances is 1.3-1 -5 tons of pig iron (with 92 per cent. Fe) for i ton of Cu; usually it is nearer 1.75-2 tons. In view of the fact that the prices paid for pyrite have been increasing, be- 424 METALLURGY OF COPPER cause of the great advances made in the extraction of the Cu by the Longmaid- Henderson process (§ 225) and the briquetting and sintering of the leached iron oxide/ it is probable that the weathering process will be abandoned and the ore sold direct to the acid-maker, 213. Leaching Ore in Place. — This has been tried at Schmoellnitz, Hungary, by Buddeus;^ in England, and at Chase Creek Canyon, Ariz., by Austin.^ The mines of Schmoellnitz drowned in 1878 were pumped out in 1904, when the Cu recovered from the mine- water paid for the pumping. Durant^ records the flooding of a mine in England from two to four times a year, recovering first Cu from the water and later iron ochre, after which the water is used again for filling the mine. At Chase Creek Canyon, Austin experimented with the leaching of dissem- inated sulphide copper ore in place, by opening the ore-body in several horizons 50 ft. apart with drifts and cross-cuts, and introducing water which, filtering through the intervening rock, dissolved the Cu. The solutions were collected in a shaft starting from the lowest level, placed below the natural drainage of the locality, and then pumped into precipitating tanks. It was found that i ton of solution contained 0.4 lb. Cu, but there was not enough of it. Adverse conditions necessitated the abandoning of experiments to increase the amotmt. 214. Leaching Sulphide Ore after Conversion into Sulphate by Sulphatizing Roasting.* — The ore to be suited for this process must be a cupriferous pyrite, rich in FeSj, poor in CuS^; if there is not enough FeS2, this will have to be supphed; sulphatizing agents such as FeS04, AI2. (804)3, NaHS04 have been added to the ore-charge. The roast has to bfe carried on slowly and at a low temperature. The ore may be in lump-form or finely divided. Lump ore is roasted in heaps or kilns, fine ore in kilns or muffle-furnaces. Lump ore roasted in heaps will always be imperfectly sulphatized. A satisfactory yield in Cu can be obtained only by supplementing the roasting by some additional process such as weathering; the same is likely to be true with kiln-roasted ore whether this is coarse or fine; fine-ore muffle-furnaces which permit a good control of temperature and air give a good extraction without necessitating any auxiliary treatment. 215. Sulphatizing Heap Roasts.— i. Rio Tinto^ is the best-known example where the process of heap roasting was carried on, until 1888 when the passing- • Hofman, "General Metallurgy," 1913, p. 628 and foil. - ^ Berg. Hiittenm. Z., 1904, ixiii, 13, 41, 73. ^Min. Melh., 1911, n, 153, 187; Comments Eng. Min. J., 1911, xcxi, Channing, 601; Henry, 6g9;Webber, 700; Webber 1911, xcil, 197.' ' " ^ Eng. Min. J., 1911, xcii, 928. ' Experiments of Schoeller on Slimes, /. Soc. Chem. Ind., 1913, xxxii, 677. " Deumi^, Bull. Soc. Ind. Min., 1887, i, 835. Launay, Ann. Min., 1889, xvi, 491; Berg. Hiittenm. Z., 1890, xldc, 229; Eng. Min. J., i8go, L, 741. Collins, Tr. Inst. Min. Mel., 1893-94, Hi i7- Courtney, Proc. Inst. Civ. Eng., 1895-96, cxxv, 135. Chalon, Rev. Un. Min., 1902, LVii, 201. Correspondent, Min. J., i9io,lxxxix, 731. LEACHING OF COPPER 425 off of the sulphurous gases into the air was forbidden by law. Along the floor on which a heap is to be erected are built horizontal dry-stone air-flues 12-15 in- wide X 15-18 in. high. Usually three flues run longitudinally, 13 ft. apart, and one transversely. At the intersections dry-stone chimneys are erected. Small heaps, 20-26 ft. in diam. and 10 ft. high, hold 200 tons of ore, biu-n two months, and require 0.21 cord of wood per ton of ore; large heaps are oblong, 56X33 ft., and 1 1-12 ft. high, hold 1500 tons of ore, biurn six months, and require only 0.07 cord; medium-size heaps, 98X16 ft. and 8 ft. high, hold 330 tons of ore. The loss in weight by roasting is about 12 per cent. Small heaps yield more CUSO4 than large. The average extraction of Cu is 84 per cent. Roasted ore is transferred in 2-ton side-dump cars to cemented and asphalted masonry vats, 30 ft. long by 8 ft. wide by 3 ft. deep, having false bottoms of rough planks; and is leached with water in five to seven consecutive washings of 36 cu. ft. of water per ton of ore, each lasting 24 hr. The later practice was to wash the ore in place as in the weathering process, and thus save handling. Table 108. — Sulphatizing Roasting and Leaching at Rio Tinto in 1802 Raw ore Roasted ore Leached ore s Fe 48.0 41.0 2.2 2-3 i-S I.O 0. 2 1-3 2-5 6.7 39-0 ^•4 2.8 1.6 o-S 13-0 31.0 3-0 7-5 S5-0 I U Cu Pb 30 1.0 0. 2 Zn As HsO 4-5 25.3 0, Loss Insoluble 2-S Total 100. 100. 100. The brown copper liquor with 120-225 g. Cu per gallon is collected in a reservoir filled with raw fines as in the weathering process (§ 212) to reduce Fe2. (804)3 to FeS04, drawn into precipitating vats, 8 ft. wide by 3 ft. deep, charged with Fe, and then through precipitating flumes, 3-6 ft. wide, 1.5 ft. deep and over i mile long with a fall increasing from 0.4 to 2 per cent. The iron con- sumption is 1.25 Fe : i Cu. The cement copper with 80 per cent. Cu is dried, packed, and shipped. The leached ore retains 0.20 per cent. Cu; it is piled in waste heaps holding as much as 50,000 tons, weathered, and leached at intervals. The roasted ore weathers very slowly. In order to hasten decomposition, small heaps of raw ore are built against the large pile, ignited, and the fumes are made to enter the pile as much as possible. The elimination of As, Sb, and Bi by roasting is, according to Gibb,' As 76.8, Sb 22.0, Bi 14.8 per cent. 2. COPPEROPOLIS, Cai..^— Slatey pyritic ore with chalcopyrite containing ^Tr. A. J. M. E., 1903, xxxm, 668. ^Bull. No. 23, State Mineralogist Cal., 1902, p. 193; Private Communications by F. H. Harvey, 1893, iQ", and G. McM. Ross, 1912; process abandoned in 1904. 426 METALLURGY OF COPPER 5-5.5 per cent. Cu is crushed to pass a S-S-in. ring, piled on two layers of cord wood, placed crosswise and covered with brushwood, to form heaps 8-16 ft. wide and 4-1 2 ft. high, holding 3000 tons and roasting four months. The roasted ore, in which 40 per cent, of the Cu is present as CuS04, is transferred to inclined wooden floors covered with tar paper, piled in heaps 300 by i5oby 15 ft., covered with undecomposed fines of the roast heap, and leached with H2O and mother liquor from the precipitating vats. The solution with 7 per cent. Cu, st)me FeO, MgO and AI2O3, is run into two cement storage vats 14 by 14 by 16 ft. and thence drawn into a wooden copper-lined horizontal precipitating barrel 30 by 5 ft. supported by trunnions and provided with two charging openings and one relief valve for the escape of H. The barrel holds 3000 gal. of liquor, is charged with iron, rotated at a speed of 4 r.p. hr., and precipitates the Cu in 2 hr. The decopperized solution and cement Cu are discharged through a coarse screen, attached to the charging opening, to retain the iron in the barrel, and through a second screen, to retain coarse particles, into one of two settling tanks; the clear mother liquor is drawn off to be used as a solvent. The cement copper, 90-97 per cent. Cu, is washed, transferred to a drainage floor, dried on cast-iron pans heated by a direct fire from cord wood, and shipped. The fines covering the roasted heap are removed and placed on the wood bed of another heap or beneath it, as they form a layer that reduces loss of solution into the ground. The yield of copper is 40 per cent., the iron consumption is 1.15 Fe : i Cu. 3. Agordo, Italy. ^1 — The distinguishing feature of the operation formerly carried out here is kernel roasting (§ 54) of pyritic ore with Cu 1.60, Fe 42.00, S 20.00, As 1.40, Insol. 5.00 per cent., which gives 13 per cent, kernels with 3-6 per cent. Cu to be smelted, and 87 per cent, rinds with < 0.5 per cent. Cu to be leached. There are 64 vats of 1400 to 1700 cu. ft. capacity. A tank is filled in part with Uquor of 14-15° Be, then charged with about 4 tons of ore, which remains in contact with the Uquor for 24 hr. The solution, now of 31- 34° Be., is drawn off, clarified, the Fe2-(S04)3 reduced by SO2 gas to FeS04, run into vats, 13 ft. 1.5 in. long by 9 ft. 10 in. wide by 4 ft. 11 in. deep, and the Cu precipitated by Fe; the FeS04 formed is recovered as green vitriol. The leached ore is washed twice, the H2O remainmg 24 hr. in contact in each wash. With an unsatisfactory extraction, the leached ore is spread as a cover over a roast heap and then leached again. When the Cu-content has been reduced to 0.25 pet cent., the ore goes to the dump. This represents an extraction of only 50 per cent. The consumption of iron is 2.5 Fq : i Cu; the high figure is due to the H2SO4 formed by the reduction of Fe2.(S04)3 with SO2; without this reduction the iron consumption was 3.27 Fe : i Cu. 4. Other Localities.— Kernel roasting followed by leaching was tried at Ducktown, Tenn.^ Other older examples are: Balan, Transylvania;' 'Mazzuoli, Ann. Min., 1876, K, 1900; Berg. Huttenm. Z., 1876, xxxv, 363. Egleston, School Mines Quart., 1887-88, ix, 124, 256. Ernst-Monaco, Berg. HiUtenm. Z., 1891, L, 26. "Wendt, School Mines Quarts, 1885-86, vii, 218. ^Flechner, Oesl. Zt. Berg. HiiUenw., 1882, xxx, 355; 1883, xxxi, 435, 463. LEACHING OF COPPER 427 Maidenpec, Servia;^ Szalathna, Hungary i^ Colorado' {Monier process); Terrino process (Heating with Fe2.(N03)6).* All the sulphatizing roasting processes enumerated suffer from the drawback that there is an imperfect control of temperature and air, with the consequence that the ore is roasted either too little or too much, and in most cases is more or less sintered. The im- perfect extraction of Cu in the first leaching requires additional treatment of the ore in order to increase the yield. 216. Sulphatizing MufHe Roasts. — Roast- ing pyritic concentrate in a muffle fiu-nace under controlled temperature and air has been carried out first in the laboratory by Warlimont,^ and then at the works of Predazzo, Italy, by Hesse.* Other laboratory data are those of HoUis- Lannon-Quayle-Grommon,^ Austin,^ Handy.' Large scale work has been carried on by Wedge in his multiple-hearth down-draft muffle fur- nace, Figs. 446-447, and Laist.'" The results of Wedge are assembled in Table 109; the work of Laist is discussed in § 219. At the works of Predazzo, Italy, ^^ is treated a 60-mesh concentrate in which the Cu occurs as chalcopyrite; it contains Cu 7.1, Fe 17.25, S 8.66, CaO 2.90, MgO o.io, PjOb 0.23, AI2O3 5.45, Insol. 49.97 per cent. The roasting furnace is a horizontal boiler-iron revolving ' Simons, Berg. HiUtenm. Z., 1885, XLiv, 58. 'Beaugay, Ann. Min., 1884, vi, 453, Berg. Hiittenm. Z., 1885, XLIV, 242. Farbaky, op. cit., 1894, mi, 175, 183, 225, 241, 249. ' Reichenecker, op. cit., 1870, xxix, 449. Trippel, Eng. Min. J., 1872, xiv, 114, 129. *Berg. HiUtenm. Z., 1888, xlvii, 171. ^ Metallurgie, 1909, vi, 83, 127. 'Op. cit., 1909, VI, 580. ' Colo. School Mines Bull., 1908, iv, 112. 'Mines and Meth., 19 12, iv, 9. 'Eng. Min. J., 1912, xcrv, 487. ^"Eighth Internal. Congr. Appl. Chem., New York, 1912, III, 151; Tr. A. I. M. E. 1913, xLiv, «i8. Op. Cil. I913, XKVI, 362. " Hesse, Melallurgie, 1909, vi, 580, with drawings. a 1 a s || 1 S 33 00 r « ^0 ) 4) ■ 2 00 .0 Tj- S 2 ^ H 1-) to < Z n 1-1 fi r- 00 00 1 10 M r^ w M -1 ro r- 00 « 1 fO WW *o r- 00 Ov £ 3 1 a .2 1 u as e a -S. M 0. 4) hi 00 QO 3 ^1 5I 00 4) li < CO to g H en on CO M d < r1 d i-i N H < a m \ OS <- 00 £ 00 00 en C ■ IT) 00 M f1 0\ 0> Ci w n N § H t^ N 00 -ft <0 (*> -* M CO H ro ■* ro (T) M B fe a <3 ■3 1 s > C c a •c > L. P s T 1 rt c c I c 1 c.^ 0) 428 METALLURGY OF COPPER cylinder, 6 ft. 6f in. in diameter and 12 ft. s§ in. long, resembling a Briick- ner cylinder, provided with ribs to raise and stir the ore. The cylinder is enclosed in a brick chamber and heated externally from three naphtha atomizing burners, 6 ft. 6f in. from the shell; the flames strike a bridge wall which deflects the gases downward that they may pass around the cylinder and leave near the top through a flue provided with a damper. The furnace makes 0.75 r.p.m., and receives a charge of 3 tons which it roasts in 13-14 hr.; discharging and recharging take 4-5 hr., thus making the capacity for 24 hr. 4-5 tons; one man per shift tends the furnace. Through the air-inlet collar is inserted an iron-constantan thermo-electric pyrometer to measure the temperature, which at first is held at 480° C. and later at 560°. The outlet collar is connected with an earthenware acid-proof suction fan which draws in the air necessary for oxidation (an occasional lack of air may be supplemented by an auxiliary fan blower) and delivers the gases to a dust flue connected with a stack. Of the Cu, 65 per cent, is water-soluble, 31 per cent, acid-soluble (hot H2SO4 of 17° Be.). With ore containing 20 per cent. Cu, 98 per cent, is soluble in acid. The brick hning of the furnace becomes incrusted in time; the crust has to be removed several times a year, but 80 per cent, of its Cu is water-soluble, and an additional 14-16 per cent, acid-soluble. The fur- nace makes 20 per cent, flue-dust; 93 per cent, of its Cu is acid-soluble. It is es- sential for good work to have just the right amount of air in addition to maintain- ing a correct temperature. At 340° C. the S begins to burn, the temperature of the ore rises to 400° C. and makes it necessary to diminish and then to shut off the naphtha flames. The air to be admitted is regulated by the damper, the amount necessary being indicated by the temperature of the charge. The metals Fe and Cu are sulphatized at the same time; but FeS04 is changed at 480° C. into Fe203.2S03, and this is decomposed at 560° C, while CuSOs changes at 670° C. into 2CUO.SO3, and this at 736° into CuO. In roasting for CuO, which is done to free CuSO* from FeS04 (see below), the temperature is held between 590 and 610° C. A charge of 3 tons of roasted ore is leached with 6-7 cbm. (212-247 cu. ft., 1585-1850 gal.) H2SO4 of 17" Be. at 40-50° C. in a Hofmann vat (§ 240) 9 ft. io| in. in diameter and 5 ft. io| in. deep; the stirrer makes 24 r.p.m. In order to prevent any settling of ore, a perforated lead pipe served with com- pressed air is attached to the side-wall f in. above the bottom. Starting the stirrer requires 5 h.p., but 2 h.p. will keep it going. Five leaching tanks are placed in series on corresponding terraces, 3 ft. iij in. high, and are connected at the sides by pipes, 3I in. in diameter, provided with valves. The clarified liquors are drawn from the bottom through the settled ore by means of a Borchers valve. This^ consists, as shown in Figs. 432-433, of a suspended in- verted cylindrical cup which can be raised and lowered by means of a handle. While the ore is being agitated, the cup is lowered, Fig. 432, so as to close the top of the discharge-pipe for clear liquor. When the ore has settled and the ' See also Borchers, Metallurgie, igos, 11, 375. LEACHING OF COPPER 429 Figs. 432-433. — Borchers valve. liquor become clear, the cap is raised so that the rim is in the liquor, and the discharge cock for liquor opened. Fig. 433. Roasted ore is charged into the top vat and fresh acid run gradually into the third; the ore travels downward in five steps, and the liquor upward, being raised by means of acid-eggs. Tank No. 4 serves to wash the leached ore, and No. S to receive decopperized washed ore. Leaching in a tank lasts 2 hr., set- tling 0.5 hr. The Cu-liquor drawn from the top tank measures 24° Be., is only slightly acid, and contains 4 per cent. Cu, 0.8-1. i per cent. Fe", and 0.03 per cent. Fe"". It is freed from Fe by the Hofmann process (§ 240) to 0.08 per cent. Fe, in two stages of 50 hr. each, and filter pressed. The residue from the press is treated with cold dilute acid, i . 5-2 per cent. H2SO4, to dissolve any excess Cu and CU2SO4. The Cu-solution is concentrated to 33° Be. and crystallized; the market vitriol is 98 per cent. pure. Experiments carried on at the works of the Shannon Copper Co., Clifton, Ariz., by Schimerka^ with a sulphatizing roast of low-grade sulphide ore (Cu 2.37, S 3.02, SiOa 58.60, Fe 8.90, AI2O3 13.90, CaO 2.10, MgO 2.38, Zn and Mn traces) resulted in an extraction of 84.5 per cent, of the Cu with a consumption of 3.19 lb. H2SO4 per ton of ore. The Bradley process^ has been tried on a large scale at Anaconda. Its leading steps are sulphatizing roast, lixiviation, treatment of liquor with CaCU, filtration of CaS04 from CuCU, precipitation of Fe(0H)3 and A1(0H)3 by CuO or Ca(0H)2, filtration of hydroxide precipitates, and precipitation of CuO by CaCOs. It is not in operation at present. It was found that in a chloride solution of Fe, Al, and Cu, the precipitant CaCOs separated first Fe2(OH)3, and then CuCl2.3Cu(OH)2 with some Al2(OH)3. The Hybinette process' is in operation in Norway. From 2 to 20 per cent. Na2S04 is added in the sulphatizing roast, the CUSO4 formed is dissolved and the Cu electrodeposited. At the Braden copper mines^ in Chile, sulphatizing in Wedge furnaces and electrodeposition of Cu is the process used for treating low-grade ores. The sulphatization of the Cu in burnt pyrite can be accompUshed by roasting in a Wedge furnace, as shown in Table 109, or a furnace with a revolv- ing hearth, as advocated by Richard.^ The same can also be accomplished with ^ Eng. Mill. J., 1913, xcvi, 1107. ^Eng. Min. J., 1912, xcra, 47, 533; Met. Chem. Ejig., 1912, x, 178; Mines and Melh., 1912, lu, 404. ' Met. Chem. Eng., 1913, xi, 6. * Yeatman, Min. Sc. Press, 191 1, cm, 769. Editor, op. cit., 1913, cvi, 932. f Chem. Z., 19 1 2, XXXVI, 565. 430 METALLURGY OF COPPER spreading the ore, moistening with mother-liquor, and exposing to the sun, as advocated by Truchot. The second method has been recommended for burnt pyrite with 'NaCl. In the presence of CU2O there is separated Cu according to 3Cu20+2FeCl2+ (:c+3)H20+3'NaCl = 4CuCl+Cu2+2Fe(OH)3+:«;H20+)'NaCl. ThisCumight act upon CuCU as shown by Cu+CuCl2+a;H20+>'NaCl=2CuCl+a;H20+ yNaCl; but experience has shown that it is advisable to convert by roasting any CU2O present into CuO before leaching. The action of FeCU upon CuO overheated in roasting is imperfect; a similar behavior with Fe2.(S04)3 has been noted by Thomas.^ The FeClj is prepared by dissolving 120 lb. NaCl in 1000 lb. H2O, adding 280 lb. FeS04+aq. to the solution, and giving 200 lb. more of NaCl. The reaction taking place is: FeS04+(y+2)NaCI+xH20 = FeCl2+ NajSOi+yNaCl+ajHaO; Na2S04 falls out of solution and is removed. The Cu chlorides formed by the action of FeCl2 are separated from the Fe(0H)3 by filtration and brought in contact with metallic Fe to precipitate Cu according to 2CuCl+CuCl2+»H20+yNaCl+2Fe = 3Cu+2FeCl2+a;H20+yNaCl. The process is not in operation although it has been tried at several places. The interest in it is therefore mainly chemical. The advantages are, absence of chloridizing roast; use with calcareous ores; and small consumption of fuel and Fe; the disadvantages, low extraction of copper (tailings assay Cu 0.5 per cent.) ; loss of CI by the formation of basic salts [6 FeCl2+xH20+3'NaCl+30 = (4FeCl3+Fe203)+a;H20+yNaCl],and by included CuO or CuCU; difficulty of filtration of Fe(0H)3 and of basic salts; and precipitation of part of the dissolved Agby Cu. (2) The Hunt and Douglas Process No. II. ^ — This process, also regenera- tive, is based upon the following: (i) The solution of CuO in H2SO4, viz., CuO+H2S04+xH20 = CuS04 +(i-|-x)H20. (2) The partial chlorination of CUSO4 by FeCl2 or CaCl2, viz., 2CUSO4+ FeCl2+xH20 = CuS04+CuCl2+FeS04+xH20 or 2CuS04+CaCl2+xH20 = CuS04+CuCl2+CaSO4+xH2O. (3) The formation and precipitation of CuCl by the forcing of SO2 (9 per cent, vol.) through the CUSO4-CUCI2 solution, with the simultaneous regenera- tion of the H2SO4, which is used again as solvent after the expulsion of dissolved SO2, viz., CuS04-|-CuCl2 + S02-f (X-|-2)H204=2CUC1-|-2H2S04+XH20. (4) The regeneration of the FeCl2, or CaCl2, by the decomposition of CuCl ' Uetallurgie, 1904, i, 8, 39, 59. ' Hunt, Tr. A. I. M. E., 1881-82, x, 11; 1887-88, xvi, 80; Eng. Min. J., 1885, xl, 37. Douglas, Min. Res. U. S., 1883-84, 279. Howe, "Production Gold and Silver in the U. S.," 1883, 790. Franke, Metallurgie, 1910, vii, 486. Mangier, op. cit., 191 1, vni, 178. Canby, Eng. Min. J., 1911, xci, 1156. Douglas, op. cit., 1911, xcii, 51. 436 METALLURGY OF COPPER with Fe, or Ca(0H)2 viz., 2CuCl+a;H20+Fe = 2Cu+FeCl2+a;H20 or aCuCl +a;H20+Ca(OH)2 = 2Cu(OH)+CaCl2+a;H20. The advantages of this modification of process No. I are: absence of Fe2(OH)3 to be filtered, low consumption of Fe, and recovery of pure Cu. The disadvan- tages, loss of some Ag,* and imperfect precipitation of CuCl by SO2. The process was used by O. Hofmann^ at the Argentine works of the Kansas City Smelting and Refining Co. for treating matte with Cu 40, Pb 11, Fe 20, Zn 2, Mn I, S 21 per cent., and Ag 250 oz. per ton. He succeeded in overcoming the many difficulties encountered, but found it advantageous to modify the proc- ess by chloridizing the CuS04-solution with HCl, reducing CuCU to CuCI with Cu, filter-pressing CuCl, decomposing it with Fe, separating FeCl2 from Cu (90-94 per cent, pure), evaporating to dryness, heating in a retort in the presence of air and steam, and recovering the HCl formed. The modified proc- ess was in operation for some time, but was replaced by the manufacture of blue vitriol (§ 240), more profitable at that time. Other modifications are those by Stahl,* using MgCl2; Laist,* absorbing SO2 in a cold CuS04-NaCl solution and then heating in a closed vessel under pressure to form CUCI2; Douglas,' decomposing CU2CI2 electrolytically; Cap- pellen-Smith, decomposing CU2CI2 by heating with Ca(0H)2 and C to a low temperature, 224. Leaching of Sulphide Ore after Chloridizing Roasting. — Chloridation of sulphide copper with < 2.75 per cent. Cu in connection with heap-roasting used to be the common practice at Rio Tinto.* Heaps 20 by 26 ft. and 10 ft. high with 800 tons of ore, or 26 by 30 ft. with 1200 tons of ore, were built over three longitudinal and two transverse air flues, 20 in. square; the smaller heaps had two, the larger three chimneys. The ore was roasted and then leached. The leached ore was removed from the tanks, mixed with raw ore, 2-3 per cent, salt, and 2-3 per cent, pyrolusite. This mixture was now placed upon an ordinary heap to a depth of 16.5 ft., when this had been fired and SO2 was coming ofi freely. The chloridation then proceeded in the usual way. When the roasting was finished, the surface of the heap was divided by ridges into leaching beds 26 ft. square and watered. A leached heap was allowed to weather, and then watered at intervals to recover additional amounts of copper. The incrusted cover was broken up when necessary. 225. Leaching of Sulphide Ore after an Oxidizing Followed by Chloridizing Roast. Longmaid-Henderson Process. — This mode of procedure was in- " Blowing hot air through dilute H2SO4 containing small amounts of FeSOj, CuCl, and HCl, causes first the formation of Fe2.(S04)3and CuCh, and then of some AgCl which is dissolved by the chlorides, and later precipitated by CU2CI2. ' Min. Ind., 1908, xvn, 296. ^Zt. angew. Chem., sSgi, rv, 24; Iron, 1892, xxxix,i66; Oest. Zt. Berg. Huttenw., 1892, XL, 88; Berg. HiiUenm. Z., 1892, Li, 61. * U. S. Patent, No. 903732, Nov. 10, 1908. ^ Eng. Min. J., 1911, Lii, 51. «Launay, Ann. Min., 1889, xvi, 502; Eng. Min. J., 1890, l, 741; Berg. Hiitlenm. Z., 1890 XLDC, 230. LEACHING OF COPPER 437 vented by Longmaid in 1842 and improved by Henderson in i860; it goes by the name of Longmaid-Henderson process.^ It is suited for burnt pyrite, quite free from gangue and running low in Cu, and is based upon the chloridizing roasting of burnt pyrite (cinder) for the conversion of Cu, and with it of any small amount of Ag and Au present, into soluble chloride, followed by the recov- ery of these metals from the solution by precipitation with Fe; the residual FejOa forming a valuable iron ore. A modification of the process, less important since the advent of the electrolytic refining of copper, is the precipitation of Ag and Au before the Cu, and the working-up of the two products independently. The composition of pyrite; roasted in kilns for the production of SO2 in sulphuric-acid plants has been given on page 85. Iron is present mainly as Fe203, then follow FeS2, Fej. (804)3, and CuFeSa. Copper is present mainly as 6U2S, then follow CuS04,CuO and lastly comes CuFeS2. The forms in which Cu is present and the respective amounts are shown in Table in. Table hi. — Copper Compounds in Kiln-roasted Pyrite" Sample Cu, total, per cent. Copper, per cent., present as No. CUSO4 CuO CuzS CuFeSi I 4.41 50.00 16. 1 2gi.I0 S-70 2 4.67 14.46 13 -I 64.47 8.03 3 4.42 31-39 22.8 45-9° 0.00 4 1.86 12,00 19.8 54-9° 13-30 S 1.06 32. 10 18.6 38.80 10.50 Average . . . 3-28 27.9 18. 1 46 . 43 751 The burnt pyrite to be treated by the process must contain little gangue (< 20, usually < 10 per cent.) as this would consume an excessive amount of salt; it should not assay over 6 per cent. Cu, as there is danger of the forma- tion of kernels in roasting, which acts unfavorably upon chloridation (it usually contains < 4 per cent.); and lastly must show i-i-S parts of S for every part of Cu to obtain a satisfactory percentage of CUCI2. Any lack in S is made up by the addition of pyrite. •Lunge, G., "Sulphuric Acid and Alkali," Gurney and Jackson, London, 1913, i, part 3, 1470-1529. Wedding-Ulrich, Zt. Berg. HuUen. Salin. Wesen i. P., 1871, xrx, 298; Berg. HuUenm.Z., 1872, xxxi, 147. Brauning, Zi. Berg. HUttcn. Salin. Wescn i. P., 1877, xxxv, 156. Howe, "Production Gold and Silver in U. S.," 1883, 774. Egleston, Tr. A. I. M. E., 1885, XIV, 198. Schelle-Semlitsch, Oest. Zt. Berg. HiiUenw., 1893, XLi, 517, 531; Berg. Hut- lenm. Z., 1894, liii, 76. Stahl, op. cit., 1894, liii, i; 1897, lvi, 185, 235, 319. Helmhacker, Min. Sc. Press, 1898, lxxvi, 417. Krutwig, Rev. Un. Min., 1899, XLVi, 35. Clemmer, Min. Ind., 1899, viii, 197 (Comment, Eng. Min. J., 1900, lxx, 361); 1900, rx, 283. Gibb, Tr. A. I.M.E., 1903, xxxiii, 669. Bahlsen, Metallurgie, 1904,1, 258. Colby,/. I. St. Inst., 1906, ni> 359. Lilja, Met. Chem. Eng., 1910, viii, 395. Kothny, Oest. Jahrb., 1910, Lvm, 97; Metallurgie, 1911, viii, 389. Franke, op. cit., 1910, vii, 488. Mengler, op. cit., 1911, vm, 179. * Kothny, loc. cit. The older data of Wedding-Ulrich (Joe. cit.) show dififerent proportions. Complete analy- ses of burnt pyrite are given by Lunge (}oc. cit.) and Schelle-Semlitsch {loc. cit). 438 METALLURGY OF COPPER The operations to be considered are: crushing and mixing of ore and salt; chloridizing roasting; condensing of gases and vapors; leaching chloridized ore with water and tower-Hquor; clarifying the copper liquor; precipitation of Cu (with Ag and Au) by Fe; washing and refining the precipitated Cu; disposition of residue (blue billy, purple ore) from leaching; disposition of waste Hquor; precipitation of Ag and Au independently of Cu; results and costs. Illustrations of older plants have been given by Wedding (Widnes, St. Helens, England), Defrance (Hemixen, Belgium), and Brauning (Oker, Germany), and of recent plants by Clemmer (Natrona, Pa.), and Colby (Newark, N. J.). In Figs. 435-436 are given outline sketches of a modern 60-ton Longmaid- Henderson plant. Ore, i.e., roasted pyrite or cinder, and salt are received in a dehvery-bin to be transferred by means of a 14-in. conveyor belt and tripper to two 100-ton cinder- and one 20-ton salt-bins. These raw materials are fed in weighed quantities to four 2-ton revolving mixers and thence dis- charged through chutes into two No. 4 Krupp ball mills' from which the mked and ground pulp passes by means of conveyors or of chutes into the boat of the elevator which empties into a 60-ton storage-bin. From this the pulp passes through a chute into a second elevator which delivers into the 2-ton bin of the Wedge s-hearth muffle furnace, detail in Fig. 446, where it is chloridized. The chloridized ore is discharged through four openings in the bottom into i-ton cars, running on an elevated track, and delivered to the 2X7 = 14 leach- ing tanks, detail in Figs. 448-449, placed in two rows. The leached ore is removed from the tanks in 3-ton buckets traveling on an overhead trolley. The copper solution is collected in the concrete copper-liquor tank, whence it is run into nine copper-precipitating tanks placed in rows of three. These tanks receive the precipitating scrap iron through overhead trolley buckets filled from a storage building. The copper precipitate is transferred onto a copper screen moving over three wash-tanks, freed from iron, washed, settled and transferred to the filter-press. The liquor freed from copper is run from the tanks into an open concrete catch-pit charged with iron, in which floating particles of copper are settled and unprecipitated ones thrown out of solu- tion. The effluent passes over a bright piece of iron which ought not to become tarnished. It will be noted that the sloping floors are made of reinforced concrete, and have upturned sides, in order that all drippings may be collected and conducted to a receiving pit. 226. Crushing and Mixing of Ore and Salt. — Pyrite roasted in coarse-ore kilns does not exceed 3 in. in size, that from fine-ore kilns 0.25 in. As it is essential for a successful chloridation that ore and salt be intimately mixed, it becomes necessary to crush the two together. The finest size is prob- ably 8-mesh, the coarsest 4-mesh; under 8-mesh makes too many fines for satisfactory filtration in leaching, over 4-mesh causes imperfect chloridation. ^ Hofman, "General Metallurgy," 1913, p. 590, LEACHING OF COPPER 439 IQO.Ton Cinder., KIB'Dia'>\ .••' 3a//- B Catch Pit- of , Concrete, Unco M^\ Copper liquor tank Concrete S'-O'wide ?'-6"cieep t Fig. 43S- — Longmaid-Henderson Plant. 440 METALLURGY OF COPPER ^■iO-,BI--'>^''°^^'^ld LEACHING OF COPPER . 441 Clemmer' states that the best results are obtained by crushing ^ of a mixture through an 8-mesh screen, the rest through a 20-mesh, and then mixing the two products. The machine cornmonly used is an edge-roller^; sometimes a continu- ous Krupp ball mill is employed. In both cases attention has to be paid to the removal of dust. An edge-roller, 9 ft. in diameter, with two runners 52 in. in diameter, weighing each 9000 lb. and making 25 r.p.m., will crush through an 8-mesh sieve in 24 hr. 100-150 tons of mixture (10 per cent. NaCl) according to moistiu-e and coarseness of the feed; a ball mill, 6 ft. 2% in. in diameter, holdmg 80-100 steel balls, 5 in. in diameter and smaller, weighing about 18 lb, each, and making 22 r.p.m., will treat with an 8-mesh screen in 24 hr. 100-120 tons of mixture (10 per cent. NaCl). The smallest amount of salt necessary for an ore with 4 per cent. Cu is given by Kothny (see below) as 7.5 per cent.; the largest range in practice is from 10 to 20 per cent.; the usual limit until recently was 12 and 15 per cent., when Wedge reduced it with his down-draft furnace (see below) to 9 per cent. The salt is generally not dried before using, although dried salt is easier to crush. At Oker, carnallite (KCl.MgCl2+6H20) was used to replace some of the salt in order to furnish the H2O necessary for the formation of HCl. The crushed mixture is screened to insure uniformity. 227. Chloridizing Roasting and Condensation of Gases. — The chlorida- tion of copper has been explained as being due largely to the presence in burnt pyrite of CUSO4, which, acting upon NaCl, formed CuCl2 and Na2S04, and to the decomposing effect of FeS04, either present as such or formed by the oxidation of FeS. The FeSO* from both sources could act upon NaCl and form FeCl2, FeCls, and Na2S04; or, after it had been decomposed by heat, the SO3 set free would convert CU2S into CUSO4; or acting upon NaCl, it would give CI (which would chloridize CU2S) and HCl in the presence of H2O (and chloridize CuO). Kothny's analyses (p. 437) prove that neither FeS nor FeS04 is present in burnt pyrite. His experiments' have shown that with burnt pyrite mixed with salt and roasted at a temperature of 500-600° C. the following reactions take place: (i) 2FeS2+702 = Fe2.(S04)3 + S02 and Fe2.(S04)3 + 6NaCl = 3Na2S04 + FeaCle; (2) 2Cu2S-|-502 = 2CuS04-f-2CuO, 2CuO-|- 2502+02 = 2CUSO4, and 3CUO+ £62.(504)3 = Fe203+3CuS04 to some extent. (3) Cu504+2NaCl = CuCl2+Na2S04, 3CuO+Fe2Cl6 = Fe203+3CuCl2. (4) Cu2S+4CH-30 = 2CuCl2+S03 and 3Cu2S-F2Fe2Cl6+90 = 6CuCl2+ 2Fe203+3503, neghgible; (5) 2NaCl+S03+H20 = Na2S04+2HCl and CuO+HCl = no reaction. (6) Ag2S04+2NaCl = 2AgCl+Na2504 and Au+3Cl = AuCl3. The CuCl2 formed may be decomposed; by 2CuCl2+02 = 2CuO+2Cl2, a reaction which is much retarded by the presence of CI and HCl; by CUCI2+H2O ' Min. Ind., 1900, tx, 283. ' The Carlin mUl, Clemm, op. cit., p. 284. ^Oest. Jahrb., 1910, LVin, 97; Metallurgie, 1911, vra, 389. 442 - METALLURGY OF COPPER = CuO+2HCl; by CuCl2+heat = CuCl+Cl, which does not take place between 350 and 550° C. in the presence of much NaCl. Kothny concludes that for a successful chloridizing roast it is essential: (i) that ore and salt be finely divided (8-mesh) and intimately mixed; (2) that there be free access of air and vigorous rabbling; (3) that the amount of S present be equal that of Cu; (4) that there be enough NaCl added to the charge, with 4 per cent. Cu not < 7.5 per cent. NaCl; (5) that the roast be not unnecessarily prolonged, and (6) that the temperature be held between 500 and 600° C. Both reverberatory and mufiSe furnaces are used for roasting, and the ore rabbled either by hand or mechanically. The leading advantages of the rever- beratory furnace are its cheapness, and the fact that it requires about half the fuel of the muffle furnace; the latter gives a more even temperature, furnishes a more concentrated gas, requiring half the condensing capacity for the towers, and has a stronger oxidizing and chloridizing effect owing to the absence of fuel gases and the consequent smaller velocity of the gas current. Most reverbera- tory and muffle furnaces are single-hearth; recently multiple-hearth muffle furnaces have come into use, both hand and mechanically rabbled, and have effected a considerable saving in salt. Hand-rabbled furnaces have a very small capacity, from 2, more commonly from 5 to 9 tons in 24 hr. They treat a charge weighing from 1580 to 9600 lb. in from 6 to 12 hr., the great variation being due to the percentage of Cu and the manner of operating. The new mechanical furnaces of Wedge treat about 70 tons in 24 hr. and furnish on account of the mechanical rabbling a product richer in CUCI2 than can be obtained with hand work. For the practice in the U. S., hand-rabbled fiurnaces need not be dis- cussed in detail; they are fully treated by Lunge and Schelle-Semlitsch. The following examples are typical forms. (i). The Reverberatory Furnace of Oker.' — This is shown in Figs. 437- 439. The hearth, 24 ft. long by 10 ft. wide, is built of grooved tiles; in the roof, 2 ft. 2f in. high at the center, are four charging hoppers, m, and on either side four corresponding rabbling doors, I. The furnace is fixed with producer gas (gaseous fuel was replaced later by solid fuel) which, arriving through /, passes into the distributing flue, j, and enters the five combustion flues, g, where it meets air drawn in through ports, h. The flame travels underneath the hearth, rises at the opposite end, where auxiliary air enters through ports k', to finish the combustion and to furnish the O necessary for the oxidation of the S, and then retmrns passing over the ore-charge to leave the furnace through flue, i, leadmg to a gossage tower. The furnace treated two 2.5-ton charges in 24 hr., or in 24 hr. close on to 42 lb. per square foot hearth area; it requu-ed three men on a shift and consumed 10-12 per cent, coal on the chloridized ore with direct firing. A charge, crushed in an edge-roller to pass a 6-mesh screen, contained 85 per cent, burnt pyrite (CuO 9.80, FeaOs 53.14, FeSa 7.13, PbO 2.25, Ag 0.008, ZnO 2.43, MnaOs 0.57, SO3 9.51, AI2O3 4.43, Insol., etc., 11.65 per cent.) and 15 per cent. Strassfurt salt (NaCl 85.1, CaS04 4.0, MgS04 3.1, MgCl2 2.6, KCl 1.7 1 Brauning, Zt. Berg. Hutten. Salin. Wesen i. P., 1877, xxv, 156; the leaching department of the smeltery has been abandoned on account of the lack of suitable ore. LEACHING OF COPPER 443 HKSO4 0.2, H2O 3.3 per cent.). In order to prevent overheating, the full ore- charge is not introduced at once, but in four parts. Supposing that of the four hoppers. No. i is near the entrance of the producer gas, | of the charge are drop- ped through hoppers, 2 and 3, spread to cover half the hearth, allowed to come to a dark red, whereupon whitish sulphurous fumes will pass off, and rabbled occasionally. The | are transferred to cover the hearth under hoppers 3 and 4, when the thu-d J of the charge is dropped through hopper No. 2; and when this has become a dark red, the remaining j through hopper No. i. Charging and Section on Line CD ^^^^^^ Section on Line EF C Section on Line AS „ D P Figs. 437-439. — Reverberatory chloridizing furnace of Oker, Germany. bringing the charge to a uniform dark red to start the oxidation take about 4 hr. The fire is now checked, and air-flues, h', are opened; the charge is rabbled con- tinuously and occasionally transferred from the hotter to the cooler part of the hearth as may seem necessary. The bluish S-flame, tinged with the yellow of the NaCl, and whitish fumes gradually diminish until they disappear com- pletely. The chloridation is finished, and the charge withdrawn and transferred to the cooling floor. During the chloridation the chemical reactions furnish most of the required heat. Toward the end more gaseous fuel is needed than at first, to keep up the necessary temperature. Bright daylight is kept from the furnace room to enable the workmen to recognize the temperature of the charge. 444 METALLURGY OF COPPER LEACHING OF COPPER 445 During the roast, tests are made for ascertaining the amount of CuO present by leaching a sample with boiling H2O, filtering, boiling the residue in dilute HCl, and supersaturating with NH4OH; occasionally the residue from the HCl treat- ment is digested with HNO3 and an excess of NH4OH added to test for unde- composed CuaS. The chloridized ore has a greenish-grayish color. The aver- age chloridation result was Cu soluble in H2O 75 per cent. (CuCU), soluble in dilute HCl 25 per cent. (CU2CI2, CujOClj, CuO), soluble in aq. reg. 5 per cent. (CU2S). In the original paper are given complete analyses of two samples of ore taken at the middle and at the end of the roast. (2) The Mottle Furnace of the Pennsylvania Salt Mfg. Co., Natrona, Pa.» This is shown in Figs. 440-444- The hearth, 35 ft. long by 7 ft. 9 in. wide! equal to 271.25 sq. ft., is built of grooved tiles. In the roof are four feed open- ings served by cylindrical charge holders closed by sliding doors; on either side are five working doors. The furnace is direct-fired, the flame passes outward over the roof of the muffle, which is of double the usual thickness for a distance of 8.5 ft. to prevent overheating of charge, descends at the opposite end, returns underneath the muffle, and passes off to the chimney through an underground flue placed beneath the ash-pit. The hot roaster gases leave the muffle through cast-iron pipes joined by cast-iron bands, which are bolted together and packed with asbestos soaked in tar or a mixture of clay and tar. The furnace treats in 8 hr. a charge of 9600 lb. Spanish burnt pyrite with 2 per cent. Cu, ground through a 20-mesh sieve, and mixed with 10 per cent, salt; or in 24 hr. 106 lb. per square foot hearth area. In a similar furnace, 22 ft. 8 in.Xi3 ft. = 293.8 sq. ft., of the Tharsis works at Widnes, England,'' 4500 lb. ore with 17 per cent, salt are treated in 6 hr., or in 24 hr. 61 lb. per square foot hearth area. The progress of this roast is given in Table 112. It shows that there is a decrease in the chlorida- Table 112. — Progress of Chloridation in Hand- worked Mupele Furnace England OF Widnes, Percentage of total Cu After I hr. After 3hr. After 6.S hr. Soluble in H2O Soluble in dilute HCl, insoluble in H2O S4 38 8 SI 42 7 IS Soluble only in HNO3 S tion of the Cu from 54 to 51 per cent, between the first and third hour, and then a gradual increase until the maximum of 75 per cent, water-soluble CUCI2 has been reached. It may be that this decrease is accidental. (3) The Wedge Single-hearth Mechanical Reverberatory Furnace WITH Top Muffle Effect. — This furnace, shown in vertical section in Fig. 445, is constructed on the same principle as the pyrite burner discussed in § 69 and illustrated in Fig. 112. The furnace is 32 ft. in diameter, has a hearth 13 ft. wide corresponding to a hearth area of 768 sq. ft.; the central shaft, 4 ft. in ' Clemmer, Min. Ind., igoo, ix, 287. 'Wedding, loc. cit. 446 METALLURGY OF COPPER diameter, has four water-cooled stirring arms with heavy cast-iron rabbles, and makes i rev. in 4 min.; with two arms the shaft would make 2 r.p.m. The fur- nace is heated with four to six oil-burners. The products of combustion and the roaster gases pass off together through the chamber covering the roof of the hearth before they enter the flue leading into the gossage tower. The burnt pyrite, crushed with 17 per cent. NaCl in an edge-roller or a ball-mill to 8-mesh, is fed mechanically near the center of the furnace and travels over the hearth in from 2 to 2.5 hr. giving a chloridation of 96 per cent. (CuCla 40 per cent., CU2CI2, CuO, etc., 56 per cent.); the temperature is held at from 600 to 650° C. The furnace requires a 5-h.p. engine, consumes 11.6 gal. oil-residuum ( = 178.6 lb. coal) per ton charge, and puts through in 24 hr. from 80 to 100 tons of charge. Fig. 445. — Wedge single-hearth mechanical reverberatory roasting furnace with top muffle effect. (4) The Wedge Single-hearth Mechanical MuPFLE Furnace. — This is of the same general construction as the reverberatory furnace, with this difference: that both the upper and lower heating chambers of the mufHe are heated by oil- burners, and that the fire- and roaster-gases pass off separately to the stack and the gossage tower. A muffle furnace consumes twice the fuel that does the reverberatory furnace, and the bottom is readily corroded if the chlorida- tion is carried through on a single hearth. For this reason the reverberatory furnace is more common than the muffle furnace with a single-hearth type in spite of the great advantages the latter offers over the former as regards control of temperature and concentration of roaster-gas. The single muffle had recently been replaced by the following furnace. (5) The Wedg;e multiple-hearth, mechanical down-draft muffle fur- nace is shown in vertical section in Fig. 446. This is a five-hearth muffle LEACHING OF COPPER 447 furnace, i8 f t. 5 in. inner diameter and 3 1 ft. high. The feeding and course of the ore are the same as in the Wedge pyrite burner (§ 69, Fig. 112). On account of the low temperature the rabble-arms are air cooled instead of water cooled. The leading novelty lies in the manner of firing. In the older mechanical muffle furnaces constructed upon the MacDougall principle as, e.g., in the Haas f urnace,^ a single flame enters beneath the bottom muffle and then travels upward in zig- zag in the flues enclosing the muffles. The result is that the bottom muffle is 6' nti- — A Fig. 446. — Wedge five-hearth mechanical down-draft muiBe roasting furnace. overheated, if the upper muffles are to be brought to the desired temperature; they are too cool, if the heat in the bottom muffle is correct. In the Wedge furnace each muffle is heated independently by having either oil- or gas- burners placed between the roof of one muffle and the floor of the next following, as shown in Fig. 446, or by having, with solid fuel, two fire-places on the main floor and conducting the fire-gases independently to the heating spaces between ' W. R. Ingalls, " Metallurgy of Zinc and Cadmium," Hill Publishing Co., New York, 1903, ?• 143- 448 METALLURGY OF COPPER the muffles. In Fig. 446 are shown the oil-burners, of which there are eight; the products of combustion pass off at the right through horizontal flues into a main downtake leading to the stack. The roaster-gases zig-zag upward and pass from the top muffle into a main leading to^the gossage tower. The ore passes through the furnace in 8-1 1 hr., being stirred by two arms on a hearth making i rev. in 2 min. The chloridation is 86 per cent. Cu as CUCI2; an addi- tional 10 per cent. Cu or more is recovered by leaching with tower liquor. The furnace requires a S-h.p. engine, treats in 24 hr. 60 tons of Rio Tinto burnt pyrite containing 3.5 per cent. Cu, ground through a 20-mesh sieve and mixed with 7 per cent, salt; and consumes 280 lb. bituminous coal per ton of chloridized ore, which corresponds to 18.7 gal. oil residuum. The following is a record of burnt Spanish pyrite, crushed through a 20-mesh sieve and mixed with 10 per cent, salt, passing through the five-hearth muffle furnace. (Water — soluble Cu o . 54 per cent. = 26 . 09 per cent, extraction. Acid— Soluble Cu 0.97 " =46.80 " extraction. Insoluble Cu 0.56 Total Cu 2 . 08 par cent. = 7 2 . 89 per cent, extraction. (Water — soluble Cu 0.66 per cent. = 3i.9 per cent, extraction. Second Hearth, $io° C. \ Kcid — soluble C^ 0.95 " =45-8 " extraction. [insoluble Cu 0.46 " Total Cu 2 . 07 per cent. = 77.7 per cent, extraction. Third Hearth, 560° C. Water — soluble Cu 1.72 per cent. = 83 . 5 per cent, extraction. Acid — soluble Cu 0.18 " = 8.9 " extraction. Insoluble Cu 0.16 " Total Cu. .2.06 per cent. = 9 2. 4 per cent, extraction. I Water — soluble Cu 1,81 per cent.=84.6 per cent, extraction. Fourth Hearth, 620° C. Acid — soluble Cu 0.29 " =13.5 " extraction. [Insoluble Cu 0.04 " Total Cu 2 . 14 per cent. = 98 . i per cent, extraction. Fifth Hearth, cooling of Water — soluble Cu 1 . 8 2 per cent. = 85 . 04 per cent, extraction. Acid— soluble Cu:....o.28 " =13.06 " extraction. Insoluble Cu 0.04 " Total Cu 2 , 14 per cent. = 98 . 10 per cent, extraction. (6) The Wedge Multiple-hearth, Mechanical Down-drapt Reverbeea- TORY AND Muffle Furnace.— This furnace, shown in vertical section in Fig. 447, has eight hearths over which the ore travels downward in the usual way in from 8 to 13 hr., being stirred with two arms on each hearth making i rev. in 2 min. as in the other furnaces of Wedge. The novel part of this furnace is the mode LEACHING OF COPPER 449 of firing from two lateral fire-places. The gases from fire-place, a, e.g., rise in the vertical flue, enter ports e and /, come in contact with the ore spread over hearths Nos. i and 2, heat and kindle it, and pass off into flue b, leading to the stack. The kindled ore is transferred onto hearth No. 3, which forms the bottom of the muffle. The heat generated by oxidation and chloridation is sufficient to make extraneous fuel unnecessary while the ore travels over hearths Nos. 3, 4, and 5. If the temperature becomes too low on hearths Nos. 6 and 7, the dampers closing the heating flues c and d, are drawn the amount required to furnish the muffles the desired amounts of heat. The gases from the six muffles pass off Fig. 447.— Wedge multiple-hearth Mechanical down-draft reverberatory and muffle furnace. into the condensation tower. By the arrangement shown, the temperature of the furnace can be regulated to suit the character of the ore that is to be chlori- dized. With a pure Spanish pyrite the dampers of flues c and d, wHl remain closed, as the temperature of 500-600° C. is suflicient to obtain with a low per- centage of NaCl a high chloridation. With burnt pyrite containing some blende or galena, the dampers wUl have to be opened more or less in order to furnish the heat necessary for the decomposition of ZnS04 and the partial dis- sociation of PbS04. This Wedge furnace is identical with that of Ramen and Beskow used m most of the modern European plants, the furnaces having been constructed mde- pendently on either side of the Atlantic. (7) Condensation of Gases.— The gases issuing from a chloridizmg fur- nace contain SO2, SO3, H2SO4 (from S02+Cl2+2H20 = H2S04+2HC1), CI, HCl, N and 0, some volatilized CU2CI2, As, Sb, flue-dust, and wxth reverberatory 29 4SO METALLURGY OF COPPER furnaces, CO2 and perhaps some CO. They ascend in a gossage tower in which water trickles down slowly absorbing the acids, condensing volatilized chlorides, and collecting particles of flue-dust. The collected water forms the "Tower Liquor" used as a solvent for CU2CI2 and CuO. The gossage tower is a square or circular shell of heavy sheet lead suspended in a wooden frame or a square brick tower lined with acid-proof brick, packed in the case of muffle furnaces with coke or quartz, in the case of reverberatory furnaces with acid-proof brick laid checkerwise, as larger interstitial spaces are necessary for the greater volume of gas. With the quickly working muffle furnaces at Natrona there is in use one tower 12 ft. square and 50 ft. high for seven furnaces having a total hearth area of 1900.75 sq. ft., or I sq. ft. horizontal condensing area for 13.2 sq. ft. hearth area, treating, in 24 hr., 1400 lb. ore mixture. With the slowly working reverberatory furnaces of Oker there are in use two towers 5.6 ft. square and 17.4 ft. high for three furnaces having a hearth area of 720 sq. ft., or I sq. ft. horizontal condensing area for 22.9 sq. ft. hearth area, treating in 24 hr. 961.8 lb. ore mixture. With the s-hearth Wedge muffle furnace having a hearth area of 1246 sq. ft. and treating 60 tons of charge in 24 hr. there is in operation a gossage tower 8 ft. 4 in. square ( = 69.4 sq. ft.) and 41 ft. 3 in. high, or i sq. ft. horizontal condensing area for 18 sq. ft. hearth area, treating in 24 hr. 1734 lb. ore mixture. With the Wedge reverberatory furnace the condensing area required is twice that for the muffle fxurnace. About 48 cu. ft. of water are required per ton of roasted ore. 228. Leaching Chloridized Ore by Water and Tower Liquor. — The leach- ing vats at present are usually 12 ft. square, 4-5 ft. deep, and hold about 10 tons of charge. ,They are made of 3-in. planks, well calked with oakum and red lead, and tied by cast-iron corner-pieces and wrought-iron girder-shaped screw bolts. The wood is painted on both sides with tar; in some instances the vatshavebeen lined with lead. The filter-bottom has been constructed in various ways. The simplest is to place on the floor close together 2 X 2-in. slats beveled at the top and cover them with a filter-bed of small pieces of coke. A better method is to protect the wooden floor with a layer of hard-burnt acid-proof perforated brick and place on this a gravel-filter 6 in. deep, made up of one 3-in. layer of pebbles 1.5 in. in diameter, followed by another of sand 0.75 in. in diameter. Figs. 448-449 show the vat of the Penna. Salt Mfg. Co. of Natrona, Pa. The 3-in. yellow-pine planks used in the construction are well tarred before being put in place. The vat consists of an outer and an inner box separated by a 3-in. layer of sand and soft pitch poured in place. The filter consists of hard-burnt, acid- proof brick laid in straw. Figs. 450-451 represent the spigot for drawing off the solution. In front of a row of tanks are two launders for strong and for weak Hquors to be delivered to the clarifying tanks on the next lower level. The English mode of operatmg is to dump the ore hot (200° C.) into the vat and then fill the vat with weak wash liquor. This remains in contact with the LEACHING OF COPPER 451 ore for about 2 hr., becomes heated, and, dissolving most of the CuCU, becomes strong (8° Be.), so that it can be drawn into the clarifying and thence into the precipitating vats. When withdrawn, the ore is washed with hot water, produc- ing weak liquor, which is stored and serves as first wash- water for another tank. The water-leaches carry at least 75 per cent, of the Cu and 95 per cent, of the Ag. The Cu extracted is purer than that recovered by means of the tower liquor with which the ore-vat is now filled, because this liquor may contain As, Sb, Bi, Pb, etc. The ore used to require as many as six treatments with tower liquor to extract an additional 20 per cent, of Cu.' The leaching is not continu- FiGS. 448-449. — Leaching vat, paved with acid-proof brick laid in straw. ous; the different washes are allowed to remain in contact with the ore for given periods which are determined in part to avoid prolonging the whole treatment beyond 48 hr. At Oker, Germany, part of the mother liquor from the precipitation of the Cu is used as first solvent after having been heated to 40° C. This liquor being used over and over becomes charged with NaCl, FeCl2 and other chlorides; it weighs 18° Be., and contains Cu 0.0x5, Pb trace, FeO 2.14, FezOs 0.15, AI2O3 o.ii, ZnO 0.06, MnO 0.31 (NiCo)O o.oi, CaO 0.12, MgO 0.52, Alk. 2.6x, CI 2.56, H2SO4 5.89, As and Sb traces, total solids 14-495 per cent. The FeCU has a chloridizing effect upon CuO; the chlorides assist the solution of ■ With mechanical furnaces more CuCU is produced than with hand-raked furnaces as long as the temperature remains the same, so that the water-leach may extract as much as 85 per cent, of the Cu. 452 METALLURGY OF COPPER 1" i. AgCl, AuCU, and CU2CI2.1 The leaching is continuous and is stopped after from 4 to 5 hr., when the solution ceases to show a bluish color. The liquor extracts from 75 to 80 per cent, of the Cu, weighs 38° Be., and contains: Cu 3.71, Pb o.oi, Ag 0.005, Bi trace, (FeAl)203 0.29, ZnO 4-97> MnO 0.58, (NiCo)O 0.04, CaO trace, MgO 0.27, Alk. 10.60, CI 12.56, SO3 8.95, As and Sb 0.32 per cent. As in time it becomes overcharged with salts, it is concentrated by storing in vats in the open, and the crystallization of salts assisted by introducing brush-wood. The tower liquor which is subsequently used as solvent is run into the ore-vat, and remains there until its dissolving power has been used up, which lasts about 4 hr. The last solvent is boiling dilute H2SO4 of 8° Be.; it remains in contact with the ore for 48 hr. The time for treatment of a charge is about three and one-half days. 229. Clarifying of Copper Liquor. — The rich copper liquor from the leaching tanks, above 18° Be. is likely to be cloudy from fine ore, PbS04, etc. The PbS04 has been found to carry down considerable Au, assaying as much as 5 oz. per ton. The liquor is run into tarred wooden settling tanks, usually 12 ft. square and 6 ft. deep, which, have a discharge through a perforated wooden block, 6X6 in., placed in the side near the bottom. The number of settling tanks is the same as that of the leaching tanks. Settling takes several hours. In front of a row of tanks is a single launder to receive the clarified liquor. 230. Precipitation of Copper by Iron. — A precipitating vat, made of wood and tarred, is 12 ft. square and 6 ft. deep. It has a false bottom of slats 2 ft. above the true bottom, to furnish a support for the iron, and a space for the collection of the cement copper; it is provided with a pipe for heating the liquor by means of live steam, and has a discharge for liquor through a 6-in. wooden block closed by means of a plug. There are half as many precipitat- ing tanks as there are leaching tanks. Each tank is filled loosely with scrap iron, the copper liquor is run in, and the steam turned on to bring it to a boil. Tanks are kept covered with boards to diminish the loss of heat and to retard the formation of oxychlorides, which increase the consumption of Fe. Precipi- tation may last only 12 hr., but usually takes a day and even longer; it is finished if a bright iron rod does not become tarnished with Cu. When this is the case, the mother liquor is run off through settling tanks, sometimes also through a horse-hair filter, to settle and catch particles of float copper. For the sake of safety the liquors from a row of vats are passed through auxiliary precipitating tanks placed in series; in the overflow of the last is suspended a bright iron rod. 'As the presence of CuzCU interferes with the precipitation of Ag by KI (Claudet method), it would be necessary to leach first with H2O and then with mother liquor, if the Ag was to be thus recovered. Plug Tap Figs. 450-451. — ■Spig9t of Leach- ing vat. LEACHING OF COPPER 453 There is consumed i lb. Fe for i lb. Cu, the low consumption being due to the CujCU present. With rich solutions, a clean-up is made once a week; with poor solutions once a month. In both cases the mud is passed over 8-mesh copper screens to remove particles of Fe. 231. Washing and Refining of Cement Copper.— The cement copper is transferred from the precipitating tanks to washing-vats, where it is freed from all chloride liquor. Careful washing is essential, as in the subsequent smelting, any CI would cause a considerable loss of Cu by volatilization. Analyses of cement copper are given in Table 113. Table 113- -Analyses of Cement Copper England Oker, Germany Hemixen, Belgium Washed Natrona, Pa. Cu 72.50 2.60 0.046 77-45 0.63 0. 10 0.006 0.04 o-iS 6.72 75-07 95-93 o.os 90 pb Ab . 35 oz. Au O.IS oz. Bi As 0.306 0. 14 0.35 Sb FejOj 4.41 FeO 5-80 2.36 2-55 Fe AkO. 0.99 1 .02 0.02 0.03 0. 10 2.71 Zn Mn ' 0.99 NiCo 0.17 O.IO CaO 1.28 MeO+Alk NaCl 0.55 2.04 NasSOi SO. 458 1. 19 0.61 3-654 Brauning 0. 16 0.21 CI 4.10 8.00 Egleston HjO Referenca Lunge Stahl, Disser- tation, 1886 Clemmer The washed cement copper is partly dried, compressed, and bagged if it is to be shipped. If it is to be treated at the leaching plant, it is charged more or less moist (8-10 per. cent H2O) into a reverberatory furnace either by itself or with the addition of pure white metal, and smelted for blister copper; if it is not sufficiently pure for this purpose, it is added to a matte charge. 232. Disposition of Residue from Leaching, and of Waste Liquor.— The residue of the leaching vat is a rich iron ore with 90+ per cent. FeaOs, usually low in P and S if it has been well washed. It goes by the name of Purple Ore or Blue Billy. It is removed from the vats by shoveling onto a sUghtly inclined platform back of the leaching vats, i.e., on the side opposite the clarifying tanks. The platform has discharge-openings through which the ore is transferred into cars after the water has been drained off. Table 114 gives a few analyses. 454 METALLURGY OF COPPER Table 114. — Analyses of Purple Ore England Oker, Germany FeaOs AI2O3 90.61 95 10 79 3 Cu s o.iS 0.08 0.18 0.07 . 3-0 , 8 P PbS04 1 .46 0.37 1.29 0.49 CaS04. CaO 2.5 MgO+alk .... 1 .0 Na2S0, 0.37 0.28 0. 29 NaCl H2SO4 5,5 Insoluble 6.30 2 13 6 Reference Lunge Lunge Brauning Purple ore is used as a flux for siliceous lead ores, as a fettling for puddling fur- naces, or as an iron ore for blastfurnaces. In the last case' it is usually first con- verted into lump form by briquetting and sintering (Grondall process), by nodu- lizing, or by mixing with fuel and agglomerating in a D wight-Lloyd machine whereby the S-content is reduced to traces. The disposition of waste liquor has to be considered in the location of a plant, as the pollution of rivers may cause serious inconveniences. Attempts have been made to recover the Na2S04, but they have not been successful (Lunge). 233. Precipitation of Copper Independently of Silver and Gold.' — Several processes have been devised for the separate recovery of the small amounts of Ag and Au present in the CuCla-solution by precipitating with suitable reagents. Since the perfection of the electrolytic refining of copper, which has reduced the cost to $4 or $s per ton of cathode copper, these processes have lost their former importance. (i) The Claudet Process. — This process is in use in many European plants, but has been given up in this country, as the precipitation was found to be incomplete, leaving, according to Clemmer, 5 oz. Ag. per ton in the copper, and expensive when compared with the price received by the electrolytic refiner who pays for 95+ per cent, of the silver-content. The process consists of precipitating Ag (Au) by Znia as Agl, and decompos- ing the separated precipitate with Zn and HCl whereby the Znl2 is regenerated. The Ag-content in the clarified copper liquor from the water-leaches in the Eng- lish method of leaching, or from the final liquor-wash in the Oker method, is determined, the solution drawn off into a precipitating vat, and diluted with 10 per cent. H2O containing an excess of Znl2 over that required for the Ag, as some Pb is precipitated as Pbl2. The dilution causes some PbS04 and CU2CI2 to separate. The presence of CU2CI2 interferes with the complete precipitation of ' Hofman, "General Metallurgy,'' 1913, p. 629, 644. ^ Stahl, Berg. HiUlenm. Z., 1892, li, 443. LEACHING OF COPPER 455 the Ag. The Agia settles in about 48 hr. ; this time has been reduced to 24 hr. at Oker by the addition of a coagulant of glue (60 g. glue +10 liters H2O) and tannin (30-40 liters) obtained by boiling white-oak bark). The precipitate consists principally of Aglj, Pbl2 and PbS04. It is removed from the vat when a sufficient amount has accumulated, washed, and treated with Zn and HCl. The loss in I is made good by addition of KI. Metallic sponge obtained from its decomposition contained Ag 5.95, Auo.o6, Pb 62.28, Cu. 0.60, ZnO 15.46, FejOs 1.50, CaO i.io, SO3 7.68, Insol. 1.75 per cent. (Lunge). (2) The Mayer Process.— Here Ag is precipitated with Nal, and the Agl treated with NajS, forming AgaS, and Nal. , The precipitate at Atvidaberg contained 10.5 per cent. Ag, that of Konigshiitte 25.30 per cent. (3) The Gibbs Process. — By fractional precipitation with H2S, nearly all the Ag is thrown down as AgaS with about 6 per cent, of the Cu, furnishing a black slime assaying about 200 oz. Ag per ton; the Cu, precipitated later with Fe, assays about 3 oz. Ag per ton. (4) The Snelus Process. — Iron sponge is blown into the solution to precipi- tate about 19 per cent, of the Cu, which carries down about 80 per cent, of the Ag. (5) The Jardine and Chadwick Process. — Dilution of the Cu-liquor is to cause falling out of AgCl, and addition of Pb(C2H302)2+3H20 to form PbS04, which quickly carries down most of the AgCl. 234. Results and Cost. — The yield in Cu is from 95 to 98 per cent.; that of Ag (Au) is about 75 per cent, with ores assaying from 0.75 to 1.2 oz. Ag and 0.02 oz. Au per ton. The cost of working a ton of burnt Spanish pyrite (in 1899), at Natrona, Pa., with a plant treating 200 tons charge per day with hand- rabbled muflSe-furnaces^ was as follows: Labor, 80 men at $1.50-2.50, $134.75; unloading cinder and salt, and loading purple ore, $35; 21 tons of salt at $3, $63; pyrite fines, $7; 20 tons of coal at $1, $20; 5.5 tons of iron scrap at $7, $38.50; repairs, depreciations, management, etc., $40; total $338.25, or $1.87 per ton burnt pyrite and $1.69 per ton mixture. As the 14 furnaces of the plant require 28 men at $1.75 = $49, the cost of treatment with a mechanical furnace ought to be considerably lower; the amount of salt required ought also to be reduced on account of the more uniform stirring. The cost of treatment at present (including grinding, f urnacing, leaching, pre- cipitating with recovery of a portion of the gold, silver, and lead values), using an eight-hearth furna,ce (Fig. 447) and recovering 47 lb. copper is, according to the best European practice, substituting American prices for labor: Labor in process $0.67; labor in repairs $0. 1 1 ; materials in process $0.70; materials in repairs $0.20; total $1.68. The cost of materials used in the process (70 cents) is made up as follows: fuel for boilers 8 cents; fuel for f urnacing 8 cents (4 per cent. coaP on burnt pyrite at $2.16 per ton); salt 38 cents; iron 10 cents; miscellaneous 6 cents. The cost of furnacing alone is $0.25 (labor 13 cents, fuel 8 cents, repairs 4 cents). ' Clemmer, Min. Ind., 1899, viii, 202. ' Some European plants use only 2 per cent. coal. 4S6 METALLURGY OF COPPER B. Leaching Copper Matte 235. Leaching of Copper Matte in General. — Wet treatment of copper matte need to be considered only when this carries precious metals; if this is not the case, the matte is usually brought forward to metal by some smelting process. Lixiviation of copper matte was more common formerly than it is at present, when matte is being more and more brought forward to metallic copper to be refined electrolytically, as not only are Ag and Au recovered, but a high-grade metallic copper is produced instead of blue vitriol, for which the return is likely to be unsatisfactory. The processes used for leaching copper matte are of two kinds. In both the matte is given a preliminary roast. In the first, the Ag is converted to AgCl (Augustin process) or Ag2S04 (Ziervogel process), dissolved, and the CuO later reduced to metal; in the second, the Ag remains intact while the CuO is dissolved by H2SO4, and either sold as blue vitriol (Freiberg, Hofmann proc- esses) or precipitated as CU2CI2 and then reduced to metal (Hunt and Douglas process, No. II). 236. The Augustin Process.' — The process has been used for ore, matte, speise, and metallic copper (§ 242). The leading steps in the process with copper matte are: oxidizing roast to produce CuO and Ag2S04; chloridizing roast to convert AgzSO^ into AgCl; solution of AgCl in hot brine; precipitation of Ag by means of Cu; recovery of Cu by means of Fe. The process is obsolete. The principal reasons for this are: the small dissolving power of brine^ for AgCl necessitating heating, storing, and handUng of large volumes of liquor; the imperfect extraction of Ag, especially in the presence of As, Sb, and Zn; the volatilization of AuCU and its insolubility in brine. The process was used at Mansfeld, Prussia, from 1840 to 1849;' Freiberg, Saxony, from 1848-1862;* at Black Hawk, Colo.,^ where it served to increase the yield in Ag from the Ziervogel process (§237); at Kosaka, Japan^ in conjunction with the Patera process; and other places. 237. The Ziervogel Process in General.'' — This process was invented in ^ A. Grutzner, "Die Augustin'scHe Silberextraction in ihrer Anwendung auf Hutten- producte und Erze," Vieweg, Brunswick, 1851. Kerl-Crookes-Rolirig, "Practical Treatise on Metallurgy," Longmans, Green & Co., London, 1868, i, 368. Rivot, L. E., "Trait6 de M^tallurgie," Dunod, Paris, 187 1, i, 405.^ Howe, "Production Gold and Silver in the U. S., 1883, p. 764. Balling, C. A. M., "Metallhuttenkunde," Springer, Berlin, 1885, p. 355. 2 Hahn, Tr. A. I. M. E., 1873-74, n, 99; Eng. Min. J., 1898, Lxv, 434. ' Grutzner, loc. cit. * Kerl-Crookes-Rohrig, loc. cit. « Egleston, Tr. A. I. M. E., 1875-76, iv, 295. * Kmvabara, School Min. Quart., 1893-94, xv, 355. ' Steinbeck, Zt. Berg. Hiitten. Salin. Wesen i. P., 1863, xi, 95. Rivot, L. E., "Traite de M6tallurgie," 1871, i, 425. Howe, "Production Gold and Silver in the U. S., 1883, p. 753. Bradford, Tr. A. I. M. E., 1903, xxxm, sp. At Mansfeld: Leuschner, Zt. Berg. Hiitten. Salin. Wesen i. P., 1869, xyn, 135; Berg. LEACHING OF COPPER 457 1840, introduced at Mansfeld in 1844, and is carried on there at present. Pul- verized silver-bearing high-grade copper matte is subjected to a sulphatizing roast to form CuO, FeaOa and Ag2S04; the Ag2S04 is dissolved with hot H2O acidulated with H2SO4, and precipitated from its solution with granulated or sheet copper. The resulting mother Uquor is used again as solvent for Ag2S04 and its Cu recovered at intervals by mearis of Fe. The leached CuO, con- taining some Fe203, and any precipitated Cu are smelted in a reverberatory furnace for blister copper. Besides in Mansfeld, the process has been (is?) in use at Swansea, Wales, and Argo, Colo, (closed in 1909). The process is suited only for a reverberatory furnace or converter matte which runs high in Cu, low in Ag, and is pretty free from impurities, such as Pb, As, Sb, Bi. The Cu must be present as CU2S; 0.06 per cent, is the hmit for metallic Cu if the sulphatization of the Ag is to be satisfactory. The Cu- content must be high in order to furnish the necessary SO3 from the decomposi- tion of CUSO4 for sulphatizing Ag2S, which in its turn must be rightly appor- tioned to the Cu-content. The compounds PbS and Sb2S3 cause sintering; As and Sb form insoluble Ag3As04 and AgSbOs, and Bi^ an insoluble Bi2.(S04)3-»;Ag2S04. The matte treated at Mansfeld up to 1874 contained 64-65 per cent. Cu; since then 74-75 per cent, has been made the standard, as the results of a series of tests carried on between 1869 and 187 1 had shown that the higher tenor gave a better yield in Ag as well as a higher grade of Cu. The leading components of the matte are: Cu 74, Ag 0.4, Pb 0.6-0.8, Fe 1.2-2.0, Ni 0.4, Co 0.1, As o.oi, Sb none, S 19-20, metallic Cu 0.03-0.05 per cent. ■ The white and the pimple metal treated at Argo assayed Cu 60 per cent., Ag 750 oz., Au 10 oz. per ton, and Cu 77 per cent., Ag 90 oz., Au 0.2 oz. per ton. The following discussion is confined to Mansfeld; the Colorado plant, which was stopped in 1909 and will not be taken up again, has been described by Pearce;^ details of English practice have not been made public. 238. Ziervogel Process at the Gottesbelohnung Works, Mansfeld. (i) Crushing. — The matte with 74-75 per cent. Cu, broken by hand, is pulverized in ball-mills to pass a i6-mesh screen; a Briickner-Sachsenberg mill with 0.5 ton of 4-in. balls pulverizes in 24 hr. 17 tons of matte; a Gruson mill with 0.7 ton of 6-in. balls, 24 tons. (2) Roasting. — In roasting, FeS is converted into Fe203 passing in part through the stages of FeS04 and Fe2S06; the behavior of CU2S is similar except- ing that CU2O is formed as long as there is present any CU2S, and has to be con- verted completely into CuO. The SO3 set free by the decomposition of CuSO* causes Ag2S and Ag2 to be converted in Ag2S04 as shown by Ag2S-|-4S03 = HUUenm. Z., 1869, xxix, 432; Report of 1881, op. cit., 1881, xl, 430; Report of 1904, Metal- lurgie, 1904, I, 229; Report of 1907, op. cit., 1908, v, 27. Egleston, 5cAoo/ Min. Quart., 1890- 91, Xli, 207. Private Notes, 1911. Private Communication by R. Franke, 1913. In Colorado: Egleston, Tr. A. I. M. E., 1876, iv, 276. Pearce, op. cit. 1889-90, xvni, 55. ' Pearce, op. cit., 1889-90, xvra, 67. 'Loc. cit. 458 METALLURGY OF COFFER Ag2S04+4S02 and 2Ag+2S03=Ag2S04+S02. Sulphatization of Ag, however, is difficult and unsatisfactory. The roasting operation is divided into the stages of rough-roast and finish- ing-roast; the latter operation is repeated once with leached matte. The rough-roast, Figs. 452-453, is carried on in a four-hearth circular mechanical fine-ore kiln, 12 ft. 5! in. in diameter, with central vertical driving shaft, which has two rabbling arms with trailing teeth to each hearth, and is geared from above to make 1.25 r.p.m. The two upper hearths are connected by a flue as well as the two lower. The heat necessary for roasting is furnished by the oxidation of the matte. The furnace is fed by charges, and therefore works intermittently. Each upper hearth receives a charge of 1.4 tons of matte, which is roasted at a temperature of 500-600° C. The charge remains 3 hr. on the upper hearth and is then transferred to the lower, where it roasts Section on Line d-b ^ .L ^ Figs. 452-453. — Mechanical rough-roasting kiln. 4 hr. 52 min. more. A furnace with its pau- of double hearths treats in 24 hr. 14 tons of matte. The rough-roasted matte retains 5 per cent. S, the Ag2S remains unchanged; FeS is converted into Fe203; 45 per cent, of the Cu is present as CU2O; the remaining 55 per cent, as CU2S, CuSO*, and CuO. If the S- or the CuzO-content varies much from the above figures, the extraction of Ag becomes unsatisfactory. The roasted ore goes to the cooling floor, where it remains 8 hr., and is then reground in the ball-mill. The life of the furnace is from 8 to 12 years. The finishing-roast is carried on in a two-hearth furnace. Figs. 454-455, of similar general construction as the rough-roasting kiln and of the same dimensions; it is, however, fired with producer gas. The step-grate of the Zahn producer is i ft. yf in. by 5 ft. lof in, the height 3 ft. iii in. and 4 ft. 7I in., the flue 4 ft. long; the flame is split to reach the upper and lower hearths; the roaster-gases pass downward into an underground main. The driving shaft LEACHING OF, COPPER 459 makes 1.75 r.p.m. The temperature is 850° C. A charge of 1.75 tons of rough-roasted matte is dropped onto the upper hearth, remains there 1.75 hr., whereby the 45 per cent. CujO are reduced to 12 per cent., and then transferred to the lower hearth where the remaining CuzO is oxidized to CuO in 3 hr., and the Ag2S converted into Ag2S04. The side doors are kept partly open to allow the air to have free access. The sulphatized matte goes to the cooling floor, where it cools in 8 hr. to 70-80° C, to be sifted at this temperature into the leaching vats. The furnace treats in 24 hr. 14 tons of rough-roasted matte and Section on Line C-d _s Figs. 434-455. — Mechanical sulphatizing reverberatory furnace. consumes 1.7 tons of bituminous coal. After every removal of a charge the path of the rabble-arms is reversed. It is essential to have uniformly strong oxidizing conditions and a correct temperature for the control of the formation of Ag2S04. An important factor is the amount of CusO present. At 850° C. an excess of CuaO over 45 per cent, begins to act upon CU2S setting free Cu; further Cu-|-2AgS gives 2Ag-|-Cu2S, and Ag2 is difficult to ;sulpbatize; again CU2O decomposes Ag2S in the dry way' as well as in the wet way: ,Cu20+ Ag2S04 = CuS04-|-CuO-f 2Ag (spangle-reaction). If the temperature is too low, CUSO4 remains undecomposed, Ag2S is imperfectly sulphatized, and the ore upon leaching begins to cake; if it is too high Ag2S04 is decomposed.^ At Argo' 2 per cent, of NaHS04 used to be added to the finishing-roast to assist in the sulphatization of the Ag. iPlattner, " Rostprocesse," p. 141. 2 An interesting side-issue at these works is the utilization of the SO3 set free in the finishing roast. The gases are passed through a long flue charged with the dust, collected in the wet condensation of the blast-furnace gases, which contains ZnO, some ZnS04, PbO, PbS04, etc., and readily absorbs SO3. The dust is worked up at the rate of 2.4 tons per day by treating with H2O and some H2SO4 in an upright stationary cyUndrical vat with mechanical stirrer to which some blue powder is added to precipitate all the Cu and Ag. The leached mud is filter- pressed, and the filtrate of 32° B6. evaporated in a Patzburg vacuum pan furnishing granular ZnS04-crystals used in the manufacture of€thopone; the residue is sold to lead works. ' Eng. Min. J., 1890, xlk, 203. 460 METALLURGY OF COPPER The progress of the roast is followed by the appearances of the solution ob- tained in leaching samples with HzO. At first, the solution will be greenish from FeS04, later blue from CUSO4; NaCl will give a slight curdy precipitate of AgCl, and its amount will increase slowly as the formation of Ag2S04 pro- gresses; bright spangles of Ag separating from the solution show that CuzO Eront Elevation Section, on. Line a-fi lieaohlng Tats Pjcecipitatlng VatB Fig.456 Fig.457 Fig.458 M. 1:100 1 2 3 4 B 6 7 S 9 10 m. Figs. 456-458. — ^Leaching and precipitating division. is decomposing Ag2S04. The roast is finished when the blue color caused by CUSO4 has nearly disappeared, the spangle reaction has ceased, and NaCl gives a heavy precipitate of AgCl. (3) Leaching. — The sulphatized matte is leached in wooden vats with false bottoms and duck filters. Figs. 456-458. A vat is 40 in. in diameter, 31.5 and 35.4 in. deep, receives a sifted charge of 0.7 ton, is leached in 4 hr. with hot Plan ^j*Tjijjy mm7\~rr~r.. Scale 5 10 m. tlll l lll I I I I I 1 1 I I I Figs. 459-461. — Auxiliary mechanical sulphatizing reverberatory roasting furnace. H2O (70-80° C.) and then with hot desilverized liquor that has been acidulated. The silver has been extracted when a bright Cu plate is not tarnished and NaCl ceases to produce any turbidity. For i kg. Ag recovered there is used i liter H2SO4 of 35° Be. The leached residue retains 0.025-0.035 per cent. (7-10 oz. per ton) Ag. An analysis' of the residue showed Cu 73.93, Ag 0.0268, Pb 1 Die Mansfeld'sche Kupferschiefer Bauende Gewerkschaft," 1907, p. 147. LEACHING OF COPPER 461 0.SS7. Ni 0.436, Co 0.140, sulphate-S 0.466, sulphide-S 0.062, total-S 0.528 per cent. In order to reduce further the Ag-content, the leached residue is roasted again at 650-700° C. in a two-hearth mechanical reverberatory furnace. Figs. Section on Line U-b Section on Line C-d Fig. 462 Horizontal Section * on Line e-f M. 1:60 l_j l_t l -l M l-L ± i m. Figs. 461-463. — Hydraulic press for cement silver. 459-460, similar to the sulphatizing furnace, in which the driving shaft makes \ r.p.m. The residue retains enough sulphide-S to effect the desired sulpha tization. The furnace is charged with from 3.4 to 3.7 tons of drained residue, remains on the upper hearth 1.75 and on the lower 3 hr., is leached in 0.7-ton charges, in 3 to 4 hr. The final residue now assays Cu 73-74, Ag Section on Line 6 -f Horizontal Section on Line d-b Figs. 464-466. — Drying and calcining furnace for cement silver. 0.016, (4.6 oz. per ton), Pb 0.5-0.8, Ni 0.4, Fe 1.2-2.0, As 0.015-0.020, S 0.5 per cent. It is smelted and refined in a reverberatory furnace (Table 93). (4) Precipitating.— The silver liquor, clarified by traveling through a pair of settling and distributing troughs, or by filtering when necessary, is freed 462 METALLURGY OF COPPER from Ag by percolating through two rows of small (24 in. in diameter, 20 in. deep, with slight taper toward bottom) wooden vessels provided with false bottom and filled with granulated Cu. The upper row of vessels contains corroded granules over which silver liquor has passed for some time. The Cu of the granules has been more or less replaced by Ag, and the bulk of the Ag has been detached by rubbing and then removed. The lower row of vessels contains fresh granules. When the Cu in an upper vessel has been entirely replaced by Section on Lme i~k Fig.467 Figs. 467-468. — Crucible-furnace for melting cement silver. Ag, the vessel is removed, the lower one transferred to the upper shelf, and another vessel filled with fresh granules placed on the lower. (5) Further Treatment of Silver.— The cement silver, about 0.996 fine, is washed, compacted in a hydrauhc press, Figs. 461-463, into cakes 7.78 in. in diameter, dried and calcined in hon boxes. Figs. 464-466, and melted in a crucible furnace. Figs. 467-468, and cast. The bars contain Ag 99.920- 99.950, Cu 0.02-0.04, Pb 0.004-0.008 per cent. In 1906 a ton of white metal (Cu 73-74 per cent.) yielded 56.5 oz. Ag; the extraction was 92 per cent; the cost of treatment was about $7 per ton of matte. LEACHING OF COPPER 463 239. The Freiberg Vitriolization Process, i— The aim of the process is to dissolve with hot dilute H2SO4 the CuO from high-grade dead-roasted argentif- erous matte and convert it into marketable blue vitriol; the insoluble silver- bearing residue is added to a lead blast-furnace charge. The process is based upon the solubility of CuO and the relative insolubility of Ag(Au) in dilute HjS04. If dead-roasted copper matte is treated with hot dilute H2SO4, CuO and ZnO will first go into solution, then follow Fe203, NiO and CoO, and to a very small extent Ag. Any CuaO present is decomposed, Cu20-FH2S04 = Cu-f- CUSO4+H2O. The Cu will precipitate some Ag that may have been dissolved, but the dead-roast is usually prolonged sufficiently to convert all CU2O into CuO. The PbO present is changed into PbS04. Arsenates and antimonates will be partly decomposed, the former being more soluble than the latter. The residue will contain Ag(Au), PbS04 and other insoluble sulphates. The process was developed at Freiberg, Saxony; it has been replaced there by the vitriolization of metallic Cu; in the United States' it is in operation at the Selby Lead Works, San Francisco, Cal., where blue vitriol is produced from matte. (i) Matte. — The matte to be suited for the process ought not to contain more than 0.2 per cent. Fe, as the FeS04 formed would crystallize with the CUSO4 and decrease its market value. However, Gin'' has invented a process for the separation of the two sul- phates. It is based upon the fact that the solubility of CUSO4 is at its maxi- mum at 100° C. and diminishes very slowly with a rise of temperature, while that of FeSOi reaches its maximum a little below 100° C. and then falls off very quickly being practically nil at 160° C. If a mixed moderately concen- trated (350-400 g. CuS04-|-aq. : i liter) solution, kept at nearly boiling tempera- ture, is pumped into a tubular boiler, having an active circulation, and heated there to 180° C. ( = about 200 lb. steam pressure), the FeS04 will sepa- rate. The CUSO4 liquor with suspended FeS04 crystals is forced from the boiler through a filter-press, with chambers kept hot by steam at 220 lb., to separate the crystals from the liquor. While the process is interesting, it seems more rational to scorify the Fe in the reverberatory smelting necessary to obtain a matte rich in Cu, than to cut short the dry process and extend the wet with its inherent complications. An analysis of Freiberg matte by ScherteP gave Pb 4.85, Ag 0.31, Cu 73.95, ' Kuhlemann, Zl. Berg. Hutten. Salin. Wesen i. P., 1871, xrx, 180; Berg. IlUUenm. Z., 1872, XXXI, 76. Capacci, Rev. Un. Min., 1881, ix, 276. Howe, "Production Gold and Silver in the U. S.," 1883, 790. Doerr, Min. Ind., 1896, v, 225. Gignoux (Lyon Mill, Dayton, Nev.), Min. Res. U. S., 1882, 297. Rickard, Min. Ind., 1908, xvii, 588, Selby Lead Works, San Francisco, Cal. ^FiSth Internal. Congr. Appl. Chemistry, Berlin, 1903, i, 597; Eng. Min. J., 1903, "ovi, 358. 'Hofman, "Lead," 1898, p. 373. 464 METALLURGY OF COFFER Bi 0.02, Sb 0.06, As 0.18, Fe 0.13, NiCo 0.21, S 18.98. The matte usually assays Cu 70-75 per cent. Fe < 0.20 per cent. (2) Crushing and Roasting. — The matte is stamped dry to pass a 6-mesh sieve, and dead-roasted in a two-hearth hand reverberatory furnace receiving charges of 3520 lb., which pass through it in about 16 hr. The roasted matte retains about i per cent. S, is ground in a buhrstone mill, and sifted through an 80-mesh screen; the oversize is returned to the mill. (3) Leaching. — The leaching-vats are hard-lead cylindrical vessels, 0.78 in. thick, 5 and 4 ft. in diameter, and 4 ft. high; they have a side discharge near the bottom. A vat is charged with 440 lb. H2SO4 of 50° Be. and 880 lb. mother liquor. This solvent is heated with superheated live steam, and 360 lb. roasted matte stirred in in 2 hr. When the solution, lasting 5 hr., is finished, the steam is shut off, the concentrated liquor diluted to 32° Be., the residue allowed to settle for i hr., and the Cu-Iiquor siphoned into a clarifying vat to remain i hr., and thence into wooden crystallizing tanks of about 90 cu. ft. capacity, lined with soft lead, 0.2 in. thick for sides, 0.4 in. for bottom. Here the CuS04-|-aq. crystallizes out in from 10 to 21 days on suspended lead strips I in. wide and 0.2 in. thick, as well as on the sides and the bottoms of the tanks. The mother liquor serves to dilute chamber acid. The bottom and side crystals, imperfectly developed, are likely to retain some insoluble resi- due; they are redissolved in H2O and mother liquor from the second cry- stallization to form a solution of 32-34° Be. The crystals of blue vitriol con- tain Fe 0.05-0.06, Ni 0.006, Zn 0.003, ^b 0.004 per cent. The residue in the solution vat is removed, settled, drawn off into a conical lead-lined wooden box, treated with H2SO4 and steam, filtered, washed, and then added to the charge of a lead blast-furnace; it forms 15 per cent, of the weight of the matte and assays, Ag 1.3-2 per cent., Cu 3-8, Pb 40-50 per cent. Mother liquor becoming too heavily charged with FeS04 to furnish clean blue vitriol is treated with Fe to precipitate cement Cu. Of the Cu charged there is recovered 96.6 per cent, in blue vitriol, 0.9 in residue, o.i in flue-dust, 0.15 in drosses, 1.3 in cement copper, 0.15 in the pre- cipitating iron, total 99.2 per cent. 240. The Hofmann Vitriolization Process.' — The process developed and put into operation at the works of the Kansas City Smelting and Refining Co., Argentine, Kan. (now dismantled), resembles in its general featxures the Frei- berg vitriolization in that CuO is extracted from roasted ore by means of H2SO4 and crystallized as CuS04-)-aq.; it differs from it in that any Fe present is pre- cipitated as Fe203, and this permits the use of matte rich in Fe as raw material. The process is of sufficient importance to warrant a full discussion. (i) The Matte. — The raw material is the leady copper matte usually produced by the concentration in the blast-furnace of the matte formed in ' Hofmann, O., Min. Ind., 1899, viii, i8g; 1900, rx, 222; 1901, x, 230; " Hydrometallurgy of Silver," McGraw-Hill Book Co., New York, 1907, p. 259. Hesse, "Works at Predazzo," Metallurgies 1909, vi, 580, drawings. LE AGEING OF GOPPER 465 smelting copper-bearing lead ores; it contains Cu 36, Pb 13, Fe, etc., 30, S 21 per cent. (2) Roasting.— The matte is crushed in a rock breaker, and pulverized in a 100-ton Krupp ball-mill to pass a 50-mesh screen. The pulp is roasted in a Hard Maple /jt Bcdiine: Blocks Fig. 469. — Hofmann lixiviation vat with mechanical stirrer. two-hearth Pearce turret furnace. Figs. 151-152, 40 ft. in diameter, having three fire-places for the lower and two for the upper hearth. The roast is carried on in such a way as to convert all the FeS into Fe203 and about 75 per 30 466 METALLURGY OF COPPER cent, of the Cu into CuO. This is accomplished by keeping the temperature low on the upper hearth, and at a cherry red (700-750° C.) on the lower. The roasting matte at first looks a bright red from the oxidation of the sulphide and becomes darker later on. It is then necessary to urge the fire, but- there is danger of overheating, which causes the yield in soluble Cu to fall off. As the formation of some lumps cannot be avoided, the roasted matte is passed again through a Krupp ball-mill with a 50-mesh screen. (3) Leaching. — ^This is carried on in the stationary tank, 12 ft. in diameter and 6 ft. high, with mechanical stirrer, shown in Fig. 469.1 The tank is of wood, has a central cone-shaped projection fastened to the bottom, which forces the matte toward the periphery, where it is held in suspension by the swiftly moving solvent. ^ The interior of the projection is filled with sand. The stirrer, a suspended vertical shaft with two horizontal propeller-shaped arms, is driven by gearing. The rim of the tank has a wooden ring to prevent splashing of the liquor. In the side near the bottom are two discharge pipes of which usually the lower one only is used. The tank is carried by a frame- work standing in a lead pan, intended to catch any leakage, which rests on a floor carried by sills. In leaching, the tank is filled two-thirds with H2O, the stirrer started, 3 per cent. H2SO4 added, and live steam introduced. The matte is fed in gradu- ally from a dump-car with covered discharge provided with a slot; at the same time H2SO4 is run in at a rate to maintain the original standard of 3 per cent, free H2SO4 in the liquor. Working in this way with little free acid causes little Fe203 to be dissolved, the solution showing only 0.7-1.0 per cent. Fe. When the liquor measures 20-22° Be., the flow of acid is shut off, and matte added in small amounts at a time until the free acid has been neutralized. (4) Refining of Copper-liquor. — With the stirrer in motion the finished charge is drawn off into an upright pressure tank and forced with 40-50 lb. pressure through a filter-press. The cast-iron pressure tank is shown in Figs. 470-472; it is lined with lead, and the latter protected from wear by wood.' In filling the tank, some compressed air is admitted with the pulp in order to prevent the latter from packing. The filter press has hard- wood frames and plates 4 ft. sq., is 25 ft. long, and holds 5 tons of residue. The filtrate flows into a collection vat, from which it is elevated by means of a pressure-tank to the top of the refining tower, shown in Fig. 473, for the precipitation of Fe203, As, Sb, Bi, Ni, Co This purification is accomplished by adding CuO (really roasted copper matte) to the hot neutral solution of CUSO4, through which is forced at the same time finely-divided air. The main reaction taking place may be expressed by FeS04-|- dissolved impurity -fO-|- CuO = Fe203+ ^ Min. Ind., 1901, x, 232. ^At the works of Predazzo, it was found necessary to force compressed air through a perforated pipe placed at the periphery to prevent the roasted ore from packing. See also Megraw, Eng. Min. J., 1912, xciv, 360. 'The tank used at Argentine had two 4rft, cylindrical sections instead of the single one shown in the figure. LEACHING OF COPPER 467 precipitated impurity+2CuS04; some basic ferric and cupric salts are formed which remain in tlie residue. The latter is treated with dilute (2.5-3 Per cent.) H2SO4, which dissolves only the Cu, and is filtered. The filtrate goes to the collection vat for the refining tower, the residue is worked with the residue of the solution tank. The refining tower, Fig. 473, is built of 4-in. staves of Cali- fornia redwood well bound by iron rods; it stands on a trestle and carries timbers anchored to the foundation by heavy guide-rods to guard agamst oscillation likely to be caused by the compressed air. The 4-in. air-inlet pipe is made of Plan showinB Size of Openinss ^1®!^ @ Sectios-on Line A B Plan of Supporting Frame Figs. 470-472. — Cast-iron pressure-tank. lead. Its horizontal arm enters the tower 18 in. above the bottom and is con- nected with a radial 6-in. lead pipe closed at the opposite end and perforated on the lower side. The vertical arm reaches to the top of the tower where it is joined through a .valve to an iron pipe reaching down to the receiver of the air- compressor. This arrangement prevents the solution from running into the compressor when the latter is not in operation. Opposite the air-inlet is a r-in. steam pipe held in the cast-iron door of the manhole; at right angles to it is the 4-in. discharge pipe provided with a hard-lead valve. On the top of the tower are a 4-in. inlet for solution and an 8-in. outlet for steam and air (not shown). The latter enters a lead-lined box with zigzag shelves to precipitate and carry 468 METALLURGY OF COPPER down particles of liquor entrained by the air. The upper third of the vat contains glass gauges to watch the filling. i TtlTfl ban Cut Iton Tl&ngu Bubtwt OoBkst ffiiMyssMA*v)?ii?-;?ASiSia/^:ss;a( Fig. 473. — Refining tower. In operating, the tower is charged with 5000 gal. liquor, steam is turned on as well as some air. The latter makes the heating proceed uniformly and causes LEACHING OF COPPER 469 470 METALLURGY OF COPPER some basic ferric sulphate (not over 50 per cent, of the Fe present) to fall out. When the temperature of the liquor has reached 70-80° C, more air is admitted, and some roasted matte fed. After from 3 to 4 hr., all the impurities will have been precipitated. The progress of the precipitation is followed by testing for Fe samples taken from a cock in the side wall. As soon as the solution is freed from Fe, all the other impiu-ities will have been eliminated, as they fall out of solution before the Fe. (5) Evaporation and Crystallization. — The refined Cu-liquor, of 24- 26° Be., and free from Ag (any dissolved Ag2S04 having been precipitated by FeS04), goes to storage tanks. These supply the evaporators, in which the gases from a fire-place pass through tubes immersed in a flat wooden lead-lined tank, 65 ft. long by 12 ft. wide by 2 ft. deep, holding the Cu-liquor; the liquor travels in a direction opposite to that of the gases and the water vapor generated is drawn off by a suction-fan. The evaporator is shown in Figs. 474-475- In operating, the pan is filled to the level of discharge pipe W, the fire started, and fan P set in motion; the level of the liquor is kept constant by occasional feeding from T. When the liquor near W is found to have reached the desired density of about 30° Be., weak liquor is turned on from T, and the amount to be fed regulated by hydrometer tests at W. The liquor from W flows into hori- zontal pressure tanks and is forced into covered troughs of California redwood, loXii in., for distribution to the crystallizing tanks. The troughs are in i6-ft. sections, which are butted together, the joints being made tight by rubber gaskets, iron flanges, and bolts. The boards of a section are joined with brass screws; the joints in it are made tight by a cement prepared by boUing together waste rubber, resin, linseed-oil, and ferric oxide. The crystallization plant is shown in Fig. 476. There are two rows of tanks between which is a traveling rotable belt elevator with copper cups, which raises the crystals, shoveled into the boot, and delivers them to the hopper. Here they drain and are discharged into the car below; the mother liquor collects in channels on the sides of the track and flows to a collecting pit. The tanks are built of two g-in. courses of acid-proof brick separated by a 2-in. space filled with a mixture of asphalt and sand. Each tank is 6 ft. deep and has a capacity of 720 cu. ft.; on top is a wooden frame carrying strips of lead 5 ft. long; crystals form on these as well as on the sides and the bottom of the vat. Crystallization requires seven days; when finished, the mother liquor is drawn off through brass tubes into the side-launders and flows to the collecting pit, whence it is elevated by pressure tanks to storage vats, to be used again in the solution tank and refining tower. The frame with strips of lead and adhering crystals is raised by block and tackle; the crystals are knocked off and go with the side- and bottom-crystals to the elevator. The car receiving the crystals delivers them to a bin, whence they are fed to a crusher consisting of a fast-moving roll and a toothed station- ary plate. The broken crystals are transferred to an inclined trough, washed with water, and sized in two hexagonal drums with brass shaft and arms and LEACHING OF COPPER 471 maple sides having openings 0.375 and 0.125 in. in diameter. The undersize crystals with the wash-water go to dissolving tanks, the oversize are dried in a brass centrifugal machine. The crystals from the neutral solution retam their bluish color longer than do those from a slightly add solution. In order to reduce as much as possible the formation of small crystals, a tank filled with concentrated copper hquor is covered with a layer of water spread about i in. thick from a flat nozzle. This prevents the formation of small crystals on the surface {salting om<) which sink to the bottom as soon as formed, a phenomenon caused by evaporation of the liquor on the surface. Electric Motor --.r^"- Fig. 476. — Crystallization Plant. K, frame with turntable A, held by pin P, and circular track; E, belt elevator swinging on shaft F; L and M, pulleys for driving elevator pulley N; B, boot of elevator. The plant at Argentine with a daily capacity of 60 tons of blue vitriol had: 3 Pearce furnaces, 8 solution tanks, i pressure tank, 5 filter-presses with storage tanks, 8 refining towers, 11 evaporators for 90,000 gal. refined copper liquor per day, and 112 crystallizing vats, each of 720 cu. ft. capacity. C. Leaching Metallic Copper 241. Leaching of Metallic Copper in General. — The leaching of metallic Cu with H2SO4 has many points in common with the similar treatment of copper matte. Leaching Cu has been replaced to a considerable degree by electrolysis, at least with pure metal. Two processes have to be considered, the obsolete Augustin, and the Harz vitriolization which has retained its place as an independent process with impure Cu, and as an auxiliary process in the preparation of the blue-vitriol electrolyte in the electrolytic refining of Cu (§ 253, 264). 472 METALLURGY OF COPPER 242. The Augustin Process.' — The underlying principles are the' same as those for the treatment of copper matte, except that metallic copper is subjected to a chloridizing roast. The process was put into operation in 1885 by Klock and Hartmann in St. Louis, Mo., but details are lacking. The best example is probably that of Tajova^ where the process was replaced in 1893 by electrolysis. Only the chloridizing roasting of metallic copper is of interest at present. Two grades of black copper were treated, containing: Cu 80-84, Ag 0.30-0.36, Sb 3-7 per cent., and Cu 70-80, Ag 0.20-0.25, Pb 9-15 per cent. The black copper tapped from the blast-furnace is ladled onto cast-iron plates, pulverized during solidification by rubbing with wooden hammers, and sifted through two screens. Of the three sizes obtained, the fine is ready for the furnace, the medium is ground in a buhrstone mill with 3-in. cast-iron grinding faces,and the coarse is first dry-stamped and then ground. The roasting-charge is made up to assay not over 7 per cent. Pb nor 0.4 per cent. Ag; it receives 15 per cent. salt. A charge of 500 lb. is dropped on to the upper hearth of a small two-hearth reverberatory furnace, and roasted at a dark-red with continuous rabbHng for 7-10 hr.; it is then transferred to the lower hearth, where the rabbling is con- tinued and 4 per cent, carbonaceous matter added in three portions at half-hour intervals to decompose a:MO.As(Sb)206; the temperature is then raised to a good red heat, and held there for 1.5-2 hr. Firing is now stopped and rabbling con- tinued for 0.5-1 hr. more. The chloridized ore is raked up to a heap, remains untouched for 0.5 hr., and is then withdrawn, having gained 30-33 per cent, in weight. It is sifted while still warm (80° C.) ; any oversize is ground and added to the next charge. The chloridation of the Ag may be expressed by 2 Ag-f- 2NaCl -fC02+0 = 2AgCl-hNa2C03 and Ag-f-CuCl2 = AgCl-|-CuCl. An addition of a small amount of Si02 helps the chloridation. 243. The Vitriolization Process.' — The process in its present form was put into operation in 1858 at the Copper Smeltery of Oker, Harz Mountains, and often goes by the name of Harz vitriolization. It is based upon the solubility of Cu in hot dilute H2SO4 in the presence of air, and the relative insolubility of Ag, Au, Pb, As, Sb, etc. The leading steps are refining and granulating silver- bearing impure metallic Cu, dissolving the Cu granules in hot dilute H2SO4 in the presence of air, separating the Cu-solution from the residue, crystallizing ' Augustin, op. cit. Egleston, Tr. A. I. M. E., 1876, rv, 295. Capacci, Rev. Un. Min., 1881, x, 201. Howe, "Production Gold and Silver in the TJ. S.," 1883, p. 764. ^ Markus, Berg. HiiUenm. Z., 1852, xi, 5; 1855, xiv, 64. Kerpely, op. cit., i87i,xxx, 190, 285. Wagner, Oest. Zt. Berg. Huttenw., 1873, xxi> 319- Balling, C. A. M., "Metallhiittenkunde," Springer, Berlin, 1885, p. 358. ' Kuhlemann, Zt. Berg. Hutten. Salin. Wesen i. P., 1871, xix, 180. Brauning, op. cit., 1877, xxv, 166. Howe, "Production Gold and Silver in the U. S.," 1883, p. 790. Egleston, op. cit., 1884, p. 600. Clement, Min. Ind., 1900, rx, 278. LEACHING OF COPPER 473 the CuS04+aq. and converting it into marketable blue vitriol, and working up the insoluble residue. (i) Refining and Granulating Black Copper.— Black copper contains 90 ± per cent. Cu. The aim in refining is to scorify Pb, Fe, Ni, Co, Zn, etc., so as to prevent their being attacked or dissolved by the acid, as, when crystalliz- ing with the blue vitriol, they would impair its quality. The slagging of Pb, Fe, and Zn is readily accomplished, that of Ni, Co and Bi (§ 193) less so. Accord- ing to Egleston^ the scorification of Ni(Co) is greatest at the period when Cu gives off the last of its S, hence some of the Ni in Cu can be concentrated in a small amount of slag. The ciurve of Wanjukow, Fig. 411, also shows that the elimination of Ni is rapid during the boiling period. A similar observation was made by Kuhlemann.'' The mode of operating is the same as in refining copper (§ 188 and foil.). An analysis of granules from Altenau' gave Cu 95.00, Pb 2.71, Fe 0.07, Ni-Co-Zn 0.048, Sb 1.53, As trace, Ag 0.30 per cent. Granules ought to be flat, 1.2 in. in diameter, and have thin walls 0.02 in. thick; they resemble somewhat flaked breakfast-food; frequently they are rounded. The form of the granule depends upon the pitch of the Cu and the granulation proper. According to Egleston* the Cu ought to be granulated at the end of the boiling period, i.e., when it has absorbed some CU2O, but has not yet reached the stage of set copper. At Oker^ the metal is tapped before it has ceased boiling; at Freiberg^ the same is the case; the reason beiiig that the liberation of SO2 causes the walls to become thin. The progress made in the refining of the black copper is carefully regulated toward the end in order to obtain just the pitch which is correct for granulating, as with the right pitch the granulation is simple and effective, while with a wrong pitch the granules are likely to be spherical and solid, and violent explosions are common. An excess of CU2S in the bath is indicated by films of CU2S flitting over the surface of the metal which has been freed from slag, and by the swelling of the slag when this is being skimmed. The excess is removed by charging small amounts of roasted white metal; or by rabbling or blowing. An excess of CU2O is indicated by the brightness of the surface of the Cu and the quickness with which the skimmed slag solidifies; a stick of sulphur thrown on the bath burns with the evolution of the brownish fumes of S-vapor. The excess of CU2O is removed by charging small amounts of white metal or of stick-sulphur. A granulated sample of Cu of the right pitch is pale red and shows no blackish specks, which indicate CU2S; purplish granules indicate an excess of CU2O. Beside the pitch, the temperature of the copper is of importance; the latter ought to be as low as will permit the metal to run in a thin stream from the tap-hole. The lower the temperature, the more effective is the expulsion of SO2. 'Tr.A.I. M.E., 1882, X, 49. '^Loc. cit., p. 205. 'Kuhlemann, loc. cit., p. 203. • Tr. A. I. M. E., 1875-76, IV, 296. 'Brauning, loc. cit., p. 163. 'Private Notes, 474 METALLURGY OF COPPER For granulating, the copper is run from the furnace in a thin stream into a deep covered water-tank, of wood, iron, or of brick, well cemented, let into the ground and provided with a steady inflow of cold water. On leaving the spout the copper either meets a strong jet of water which scatters the metal and thus assists in the forming of flat granules, or it drops on to a pole of green wood which breaks up the stream; in the latter case the basin is filled with hot water (Frei- berg). The basin must be deep. Granulating with a jet of water requires cold water in order that the granules shall have become solid before they reach the bottom; if this is not the case, there is danger of serious explosions, and of the granules adhering to one another and forming lumps. The granulating basin is always covered, as explosions of more or less violence are always likely to occur. The basin must be deep enough so as not to be more than half filled with granules by a furnace charge. At Oker about 3 tons of copper are granulated in an oval wooden tank 8 ft. 2§ in. by 4 ft. 11 in. and 4 ft. 7 in. deep; at Freiberg about 10 tons in a circular boiler-iron tank 6 ft. 10 in. in diameter and 9 ft 10 in. deep. In the tank is placed a basket connected by chain or wire-rope with an overhead traveling-pulley to remove the granules and transfer them to the solution tank. (2) Dissolving of Granules. — The main reaction taking place in the solu- tion of the copper is Cu-|-H2S04+0 = CuS04+H20; a secondary reaction is 2CuS04+2Cu = 2Cu2S04 and Cu2S04+H2S04+0 = 2CuS04+H20, i.e., some of the CUSO4 formed by the main reaction acts upon Cu and is reduced to CU2SO4, but the latter is oxidized again to CUSO4 in the presence of H2SO4 and O. Thus the Cu is dissolved by the direct action of H2SO4 and O, and the indirect action of CUSO4. The befiavior of foreign metals with hot dilute H2SO4 is similar to that in treating roasted matte. There will go into solution Cu, Zn, Fe, Ni, Co, and small amounts of AS2O6 and Sb206; the residue will contain Ag, Au, PbS04, most of the AS2O6 and Sb206, some Pb3As208, and basic sulphates of Sb, Sn, and Bi. The manner of operating varies considerably as shown in § 244 by several examples. (3) Crystallization. — Trade demands that the crystals of blue vitriol shall be pure, large, and of a correct color. The conditions are fulfilled by hav- ing a clean and clear solution of the right concentration (28-29° Be.) with not over I per cent, free acid, by a slow crystallization (six to eight days) in covered tanks in which are suspended strips of lead from cross-bars, by freeing the crystals from mother liquor through washing, and by drying the washed crystals. In allowing blue vitriol to crystallize out of solution in the usual way, large crystals form on the lead strips (simUar to rock-candy on threads), smaller crystals on the sides of the tank, and the smallest on the floor. As the last bring only a low price, it is important to hinder their formation as much as pos- sible; which is accomplished by the method of O. Hofmann given in § 240. Usually the bottom-crystals are redissolved and recrystallized. (4) WoRKiNG-up or Insoluble Residue. — This is briquetted and smelted in a suitable furnace for lead bullion or copper matte as the conditions may require. LEACHING OF COPPER 475 244. Examples of Vitriolization. (i) Oker, Haez Mountains.i— Figs. 477-481 represent the leading apparatus in use at Oker. Fig. 477 gives the stor- age tank for acid, 11 ft. 10 in. by 6 ft. 5^ in. and 4 ft. deep, made of wood and lined with 8-10 lb. lead; it has a steam coil for heating, and siphons with roses for drawing off acid. The conical solu- tion-vat, 5 ft. 3! in. high and 2 ft. 11 in. and 2 ft. 4! in. in diameter, also of wood and lined with 16-18 lb. lead, has at the side a discharge-spout, 8 in. square and i ft. 3 in. long. The false bottom is made of a heavy sheet of per- forated lead carrying two blocks of wood, 6 in. high, boiled in oil, upon which rest bars of copper. The vat receives about i ton of granules, which fill it to a height of about 3 ft. 3 in. The solvent, H2SO4 of 29-30° Be. heated to 70° C, is sprayed on the copper for S min. and then shut off; now air is allowed to ascend through the warmed and acidulated copper for |-f hr. to Fig. 477. — Acid-tank and solution-vat. -63'«^^- DrainluK Table "^r \i Ejector ^ Tank 1 it J_ / Draining Table 31 31 -41— Fig. 478. — Solution-vats and crystallizing-troughs. furnish the necessary for the formation of CUSO4; acid is again admitted 'Brauning, Zt. Berg. Hutten. Salin. Wesen i. P., 1877, xxv, 165. Egleston, "Production Gold and Silver in the U. S.," 1884, p. 600. Private Communication, 1913. 476 METALLURGY OF COPPER for 5 min., in which time the CUSO4 will be washed down with the insoluble residue and will flow off through the discharge; the solution at first blue soon becomes colorless. The operations are repeated, and the dissolved granules replenished so as to keep the vat filled. A clean-up is made at intervals when the insoluble residue has choked up the open spaces between the copper bars, or when the bars have to be renewed. Six vats will handle about 600-800 lb. granules in 24 hr. An average analysis of the granules treated in 1902 gave: Cu 91.47, Ag 0.462, Au 0.0154, Pb 0.887, Bi 0.286, As 3.863, Sb 2.133, Fe 0.028, Ni-Co 0.750. From a unit of six dissolving vats. Fig. 478, the copper solution and insoluble residue pass through a series of zigzag launders, 347 ft. 4! in. long, in which the insoluble part sinks to the bottom and blue vitriol crystallizes owing to the cooling effect of the air. Most of the mud sinks to the bottom in the head launder; the vitriol collecting near the tail-end is rich in CaSOi, PbsAsaOs, PbsSbaOs. The mother liquor flows from the last launder into a collec- tion tank, 7 ft. 105 in. square, whence it is raised by an injector to the storage tank for" solvent, to be brought to the right concentra- tion by addition of chamber acid (50° Be.). In 1904 a liter mother liquor (specific gravity 1.7 19) contained in g :Cu 15.000, Ni 22.980, Fe 3.052^ 1 jj|-- I0'».)ii'-^g c 1 ' 1 ' 1 1 1 1 1 1 1 1 1 1 1 III I ' I i 1 1 1 1 1 1 1 1 1 1 III 1 ' 1 ' 1 ' 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 III r 1 1 1 1 1 1 1 Fig. 479. — Vitriolization at Oker, Germany. Figs. 480-48 1 . — Concentration-pan. free H2SO4 203.880. A cross-section through a launder is given in Fig. 479: two wooden lead-lined launders are separated by a gable-shaped division covered with sheet lead. The crude vitriol is shoveled out once a week onto the sloping planes of the division from which the mother liquor flows back into the launder. The vitriol is washed superficially at intervals with clean water to remove adhering acid mother liquor. The crude vitriol, which is unsightly and encloses insoluble residue as well as mother liquor, has to be purified. It is removed to a lead pan shown in Figs. 480-481, dissolved in weak mother liquor, and the mud allowed to settle. The pan, II ft. 6 in. long by 10 ft. io| in. wide and i ft. 11 in. deep, is of f-in. lead; it rests on |-in. cast-iron plates, beneath which are heating-flues for the passage of the gases from the fire-place. The pan is charged to a depth of i ft. 3I in. with mother liquor and water, to density of 18-19° ^^., and heated to 70° C; crude LEACHING OF COPPER 477 vitriol is stirred in and dissolved in about i hr. until the liquor has been raised to 28° Be. Some fine sheet copper used to be added to the pan-charge to precipitate traces of Ag that might have gone into solution. At present an emulsion of ground galena is added for this purpose. The fire is withdrawn from the grate, the pan is covered with boards to retard the cooling, and the mud allowed to settle 6 to 8 hr. The clear copper liquor, still at 50° C, is with- drawn through the upper tap, b, into wooden lead-lined crystallizing tanks, 9 ft. 6 in. long by 4 ft. 7 in. wide and 3 ft. '3 in. deep, in which are suspended strips of lead, 0.15 in. thick and 0.78 in. wide, from 25 cross-bars, five to a bar, reaching to near the bottom. The crystallization, which must be slow if well- developed crystals are to form, lasts from 8 to 12 days.' The largest crystals form on the strips, then follow the side-, and lastly the bottom-crystals. Strip- and side-crystals are washed with water to remove adhering mother liquor, drained on inclined tables (better in a centrifugal machine), dried for eight days in a dark room on shelves 6 ft. 6 in. wide at a temperature ranging from 20 to 25° C. (above 25° C. the transparent crystal becomes opaque), sorted according to size, and barreled. Bottom crystals are recrystallized. An analysis of blue vitriol of Altenau, where the mode of operating is the same as at Oker, gave CuS04+aq, 99.9700, Fe 0.0107, Sb 0.0123, As 0.0064, Ag 0.0006, Zn trace, Ni trace.^ The blue vitriol produced in 1900 is less pure; it analyzed: Cu 24.721, Fe 0.054, Ni 0.255, Co trace. As 0.068, Sb trace, Bi none, Ag 0.0015, free H2SO4 0.913 per cent. The yield of Cu in the form of blue vitriol is 99.76 per cent. ; that of Ag is 99.4 per cent. The mother liquor from the crystallizing tanks serves to dissolve crude vitriol. The residue in the concen- tration pan is withdrawn at intervals through the lower tap, c, which is flush with the bottom, washed with water, filter-pressed, dried, and smelted with a lead charge. The residue produced in 1901, forming 17.5 per cent, of the weight of the granules, averaged Ag 4.010, Au 0.1066, Cu 1.38 (1.15 as CuSO^, Pb 16.08, Bi 8.02, As 17.06, Sb 19.71, Fe 0.23, NiCo 0.45, SO3 10.28, H2O 0.95, Insol. 2.62 per cent. The mother liquors are used over and over until they become so heavily charged with Fe and Ni that they interfere with the purity of ibe blue vitriol; they are then worked for Cu and Ni. A unit of 6 solution vats, 2 concentration pans, and 24 crystallizing tanks produces per day 3000 lb. blue vitriol; 100 lb. Cu give 380 lb. CuS04+aq, require 160 lb. H2SO4 of 50° Be., and about 310 lb. coal.' (2) Freiberg, Saxony.^ — The solution vessel is a lead-lined wooden tank, 4 ft. rrj in. square and 5 ft. 11 in. high. On the true bottom is placed an inclined floor of gypsum, and this covered with sheet lead with edge turned up and beaten against the sides. The bottom is 5^ in. thick on one side and tapers to the discharge opposite, to facilitate the removal of insoluble residue. The vessel has a false bottom of perforated (0.4-in. conical holes) sheet lead, 73 in. above the true bottom, supported by posts. The discharge is in the side on 'See Hofmann's addition of water, § 240. 'Kuhlemann, loc. cit., p. 213. 'Private Notes. 478 METALLURGY OF COPPER the true bottom. A steam injector under the false bottom furnishes the air and heat necessary for the solution of the copper; the H2SO4 of 12° Be. and 60-75° C. is sprayed on the granules by a rotary feeder working upon the reaction-wheel principle. This Haege wheel has a small glass disc on the lower side and rotates on a needle attached to a lead-coveredironbar placed across the vat. The vat holds 5 tons granules assaying Cu 97-98 and Ag 0.3-0.4 per cent. An analysis of 1905 gave: Au 0.0051, Ag 0.38, Cu 97.78, Pb 0.27, Bi 0.14, Fe 0.12, Ni 0.09, Co 0.03, As 0.33, Sb 0.47, Sil trace, Insol. 0.12, S o.io, O 0.13 per cent. According to the desired rate of working, governed in part by the concen- tration of the acid, there are treated in a vat from 1650 to 3300 lb. granules in 24 hr.; the acid correspondingly contains from no to 400 g. H2SO4 per liter ( = 12-28° Be.). The copper liquor produced from 37-200 g. H2SO4 has a density of 32-35° Be. and averages 35-40 g. free H2SO4 in a liter. The copper liquor flows into settling tanks having partition walls to retard the current and to force it to pass through and not simply along the top of the tank. The clarified liquor is crystallized; the crystals are freed from mother liquor, washed and dried in a centrifugal machine; they contain 0.025 Fe, 0.012 Ni, from o to o.io per cent, free H2SO4, and are free from As and Sb. The mother liquor contains in a liter 1.026 g. Fe, 2.416 NiCo, 0.035 Bi, 5.87 As, 0.529 Sb, 0.049 g- Pb. The insoluble residue is drained, washed with hot water, filter-pressed, treated with water in the press until the effluent measures 2-3° Be., and dried. It assays Ag 5-6.5, Pb 7, and Cu 40-65 per cent. It is roasted sulphatizingly in a muffle furnace until a leaching test shows that the water-soluble CUSO4 has been removed from the residue down to 0.2-0.3 P^r cent. Cu. The roasted residue is treated with hot dilute H2SO4 in the presence of sheet copper to decompose all Ag2S04 that may have been dissolved. The resulting second residue, after washing and drying, assays Au 0.28, Ag 22.5, Pb 28, Cu 0.2-1.0 and Bi o.i-io.o per cent.; and is charged on the lead bath of a cupelling furnace. Of the Cu charged there is recovered in the blue vitriol 99.85, in the residue 0.05, in the dross o.io per cent. There are produced per day about 30,000 lb. blue vitriol. The plant consists of the following parts: (i) Solution of Granules: 3 acid tanks, 4 solution vessels, 2 small and 4 large clarifying tanks, 6 collecting tanks for liquor of 32° Be. (2) Retreatment of Vitriol: 2 solution tanks, 3 clari- fying tanks, I residue tank. (3) Treatment of Residue: 2 acid tanks, i solution vessel, 2 drying boxes, i filter-press, i muffle furnace for roasting, i collecting tank for residue, i collecting tank for liquor, i filtering box. (4) Crystallization: 150 tanks. (5) Drying of Blue Vitriol: 2 centrifugal machines, i tank for liquor. (6) Additional: i compressor, acid-eggs, collecting tanks for mother liquors and acid. (3) Eastern United States. 1— The solution tank ("oxidizer," "dissolver") for granulated copper is usually 3 ft. by 3 ft. 3 in. in diameter and 6 ft. 6 in. high; it is of wood lined with is-lb. lead, and has an opening i ft. square in the front ^ Clemmer, Min. Ind., 1900, rx, 277. LEACHING OF COPPER 4^^ near the bottom; a false bottom of f-in. iron pipes, covered with heavy lead pipe and closed at the ends by overlapping lead, rests i ft. above the true bottom upon acid-proof brick. The granulated copper filling the tank is covered by an acid-distributor which is a piece of sheet-lead having up-turned edges and two i-in. holes per square inch surface. The acid, of io° Be. and 80° C, is fed by the oscillating trough, 34 in. long and 15 in. wide, shown in Fig. 482. It is made of i-in. boards, the axis F extends 3 in. beyond the trough; G and D are stops. The copper solution flows from the dissolving tank into a receiving vessel and is drculated until it measures 34-35° Be. at 16° C; it is then run into a settling tank and thence into a crystallizing vat. The copper in the solution tank is replenished without shutting off the acid; once a month the solution tank is washed out by water under pressure. The slime collecting under the false bottom and in the settling tank is removed periodically. Six solution tanks K Bib Spigot Add Stoneware Fig. 482. — Oscillating feed-trough. furnish 5000 lb. blue vitriol per day, for which 10 crystallizing vats 10 ft. square and 3 ft. deep are necessary. The crystallizing tanks are of wood, lined with 8-lb. lead on the sides and lo-lb. lead on the bottom; the sides often receive a taper of i in. to prevent buckling of the lead. The mother liquor is drawn from the bottom through an opening having a plug-seat and plug of hard lead (10 per cent. Sb). At the end of the crystallization (which favors layering), it is sometimes advantageous to draw off the top (or lighter) liquor separately from the bottom (or heavier) liquor which is richer in CUSO4, and recover the Cu from the lighter by means of Fe. The crystals from a tank are dried by passing them through a housed revolving brass screen with J-in. holes; the chamber is heated by live steam and is connected with a suction fan to draw off the vapors. The crystals are fed continuously into the screen, the fines are sifted out to be redissolved, and the coarse crystals {> \ in.) are discharged direct into packing barrels. A plant is usually terraced. On the top-level are the acid tanks, lined with 10-lb. lead and provided with steam-coils; next come the dissolving tanks; these are followed by receiving and settling tanks lined with 8-lb. lead; on the fourth level or ground floor are the crystallizing vats in double rows with a 2-ft. passage way, and the wash-floor for crystals, covered with lo-lb. lead. The mother-liquor tanks, lined with 8-lb. lead, and the sump, lined with 6-lb. lead, 48o METALLURGY OF COPPER are let into the ground floor; drainage gutters leading to the sump are lined with 6-lb. lead. The cost of producing 5000 lb. blue vitriol per day near New York City in a plant which manufactures acid and smelts copper was 3.73 cents per pound in 1900; the totals were: Copper and acid $152.63; foreman and labor $11.50; packing $4.80; fuel $9.00; repairs, insurance, depreciation, etc., $6.00; in all $i83-93- (4) Special Forms or Oxidizers in the Eastern United States. — (a) Works A. — The solution vessel is made up of two flanged sections of glazed sewer-pipe, 2 ft. 10 in. high and 3 ft. 4 in. inner diameter. The lower section rests on a heavy flanged lead plate with i|-in. holes, covered with 6 in. of slag serving as a support for the granules. On the upper section is placed a wooden frame supporting a flanged lead plate with f-in. holes. Hot acid is admitted periodically onto the distributing plate through a circular lead pipe, having perforations on the inner side, by mechanically opening and closing a cock. (b) Witherell Down-draft Oxidizer. — A wooden lead-lined tank, 12 ft. in diameter and 6 ft. 6 in. deep, has a radial grate surface, 3 ft. 6 in. down from the top, and in the center a suspended ventilating chimney, 30 in. in diameter. The grate is filled with granules to 6 in. from the top of the tank; hot acid is turned on periodically through two radial horizontal wooden troughs, with staggered holes, moving around the chimney. Steam is admitted at intervals. The copper liquor overflows through a trapped discharge-spout. When the granules have been dissolved down to 12-18 in. from the grate, the suspended chimney is raised, a man enters the tank, shovels over the granules, and removes the mud with a hose. The tank is then recharged. (c) The Walker Dissolving Machine. — In order to hasten the solution of granules, A. L. Walker^ introduced at the Perth Amboy plant of the American Smelting & Refining Co. his mechanical device, by means of which copper granules are alternately immersed in and withdrawn from a tank holding hot dilute H2SO4, fresh acid being fed and copper-liquor removed continuously. The essential parts are: (i) three walking beams, set at 180°, having at one end an arc-shaped piece, at the other a pitman driven from an eccentric; (2) a stout basket of hard-lead, 6X4 ft. and i ft deep, charged with 1 ton of copper granules, carried by a lead-covered copper frame which is suspended by a chain riding on the arc-shaped piece, and immersed vertically in (3) a solu- tion-tank, 15 by 15 ft. and 5 ft. deep, provided with a steam-coil and a ver- tical partition reaching to near the floor, which separates the immersion-division, fed continuously with dilute H2SO4 so as to have 10 per cent, free acid, from the discharge-division provided with an outlet for the continuous outflow of concentrated copper-liquor. The movement of the walking-beam is so appor- tioned that the basket remains half a minute in the air for the purpose of oxi- dation and is immersed for a half minute in the tank for the purpose of solu- tion. The outflowing copper-rich liquor runs through a series of lead-lined lU. S. Patent No. 788862, May 2, 1905. LEACHING OF COPPER 481 launders, 24 in. wide and 6 in. deep with wooden sides and sheet-iron bottom in which small crystals of blue vitriol separate, and enters a second solution tank equipped in the same manner as the first and 10 in. below it. The copper- liquor from the second tank passes through a second series of launders where crystallization again takes place, and enters a third solution-tank whence the copper-liquor passes into a third series of crystallizing launders. As the three solution-tanks are provided with steam-coils and evaporation takes place all the time, the liquor leaving the third series of crystallizing launders is very much reduced in volume. It is fortified with acid to bring the free H2SO4 to the Standard 15 per cent, and returned to the first solution-tank. The small crystals and the insoluble residue settling in the crystallizing launders are re- moved at intervals, charged into ordinary tanks where the crystals are dissolved in hot water and the mud is settled. The clear blue-vitriol solution is run into crystallizing tanks, and the mud carrying precious metal is worked up in a suitable manner. The granulated copper is dissolved six times as fast as by the ordinary method using a steam-injector. The method was aban- doned in 1907, but put into service again later on. 31 CHAPTER IX ELECTROLYSIS OF COPPER 245. In General. — The raw materials from which Cu might be extracted by means of an electrolytic process are. ore, matte, speise, and lastly metal which is to be refined to a higher degree than is practicable by dry methods. 246. Electrolysis of Ore. — All direct as well as most indirect processes have been failures. A direct process is one in which the ore forms the anode. The reasons why such a treatment cannot be successful have been given in § 156. An indirect process is one in which the raw or roasted ore is treated with a solvent, and the solution electrolyzed, not in contact with ore. The methods advocated by Laszcynski (§205), Carmichael (§206), Greenawalt (§207), Siemens-Halske (§ 218), Hoepfner (§222) have already been discussed. 247. Electrolysis of Copper Matte.' — After the electrolytic refining of riietallic copper had proved a commercial success, it seemed, natural to go one step back and attempt the electrolysis of copper matte in a bath of acidulated blue vitriol with cast matte as anode and sheet copper as cathode. This process was patented by Marchese in 1882^ and carried out on a large scale with 30- per-cent. matte at the Casarza works near Sestri-Levante, Genoa,' and proved a failure. Later it was taken up again at Stolberg, Westphalia,'* with matte of 15-16 per cent. Cu, but proved again to be unsuccessful. It was held that the matte would be decomposed in part by electrolysis, in part by the Fea-CSO^s, formed in the bath, and that a potential of about i volt would be suflident for the work. The process failed because the e.m.f. required reached 5 volt.; because the anode became coated with non-conducting matter (S); because decomposition was unequal and disintegration of the anodes followed; because more Cu plated out than went into solution and the electrolyte was enriched in FeS04, which offered a greater resistance than CuS04 and required fre- quent renewal of the bath; because the character of the deposited Cu was infe- rior; and because the deposited Cu was redissolved by the Fe2. (804)3 formed. The remarks in § 205-207, 218 and 222 relating to the processes for treating ore hold good also for matte. In order to make electrolysis of matte in aqueous solution successful, it is necessary that the matte be nearly pure CuzS which leaves behind only a small ^ Borchers-McMillan, op. cit., p. 247. = Rev. Un. Min., 1883, xiv, 331; 1884, xv, 668, 1885, xvii, 563. ' Badia, La lumUre electrique, transl. in Sc. American Supplement, 1885, xrx, 7623, 7646, Nos. 478 and 479; also Berg. Hiittenm, Z., 1885, XLiv, 330; Eng. Min. J., 1885, XL, 21. Zopetti, II Politecnico, Nov. and Dec, 1885, transl. in Rev. Un. Min., 1886, xix, 197; XX, 94; also Berg. Hiittenm. Z., 1886, XLV, 207, 221, 538. * Cohen, Berg. Hiittenm. Z., 1888, XLVii, 406; 1894, liii, 328; Zt. Electrochemie, 1894, p. 50- 482 ELECTROLYSIS OF COPPER 483 amount of residue (S), and at the same time will not contaminate the electrolyte with inpurities (Fe). This has been done at Mansfeld, Germany, with the Borchers-Franke-Giinther process.^ The matte with Cu 72-76 per cent, is cast into anodes, 40 in. square by 2 in. thick, which have to be annealed to pre- vent their cracking. They are suspended in parallel alternating with Cu cathode sheets in the usual wooden lead-lined electrolyzing vat for metallic copper, by T-shaped pieces of Cu imbedded in the matte and tinned at the ends. The electrolyte is an acidulated solution of blue vitriol held at 70° C, circulated and aerated; the current density 7 amp. per square foot at a pressure of 0.75 volt. The Cu is dissolved and goes to the cathode, S and other insoluble matter go into the anode mud; if the current density is too low, only half of the Cu goes into solution. The anode has to be removed before it becomes too thin and begins to crumble. The mud is filter-pressed, treated with a hot solution of acetylene tetrachloride to dissolve the S, which separates upon cooling, and then worked by the Ziervogel process (§ 237) to recover the Ag. While metallurgically the process was successful, the cost of treatment was too high; the process was there- fore abandoned. 248. Electrolysis of Speise. — Considering that all attempts at the elec- trolysis of ordinary copper matte have been unsuccessful, it is not to be expected that a similar treatment of speise will be anything else but a failure. A process was suggested by Andre in 1877," but there is no record of any- thing more. 249. Electrolysis of Metallic Copper in General.' — The aim of the elec- trolytic refining of copper is to produce pure copper from a high-grade blister copper by means of the selective chemical action of the electric current. The current is intended to dissolve copper from the blister copper suspended as anode in a solution of CUSO4 acidulated with H2SO4, and to deposit it on a copper cathode. Most of the usual impurities will remain behind and form the anode mud, some will be dissolved and become concentrated in the electrolyte, some may be deposited on the cathode. The possibility of such a process was proved experimentally in 1847 by von Leuchtenberg, but its industrial applicability was first recognized by 'Wagner-Primrose, Eng. Min. J., 1907, lxxxiv, 673. Editor, Melallurgie, 1908, v, 29. ' Dingl. Polylech. J., 1879, ccxxx, 281; 1880, ccxxxvi, 413. 'T. Ulke, "Modern Electrolytic Copper Refining," Wiley, New York, 1903. W. Borchers and W. G. McMillan, "Electric Smelting and Refining," Lippincott, Philadelphia, 1904, p. 187-245. J. Billiter, "Die Electrochemischen Verfahren der Chemischen Grossindustrie." Knapp, Halle, 1909, l, pp. 37-139. Keller, Min. Ind., 1898, vil, 229. Crocker-Arendt, School Mines Quart., 1903, xxv, 3. Bancroft, Tr. Am. Electrochem. Soc, 1903, iv, 175; Electra- chem. Ind., 1902-03, i, 484, 584; Mines and Minerals, i903,xxiv, 182, 229; Eng. Min. J., 1903, Kxvi, 740; Melallurgie, 1904, l, 670. Schwab-Baum, Jl.Phys. Chem., I903,vn,493. Burgess, Tr. Am. Electrochem. Soc, 1905, vii, 51; Electrochem. Metal. Ind., 1905,™, 173- Addicks, Jl. Franklin Inst., 1905, Clx, 421; Min. Sc. Press, 1906, xcil, 38. Bennett, Electrodeposition (plating) of Copper, Tr. Am. Electrochem. Soc, 1913, xxni, 233; Met. Chem. Eng., 1913, H, 284. Burns, Tr. A. I. M. E., 1913, xlvi; (Great Falls Plant); Discussion, Met. Chem. Eng., 1913, XI, 670 (Motherwell, Bums). 484 METALLURGY OF COPPER Elkington who patented in 1865 the multiple system of the process, and erected the first successful plant in 1869 at Pembrey, Wales. His example was soon followed by others in England, Germany, and elsewhere. In the United States the Balbach Smelting 81 Refining Co.^ in 1883 was the pioneer of the industry, which has grown to such a degree that in 1912, with the United States furnishing over one-half of the world's copper, 81 per cent, of its product was electrolytic copper. The second form of the process in operation to-day is the Hayden or Series System introduced in 1886. Table 115^ gives the leading works of the United States and their capacities. Table 115. — Electrolytic Copper Repineries of THE United States Works Location Process Capacity million pounds Nichols Copper Co Raritan Copper Works Baltimore Copper Smelting & Rolling . . Co. American Smelting & Refining Co U. S. Metals Refining Co Laurel Hill, N. Y Perth Amboy, N. J . . Canton, Md Maurer, N. J Chrome, N. J Newark, N.J Great Falls, Mont.... Tacoma, Wash Buffalo, N.Y Series Multiple. . . . Series and multiple. Multiple. . . . Multiple Multiple. . . . Multiple .... Multiple. . . . Multiple. . . . 330' 360' 312' 192'' 180' Balbach Smelting & Refining Co 48^ 65' Tacoma Smelting Co ^6^ 5S' When CUSO4 is dissolved in acidulated water, it is in part dissociated into Cu" and SO4" ions. If the solution is electrolyzed using copper electrodes, the SO4" ions migrate to the anode, are deposited, give up their charges to the anode, and combine with an equivalent amount of copper. Similarly, the Cu" ions migrate to the cathode, give up their charges to the cathode, and are deposited as metallic copper. Another way of expressing the same idea is that the Cu at the anode receives two positive charges (Cu+2-|- = Cu"), is converted into Cu", and goes into the solution; at the cathode the two positive charges of Cu" are neutralized by two negative charges of the current (Cu"+2— =Cu), and Cu falls out of solu- tion. The second statement shows that in a process with soluble copper electrodes there is only a transference of Cu from anode to cathode, hence there is required little energy or only a small electromotive force to set in motion, by overcoming the resistance of the solution, large quantities of Cu or SO4 on their paths to the cathode, or anode. The quantity of current, ' XJlke, Electrochem. Ind., 1903, i, 240. Editor, 1904, II, 303. " Min. Ind., 1912, xxi, 288. ' Estimated. ^ To be increased to 480,000,000 lb. 'OflBcial figures, ELECTROLYSIS OF COPPER 485 on the other hand, has to be proportional to the number of Cu" ions, or, in industrial work, the amperage will be high as compared with the voltage. This regularity of equal solution at anode and deposition at cathode may be disturbed by the formation of CuaO^ with a low current density, e.g., if below 0.09 amp. per square foot. Here Cu" ions in the electrolyte, instead of re- ceiving two negative charges (Cu"+2— = Cu), receive only a single one (Cu"+i— =Cu'), with the result that CuSO* is reduced only to CU2SO4 instead of to Cu, and Cu2S0n^Cu+CuSO4 or 2Cu"?^Cu+Cu". This dis- turbance is favored by a high temperatiu:e and a high concentration of the electrolyte. It forms the explanation^ of the fact that in the electrolysis of copper in CuS04-solution there is always set free finely divided Cu which col- lects in the anode mud and causes this to assay 40-50 per cent. Cu. But CU2SO4 may be and is in part converted into CUSO4 by the reaction CU2SO4+ 0-|-H2SO4=2CuSO4+H2O, and this is one of the reasons for the neutralization of the free acid in the electrolyte. . 250. Behavior of Individual Impurities. — In industrial electrolysis the anode is not pure copper, but blister copper with 98-I- per cent. Cu. The behavior of impurities likely to occur in such material was first studied by Kiliani' with a solution containing 15 per cent. CuS04-f5 aq (=3.81 per cent. Cu) and 5 per cent, free H2SO4, with a current density of 1.8 amp. per square foot, and an electrode distance of 2 in. Though the composition of this electro- lyte resembles that in use at present, the current density is very low; never- theless the data of Kiliani may serve as a general guide, and will be supplemented by later information. According to their behavior in general, impurities are conveniently grouped under four heads: I. Ni, Co, Fe, Mn, Zn, Pb, Sn. II. Au, Ag, Pt, Se, Te. III. CU2O, CuaSe, CuzTe, CujS. IV. As, Sb,Bi. A general idea of the amounts of most of these elements and compounds present in the anode, and of their distribution in solution and residue after electrolysis, is given in Table 116.^ The unpurities of group I are all electropositive to Cu, and will be therefore dissolved before Cu and concentrated in the electrolyte at the expense of Cu. The first five, Ni, Co, Fe, Mn, and Zn are also attacked chemically by the free H2SO4 and therefore neutralize it. The Fe goes into solution as FeS04, is changed into Fe2.(S04)3 by anodic oxidation (2FeS04 -t-0-FH2S04 = Fe2.(S04)3 +H2O), and requires an addition of H2SO4; the Fe2.(S04)3 formed is again re- duced to FeS04 at the cathode by deposited Cu, viz., Cu-|-Fe2.(S04)3 = CuS04 ' Forster-Seidel, Zt. Electrochem., 1897, in, 479- ' Rossler, Dingier Polyt. J., 1881, ccxlii, 286. WoMwiU, in Borchers-McMiUan, op. cit., p. 199; ^t- Electrochem., 1903, ix, 311. ' Berg. HuUenm. Z., 1883, XLii, 23s, 250, 375, 399, 423; i88S> XLiv, 249, 261, 273. •Keller, Eng. Min. J., 1897, LXiv, 514; Min. Ind., 1898, vii, 239. 486 METALLURGY OF COPPER Table ii6.- —Analyses and Distributions of Anode Impurities Example No. i Example No. 2 Element or Analysis Distribution in Analysis Distribution in compound Anode, per cent. Residue, per cent. Solution, per cent. Residue, per cent. Anode, per cent. Residue, per cent. Solution, per cent. Residue, per cent. As 0.3432 0.00173 99-30 0.0093 0.0320 9.0651 0.0586 0.0025 i.oo7S 53-894 0.2959 II. 010 0.910 3-930 6.250 2.107 0.394 i.174 99-93 100 100 0.07 0.3444 99.40 55-150 0.198 13.820 2.070 0.340 2.440 1.090 0.718 0.892 0.800 10.680 99.914 100 Au Cu Pb 100 0.086 Bi 21.78 38.86 77.10 78.22 61 .14 2. 290 100 100 0.0035 0.0510 0.0180 . 6045 0.0056 (?)' 39-29 70.10 62, 16 60.71 29.90 37-84 Sb As Se Te 100 Fe SO4 5.268 2.36s H2O (250° C). + 2FeS04. In general, impurities electropositive to Cu offset its chemical corrosion; they may be present in considerable amounts before they affect the cathode deposit. It has been noticed that if the anode contains much Ni, the residue formed may assay as high as 10 per cent, nickel, the NiO going into the mud, the Ni into the electrolyte.^ Lead is converted into PbS04 which is practically insoluble and goes wholly into the anode mud. Tin goes into solution and then falls out as a basic sulphate liberating free acid. It has been noticed that Sn acted favorably upon the smoothness of the cathode copper. This is probably due to its causing the reduction of copper arsenate to arsenite, with which it forms an insoluble compound and thus puri- fies the bath.^ At the works of the Chicago Copper Refining Co., the late H. L. Bridgman used to add 25 lb. Sn to 100 tons of Cu in treating copper rich in As; the practice was too expensive for ordinary work. At present the arsenic is eliminated by smelting methods, and no high-arsenic copper is used as anode. Of the metals in group II, Au and Pt are not dissolved; Ag also is insoluble under normal conditions of electrolyte. Any Ag found in the cathode has been carried to it mechanically. Should the electrolyte become neutral, Ag will be dissolved and deposited. The rare elements Se and Te go completely into the anode mud. Of the compounds assembled in group III, CU2O is always present in the anode; it is not attacked electrolytically and sinks to the bottom of the vat; there it may be dissolved chemically by the free acid and thus reduce the amount of the latter. The compounds CuaSe, CuaTe, and CuaS are attacked neither electrolytically nor chemically. ^L. Addicks, 191 2. ' Ulke, Min. Ind., 1897, vi, 242; Peters, "Modern Copper Smelting," 1895, p. 600. ELECTROLYSIS OF COPPER 487 In group IV are collected the three metals which are partly dissolved and which partly fall again out of solution as basic sulphates, or may be deposited with the copper on the cathode. The electromotive force necessary for de- composition is, for Cu 0.30, for As 0.27, for Bi 0.21, for Sb o.io volt.i Arsenic in the metallic state is dissolved as As2.(S04)3, and this'saltlT^ore or less decomposed by hydrolysis: As2.(S04)3+6H20<=i2H3As08+3H2SG4, when H3ASO3 (or As203-|-aq.), being only slightly soluble, falls out of solution and goes into the mud. Highly arsenical copper becomes coated with greenish, slimy CU3AS2O8 which increasing the resistance has to be brushed off in case the coating becomes heavy.^ If with a difference in potential of 0.3 volt, a current density of 15-20 amp. per square foot, and a bath temperature of 40- 50° C, the As-content of the electrolyte reaches 2 per cent.. As is precipitated upon the cathode; hence the aim is to keep the As-content below 1.25 per cent. 101.0 10fl.R Arsenic Content of Electrolyte, Eer.ccnt 0.1 0.2 0.3 0.4 O.B 0.6 0.7 0.8 0.9 1.0 1.1 1.2 1.3 1.4 1. 100 6 \ 100 4 \, 100.2 \ \ V \ 99 8 N 99.0 \ \ 99.4 o o v^ o o ^ ^ - 99.0 Fig. 483. — Relation of arsenic-content of electrolyte and conductivity of cathode copper (Addicks). Deposition also takes place when the Cu-content falls below 2.8 per cent. With an anode containing 0.3 per cent. As, from 30 to 35 per cent, of the As goes into the mud, and from 70 to 65 per cent, into the electrolyte. This general figure is changed by the As-content of the electrolyte, as with a bath ahready containing a considerable amount of As, more of this metal will go into the mud than into the solution. With very small quantities of As, as in re- fining some Lake copper, all the As goes into the mud. According to Wickes,' deposition of As is largely governed by the degree of hydrolization. The relation between As-content of electrolyte and conductivity of cathode- Cu is shown in Fig. 483 (Addicks). ' Neumann, Zt. phys. Chem., 1894, xiv, 229. 'Ulke, Min. Ind., 1897, vi, 240. 'Tr. A. I. M. E., 190S, XXXV, 40 488 METALLURGY OF COPPER The compound CusAsaOg being a non-conductor^ goes into the miid, where it is in part decomposed by the free H2SO4 which dissolves the AsaOs; some AsaOs is also formed by oxidation of AS2O3 at the anode. From a neutral solution As is readily deposited on the cathode. Wen^ found that additions of small amounts of HCl, Na2S04, AICI3, and NaCl improved the cathode copper chemically in hindering the deposition of As and Sb, and physically in preventing the formation of trees. Of these in- organic additions, NaCl is the most effective.' Organic agents, such as gelatine and tannin, aid in furnishing smooth deposits; this is not the case with peptone.^ The combined addition of o.oi per cent. CI in the form of NaCl, and of o.oi- 0.02 per cent, gelatine gives a smooth ductile deposit of great purity with an electrolyte containing CUSO4+S aq., 15 per cent, and free H2SO4 10 per cent, held at 40° C, even when this contains as much as 6 per cent. As, the current density being 40 amp. per square foot and the pressure about 0.5 volt. It is common practice to add a very small amount of glue to the storage tank for electrolyte. Even the usual ridiculously small addition decreases the resistance of the bath; further it has been found that it takes a much longer time in the refining furnace to bring such cathodes to the stage of set copper, than if no glue whatever has been used. Antimony.^ — The behavior of Sb is similar to that of As, but Sb is less readily deposited. In large-scale work some insoluble dark antimony compound is often seen floating on the surface of the electrolyte; there is then danger of its adhering to the cathode and becoming entrapped. Care is usually taken to collect it from the last of a series of vats forming cascades by placing a screen across the outlet or beneath the overflow. The antimony content of the electro- lyte appears to remain approximately constant at 0.03 per cent. The effects of adding organic agents have been noted under arsenic. Bismuth. — This stands between As and Sb as regards its behavior in electro- lysis (p. 487); it is more readily deposited than As. According to Terrill' an addition of a drop of Br- water to a sample of electrolyte will indicate by the formation of a white cloud the presence of Bi; if the cloud appears at once, the danger-point of electrodeposition has been reached; if it takes about i min. to form, the danger-point will be reached in about 48 hr. Usually Sb, and especially Bi, occur in quantities too small to cause any trouble with the cathode copper.' 251. The Current. — The efficiency of the refining process is dependent ' Johnson, Electrochem. Ind., 1904, n, 207. ^ Dissertation, Columbia University, 1911; Tr. Am. Electrochem. Soc, 1911, xx, 121. ' See also Speer, op. cit., 1912, xxn, 281. * Jarvis, School Mines Quart, 1909, xxx, 100. 'Hampe, Eng. Min. J., 1892, Liv, 78; Berg. Hiittenm. Z., 1892, li, 177; Chem. Z., 1892, XVI, 417. Sprent, C. "Das Verhalten von Antimon bei der Kupferraffination," Dissertation, Dresden, 191 1. ' Trans. Inst. Min. Met., 1897-98, vi, 215. 'See also, Motherwell, Met. Chem. Eng., 1913, xi, 670. ELECTROLYSIS OF COPPER 489 upon the character and temperature of the electrolyte, the current density, and voltage. For a given temperature the conductivities of a CUSO4- and a H2S04-solution increase within certain limits with the CuSOi and H2SO4 present, but by the addition of H2SO4 to the CuSOi-solution the dissolving power of the latter for CuS04 is diminished, as well as the dissociation of the salt, i.e., the conductivity of the electrolyte. In the same manner the dissociation of H2SO4 in the electro- lyte is diminished by the presence of CUSO4. Tables 117 and 118 give the ex- perimental results of Richardson and Taylor.^ Table 117. — CoNDUCTrviTiES of Mixtures of CUSO4+5AQ. and H2SO4 in Reciprocal Ohms per Cu.cm. Temperature 25° c. 45° C. Gram HjSOi in TOO c.c S 10 IS 20 5 10 15 20 Gram CuSOi +5 aq.inioo c.c S 10 IS 20 0-953 0.0221 0.0343 0.0423 0.208 0. 204 019s 0.189 0.182 0.410 0.388 0-350 0.338 0.319 0.565 0.531 0.500 0.458 0.433 0.683 0.646 0.600 0.558 0.0205 0.0294 . 0468 0.0574 0. 246 0.242 0. 222 0.217 0. 212 0.492 0.461 0.422 0.381 0.378 0.683 0.643 0.606 0.54s 0.521 0.839 0.791 0.738 o.6go 0.643 Table 118. — Conversion of Data in Table 117 to Practical Notations Gram CuSOj+saq. Cu, . CuSOi+s aq. GramH2S04. H2S04, in TOO c.c. per cent. per cent. in 100 c.c. per cent. 4.0 I 3.91 ■ 515 5 8.2 2 7.82 10.17 10 17-4 3.99 15.64 16.5 15 27.4 5.99 23.46 22.8 20 32.9 6.98 27.37 The conductivities of working solutions are about 15 per cent, smaller than those found by the experimental work with mixtures of pure CUSO4+S aq. and H2SO4. Table 117 has shown that a rise in temperature of bath increases the conduc- tivity. The relation between voltage and current density in a bath with 16 per cent. CuS04H-s aq. (or 4 per cent. Cu) and 9 per cent, free H2SO4, for tem- peratures ranging from 20 to 90° C, is shown in Fig. 484.^ The voltage does not rise in the same ratio as the amperage, as might be expected, and give a straight-line curve; the curves converge toward the current density axis, and do this the more the lower the temperature of the bath. The relation between voltage and temperature for different current densities ' Trans. Am. Electrochem. Soc, 1911, xx, 179; Mel. Chem. Eng., 1911, k, 536. ' Schwab-Baum, /. phys. Chem., 1903, vii, 497. 49° METALLURGY OF COPPER 2.0 2.5 8.0 Amperes Fig. 484. — Relation of current and tem- Fig. 485. — Relation of voltage and tempera- perature in electrolyte with 16 per cent. ture for different current densities in electroljrte CuS04+5aq., 9 per cent, free H2SO4. with 16 per cent. CuSOi+saq., 9 per cent, free H2SO4. 20° 30° 40 50° 60° Temperature 70 80° 90 20° 30° 70° 80 90' r 50° 00 Temperature Fig. 486.— Relation of voltage-ratio and tem- Fio. 487.— Relation of watt-hour and tempera perature for different current-densities in electro- ture for different current-densities in electrolyt lyte with 16 per cent. CuS04+saq., 9 per cent. with 16 per cent. CuS04-f-saq., 9 per cent, fre free HjSO,. H2SO4. ELECTROLYSIS OF COPPER 491 is given in Fig. 485. The voltage, measured with electrodes i cm. apart, decreases as the temperature rises; the decrease is rapid at low temperatures, and becomes less and less as the temperature rises. In order to make the curves in Fig. 485 independent of the electrode distance, the curves in Fig. 486 have been drawn by Schwab and Baum, in which the voltage ratio has been plotted as ordinate instead of the real voltage, voltage at 20° C. having been made the standard. If the voltage at one temperature is known, that for another temperature is found through the curve. Theoretically i amp.-hr. deposits from a CuS04-solution i.r86 g. Cu, or in order to deposit i lb. av. Cu there are required 382.4 amp.-hr. ' In practice, 400 to 500 amp.-hrs. are necessary. The relation between the watt-hour required for the deposition of i g. Cu, and the operating temperature based on the laboratory experiments of Schwab and Baum is given in Fig. 487. A. Multiple System 252. The Multiple (Elkmgton) System in General.' — In this process, Figs. 488-489, the anodes, a, of high-grade copper, and cathodes, b, of pure copper, are connected in multiple, and suspended crosswise in an oblong vat, c, charged with a solution of blue vitriol containing free sulphuric acid. (A current of suitable strength passes from the anodes through electrolyte to cathodes, WM/MMM/M/ M/Z/MmM/m/M' , ''/////////^////^///W/////////////////////////////7f Figs. 488-489. — The multiple system. dissolves copper from the anodes and deposits it on the cathodes, while insoluble impurities collect on the bottom of the vat as a residue called anode mud or anode 'Anaconda: Editor, B,ng. Min. J ., 1896,1x11, 271. Hering, Berg. Hiiltenm. Z., 1893, in, S4- Hofman, Tr. A.I.M. E., 1904, xxxiv, 308. Raritan No. i : Addicks, Min. Ind., igoo, X, 261 (remodeled in 1912 on lines of No. 2). Balbach S. & R. Co.: Electrochem. Ind., 1904, II, 303 (remodeled 1910). Great Falls: Hofman, Tr. A. I. M. E., 1904, xxxiv, 308. Burns, op. cit., 1913, XLVi. Chrome: Addifks, Min. Ind., 1906, xv, 301; Eng. Min. J., 1907J Lxxxiii, looi. Raritan No. 2: Easterfirooks, Electrochem. Met. Ind., 1908, vi, 181, 245, 877- Lithgow, N. S. W.: Blakemore, Tr. Austral. Inst. Min. Eng., 191 2, xv. 36; Eng, Hin. J., 1910, xc, 717, 769. Port Kembla, N. S. W.: Casey, Eng. Min. J., 1910, xc, iiii. 492 METALLURGY OF COPPER slime. The deposited copper with its cathode is removed at intervals, melted down in a reverberatory furnace, toughened, and cast into suitable forms. The anode mud containing the electronegative precious metals and insoluble impurities is refined, and cast into bars. Occasionally other substances, such as selenium and palladium, are recovered as by-products; tellurium, for which there is no market at present, goes to waste. The uncorroded part of the anode goes back into the furnace from which the anodes are cast. 253. Electroljrte — Composition, Temperature, and Circulation. — The range of composition of the electrolyte is CUSO4+S aq., 12-16 per cent. ( = 3-4 per cent. Cu) and free H2SO4, 5-13 per cent.; the usual figures are: Cu 3 and free H2SO4 12 per cent. The Cu-content is never allowed to fall below 2§ per cent., as otherwise there is danger of As being plated out; with over 13 per cent, free H2SO4 the bath is decomposed electrolytically and polarization is likely to offset increased j:onductivity.* In Table 119 are given analyses of electrolytes of different de- grees of purity and concentration. Table iiq . — Analyses DF Electrolyte Great Falls, Mont. Pertli Amboy, N. J. Maurex, N.J. Refining Tank Starting- sheet Tank Room I Room II Cu 3.280 0.500 0.041 0-377 0.016 0.021 0.600 None None Trace 0.418 None 3-404 0.425 0.034 3-91 1.03 0.009 0.38 4.00 1.02 0.009 0.34 3 - 1 Asr 1 .0 Sb 0.03 Ni Co Bi Fe 0.383 0.096 0.104 0.25 Se Te Pb ... Zn Ag, Au CaSOj 0.13 1.26 12.85 19-73 0.19 0.85 12 .24 18.89 Na2S04 Free H2SO4 . 13.030 10.213 Total H2SO4 CI 0.004 1 . 220 (2) 0.0034 I-I7S (2) 0.003 1 .2 (4) Specific Gravity Reference I -255 (3) I -255 (3) The different soluble metal sulphates appear to act cumulatively as regards conductivity, thus, e.g., the conductivity of a bath with 3 per cent. Cu and 0.5 per cent. Ni is approximately the same as one with 3.5 per cent. Cu. 1 Addicks, "Rapid Measurement of Conductivity," Ekctrochem. Ind., 1904, n, 306. ^ Burns, Tr. A. I. M. E., 1913, XLvi. ' Private Communication, C. H. Aldrich. ' Private Communication, H. H. Alexander. ELECTROLYSIS OF COPPER 493 In working, the normal composition of the bath is likely to be changed. It is impoverished in Cu because the electropositive metals Zn, Fe, Ni, Co, and Mn go into solution and replace equivalent amounts of Cu; it is enriched in Cu by the chemical action of the free H2SO4 and by the dissolving effect of CuSOi upon Cu (§ 243), which takes place largely at the surface of the bath; the total amounts to from 0.5 to i per cent, of the Cu deposited. With the high- grade anode in common use at present, any impoverishment in Cu is more than balanced by enrichment; hence Cu is removed at intervals either in the metallic state, by plating out, or as blue vitriol by crystallizing out. In general from i to 2 per cent, of the cathode capacity has to be removed. However, in the presence of much Fe and Ni, more Cu may be deposited than is dissolved. The electrolyte is always impoverished in its content of free H2SO4 because the separation of impurities as normal or basic sulphates, and to the chemical action of the free H2SO4 upon Cu and CU2O. The acid has therefore to be replenished. The temperature of the bath ranges from 40 to 60° C. The hotter the bath, the lower the resistance and the smoother the cathode deposit, but the chemical action of the acid is also greater. The electric energy raises the tem- perature to about 34° C. ; for a higher temperature heating by steam-coils (1-2 in. in diameter) in storage tanks is required. With 60° C. the evaporation in 24 hr. in a tank is about 22 lb. water per square foot solution exposed. The fall in temperature of the electrolyte in passing through a cascade is about 5° C. in summer, and 10° C. in winter. The idea of covering tanks^ to diminish the re- duction of temperature owing to radiation is at present impracticable.* Experiments are, however, under way to make it practicable. Continuous circulation is essential to correct differences in composition of electrolyte caused by the process. At the anodes, where copper goes into solu- tion, the electrolyte is heavier than at the cathodes, where it goes out of solution; the heavier part sinks and the lighter rises, causing layering in the bath; the cur- rent passing mainly through the heavier solution causes uneven corrosion of the anode and irregular deposition on the cathode. The greater the current density and the higher the temperature of the bath, the more rapid the circu- lation required. LThus, with a current density of 40 amp. per square foot of cathode area, the electrolyte is exchanged once in 3 hr.; with 15 amp. once in 4 hr., with 10 amp. once in 5 or 6 hr., the rate of flow ranges from 6 to 3 gal. per minute.*} It is important that the flow be sufficiently slow to permit all anode residue to settle, and its path in such a direction as to leave settled residue undisturbed. The rate of circulation is also governed by the impurity and precious metal of the anode. Thus, the higher the As-content the greater has to be the rate of circulation, if the deposition of As is to be avoided; on the other hand the higher the content in precious metal, the slower has to be the ' Schwab-Baum, //. Phys. Chem., 1903, vn, 493. 'Addicks, Eleclrochem. Ind., 1903, i, 487. 'Tests at Great Falls by Burns, Tr. A. I. M. E., 1913, xlvi. 494 METALLURGY OF COPPER circulation in order to prevent stirring up the large amount of mud which settles on the bottom of the vat. Kiliani^ measured the differences in potential arising when working with and without circulation. The usual method of circulation is to have rows of vats on wide terraces with steps 2-3 in. high, and to let the electrolyte overflow from the vats on the top row into those on the next row below and so on. In Fig. 490 the elec- trolyte is raised from a well into a main whence one part flows into a distrib- uting box for the electrolyzing vats arranged on either side in cascades, while another is diverted to the liberators, i.e., vats with insoluble anodes in which electrolyte is freed from Cu, As, etc., and then returned as fresh acid to the main circuit. From the last row of electrolyzing vats the electrolyte flows into a trough emptying into the well or sump connected with the pump. Each of the vats in Fig. 490, as well as in Figs. 505-506 (Great Falls, Mont.). Fig. 490. — Circulation of electrolyte. shows a partition at the discharge-end which reaches to within 6 or 8 in. from the bottom so as to leave room for the settling of the anode mud. The elec- trolyte is received on the top of the bath at one end, and withdrawn from near the bottom at the other; it thus has to travel diagonally through the vat whereby a uniform density is maintained.^ Another arrangement. Figs. 488-489, is to have the partition, d, at the feed-end; when the electrolyte is delivered back of the partition, sinks downward and travels diagonally upward to the dis- charge spout, e, at the delivery end. The arrangement at the Raritan plant No. 2, Perth Amboy, N. J., is shown in Fig. 491. The electrolyte is delivered to a tank back of a semicircular lead partition at the inflow-side and passes to the bottom; the bulk of it over- flows at the top at the opposite end, but a small amount of heavier solution is withdrawn near the bottom through three rows of small holes in the lead parti- tion. Attention may be called in connection with Fig. 491 to the means em- ployed for preventing any countercurrent or stray electric current from interfering with the main current. The delivery and receiving mains rest ' Berg. Hutlenm. Z., 1885, XLiv, 273. 2 Inefficiency of method with high current density at Great Falls; Burns, Tr. A. I. M. E., 1913, XLVI. ELECTROLYSIS OF COPPER 49 S upon glass; the branch delivery and receiving lead pipes are cut and con- nected by non-conducting hose, and are protected by vitrified brick. ^ If the tanks are all on one level they may either receive their solutions severally from a common feed-trough, as was the case with the original plant ^^^'"^^"'^"^^^^^^^■w'""^^""'"!!^^ 3^ Fig. 491. — Circulation of Electrolyte at Raritan Fig. 492. — Siemens-Borchers circula- Plant No. 2. tion of electrolyte. of Great Falls, Mont.,*, and deliver into a common discharge trough, or the solution of each tank may be circulated independently by the Siemens-Borchers apparatus^ sketched in Fig. 492. An L-shaped lead pipe, a, is lowered at one end of the vat so that the horizontal arm shall lie on the bottom and underneath TiAkb Jlned wnb 6 Lb, Xead well dreaaed up to OutaidQ BozlDg over Toq nod doirn Outsldo Xonsitudinal Section of Tanks Fig. 493. — Circulation of electrolyte at Lithgow, N. S. W. the tray, b, which is to receive the anode mud. A smaller lead tube, c, drawn down at the bottom, is inserted into the vertical arm. Air under 3 or 4 lb. pressure is forced down tube c, rises between a and c, acts as an air-lift pump, and causes the solution to overflow from pipe a; a corresponding amount of ' Burns, Tr. A. I. M. E., 1913, XLVi. ' Borchers-McMUlan, op. cit.,p. 221; Zt. Ekclrochemie, 1904, P- 221. 496 METALLURGY OF COFFER course must enter at the bottom. In this manner the solution of each vat is circulated independently of its neighbor, and at the same time aerated. The aeration of warm solution will cause ferrous salt to be converted into insoluble basic ferric salt, and the electrolyte to become purifiedX^Schneider-Szontag i slightly modified the above device' at Maurer, N. J. The rate of circulation by Fig. 494.— Pohle air-lift pump, Great Falls, Mont. the apparatus is not sufficiently large for the current density used in the United States, and the method has therefore been given up.^ A modification is in operation at Lithgow, N. S. W.,^ where a vertical 5-in. 'Ulke, Eng. Min. J., 1896, Lxn, 464. " Tr. Aust, Inst, Min. Eng., 191?, xv,36; Eng. Min. J., 1910, xc, 717. ELECTROLYSIS OF COPPER 497 copper pipe connected with air under 5 lb. pressure at the upper end and turned up at the lower, discharges into the lower end of a f-in. vertical lead pipe reaching to within 6 in. from the bottom of the tank; the compressed air acts as a Pohle air-lift pump and raises the bottom part of the solution to the top; this is done in addition to the regular circulation of the electrolyte down the cascade. ^ The fall in temperature of a bath varies with the size of vat, and the rate of circulation, which in its turn is dependent upon the current density. As shown on page 493 the range of temperature is from 5 to 10° C, varying with the season of the year. The electrolyte from the lowest numbers of the cascades is collected in a sump and pumped into a distributing tank. Formerly lead-lined acid-eggs and plunger pumps were used for this pur- pose; at present the Pohle air-lift pump' and the Antisell centrifugal pump have replaced the older apparatus. The Pohle air-lift pump in use at Great Falls, Mont., is shown in vertical section in Fig. 494. The feed and de- livery pipes, connected by a return- bend, are 6 in. inside diameter, have |-in. walls of hard lead, are cast in 4-ft. lengths with flanges (the air-pipe is f in.) ; the whole is held in a cenient-Iined well, 18 ft. 9 in. deep and 4 ft. in dia- meter. Working against a head of 14 Figs. 495-496.— Antisell circulating pumps, ft. 8 in., 160 gal. electrolyte (specific Raritan Works, gravity 1.22) are raised per minute with 80 cu. ft. free air of 16 lb. pressure. The Antisell centrifugal pump. Figs. 495 and 496, of the Raritan Copper Co., Perth Amboy, has a hard-lead (12 per cent. Sb) cylinder, 4 ft. 6f in. high and 14 in. in diameter, above which are the bearings of the driving shaft carrying at its lower end the rotor making 750 r.p.m. The inlet and outlet pipes are 8 in. in diameter. The pump handles 46.5 cu. ft. solution per minute, has a capacity of 66 cu. ft., and requires 5 h.p. 254. The Current. — The drop in potential between vats ranges from 0.2 to 0.4 volt. Magnus^ found at Anaconda that of this total fall as much as 22.5 per cent, was due to contact resistances and current leakages. At Great Falls, Mont.,' the drop between anode busbar and anode is 7.40 per cent, of total voltage, between cathode busbar and cathode 9.24 per cent., and across the electrolyte 83.36 per cent. Addicks* distributes the resistances in a tank as ' Hofman, " General Metallurgy," 1913, p. 710. " Electrochem. Ind., 1903, i, 561- ' Burns, loc. cit. *Jl. Franklin Inst., 1905, clx, 431; Private Comm., 1912. See also Hutchinson, Electrochem. Ind., 1904, 11, 13, Addicks, op. cit., p. 180; Spalding, Min. World, 1910, xxxii, 102 (Power). 32 498 METALLURGY OF COPPER follows: metallic lo per cent., liquid 60, transfer 5, contact 15, counter 5, slime 5. With well-insulated tanks and broken connections in circulating pipes, the current shunted arotmd by grounds should not exceed i per cent, of this total; short-ckcuits between anodes and cathodes or by indirect contacts between electrodes and tank- walls amount to 5 + per cent. ; hence the efficiency shows a range of go to 95 per cent. It is essential to keep contacts bright by scouring with emery; the bright surfaces are sometimes coated lightly with oil to retard corrosion. The current density shows a range of 12 amp. per square foot (Calumet & Hecla) to 34 amp. (Great Falls);' the former average of Eastern refineries of 15 amp. has been raised to 18 to 20 and 21. The low current density at the Calumet & Hecla refinery is due to the desire of preventing even a trace of As in the anode from passing to the cathode.^ The high density of Great Falls finds its explanation in the cheap water-power of the Missouri River which compensates by the large output for the given plant for the loss of energy. With a large density it is necessary to exchange the cathodes more frequently than with a low in order to prevent short-circuiting. Thus at Great Falls, in 1904, with 40 amp., exchanging cathodes every second day gave an ampere efficiency of 91 per cent., while with four-day cathodes this fell to 85 per cent. The more recent data by Burns,' dealing with current densities of 32.6-36.9 amp. per square foot, are given in Table 120. Table 120. — Relation of Age of Cathodes and Number os Electrodes in Tank Age of cathodes, Electrodes per tank Average amperes Av. amperes per sq.ft Ampere efficiency, per cent. Cu per kw-hr., lb. Cathode, oz. Ag per ton Cathode, per cent. days Anodes * Cathodes As-l-Sb 4 20 20 9,300 36.9 88,0 3-93 1.32 0.0036 2 20 20 8,808 3S-0 90.85 3 72 0.83 . 0030 3 20 20 8,878 35-2 89.00 3 75 0.83 0.0032 2 21 21 9,03s 34 I 90.90 3 84 0.89 0.0032 3 21 21 9,223 34-8 89.40 3 87 1 .02 0.0029 2 22 22 9,071 32.6 90.50 4 02 0.89 . 0033 3 22 22 9,167 33-0 88.80 4 07 0.95 . 0030 With a density of 15 to 20 amp., two cathodes serve for one anode; with 10-12 amp., they are exchanged only when the corroded anode is ready to be removed. Large plants are divided into several sections, each of which is served by one generator. Numerical examples are shown in Table 1 28, where details are given with the descriptions (§ 266) of the Anaconda, Great Falls, and Raritan No. 2 plants. ' Forty amp. in 1904. ' It is the intention to raise the density. 'Loc. cit. * Converter anodes: Cu 99.13, As 0.127, Sb 0.055 P^r cent.; Ag 33.91 and Au 0.22 oz. per ton. ELECTROLYSIS OF COPPER 499 255. Anode. — The anode ought to be of such a character that it is evenly- corroded and does not affect the cathode deposit. Even corrosion is possible only if the amount of impurity present is small and the pitch of the copper right. An anode rich in precious metals usually assays 97.5 per cent. Cu; one with little silver, 99 per cent. Cu and over; 95 per cent. Cu is probably the lowest permissible figiure. Analyses of anodes are given in Table 120. Table 120.- -Analyses OP Copper Anodes' Range, U. S. Average, U. S. Kosaka, Japan Maurer, N. J. Great Falls, Mont. Raritan, Perth Amboy, N. J. Cu 98-99 -s 0-300 0-40 0-2 99-2S f incl. in. \ Cu 0.05 o.os O.OI O.OI 0.08 99-034 42.19 0.4509 0.044 o.o6s 0.049 0.004 99.1300 39-98 0.23 0.1183 OOS34 0.0038 O.OIIO 0.0420 0.0018 0. 2610 0.0090 0.0170 98.986 90.00 I-2S 0.196 0.017 0.014 0.047 0.093 99-S74I 7S-00 Agoz Au oz 300 As Sb 0.08-0. IS 0.048 Bi o.ooss 0.019 Fe 0.03 Ni Co S 0.003 0.03 0.03 o.os 0.02 O.OI 0.009 0.177 0.006 0.014 Se ... some some on Te 0.013 Si Pb 0. 109 O.OOI 0.006s 0.003s U.140 0.013 Zn Reference Addicks, Jl.FrankHn ImtU,igos, clx, 422. Addicks, private communi- cation, 19 1 2 Private communi- cation, 1910 Private communi- cation, 1912 Burns, Tr. A.I.M.E., 1913, xlvi Private communica- tion, 19 1 2 If an excess of harmful impurity is present in any shipment, it is good policy to mix this with material of higher grade from another source in the casting furnace instead of running chances of over-charging the electrolyte with harmful metal and obtaining inferior cathode copper. In the United States most of the anode material is converter copper, hence high-grade, and low in As, say 0.05-0. 10 per cent. Formerly with reverberatory copper the As-content frequently exceeded i per cent. Anodes cast direct from the converter are less desirable than after the copper has been transferred to a reverberatory furnace and poled to reduce the 0- and SOj-content, as the anode is irregular in thickness calling for wide spacing in the tank, is unevenly corroded, gives much scrap, and fxu-nishes an anode mud rich in Cu. Thus experiments at Great Falls, Mont.,^ showed that converter copper with Cu 99.27 and As, Sb 0.071 per cent., Ag 61.14 and Au 0.20 oz. per ton, gave 8 per cent, scrap, screened (40-mesh) mud with Cu 40.3 per cent, Ag 6755 and Au 18.34 oz. per ton, and cathodes with 1.25 oz. Ag per ton; ' See also p. 486. "Hofman, Tr. A. I. M. E., 1904, xxxiv, 310; Burns, op. cit., 1913, xlvi. Soo METALLURGY OF COPPER Fig. 497. — Anode with shoulder. while similar copper with Cu 99.27, As, Sb 0.071 per cent, Ag 61.14 and Au 0.219 oz. per ton poled in the reverberatory furnace gave 5.30 per cent, scrap, mud with Cu 18.80 per cent., Ag 14079 and Au 38.45 oz. per ton and cathodes assaying 0.95 oz. per ton. If, nevertheless, anodes are frequently cast direct from the converter, the reason is that saving the expense of the work in the reverberatory furnace more than balances the dis- advantages of a higher percentage of scrap, of a greater cost of treating the mud, and of the loss of Ag in the cathode copper. In preparing anode material in the refining furnace, the poling is carried only to plate-pitch, i.e., the cast plate shall have a level surface, and for this the copper must retain a considerable amount of CuaO. The size of the anode is largely determined by the convenience of handling. At first, anodes were made small, 30X24 in. and i in. thick, and weighed about 230 lb. They were raised and lowered singly by a block and tackle suspended from an overhead track. At present, they are usually larger, about 3 ft. square, owing to the fact that the twenty odd anodes of a vat are raised and lowered together by means of an overhead electric traveling crane. The width of the anode is limited in part by the tendency of the opposing cathode to curl before it has thickened sufficiently to become rigid. In fact, a new cathode is usually removed after it has been in the vat for two days, straightened in order to make the elec- trode-distance uniform, and lowered again in place. This curling can be avoided by using two cathodes for one anode as was done at Anaconda, Figs. 530-534, and is still the cus- tom at the Balbach works, but this compli- cates the handling by means of a crane. The length may be influenced by the per- centage of precious metal present, as the lower end of the cathode is likely to become richer in Ag than the upper because of con- tamination with falling slime. Anodes con- taining much precious metal will be made shorter than those containing little; the other remedy, deepening of tank, means a larger crane-lift. The anode situated near the bottom-outlet of the electrolyte. Figs. 505-506, is sometimes made shorter than the rest, to counteract any stirring-up of settled slime. The thickness depends again upon the amount of precious metal as well as upon the cost of handling. A thick anode takes a long time for corrosion and I 31*- Solution /Line 4H!= i T A T • 2'oyf Great Falls Standard Anode with Jlorrow Clip Figs. 498-499. -»! 1 //r 2^i ELECTROLYSIS OF COPPER SOI represents a large average for tank resistance, but requires only a single hand- ling; a thin anode furnishes a large percentage of scrap to be resmelted because of the disproportionate weight of the supporting lugs. An anode isexpected =====^1 •£ Fi->r--»~-»h7R-8)»-+--f*4'--"---*i«'-r •s ifil .„ JSli I . if iti s^-l Anode Transfer Car Figs. 500-302. — Transfer-car for anodes. to be corroded in from two to six weeks, and is made from i to 1.5 in. thick. Details of some anodes are given in Table 128. The manner of suspension and with it the form of the upper part of the anode varies at different works. Ordinarily the anode is cast with a projection or 502 METALLURGY OF COPPER arm or lug on either side, as in Fig. 497, representing the Anaconda anode, by which it rests on the conductor bars. C Sometimes it is cast with two per- forated lugs. Figs. 530-534, by means of which it is suspended with heavy copper-wire hooks from the cross-barD A third method is represented by the one using the Morrow clip. Figs. 498-499, found at Great Falls, Mont., and Maurer, N. J., a loop of J-in. copper rod the ends of which are placed in the anode mold before filling with copper. The electric contact is thus made perfect. Each form has its d^enders; the main consideration, besides perfection of electric contact, is to reduce as much as possible the amount of metal not im- mersed in the electrolyte, which has to go back to the anode furnace. At Great Falls^ with the anode shown in Figs. 498-499, and weighing 500 lb., the scrap formed in 1912 was 5.9 per cent. The manner of casting has undergone many changes. At first, the open cast-iron anode molds placed on the floor were filled by hand-ladles;^ later suspended (bull) ladles came into use, and these are still common in plants dealing with charges of 30 tons of copper and less; for some time the metal was tapped into sand molds. The advent of the Walker casting-machine. Figs. 425-426, which permitted reverberatory-furnace charges of 200-t- tons of cop- per, did away with ladling in large plants. Forms of rotating tables other than the Walker or Clark are in use at Great Falls (Kleppinger machine. Figs. 353- 354) and Brooklyn; link-belt machines (p. 400) are found at Perth Amboy and Chrome, N. J., and Anaconda, Mont. Special attention has been given to the details of the molds to prolong the life, to insure a smooth bearing surface of the arm, and to obtain an easy release of the anode. An anode is released from the mold either by a knock-out pin (p. 400) raised mechanically from the center of the mold, or by prying after having removed near the top a detachable part of the rim of the mold. A released anode is picked up by a compressed-air lifting apparatus and lowered in an iron water-tank the sides of which have notches to receive the shoulders of the anodes. A tank is of the same size as a depositing vat, and the distance between the notches is equal to the electrode distance. The anodes are cooled and scaled, i.e., freed from CU2O, by the im- mersion. From the immersion tank the anodes are removed by an overhead electric crane and deposited on a transfer car. Figs. 500-502, on which they are hauled to the tank room. The anode molds used are always open. Truswell has developed a closed mold.' 256. Cathode. — The cathode, or starting sheet, is a thin plate of copper deposited upon rolled sheet copper, \-\ in. thick, in a special set of "stripping tanks" which usually are made deeper than the regular corroding tanks. Figs. 503-504 show such a " stripping sheet" riveted to a pair of cross-bars. In order to prevent the deposited metal from adhering to the rolled sheet, the latter is ^ Burns, loc, cit. ' Illustration, Min. Sc. Press., 1899, lxxdc, 266. ^ Eng. Min: J., 1906, lxxxi, 853. ELECTROLYSIS OF COPPER 503 i-a' y ' ^ ■' '. =iP? -a'lX' a. «-, ■ 'H J L d Figs. 503-504. — Cathode stripping-sheet. greased and then sprinkled from a pepper-box with well-conducting graphite or painted with a low-grade mineral oil containing artificial graphite, i gal. oil: 0.5 lb. graphite. Ordinarily the rim of the rolled sheet is coated for the width of I in. with asphalt in order to prevent the plating-out of any copper and thus facilitate the stripping of the deposited copper when this has reached a thickness of about iV in. The coating of asphalt is applied with a brush, or the sheet is dipped. Sometimes grooved strips of wood slipped over the sheet take the place of the asphalt. Instead of giving the sheet copper approximately the shape of the starting cathode, a small groove, tt in. deep, is made in the stripping sheet, \ in. from the rim, a-a on one side and a' -a' on the other in Fig. 503, which traces the outline of the starting cathode; then the sheet is greased, peppered, and suspended in a corroding vat supplied with the regular anodes. The deposited copper will ;part readily along the groove when the plated metal is being removed. This makes the stripping, first of the metal and then of the border by means bf a chisel-pointed bar, easy, and the starting cathodes are well trimmed at the same time. According to Huntington^ the lines of crys- tallization of deposited metal are at right angles to the surfaces on which the deposit is made; hence in a groove there will be discontinuity of the two sets of crystals and a line of weakness will be developed, a phenomenon resembling the line of weakness in a rectangular casting in which the crystals arrange and group themselves with their principal axes in lines perpendicular to the cooling surfaces.^ In order to give the thin starting cathodes the tensile strength' necessary to carry the weight of the copper to be deposited, it is important that the copper be plated out slowly; hence the cxurrent density is made lower than in regular work, e.g., one-half at Great Falls. This is accomplished by dividing the cur- rent, and by increasing the resistance of the electrolyte, either by reducing the copper and acid-content or by adding gelatine to the bath. The ampere effi- ciency is about 85 per cent.* There is provided one starting sheet tank for 6 to 12 corroding tanks. The time required for preparing a starting cathode of a given thickness is ascertained from the fact that i amp. per square inch (or 144 amp. per square foot) will give in i hr. a deposit 0.008104 in. thick. Ordi- narily it takes 10 days to prepare a sheet; at Great Falls the time is 12 hr. The cathode is usually made slightly longer and wider than the anode. Figs. 505- 506, in order to prevent or diminish the formation of excrescences. The cathode extends downward to 6 or 8 in. from the bottom of the tank, in order to leave room for the accumulation of anode mud and for the passage of the i£ng. Min. J., 1905, Lxxx, 1109, photomicrographs. 2 Microphotographs of Waser and Kuhnel, Electrochem. Zt., 1912, xviii, 151, 211. ' Forster, Berg. Huttenm. Z., 1899, LVlii, 473. * Table by, Burns, loc. cit. S04 METALLURGY OF COPPER electrolyte; it is about 2 in. narrower than the tank and furnishes a r-in. space on either side for the electrolyte. The form of the upper part of a starting cathode depends upon the manner in which it is suspended from its cross-bar. This is always of copper, usually I in. in diameter and flattened at the ends to furnish the necessary contact surface. Formerly the cathode was a rectangular sheet of which one end was bent 180° C. and hooked on the bar; at present there are usually two strip-like extensions at the suspension end, as shown in Fig. 507, by means, of which the Figs. 505-506.— Tank details of Great Falls. cathode is connected with the cross-bar^ Ordinarily the strips are made suffi- ciently long to serve as flat hooks after having been bent 180°. At Anaconda, Figs. 530-534, the upper rim of the small (11X33 in.) starting sheet used to be bent and clamped over the ends of a loop of sheet copper, 11X4.5 in., by a machine similar to one used in making stove-pipe. At Great Falls and Perth Amboy (Raritan No. 2), the Morrow clip, Figs. 508-509, is in use: a loop of deposited copper is fastened to the starting sheet by a machine which on one side punches a hole through the loop and sheet, bends over the protruding ends at the opposite side and forces them down. At Maiirer, N. J., two holes ELECTROLYSIS OF COPPER 505 are punched through loop and sheet to insure against accident. Whatever may be the manner in which the starting cathodes are suspended, care must be taken to have them straight before they go into an electrolyzing vat; after having been in a vat for two days they are taken out singly, placed on an in- clined wooden plane, held on a movable wooden support, and straightened with suitable wooden beaters. The cathodes of a tank are removed together by means of an overhead electric crane after seven to fourteen days, dipped into hot water, and then de- Solution Line I SI. -TL =s :rj -2V Fig. 507. — Starting-sheet with exten- sions to be bent over. Great Palla Slarfing Sheet with "Morrow Clip Figs. 508-509. posited on a transfer car to go to the refining furnace to be melted, fire- refined, and cast. The cast copper generally shows an electric conductivity Jower by i to 1.5 per cent, in comparison with the cathode copper.' This is due to occluded electrolyte and anode mud, and to the absorption of impurity (S) from the furnace fire-gases. 257. — ^JVfanipulation of Electrodes.^ — An electrolyzing vat holds at present 2o-|- anodes and one cathode in excess of the number of anodes; as many as 60 electrodes are permissible. Formerly each electrode was handled by itself; at present the anodes of the vat are charged and removed together, as well as the cathodes. Thus, as already indicated, the anodes for a tank are stacked up- right on a skeleton transfer car. Figs. 500-502, which is hauled by horses, steam, or electric locomotives to the tank house. Here they are raised together by an overhead electric traveling crane, transferred to the tank for which they are intended, and lowered. The remains of the corroded anodes of a tank are removed in the same way, transferred to a washing tank where they are brushed to free them from anode mud, and removed to the anode furnace room. The starting sheets, prepared in a division of the tank house, are also stacked, • Emrich, Tr. A. I. M. E., 1912, xliii, 453. 'Editor, Eng. Min. J., 1911, xcii, 50- So6 METALLURGY OF COPPER taken by the crane, and lowered in the vats. The finished cathodes are handled similarly. At Great Falls the cathodes are first dipped into water and drained, then 30 per cent, of the sheets making up a charge are dipped into milk of lime and allowed to dry. The coat of lime protects the plates from the sulphur in the fire-gases while melting. The ELECTRODE DISTANCE is Usually about 2 in.; the fall in potential be- tween a pair placed in series is about 0.3 volt. Extremes in electrode distance are 1.5 and 2.5 in., the larger figure will be found with anodes that are impure or that run high in precious metal. If the distance is too small, fragments of copper, anode mud, excrescent growths, etc., tend to bridge the space and cause short-circuiting; if it is too large, there is an unnecessary increase in resistance, less copper is deposited, and electric energy is lost by being con- verted into heat. 258. Depositing Vat. — Formerly the depositing vats were arranged in single rows, Figs. 488-489. ^The vats, Fig. 510, were built almost universally of 3-in. Fig. 510. — Depositing vat. planks connected by rabbet, a, tongue and groove, b, or feather, c, joints; threaded tie-rods, d, passed at the ends through the projecting side planks, I^'igs- 505-506. This method is still common with vats that stand isolated. Sometimes the vats used to be coated with asphalt, rubber, or some other im- pervious paint; but more generally they were, and are to-day, lined with 6- or 8-lb. lead which extends over the top, Fig. 506, to prevent the wood from absorbing the electrolyter\ In the newer Eastern plants of the United States, the Walker tanks and their arrangement in blocks have met with general favor. The details of the tank construction at the Raritan works are given in Figs. 511-515. The inner boards are only 1.75 in. thick, the outer planks the usual ELECTROLYSIS OF COPPER 507 3 in.; both boards and planks are connected by feather-joints. In a block two adjoining tanks are separated by an open space 1.75 in. wide with air-holes for ventilation; 3 tie-rods pass through such a space tying channel-iron buck- stays, while in the two outer tanks of a block the rods are placed in openings drilled through the planks. The side walls do not extend to the top of a tank, but Figs. 511-515. — Walker system of depositing vats. are morticed in a cap, 9.25 in. high by 5.25 in. wide, which carries a maple board, 7s by i| in., supporting a triangular bar, i| in. at base; the strong cap carries the large weight of the electrodes. The tanks rest on longitudinal sills, 8 by 8 in., standing upon isolating glass plates, 14 by 14 by i in., carried on 12- by 12- in. brick posts laid in cement. A tank bottom is usually lined with j-in. boards, So8 METALLURGY OF COPPER placed lengthwise, to protect the lead-lining from falling pieces of anode. At each end of the tank is a cross-board, held in place by brackets, which prevents the longitudinal boards from being floated. The tanks of the Lithgow plant, N. S. W.,^ given in Figs. 493, 516-518, are 4 ft. 2.75 in. by 2 ft. 6.25 in. and 2 ft. 9.5 in. deep; they are arranged in single rows of three tiers, each tier forming a cascade of eight tanks with 5 in. be- tween adjoining tanks. U— lOK"-* Top Plan View — MJ# >j xli Batten OS Top of lank to luppoct Agltntot LeaA OTettlotr Hpe t 4"z i Lead Lqiudtt ll.Elli'iSj" Position of Cathode Sheets Position of Anode Figs. 516-518. — Depositing vat of Lithgow, N. S. W. Some interesting features of construction are to be noted. The tank is made of Oregon pine; there are an outer and an inner box, 3 and 2.25 in. thick, between which is placed the 6-Ib. lead lining; the planks have been soaked for 10 min. in paraffine wax held at 60° C. The tanks rest on wooden longitudinal bearers, 5 by 3 in., carried by glass insulators which are supported by wooden blocks, 4 by 4 by 3 in., standing upon brick piers, 9 by g in. The length of a tank depends upon the number of electrodes it is to hold. • Blakemore, Tr. Aust. Inst. Min. Eng., 1912, xv, 36;' Eng. Min. J., 1910, xc, 717, 769. ELECTROLYSIS OF COPPER 509 A common figure is perhaps 10 ft. with 20-22 anodes; at Chrome' there are at present 26 anodes in a tank. The width and depth vary with the size of the electrode. Ordinary dimensions are 2 ft. 8 in. and 3 ft. 9 in.; in recent years^ 'Walker, Min. Ind., 1910, xix, 218. 'Walker, Eng. Min. J., 1911, xci, 41. Sio METALLURGY OF COPPER anodes have been increased from 3 ft. wide by 2 ft. deep to 3 by 3 ft., which in- creases the necessary depth of tank by 12 in. In all modern plants the tanks are placed on piers so as to leave head-room of about 9 ft. beneath for ventilation, discovery of leaks, etc. The basement floor is made acid-proof by being built of chemical brick with joints filled with pitch; it slopes toward troughs ending in a sump. A common electric connection of the depositing vats is shown diagram- matically in Figs. 488-489. It is typical for tanks arranged in single rows. Figs. 521-522. — Thofern electric connection. The advantage is ready accessibility of contacts; the disadvantage, require- ment of much floor space. Great Falls, Figs. 519-520, and Lithgow, Figs. 493, 516-518, may serve as examples. The Thofern,' also intended for single rows of vats, is shown in Figs. 521-522. Two busbars are placed on the tops of the sides of a vat. An anode is suspended from two positive cross-bars and a cathode from two corresponding negative bars. No isolating blocks are required; the corrosion of the anode is said to Fig. 523. — Arrangement of vats in pairs. be more even, and the deposition upon the cathode more uniform. Whatever truth may lie in these claims, the facts are that an excessive amount of copper is tied up, and that twice as many contacts have to be kept bright as is necessary in the single-busbar arrangement. This method is the one favored in eastern refineries, as it gives a good contact, and locks up the least copper. The bars have to be well isolated from the vats; their supports are therefore made non- ' Hering, Berg. HiiUenm. Z., 1893, Lii, 54. ELECTROLYSIS OF COPPER Sii conductive by soaking wood in paraffine or similar substance, or by using glass. The cross-sectional area of a bar depends upon the amount of current that is to pass through it. With rectangular bars i sq. in. for 300 amp. used to be the standard; this was raised to 550 amp., but even with 1000 amp. they remain sufficiently cool; with triangular bars 400 amp. at the contact gives no trouble. The cross-bars from which the electrodes are suspended are usually round bars of copper, i in. in diameter, flattened at the ends. Formerly they .were made of soft steel copper-plated. 259. Corrosion of Anode. — According to their places in the electrochemical series, the metals electropositive to copper ought to go into solution before the copper, but with an anode of 98 per cent. Cu and over, there can be only a ten- dency in this direction. On the surface the more positive metal will be dissolved first, but the copper will follow closely. As the positive impurities as well as the CujO are not evenly distributed, and parts of an anode will have cooled more quickly than others and become harder, the anode will not be evenly corroded; corrosion-pittings or hollows form; the free acid also acts chemically, especially at the contact of copper, electrolyte, and air. The result is that the anode becomes honey-combed, and even spongy if it was impure. Particles of copper fall off and collect in the anode mud. The purer the anode, the more even the corrosion. Usu- ally the anode is removed in from 2 to 3 or 4 weeks, i.e., before parts have become so thin that there is danger of a piece falling and "thereby injuring the bottom lining of the vat or causing a short cir- cuit, or both. In most cases the bottom of a vat is protected by boards (page 507). The weight of the corroded anode which goes back to the anode furnace amounts to about ro per cent, of the original anode. Such anode scrap is placed in a hot-water tank and scrubbed with long-handle brushes to remove adhering anode mud. The necks are then cut off with shears to allow more compact charging in the re- verberatory furnace. 260. Deposition on Cathode. — The purity of the cathode copper depends upon the purity of the anode, the slow- ness of deposition, the constancy of current, the composition, clarity, tempera- ture, and circulation of the electrolyte, and lastly upon the prevention of anode matter coming mechanically in contact with the cathode. With a low current density the surface is crystalline,^ solid, free from pin-holes, blisters, streaks, 'Huntington, Eng. Min. J., 1905, Lxxx, 1109. ^^! ^^^^^\ ^'^ " iC^^^^^^^^H ^wH r ^iv/' - SKUjilwIirt ' "^-^l^^^m Bftv w % '^^Rjv^rwS^ ^a Htt^t. »i^ "!^HHMHHHiH m^ ' Fig. 524.- -Electro-deposition crystals of copper. 512 METALLURGY OF COPPER etc. With the usual high current density, the deposit, smooth at first, soon becomes rough, shows knobs, especially near the edges, unless an addition of o.ox per cent, of glue is made, which keeps the deposit smooth until the cop- per-content of the electrolyte falls below 2.5 per cent. This is accomplished, however, at the expense of voltage, which grows, probably on account of an increased transfer-resistance. The increase in voltage may reach 20 and even 30 per cent, with the usual extremely small amounts of glue charged. In working with an insoluble anode, i.e., with a potential of 2-2.5 volt, in purifying fouled electrolyte, there are formed at the edges of the cathode large crystals' as shown in Fig. 524. / / ^ ^ / / ^ 4 / y^ y^ / ^^ t^ / y- ^ / • y / , / / / / 0.2 0.4 0.6 0.8 1.0 1.2 1.4 1.6 1.8 2.0 2.2 2.4 2.6 2.8 , Per cent Au and Ag Present in Anodes which is Lost in Cathodes 3.0 Fig. 525. — Relation between current-density and precious-metal losses (Addicks). The deposition of copper in the form of tube, sheet, or wire is carried out on a large scale in Europe.^ A discussion of details lies outside of the scope of this treatise. The impurities in the cathode copper are either electrodeposited or mechan- ically occluded.* Under normal conditions no impurities are electrodeposited. Electrolyte enclosed between coarse crystals of copper does not play an im- portant r61e, but if fouled by i per cent. As, it assists in the reduction of the electric conductivity. Cathode copper rarely contains n 1 00 00 M M (*) m M in 000 t- '* o> ^ •* f*J N WHO i 00 d 00 N 1- 00 d d 00 \o d H fn a •* N M a d ^d ri ro d M 06 »n 1 00 r- H r*5 MO H ro o> m 00 ■* H § & N t> 00 N 00 N m m QO in 10 ■* 5 a! •0 00 00 WON W f*5 H in 00 M in n vO m M d ro m r*) m 0, w N rO M -* 00 00 m fo N d Tt M m N d %d t^ M d d g S en ■N m ro CO H Oi r- 1^ 4 H n £ S OOOmO 000000000000 ot-ooo oooo-orroooi-O'O Tt-OONO-O OfOMoGO'OOiOOr^HN rOHHOr- NfOMOOf^OWHHWO rot^OwO f^mOOOOOMwOroo Oi m d c 00 M m vO 000 N -O I- M M " Oi M m m m fo w m Ih ... l-l " 5 < °< J3 ^ < C£ 2 £ eg w 6 CO D Q M i > •c ^ a M 1 „ Fl l=! n S rrt u (U 1 > £ii Oi u .3 % t1 ? ■a ■n o\ < a M « u Wl 10 n 8 > Hi X M 8 ^ S •^ rO 8 ^ bq s M ^ k. Id S U ►-H ■-' M T1 r^ • ■d -^ f:;< .; 3 !? ^ FQ " H " B -c ELECTROLYSIS OF COPPER 515 particles of anode copper; the screened mud will then be raised to a storage tank by a centrifugal pump, after having been stirred up occasionally with an oar-Uke rod to prevent its packing down. 262. Treatment of Anode Mud.— Of the various methods suggested for treat- ment^ four only are at present in use: direct fusion, lead-soaking, solution in H2SO4 in the presence of air and NaNOa, roasting followed by solution in H2SO4. (i) Direct Fusion. — ^This is carried on in a reverberatory furnace with movable hearth fired with two oil-burners. The oval hearth is 8 ft. long, 6 ft. wide and 10 in. deep; consists of a cast-iron bed-plate and a |-in. wrought-iron ring; and is rammed with a mixture of 2800 lb. dolomite, 1800 lb. cement, and 1150 lb. fire-clay to a thickness at the sides, front, and back of 12 in. and at the bottom of 4 in. The side lining encloses a 2-in. water-cooled pipe. The stored mud is stirred up, drawn into a pressure-tank, and filtered in a press 3 ft. in diameter with 14 plates. The filtrate passes through 8 settling tanks placed in series, which collect any mud that may have passed through the cloth, and flows into the storage tank of the tank room. The 3-in. cakes are discharged into a car, to which is added soda-ash and siliceous ore rich in precious metal. The mixture is fed into the furnace at intervals at the sides near the burners; the slag flows off at the front. A furnace treats in 24 hr. about 5 tons of mud (Table 122, New Jersey), burns 500-600 gal. oil, makes 2.5 tons slag (Table 124, New Jersey), o.g ton flue-dust (Table 125), and dore silver, 0.820-0.860 fine. The slag goes back to be smelted, the silver is further purified in a lead cupelling furnace, and the flue-dust is passed through a suspended sheet-iron oval flue (4 by 3 ft. and 200 ft. long) with discharge doors in which the temperature is reduced to 125° C. before the gases enter a wet scrubber which collects most of the remaining dust and fume. The collected dust is treated in the dry way, by mixing with litharge and smelting in a reverberatory furnace, or in the wet way (see below). Direct fusion is applicable only in a refinery that is connected with a lead plant to take care of the intermediary products. (2) Lead-Soaking. — The filter-pressed cakes of slime are partly dried and charged in paper bags, from 10 to 15 lb. at a time, onto the lead bath of a cupel- ling furnace, when the copper and other impurities are readily oxidized and scorified. As coppery Utharge has a strongly corroding effect, the cupel should be water-jacketed and in addition lined with magnesite brick. (3) Solution in H2SO4 with Air and NaNOs. — The steps taken are: solu- tion of copper with disposal of blue vitriol; filtering, washing, and drying of residue; oxidizing fusion in a reverberatory furnace with blast and niter, and withdrawal of slags until the silver or dore silver is fine;^ and casting into marketable form or into plates suited for parting. A few examples may serve to show details: Example I. — The slime is screened through a i-in. and then an 8-mesh ' Ulke, Min. Ind., 1893, 11, 278. 'The alloy-series Ag-CuuO has been studied by Mathewson and Stokesbury, Internal, Zt. Metallographie, 1914, v, 193. Si6 METALLURGY OF COPPER sieve placed on a circular stationary vat to remove coarse anode copper; the remaining metallics are screened out by passing the mud through a centrifugal machine with a 40-mesh sieve. The slimy mud assaying Ag 50 and Au 0.5 per cent, is transferred into an oblong wooden lead-lined vat on the bottom of which lies a zig-zag lead-pipe perforated on the upper side. Dilute H2SO4 (1:4) is drawn into the vat and heated air injected through the pipe. This stirs the mud, heats the solution and furnishes the necessary O. Solution and mud are discharged into a settling tank; the clarified liquor is siphoned off; the mud is washed three or four times by decantation, filter-pressed, and washed until sweet. The cakes from the press are partly dried and melted in a water- jacketed cupeUing furnace. A small amount of slag, assaying Pb 20 per cent, and Sb < 10 per cent., is formed and drawn off; then niter is charged, which forms a slag with as much as 20 per cent. Te.^ This is drawn off, and more niter is added until the silver is fine and ready for casting. Example II. ^ — The slime is screened through a copper plate with i-in. holes and a lead plate with slots equivalent to a 40-mesh wire screen, into a tank, whence it is transferred by a centrifugal pump to three storage tanks 4 by 4 by 10 ft. in the silver-refinery,^ where the slime, assaying 50 per cent. Ag and 30 per cent. Cu, settles, while the overflow goes back to the tankhouse. The mud is transferred into seven circular hard-lead agitators, 6 ft. in diameter and 5 ft. deep, with vertical steel lead-covered shafts and hard-lead paddles making 15 r.p.m. A charge consists of 6,000 lb. wet mud ( = 3,000 lb. dry), 2,600 lb. H2SO4 66° Be., 250 lb. niter, and water; it is heated with live steam and air is pumped in to assist the oxidation. Boiling, which usually lasts 8 hr., is continued until silver goes into solution, when the necessary correction is made by adding fresh mud. The solution (Cu 4-5, As 0.4, free H2SO4 15 per cent.) and decopperized mud are drawn into a lead-lined cast-iron acid egg and forced through a Bushnell filter press, 27 by 27 in., with 40 plates and frames of hard lead. Table 123 gives analyses of raw and of treated mud. Table 123. — Analyses of Raw and op Treated Mud Mud Se Te As Sb 1 Bi Pb 1 Cu Pe Ni S SiOi Ag Au I Raw".... 3.82 4. 10 7.13 16.21 |1.62 10.94 12.86 0.17 0.09 1.73 0.S3 26.58 0.36 Treated.. 4.20 4. 10 7.73 19.81 jl. S3 13.18 2.81 U.09 u.09 1.92 0.40 30.72 0.42 11 Raw". . . . 2.30 3.56 S.42 6.76 ,0.4s 3.58 12.26 0. 14 0.03 i.15 "93 53.68 0.28 Treated. . 2. 54 4.28 3.61 7.21 jo. 54 4.10 3-02 0.13 0.03 1.29 1.02 63.36 0.31 The mud, which usually contains < i per cent. Cu, is washed, discharged, and smelted with the addition of niter in an air-blown, oil-fired (ij-in. Rockwell burner) reverberatory furnace* with a hearth 8 ft. 3 in. by 5 ft. 7 in., lined with 9 in. magnesite, sides and bottom. A charge of 17,500 lb. moist (?o per cent. H2O) slime is smelted in 48-72 hr. with 2,000 lb. niter and 200-250 lb. soda- ' Whitehead, /. Am. Chem. Soc, 1895, xvn, 849. ' Raritan: Addicks, Min. Ind., 1900, rx, 261; Private Notes, 1907 and 1913. ' Plan given in Eleclrochem. Met. Ind., 1908, vi, 277. " Have been given in Table 122. * Drawing, Eleclrochem. Mel. Ind., 1908, vi, 277. ELECTROLYSIS OF COPPER S17 ash, with a consumption of 200-250 gal. oil; it furnishes about 5,000 lb. dor6 silver of a fineness of Ag 0.984, Au, 0.15, Cu o.i, which is cast into anodes 11.5 by 7.5 by 0.5 in. to be parted electrolytically by the Balbach-Thum process. (4) Roasting and Solution in H2SO4.— The slime freed from metallics by a 40-mesh sieve is filter-pressed; the disintegrated cakes are charged into a muffle furnace and heated to oxidation, which begins below a visible red. When once started, the smoldering roast progresses of its own accord. "• The roasted slime is pulverized, treated with hot dilute H2SO4 for 8-10 hr. in a tank, say 8 ft. in diameter and 8 ft. high, having a mechanical stirrer. A total of 13,000 lb. leached slime is melted in an oil-fired reverberatory furnace with hearth 5 by 12 ft. and 15 in. deep, lined with magnesite brick; it takes 30 hr. to melt the material, charged at short intervals. Niter flux from a preceding charge is added to assist in liquefying the slag, which is skimmed and goes to the anode furnace (3000 oz. per ton) or the ore-furnace (1000 oz. per ton). There forms on the metal bath a dark cherry-colored selenide of copper and silver (Cu 33, Ag 33) Se 33 per cent.) which is oxidized by rabbling and by forcing air through the bath by means of an iron pipe.^ The flue-dust collected, rich in precious metal, contains from 30 to 50 per cent. Se and Te, both of which are recovered. To the bath is added soda-ash to purify the metal. The slag formed is skimmed, and fresh flux is added until the slag ceases to become dark, whereupon soda-ash and niter are charged to fine the silver. After the first 13,000 lb. of decopperized anode mud have been thus treated and a bath of more or less pure silver has been obtained, more mud is charged and refined until the furnace has received its complement of 22,000 lb. mud, the total treatment of which takes four and one-half days and furnishes 7000 lb. dore silver. Table 1 24 gives analyses of reverberatory slags obtained in the fire-refining of anode mud, and Table 125 analyses of flue-dust and scrubber-sludge. Table 124. — ^Reveeberatory-slags from Anode Mud No. As Sb Se 2.3 Tc 1.2 Cu 8.4 ' iS-4 9.6 7-3 6.56 1 Pb Bi Ag Au I II 3-8 7-3 10.4 14.6 12.2 7.0 14-05 ■2-25 11.604 8.747 3.220 5-396 2.28 0.0003 0.0014 III 0.0031 0.0048 0.0063 IV 1 V 3 -OS 14-3 2-3 4-5 Table 124. — Reverberatory-slags from Anode Mud — {Continued) No. SiOs Insol 1 FeO CaO MgO AI2O3 Ni s Mn I 14.2 13-3 16.9 15.0 14.4 34-0 2-5 1-43 II III IV 1.8 2.25 0.8-2.1 4-75 V 2.34 2.50 0.03 0.27 trace ■ See also, p. 478. ' Behavior of Ag and O: Kohlmeyer, Chemiker Z., 1912, xvni, 151, 211. Si8 METALLURGY OF COPPER Table 125. — Analyses or Flue-dust and Scrubber Sludge from Refining Anode Mud As Sb Se Te Cu 1 Pb Bi Ag Au SiOj 1 FeO CaO Nj S Flue-dust. 4.80 9.00 12.18 13-90 25.30 39.01 S.20 3.00 4.50 3.10 1.60 3.00 1. 00 0.70 4.018 0.002 Flue-dust Scrubber sludge. . . . 0.61 9.34 2.88 1.98 0.013 0.3s 8.10 0.40 trace 0.97 The gases and fumes from the reverberatory furnace treating anode mud are always passed through cooling flues and thence through scrubbers. At the refinery of the Raritan Copper Co. the gases leaving the scrubbers are further treated in a Cottrell electric condenser.^ In its leading features it resembles the apparatus used at theBalaklala smeltery, discussed in connection with Figs. 246-247; in detail it differs to meet the new conditions. Figs. 526-528 give a plan and horizontal section, a section and side-elevation, as well as a section and end-elevation of the condenser with its connections. It consists of a lead-lined plate-iron tank, 18 ft. long by 8 ft. 55 in. wide and 6 ft. deep; the bottom is hopper-shaped to catch and deliver the sludge to a settling tank; the top is covered with planks provided with handles near the ends. The tank has two sets of 14 longitudinal parallel iron bars, about 3 ft. 6 in. long, resting on cross- pieces connected at the ends with the two branches of the positive pole. From each bar are suspended rod-shaped iron positive discharges about 5 ft. long. The number of positive poles, 26 used at the start, has been reduced to 18. Parallel with the longitudinal positive bars are suspended sheet-lead negative grounded poles; eight of these plates are held in the slots of two lead-covered rectangular rolled iron bars. The motor and alternator are in separate buildings. They are combined as a rotary converter, the direct-current end being wound to 220 volts, the alternating end to about 158 volts, single phase. The rotary converter is rated at 2 kw., and is mounted on a shaft. Connected to the rotary converter shaft is a micanite disc carry- ing two segments of copper. The brushes carrying the current to the rectifier have a rotary adjustment for properly regulating-over the lag of the current. The transformer is oil- cooled, 7.5 kw. capacity, with 10 taps for transforming the current from 35,000 to 50,000 volts. The input direct current is about 10 amp., and, of course, 220 volts, while the rectified current, is a few milliamperes at about 40,000 volts. The condenser is designed to treat 10,000 cu. ft. gas per minute; at present about 6500 cu. ft. gas with 0.0258 g. solid matter per cubic foot pass through it; a clearance of about 90 per cent, of solids is effected. The condensed matter runs high in Ag and low in Au. 263. Recovery of Tellurium and Selenium.— It was shown on p. 516, that the niter slag obtained in refining anode mud in the reverberatory furnace con- tained as much as 20 per cent. Te. There are two ways of recovering the Te although they are not much carried on because so far there is little industrial use for the metal. One method consists in dissolving the slag in hot water, filtering and evapo- rating, if necessary, to the right degree of concentration, and then electrodepos- iting the metal, using lead electrodes. ' Hofman, "General Metallurgy," 1913, p. 858. ELECTROLYSIS OF COPPER 519 __ Fig-. 528 Section and End Elevation Figs. 326-528.— Cottrell electric condensation of the Raritan Copper Co. 520 METALLURGY OF COPPER The other method is based upon the reaction which takes place when the filtered hot alkaline aqueous solution is just acidified with H2SO4, and causes Na2Te203 to fall out. From the filtrate, Se can be precipitated by means of SOa-gas. Most Se on the market is produced from the grayish flue-dust col- lected in the cooling flues of the reverberatory furnace treating anode mud. Scrubber liquors which contain Se must be concentrated to about 20° Be. before they can be treated for Se; the present price of the metal does not warrant treating the liquors. Flue-dust with 20-25 per cent. Se is treated in charges of 70 lb. in a 125-liter jar with 15 lb. commercial HCl, 10 lb. KCIO3 or NaNOs, and water, stirred, and allowed to settle. When clear, the supernatant liquor is siphoned off, the residue washed with H2O by stirring, settling, and decanting or siphoning until the water- wash has a density 5° Be. or less. To 4 vol. of Se-solution are added 2 vol. of commercial HCl, or enough to have 15 per cent, free HCl; the whole is warmed and SOa-gas, generated by burning S, passed through the vat, when at about 80° C. practically all the Se will be precipitated as a reddish powder. The temperature of the solution rises during the precipitation. The precipitate is settled, filtered through a wet filter, washed with cold water, and dried quickly in porcelain dishes on a hot steam table. The red precipitate shrinks to a bluish to brownish-black cake, and squeezes out some water charged with impurity. The dried Se is placed in a kettle, heated for about 4 hr. on a stove the top of which becomes barely red hot. In melting, a crust forms on the fluid metal; the latter is poured off into an enameled mold, the crust is brought to a higher temperature whereupon more fluid Se is eliquated; finally the crust it- self is liquefied and cast into molds. Details of the practice at the few plants producing Se vary considerably and are not available. ,/ 264. Foul Solutions.' — The electrolyte in time becomes overcharged with impurities such as As, Sb, Bi, Ni, Fe, which interfere with the quaUty of the cathode deposit, and with the blue vitriol which is recovered by crystallization. A fouled electrolyte from Great Falls^ contained per liter, Cu 51.8, Fe 13.2, As 14.02, Sb 0.62, H2SO4 48 g.; the low percentage of Sb is due to the addition of enough HCl to the head tank to maintain 0.04 g. CI per liter in the solution. There are two general methods for pmrifying the electrolyte; one is direct removal of impurities by precipitation, the other withdrawal of part of the foul electrolyte and its replacement by blue vitriol and the necessary free H2SO4 while the foul solution is being treated. The direct removal of impurities by precipitation has not been sufficiently successful in practice to become adopted. Thus, boiling with metastannic acid in order to precipitate As; filtering through oxidized granulated copper to throw down Sb and Bi; blowing air through the solution heated to 35° C. and over, to 1 Ulke, Min. Ind., 1897, vi, 239; Zl. Eledrochemie, 1898, iv, 309; Berg. Huttenm. Z., 1898, LVII, 264. Burns, Tr. A. I. M. E., 1913, XLVI. 2 Hofman, Tr. A. I. M. E., 1904, xxxiv, 312. ELECTROLYSIS OF COPPER 521 oxidize FeS04 and cause it to fall out as a basic ferric sulphate, accompanied perhaps by Sb- and Bi-salts, have all been tried and given up again. The general practice is to withdraw continuously one or more per cent, of the electrolyte from the main stream, purify it, and allow fresh electrolyte or puri- fied solution to flow into the head tank. The following methods of purification may be considered as covering the ordinary modes of operating: Crystallization, suited for an electrolyte practically free from Ni. Crystallization followed by electrodeposition, suited for an electro- lyte heavily charged with Ni. Electrodeposition, suited for an electrolyte lightly charged with Ni. (a) The Crystallization Method. — The free acid is neutralized, or rather reduced to below i per cent., by dissolving in it granulated copper in the presence of air, Vitriolization Process, § 243. The neutral or slightly acid liquor is concentrated in lead pans by steam coils to 43° Be. and then crystallized, whereafter, in the absence of NiS04, from 85 to 90 per cent, of the blue vitriol can be recovered of a sufficient purity to permit re-solution and crystallization for a marketable product. Following is an analysis of impure crystals:^ Cu 32.78, Fe 0.589, Ni 0.0496, Pb 0.0122, Bi 0.0640, Sb 0.2920, As 0.2470 per cent. The copper remaining in the mother liquor is recovered in two ways. It is precipitated by Fe with the As and Sb that is present; the first precipitate is kept separate from the last, as this may contain as much as 60 per cent. As and form the raw material for the manufacture of arsenical compounds. In the second method the liquor is concentrated to the point at which both sulphates of Cu and As crystallize together; the crystaUine mass, separated from the mother liquor, is treated with just enough water to dissolve the copper sul- phate; the residual arsenical salt is suited for the production of copper-arsenic salts. Drawings of the Sulphate Building of the first electrolytic refining plant of Great Falls, Mont., have been published by Burns {loc. cit.). As regards the separation of CUSO4 and NiS04 by fractional crystallization, it is held^ that if an acid solution at 35° C. contains an excess of CUSO4 over a presupposed cryohydrate of 2NiS04+a;H20: iCuS04+3'H20, CUSO4-I-ZH2O will separate, but that with an excess of NiS04, there will be formed crystals of NiS04+7H20. The ordinary CUSO4+SH2O' gives" up 2 mol. H2O at 28° C, and NiS04+7 H2O loses 2 mol. H2O only, at 40° C. (b) Crystallization Followed by Electrodeposition.— The electrolyte is concentrated to 46-47° Be. in wooden lead-lined tanks with steam coils carried by cast-lead frames. A tank 16 by 15 ft. and 4 ft. 10 in. deep with 350 ft. of 1.5 in. lead pipe will evaporate in 24 hr. 1000 cu. ft. liquor of 1.240 sp. gr. to 340 cu. ft. 1.460 sp. gr. The concentrated solution is run into a crystallizing tank ' Keller, Min. Ind., 1898, vii, 238. ' Private Communication by C. S. Witherell, 1912. ' Hofman-Wanjukow, Tr. A.I. M. E., 1912, xuii, 523- 522 METALLURGY OF COPPER constructed to cause a rapid separation of crystals. Such a tank 12 by 6 ft. and 4 ft. deep with eight cross- timbers, 6 by 6 in., each carrying eight rows of vertical zigzag i-in. water-cooled lead pipes will cause 82 per cent, of the copper to crystallize in 48 hr. in the form of small crystals which carry some NiS04+aq. The mother liquor goes to the liberating tanks; water is run into the crystallizing tank, steam is turned on, the crystals are dissolved, mud is allowed to settle, and the liquor run to the tank house. The liberating tanks receiving the mother liquor are depositing tanks, with sheet-lead anodes and sheet-copper cathodes, in which all the copper and some of the arsenic are deposited in a suffi- ciently coherent condition to permit scraping off the deposit with a chisel-pointed bar, and are turned over to the smelting department. A liberating tank has the same dimensions as a depositing tank. There are usually provided 1.5 to 2 liberating tanks for every 100 depositing tanks. The electromotive force required ranges from 2 to 2.5 volts. In order to hold back in the bath the choking fine particles of H2SO4 which are' carried off into the air with the O liberated at the anode, the electrolyte is cov- ered with a layer of oil. In some instances accidents have occurred on account of the formation of AsHs; hence at several works these tanks are placed in an open shed. The further treatment of the copper-free solution containing, e. g., Ni i-|- per cent., 'As i per cent., free H2SO4 15 per cent., varies somewhat. The As206-salts are first reduced to the As203-stage with S02-gas by allow- ing the solution to run down one covered cascade in which the gas ascends, and then AS2S3 is precipitated by H2S in a second cascade. The arsenic-free solution is concentrated in a series of lead pans (twelve, 6 by 4 ft. and 13 in. deep) followed by iron pans (four round-bottomed of the same dimensions), with the fire-place beyond the last iron pan. The fire-gases pass under the iron pans and then under the lead pans, while the liquor flows in the opposite direction leaving the last iron pan at a concentration of from 72-75 per cent, free H2SO4 to be collected in an iron pan where, upon cooling, NiS04 contaminated with 1-2 per cent. FeS04 falls out. The nickel-free acid goes to the tank house. The NiS04 and FeS04 are dissolved in H2O, the FeS04 is oxidized with CaC104 +CaCl2 in the cold and precipitated with CaCOs. The purified NiS04 is evap- orated to dryness in a pan, and calcined in a reverberatory furnace to drive off H2O and SO3. The NiO is mixed with charcoal and smelted in an oil-fired reverberatory furnace at the rate of 1.5 tons in 24 hr. and is either cast into ingots in upright split molds, or shotted after the C it had absorbed has been removed by additions of NiO. Instead of concentrating the copper-free Hquor in a single operation, two steps are taken, steam-concentration to 50° Be, and direct-fire concentration in a V-shaped boiler-iron tank to 72 per cent. H2SO4. The resulting NiS04 may be freed from part of its H2SO4 by placing on a quartz filter, transferring to an inclined lead-covered drainage-floor, placing on a perforated lead plate provided with suction, and adding a small amount of wash- water. The crystals will contain Ni 31 per cent, and free H2SO4 8 per cent. The NiS04-crystals ELECTROLYSIS OF COPPER 523 from the solution with 72 per cent. H2SO4 may be freed from most of their H2SO4 by washing with a little water in a centrifugal machine. Copper- and nickel- free concentrated acid may be freed from As by boiling with charcoal, which reduces AszOb to AS2O3, and precipitating by diluting to 40° Be.; the AsaOs settles readily and with it the coloring C^H,, formed in the treatment with charcoal. At Great Falls, Mont.,i the electrolyte is purified in the following man- ner. From the 320 tanks (9 ft. 7 in. by 2 ft. 4 in. and 3 ft. 9 in. deep) there are withdrawn daily 25,000 liters electrolyte to be purified. The solution, concen- trated to 48° Be., is drawn into crystallizing tanks, and remains there 4 days, during which 82 per cent, of the Cu crystallizes out. The resulting mother liquor, containing H2SO4 47S> Cu 17.4, As 20.2, Sb i.i, Fe 15.2 g. per liter, is electrolyzed in four purifying tanks, of the above dimensions, containing lead anodes and copper cathodes. With a circulation of 7 liters or nearly 2 gal. per minute (depositing tanks have one of 6 gal. = 22.5 liters) there are removed 99 per cent, of the Cu, 78 of the As, 91. i of the Sb with an ampere-efficiency of so per cent. The more or less slimy cathode deposit contains H2O 9.66, Cu 46.30, SiOa 0.38, FeO 1.66, AI2O3 0.4, CaO 1.08, S 5.02, As 21.48, Sb 2.28, Ni 0.3s, Zn 0.32 per cent., Ag 3.61 and Au 0.03 oz. per ton. The changes taking place in the electrolyte of the four tanks placed in series, with a circulation of 4 liters per minute, are shown in Table 126. Correcting Table 126. — Removal of Cu, As, and Sb from Electrolyte in iNsonraLE-ANODE Tanks (Circulation, 4 literj per min. — 9000 amp; 31.8 amp. per sq. ft. Grams per liter Volts per tank Tempera- ture, deg. C. Tank HjSOj Cu Fe As Sb Inlet tank No. i 144 184 194 208 216 37.060 7-376 0.504 0.088 0.048 6.242 6.813 7-364 7.701 7-915 3.200 2. 240 0.400 0.056 0.028 0.463 0. 260 0.061 0.038 0.028 17 42 S7 64 65 Outlet tank No. i Outlet tank No. 2 Outlet tank No. 3 Outlet tank No. 4 2. 22 2.25 2.25 2.25 Table 127. — Analyses op Table 126 Corrected to Basis or Constant Volume of Electrolyte Tank Grams per liter Percentage elimin- ation of original amounts Ampere- efBciency, HaSOi Cu Fe As Sb Cu As Sb Inlet tank No ± 144 169 165 169 170 37.060 6.760 0.427 0.071 0.038 6. 242 6. 242 6. 242 6.242 6. 242 3.200 2.050 0-339 0.045 0.022 0.4630 0.2380 0.0517 0.0308 0.0220 Outlet tank No. i Outlet tank No. 2 Outlet tank No. 3 Outlet tank No. 4 81.8 17. 1 o.g 0. 1 3S-9 53-S 9-2 0.7 48.7 40. 2 4-7 1-7 71.70 19-50 1.68 0.15 Totals and averages. 99-9 99-3 95 -3 23.26 ' Burns, loc. cil. 524 METALLURGY OF COPPER the analyses for a basis of unchanged volume of solution (in which Fe = 6.242 g. per liter) gives the data in Table 127. This shows that while the percent- age-deposition of Cu, As, and Sb with a circulation of 4 liters per minute is much higher than with one of 7 liters, the ampere-eflSciency has fallen from 50 to 23.26 per cent. The electrolyte freed from most of its Cu, As, and Sb, stUl retains Fe, Ni, B,i and Zn. In order to remove these, the liquor is transferred to a lead-lined tank (13 ft. in diameter and 4.5 ft. deep, lined with 12-lb. chemical sheet lead, provided with 600 ft. of i-in. 8-lb. lead pipe), concentrated to 55° Be., run into an open tank 10 by 4 ft. and 3 ft. deep, allowed to stand for 4 days; during which Fe, Ni, Bi, and Zn will crystallize as sulphates, leaving behind a mother liquor with H2SO4 HOC, As I, Sb 0.2, Fe i, Ni 5.3, Zn 1.5 g. per liter. Originally the electrolyte cut out from the main stream was run direct into the insoluble-anode tanks. The deposit was in the form of a black slime which in part adhered to the cathode and in part collected on the bottom of the tank. Its composition was H2O lo.o per cent., Cu 55.1, Si02 i.i, FeO 0.4, AI2O8 0.4, CaO 0.3, S 4.1, As 10.3, Sb 2.5, Ni 0.35, Zn 0.32 per cent., Ag 3.4 and Au 0.02 oz. per ton. The method was abandoned because the ampere-efficiency was much lower, and the amount of slime produced much higher than when 82 per cent, of the Cu had been first removed by crystallization. (c) Electrodeposition. — In the liberating tanks as much pure copper as is feasible is deposited, leaving only a small amount of impure copper containing most of the arsenic and antimony. This fractional deposition is accomplished by retarding the flow of solution in the liberating tanks in which the arsenic is to come down. The separated arsenical copper, and the impurities, in part adhere to the cathode, and the remainder falls to the bottom and forms a dark mud. The acid freed from As and Sb goes to the tank-house. If it should be too rich in Fe, it is concentrated to about one-third its volume and cooled, whereby most of the Fe will crystallize. 265. Costs. — The cost of a multiple electrolytic plant is great on account of the large amount of copper and blue vitriol locked up. Thus at Great Falls, Mont.,1 with 300 tanks, 9 ft. 7 in. by 2 ft. 4 in. and 3 ft. 9 in. deep, each holding an electrolyte with 3. 28oper cent. Cu, 22 anodes weighing 5001b. and 22 cathodes weighing 2.5 lb., and with a daily production of 174,000 lb. copper, employ- ing a current density of 34 amp. per square foot and 2-day cathodes, there are locked up in anodes 2,300,000 lb. Cu, 44,000 oz. Ag and 316 oz. Au; in slime 227300 lb. Cu, 140,000 oz. Ag. and 850 oz. Au; in cathodes 180,000 lb. Cu; in solutions 95,000 lb. Cu; or a total of 2,597,300 lb. Cu, 184,000 oz. Ag, and 1166 oz. Au. It is generally held that a plant having a daily capacity of 100 tons of copper, casting anodes as well as cathodes, costs about $450,000 excluding the precious metal that is tied up. The cost of refining by the multiple process at Ana- conda in 1897-98 with a yearly output 30,000 tons of copper was 0.75 ' Burns, loc. cit. ELECTROLYSIS OF COPPER 525 cents per pound or $15 per ton of copper produced.' With increase of size of plant and improvement in the methods of handling and of operating, the cost in Eastern refineries with a daily capacity of 200 tons and over is from $4 to $5 per ton of copper, excludmg all overhead charges. The old rule that i ton of coal is required for i ton of cathode copper still holds good to-day. The curve given in Fig. 529, drawn by L. Addicks,^ shows the relation that exists between the current density and the cost of power in the different plants of the United States using the multiple process. 4 i Dollars per K.W.-Year 12 16 20 24 28 32 36 40 44 48 52 56 60 44 40 36 82 \ \ \ V ' 28 1" 12 \ \ \ "~^ -^ - n Fig. 529. — Current density vs. cost of power (Addicks). 266. Examples of Multiple Process. — Three examples may serve to show the general arrangements of multiple plants: Anaconda, erected in 1904; Great Falls, remodeled in 1896; and Raritan No. 2, erected in 1908. (a) The Anaconda Plant (Thofern System),' shown in Figs. 530-534, has 1400 tanks divided into 7 sets of 200 each. Each set, made up of 2 blocks of 100 tanks, receives from its generator a current of 4000 amperes at 60 volt to be distributed according to the Thofern system (Figs. 5 2 1-5 2 2) . * The positive current, arriving between two blocks, traverses one block to the left of the North- South aisle in zig-zag as indicated, passes to the other block on the right of the aisle, supplies its tanks with electricity along similar paths as the first, and leaves 1 Keller, Min. Ind., 1898, vil, 236. 2 See also Met. Chem. Eng., 1914. xii, 91. 'Corresp., Eng. Min. J., 1896, lxii, 271; Hofman, Tr. A. I. M. E., 1904, xxxiv, 308; D. H., Zt. Eleclrochem., 1901, vii, 793! not operated at present. * Fontaine, Eng. Min. J., 1892, liii, 669. Bering, Berg. HiiUenm. Z., 1893, lii, 53. S26 METALLURGY OF COPPER I ! I I I I I M ' I ' Ml ' ' ' I I I I Llj — . Hi C 75, 8S- ELECTROLYSIS OF COPPER 527 the set opposite the place of entrance. The electrolyte, 40 g. Cu and 150 g. H2SO4 per liter, brought to 50° C. in the head-tank of a block by a steam-coil, flows through a 4-in. pipe into 10 communicating distributing boxes. Each box feeds 10 electrolyzing vats arranged in two cascades on either side. The last vats of the cascades on both sides discharge into two collector boxes which, con- nected by a pipe, feed the Pohle air-lift pump which raises the electrolyte to the head-tank. The circulation between the individual vats is effected by drawing off the electrolyte at the side of one tank, 10 in. above the bottom, through an inclined lead pipe, broken and connected with rubber hose to prevent any stray current from passing, and delivering it on the same side near the top of the next vat back of a distributor of perforated sheet lead. With a current density of 10 amperes per square foot cathode area, the circulation is only 3 gal. per minute. The individual vats, 8 ft. 2 in. by 4 ft. 5 in. and 4 ft. 6 in. deep, are of 3-in. planks lined with 8-lb. lead; the bottoms are covered with i-in. boards. In the side of a vat is a cock, 4 in. above the flo6r, to draw off electrolyte, and in the bottom a 2-in. plug to close the opening for removing anode mud (Cu 10 per cent., Ag 18,000, and Au 100 oz. per ton), into a trough leading to a collecting tank, to be screened and forced by an acid-egg into the silver department. The furnace- refined converter anodes, 32I in. long by 241 in. wide by ij and i in. thick, weighing 230 lb. (Cu 99 per cent., Ag 90 and Au 0.5 oz. per ton) are suspended by two hooks in pairs from iron copper-coated cross-bars; they are corroded in 37 days and make 7 per cent, scrap. The starting cathodes, 33 by 1 1 in., weighing f-i lb., are, clamped to strips of copper and suspended in fours from a cross-bar; they are removed when their several weights have reached about 100 lb., or when the anodes have been completely corroded. An analysis shows Cu 99.96, As 0.0009, Sb 0.0023 per cent., Ag 0.25 oz. per ton. The fall in potential between vats is 0.3 volt. The 38 anodes or the 80 cathodes of a vat are put in place and removed together; the cathodes are taken out individually, two days after they have been placed, in order to be straightened. (b) The Great Falls Plant.' — This plant was erected in 1903 and has been in almost continuous operation since then. It has undergone several changes to meet the demands of the increased output of the smelting depart- ments. In 1893 there were produced 3179 tons of copper; in 1912 the product was 31,596 tons, which meets the market requirements west of Chicago, 111.; the greater part of the copper of the smeltery is refined at the works of the Rari- tan Copper Co., Perth Amboy, N. J., discussed below. The flow-sheet of the refinery is given in Fig. 535, and plans and sectional elevations of the tank and sulphate buildings are represented in Figs. 536-541. The tank building contains 20 double rows (Figs. 519-520) of 16 tanks (Figs. 505-506), each with 22 converter anodes with 40 oz. Ag and 0.24 oz. Au per ton (Figs. 498-499) and 22 cathodes (Figs. 508-509). A double row of tanks forms a cascade (Figs. 519-520, and 541). At the head of the tank room is the pump room with three sumps for electrolyte and corresponding Pohle air- 'Hofman, Tr. A.I. M. E., 1904, xxxiv, 308. Burns, op. cit., 1913, XLVI. 528 METALLURGY OF COPPER Table 128. — ^The Multiple Process in the Raritan Copper Co- Perth Amboy, N. J. American Smelting and Refining Co., Maurer, N. J. Electrolyte: 3-5 12.5 Composition | ^^^ ''^^■ ^^^ ^^364 ::::::.:::::. 60 35 Antisell pump Antisell pump Current: 18 12-13 0.30 7000X128 6500X132 4100X125 7300X125 4150X125 Anode: Cu 99-3 Cu 97-98, Ag i.o 36X28Xii t = b 36X32X1 t|">b Weight, pounds ■ 500 450 Mode of suspension Shoulder Shoulder Corrosion, exchanged after dcys 28 28 12-15 14 Cathode: Starting sheet, length X width X thickness, inches 30X37X?^ 38X34X3V Weight, pounds II 11-12 Morrow clip Morrow clip 7 Weight, pounds 125 185 Overhead electric crane Overhead electric crane Deposition vat: I2'X3'4"X3'6" li'X3'6"X3'6" 26, 27 28, 29 - Electric connection Walker Walker Busbar, cross-section 9Xii Amp. per sq. in. cross-section 620 Anode mud: Per cent, of anode 0.6 Composition, per cent Ag-60 Cu-20 Removed, after days 28 14 Referred to on pages: ELECTROLYSIS OF COPPER jElECTROLYTIC REFINERIES OF THE UNITED STATES 529 1' " U. S. Metals Refining Co., Chrome, N. J. Balbach Smelting and Refining Co., Newark, N. J. Calumet & Hecla Mining Co., Buffalo, N. Y. Boston & Montana Consolidated Copper and Silver Mining Co., Great Falls, Mont. Anaconda Copper Mining Co., Anaconda, Mont. 3.0 12.0 16.0 4.5 13.8 3-3 13.0 4.0 15.0 54 50-57 49-40.5 64 SO 35 4 3 6 3 Antisell pump Pohl6 air-lift pump Pohl6 air-lift pump Pohl« air-lift pump 20 20 II 34 10 0.38 0.3 0.22 0.6 0.3 10,500 X 100 5000X75 2000 — 2200X250 4500X180 4000 X 60 Cu 98-99 variable Ag AuCAs + Sb)>.2 Ni.etc. Cu 99 . 1 Ag 0.13 Cu 99.25 Ago.3 36X36Xt = b 36X24X1! ti">b 28.SX24X1.8 28X30X1. 5 t = b 35X24 3"t 2i"b 32JX241 l}"t l"b 475 350 400 500 230 Shoulder Cast lugs Lugs with hooks Morrow clip Lug 20 24± 60 18 37 IS 15-20 8 8 7 4o!X38}Xo.ll4 37X2S(XA?) 29iX2S 29X31 361X26 33X11 8-10 7 6 and 7 2i 5-1 Loop Loops Straps Morrow clip Clamped loop 10 10 30-1- 2 37 140 170 55 Overhead elec- tric crane Overhead trolleys 2 hand hoists Cranes Overhead crane Overhead elec- tric crane Il'X3'6J"X 3' 7!" 9'X28"X44" 9'li"X33l"X32" 9' IJ"X33J"X38" 9'9"X2'4"X3'9" 8'2''X4'S''X4'6' 30, 31 21, 22 20, 21 21, 22 38, 80 Walker Cross bars Double rows Thofern system 20 sq. in. Mains 2}"Xl}" On tanks 2"Xl" 4S and 3i"XiJ" 700-900 1-25 I 0.20 Ag 34 Cu IS Ag + Au 25-40 Cu is-20 Rest As, Sb, Pb, etc. Cu-35 Sn-l As-3 Ag-5S Small amounts of Pb-6 Se.Te, Sb. Ni, Co Cu43 Ag 17 Cu 10 Ag62 10 30 60 60 34 53° METALLURGY OF COPPER lift pumps and distributing tanks. Next to the pump room are one set of starting-sheet tanks of the same dimensions as the depositing tanks, each with 22 anodes and 21 cathode-blanks of i-in. sheet copper 27.5 by 39.5 in.; four purifying tanks; and two reducing tanks. The crane room contains a 7-ton crane for handling cathodes and scrap. Anodes and cathodes are hand- led four at a time with a i-ton Yale & Towne triplet chain block and multi- ple hook; they arrive and are deposited in buggies, and are weighed in the crane and scale room. The electrolyte (analysis, Table 129) contains per liter 40 g. Cu and 160 g. H2SO4, and is circulated at the rate of 6 gal. per minute. The current density is Origin of Materia] Treated. Klectrolytic Slime. 21%H20 V Steam Drying Tables. I Electrolytic Slime. Silver Refinery. Fig. 535- — Flow-sheet of Great Falls electrolytic refinery. 34 amp. per square foot cathode area. Cathodes are removed every 2 days. An anode of 500 lb. weighs 30 lb., when returned to the converter. The anode mud is taken out once in 60 days when about 5.5 in. will have collected. A cop- per jumper. Fig. 521, serves to disconnect the tanks that are to be cleaned; a bronze steam-injector to transfer the mud over |- and |-in. screens into a lead-Uned tank, 12 ft. in diameter and 4 ft. deep, where it is washed with water, drawn off into a cast-copper lead-lined montejus, and forced through a Bushnell filter-press, containing hard-copper plates and 8-oz. duck filters. This furnishes 13 cakes, 26 in. in diameter and 1.25 in. thick, which retain 20 per cent. H2O. The cakes are dried on copper steam-drying tables, Fig. 538, ELECTROLYSIS OF COPPER 531 Scale Room SECTION OF TANK ROOM ON LINEA-A 532 METALLURGY OF COPPER until half of the water has been expelled; crushed with a roller on a cement floor; and sampled in 22,000-lb. lots for shipment. The sulphate building, Figs. 538-540, with the additional starting-sheet tanks and the machinery for preparing the sheets, contains three boiling tanks for con- centrating electrolyte, the necessary crystallizing vats for blue vitriol, and precipitating tanks for treating mother liquor freed from part of its copper in the form of blue vitriol. The current is furnished by two Westinghouse, shunt-wound, engine-type, 8io-kw. generators having a normal capacity of 4500 amp. at 180 volt with a speed of 130 r.p.m. The average output of the machines is 4600 amp. each at 222 volt or 1021 kw., equivalent to 26 per cent, overload. The genera- tors are driven by water-power. The current is conducted to the tank build- ing through slabs of cast copper joined by overlapping ends fastened with iron bolts; expansion joints in the form of cast-copper arches take care of changes of temperature. (c) The Raritan Plant, No. 2 (Walker System). — The plan of the tank house is given in Fig. 542. The building, 582 ft. long by 149 ft. 6 in. wide, has in the tank room on the main floor three parallel rows of tanks, each of which is served by two lo-ton three-motor overhead cranes, 19 ft. 8 in. above the floor, for handling anodes and cathodes. Beneath the main floor, supported by con- crete pillars (Fig. 512), is a light cellar 9 ft. 9 in. high with acid-proof floor consisting of 6 in. concrete covered with pitch and overlain by a 2-in. course of chemical brick with pitch joints. The floor drains through gutters to a sump. The building is warmed with exhaust steam and has artificial ventilation to keep everything dry and thus prevent leakage of current. The air is changed by a rotary fan once in 20 min.; 75 per cent, of the air goes through the cellar, 25 per cent, through the room. There are 1188 tanks in the room grouped in three rows of 396 tanks; a row has 36 nests, and a nest 11 tanks, 10 ft. long by 2 ft. 8 in. wide by 3 ft. 11 in. deep. There is one liberating tank for every 44 elec- trolyzing tanks. The engine room has four generators, one of which is held in reserve; a generator requiring 1250 h.p. furnishes 396 tanks with a current of 7500 amp. at 145 volt. The heavily-dotted lines show the passage of the current from the generators to each of the three rows of tanks, through which it travels lengthwise. The current density is 20 amp. per square foot cathode area. The main conductors have a cross-sectional area of 12.75 sq. in. and receive a current of 500 amp. per square inch. The fall in potential from anode to cathode is 0.210 volt, and from tank to tank 0.360 volt. At the ends of the bmlding, opposite the stripping benches, are the vats for preparing the starting sheets; nearby are shears and Moore looping machines. In the cross-aisles are wash- ing boxes to clean corroded anodes before they are returned to the anode furnace, and to dip cathodes twice in hot water to wash off adhering electrolyte; in the same aisles are stands or racks for a complement of anodes and cathodes to be taken to or from the vats. In the pump room are six Antisell centrifugal pumps (Figs. 495-496). The return electrolyte flows into six sheet-iron lead- lined pump boxes of 4 cu. ft. capacity which overflow either into three emer- ELECTROLYSIS OF COPPER S33 534 METALLURGY OF COPPER gency tanks, i6 ft. in diameter and 12 ft. deep of sheet iron and lead-lined, or normally into the inlet pipes of the pumps. The electrolyte is raised 27 ft. and delivered to one of the 6 heating tanks, 28 by 4 ft. and 4 ft. deep, pro- vided with 12 lead steam pipes, 1.5 in. diameter, to raise the temperature to about 50° C. For the circulation of the electrolyte, containing 3 per cent. Cu and 12 per cent, free H2SO4, the tank room is divided cross- wise into six units, two at the ends, and four in the center, each with 198 tanks. A unit with its six cross-rows of tanks is served from a centrifugal pump; two neighboring cascades are fed from one branch pipe. The rate of circulation is 4 gal. per minute, and the system is shown in Fig. 491. The construction of the tanks has been given in Figs. 511-515; a tank contains 24 anodes, 36 by 28 by 2 in., and 25 cathodes, 37 by 30 in.', giving it an active cathode surface of 369.5 sq. ft. An anode weighs 475 lb., and has shoulders i in. thick narrowed toward the ends so that the center of gravity lies near them. A starting cathode is ready after 42 hr. deposition. An anode remains in the vat for 28 days and furnishes from 12 to 13 per cent, scrap of which 5 per cent, is in the shoulders; a cathode remains 14 days and weighs 160 lb. A tank is cleaned up, when the anodes are exchanged. The working- up of the mud has been discussed in § 262. There are locked up in the plant, in rods, plates, busbars, leads, etc., 581,000 lb. Cu, in the electrodes under treatment 10,000,000-11,000,000 lb., and in the electrolyte 240,000 lb. The lead used in construction, pipes, anodes, etc., totals 1,370,000 lb. (d) Tabulated Data. — Table 128 contains the leading facts of the principal electrolytic plants of the United States using the multiple process. The refineries of Lithgow^ and Port Kembla^ have been described in the accompanying references. ' B. Series System 267. Series (Hayden) System in General. — In this process cast or rolled electrodes of high-grade copper are placed vertically in series in an electro- lyzing vat charged with -acidulated blue vitriol so as to fit closely the sides. Fig. 543 gives a diagrammatic sketch. As the current passes through the vat, the electrodes, with exception of those at the ends, become negatively charged on the sides facing the entrance and positively on the sides facing the Fig. S43-— The series (Hayden) system, exit of the Current. The positive current entering through one end electrode, which is solely anode, causes Cu to be dissolved and to be deposited on the negative side of the next following intermediary electrode, while on the positive side of the latter Cu goes into solution, and so on through the vat to the last ^ Eng. Min. J., 1910, xc, 717. 'JIfeJ. Cfew. £Mg., 1912, X, 694. ELECTROLYSIS OF COPPER 535 electrode which, solely cathode, is connected with the exit wire. In this manner the copper is dissolved and deposited until the original intermediary electrodes, anodes on one side and cathodes on the other, have been changed into electrodeposited copper, and the end electrodes have become lighter or heavier. The insoluble impurities and precious metals collect on the bottom of a vat as anode mud. The pure copper and the anode mud are worked up as in the multiple process. The Hayden process' has outlived the other two series processes of Smith and of Randolph. In the former, the electrodes were placed horizontally and sr — II I ' 1 OoBoroce I Figs. 544-548. — Tank-details of the Baltimore electrolytic refinery. separated by diaphragrns, and the current entered at the top; in the latter it en- tered at the bottom. Stalmann's idea^ of riveting sheets of copper to the nega- tive sides of the vertical intermediary electrodes of Hayden, and thus facilitating the removal of the last of any undissolved electrode- from the newly deposited copper, has been found to be unnecessary. In the Hayden process, then, the vertical electrodes of a vat are connected in series, and the vats in multiple. Series processes are in operation in the United States at the works of the Nichols Copper Co., Brooklyn, N. Y., and the Baltimore Copper Smelting & Rolling Co., Baltimore, Md. ' Badt, Eng. Min. J., 1892, liv, 126. ^U. S. Patent, No. 467350, 467484, January iS, 1892. 536 METALLURGY OF COPPER 268. The Baltimore Plant. — As little information is available regarding the Nichols reiinery/ the present discussion is confined to the Baltimore plant, of which tank details are given in Figs. 544-548. 269. Electrolyte. — The composition, temperature, and circulation of the electrolyte are about the same as in the multiple process. It contains 1 2 per cent, blue vitriol and 9 per cent, free H2SO4; the circulation is 2.5 gal. per minute; the temperature 40-43° C. The electrolyte is siphoned from the bottom of the vat at the discharge end instead of being made to overflow. The siphon is shown in Fig. 546; it has an orifice at the top which serves as an overflow and starts the siphon, whereupon the regular flow begins near the bottom of the tank. 270. Current. — The current connection is made through conductor bars at the ends of a tank. Figs. 544-546. The current density is 21 amp. per square foot and the electromotive force 22 volt with a vat holding 135 elec- trodes; the fall in potential from plate to plate is about | volt. On account of the high voltage there is a strong leakage of current, which reduces the ampere eflicienfcy to from 65 to 67 per cent. 271. Electrodes. — The electrodes are rolled sheet copper. In order to permit rolling, the copper must be of good quality and may not contain too much CU2O; hence blister copper from the reverberatory furnace or the converter has to undergo a partial fire-refining before it may be cast into suitable cakes. A small addition of Pb, not over o.i per cent., to the refining charge improves the rolling quality of the copper. Table 129.^ — Copper Suited and Unsuited for Rolling Electrodes'' Copper Pb Bi Sb As Se and Te Ag, oz. per ton Au, oz. per ton Suited tor rolling Suited for rolling Unsuited for rolling Unsuited for rolling Suited for rolling after adding Pb , Suited for rolling after adding Pb, 0.0082 0.0093 0.0558 0.0073 o . oo6g trace 0.0025 o . 03 20 0.0274 0.0340 0.0095 0.0055 0.0443 0.0651 0.1245 0-^350 0.0602 0.0370 0.0068 0.0586 0.1160 0.0582 o . 03 1 2 0.0255 0.0071 o . 0098 0.0153 0.1067 0.0527 0.0365 100.60 229.40 156-30 1 7 2 . go 0.45. 0.12 0.48 0.60 Toughened copper from a 250-ton reverberatory furnace is run into ladles, of 6000 lb. capacity, placed on trucks; they discharge their contents into a retaining tilting-ladle of 20,000 lb. capacity, from which billets weighing 490 lb. are cast at the rate of 60 tons per hour by means of a rotating-table casting machine with 40 molds lOf by 42 by 35 in. The red-hot billets are dropped onto a conveyor which delivers them to the first of a series of 5 two-high con- tinuous rolls, 21 in. in diameter, which roll the billets into sheets yV in. thick at the rate of 10 tons per hour.' From the fifth roll, the sheet is transferred to • Cast electrodes 10 by 59 by | in., weighing 67 lb.; wooden vats lined with pitch; electro- lyte kept cool to prevent flowing of pitch. ^ Keller, Min. Ind., 1898, viil, 233. 'Editor: "Rolling Copper," Iron Age, 1907, lxxx, 507. ELECTROLYSIS OF COPPER 537 a sixth roll, in line with the fifth but moving in the opposite direction, which de- livers the sheet into a water-trough provided with rollers. The cooled sheet goes to crocodile shears which cut it into electrodes ii by 24I by A in., forming 6 per cent, scrap from the ends. The electrodes are straightened under a drop- hammer and go to the frame division. Figs. 547-548, where two are placed by hand between a pair of grooved wooden strips; the joints on the positive sides are then painted with tar to facilitate the removal of the deposited copper from any remaining electrode material. The difference between the commercial cathode and the cathode freed from all electrode scrap is shown in Table 130.' Series-copper usually contains 1.5-2.0 oz. Ag per ton; multiple-copper 0.3 oz. Table 130. — Cathode from; the Hayden System Cathode Pb Bi Sb As Ag. oz. per ton Commercial Freed from all scrap 0.00047 0.000s 5 0.00018 0.00136 0.00094 u. 00059 0.00025 1 .11 0.36 272. Depositing Vat. — On account of the high voltage in the series process and the consequent danger of shortcircuiting, the usual construction of the vats of wood lined with lead is impracticable. Formerly the vats were of slate with joints made tight by a tar cement; the sides were coated with tar, and the tops covered with slats. At present. Figs. 544-546, the tanks, 11 ft. 6 in. long by 25 in. wide by 26 in. deep, are composed of a mixture of asphalt, asbestos, and sand molded in place; 66 vats form a block, which rests on square glazed drain- tile pipes each carried by two courses of brick and concrete walls; between brick and wall is placed sheet lead to deflect any leakage of electrolyte. Between the several tanks are spaces 3 in. wide through which pass tie-rods, enclosed in lead pipe, connecting wooden buck-stays and channel-iron washers, which take up the end- thrust of the tanks through plates let into the tanks. The side- thrust of a block is taken up by a 4-in. wall built of 2-in. strips, the wall being held in place by posts buried in the ground. The spaces between the single tanks are filled with broken stone, and the interstices closed by pouring in molten sulphur. The 66 tanks of a block are connected in parallel and receive a cur- rent of 500 amp. at 220 volt. 273. Corrosion and Deposition. — Both proceed uniformly, and an elec- trode is corroded in 17 days. On the edge of the deposited copper there is usually found a small strip of electrode material, which is pulled off with nippers. At the center there remains sometimes a skeleton-like undecom- posed patch of electrode which has to be removed. 274. Clean Up.— When the electrodes in a vat have been decomposed they are disconnected; the electrolyte is siphoned off; the cathodes are washed with a hose, being turned over like leaves in a book, removed, and the strips taken off; the copper is then transferred to the reverberatory furnace. The mud is sluiced out and worked up, as well as the fouled electrolyte, as in the multiple process. 1 Keller, Min. Ind., 1898, vii, 241. 538 METALLURGY OF COPPER C. Multiple Versus Series System 275. Multiple and Series Systems Compared.' — The advantages claimed for the multiple system are: (i) Treatment of anode copper rich in precious metals and high in impurities; (2) handling of material in large units at low cost; (3) permissible variation in composition of electrolyte. The advantages claimed for the series system are: (i) Smallness of power required per unit of deposited copper; (2) small amount of copper and precious metal locked up; (3) low percentage of scrap produced" (4) little space re- quired per unit of copptr deposited. (i) In the multiple system anodes with over i per cent, impurity and 400- 1000 oz. AgAu per ton are frequently treated. In the series process the rolling of electrodes requires pure electrode material; the small electrode-distance, favoring the settlement of slime on the deposited copper, and the adhesion of anode mate- rial would cause the market copper to be rich in precious metal. In the series system therefore the electrodes should not contain over 0.166 per cent, impurity and not over 100 oz. AgAu per ton; 70 oz. is preferred. (2) In the multiple system the electrodes and anode scrap are handled mechanically in large units, but in the series system the manipulation of elec- trodes and cathode copper is entirely by hand. (3) In the multiple system there are open spaces between the vat and the electrodes, and the electrode distance is large, both of which permit a rapid circulation without danger of stirring up anode mud. The more rapid the circulation, the more impure can be the electrolyte without endangering the cathode copper, and the less frequent has to be its renewal. (i) In the series system, as the electrode distance is small, the fall in poten- tial is low; it is about one-half that of the multiple system, or 0.15 vs. 0.30 volt. This means that in the series system half the power will be required to deposit a given amount of copper as in the multiple. This advantage is offset by the great cost of casting thin electrodes or of rolling cakes into sheets when compared with the casting and handling in the multiple system. (2) In the series system part of the copper in the busbars and all in the cross- bars is saved. The amount of copper and precious metal locked up in tanks is equal to the daily product multiplied by the time interval of the clean-up periods; i.e., in the series system the factor is 15, in the multiple it is 26 for copper and 33 for precious metal, allowing for scrap in both cases. With' average copper bul- lion the interest on the metal locked up in the series system is one-half that in the multiple. (3) The scrap produced in the series system amounts to 3 to 6 per cent, of the weight of the electrode; in the multiple the usual figure is 10 per cent, and often 13 per cent. 1 Peters, "Modern Copper Smelting," 1895, p. 577. Keller, Min. Ind., i8g8, vu, 229. Haber, Zt. Eledrochem., 1903, ix, 384. Editor, Eledrochem. Met. Ind., 1908, vi, 223. Walker, Min. Ind., 1908, xvn, 327. ELECTROLYSIS OF COPPER 539 (4) The series system formerly required much less floor space than the mul- tiple. This does not hold good any longer, since refiners using the multiple system have increased the depth of the immersed anode; thus the U. S. Metals Refining Co. has anodes 3 by 3 ft., and requires 330 sq. ft. of tank room per ton of copper produced per day, a figure which is lower than in the series system. Summary. — The cost of operating by the two systems is about the same; the multiple has the advantage of being able to treat almost all classes of copper bullion, and of requiring less care in conducting tank room operations Of the nine electrolytic refineries in the United States (Table 115), two have the series system, the rest the multiple. INDEX Accessories, converter, 314 Acid-feeder, vitriolization metallic copper, 479 Add vs. basic converter, 334, Copper Cliff, 344 Addicks, charging machine, 381 -Marks system, melting cathodes, 401 Addition-agent, effect upon deposition of copper, 488 Additions, fire-refining copper, 387 Agglomeration, flue-dust, 233 slime, with native copper ore, 370 Agordo, Italy, sulphatizing heap-roast, 426 Aguas Calientes converter, 301 Aich metal, 37 Air-cooling bottom, copper-refining furnace, . 373 . -drill, lining acid converter, 311 partial pyritic smelting, igg -pressure, atomizing oil, 269 pyritic smelting, 193 -rabbling, copper- refining furnace, 382 vs. steam, atomizing oil, 269 Ajax metal, 45 Allen-O'Hara roasting furnace, 124 pyritic process, 198 AUotropic copper, 5 Alloys, copper, 24 Altitude and blast-roasting, 146 Alumina, partial pyritic slags, 199 Aluminum brass, 36 bronze, 48 American process, blast-furnace smelting, 64 Ammonia, as solvent, oxide ore, 404 Ampere-efficiency, multiple system, 498 series system, Baltimore, 536 starting-tank, 503 Anaconda anode, 500-502 basic converter, 345 blast-furnace, 152 briquetting flue-dust, 234 cathode, 504 converter, acid lining material, 310 lining, 313 converting, 191 1, 306 dust flue, 220 early converter, 301 Evans-Klepetko furnace, 107 feeding blast-furnace, 165 flow-sheet, 354 Laist process, leaching tailings, 43 1 length of reverberatory with direct firing, 267 multiple plant, 525 old converting plant, 301, 360 Anaconda reverberatory furnace, 1908, 243 claying, fettling, 284 heat balance, 280 starting, 265 Anaconda settler, 171 temperature record, 282, 283 Annealing brass, 31 Anode, corrosion, 511 insoluble, 512 use of oil in tanks, 522 mud, S13 treatment, 515 multiple process, 499 Antifriction metal, 45 Antimonial speise, 217 Antimony and copper, 19 behavior, fire-refining copper, 393 electrolysis, 488 elimination in blast-furnace, etc., see Elimination Antisell centrifugal pump, circulation elec- trolyte, 497 Apparatus, roasting, 70 Area, roast-yard, 74 Argentine, Kans., vitriolization plant, 471 Argo, cleaning reverb, slag, 286 process, 295 Arizona, copper blast-furnace, 359 experiments, leaching oxide ore, 414 pig-copper, 361 practice, smelting oxide ore, 360 Arrangement converting plant, 320 of vats, multiple process, 506 oil burners, Steptoe Valley, 273 reverb, furn. plant. Copper Queen, 286 Arresting devices, flue-du^t, 224 Arsenic and copper, 18 behavior electrolysis, 487 fire-refining copper, 393, 394 condensation, 147 elimination in blast-furnace, etc., see Elimination Atacamite, leaching, 420 Atomizing burners, oil, reverb, furn., 269 Augur machine. Chambers Bros., briquetting flue-dust, 234 Augustin process, matte, 456 metallic copper, 472 Automatic roasting furnace, 87, 138 Babcock & Wilcox vs. Stirling waste-heat boiler, 243 BaflJes, flue-dust, 221 Baghouse, Mammoth smeltery, 228 Bag-shaker, Benedict, 229 Balaklala, condensation flue-dust, 229 converter, 306 Balan, Transylvania, sulphatizing roast, 426 Balbach S. & R. Co., copper-refining furn, 378 pioneer electrolysis metaUic copper, 484 Ball converter, Shelby, 306 Baltimore series plant, 536 541 542 INDEX Barrel converter, 305 work, native copper ore, 361 Base and silica required for slag formation, i8s Basic bottom, reverb, furn., 260 converter, 334 Anaconda, 345 Cananea, 347 copper loss, 343 Great Falls, 349 magnetite lining, ^^^ Peirce-Smith, examples, 343 Steptoe Valley, 346 vs. acid converter, 334 Copper Cliff, 344 Bath of matte, reverb, furn., advantages, 282 Bearing bronze, 45 Bedding and reclaiming MacDougall charges, Cananea, 112 Bell metal, 46 Benedict bag-shaker, 229 Bennett matte-mold, 175 Bennets pouring-spoon, 317 Best selecting process, 293 Bibliography, copper, 3 Bingham Canyon, blast-roasting, 140 Bismuth and copper, 16 behavior in electrolysis, 488 elimination in blast-furnace, etc., see Elimination Bituminous shale, heap-roast, 79 Black copper, Arizona, 361 Blagodatny, blast-roasting, 142 Blast, blast-furnace, 182 converter, 314 partial pyritic smelting, 199 preheating, 182 pyritic smelting, 193 trapped, 157 Blast-furnace, Anaconda, rS2 and accessory apparatus, 148 Arizona, oxide ore, 359 blast, 182 bosh, 161 building, 152 Cananea, 152 charge calculation, 185 cost of smelting, 233 effect of briquettes on flue-dust, 234 fore-hearth, 169 flue-dust, 218 amounts, 234 fuel, 182 gas, temperature, velocity, 218 Great Falls, 152 hearth, 157 accretion, 235 heat-efficiency, 189 Herreshoff, 149 production of metallic copper from matte, 236 Mount Lyell, 152 native copper ore and by-products, 370 reducing smelting, 178 settler, 169 shaft, 61 slag, 217 Blast-furnace, slag, reducing smelting, 179 smeltery, 148 smelting power, 236 table, 176 oxide ore, 360 vs. reverb, furn., matting, 287 water-jackets, 161 withdrawal of gases, 165 yield in metal, 235 Blast-roasting, 138 Casapalca, 142 down-draft, 143 first experiments, 140 Garfield, 139 matte, 142 Morenci, 140 Spain, 141 speise, 217 up-draft, 139 Wallaroo, 141 Blister copper, 291 forming stage, converting, 324 furnaces, 291 Blowing in, partial pyritic furnace, 210 Blue Billy, 4S3 metal, 295 slag, 29s vitriol, 55 crystallization plant, Hofmann, 470 requirements for crystallization, 474 salting out of small crystals, 471 Boiling point, copper, 7 Bone-ash wash, copper mold, 397 Borchers-Franke-Giinther process, electrol- ysis of matte, 483 Boron oxide, addition, fire-refining copper, 387 Bosh, blast-furnace, 161 pyritic, 190 Bottom-blown converter, 300 Bottom, copper-refining furnace, 373 Bottoms, made in reverb, furn., 293 Braden copper mines, leaching ore, 429 Bradley process, leaching copper ore, 429 Brass, 28 aluminum, 36 iron, 36 manganese, 38 nickel, 39 regular, 34 special, 36 test, fire-refining copper, 386 tin, 38 Bringing forward matte to copper in blast- furnace, 236 in reverb, furn., 288 Briquettes on blast-furnace charge, 235 flue-dust, 234 slime, native copper ore, 370 British Columbia Copper Co., smeltery, 355 Brochantite, leaching, 420 Broken connections, circulation electrolyte, 527 Bronze, 39 machinery, 45 manganese, 38 special, 46 INDEX 543 Bronze, Tobin, 37 Brower converter hood, 314 Brown reverb, smelting furn., 255 Horseshoe furnace, 129 straight-line reverb, roasting furnace, 124 BrUckner roasting furnace, 135 Building of the blast-furnace, 152 roast-heap, 75 Bull-ladle, wire-bar copper, 398 Bullwhacker works, leaching oxide ore, 414 Burner, oil, reverb, furn., 269 Shelby, Sorensen, Steptoe, 272 Busbar, cross-section, 511 Butte & Boston, converter lining, 312 Butte-Duluth Mg. Co., leaching oxide ore, 414 Butte dumps, tailings, leaching, 413 mine-waters, 408 Reduction Works, settler, 172 Calcines, hot vs. cold, reverb, furn., 280 Calcium hydroxide, precipitant for copper, 40s '' . . sulphide, precipitant for copper, 405 Calculation, air, pyritic smelting, 193 blast-furnace charge, 185 partial pyritic charge, 202 reverb, furn. charge, 279 thermal balances, 204, 327 California copper ore, 61 Calumet & Hecla, reverb, furn., 366 Canadian Copper Co., reverb, furn., dust- fired, 24s smeltery, 355 Cananea, basic converter, 347 blast-furnace, 152 converter, acid lining material, 31 r dust flue, 219 feeding blast-furnace, 166 MacDougall furnace, in, 112 oil-burner, 272 reverb, furn., claying, fettling, 284 settler, 172 smelting flow-sheet, 355 reverb, furn., 247 treatment flue-dust, 233 Capital, establishment and working, sul- phide plant, 358 locked up in multiple plant, 524 Carmichael process, leaching ore, 415 Casapalca, blast-roasting, 142 Casarza works, electrolysis of matte, 482 Cast brass, 36 Casting anodes, multiple process, 502 by hand, refined copper, 396 by machine, refined copper, 398 machine, converter copper. Great Falls, 317 refined copper, 395 temperature, refined copper, 398 Cast-iron top-hearth, MacDougall furn., 108 Cathode, deposition, multiple process, 511 multiple process, 502 Cathodes, changing of, 498 Cavity, acid converter, 311 Cement copper, Longmaid- Henderson process, . 453 silver, pressing, drying, melting, 462 Cerro de Pasco, blast-roasting, 146 Chalcopyrite, decomposition by heating in closed vessels, 433 Chambers augur machine, briquetting flue- dust, 234 Changes, fire-refining black copper, 391 cathode copper, 394 Charcoal in blast-furnace, 182 Charge, feeding, blast-furnace, 165 Charging copper-refining furnace, 381 Charles method, warming converter fining, „, . 313 Chemical analysis, Anaconda, calcine, loS anode and anode mud, 486 mud, 514 raw and treated, 516 multiple process, 514 basic converter slag. Anaconda, 346 MgO vs. Si02, 341 bituminous coal. Great Falls, 254 black copper, 391, 392, 472 Arizona, 361 blanket slag, reverb, furn., 278 blast-furnace slags, oxide ore, 361 roasting gases, 142 ore, 142 sinter, 141, 143 blister copper, 292 converter, 331, 332 Blue Billy, 454 metal, 295 vitriol, pure, 464, 477, 478 impure, 521 bottoms, 296 burnt pyrite, 442 Butte concentrate, calcine, flue-dust, 123, 124, 176 calcine. Anaconda, 108 Cananea converter, acid lining mate- rial, 311 cathode, Hayden system, 537 cement copper, Longmaid-Henderson process, 453 chrome brick, 157, 172 concentrate and calcine, MacDougall furnace, Steptoe Valley, 109 calcine, Edwards reverb, furn. Gold- field Cons., 120 flue-dust, Cananea, in, 112 Garfield, 109 Morenci, 140 Copper Queen, no Hayden, no, in Tooele, 109, no Utah Copper Co., 339 converter anode, Anaconda, 527 copper, 331, 332 gas, 333 linings, 310 matte Anaconda, 346 mattes, 327 slag, 332 acid, 324 basic, 341, 342 Anaconda, 346 Great Falls, 353 544 INDEX Chemial analysis, copper granules, for vitriol- ization, 473, 476 matte, 457 precipitate, Butte mine- water, 409 refined, 12, 13, 392, 393 refinery slags, 383 dust, basic converter, Garfield, 343 electrolyte, 492 floater, reverb, furn., 282 flue-dust. Anaconda, 220, 221 basic converter, 343 converter, 332, 333 Garfield converter, 343 Great Falls, 227 MacDougall furn., Anaconda, 104 Garfield, 109 Mammoth smeltery, 229 Mason Valley, 233 scrubber sludge, anode mud, treat- ment of, 518 furnace-bottom,' loosened, "floater," 282 gas, from stack, Great Falls, 225 head- and tail-liquor, Eisbee, 413 Butte, 412 Rio Tinto, 423 Schmoellnitz, 408 leached residue, Ziervogel process, 460, 461 lining material, acid converter. Ana- conda, Great Falls, 310 matte, 213 average, Anaconda, 346 Freiberg vitriolization process, 463 metal slag, 290 Miami concentrate, 347 mine-water, Butte, 412 freed from copper, 413 and tail-water. Copper Queen, 413 Montana copper ore, 61 Nevada copper ore, 61 ochre, from mine-water, 407 ore, Rio Tinto, 421 slag, reverb, furn., Argo, 297 oxide copper ore, 360 Predazzo,Italy, sulphide copper ore,427 producer gas. Great Falls, 254 wood, 268 purple ore, 454 pyrite, kiln-roasted, 437 refined copper, 12, 392 regule, 296 residue, vitriolization metallic copper, 477 reverb, furn., blanket slag, 279 slag. Step toe Valley, 270 Rio Tinto, raw,roasted, leached ore,425 roaster gas, MacDougall furn., 147 slag, 296 Schmoellnitz, mine-water, 408 screenings, fine concentrate, for Evans- Klepetko furn., Great Falls, 104 Shannon Copper Co., low-grade ore,429 oxide ore, 430 silica sand, working botom, 261 siliceous ore, basic converter, 340, 344 Chemical analysis, siliceous ore, Anaconda, 346 Great Falls, 352 silver bar, Mansfeld, 462 slag, Knudson process, 196 pyritic. Mount Lyell, 193 reverb., 275, 277 smelting oxide ore, 361 - slime, electrodeposition with insoluble anodes, 523, 524 speise, 216 Stassfurt salt, 442 Sulitjelma ore, 196 t o w e r-Uquor, Longmaid-Henderson process, 4Si, 452 Utah copper ore, 61 white metal, 290 changes, fire-refining copper, 391 properties brass, 32 bronze, 43 copper, 9 Chemistry, converting, 323 partial pyritic smelting," 201 pyritic smelting, 194 reducing smelting, iDlast-fumace, 183 reverb, furn. smelting, 278 Chloridation of burnt pyrite, 441 progress, 445, 448 Chloridized ore, Longmaid-Henderson, leach- . ing, 450 Chloridizing furnaces, 442 mechanical muffle. Wedge, 446 reverb.. Wedge, 445 muffle, Natrona, 445 multiple-hearth mechanical down-draft muflSe, Wedge, 446 reverb, and muffle. Wedge, 448 Ram6n-Beskow, 449 reverb., Oker, 442 roast, condensation of gases, 449 Longmaid-Henderson process, 441 Chrome brick, copper refining furn., 374 reverb, furn. bottom, 261 Canadian Copper Co., 246 Chromium and copper, 50 Circulation electrolyte, 493 Anaconda plant, 527 Clarifying copper liquor, Longmaid-Hender- son process, 452 Clark-Antisell, charging machine, 381 Clark casting machine, refined copper, 400 Claudet, process, 454 Clay brick, reverb, furn., 260 Claying, Cananea, 284 Clean-up, multiple system, 513 series system, Baltimore, 537 Coal, reverb, furn., 265 Coarse metal, 288 Cobalt, behavior electrolysis, 485 fire-refining copper, 392 and copper, 18 Coke in blast-furnace, 182 -iron couple, precipitation of copper, 405 use in refining copper, 398 Collins, blast-roasting pot, 140 Color, copper, 4 Colorado Smelting Co., reverb, furn., 238 sulphatizing roast, 427 INDEX 545 Coming to nature, refined copper, 395 Common brass, 34 Comparison, heaps, stalls, kilns, 86 roasting methods, 147 Composition, electrolyte, 492 Compressed air, copper-refining furn., 382 Concrete foundation, reverb, furn., 243 Condensation, Anaconda, 220 Balaklala, 229 Cananea, 219 Copper Queen, 219 devices, flue-dust, 224 of gases, chloridizing roasting, 449 Great Falls, 221 Mammoth Works, 228 Conductivity bridge for copper, 386 , electrolyte, 489 Connections, broken, electrolysis copper, 49s, 498 'Consolidated Arizona S. Co., Edwards furn., 120 Converter, accessories, 314 acid vs. basic, 344 lining, 300, 310 table, 301 Aguas Calientes, 301 Anaconda, 1911, 306 anode, cast vs. refined, 499 Balaklala, 306 barrel, 305 basic, 334 products, loss, cost, 342 blast, 314 cavity, 31: charge, addition of Si02, 326 chemistry, 323 copper, 331, 332 casting machine, 317 ingot, 317 David-Manh6s, 305 Eguilles, 301 gases, 326, 333 Great Falls, 1909, 301 Haas, 308 hood, 314 horizontal, 305 Knudsen, 197 Leghorn, 305 lining acid, 300, 310 losses, 331 Manhfis, 300 manipulation, 309 operation, 323 Parrot, 301 Peirce-Smith, 33s plant. Great Cobar, 32X products, 331 Shelby, ball, 306 slag, 332 basic, 342 Great Falls, 353 Stallman, 301 table, acid vessels, 302, 303 thermal balance, 327 trough, 30S upright, 300 vs. horizontal, 308 35 Converting, basic, 334 temperature, 341 chemistry, 323 copper loss, 333 cost, 331, 333 elimination, impurities, 326 losses, 331 matte, 298 plant, 320 vs. roast-smelting, elimination of im- purities, 326 silver loss, 333 speise, 217 Coin bronze, 46 Copper and aluminum, 48 and aluminum manganese, 5° and antimony, 19 and arsenic, 18 and bismuth, 16 and chromium, 50 and cobalt, 18 and cuprous oxide, 13 and gold, 23 and gold-silver, 23 and iron, 16 and lead, 15 and lead-silver, 22 and manganese, 17 and minor metals (Ca, Mg, Cd, Tl, Pd, Pt, Va, W, Ti, Cr, Cd-Sb), 23 and nickel, 17 and oxygen, 13 and phosphorus, 25 and quicksilver, 50 and selenium, 21 and silicon, 26 and silver, 22 and sulphur, 21 and tellurium, 21 and tin, 39 and titanium, 50 and tungsten, 50 and vanadium, 50 and zinc, 28 alloys, 24 best selected, 293 bibliography, 3 casting machine. Great Falls, 317 Walker and others, 398 chemical properties, 9 Cliff, basic converter, Peirce-Smith, 344 content of slag, length of reverberatory furnace, 267 cost of production, 2 cuprous oxide, 13 deposition per amp.-hr., 491 dissolved gas and spec, grav., 384 effects of impurities, 12 grades, 2 _ <- o granules, for vitriohzation, 473, 47°, 47a history, i ingot, 396 bar, 396 . in matte, necessary for precious metal, 216 liquor, refining, Hofmann process, 40& loss, basic converter, 343 546 INDEX Copper loss in blast-furnace, i8g, 235 in converting, ziS in slag, 217 matte, converting, 298 electrolysis, 482 metallic, leaching, 471 of commerce, 10 Copperopolis, Cal., sulphatizing heap-roast, 42s Copper ore, electrolysis, 482 leaching, 402 marketing, 62 metallurgical treatment, 62 smelting, 63 ores, 60 physical properties, 4 pig-mold, 359 poisonous effect, 232 precipitate, 406 Butte mine- water, 409 production. United States, world, 2 Queen, arrangement of reverb, furn. plant, 286 barrel converter, 305 dust flue, 2ig MacDpugall furn., no mine-water precipitating plant, 412 system, ore bedding, 357 treatment of flue-dust, 232 rain,_383 refining, chemical changes, 391 furnace, gas-fired, 377 operation, 381 temperature, 395 furnaces, 374, 388, 390 selenide, behavior electrolysis, 486 silicates, 52 -silver selenide, 517 smelting for matte, in reverb, furn., 288 specifications, 12 statistics, 2 sulphide-iron sulphide, 213 sulphide, roasting, 65, 67 sulphides, 53 teUuride, behavior electrolysis, 486 tough-pitch, 38s uses of, II Corrosion and deposition, series system, Balti- more, S3 anode, 511 brass, 32 bronze, 43 converter lining, 311 water-jackets, 164 Cost, acid converting, 333 basic converting, 342, 343 blast-furnace smelting, 235 heap-roast, 78 Knudsen process, 198 leaching Anaconda tailing, 43 1 Longmaid-Henderson process, 455 multiple plant, refining copper, 524 power vs. current density, multiple process, 525 precipitating mine-water, Butte, 412 producing blue vitriol, 480 production of copper, 2 Cost, refining by multiple process, 525 roasting in Allen-O'Hara furnace, 126 Bruckner furnace, 138 hand reverb, furn., 116 MacDougaU furn.. Anaconda, 108 Pearce furnace, 133 Wedge furn., 114 Wethey furn., 124 smelting Lake Superior copper ore, 371 oxide ore, 360 plant, 3S7 reverb, furn., 286 stall-roast, 81 Ziervogel process, 462 Cotton furnace, reverb, smelting, 255 Cottrell electric condensation. Anaconda, chloridizing roasting, 431 Balaklala, 229 Garfield, 231, 343 Raritan, 518 Counter-current system, leaching copper ore, 406 Cross-bar, multiple process, 504 washing, etc., 513 Crucible furnace, melting cement silver, 462 Crushing ore, roast-heap, 71 and mixing ore and salt, Longmaid- Henderson proc, 438 Crystallization and electrodeposition, treat- ment foul electrolyte, 521 blue vitriol, requirements, 474 copper, 4 method, treatment foul electrolyte, 521 plant, blue vitriol, Hofmann, 470 Crystals deposited copper, orientation, 503 Cupola copper, smelting native copperore, 362 matte, melting for converter, 320 Cupric carbonate, 52 chloride, 58 as solvent, oxide ore, 404 oxide, 52 decomposition by iron, 278 silicates, 53 sulphate, 55 sulphide, S4 Cupro-magnesium, addition fire-refining copper, 387 Cuprous chloride, 57 oxide, 51 behavior electrolysis, 486 in tough-pitch copper, 387 silicate, 52 sulphide, 53 -ferrous sulphide, 213 behavior in electrolysis, 486 Current and voltage vs. temperature, elec- trolyte, 480 density, electrolysis copper, 497 vs. cost of power, multiple process, 525 vs. precious metal loss, 512 electrolysis, 488 series system, Baltimore, 536 Data, acid converter, 302 blast-furnace, native copper ore, 371 oxide copper ore, 360 sulphide copper ore, 176 INDEX 547 Data, electrolysis, multiple process, 528 reverb, furn., fire-refining, 388 native copper ore, 363 reverb, turn., sulphide copper ore, 256 David-Manhcs converter, 305 David selecteur, 308 Dawson furnace, reverb., smelting, 255 Day, apparatus, blowing concentrate, 340, 347, 348 Decopperization of residue, Freiberg vitrioli- zation, 478 Delta metal, 37 Dense-poling, copper-refin. furn., 384 Depositing vat, multiple process, 506 series system, Baltimore, 537 Deposition on cathode, multiple proc, 511 Depreciation, copper smeltery, 358 Detroit Copper Co., blast-roasting, 140 Direct fusion, anode mud, 515 matte, 320 process, 293 Dissolver, vitriolization of copper, 478 Distance, bottom and sides of tank to cathodes, multiple process, 503 Doetsch process, leaching sulphide ore con- version into chloride, 432 Dorell converter hood, 315 DoubUng in converting, 325 Douglas r6asting furnace, 138 Down-draft blast-roasting, 143 Draft, natural vs. forced, reverb, furn., 267 oil-fired reverb, furn., 269 Drying furnace, pressed cement-silver, 461 Ductility, bronze, 41 copper, 6 Ducktown, Tenn., kernel-roast, 426 Dumps, leaching, 413 Durana metal, 38 Dust-arresting devices, 224 Dust-chamber, efficiency, flue-dust, 221 deposition in chamber. Anaconda, Can- anea. Copper Queen, 219, 220, Great Falls, 225, flue and chambers, blast-furnaces, 219 Dwight-Lloyd, agglomeration flue-dust, 233 roasting machine, 143 sintering purple ore, 454 -Messiter, ore-bedding system, 356 Dyblie converter valve, 306 Early converting, 299 Eastern U. S., vitriolization plant, 478 Economizers, reverb, furn., Copper Queen, 287 Garfield, 244 Edwards furnace, 116 Efliciency, blast, converter, 314 blast pyritic smelting, 196 dust-arresting devices, 221 Eguilles converter, 301 Elastic limit, copper, s Electric condensation, Balaklala, 229 Garfield, 231 converter, 343 Raritan, 518 conductivity brass, 32 cathode copper, cast vs. not cast, 505 Electric conductivity copper, 6, 7 tests, fire-refined copper, 386 connection, depositing vats, 510 converter, 309 furnace, refining copper, 372 Electrodeposition copper, Laszczynski proc- ess, 414 from solutions, 405 method, treatment foul electrolyte, 524 tubes, sheets, wire, 512 Electrodes, multiple system, manipulation, . s°s series system, Baltimore, 536 Electrolysis, as solvent, oxide ore, 404 copper, 482 matte, 482 metallic copper, 483 ore, 482 speise, 483 Electrolyte, arsenic vs. conductivity of cathode copper, 487 Electrolyte, Baltimore series plant, 536 composition, temperature, circulation, 492 conductivity, 489 foul, treatment, 520 Electrolytic plants U. S., 484 Electrolyzing, vat, multiple system, 506 series system Baltimore, 537 Elevation, effect on roasting, 136 smelting, 182 Elimination, As, Sb, Bi, in blast-furnace, 189 impurities, converting, 326 electrolysis metallic copper, 485 fire-refining copper, 383 heap-roast, 425 pyritic smelting, 196 reverb, furn., 279, 291 roasting coarse metal, 289 roast-smelting, 293 smelting roasted coarse metal, 291 through bottoms, 295 weathering, 421 Elkington process, electrolysis metallic cop- per; 484 Elongation, copper, 5 El Paso, tilting reverb, furn., casting copper, 401 E.m.f., deposition Cu, As, Bi, Sb, 487 Enrichment in acid and copper of electrolyte, 493 Evans-Klepetko furnace, 77, 105 Evaporating pan, Hofmann, 469 Evolution, reverb, matting furn., 237 Exchanging cathodes, 505 Expansion, linear, copper, 7 Extra process, reverb, smelting, 293 Feeding charge, blast furn., 165 reverb, furn., 281, 283 siliceous ore to acid converter, 325 Ferric chloride, as solvent for oxide ore, 403 sulphate as solvent for oxide ore, 403 reduction to ferrous sulphate, by pyrite, 422 by sulphurdioxide, 405 548 INDEX Ferrites, copper refinery slags, 383 Ferro-silicon, addition, fire-refining copper, 387 Ferrous chloride, as solvent for oxide ore, 403 Ferrous sulphide-cuprous sulphide, 213 Ferrous sulphide-iron, 212 Fettling, Anaconda, 284 Cananea, 284 Filter-bottom, leaching-vat, Longmaid-Hen- derson process, 450 Filtering, copper solutions, 406 Fine ore, roasting in shaft furn., 87 Fining, copper refin. furn., 382 Finishing on rod, basic converter, 342 roast, mechanical reverb, furn., Ziervogel process, 458 Fink smelter (reverb, smelting furn.), 255 Fire-refining impure copper, 372 Firing roast-heap, 77 Flapping, copper refin. furnace, 382 Floater, fusion by conversion, 283 Flow-sheet, Great Falls multiple plant, 527 Flue-dust, agglomerating in matte, 236 acid converter, 332 amount with blast-furnace, 234 arresting devices, 224 basic converter, 342 blast-furnace, 218 Great Falls, 221 roasting furnaces, collection, 147 smelting anode mud, electric condensa- tion, 518 treatment in the dry way, 232 in the wet way, Mansfeld, 459 Flume, precipitation mine-water, 408 Foaming of slag,acid converter, 325, basic converter, 340 Fore-hearth, blast-furn., see also settler, 169 Orford, 150 Formation temperatures ferro-calcic silicates, 181 Forms, copper of commerce, 10 Foul solutions, electrolysis, 520 Foundation, reverb, furn., concrete, 243 Fractional crystallization, nickel and copper sulphate, 521 Freeland charging-machine, 166 Freiberg kiln, 85 vitriolization plant, 477 process, 463 Freudenberg plates, flue-dust, 221' Froelich process, leaching sulphide ore, con- version into chloride with ferric chloride, 433 Fuel, blast-furnace, 182 consumption refining copper, 343 dust, reverb, furn., 267 reverb, furn. Canadian Copper Co., 24S partial pyritic smelting, 199 pyritic smelting, 193 reverb, furn., 265 Fulton, reverb, slags, 275 Furnace, blister, 291 metal, 289 Furnaces, chloridizing burnt pyrite, 442 copper refining, 388-390 Garfield basic converter, 343 blast-roasting, 139 electric condensation, 231 flue-dust, 343 MacDougall, furn., 109 Peirce-Smith converter, 343 Garretson furnace and process, 198 Gas absorption, brass, 32 bronze, 43 copper, 6 alloys, 24 analyses, reducing smelting, blast- furnace, 184 partial pyritic smelting, 201 poling, copper refin. furn., 385 Gases, blast-furnace, temperature, velocity, 218 withdrawal, 165 condensation, chloridizing roasting, 449 converter, 320 pyritic furnaces Mount Lyell, 194, 195 temperature. Great Falls, 225 velocity, volume, etc., Great Falls, 226 General arrangement Great Falls basic- converter, 349 sulphide plants, 353 German process, reducing smelting in blast- furn., 64 Gerstenhofer furnace, 87 Gibbs process, 455 Gin process, separation ferrous and cupric sulphates, 463 Giroux, hot-biast, 166, 182 Gmahlin-Shelby fettUng-device, 284 Gold and copper, 23 behavior electrolysis, 486 in matte, 215 loss in blast-furnace, 235 silver and copper, ternary alloys, 23 Good merchant brand, G. M. B., 294 ordinary brand (G. O. B.), 294 Gossage-tower, Longmaid-Henderson process, 4SO Grades of copper, 2 native copper ore, 361 Granby charging car, 166 Granularing refined block copper for vitrio- lization, 473 Great Cobar blast-furnace and converting department, 321 Great Falls anode, 502 basic converter, 349 blast-furnace, 152 bottom-blown converter, 300 casting machine, converter copper, 317 cathode, 504 converter, 301 acid lining material, 310, 311 copper refining furn., 377 depositing vat, 509 dust in chambers, 225 foul electrolyte, 5 20 gases, temperature and draft, 226 feeding blast-furn., 165 multiple plant, 527 new reverberatory furn., 254 new flue-system, ?2i INDEX 549 Great Falls plan of works, 222 plant, circulation of electrolyte, 495, 497 purification foul electrolyte, 523 reverb, furn., 253 settler, 169 slag-casting machine, 316 smeltery, 355 stack, 225 tank details, 504 Greenawalt process, leaching oxide ore, 419 Greenwood auxiliary slag-bowl, 175 smeltery, 355 Grondal, briquetting purple ore, 454 Ground, roast-yard, 74 Gumeschevsky, mine dump, leaching, 414 Gun metal, 45 Haas, vortex converter, 308 Haege wheel, vitrioUzation of copper, 478 Hand-converter, 309 Handling liquors, 406 Hardness, copper, $ Hayden, MacDougall furn., no series system, 534 tilting reverb, furn. casting copper, 401 Harz vitriolization process, 472 Headings, native copper ore, 361 Heap-roast, sulphatizing, 424 -roasting, 75 Heap, roasting i^i, 71 stall and kiln, compared, 86 Hearth-accretion, blast-furn., 235 -area, blast-furn., 161 blast-furnace, 157 -furnace, refining copper, 372 Heat balance, converter, 327 reverb, furn. Anaconda, 280 blast-furnace, 204, 208 efiiciency blast-furn., 189 Heberlein, agglomeration of flue-dust, 233 Heinze-Freeland furnace, 96 Hercules metal, 36 Herkules furnace, 96 Herreshoff blast-furnace, 149 roasting furnace, 92 High brass, 34 High-Ore, mine-water, precipitation plant, Butte, 409, 412 Hinckley Filter Co., roasting furn., 138 History, converting copper matte, 299 copper, I pyritic smelting, 191 Hoepfner process, leaching sulphide ore, 434 Hofman-Hayden Hallowell, chemical and physical changes for refining cath- ode copper, 394 Hofmann evaporating pan, 469 vitriolization process, 464 Hood, converter, 314 Hoopes conductivity bridge, 386 Horizontal converter, 305 Horizontal vs. upright converter, 308 Hot-blast, Giroux, 166 Hunt-Douglas process leaching ore, chlorina- tion with ferrous chloride, 434 Huntington-Heberlein pot, 139 Hybmette process, leaching copper ore, 429 Hydraulic converter, 309 press, cement silver, 461 Hydrochloric acid as solvent, oxide ore, 402 Hydrogen sulphide, precipitant for copper, 40s Hydrolization, effect upon deposition of arsenic, 487 Impoverishment in copper of electrolyte, 493 Impurities, anode and anode mud, 486 behavior, electrolysis of copper, 485 cathode copper, deposition, 512 effects on copper, 12 elimination, converting, 326 see elimination. Incorporation, flue-dust in blast-furnace slag, 233 converter slag, 232 Ingot-bar, 396 converter copper, 317 copper, 396 Insoluble anode, 512 International S. & R. Co., plant, 354 Iron and copper, 16 behavior electrolysis, 485 fire-refining copper, 392 -brass, 36 precipitant for copper, 404 Longmaid-Henderson process, 432 sulphide-iron, 212 -cuprous sulphide, 213 precipitant for copper, 405 roasting, 65 Jardine-Chadwick process, 455 Jerez Lanteira, barrel converter, 305 Jones-Bennetts waste-slag car, 175 Jones roasting furnace, 138 Kalchoid, 45 Kansas City S. &. R. Co., vitriolization plant. .471 . Karkarlinsk mine, Laszczynski process, 414 Kauffmann furnace, 96 Keeping hot empty converter, 321 Keller multiple-hearth roasting furnace, 126, 128 Kelley slag-casting machine, 316 Kelly-Thomson rabble-arm, 107 Kernel-roast, Agordo, Italy; Ducktown, Tenn., 426 -roasting, 67 Kiddie hot blast, 182 KiUani, behavior impurities, electrolysis, 485, 494 Kiln, automatic, for roasting, 87 Evans-Kelpetko, 97, 105 Freiberg, 86 Gerstenhofer, 87 heap, stall, compared, 86 Heinze-Freeland, 96 Herkules, 96 Herreshoff, 92 Kauffman, 96 MacDougall, 91 Maletra, 89 55° INDEX Kiln, Merton, 97 O'Brien, 95 Oker, 83 roasted pyrite, compounds in, 437 roasting in, 82 fine ore, 87 rough-roasting matte, Ziervogel process, 458 Sjostedt, 97 Spence, 90 Wedge, 112 Klepetko blast-furnace, 152 Klepinger copper- casting machine, 317, 319 Knapp-Kunze blast-roasting process, 138 Knudsen process, ig6 Kothny, chloridation of burnt pyrite, 441 Krohnke process, 404 Labor, basic converter, 343 Ladle, casting copper, 397 Laist process, leaching tailings. Anaconda, -Tanner converter hood, 3rs Lake Superior reverb, furn., 363 Large charges and long intervals, reverb. furn., 281 Laszczynski process, 4r4 Latent heat of fusion, copper, 7 Launder, wood-iron, vitriolization, 481 Leaching and precipitation, Ziervogel proc- ess, Mansfeld, 460 apparatus, method, 406 atacamite, 420 bronchantite, 420 chloridized ore, Longmaid-Henderson process, 450 copper ore, 402 outline of processes, 406 matte, 456 metallic copper, 471 mill tailings, 413 miiie dumps, 413 oxide ore, 414 Rio Tin to, 421 sulphate ore, 407 sulphide ore, conversion into chloride by oxidizing and chloridizing roast, 436 by cupric chloride, 434 by ferric chloride, 432 conversion into oxide by roasting, 43 1 sulphate, sulphatizing roast, 424 weathering, 420 with ferric sulphate, 430 in place, 424 oxidizing-roasting and chlorinating with ferric chloride, Hunt-Douglas, 434 vat, Longmaid-Henderson process, 450 Lead, addition, fire-refining copper, 387 and copper, 15 behavior electrolysis, 486 fire-refining copper, 393, 394 silver and copper, ternary alloys, 22 soaking anode mud, 515 Leghorn converter, 305 Length, reverb, furn. direct firing, 267 Leonard plant, mine-water, Butte, 412 Lime-coat, for cathodes, 381, 506 Lime, neutralization of SO3, Sprague proc. 229, precipitant for copper, 405 Limestone, unchanged in MacDougall kiln, 104 Lining acid converter, 310, 311 ore used, 310, 311 Anaconda basic converter, 345 Peirce-Smith, basic converter, 338 Steptoe basic converter, 346 Link-belt casting machine, refined copper, 400 Lithgow depositing vat, 508 multiple plantj 534 plant, circulation of electrolyte, 493 Lixiviation vat, Hofmann, 465 Lloyd, altitude and blast-roasting, 146 smelting at high altitude, 182 Longmaid-Henderson process, 436 Loss in copper, acid converter, 331 basic converter, 342, 343 blast-furnace, 189, 196, 235 slag, 217 reverb, furn., 286 precious metal vs. current density, 5x2 Low brass, 34 Lump ore, roasting in shaft" furnaces (kilns), 82 Luster, copper, 4 MacDougall furnace, 91 air-cooled, 108, log, no Machine, ramming acid converter lining, 311 Machinery bronze, 45 Magnesite bottom, reverb, furn., 261 brick, copper refining process, 3 74 hearth, reverb, furn., Canadian Copper Co., 24s Magnesium, addition, fire-refining copper, .387 Magnetism, copper, g Magnetite lining, basic converter, Wheeler- Krejci, 353 Maidenpec, Servia, sulphatizing roast, 427 Maletra shelf-burner, 89 Malleability, copper, 6 tests, fire-refining copper, 386 Mammoth smeltery, flue-dust collection, 228 Management, acid converter, 323 basic converter, 339, 343 blast-furnace, 189 partial pyritic furnace, 202 pyritic furnace, 196 reverb, furn., 280 Manganese and copper, 17 behavior electrolysis, 485 brass, 38 bronze, 38, 47 copper, addition, fire-refining copper, 387 sulphide, roasting, 65, 67 Mangano-silicon, addition, fire-refining cop- per, 387 Manh6s converter, 300 Manipulation, converter, 309 electrodes, multiple process, 505 INDEX SSI Mansfcid, blast-furnace gas, 184 tieatment of flue-dust in the wet assay, 4S9 Ziervogel process, 457 Manufacture, brass, ^^ bronze, 44 Marchese process, 482 Marketing copper ore, 62 Mason Valley, agglomeration flue-dust, 233 Mass, native copper ore, 361 Mathewson blast-furnace, 153 method of charging reverb, furn., 283 Matte, 212 agglomerating flue-dust, 236 bath of, in reverb, furn., 282 converting, 298 disposal, 175 electrolysis, 482 leaching, 456 mold, 175 precious metal, 215 pyritic concentration, 236 vitriolization, 463 Hofmann proc, 464 Matthiessen standard, 8 Matting, reverb, furn., 236 ■Maurer cathode, 504 Mayer process, 455 McMurtrie-Rogers blast-roasting proc, 138 Mechanical charging copper refin. furn., 381 down-draft multiple-hearth muffle-furn.. Wedge, 446 -hearth rever. and mufBe-furn., Wedge, 448 muffle chloridizing furn., 446 properties brass, ordinary temp., 30 properties, brass, varying temperatures, 31 bronze, ordinary temperature, 41 varying temperatures, 42 reverb, chloridizing furn.. Wedge, 44S stirring, leaching copper ore, 406 Medal bronze, 46 Melting cathodes, system Addicks-Marks, 401 cold matte in converter, 321 copper refin. furn., 382 point, copper, 7 points, common alloys, 25 Merton furnace,, 97, 121 Metal, coarse, 288 furnace, 289 slag, 289 Metallic copper production in blast-furn. from matte, 236 copper, vitriolization, 472 Metallurgical treatment, copper ore, 62 Michigan copper ore, 62 smeltery, reverb, furn., silica-brick bottom, 366 separation of smelting and refining, 369 Microscopic tests, fire refining copper, 386 Miedzianka mine, Laszczynski proc, 414 Mill tailings, leaching, 413 Mine and tail water, Bisbee, 413 Butte, 412 Copper Queen, 413 Mine and tail water, Schmoellnitz, 408 dumps, leaching, 413 la Motte, blast-roasting, 140 water, acid, freed from copper, 413 recovery of copper, 407 Mineral, native copper ore, 361 Mirror metal, 46 Mode of operating reverb., smelting native copper, 368 Mold, copper ingot, 396 pig, 359 wire-bar, 397 Monier process, sulphatizing roast, 427 Montana copper ore, 60 Montejus, acid liquor, 466 Morenci, blast-roasting, 140 Morrow clip for anode, 502 cathode, 504 Moss-copper, 215 Mount Lyell blast-furnace, 152 converter, 301 converting plant, 320 Movable settler, 172 Mud, anode, 513 Mufiie chloridizing furnace, Natrona, 445 roast, sulphatizing, 427 roasting furnace, 138 vs. reverb, furn., chloridation burnt pyrite, 442 Multiple plants of the U. S., 528 process or system, electrolysis, 491 vs. series copper, silver content, 537 system, 538 Native copper ore, 62 smelting, 361 Natrona, chloridizing muffle furn., 445 Neill process, leaching ore, 418 Neutralization SO3, Sprague process, 229 Nevada copper ore, 61 Nickel, action in electrolyte, 492 and copper, 17 behavior electrolysis, 485 brass, 39 scorification in refining black copper, 473 Nichols Copper Co., series plant, 535 Nichols-James process, 293 Nickel, behavior, fire-refining copper, 392, 394 O'Brien furnace, 95 Ochre, from mine water, 409 O'Hara roasting furnace, 124 Oil-burner, Shelby, Sorensen, 272, 273 -poling, copper refin. furn., 385 Oil in blast-furnace, 182 in partial pyritic smelting, 201 reverb, furn., 269 use as cover in depositing tanks with insoluble anodes, 522 Oker, chloridizing reverb, furn., 442 kib, 83 vitriolization plant, 475 Opening roast-heap, 78 Operation of reverb, furn., smelting native copper ore, 368 Ordinary process, reverb, furn., 288 SS2 INDEX Ore-bedding, system Copper Queen, 357 Dwight-Messiter, 356 Ore, charged, basic converter, 339, 340 crushing and sizing, for roast-heap, 74 for acid lining of converter, 310 siliceous, basic converter, 339 slag, smelting native copper, 562 Orford fore-hearth, 150 Orientation, crystals electro-depositedcopper S03 . Overblowing in converting, 325 Overflow slag-pot, 173 Overpoling, fire-refining copper, 387 Oxide ore, 61 reduction and concentration, 359 smelting, 358 Oxidizer, vitriolization of copper, 478 Oxygen and copper, 13 behavior, fire-refining copper, 393 Packing of matte in Hofmann lixiviation tank, 466 Pan evaporator, Hofmann, 469 for solution of blue vitriol, 476 Parral tank, leaching copper ore, 406 Parrott converting plant, 320 Partial pyritic smelting process, 64, ig8 Patching basic converter, Anaconda, 346 Pearce gold separating process, 298 turret furnace, 132 Peirce, mechanical casting refined copper, 398 -Smith converter, 335 Percolation, method of leaching, 406 Perth Amboy cathode, 504 Peyton Chemical Works, reverb, furn., 255 Phosphor bronze, 46 copper, addition, fire- refining copper, 387 Phosphor-tin, 44 Phosphorus and copper, 25 Physical changes, fire-refining cathode copper, 394 properties, copper, 4 Plant, Dwight-Lloyd blast-roasting, 145 Longmaid-Henderson, old and new, 438 smelting, cost of, 357 Plate-baffles, flue-dust, 221 -pitch for anodes, 500 slag, 281 Platinum, behavior electrolysis, 486 Pohl6 air-lift pump, 526 circulation electrolyte, 497 Poisoning furnace bottom, 294 Poisonous effect of copper, 232 Poling, copper refin. furnace, 384 Pollock waste-slag car, 175 Porphyritic slag, reverb, furn., 275 Port Kembla multiple plant, 534 Pouring spoon, Bennetts, 317 temperature, alloy, 25 Precious metal in matte, 215 loss in blast-furnace, 235 vs. current density, 512 Precipitants of copper, 404 Precipitation of copper, Longmaid-Hender- son process, 452 method and vat, 406 Precipitation of silver and gold, Longmaid-, Henderson process, 454 Ziervogel proc, 461 vat, Longmaid-Henderson proc, 452 Predazzo, Italy, sulphatizing muffle roast, 427 Preheating air, pyritic smelting, 193 blast, 182 Preparation, alloys, 25 Pressure, air and steam, atomizing oil, 269 tank for acid liquor, 466 Process, Augustin, matte, 456 metallic copper, 472 Borchers-Franke-Giinther, 483 Bradley, leaching copper ore, 429 Carmichael, leaching ore, 415 Claudet, 454 direct, 293 Doetsch, leaching sulphide ore, 432 Elkington, 484 extra, 293 fire-refining copper, 372 FroeUch, leaching sulphide ore, 433 Gibbs, 45 s Gin, separation ferrous and cupric sul- phates, 463 Greenawalt, leaching oxide ore, 419 Harz vitriolization, 472 Hoepfner, leaching sulphide ore, 434 Hunt-Douglas, leaching ore, 434 Hybinette, leaching copper ore, 429 Jardine-Chadwick, 455 Laist, leaching Anaconda tailings, 431 Laszczynski, leaching and electrode position, 414 Longmaid-Henderson, 436 Marchese, 482 Mayer, 455 Monier, sulphatizing roast, 427 multiple, electrolysis, 491 vs. series, 538 Neill, leaching ore, 418 Nichols-James, 293 ordinary in reverb, furn., 288 Roessler, refining jeweler bars, 298 selecting, 293 Siemens-Halske, leaching and electro- deposition, 430 Snelus, 455 Slater, solvent ferric chloride, 434 smelting native copper ore, 362 Stadtberge, leaching ore, 414, 419 Terrino, sulphatizing with ferric nitrate, 427 the Argo, 295 the Pearce gold-separating, 298 Van Arsdale, leaching ore, 418 Ziervogel, 456 Processes, leaching copper ore, outline, 406 Producer gas, from coal, reverb, furn., 265 from wood, reverb, furn., 268 ignition in blast-roasting, 146 Production of copper, 2 Progress, chloridation burnt pyrite, 445, 448 Prosser-Ladd, charging machine, 381 Punching tuyeres, basic converter, 340 Purification fouled electrolyte, 520 INDEX 553 Purple ore, 453 Putting in, sand bottom, 261 Pyrite, kiln-roasted, compounds in, 437 Pyritic bosh, 190 concentration of matte, 236 process, 64 smelting, i8g chemistry, 194 Quartz, siUca bottom, 261 Quicksilver and copper, 50 Rabbling, copper refin. furn., 382 Ragging, 75 Ramfin-Beskow, chloridizing furn., 449 Ramming acid converter lining, 311 Randolph series system, 535 Rare metals in converter copper, 332 Raritan, anode-mud treatment, 516 depositing vats, 506 multiple plant No. 2, 532 plant, circulation of electrolyte, 494, 497 Rath, first conversion of matte, 299 Rating, blast-furnace smelting process, 236 Reactor process, 293 Record, automatic, reverb, furn., 285 low vs. high draft, oil-fired reverb, furn., 270 starting Anaconda reverb, furn., 265 Red metal, 295 Reducing smelting, blast-furnace, 178 chemistry, 183 References to blast-furnaces, 176 to reverb, smelting plants, 255 Refined vs. converter anode, 499 Refinery-slag, 289 Refining black copper for vitriolization, 473 fire, impure copper, 372 slag, smelting native copper ore, 362 tower for copper liquor, 466 Regenerative reverb, furn., Great Falls, 253 Regular brass, 34 bronze, 44, 45 Regule, 29s Regulus, spongy, 295 Reinartz, research into Carmichael proc, 416 Removal anode mud, 513 Removing cathodes, 505 Repath casting machine, refined copper, 400 Requirements, chloridation of pyrite, 442 pyritic smelting, 190 Residue Freiberg vitriolization, retreatment, 478 Longmaid-Henderson proc, 453 Results, Longmaid-Henderson process, 455 partial pyritic smelting, 202 pyritic furnace, 196 Reverb, chloridizing furn., Oker, 442 furnace. Anaconda, 1903, 243 1908, 243 heat balance, 280 bringing forward matte, 288 calculation of charge, 279 Calumet and Hecla, 366 Canadian Copper Co., dust-fired, 245 Cananea, 247 clay-brick bottom, 260 cleaning slag, 286 Reverb, furnace, coal, 265 Colorado Smelting Co., 238 cost of smelting, 286 elimination of impurity, 279 fettling, 284 fuel, 265 fuel-dust, 267 Great Falls, 253 height of roof, oil and coal-fired, 270 hot vs. cold calcine, 280 large charges, long intervals, 281 length with direct firing, 267 losses, 286 loss in copper, 286 management, 280 mechanical, finishing-roast, Ziervogel process, 458 natural vs. forced draft, 267 oil, 269 Peyton Chemical Works, 255 plant of Copper Queen, 286 producer gas from coal, 265 from wood, 268 references, 255 slag, 275 small charges and short intervals, 283 smelting, chemistry, 278 native copper ore, 362 Steptoe Valley, 271 tables, 256-259 temperature, record, Anaconda,' 282, tilting, for casting copper, 401 working bottom, 260 settler, blast-furnace-slag, 232 smelting, 236 wood, 268 vs. blast-furnace matting, 287 vs. muffle-furn., chloridation burnt pyrite, 442 Revolving barrel, leaching copper ore, 406 Rhodes matte-mold, 175 Rio Tinto, leaching, 421 sulphatizing heap-roast, 424 Roast-heap, 75 smelting vs. converting, elimination of impurities, 326 speise, 216 Roaster slag, 289, 292 Roasting, 291 and reduction proc, blast-furn., 64 and solution, anode mud, 517 apparatus, 70 effect of elevation, 136 fine ore in shaft furnace, 87 furnace, Allen-O'Hara, 174 automatic, 87, 138 Brown horseshoe, 129 Brown straight line, 124 Bruckner, 135 Bouglas, 138 Edwards, 116 Evans-Klepetko, 97, 105 Gerstenhofer, 87 hand reverb., 115 Heinze-Freeland, 96 Herkules, 96 Herreshoff, 92 554 INDEX Roasting furnace, Hinckley Fiber Co., 138 Jones, 138 Kauffraann, g6 Keller multiple-heartli, 126, 128 MacDougall, 91 Maletra, 8g Merton, 97, 121 muffle, 138 O'Brien, 95 O'Hara, 124 Pearce turret, 132 reverberatory, 115 Ropp, 122 see kiln, Sjostedt, 97 Spence, 90 Wedge, 112 Wethey, 123 multiple-hearth, 126 in automatic furnaces, 138 heaps, 71 kilns, 82 shaft furnaces, 82 stalls, 79 lump-ore in shaft furnaces (kilns), 82 sulphide copper ore, 65 summary, 147 Roast yard, 71 Rochlitz, leaching ore with hydrochloric acid, 419 Roesing wire-system, Great Falls, 221 Roessler process, refin. jeweler bars, 298 Rolling electrodes, copper suited, 536 Ropp furnace, 122 Rosette copper, 372 Rough roasting kiln, Ziervogel process, 458 Ruebel alloy, 39 Sacio, smelting at high altitude, 183 Sampling copper, 22 Sand bottom, reverb, furnace, 260 Salting out of small crystals, blue vitriol, 47J: Savelsberg pot, 142 ScaUng anodes, multiple process, 502 Schmoellnitz, mine-waters, 407 Scrap formed, multiple process, 499, 502, S" . Screen analysis: calcine, Anaconda, 108 Edwards, furn., Goldfield cons., 121 concentrate, MacDougall furn., Hayden, no concentrate Morenci, 140 flue-dust. Mason Valley, 233 Scrubbing flue-dust, anode-mud furnace, SIS Selecteur, David, 308 Selecting process, 293 Selenide of copper and silver, 517 Selenium and copper, 21 behavior electrolysis, 486 recovery from flue-dust, 518 Series system, refining copper, 534 vs. multiple copper, silver content, 537 system, 538 Set copper, copper refin. furnace, 383 Settler, biast-furnace, 169 Settler, oil-fired, reverb furn., 232 Sewer pipe, solution vat, vitriolization, 480 ' Shaft, blast-furnace, 161 furnace, roasting in, 82 fine ore, 87 Shannon Copper Co., sulphatizing heap-roast, 429 . . sulphatizing oxide ore with ferric sulphate, 430 Shelby, ball converter and valve, 306 blast-furnace, 157 matte-ladle, 175 oil-burner, 272 Shelf-burners, 89 Siemens-Borchers, circulation of electrolyte, 49S -Halske proc, leaching, electro-deposi- tion, 430 regenerative chambers, reverb, furn., Great Falls, 253 Siphon, series tank, Baltimore, 536 Silica, addition to converter charge, 326 and base reqiured for slag formation, I8S bottom, reverb, furnace, 260 brick, bottom reverb, furn., 265 Mich, smeltery, 366 content, converter slag, 340 Silicon, addition, fire-refining copper, 387 bronze, 47 and copper, 26 Silver and copper, 22, 515 and oxygen, 517 behavior, electrolysis, 486 fire-refining copper, 393, 394 Bow plant, mine water, Butte, 412 -copper selenide, 417 gold and copper, 23 in matte, 215 lead and copper, 22 loss, converting, 333 in blast-furnace, 235 Sintering purple ore, 454 Size of anode, multiple process, 500 Sizing ore, roast-heap, 74 Sjostedt furnace, 97 Skimming slag, converter, 332 Slag, basic converter, 340 blast-furnace, 217 * casting machine, 175, 316 copper content and length of reverb. furn., 267 formation, silica and base required, 185 forming stage, converting, 323 fusion anode mud, 517 oxide ore, 361 pot, 174 partial pyritic smelting, 198 pyritic smelting, 192 reducing smelting, blast-furnace, 179 reverb, furnace, 275 loss of copper, 286 Slags, copper refin. furn., 383 Slater process, use of ferric chloride, 434 Slime concentrate, native copper ore, agglo- meration, briquetting, 370 Small charges and short intervals, reverb, furn., 283 IiSDEX 5SS Smeltery, Anaconda, 354, 355 British Columbia Copper Co., 355 Canadian Copper Co., 355 Canancii, flow sheet, 355 cost, 357 Smeltery, (Ireat Falls, 222, 355 Smelting column, blast-furnace, 162 copper ore, 63 in converter, 298 in the blast-furnace, 148 native copper ore, 361 and by-products in blast-furnace, 370 in reverb, furn., 362 plants. Lake Superior ore, 361 oxide ore, 358 power, blast-furnace, 236 pyrite, 189 and refining, native copper ore in sepa- rate furnaces, 369 reverb, furn., 236 sulphide ore, 63 Smith series system, 535 Snelus process, 455 Sodium sulphide, precipitant for copper, 405 Solution of anode mud in sulphuric acid, 515 vats, special forms, vitriolization, 480 Solvents, for copper ore, 402 Sorensen oil-burner, 273 Sow, blast-furnace, 23 s Spain, blast-roasting, 140 Special brass, 36 bronzfe, 44, 46 Specifications, copper, 11 Specific heat, copper, 7 Speculum metal, 46 Speed of elimination of components, con- verting, 326 Speise, 216 electrolysis, 483 Spence furnace, 90 Spongy regulus, 295 Spout, trapped of blast-furnace, 157 Sprague process, 229 Stack, Great Falls, 225 Stadtberge, leaching oxide ore, 414, 419 Stahl, chemical changes, fire-refining copper, 393 . leaching ore with hydrochloric acid, 414 Stall, heap, and kiln, compared, 86 roasting, 79 Stallman converter, 301 Standard brand of copper, 294 Starting Anaconda reverb.. furn., 265 sheet, multiple process, 502 tanks, number, 503 Statistics of copper, 2 Steam-pressure, atomizing oil, 269 Steam vs. oil, atomizing oil, 269 Steptoe Valley, arrangement oil burners, 273' basic 'converter, 344, 346 MacDougall furnace, 108 oil-burner, 273 Peirce-Smith converter, 344 smelting, record low vs. high draft in oil-fired reverb, furn., 270 Sterro metal, 37 Sticht blast-furnace, 157 Stiriing vs. Babcock and Wilcox waste-heat boiler, 243 Stolberg works, electrolysis of matte, 482 Straightening cathodes, 500, 50.5 Stripping in Pearce gold process, 298 roast-heap, 78 sheet, multiple process, 502 tank, multiple process, 502 Structure of sample, fire-refin. copper, 386 Sulphatizing heap roast, 424 muffle roast, 427 Sulphide copper ore, 60 ore leaching after conversion into sul- phate, sulphatizing roast, 424 weathering, 420 after oxidizing and chloridizing roast, 4-36 conversion into chloride with cupric chloride, 434 with ferric chloride, 432 oxide by roasting, 431 sulphate with ferric sulphate, 430 oxidizing-roasting and chlorinating with ferrous chloride, Hunt-Doug- las, 434 Plant, general arrangement, 353 roasting, 65 smelting, 63 Sulphur and copper, 21 behavior, fire-refining copper, 392 dioxide, precipitant for copper, 405 recovery from roast-heap, 77 Sulphuric acid, as solvent 'oxide ore, 402, 403 Support, bottom, copper refin. furn., 373 Suspension of anodes, multiple process, 501 of cathode, multiple process, 504 Swedish process, reducing smelting in blast- furnace, 64 System, multiple, electrolysis, 491 series, electrolysis, 534 Szalathna, Hungary, sulphatizing roast, 427 Table, blast-furnace, 176 converter, 302 reverb, matting furn., 256, 259 Tacoma, tilting reverb, furn., casting copper, 401 Tailings, leaching, 413 Tamping converter Uning, 311 hand vs. machine, 313 Tanks, covering in electrolysis of copper, 497 deposition, multiple process, 506 series process, 535 Tellurium and copper, 21 behavior electrolysis,486 recovery from flue-dust, 518 Temperature, basic converting, 341 electrolyte, 493 fire-refining copper, 49s flue gas. Great Falls, 225 partial pyritic slags, 198 record, coal-fired reverb, furn. Anaconda, 282, 283 oil-fired reverb, furn., Steptoe \'alley, 272 Tennessee Copper Co., blast-furnace top, 167 SS6 INDEX Tennessee Copper Co., settler, 170 Tensile, strength, bronze, 41 starting sheet, multiple process, 503 Terrino process, sulphatizing with ferric nitrate, 427 Tests, fire-refining copper, 386 Teziutlan, feeding blast-furnace, i"67 Tharsis, chloridizing muffle furnace, 445 Thermal balance, converter, 327 partial pyritic matte charge, 208 ore charge, 204 reverb, furn.. Anaconda, 280 conductivity copper, 7 Thickness of anode, multiple process, 500 Thofern electric connection of depositing vats, sio Throat area, blast-furnace, 161 Tin, behavior electrolysis, 486 brass, 38 Titanium and copper, 50 Tobin bronze, 37 Tonnage and length of reverb, furn., 267 Tooele, plant, 354 smeltery, MacDougall furn., 109 Tough-pitch copper, 395 poling copper, 385 _ Tower-liquor, Longmaid-Henderson proc, .45° precipitation for mine-water, 409 Traylor waste-slag car, 174 Tungsten and copper, 50 Torsion tests, fire-refined copper, 386 Tough-pitch copper, 385 Transfer-car, electrodes, multiple process, 501 Treatment anode mud, 515 flue-dust, 232 fouled electrolyte, 520 speise, 216 Triaxial diagram, matte, 214 Tridymite, silica bottom, 261 Trough converter, 305 Truswell, closed anode-mold, 502 Tuyere-stock, blast-furnace, 165 Turning roast-heap, 78 Tymp, blast-furnace, 159 Uchatius gun metal, 45 Ultimate strength, copper, 5 Underground flue, reverb, furn., 244 United States Metals Co., copper refin. furn., 377 United Verde Co., Wedge furnace, 114 Up-draft blast roasting, 139 Upright converter, 300 ■OS. horizontal converter, 308 U. S., electrolytic plants, 484 Uses of copper, 1 1 Usual form, converter lining, 312 Utah copper ore, 61 Vanadium and copper, 50 Valve, Dyblie, 306 Shelby, 306 Vitriolization at Freiberg, 477 at Oker, 475 of matte, 463 Hofmann proc, 464 of metallic copper, 472 Vitriolization plant, Eastern U. S., 478 special solution vats, 480 V-method, heap-roast, 79 Volatilization, copper, 7 Van Arsdale process, leaching ore, 418 Voltage and temperature vs. current density, 480 drop, electrolysis copper, 497 ratio and tem. vs. current density, 480 Walker casting machine, refin. copper, 398 dissolving machine, vitriolization, 480 -Murphy air-jacket, warming blast, 360 system of depositing vats, 507 Raritan plant No. 2, 532 Wallaroo, blast-roasting, 141 tailings leaching, 413 Wanjukow, chemical changes, fire-refin. copper, 391 solubility CU2S in slag, 180 Waste-heat boiler, 243 liquor, Longmaid-HendersOn proc, 454 slag, cupola, native copper ore, 362 disposal, 173 Water as solvent, oxide ore, 402 Water-jackets, blast-furnace, 161 Watt-hour and temperature vs. current density, 480 Wear of converter lining, 311 Weathering sulphide ore, 420 Wedge furnace, 112 mechanical muffle chloridizing furnace, 446 reverb, chloridizing furnace, 445 multiple-hearth mechanical down-draft muflle furnace, 446 reverb, and muffle furnace, 448 sulphatizing muffle roast, 427 Welsh process, 236 Wethey furnace, 123 multiple-hearth roasting furnace, 126 Wheeler-Krejci, magnetite lining, basic con- verter, iSi White brass, 36 metal, 45, 289, 290 Willyoung coilductivity bridge, 386 Witherell down-draft oxidizer, vitriolization, 480 Wood in blast-furnace, 182 reverberatory furnace, 268 Working bottom, reverb, furn., 260 height, blast-furnace, 162 Widnes, chloridizing muffle furn., 445 Wire-baffles, flue-dust, 221 -bar, copper, 397 -system, flue-dust. Great Falls, 221 Yampa smeltery, blast-roasting, 140 Edwards furnace, 116 Yard, roast-heap, 71 Yield in metal, blast-furnace, 235 Ziervogel process, 462 Zinc and copper, 28 behavior electrolysis, 485 oxide, neutralization SOs, Sprague proc- ess, 229 Ziervogel process, 456