»:^<^>^>^^^:>'siX:■•^^c•'':M^,/ .♦^"V. \'^'1V* .^'' '^. / '^^ -^ aS' IC 8924 Bureau of Mines Information Circular/1983 Updated Process Flowsheets for Manganese Nodule Processing By Benjamin W. Haynes, Stephen L. Law, and Riki Maeda UNITED STATES DEPARTMENT OF THE INTERIOR fC^ !*!*'• ^MUM^Jl^^ Information Circular; 8924 w Updated Process Flowsheets for Manganese Nodule Processing By Benjamin W. Haynes, Stephen L. Law, and Riki Maeda UNITED STATES DEPARTMENT OF THE INTERIOR James G. Watt, Secretary BUREAU OF MINES Robert C. Horton, Director As the Nation's principal conservation agency, the Department of the Interior has responsibility for most of our nationally owned public lands and natural resources. This includes fostering the wisest use of our land and water re- sources, protecting our fish and wildlife, preserving the environmental and cultural values of our national parks and historical places, and providing for the enjoyment of life through outdoOT recreation. The Department assesses our energy and mineral resources and works to assure that their development is in the best interests of all our people. The Department also has a major re- sponsibility for American Indian reservation communities and for people who live in Island Territories under U.S. administration. This publication has been cataloged as follows: Haynes, Benjamin W Updated process flowsheets for manganese nodule processing. (Information circular ; 8924) Bibliography: p. 11 Includes index. Supt. of Docs, no.: I 28.27:8924. 1. Manganese— Metallurgy. 2. Manganese nodules. 1. Law, Stephen L, 11. Maeda, Riki, III. Title. IV. Series: Information circu- lar (United States. Bureau of Mines) ; 8924. TN295.U4 [TN799.1V13] 622s [669'.732] 82-600368 For sale by the Superintendent of Documents, U.S. Gorernment Printing Office Washington, D.C. 20402 <» CONTENTS Page Page ^ Abstract 1 Detailed process descriptions 10 -7 Introduction 2 Major assumptions 10 "y Acknowledgments 2 Materials handling 10 Manganese nodule processing overview 2 Summary 10 Generic types of processes 2 References 11 Processes most likely for first-generation commercial use 3 Appendix A. — Gas reduction and ammoniacal leach Summary process descriptions 3 process 12 Gas reduction and ammoniacal leach process 3 Appendix B.—Cuprion ammoniacal leach process 28 Cuprion ammoniacal leach process 5 Appendix C— High-temperature and high-pressure High-temperature and high-pressure H2SO4 leach H2SO4 leach process 44 process 6 Appendix D. — Reduction and HCI leach process 62 Reduction and HCI leach process 7 Appendix E.— Smelting and H2SO4 leach process 79 Smelting and H2SO4 leach process 8 ILLUSTRATIONS 1. Gas reduction and ammoniacal leach process 4 2. Cuprion ammoniacal leach process 5 3. High-temperature and high-pressure H2SO4 leach process 6 4. Reduction and HCI leach process 7 5. Smelting and H2SO4 leach process 9 A-1. Key to symbols 16 A-2. Ore processing and drying 17 A-3. Reduction 18 A-4. Leaching-aeration 19 A-5. Solid-liquid separation 20 A-6. Liquid ion exchange-extraction 21 A-7. Cobalt stripping-organic purge 22 A-8. Liquid ion exchange-stripping 23 A-9. Copper electrowinning — commercial 24 A-10. Nickel electrowinning — commercial 25 A-11. Cobalt recovery 26 A-1 2. Ammonia recovery 27 B-1. Key to symbols 32 B-2. Ore preparation 33 B-3. Reduction-leach 34 B-4. Oxidation-leach 35 B-5. Solid-liquid separation 36 B-6. Liquid ion exchange-extraction 37 B-7. Cobalt stripping-organic purge 38 B-8. Liquid ion exchange-stripping 39 B-9. Copper electrowinning — commercial 40 B-10. Nickel electrowinning- -commercial 41 B-11. Cobalt recovery 42 B-1 2. Ammonia recovery 43 C-1. Key to symbols 48 C-2. Ore processing 49 C-3. Leaching 50 C-4. Solid-liquid separation 51 C-5. Pregnant liquor pH adjustment 52' C-6. Copper liquid ion exchange 53 C-7. Cobalt stripping-organic purge 54 C-8. Copper electrowinning — commercial 55 C-9. Copper raffinate pH adjustment 56 C-10. Nickel liquid ion exchange-extraction 57 C-11. Nickel liquid ion exchange-stripping 58 C-1 2. Nickel electrowinning — commercial 59 C-1 3. Cobalt recovery 60 C-1 4. Ammonia recovery 61 ILLUSTRATIONS— Continued Page D-1. Key to symbols 66 D-2. Ore processing and drying 67 D-3. Hydrochlorination 68 D-4. Leaching and washing 69 D-5. Copper liquid ion exchange 70 D-6. Copper electrowinning — commerical 71 D-7. pH adjustment and cobalt extraction 72 D-8. Nickel liquid ion exchange 73 D-9. Nickel electrowinning — commerical 74 D-10. Manganese recovery 75 D-11. Cobalt recovery 76 D-12. HCI recovery 77 D-1 3. Waste recovery 78 E-1. Key to symbols 84 E-2. Ore preparation and drying 85 E-3. Reduction 86 E-4. Smelting 87 E-5. Converting 88 E-6. Ferromanganese reduction 89 E-7. Matte leaching 90 E-8. pH adjustment 91 E-9. Copper liquid ion exchange 92 E-10. Cobalt stripping-organic purge 93 E-11. Copper electrowinning — commercial 94 E-1 2. Copper raffinate neutralization 95 E-1 3. Nickel liquid ion exchange-extraction 96 E-1 4. Nickel liquid ion exchange-stripping 97 E-1 5. Nickel electrowinning — commercial 98 E-1 6. Cobalt recovery 99 E-1 7. Ammonia recovery 100 TABLES A-1. Operating parameters for gas reduction and ammoniacal leach process 13 B-1. Operating parameters for Cuprion ammoniacal leach process 29 C-1. Operating parameters for high-temperature and high-pressure H2SO4 leach process 45 D-1. Operating parameters for reduction and HCI leach process 63 E-1. Operating parameters for smelting and H2SO4 leach process 80 UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT pet percent psi pound per square inch psig pound per square inch, gage tpd ton per day tpy ton per year A/m^ atm °C gpi hr ampere per square meter atmosphere degree Centigrade gram per liter hour in inch kg lb kilogram pound UPDATED PROCESS FLOWSHEETS FOR MANGANESE NODULE PROCESSING By Benjamin W. Haynes,^ Stephen L. Law,^ and Riki Maeda^ ABSTRACT The Bureau of Mines, in cooperation with the National Oceanic and Atmospheric Adminis- tration (NOAA), has updated a 1977 NOAA report prepared by Dames & Moore entitled, "Description of Manganese Nodule Processing Activities for Environmental Studies." This updated report contains detailed flowsheets and descriptions of the five potential first- generation nodule recovery schemes now considered most likely to be used by industry; they are high-temperature gas reduction and ammoniacal leach, Cuprion ammoniacal leach, high-temperature and high-pressure H2SO4 leach, reduction and HCI leach, and smelting and H2SO4 leach. The first three processes are three-metal recovery schemes (Cu, Ni, and Co) with the option of Mn recovery from the tailings. The remaining two processes are four-metal (Cu, Ni, Co, and Mn) recovery schemes. All except the HCI process are assumed to use a nodule feed rate of 3 million tons per year (dry basis). Final metal products are Co powder and cathode Cu and Ni. A minor amount of Ni is also recovered as powder, and some Cu and Zn are recovered as mixed sulfides. Manganese in the four-metal processes is recovered as either manganese metal or ferromanganese. ^Supervisory research chemist, Avondale Research Center, Bureau of Mines, Avondale, Md. ^Research supervisor, Avondale Research Center, Bureau of Mines, Avondale, Md.. ^Chemical engineer, Avondale Research Center, Bureau of Mines, Avondale, Md. INTRODUCTION This report is one in a series of reports issued by the Bureau of Mines as part of a research project entitled, "Analysis and Characterization of Potential Manganese Nodule Processing Rejects." Deep seabed mining for manganese nodules, includ- ing the processing of nodules to recover value metals, raises a variety of environmental, social, and economic considerations. To address the waste management aspects of the recovery of value metals from nodules, the National Oceanic and Atmo- spheric Administration (NOAA), U.S. Department of Commerce, the Environmental Protection Agency (EPA), and the Bureau of Mines and the Fish and Wildlife Service, U.S. Department of the Interior, after consultation with affected and concerned interests, have agreed to embark on a multiyear cooperative research program that has the following overall objective: "To provide information needed by Federal and State agencies in preparation for receipt of industry's commer- cial waste management plans." The NOAA-funded research conducted by the Bureau of Mines has the objective of obtaining a first-order chemical and physical characterization of reject waste materials (tailings) from the types of manganese nodule processing techniques representative of those being developed by industry. The result of this research is expected to be a technical report that can be used by (1) environmental scientists in subsequent research to assess the potential effects of waste management alternatives, and (2) regulatory agencies in the determination of what standards and test requirements must be met. This is expected to facilitate the development of a basic framework that accommodates the desire to assure protection of the environment and the development of a new minerals process- ing industry. In order to adequately assess the potential effects of dispos- ing of reject waste materials from manganese nodule processing, appropriate process flowsheets must be developed so that the processes representative of first-generation nodule process- ing plants can be simulated and waste can be generated in the laboratory. During August 1977, NOAA published a report prepared under contract by Dames & Moore entitled "Description of Manganese Nodule Processing Activities for Environmental Studies" (7)^ This report considered five different processing techniques or "roadmaps" (flowsheets). Three of these five, designed to produce Cu, Ni, and Co as primary products, are called "three-metal plants" and two are "four-metal plants" designed to produce manganese as a fourth primary product. It should be noted that three-metal plants could be designed to produce some manganese if market conditions are favorable. The processing techniques identified in the NOAA report are as follows: 1 . Gas reduction and ammoniacal leach process. 2. Cuprion ammoniacal leach process. 3. High-temperature and high-pressure H2SO4 leach proc- ess. 4. Reduction and HCI leach process. 5. Smelting and H2SO4 leach process. This Bureau of Mines report has taken the Dames & Moore 1 977 flowsheets and used the input from industrial and other concerned parties to update the report and present, where necessary, changes and modifications in the flowsheets. The flowsheets presented in this report are adaptations of the original flowsheets presented in reference 7. Only minor changes have been made in the sections common to all flowsheets. All flowsheets, however, are presented for completeness. This report, unlike the 1 977 study, contains no energy balance or material balance figures. These flowsheets were used to design, construct, and operate bench-scale sys- tems to generate reject waste materials. Reject waste materi- als are being generated and will be characterized to determine chemical and physical parameters important to future environ- mental and economic considerations. The results of the char- acterizations will be published in a separate report. ACKNOWLEDGEMENTS The authors wish to acknowledge Francis C. Brown, of F. C. Brown Associates, Inc., for providing copies of the original flowsheets and for manuscript review. Also the technical staffs of Ocean Minerals Co., Ocean Mining Associates, Ocean Man- agement Inc., and Kennecott Minerals Co. are acknowledged for manuscript reviews. Their assistance has been invaluable in updating the 1977 Dames & Moore and EIC Laboratories report. MANGANESE NODULE PROCESSING OVERVIEW GENERIC TYPES OF PROCESSES In the work performed by Dames & Moore in 1 977 (7), a literature search revealed several methods for recovering the valued metals (Mn, Ni, Co, Cu) from manganese nodules using pyrometallurgical and hydrometallurgical processes, or combinations of the two. Furthermore, the extraction tech- niques can be classified by the type of lixiviant used to solubi- lize the metals of interest. These lixiviant types are ammonia-. chloride-, and sulfate- based. Using this format. Dames & Moore outlines 1 2 potential routes to metals recovery. These routes are Ammoniacal systems 1 . Gas reduction and ammoniacal leach "Italicized numbers in parentheses refer to items in the list of references preceding the appendixes. 2. Cuprion ammoniacal leach 3. High-temperature ammonia leach Chloride systems 1 . Reduction and HCI leach 2. Hydrogen chloride reduction roast and acid leach 3. Segregation roast 4. Molten salt chlorination Sulfate systems 1 . High-temperature and high-pressure H2SO4 leach 2. Smelting and H2SO4 leach 3. H2SO4 reduction leach 4. Reduction roast and H2SO4 leach 5. Sulfation roast Ammoniacal systems are used in processing of land-based nickeliferous laterites which are, in some respects, similar to manganese nodules. It is well known that Cu, Ni, and Co are soluble in ammoniacal ammonium carbonate (Caron process) and ammonium sulfate solutions. These processing routes involve selective reduction of the metals from their oxide states and disruption of the manganese nodule matrix to permit rapid, complete dissolution of the valuable metals. Acid chloride solutions are also capable of solubilizing the metal values of interest including manganese, and a substan- tial body of literature exists describing process conditions for nodule reduction and metals recovery, separation, and purification. Copper, nickel, and cobalt are also soluble in strong acid sulfate systems and serve as the initial step for various possi- ble process routes. The high-temperature H2SO4 leach proc- ess technology is used in recovering nickel from laterites, where the high temperature increases the rates of the dissolu- tion of Cu, Ni, and Co, and limits the solubility of undesirable compounds such as Fe and Mn. Alternative routes involving the acid sulfate lixiviant system include the selective, high- temperature reduction of the nodules, separation of manganese- rich slags from the metallic phases, sulfidation of metallic phases, and subsequent selective leaching of the sulfide materials. A ferromanganese product could also be recovered by further selective reduction of the manganese-rich slag phases. PROCESSES MOST LIKELY FOR FIRST-GENERATION COMMERCIAL USE Of the 12 generic process types presented previously, 7 have sufficient technical problems to preclude the likelihood of commercial development. Flowsheets have been developed for the five process options that are considered as first-generation choices, both from the published literature and by analogy to the processing of land-based ores (7). These five process options are as follows: 1. Gas reduction and ammoniacal leach. 2. Cuprion ammoniacal leach. 3. High-temperature and high-pressure H2SO4 acid leach. 4. Reduction and HCI leach. 5. Smelting and H2SO4 leach. The five processes can be broken down into three-metal and four-metal recovery systems with the three-metal sys- tems having an option to recover manganese from the tailings. The basic three-metal systems are (1) gas reduction and ammoniacal leach process, (2) Cuprion ammoniacal leach process, and (3) high-temperature and high-pressure H2SO4 leach process. The remaining two are considered four-metal The recovery of manganese from the tailings of the three- metal systems involves two basic types of treatment. For the ammoniacal leach tailings, manganese can be recovered to a limited extent by the flotation of MnCOa. The f^/lnCOa could then be further processed to produce a manganese oxide product and sold as such or further processed to produce ferromanganese or other manganese alloys. The residue from the H2SO4 system would require dissolu- tion of the tailings and subsequent chemical manipulation to recover the manganese as Mn02 or another form. The oxide product could be separated and sold or further processed to make ferromanganese or other manganese alloys. The possi- ble direct use of the tailings from either the ammonia-based systems or the H2SO4 system as a feed for ferromanganese production would require some purification of the tailings. Trace metals levels as well as sulfur and possibly phosphorus levels would be too high for direct processing of the tailings by conventional methods. For the purpose of this report, the add-on options to produce manganese products (ferromanganese, silicomanganese, and/or other oxide products) from the three-metal systems will be presented briefly with no process outlines. The proposed descrip- tions serve only as choices from several possible alternatives and, implementation or sehous consideration would be depen- dent on many conditions in addition to technical feasibility. SUMMARY PROCESS DESCRIPTIONS This section contains brief descriptions of each of the five processes as mentioned in the preceding section along with block diagrams for each process. Detailed process deschp- tions of each process are given in the appendixes. GAS REDUCTION AND AMMONIACAL LEACH PROCESS Copper, nickel, and cobalt can be recovered from nodules by a process involving carbon monoxide gas reduction fol- lowed by an ammoniacal leach process. A simplified block diagram for this process is shown in figure 1 . The first step in this process is the high-temperature (625° C) reduction of manganese dioxide (Mn02), the major com- pound in nodules, to manganese oxide (MnO) by a carbon monoxide-rich producer gas. The effect of this reduction is to disrupt the mineral structure and release the contained metals. The metal values solubilized are dissolved from the reduced nodules with a strong aqueous solution of ammonia (10 pet) and carbon dioxide (5 pet), at low temperature (40° C) and atmospheric pressure. The metal-bearing solution is decanted from the nodules and treated with a series of organic extraction steps, which selectively remove the copper and nickel from the aqueous solution. The metal values are selectively stripped from the organic extract with acidified aqueous solutions. The metal products, copper and nickel, are produced from these acidic solutions by electrowinning. Cobalt is then recovered from the aqueous ammonia-carbon •^""l""' carbon dioxide- Nickel f Copper f Cobalt t Electro- winning -| p Electro- winning Grinding and drying Chemical reduction Nickel stripping J L Copper stripping ^ Leach Nickel-copper liquid ion exchange Cobalt recovery Waste containment Liquid-solid separation Metal-bearing solution 1 f Hvdroaen sulfide Ammonia recovery Makeup IMakeup 1 ammonia Figure 1. — Gas reduction and ammoniacal leach process. dioxide solution by contacting it with hydrogen sulfide, which precipitates the insoluble sulfides of Co as well as small amounts of residual Cu, Ni, Zn, and other metals not removed in previ- ous steps. The solids are removed from the aqueous ammonia- carbon dioxide solution and contacted with air and hot (1 00° C) H2SO4 to selectively redissolve the cobalt and the small amount of nickel present. The undissolved sulfides are sold as minor products, and the cobalt and nickel are recovered from solu- tion in powder form by selective reduction with hydrogen at high pressure (34 atm) and temperature (185° C). The nodule residue, from which the major portion (98 pet) of the soluble metals has been removed, is contacted with steam (at 120° 0, 2 atm) to remove residual ammonia and carbon dioxide. The ammonia-carbon dioxide-steam mixture is con- densed and, together with the aqueous ammonia-carbon diox- ide mixture from which cobalt was removed, is recycled to extract more metal values from freshly reduced nodules. The ^eam-stripped nodule residues may be combined with smaller amounts of other process solid and liquid wastes and sent to containment. The high-temperature reduction of nodules with simulated producer gases and subsequent extraction of metals with ammonia-carbon dioxide solutions is basically the same approach £is is currently used in recovering nickel from laterites by the Caron process, and can be considered a variation of currently available technology. The metal separation and puri- fication scheme, however, is specific to nodules and is compli- cated by the chemical similarity of Cu, Ni, and Co. Separation and purification of copper and nickel by selec- tive extraction with organic compounds (liquid ion exchange reagents) is currently practiced in the extractive metallurgy of copper and nickel. However, in these cases the aqueous solutions contain primarily one metal, the others being treated as impurities, not products. Cobalt recovery from precipitated mixtures of nickel and cobalt sulfides derived from laterites is also currently practiced. The details of the procedures used to purify the leach solutions prior to reduction, however, would differ somewhat from those used for nodules because of the differences in amount and content of impurities. Plant services include facilities for generating the producer gas used in nodule reduction, raising the necessary steam and part of the power required for process use, supplying the makeup and cooling water required, and providing for materi- als handling for process materials and supplies. The genera- tion of producer gases from coal or oil for the reduction of nodules and all other plant services represent the utilization of known technology essentially without adaptation. CUPRION AMMONIACAL LEACH PROCESS Copper, nickel, and cobalt can be recovered from nodules by the Cuprion process employing a reducing ammoniacal leach. A simplified block diagram of this process is shown in figure 2. The first step in this process is a low-temperature (50° C) hydrometallurgical reduction of manganese dioxide (MnOa), the major compound in nodules, to manganese oxide (MnO) by an aqueous ammoniacal solution containing an excess of cuprous ions (Cu*). The effect of this reduction is to disrupt the mineral structure and release the contained metals. The metal values are solubilized from the reduced nodules with a strong aqueous solution of ammonia and carbon dioxide at low temperature and pressure. The metal-bearing solution is decanted from the nodules and treated with a series of organic extraction steps that selectively remove the copper and nickel from the aqueous solution. The metal values are in turn selectively stripped from the organic extract with acidified aqueous solutions. The metal products, cathode copper and nickel, are produced from these acidic solutions by electrowinning. Cobalt is then recovered from the aqueous ammonia-carbon dioxide solution by precipitation with hydrogen sulfide, which also precipitates small amounts of residual Cu, Ni, Zn, and other metals not removed in previous steps. The solids are removed from the aqueous ammonia-carbon dioxide solution and contacted with air and hot (100° C) H2SO4 to selectively redissolve the cobalt and small amount of nickel present. The undissolved sulfides are sold as minor products, and the cobalt and nickel are recovered from solution in powder form by selective reduction with hydrogen at high pressure (34 atm) and temperature (185° C). The nodule residue, which has the major portion (98 pet) of the soluble value metals removed, is contacted with steam (at 120° C, 2 atm pressure) to remove residual ammonia and carbon dioxide. The aqueous ammonia-carbon dioxide mixture, from which cobalt was removed, is also steam stripped to recover a high-strength ammonia solution for recycle to the reduction step and to provide fresh wash solution for recycle to extract more metal values from freshly reduced nodules. The steam-stripped nodule residues may be combined with smaller amounts of other process solid and liquid wastes and sent to containment. The hydrometallurgical reduction of nodules in an ammoniacal- ammonium carbonate solution has been disclosed in the patent literature. While this approach differs from the pyrometallurgi- cal reductions used in the well-known Caron process for recovering nickel from laterites, the basic outline for the Cuprion process is similar. The metal separation and purification scheme, however, is specific to nodules and is complicated by the chemical similarity of Cu, Ni, and Co. Nodules Nickel Copper Cobalt t 1 oaroon monoxiae 1 Electro- winning Electro- winning Grinding Chemical reduction Soli ition 4 Nickel stripping ♦ Copper stripping - recycle Leaching- reduced pulp preparation ' Nickel-copper liquid ion exchange Cobalt recovery Waste containment Leaching- liquid-solid separation Metal-bearing solution M t Ammonia recovery MakeuF water ) ^ Make amm( up )n a Figure 2. — Cuprion ammoniacal leach process. Separation and purification of copper and nickel by selec- tive extraction with organic compounds (liquid ion exchange reagents) is currently practiced in the extractive metallurgy of copper and nicl625° C up to 1 ,000° C). The hot, reduced nodules are then charged to an electric furnace, along with the coke and silica. In this step most of the Cu, Ni, Co, and Fe and some of the Mn are reduced (at 1 ,425° C) and form a molten alloy phase, which separates by gravity from the unreduced manga- nese slag. The hot alloy is transferred to converter vessels where, with additional silica, the manganese and most of the iron are re-oxidized with air, separated as a slag, and returned to the electric furnace. Gypsum and coke or possibly sulfur are then added to the alloy, producing a metal sulfide "matte" phase that contains the Cu, Ni, and Co. A second liquid-liquid separa- tion is made in the converter, with the slag returned to the electric furnace and the matte granulated by quenching it in cold water. The electric furnace manganese slag, with recycled iron- rich slags, may be further reduced (at 1 ,480° C) with additional coke in an electric furnace to produce a molten ferromanga- nese alloy, which separates by gravity from the unreduced manganese slag. The ferromanganese is cast for sale, and the waste slag is granulated for disposal. The metals are recovered from the granulated matte by dissolution into strong (5 pet), hot (110° C) H2SO4 solution in the presence of oxygen (at 10 atm pressure). The metal- bearing solution is treated by a series of purification steps in which it is contacted with an organic extractant that selectively removes the copper and nickel from the aqueous solution. Ammonia is added to the solution to control the pH during the separations. The metal values are, in turn, selectively removed from the organic extract and transferred to acidified aqueous solutions, which accumulate copper and nickel sulfates. The metal products, cathode copper and nickel, are produced from these acidic solutions by electrowinning. Cobalt and small amounts of copper, nickel, and other met- als not removed in previous steps are recovered from the aqueous ammonium sulfate solutions by hydrogen sulfide precipitation. The solids are removed from the aqueous ammo- nium sulfate solution and contacted with air and hot (100° C) H2SO4 to selectively redissolve the cobalt and the small amount of nickel present. The undissolved sulfides are sold as minor products, and the cobalt and nickel are recovered from solu- tion in powder form by selective reduction with hydrogen at high pressure (34 atm) and temperature (185° C). Lime is then added to the metal-free ammonium sulfate solution, and the mixture is contacted with steam (at 120° C, 2 atm) to recover ammonia for reuse in the process. The gyp- sum formed in this step is combined with other process solid and liquid wastes and sent to containment. While detailed design information on the process implica- Reducing gas Chemical reduction Grinding and drying Nodules Electric furnace smelting Ferro- manganese reduction ^Ferromanganese Cobalt^ Gypsum gen Copper t Nickel t Oxidizing sulfiding Waste treatment Electro- winning Electro- winning -1 PH adjustment Copper liquid ion exchange Neutralization Nickel liquid ion exchange 1 Leaching- llquid-solid separation 1 1 1 t Ammonia recovery * Cobalt recovery Waste c ontainment k Hydrogen sulfide Figure 5.— Smelting and H2SO4 leach process. tions involved in the smelting of nodules has not been published, enough is known about the thermodynamics of the system to permit a process outline to be constructed. Electric furnace smelting is a well developed technology, and copper and nickel are currently recovered by treatment of mattes formed during the smelting of sulfide ores. Ferromanganese of high purity is currently produced directly from high-quality ores. Thus, the reductive smelting of nodules to ferromanganese with subsequent sulfidizing of the alloy phase to form a matte is a synthesis of technologies from different areas of extractive metallurgy. Detailed information on slag properties (particularly viscosity-composition-temperature relationships); the efficien- cies of materials (coke, gypsum); energy consumption; and the distribution of minor metals and impurities among dust, slag, and matte phases under smelting conditions for this system, however, is lacking. The oxidative dissolution of sulfide ores and mattes is well known, but the metal separation and purification schemes are specific to nodules and are complicated by the chemical simi- larity of Cu, Ni, and Co. Separation and purification of copper and nickel by selective extraction with organic compounds (liquid ion exchange reagents) is currently practiced in the extractive metallurgy of copper and nickel. However, in these cases the aqueous solutions contain primarily one metal, the others being treated as impurities, not products. Cobalt recovery from precipitated mixtures of nickel and cobalt sulfides derived from laterites is also currently practiced. The details of the procedures used to purify the leach solutions prior to reduction will differ somewhat from those used for nodules because of the difference in amount and content of impurities. Plant services include facilities for generating the producer gas used in nodule reduction, generating the necessary steam and part of the power required for process use, supplying the makeup and cooling water required, and providing for materi- als handling for process materials and supplies. The genera- tion of producer gases from coal (or oil) for the reduction of nodules and all other plant services represent the utilization of known technology essentially without adaptation. 10 DETAILED PROCESS DESCRIPTIONS Each of the five processes are presented in detail, in the appendixes, with flowsheets for each unit operation. Many sections of the processes are similiar or identical but flowsheets are presented to preserve the process continuity. Major assump- tions used in calculating operating parameters are presented to show major changes from the Dames & Moore 1 977 report (7). Materials handling for all processes is similar and is presented prior to the detailed process descriptions without flowsheets. Plant services are presented without flowsheets for each of the five processes. MAJOR ASSUMPTIONS Certain assumptions were used in the original Dames & Moore report and are used in updating the process flowsheets in this report. The three three-metal processes and the smelting process are assumed to operate on a 3-million-tpy feed rate (dry basis). The four-metal HCI process is assumed to operate at 1 million tpy (dry basis). The final products are assumed to be Cu and Ni cathodes, some Ni powder, Co powder, small amounts of Cu and Zn sulfides, Mn metal, and ferromanganese. No energy balances or material balances are made in this report. Several significant changes from the Dames & Moore report are detailed here. In the 1977 Dames & Moore study (7), a moisture content of 37.5 pet was used as the water value of nodules fed to the processing plant. This value represents essentially the water content of as-mined nodules. Consider- ing the porosity of nodules (60 pet), it is likely that a substantial amount of water will be removed during transport from the mine site. Ships currently in use equipped with a Marconoflo- type^ dewatering system should be able to dewater the nodules. A recent report described the use of such a system in transport- ing slurried coal (30). A more reasonable value for moisture content of nodules received at the port facilities may be 1 5 to 20 pet. This fact would decrease the size of several components of the materi- als handling section, and lower the energy cost for drying the nodules. For this report a moisture content of 20 pet is used for nodules fed to the plant. This lowers the 3-million-tpy (dry- basis) plant feed rate from 12,500 to 10,900 tpd, and the 1-million-tpy (dry-basis) plant feed rate from 3,750 to 3,640 tpd. A second major change from the 1977 study (7) is the increase in size of the smelting plant from 1 million tpy (dry basis) to 3 million tpy (dry-basis) to allow use of conventionally sized furnaces. Also, particle size of the nodule feed is increased to minus 65 mesh instead of minus 200 or minus 325 mesh as previously used. The previous study contained an extensive bibliography and it will not be duplicated in this report. However, certain pertinent publications since 1 977 are included in the reference section of this report (7-6, 8-29, 31-32). Certain publications apply directly to one or two processes, while others are more general in nature. Updated details of the Cuprion process (1-3, 20, 26) and modifications of the Caron process (5, 9-12, 24-25, 32) have been published. Several articles on different aspects of H2SO4 leaching of laterites and nodules (6, 8, 10, 18, 23) are available as well as those dealing with smelting processes (13, 21-22, 29, 31-32). Several review articles have also been published (4, 14, 17, 27-28). MATERIALS HANDLING Facilities are provided within the plant for receiving and reclaiming raw nodules, coal, lime and limestone, ammonia, other process materials, and fuel. Provisions for handling hydrogen chloride and chlorine are also required for the reduc- tion and HCI process, and provisions for handling silica, gypsum, and acids are made for the smelting and H2SO4 process. In the 1977 Dames & Moore study (7), the proposed method of nodule transport from the port facilities was a slurry pipeline. Assuming nodules can be dewatered to 1 5 to 20 pet moisture, the use of slurry pipelines may not be cost effective. At this moisture content, the use of conveyors for transport appears more likely. Conveyors are routinely used in mining industries to transport materials over both long and short distances in many areas. The use of conveyors for nodule transport would involve known technology with little adaptation. Because all processes (except Cuprion) require the water associated with nodules to be completely removed, the use of dewatering during ship transport, coupled with conveyor transport, should result in an overall net energy savings. SUMMARY This report is an update of the 1 977 study (7) by Dames & Moore and EIC Corporation entitled, "Description of Manga- nese Nodule Processing Activities for Enviromental Studies," but does not include the energy and material balances con- tained in the initial study. The flowsheets presented in the appendixes of this report are adaptations of those in the previ- ous report with the changes since 1977. The five manganese nodule processing options outlined in this report are gas reduc- tion and ammoniacal leach process, Cuprion ammonical leach process, high-temperature and high-pressure H2SO4 leach process, reduction and HCI leach process, and smelting and ^Reference to specific trade names or equipment does not imply endorse- ment by the Bureau of Mines. H2SO4 leach process. The first three processes are designed to recover three metals (Cu, Co, and Ni), and the latter two are designed to recover four metals (Cu, Co, Ni, and Mn). The three-metal processes have the option of recovering manganese from the tailings if economically feasible. This report differs from the 1 977 study (7) by using a larger mesh size of minus 65 for the feed material for all processes except smelting (where minimal size reduction is required); by using a moisture content of 20 pet rather than 37.5 pet for the feed nodules; by upgrading the smelting process to a 3-million- tpy nodule feed rate (dry basis) to allow for more convention- ally sized furnaces; and by recognizing the use of dewatering of the nodules during ship transport, coupled with conveyors instead of slurry pipeline transport of nodules from the port to the plant. A summary table of operating parameters is presented for each of tfie five processes and each process is brol MnO + COz FejOs + CO — > 2FeO + CO2 NiO + CO— > Ni + CO2 CuO + CO— >Cu + CO2 CoO + CO— > Co + CO2 Conversion, pet: Co Cu Fe Mn Ni LEACHING-AERATION (FIG. A-4) Number of stages Temperature Time per stage Leach solution composition, gpl NH3 CO2 Pressure Solubilization of metals, pet Co Cu Fe Mn Mo Ni Zn SOLID-LIQUID SEPARATION (FIG A-5) Underflow density Wash ratio Wash recovery Wash liquor compMSsition, gpl pet solids kg/kg liquor pet CO, LIQUID ION EXCHANGE-EXTRACTION (FIG. A-6) Extraction: Extractant Number of stages Organic-aqueous ratio.. Metals extraction, pet: Co Cu Ni Zn Washing (primary): Washing agent pet NHa.. Number of stages Organic-aqueous ratio 3 in organic gpl.. LIQUID ION EXCHANGE-STRIPPING (FIG. A-( Washing (secondary): Wash composition, gpl: (NH4)2S04 H2SO4 10,900 -65 150 18.2 7.8 10.9 49 10.1 3.6 825 625 95 125 100 100 98 100 100 50 1 35-40 2 100 50 'LIX64N 3 2:1 99.9 99.9 10 Washing (secondary) — Con. Number of stages Organic-aqueous ratio Residual NH3 in organic Ni stripping Stnp solution composition, gpl Cu HzSO* Ni Zn Number of stages Organic-aqueous ratio Stripping, pet Co Cu Ni Zn Cu stripping: Strip solution composition, gpl: Cu H2SO4 Ni (max) Zn Number of stages Organic-aqueous ratio Stnpping, pet Co Cu Ni Zn Temperature gpl 2 1:1 0.01 < 0.001 40 50 «0 3 5:1 0.3 <.004 98.8 «0 40 160 10 5 2 3:1 0.2 87 «0 100 40 COPPER ELECTROWINNING— COMMERCIAL (FIG. A-! Current density Current efficiency Temperature Cu in-out H2SO4 in-out A/m^. pet. °C. gpl gpl NICKEL ELECTROWINNING— COMMERCIAL (FIG Current density Current efficiency Temperature Ni in-out. H2SO4 in-out Na2S04.. H3BO3.... A/m^. pet, "C, gpl gpl gpl 53-40 160-180 180 93 60 75-50 0.016-40 100 15 COBALT RECOVERY (FIG A-11) Precipitating agent Precipitation, pet: Co Cu Ni Zn Temperature Clarifier density Wash ratio Co leaching slurry Leaching agent Evaporation-crystallization water removal Co oxidation Temperature .pet NH4HS.. Co reduction Temperature Pressure Reductant "C pet solids pet pet H2SO4 pet °C psig °C psig 40 70 70 100 150 175 500 H, AMMONIA RECOVERY (FIG A-1 2) CO2 Temperature Pressure Efficiency for CO2 Number of stages °C atm pet ^Reference to specific trade names does not imply endorsement by the Bureau of Mines. 14 Table A-1 . — Operating parameters for gas reduction and ammoniacal leach process — Continued Parameter and unit Value Parameter and unit Value AMMONIA RECOVERY (FIG. A-1 2)— Con. AMMONIA RECOVERY (FIG. A-1 2)— Con. NH3 absorber: Temperature °C.. Pressure atm.. Number of stages 35 1.2 1 NHg stripper: Pressure atm.. Recovery of NH3 pet.. 1.5 99 2 ment to sale. The greater portion of the wealt electrolyte is recycled to the LIX section for stripping. A small amount of nickel is stripped along with the copper not deposited in electrowinning, and must be purged from the system. The purged electrolyte passes through purification cells where copper is removed by electrowinning to depletion. The decopperized electrolyte passes to vacuum evaporators where water is removed and nickel sulfate is precipitated from the resulting highly acidic solution. The nickel sulfate is removed and sent to cobalt recovery where it is combined with other purge streams. The acid is retumed to the process where, with makeup acid, it is used to dissolve scrap copper for return to the commercial cells for deposition. Sufficient steam, wash water, and makeup water are added to the circuit to offset water vaporized or carried off with evolved oxygen during electrowinning. NICKEL ELECTROWINNING— COMMERICAL (FIG. A-10) Nickel is recovered from the strong electrolyte by electrowin- ning in a manner similar to that used for copper recovery. In nickel electrowinning, however, cathode bags are used, and sodium sulfate and boric acid are added to the electrolyte to control its conductivity and pH. Dissolved organic carried from the LIX step is removed by adsorption on activated carbon prior to any electrolyte passing to electrowinning. Also, the starter sheets are pickled in H2SO4 prior to use in the commer- cial cells, and nickel scrap is dissolved in ammonia-containing raffinate and recycled to the pregnant liquor, rather than to the recycled or makeup ackJ solutk>n. The electrolyte purge required to remove impurities from the electrowinning circuit passes to a sulfide precipitation reactor and then to cobalt recovery for recovery of metal values. COBALT RECOVERY (FIG. A-1 1) Along with unextracted Cu, Ni, and Zn, Co is recovered from the LIX raffinate by precipitation with ammonium hydrosulfide (produced by sparging hydrogen sulfide into an excess of ammonia solution). The sulfide precipitate is separated from the raffinate in a clarifier, with the clarifier overflow being recycled for tailings washing and the underflow passing to stripping for ammonia recovery. The sulfide slurry is mixed with electrolyte purges from Cu and Ni electrowinning, and with Co recovered from stripping the LIX reagent. The mixture is pressure leached with air to preferentially dissolve the Ni and Co sulfides, leaving the Cu and Zn sulfides in the residues. The latter are removed by filtration and sold as minor products to smelters for recovery of metal values. Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any Zn and Cu solubilized in the first leach, the Ni-Co sulfate solution is heated and autoclaved, and Ni is reduced with hydrogen. Sufficient ammonia solution is added during reduction to neutralize the acid formed. Only a portkxi of the nk^el is removed per pass to prevent oven-eductkxi and subsequent contamination of the nickel powder with cobalt. After densification through repeated recycle, the nickel pow- der is removed, washed, and passed to drying and briquetting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator-crystallizer where the remaining nickel and cobalt are precipitated as the double salts with ammonium sulfate. Excess ammonium sulfate is purged, and the salts are redis- solved in strong ammonia solution. The cobalt in solution is oxidized to the cobaltic state (Co^^) with air. This permits the cobalt to remain in solution when the stream is acidified to remove the nickel salts, which are separated and recycled to the pH adjustment step. The nickel-free solution is then heated and autoclaved for removal of cobalt by hydrogen reduction. Sufficient ammonia is added to neutralize the acid generated. The cobalt powder is dried and briquetted for sale, while the ammonium sulfate is purged to the ammonia recovery process. AMMONIA RECOVERY (FIG. A-1 2) Tailings from the CCD wash circuit are preheated and stripped for ammonia recovery by countercurrent contact with steam in stripping towers. The steam and ammonia vapors from strip- ping are combined with vapors from other ammonia strippers and pass to an ammonia absorber-condenser tower where nx)st of the ammonia and amnronium carbonate are condensed. Vent gases pass to a cart}on dioxide absorber where they are absorbed, along with the required makeup carbon dioxide obtained from boiler offgases, in makeup scrubbing water. Ammonia is recovered from all process vents by countercur- rent contact with makeup scrubbing water in an ammonia absorber and vent scrubber. The ammonia-free gases pass to stacks for disposal. Ammonia is recovered from the ammonium sulfate purges from cobalt recovery and the LIX washing step by reaction with slaked lime. Steam is blown into the mixture to strip the evolved ammonia, and the vapor is condensed. The condensate is retumed to aqueous ammonia storage for recycle within the process. The gypsum slurry from the lime boil is cooled, along with the slurry from the tailings stripping, and passed to tailings surge. This is combined with process solid and liquid wastes and plant runoff, treated for pH control, as required, and pumped to the tailings impoundment area for disposal. 15 PLANT SERVICES PROCESS ALTERNATIVES Plant services include process and cooling water supply and treatment, steam raising and power generation, gas treatment, and reducing gas preparation. Makeup water is clarified and softened for distribution to the process, as required. Additional treatment is required for cooling tower water makeup, boiler feed water makeup, and for supplying plant potable water. Offgases from manganese reduction are burned, along with additional coal, in the main boilers to raise the required process steam and generate a portion of the power required in the process. Following particulate removal, the flue gases are combined with other process offgases and pass to gas treatment, where sulfur oxides and other acidic constituents are removed by scrubbing with limestone. The scrubbed offgases are reheated, combined with scrubbed vents from ammonia recovery, and are passed to stacks for disposal. Gas for nodule reduction is produced in a two-stage entrained-flow gasifier in which coal is mixed with preheated air and high- temperature steam for the production of a cartx)n monoxide- rich reducing gas. The gasifier product passes to reduction following particulate removal and energy recovery, with sulfur removal taking place after combustion and with other boiler flue gases. MANGANESE PRODUCTION ADD-ON The recovery of manganese from ammoniacal leach tails is an option that has been investigated. Manganese may be partially recovered from tailings by flotation of the fine MnCOa present. The MnCOa could then be converted to an oxide form for feed to produce ferromanganese or other manganese alloys. The recovery of manganese from the process tails would be contingent on the market conditions and process economics. Some intermediary product may have market value, thereby avoiding energy-intensive steps to produce manga- nese alloys. Because of the relatively new technology required for manganese recovery from ammoniacal leach tails, and the dependency upon economics and technical conditions, no detailed flowsheets are presented for this process. A major processing alternative, involving the hydrometallur- gical reduction-ammoniacal leaching of nodules is fully docu- mented in appendix B. Other than the metals separation and purification steps, the process configuration is based on the well-known Caron process, and there is little reason to believe, in the absence of actual operating data, that the fundamental approach should differ. It has been assumed that the reduced nodules can be carbonated in a fluid bed cooler prior to leaching. If carbona- tion is not complete, the nodules would absorb carbon dioxide directly from the leach liquor. Then, to maintain the required pH, the wash liquor would have to be recycled with a higher ammonium carbonate content. This, in turn, would be obtained by scrubbing a much larger fraction of the boiler flue gases in the ammonia recovery section. The use of fluid bed dryers and reactors represents an advance in processing technology over the use of rotary kilns and multiple hearth furnaces and should not drastically affect any material balance. It has been assumed that the leached tails can be stripped for ammonia recovery in the same way as is done in laterite processing, although unexpected fouling tendencies could require the use of less efficient stripping devices, thereby increasing steam (fuel) consumption. The ammonium sulfate solutions purged from the liquid ion exchange-extraction and cobalt recovery sections have been treated with lime for the recovery of ammonia for recycle. An alternative approach would involve direct recovery of ammo- nium sulfate by evaporation-crystallization, for sale as a byproduct. Alternative configurations of the metals separation steps are possible, such as selective extraction-selective stripping, but their impact on overall plant material and energy balances should be minor. Many variations are possible in the details of the scheme used for recovery of cobalt from a mixed sulfide precipitate. Ultimately, however, their impact on plant require- ments would not differ greatly from the approach used here, because the sulfides will still be oxidized to sulfates, purged, and hydrogen reduced. It does not appear likely that solutions could be purified easily enough to permit recovery of electro- lytic cobalt. 16 ?? 3 0> (0 CO I I a < o -o o 11 o 8 Pv o n, rN>_ O \> 17 J? 19 ?? 20 21 22 "QD ! II 23 '^-^5:^ 24 25 27 / I / i i \ I n APPENDIX B.— CUPRION AMMONIACAL LEACH PROCESS The Cuprion ammoniacal leach process is a three-metal process in which Cu, Ni, and Co are liberated in an ammoniacal- ammonium carbonate leach following a reduction-leach step. Carbon monoxide is used to regenerate the cuprous ion which reduces manganese dioxide. Copper and nickel are co-extracted by liquid ion exchange (LIX) reagents and are selectively stripped and recovered as electrowon cathodes. Cobalt is removed from the raffinate by precipitation with hydrogen sulfide and is recovered from the sulfide precipitate, along with some Ni, Zn, and Cu, by selective leaching and hydrogen reduction. The metal-free raffinate is steam stripped to recover a high-strength ammonia solution for recycle to the tailings wash step, together with ammonia and ammonium carbonate recov- ered by steam stripping the leach tailings. Detailed descriptions of each segment of the process are given for the following flowsheets. A summary of parameters for each section is given in table B-1 . A key to the flowsheet symbols is given in figure B-1 . ORE PREPARATION (FIG. B-2) Wet nodules are reclaimed from storage and fed through a primary cage mill where they are reduced to minus % in. They then pass to a rod mill for wet grinding to minus 65 mesh. Oversized nodules are separated and returned via the rake classifier, hydrocyclone, water-recycle circuit. The ground nod- ule slurry passing through the cyclone is held in an air-agitated surge tank to prevent settling before being pumped to the reduction step. REDUCTION-LEACH (FIG. B-3) The nodule slurry is mixed with recycled strong ammonia solution and fed, with a recycled solution containing an excess of cuprous ion (Cu*), to a train of agitated reduction reactors. The dilute slurry is contacted with a carbon monoxide-rich gas derived from gasified coal, the manganese dioxide is reduced and converted to manganese carbonate, and most of the value metals are solubilized. The heat of reaction is removed in shell and tube exchangers to maintain the reduction temper- ature at 50° C, and excess gases pass to ammonia recovery. The reduced, dilute nodule slurry is thickened with the over- flow recycle, and the thickened pulp passes to the oxidation leach. OXIDATION-LEACH (FIG. B-4) The reduced nodule slurry passes to a countercurrent oxidiz- ing leach in which air is sparged into the leach slurry to oxidize residual Cu * to Cu^ \ Co^ * to Co^ ' , and Fe^ * to Fe^ * , precipi- tating the Fe as an insoluble ferric hydroxide in the manga- nese carbonate tailings. Liquid-solid separations are made in thickeners, which also provide residence time for leaching. Offgases from aeration pass to ammonia recovery, leached nodules to washing, and pregnant liquor to LIX for metal separation and purification. SOLID-LIQUID SEPARATION (FIG. B-5) Metal values that have been solubilized in leaching and absorbed on tailings are recovered from the tailings by wash- ing in a countercurrent decantation (CCD) circuit using cov- ered thickeners. Wash liquor consists of recovered ammonia and ammonium carbonate along with raffinate from cobalt recovery. Washed tailings pass to stripping for ammonia and ammonium carbonate recovery while the wash overflow passes to leaching. LIQUID ION EXCHANGE- EXTRACTION (FIG. B-6) Pregnant liquor from leaching is filtered and passes to a three-stage, countercurrent LIX extraction circuit where copper, nickel, and some ammonia are removed from the pregnant liquor. Most of the ammonia is removed from the organic phase by washing with a weak aqueous ammonia solution. This ammonia is recovered by steam stripping a portion of the scrub solution, with the vapors passing to the ammonia recov- ery section. Provision is made for periodically cleaning the mixer-settler units used in extraction and stripping and for recovering organic and aqueous phases for recycle. Degraded organic, dust, and other forms of "crud" are removed from the organic and incinerated. Because a small amount of Co is co-extracted and is not stripped with Cu and Ni, it must be removed from the LIX reagent to prevent its buildup. This is accomplished by precipitating the cobalt, with hydrogen sulfide, from a purge stream of organic. The precipitated solids are washed from the organic and pass to cobalt recovery, while the purified organic is returned to the extraction loop (fig. B-7). LIQUID ION EXCHANGE- STRIPPING (FIG. B-8) The remainder of the ammonia is removed from the loaded organic by washing with a slightly acidic ammonium sulfate solution. The ammonium sulfate formed in scrubbing is purged to ammonia recovery. Nickel is then selectively stripped from the organic by countercurrent contact at controlled pH with depleted electrolyte from nickel electrowinning. Because elec- trowinning conventionally occurs at a higher temperature than operation of the LIX circuit, the strip solution is heated passing to electrowinning and cooled passing from electrowinning. Copper is then removed from the nickel-free organic by coun- tercurrent stripping with depleted electrolyte from copper electrowinning. The copper electrolyte is also heated-cooled on passage to-from electrowinning to permit operation of elec- trowinning at a higher temperature. Makeup organic is added to the stripped reagent to offset degradation and soluble organic losses to the pregnant liquor, and the metal-free organic is recycled to extraction. COPPER ELECTROWINNING— COMMERCIAL (FIG. B-9) Cathode copper is recovered from the LIX strong electrolyte using conventional technology. Starter sheets are deposited 29 Table B-1.— Operating parameters for Cuprion ammoniacal leach process Parameter and unit Value Parameter and unit Value ORE PREPARATION (FIG. B-2) LIQUID ION EXCHANGE-STRIPPING (FIG. B-8) Feed size to reduction ....mesh.. —65 WdsnjnQ (SGConoflry)' REDUCTION-LEACH (FIG. B-3) H2S04. 1 Reductant gas composition, pet CO 40-60 Ha 3^45 H2O 6-12 N2 1 Temperature ° C 50 Pressure atm. 1 Mn02 + CO— >MnC03 FejOa + CO-> 2FeO + COj NiO + CO->Ni + C02 CuO + CO— >Cu + CO2 CoO + CO->Co + C02 NH3 100 Organic-aqueous ratio 1 :1 Residual NH3 in organic gpl 0.01 Ni stripping: Strip solution composition, gpl Cu <0.001 H2S04. 40. Ni 50 Zn so Number of stages 3 Organic-aqueous ratio 5:1 Stripping, pet Co 0.3 Cu <.004 Ni 98.8 Zn *0 CO2 25 Cu 3-5 Number of stages 6 Reduction, pet Co 50 Cu 80 Mn 97 Mo 80 Ni 90 Zn 40 Slurry density pet solids 20-30 Cu stnpping Stnp solution composition, gpl: Cu 40 H2SO4 160 Ni (max) 10 Zn 5 Number of stages 2 Organic-aqueous ratio 3:1 Stnpping, pet Co 0.2 Cu 87 OXIDATION-LEACH (FIG. B-4) Ni so Pregnant liquor composition, gpl: NHa 100 Zn 100 Temperature ° C 40 COi 25 COPPER ELECTROWINNING— COMMERCIAL (FIG B-9) Cu 4-8 Ni 5-10 Temperature " C 50 Pressure atm 1 Recovery, pet Co 50 Current density A/m^ 180 Cun-ent efficiency pet 94 Temperature ° C 50 Cu in-out gpl 53-40 H2S04in-out gpl 160-180 Cu 90 NICKEL ELECTROWINNING-COMMERCIAL (FIG B-10) Ni 90 Cun-ent density A/m^ 180 Current effiaency pet 93 SOLID-LIQUID SEPARATION (FIG B-5) Wash ratio kg/kg liquor 2 Wash recovery pet 98 Number of stages 6 Wash liquor composition, gpl NHa 100 Temperature ° C 60 Ni in-out gpl 75-50 H2S04in-out gpl 0.016-40 Na2S04 gpl 100 HaBOa gpl 15 COBALT RECOVERY (FIG. B-1 1 ) CO2 50 Precipitating agent pet NH4HS 30 LIQUID ION EXCHANGE-EXTRACTION (FIG B-6) Predpitation. pet: Exti action Extractant 'LIX 64N Number of stages 3 Organic-aqueous ratio 2:1 Metals extraction, pet Co 1 Cu 99.9 Ni 99.9 Zn 10 Washing (primary) Washing agent pet NH, 1 Number of stages 2 Organic-aqueous ratw 3:1 Residual NH, in organic gpl 0.1 Co 98 Cu 99.9 Ni 99 Zn 99.9 Temperature ° C 80 Clanfier density pet solids 5 Wash ratio 2:1 Co leaching sluny pet 40 Leaching agent pet H2SO4 70 Evaporation-crystallization water removal pet 70 Co oxidation Temperature "C 100 Pressure psig 150 Co reduction Temperature °C 175 Pressure psig 500 Reductant Hj 'Reference to specific tradenames does not impty endorsement by the Bureau of Mines. 30 Table B-1 .—Operating parameters for Cuprion ammoniacai leach process— Continued Parameter and unit Value Parameter and unit Value AMMONIA RECOVERY (FIG. B-1 2) AMMONIA RECOVERY (FIG. B-1 2)— Con. CO2 absorber: Temperature Pressure EfficierKy for CO2 Number of stages NH3 absorber: Temperature ..°C.. atm.. ..pet.. 40 1.2 99 1 35 NH3 absorber— Con. Pressure Number of stages. NH3 stripper: Pressure Recovery of NH3.. Number of stages. atm.. atm.. pet.. 1.2 1 1 5 ..°C.. 99 2 on titanium blanks from strong electrolyte and are removed, washed, looped, and returned to the commercial section as starters. Full-term cathodes produced in the commercial sec- tion are washed, unloaded, and prepared for shipment to sale. The greater portion of the weak electrolyte is recycled to the LIX section for stripping. A small amount of nickel is stripped along with the copper not deposited in electrowinning, and must be purged from the system. The purged electrolyte passes through purification cells where copper is removed by electro- winning to depletion. The decopperized electrolyte passes to vacuum evaporators where water is removed and nickel sul- fate is precipitated from the resulting highly acidic solution. The nickel sulfate is removed and sent to cobalt recovery where it is combined with other purge streams. The acid is returned to the process where, with makeup acid, it is used to redissolve scrap copper for return to the commercial cells for deposition. Provisions are made for recycling copper, in ammo- nium carbonate solution, to reduction if required. Sufficient steam, wash water, and makeup water are added to the circuit to offset water vaporized or carried off with evolved oxygen during electrowinning. NICKEL ELECTROWINNING— COMMERCIAL (FIG. B-10) Nickel is recovered from the strong electrolyte by electrowin- ning in a manner similar to that used for copper recovery. In nickel electrowinning, however, cathode bags are used, and sodium sulfate and boric acid are added to the electrolyte to control its conductivity and pH. Dissolved organic carried from the LIX step is removed by adsorption on activitated carbon before the nickel electrolyte passes to electrowinning. Also, the starter sheets are pickled in H2SO4 prior to use in the commercial cells, and nickel scrap is redissolved in ammonia- containing raffinate and recycled to the pregnant liquor, rather than to the recycled or makeup acid solution. The electrolyte purge required to remove impurities from the electrowinning circuit passes to a sulfide precipitation reactor and then to cobalt recovery for recovery of metal values. COBALT RECOVERY (FIG. B-11) Along with unextracted Cu, Ni, and Zn, Co is recovered from the LIX raffinate by precipitation with a slight excess of ammo- nium hydrosulfide (produced by sparging hydrogen sulfide into an excess of ammonia solution). The sulfide precipitate is separated from the raffinate in a clarifier, with the clarifier overflow being recycled to ammonia recovery and the under- flow passing to stripping for ammonia recovery. The sulfide slurry is mixed with electrolyte purges from Cu and Ni electrowinning, and with the Co recovered from strip- ping the LIX reagent. The mixture is pressure-leached with air to preferentially dissolve the Ni and Co sulfides, leaving the Cu and Zn sulfides in the residues. The latter are removed by filtration and sold as minor products to smelters for recovery of metal values. Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any Zn and Cu solubilized in the first leach, the Ni-Co sulfate solution is heated and autoclaved, and Ni is reduced with hydrogen. Sufficient ammonia solution is added during reduction to neutralize the acid formed. Only a portion of the nickel is removed per pass to prevent overreduction and subsequent contamination of the nickel powder with cobalt. After densification through repeated recycle, the nickel pow- der is removed, washed, and passed to drying and briquetting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator-crystallizer where the remaining nickel and cobalt are precipitated as the double salts with ammonium sulfate. Excess ammonium sulfate is purged, and the salts are redis- solved in strong ammonia solution. The cobalt in solution is oxidized to the cobaltic state (Co^^) with air. This permits the cobalt to remain in solution when the stream is acidified to remove the nickel salts, which are separated and recycled to the pH adjustment step. The nickel-free solution is then heated and autoclaved for removal of cobalt by hydrogen reduction. Sufficient ammonia is added to neutralize the acid generated. The cobalt powder is dried and briquetted for sale, while the ammonium sulfate is purged to the ammonia recovery process. AMMONIA RECOVERY (FIG. B-1 2) Tailings from the CCD wash circuit are preheated and stripped for ammonia and carbon dioxide recovery by countercurrent contact with steam in stripping towers. The steam and ammonia- carbon dioxide vapors from tailings stripping are combined with vapors from cobalt sulfide slurry stripping, condensed, and combined with ammonia and carbon dioxide scrubbed from process vent streams. Makeup carbon dioxide is obtained from boiler offgases by countercurrent contact with scrubbing water; the ammonia-free gases pass to stacks for disposal. The ammonia-rich ammonium carbonate solution required in reduction is obtained by stripping only a portion of the ammo- nia contained in the raffinate stream from cobalt recovery. These ammonia-rich vapors are combined with ammonium- steam vapors from the LIX ammonia recovery and lime boil steps, makeup ammonia, and ammonia-rich vapors from reduc- tion and oxidation vents, and are condensed and recycled to reduction. The raffinate is then stripped further, with the ammo- nium carbonate recycled to tailings washing and the stripper bottoms used for process water makeup. Ammonia is recovered from the ammonium sulfate purges from cobalt recovery and the LIX washing step by reaction with slaked lime. Steam is blown into the mixture to strip the evolved ammonia, and the vapor is condensed, with the condensate returned for recycle within the process. The gypsum slurry from the lime boil is cooled, along with the slurry from the tailings stripping, and passed to tailings surge. This is com- bined with the process solid and liquid wastes and plant runoff, treated for pH control as required, and pumped to the tailings impoundment area for disposal. PLANT SERVICES Plant services include process and cooling water supply and treatment, steam raising and power generation, gas treatment, and reducing gas preparation. Makeup water is clarified and softened for distribution to the process as required. Additional treatment is required for cooling tower water makeup, boiler feed water makeup, and for supplying plant potable water. The carbon monoxide required for manganese reduc- tion is produced in a low-pressure gasifier in which coal is mixed with oxygen and steam and partially combusted. Acid gases are removed from the reducing gas, following particu- late removal, and are sent to the main boilers to oxidize reduced components for subsequent recovery. High-strength carbon monoxide is then removed from the sweetened gases for use in reduction, and the remaining gases, mainly hydrogen, are sent to the main boilers except for the amount required in cobalt reduction. These and other combustible process offgases are burned, along with additional coal, in the main boilers to raise the required process steam and to generate a portion of the power required in the process. Following particulate removal, the flue gases are combined with other process offgases and pass to gas treatment, where sulfur oxides and other constituents are removed by scrubbing with limestone. The scrubbed offgases are reheated, combined with scrubbed vents from ammonia recovery, and are passed to stacks for disposal. MANGANESE PRODUCTION ADD-ON The recovery of manganese from ammoniacal leach tails is an option that has been investigated. Manganese may be partially recovered from tailings by flotation of the fine MnCOa present. The MnCOa could then be converted to an oxide form for feed to produce ferromanganese or other manganese alloys. The recovery of manganese from the process tails would be contingent on the market conditions and process economics. Some intermediary product may have market value, thereby avoiding energy-intensive steps to produce manga- nese alloys. Because of the relatively new technology required for manganese recovery from ammoniacal leach tails, and the dependency upon economic and technical conditions, no detailed flowsheets are presented for this process. PROCESS ALTERNATIVES A major process alternative, involving the pyrometallurgical reduction-ammoniacal leaching of nodules, is fully documented in appendix A. While the pyrometallurgical reduction-ammoniacal leaching process is closely based on the well-known Caron process, the Cuprion process differs in two important respects: the metals separation scheme is specific to nodules processing and the dififering ammonium carbonate concentrations required in reduction and tailings washing complicate the ammonia recovery step. Alternative configurations of the metals separation steps are possible, such as selective extraction-selective stripping, but their impact on overall plant material and energy balances should be minor. Many variations are possible in the details of the scheme used for recovery of cobalt from a mixed sulfide precipitate. Ultimately, however, their impact on plant require- ments would not differ greatly from the approach used here, because the sulfides will still be oxidized to sulfates, purged, and hydrogen reduced. It does not appear likely that solutions could be purified easily enough to permit recovery of electro- lytic cobalt. Tailings stripping for ammonia and carbon dioxide recovery has been based on the Caron practice, but the raffinate strip- ping operation is unique to this process. No operating data are available to support the energy-intensive concept used. It is possible that other techniques, such as vacuum or pressure stripping, pH adjustment prior to stripping, extractive distillation, etc., may be more advantageous if other process constraints, such as the overall water balance, can also be met. The ammonium sulfate solutions purged from the liquid ion exchange-extraction and cobalt recovery sections have been treated with lime for the recovery of ammonia for recycle. An alternative approach would involve direct recovery of ammo- nium sulfate by evaporation-crystallization, for sale as a byproduct. It has been assumed that, except for the cobalt reduction requirements, byproduct hydrogen from carbon monoxide pro- duction would be burned for its fuel value, which is a relatively poor use. A small part (1 pet) of the hydrogen could be burned with 4,000-tpy sulfur to produce the required amount of hydro- gen sulfide. 32 2 5) 34 35 \ o ^ s li! .s i" 2 ^ • ? II 36 37 o o 40 42 43 ■f r— / / p z / - i 1 '\ E s o m c o > > 1 44 APPENDIX C— HIGH-TEMPERATURE AND HIGH-PRESSURE H2SO4 LEACH PROCESS The high-temperature and high-pressure H2SO4 leach proc- ess is a three-metal process in which Cu, Ni, and Co are selectively leached from the nodules by strong H2SO4 at high temperature and high pressure. After separation of the leach- ing residue and metalliferous solution by washing, the copper and nicl N1SO4 + H2O (NH4)2S04 200 CuO + H2SO4 — > CUSO4 + H2O H2SO4. 1 CoO + H2SO4 — > C0SO4 + H2O Number of stages 2 Mn02 + H2SO4— > Mn SO4 +H2O + 0.5 O2 Organic-aqueous ratio 1:1 FezOa + 3H2SO4— > Fe2(S04)3 +3H2O Residual NH3 in organic gpl 0.01 SOLID-LIQUID SEPARATION (FIG. C-4) | Temperature Nickel stripping °C 40 6 Strip solution composition, gpl: Cu Efficiency pet.. 98 < 0.001 Underflow density pet solids 35-40 H2S04 40 Wash ratio kg/kg liquor.. 2 Ni 50 PREGNANT LIQUOR pH ADJUSTtWiENT (FIG. C-5) j Number of stages Organic-aqueous ratio 3 5:1 Adjustment agent CaCOs Stnpping, pet Co H2SO4 concentration, final gpl.. 0.5 0.3 Entrained solids ppm.. =100 Cu Ni <.0O4 COPPER LIQUID ION EXCHANGE (FIG. C-6) | 98.8 Extraction: LIX64N NICKEL ELECTROWINNING— COMMERCIAL (FIG C-12) Extractant Current density A/m^ 180 Number of stages 3 Current efficiency pet 93 Organic-aqueous ratio 1:1 Temperature °C 60 Metals extraction, pet Ni in-out gpl 75-50 Co 0.1 H2SO4 in-out gpl 0.016-40 Cu 99.5 Na2S04 gpl 100 Fe .1 .1 .1 1 H3BO3 gpl 15 Ni COBALT RECOVERY (FIG .c-1 3) Zn Precipitating agent pet NH4HS.. 30 Ammonia wash pet NH3 25 Precipitation, pet: Residual NH3 in organic gpl 1 Co 98 Stripping Cu 99.9 Strip solution concentration, gpl: Ni 99 H2SO4 160 Zn 99.9 Ni (max) 10 Temperature "C 80 Cu 40 Clanfier density pet solids 5 Zn 5 Wash ratio 2:1 Number of stages 2 Co leaching slurry pet 40 Organic-aqueous ratio 3:1 Leaching agent pet H2SO4 70 Temperature ° C 40 Evaporation-crystallization water removal pet 70 Stripping, pet Co oxidation Co 0.2 Temperature "C 100 Cu 87 Pressure psig 150 Ni .9 Co reduction Zn 100 Temperature Pressure Reductant psig 175 COPPER ELECTROWINNING— COMMERCIAL (FIG C-8) 500 Current density A/m^ Current efficiency pet 180 94 50 AMMONIA RECOVERY (FIG. C-1 4) Temperature ° C NH3 absorber: Cu in-out gpl 53-40 Temperature "C. 35 H2S04 in-out gpl 160-180 Pressure Number of stages NH3 stripper: Pressure atm.. 1.2 COPPER RAFFINATE pH ADJUSTMENT (FIG. C-9) 1 Number of stages 4 atm.. 1.5 Recovery of NH3 Number of stages pet.. Final pH =4.0 2 'Reference to specifictrade names does not imply endorsementby the Bureau of Mines. 46 extraction in three stages of pH-controlled countercurrent extraction. The nickel-depleted aqueous phase, containing primarily cobalt, proceeds to cobalt recovery The entrained aqueous, containing dissolved ammonia, extracted nickel, and trace metals, is removed from the organic phase in a liquid separation step. The ammonia removal requires two stages of recovery. The organic is countercurrently washed in two stages, and then the aqueous wash phase is steam stripped of ammonia. NICKEL LIQUID ION EXCHANGE- STRIPPING (FIG. C-11) The partially ammonia-stripped organic passes to a second step for reaction with H2SO4 to draw the remaining ammonia into an aqueous phase. The phases are countercurrently con- tacted and separated in two stages. The aqueous phase returns to nickel extraction. The organic phase, containing nickel and trace metal impurities, enters two countercurrent stages of nickel stripping with weak electrolyte from electrowinning. This electrolyte is heated passing to and cooled passing from nickel electrowinning which is operated at higher temperatures. Losses of the organic extractant during the stripping of nickel are replenished with fresh organic. A small amount of organic is purged to a trace metals removal step to prevent a buildup. The "cleaned" organic purge and makeup organic are com- bined with the nickel-free organic for recycle to the nickel LIX extraction step. NICKEL ELECTROWINNING— COMMERCIAL (FIG. C-1 2) Nickel is recovered from the strong electrolyte by electrowin- ning in a manner similar to that used for copper recovery. In nickel electrowinning, however, cathode bags are used, and the strong electrolyte is chemically conditioned with sodium sulfate and boric acid to control its conductivity and pH. Dis- solved organic, carried from the LIX step, is adsorbed on activited cartx)n prior to any electrolyte passing to electrowinning. Nickel is recovered in a manner similar to that used for copper recovery. The starter sheets are pickled in H2SO4 prior to use in the commercial cells, and nickel scrap is dissolved in ammonia- containing raffinate and recycled to the ion exchange process. The electrolyte purge required to remove impurities from the electrowinning circuit passes to raffinate neutralization. The strongly acidic, weak electrolyte returns to the nickel stripping circuit. COBALT RECOVERY (FIG. C-1 3) Along with any unextracted Cu, Ni, and Zn, Co is recovered from the Ni LIX raffinate by precipitation with ammonium hydro- sulfide (produced by sparging hydrogen sulfide into an excess of ammonia solution). The sulfide precipitate is separated from the raffinate in a clarifier. The clarifier overflow is filtered of entrained precipitate and sent to ammonia recovery. The underflow is mixed with electrolyte purges from Cu and Ni electrowinning as well as the Co recovered from stripping the LIX reagent. The mixture is pressure leached with air to prefer- entially dissolve the Ni and Co sulfides, leaving the Cu and Zn sulfides in the residue. The latter are removed by filtration and scid, as minor products, to smelters for recovery of metal values. Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any Zn and Cu solubilized in the first leach, the Ni-Co sulfate solution is heated and autoclaved, and Ni is reduced with hydrogen. Sufficient ammonia solution is added during reduction to neutralize the acid formed. Only a portion of the nickel is removed per pass to prevent overreduction and subsequent contamination of the nickel powder with cobalt. After densification through repeated recycle, the nickel pow- der is removed, washed, and passed to drying and briquetting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator-crystallizer where the remaining nickel and cobalt are precipitated as the double salts with ammonium sulfate. Excess ammonium sulfate is purged, and the salts are redis- solved in strong ammonia solution. The cobalt in solution is oxidized to the cobaltic state (Co^*) with air to allow cobalt to remain in solution. The stream is acidified to remove the nickel salts that are separated and recycled to the pH adjustment step. The nickel-free solution is then heated and autoclaved for removal of cobalt by hydrogen reduction. Sufficient ammo- nia is added to neutralize the acid generated. The cobalt powder is dried and briquetted for sale, and the ammonium sulfate is purged to ammonia recovery. AMMONIA RECOVERY (FIG. C-1 4) The raffinate from cobalt recovery, containing process ammo- nium sulfate and ammonia, is countercurrently contacted with slaked lime in a boil to recover the ammonia value of the sulfate. The precipitate from the lime boil enters a thickener, then overflows to the waste surge tank for disposal in a tailings pond. The overflow liquor is recycled to the CCD circuit. Steam sparged into the lime boil strips ammonia, which passes to a three-stage ammonia condenser-absorber circuit. The aque- ous absorbed ammonia solution returns to the process. The vent gases of the entire process are passed through ammonia recovery. Unabsorbed vent gas is discharged to the stack for disposal. PLANT SERVICES Plant services include process and cooling water supply and treatment, steam raising and power generation, and stack gas treatment. Makeup water is clarified and softened for distribution to the process, as required. Additional treatment is required for cooling tower water makeup, boiler feed water makeup, and for supplying plant potable water. Offgas hydro- gen from cobalt recovery along with coal is burned in the main boilers to raise the required process steam and generate a portion of the power required in the process. Following particulate removal, the flue gases are combined with other process offgases and pass to gas treatment, where sulfur oxides and other acidic constituents are removed by scrub- bing with limestone. The scrubbed offgases are reheated (to avoid condensation of vapors), combined with scrubbed vents from ammonia recovery, and are passed to stacks for disposal. MANGANESE PRODUCTION ADD-ON The recovery of manganese from H2SO4 leach tails is an option that has been investigated. Recovery of manganese from the tailings would require dissolution of the residue and chemical manipulation to precipitate manganese as an oxide, 47 possibly MnOa, involving several separation-purification steps. The Mn02 could be used directly in steel production if specifica- tions are met.or possibly could be used as a feed material to produce ferromanganese or other manganese alloys. The recovery of manganese from the process tails would be contin- gent on market conditions and process economics. Because of the relatively new technology required in manganese recov- ery from H2SO4 leach tails, and the dependency on economic and technical conditions, no detailed flowsheets are presented for this process. PROCESS ALTERNATIVES A major alternative to the proposed process configuration involves a low-temperature and low-pressure treatment of slurried nodules in acid solution. There are, however, major drawbacks, such as solubilization of an undesirable amount of Fe that would need to be removed, longer treatment time, and lower recovery of Ni, Cu, and Co. Other than the metals separation and purification steps, the process configuration is based on the Moa Bay, Cuba, process, formerly operated by the Freeport Nickel Co., of New Orleans, La. In the absence of actual operating data, there is little reason to believe that nickel recovery from ocean floor nod- ules will differ significantly from nickel recovery in Moa Bay iron-laterite ores. A slurry, whether produced from nodules or latente ore, when leached at high temperature will substantially extract the metals of interest as well as undesirable trace metals in amounts depending on the source.The leach reactors are very fundamental in design and of proven operational reliability. The special alloy flash-down values used in this process are economically and mechanically feasible and present the most convenient arrangement. The Cu raffinate neutralization step, which assumes the use of enough ammonia to generate a basic solution and oxidation of Co to prevent coextraction with Ni, could be eliminated if a reagent could be found to selectively remove Ni in the pres- ence of Co. The precipitation and separation of basic salts of Fe, Mn, Mg, and Al would be eliminated, with a savings of necessary equipment. The ammonium sulfate solutions purged from the LIX extrac- tion and the copper raffinate neutralization have been treated with lime for the recovery of ammonia for recycle. An alternative approach would involve direct recovery of ammonia sulfate by evaporation-crystallization for sale as a byproduct. The alter- native is highly energy intensive because of the number of crystallizing-evaporating steps necessary to produce a rela- tively pure byproduct from a stream containing many entrained process impurities. The production of ammonium sulfate is outside the interest of the process and could be severe on the local market, because this production would be approximately 15 pet of present U.S. production and could affect market prices. Alternative configurations of the metals separation steps are possible, such as selective extraction-selective stripping, but their impact on overall plant material and energy balances should be minor. Many variations are possible in the details of the scheme used for recovery of cobalt from a mixed sulfide precipitate. The impact on plant requirements would not differ appreciably from the present approach, because the sulfides will still be oxidized to sulfates, purged, and hydrogen reduced. It does not appear likely that the solution could be purified easily enough to permit recovery of electrolytic cobalt. 48 (/> iir fi T3 O a> >< w 0) 58 [>-- [> 49 50 51 52 ? 5. « o i I (0 c r Q. i I < in S a> 53 54 55 56 57 1 "1 i = 1 s Sx" , »i 1 ' 1 E L \ s s \ 1 I 5 i if II s • ^-^3*^ 61 62 APPENDIX D.— REDUCTION AND HCI LEACH PROCESS The reduction and HCI leach process is a four-metal pro- cess in which Mn, Cu, Ni, and Co are liberated from dried nodules by a high-temperature (500° C) gaseous hydrogen chloride treatment of nodules. Hydrogen chloride reduces manganese dioxide to manganous chloride (liberating chlorine gas) and also reacts with other metal oxides to form soluble chloride salts. A hydrolysis reaction and quench follow, where water is sprayed on the nodules and the iron is precipitated as ferric hydroxide. The nodules are leached with water and HCI, forming a concentrated pregnant liquor of chloride salts. Copper is extracted by liquid ion exchange (LIX) reagents from the pregnant liquor, and is stripped and recovered as electrowon cathodes. Cobalt is solvent-extracted from the copper raffinate, stripped, and separated by precipitation with hydrogen sulfide. It is recovered from the sulfide precipitate, along with some Ni, Zn, and Cu, by selective leaching and hydrogen reduction. Nickel is extracted by LIX reagents from the cobalt raffinate, stripped, and recovered as electrowon cathodes. The nickel raffinate is evaporated, crystallizing man- ganese chloride as well as the other remaining chloride salts. The salts are dried using combustion gases in a countercur- rent dryer. The dried salts are charged to a high-temperature fused salts electrolysis furnace, where molten manganese metal is tapped and cast as product and chlorine gas is liberated. Excess hydrogen chloride gas in the process is recovered and recycled. Generated chlorine gas is recovered, dried, and delivered to a local chemical complex which, in exchange, returns makeup hydrogen chloride to the process. Detailed descriptions of each segment of the process are given for the following flowsheets. A summary of operating parameters for each section is given in table D-1 . A key to the flowsheet symbols is given in figure D-1 . ORE PROCESSING AND DRYING (FIG.D-2) Wet nodules are reclaimed from storage and fed through a primary cage mill where they are reduced to minus 7/8 in. They then pass to a fluid-bed dryer where surface and pore water is removed at 175° C by direct contact drying with combustion gases. Bed overflow is reduced to minus 65 mesh in a secondary cage mill, and dryer and mill fines are removed from offgases by cyclones and an electrostatic precipitator. Offgases pass to gas treatment for scrubbing, while dried nodules are transferred hot in an enclosed conveyor to reduction. HYDROCHLORINATION (FIG. D-3) Dried nodules are reacted in a fluidized bed hydrochlorination reactor by contact with the hydrogen chloride gas from the HCI surge, which is preheated to 1 75° C. The exothermic reactions maintain the reactor at 500° C. Manganese dioxide is reduced to manganous chloride, and chlorine is produced. Essentially all the Cu, Ni, and Co, as well as 90 pet of the alkali and alkaline earths present in the nodules, react to form chloride salts. Approximately 100 pet excess HCI gas is used in the reaction. This HCI gas, as well as the chlorine and water liberated in the reaction are returned to HCI-CI2 recovery. The reduced-chlorinated nodules are delivered to a second fluid- ized bed where any iron chloride is hydrolyzed by a water- spray quench, forming insoluble ferric hydroxide, and the nodules are cooled to 200° C. The offgases from both fluidized beds are sent to a electrostatic precipitator system for dust removal. The hydrolysis offgas is also passed through a waste heat recovery system. LEACHING AND WASHING (FIG. D-4) The chlorinated nodules are leached with water and HCI in a tank where the soluble chlorides are dissolved to form a liquor with a pH of 2. The solution is cooled to 40° C by circulating the liquor through an external heat exchanger. The slurried nod- ules are washed countercurrently in a six-stage thickener circuit to remove 98 pet of all soluble metal values, forming a pregnant liquor. The wash water is recycled water from proc- ess water surge. Washed tailings are sent to the waste treatment area. Flocculant is added to the thickeners to improve the settling properties of the solids. COPPER LIQUID ION EXCHANGE (FIG. D-5) Pregnant liquor is filtered and passed to a three-stage coun- tercurrent LIX extraction circuit where copper is removed from pregnant liquor. Some of the other chloride salts are physically entrained in the organic phase. These entrained chlorides are removed by washing with a water solution in a two-stage wash circuit. A purge from this wash solution is returned to the pregnant liquor surge, and makeup water is added to the circuit equal to this purge. The copper is stripped from the organic by countercurrent contact in two stages at controlled pH with depleted electrolyte from copper electrowinning. Because electrowinning conven- tionally occurs at a higher temperature than the operation of the LIX circuit, the sthp solution is heated passing to electro- winning and cooled passing from electrowinning. Provision is made for periodically cleaning the mixer-settler units used in extraction and stripping and for recovering organic and aqueous phases for recycle. Degraded organic, dust, and other forms of "crud" are removed from the organic and incinerated. Because a small amount of cobalt is coextracted and is not stripped with the copper, it must be removed from the LIX reagent to prevent its buildup. This is accomplished by precipi- tating the cobalt from a purge stream of organic with hydrogen sulfide. The precipitated solids are washed from the organic and passed to cobalt recovery, while the purified organic is returned to the extraction loop. Makeup organic is added to the stripped reagent to offset degradation and soluble organic losses to pregnant liquor and water wash, and the metal-free organic is recycled to extraction. COPPER ELECTROWINNING— COMMERCIAL (FIG. D-6) Cathode copper is recovered from the strong electrolyte from LIX using conventional technology. Starter sheets depos- ited on titanium blanks from strong electrolyte are removed, washed, looped, and returned to the commercial section as starters. Full-term cathodes produced in the commercial sec- tion are washed, unloaded, and prepared for shipment to sale. The major part of the weak electrolyte is recycled to the LIX section for stripping 63 Table D-1.— Operating parameters for reduction and HCi ieach process Parameter ar>d unit Value J Parameter and unit Value ORE PROCESSING AND DRYING (FIG. D-2) | pH ADJUSTMENT AND COBALT EXTRACTION (FIG. D-7)— Con. Feed rate, wet basis (330 days/yr; 24 hr/day) .... tpd.. 3,640 Metals extraction, pet— Con. Feed size to reduction mesh.. -65 Mn 10 Drying temperature "C. 150 Ni Zn Co stnpping Stnpping solution pH HYDROCHLORINATION (FIG. D-3) | Temperature °C.. 500 4 Pressure atm.. 1 Number of stages 2 Reagent HCI Organic-aqueous ratio 5:1 Reactions: Stnpping, pet MnOz + 4 HCI — > MnCl2 +2H2O + Ct Co 99 FejOa +6HCI-> 2FeCl3 + 3H20 Mn 99 CuO + 2HCI— > CUCI2 H-HaO Co recovery NiO + 2HCI — > NiClz + HjO Precipitating agent H2S COzOa + 6HCI —> 2C0CI2 + SHjO + CI2 Precipitation pet 100 Temperature °C 80 Temperature 'C. atm 200 1 Pressure ...atm.. 1 Pressure HydrochlonnatKXi, pet Co NICKEL LIQUID ION EXCHANGE (FIG. D-8) 100 Extraction: Cu 96 Extractant 'KelexlOO Fe 27 Number of stages 2 Mn 94 Organic-aqueous ratio 2:1 Mo 96 Temperature °C 40 Ni 100 Metals extraction, pet Ni Co Cu LEACHING AND WASHING (FIG D-4) | 99.5 99.5 RnalpH 2 99.5 Underflow density pet solids 15 pH (with NaOH) 4 Wash ratio 2:1 Washing Number of stages 6 Washing agent H2O Soluble metals removed pet- 98 2 Stripping: Strip solution composition, gpl: H2S04 COPPER LIQUID ION EXCHANGE (FIG. D-5) | Extraction: 1 40 Extractant 'KelexlOO 1 Ni 50 Cu feed- gpl *8 Number of stages 2 Number of stages 3 Organic-aqueous ratio 6:1 2:1 Ni stripping pet 99 Cu extraction Washing: w&sn composftion pet 99.5 H2O NICKEL ELECTROWINNING-COMMERCIAL (FIG. D-9) Current density A/m^.. 180 Number of stages 2 Current effiaency pet.. 93 Organic-aqueous ratio 3:1 Temperature °C.. 60 Temperature °C 40 Ni in-out gpl- 75-50 Metals extraction, pet H2SO4 in-out gpi- 0.016-40 Co 99.5 Na2SO« gpi- 100 Cu 99.5 99.5 HaBOa gpi- 15 pH (with NaOH) MANGANESE RECOVERY (FIG. D-10) Stripping: Trace elements removal agent H2S Stnp solution composition, gpl: Evaporation-crysfallization water removal ....pct.. 99.9 H2SO«.. 160 Fused salt electrolysis: Cu 40 Mn recovery.... pet.. 90 Number of stages 2 Temperature, metal »C.. 1,300 Organic-aqueous ratio 4:1 Temperature, salt °C.. 800 Cu stripping pet 95 Cun-ent density A/m^.. 46 Temperature •c 40 COBALT RECOVERY (FIG D-11 ) COPPER ELECTROWINNING— COMMERCIAL (FIG D-6) | Sluny feed, solids pet 40 Current density A/m^ 180 Evaporation-crystallization water removal pet 70 Cun-ent efficiency pet 94 Co oxidation Temperature "C 50 Temperature "C 100 Cu in-out gpl 53-40 Pressure psig 150 HjSO* in-out gpl 160-180 Co reduction Temperature Pressure Reductant °C psig pH ADJUSTMENT AND COBALT EXTRACTION (FIG. D-7) | 175 500 Cu raffinate pH adjustment: H2 pH adjustment agent NaOH LoficninQ 8QGnt .pet HaSO^.. 70 Number of stages Final pH 1 4 HCI RECOVERY (FIG. D-1 2) Co liquid ion exchange extraction Extractant HC\ Ahsnrhinn annnt H2O H2SO4 TIOA Gas drying agent Number of stages Organic-aqueous ratio 3 2:1 WASTE RECOVERY (FIG. D-1 3) Metals extraction, pet Co 99 NHa recovery ....pet.. 99 Cu 100 'Reference to specific trade names does not imply endorsement by the Bureau of Mines. 64 A small amount of nickel is stripped along with the copper not deposited in electrowinning, and must be purged from the system. The purged electrolyte passes through purification cells where copper is removed by electrowinning to depletion. The decopperized electrolyte is sent to cobalt recovery for further processing. Makeup acid is used to redissolve scrap copper for return to the commercial cells for deposition, and sufficient steam, wash water, and makeup water are added to the circuit to offset water vaporized or carried off with evolved oxygen during electrowinning. pH ADJUSTMENT AND COBALT EXTRACTION (FIG. D-7) The raffinate from copper extraction is adjusted in a surge tank to a pH of 4 by the addition of NaOH solution. This stream then passes to a two-stage countercurrent solvent extraction circuit where cobalt is extracted as a tetrachloro-complex by the organic reagent. The preferred organic reagent is a tertiary amine such as tri-isooctylamine (TIOA). The cobalt is stripped from the organic by countercurrent extraction contact in two stages with recycled process water. Because of the high man- ganese concentration in the raffinate, a portion of it is extracted with the cobalt. The cobalt is separated by addition of hydro- gen sulfide, which reacts to form cobalt sulfide precipitate which is centrifuged from the liquor and sent to cobalt recovery. The liquor, which is basically manganese chloride solution, is sent to manganese recovery. NICKEL LIQUID ION EXCHANGE (FIG. D-8) Nickel is recovered in a manner similar to that used in copper extraction. The cobalt-free raffinate passes to a three- stage countercurrent LIX extraction circuit where nickel is removed from the raffinate. Sodium hydroxide solution is added at interstages to keep the pH at 4 to ensure good nickel extraction. The loaded organic is washed to remove chloride in a two-stage operation. The nickel is stripped from the organic by countercurrent contact in two stages with depleted electro- lyte from nickel electrowinning. Because electrowinning con- ventionally occurs at a higher temperature than the LIX circuit, the strip solution is heated-cooled passing to-from electrowinning. Provisions are also made for cleaning the mixer-settler units and sending the resultant "crud" to incineration. An organic purge is also taken to ensure against buildup of unstrippable metals in the organic. Makeup organic is added to the stripped reagent to offset degradation and soluble organic losses. NICKEL ELECTROWINNING— COMMERCIAL (FIG. D-9) Nickel is recovered from the strong electrolyte by electrowin- ning in a manner similar to that used for copper recovery. In nickel electrowinning, however, cathode bags are used, and sodium sulfate and boric acid are added to the electrolyte to control its conductivity and pH. Dissolved organic carried from the LIX step is removed by adsorption on activated carbon prior to any electrolyte passing to electrowinning. Also, the starter sheets are pickled in H2SO4 prior to use in the commer- cial cells, and nickel scrap is redissolved in HCI-containing raffinate and the solution recycled to the cobalt-free raffinate surge. The electrolyte purge, required to remove impurities from the electrowinning circuit, passes to a sulfide precipita- tion reactor and then to cobalt recovery for retrieval of metal values. MANGANESE RECOVERY (FIG. D-10) The raffinate from nickel extraction is combined with the wash from cobalt extraction, and these are reacted with hydro- gen sulfide to precipitate any remaining metal values that may have passed through the series of extractions. These precipi- tates are sent to cobalt recovery. The resultant solution is evaporated in a triple-effect evaporator using steam as the initial heat source. The overhead water is cooled and sent to the process water surge for recycle. The wet crystallized chlo- ride salts are sent to a dryer, where the salts are passed countercurrently with combustion gases to drive off the remain- ing surface water as well as any water of hydration. The vent from the dryer is sent to gas treatment. The dried salts are conveyed in covered systems to a surge, which is blanketed with inert nitrogen gas to prevent reabsorption of water. From this surge, the salts are fed to a high-temperature (1 ,300° C) fused-salt electrolysis furnace where molten manga- nese forms and collects at the bottom of the reactor. The manganese is tapped periodically and cast into molds and sold as metal product. Chlorine gas that evolves from the electrolytic reaction is collected, cooled by a water quench, and sent to HCI-CI2 recovery. Fused salts are also tapped from the reactor. These salts are mold cooled and sent to the waste disposal area. A certain amount of the manganese that does not pass inspection is recycled to the furnace with additives that assist in carrying out a complete recovery of the manganese. The lining of the reactor is cooled to prolong its life in this harsh environment. Areas surrounding the furnace are hooded, and constant venting is maintained to collect fugitive fumes from the furnace and the products tapped from the furnace. The graphite anode is also continuously replaced. COBALT RECOVERY (FIG. D-11) Several streams are merged to form the input to the cobalt recovery scheme. These are the slurry from Co extraction, purge from Mn recovery, the purges from Cu and Ni electrowinning, and the solids from organic stripping in Cu and Ni extraction circuits. The mixture is pressure leached with air to preferentially dissolve the Ni and Co sulfides, leaving the Cu and Zn sulfides in the residues. The latter are removed by filtration and sold, as minor products, to smelters for recovery of metal values. Following pH adjustment and reprecipitation with hydrogen sulfide for final removal of any Zn and Cu solubilized in the first leach, the Ni-Co sulfate solution is heated and autoclaved, and Ni reduced with hydrogen. Sufficient ammonia solution is added during reduction to neutralize the acid formed. Only a portion of the nickel is removed per pass, to prevent overreduction and subsequent contamination of the nickel powder with cobalt. After densification through repeated recycle, the nickel powder is removed, washed, and passed to drying and briquetting for sale. The largely nickel-free cobalt sulfate solution passes to an evaporator-crystallizer where the remaining nickel and cobalt are precipitated as the double salts with ammonium sulfate. Excess ammonium sulfate is purged, and the salts are redis- solved in strong ammonia solution. The cobalt in solution is oxidized to the cobaltic state (Co^^) with air. This permits the cobalt to remain in solution when the stream is acidified to remove the nickel salts, which are then separated and recycled to the pH adjustment step. The nickel-free solution is then heated and autoclaved for removal of cobalt by hydrogen reduction. Sufficient ammonia is added to neutralize the acid generated. The cobalt powder is dried and briquetted for sale, and the ammonium sulfate is purged to the lime boil. HCI RECOVERY (FIG. D-12) The offgas from the hydrochlorination reactor contains a mixture of unreacted HCI, chlorine, and water. The HCI is absorbed by water, forming a highly concentrated aqueous HCI solution. The overhead gas from the tower is combined with the offgas from electrolysis and is dried by passing through a H2SO4 solution, which strongly absorbs the water from the gas, leaving dry chlonne for delivery as product. The HCI gas from the hydrolysis reactor is absorbed in the second tower, producing strong HCI solution. The offgas from this tower is mainly humid air and is sent to gas treatment. In both absorp- tion towers, the dissolution of HCI is highly exothermic, and the heat is removed with cooling water. A series of lean HCI vents is scrubbed to remove the last remaining HCI in a third tower. The overhead gas from this tower is also sent to gas treatment. All the HCI solutions are combined with H2SO4, and the HCI is stripped and sent to a HCI surge, where makeup HCI is added. The HCI is then sent to the hydrochlorination reactor. The bottoms from the HCI stripper are sent to a water stripper, where water is taken overhead and condensed. The remaining H2SO4 is filtered to remove any particulate matter that may have been entrained in the gas streams entering the HCI recovery section. The resulting strong H2SO4 is recycled. An aqueous HCI stream is split from the main aqueous HCI stream to provide acid streams needed at various stages in the process. Excess water devel- oped in HCI recovery is returned to the process water surge. WASTE RECOVERY (FIG. D-13) Ammonia is recovered from the ammonia sulfate purges from cobalt recovery by reaction with slaked lime. Steam is blown into the mixture to strip the evolved ammonia. The vapor is combined with other ammonia vents, and they are scrubbed with water to form an aqueous ammonia solution that is returned to the aqueous ammonia storage for recycle within the process. The gypsum slurry from the lime boil is cooled and combined with the slurry from stack gas treatment. After liquid-solid separation in a thickener, the solids are com- bined with other process solid and liquid wastes and plant runoff, treated for pH control as required, and pumped to the tailings impoundment area for disposal. The overflow from the thickener is returned to the process water surge, where it is combined with other water returns and makeup water to sat- isfy the process water needs. PLANT SERVICES Plant services include process and cooling water supply and treatment, steam raising and power generation, stack gas treatment, and combustion gas preparation. Makeup water is clarified and softened for distribution to the process, as required. Additional treatment is required for cooling tower water makeup, boiler feed water makeup, and for supplying plant potable water. Process combustible wastes are burned along with coal in the main boilers to raise the required proc- ess steam and generate a portion of the power required in the process. Following particulate removal, the flue gases are combined with other process offgases and pass to gas treatment, where sulfur oxides and other acidic constituents are removed by scrubbing with limestone. The scrubbed offgases are com- bined with scrubbed vents from ammonia recovery and are passed to stacks for disposal. PROCESS ALTERNATIVES While data on the reduction of nodules with HCI and subse- quent recovery of metal values from acid chloride solution have been reported in the patent literature, this technology currently has no direct analog in commercial extractive metallurgy. Also, while the thermodynamics of the reduction step are well known, and reports on laboratory studies of some separations of metals from chloride solutions are available, no data exist to indicate the expected properties of the nodule residues (e.g., filtering rates) or the problems associated with recovering a salable manganese metal from impure chloride solutions. Thus, the proposed process configuration has been established based on literature data that contain little detailed engineering and design information and without the benefit of insights that might be gained from analogies to conventional technology in related operations. This means that alternative process configurations might well be technically and economi- cally more attractive, but access to more data would be required to make such a judgment. A major process alternative is to use a low-temperature aqueous reduction and HCI leach. The stoichiometry of the reactions would consume the same amount of HCI as the high-temperature gaseous reduction. However, the iron would be dissolved as ferric chloride and would necessitate more expensive solvent extraction steps to remove the iron before recovering the valuable metals products and a spray roast of the resulting iron chloride solution to recover the HCI. On the positive side,an aqueous process would not require the drying step. A number of possible alternatives is available for recovering manganese. Direct electrowinning of manganese has been proposed, but it is questionable that the final pregnant liquor would be pure enough to develop a good electrowon product. A cementation on aluminum to obtain manganese has also been proposed. The resulting aluminum would be spray roasted to recover HCI and form AI2O3 as a product, which must be sold or disposed. In addition, the manganese product may be contaminated with aluminum. Other schemes propose produc- ing manganese hydroxide as a product which, in effect, side- steps the problem of obtaining a pure metal product. A number of metal separation schemes are possible. These include using different organic reagents as solvent extractants or LIX reagents. (0 5-0 0) a>3 3 O) ;/ / E > ; \ E I •o o o8 (M r\_ o rs>_ o r\ L^ [> 67 ?5 J? S c 68 -e ns^ ^ ^ rm^ r^ [^ r^ 4 III 'lli 70 71 72 a \ ^ 11 o So / -W C.H o / oj c S) 73 II 1 74 75 76 4r n W!WMMtf#; ' .WWlM«)Mwwm] i i i i» i 78 => V) i 5 £ 5 I s o o. 79 APPENDIX E.— SMELTING AND H2SO4 LEACH PROCESS The smelting and H2SO4 leach process is a combination pyrometallurgical and hydrometallurgical treatment of nod- ules to recover the value metals Ni, Cu, and Co, with the option of recovering ferromanganese or a storable byproduct of Mn and Fe. The smelting process produces a slag, from which ferromanganese is recovered, and a metal alloy matte com- posed primarily of Ni, Cu, Co, and S. The matte is granulated, slurried, and selectively leached with H2SO4 at elevated temperature and pressure. The leach residue and metalliferous solution are separated by a series of filtering and washing stages. After liquid-solid separation, cop- per and nickel are selectively extracted by liquid ion exchange (LIX) reagents, stripped from the ion exchange liquid into a weak electrolyte, and recovered as electrowon cathodes. Cobalt is separated from the raffinate by precipitation with hydrogen sulfide and recovered from the sulfide precipitate, along with some Ni, Cu, and Zn, by selective leaching and hydrogen reduction. Ammonia consumed in the process is recovered by lime boil and recycled to the process for use in pH control. Detailed descriptions of each segment of the process are given for the following flowsheets. A summary of operating parameters for each section is given in table E-1 . A key to the flowsheet symbols is given in figure E-1. ORE PREPARATION AND DRYING (FIG. E-2) Wet nodules are reclaimed from storage and fed through a primary cage mill, where they are reduced to minus % in. They then pass to a feed bin for delivery by a feed belt supplying nodules to a direct heated, fluid-bed drier for water removal. REDUCTION (FIG. E-3) The dried nodules are combined with coke, and the mixture is fed to a fluid-bed roaster for reduction with producer gas. After the roaster reduction, cyclones remove and return large particulates, while fine dust and hot gases pass to waste heat recovery and electrostatic precipitation. The reduced dust is delivered back to the process to recombine with the reduced products of the roaster. The reduction products are blanketed by an inert gas (such as nitrogen) and delivered by hoppers to the smelting furnace. SMELTING (FIG. E-4) Reduced nodules, comprised of MnO, FeO, asmall amount of metallic Fe, the value metals Ni, Cu, and Co, and the less volatile components of nodules, are smelted with silica flux at about 1 ,425° C. An electric furnace of conventional design would be used. Recoveries of Fe, Ni, Cu, and Co are in the range of 70 to 95 pet along with minor amounts of Mn. The iron alloy is subse- quently removed and recycled to the smelting furnace as a molten silicate. The slag is comprised mainly of Mn, Fe, and Si02-Ca in the proper ratio for good slag fluidity and for subse- quent production of ferromanganese. CONVERTING (FIG. E-5) Manganese reduction in the alloy is held at a level not exceeding 2.0 pet because it must be removed to less than 0. 1 pet prior to reacting the Ni, Cu, and Co with S. Removal of the manganese and some iron is accomplished by addition of quartzite in the proper ratio to produce an eutectic mixture of the low-melting silicates. Oxidation of iron and manganese with 95 pet O2 conserves heat for subsequent processing. After the Fe-Mn-Si02 slag is removed as a first step, a near stoichiometric amount of S must be reacted to form Ni3S2, CU2S, and CogSe. These sulfides are stable with respect to Fe and FeO at 1 ,200° to 1 ,400° C. The objective is to remove as much iron as possible without a major loss of the value metals to the exhaust gas or slag. Gypsum, reduced with coke, sup- plies the sulfur, and a fuel oil-oxygen burner supplies the heat for the reaction. A top-blown rotary converter (TBRC) was selected for con- verter operations to provide intimate slag-metal-gas contact. All of the converting could be done in the same vessel, but two are shown in the flowsheet to assist visualization of the four- step process. The third step is to add silica flux and blow the matte-iron alloy with air until the proper amount of iron is removed to balance its requirements in the ferromanganese alloy that is produced in another section of the plant. Alternatively, if iron in the nodules were low enough in relation to manganese, all of the iron silicate slag could be recycled to smelting. As a final converting step, iron is lowered to 5 pet by a series of alternate blowing-slag removal-fluxing-blowing operations. The reaction is highly exothermic, requiring some care in prevention of Fe304 production. The finished matte contains about 90 pet of the Ni, Cu, and Co and is removed by ladle to granulation. FERROMANGANESE REDUCTION (FIG. E-6) Ferromanganese can be produced from the smelter slag in an open air electric furnace with a reductant. Coke or several other carbonaceous materials would be suitable as reductants. The reduction reaction requires intensive energy input from the electric arc. Because the charge is primarily molten, the bath surface will be exposed in molten form. Standard ferro- manganese production is from cold ore, and a crust is present on the bath surface. Medium carbon ferromanganese is produced, and a slag containing about 8 pet manganese is discarded as the final waste product from the hot metal operations. MATTE LEACHING (FIG. E-7) The molten alloy from converting is quenched in a granula- tion unit. The granulated alloy is rake classified, with oversize granules going to wet milling for final size reduction while classifier fines overflow to a clarifier, where they are settler thickened. The clarifier underflow recombines with wet mill effluent in a surge tank, while overflow returns to granulation. The matte slurry is pumped to an autoclave leaching vessel operated at 1 50 psig and 110° C. The leach products, die- 80 Table E-1.— Operating parameters for smelting and H2SO4 leach process Parameter arxj unit ™. 1 Parameter and unit Value ORE PREPARATION AND DRYING (FIG. E-2) | FERROMANGANESE REDUCTION (FIG. E-6)— Con. Feed rate, wet tMSis (330 days/yr; 24 hr/day Feed size tpd.. in.. °C.. 10.900 -7/8 150 Slag composition, pet CaO MnO NaaO MgO Drying temperature 15 REDUCTION (FIG. E-3) | 22 8 Reduction gas: Producer gas ..pet CO.. pet coke.. S20 »4.5 90 90 100 100 20 90 725 925 4 3 MATTE LEACHING (FIG. E-7) Reduction reactions: + C02 O + CO2 CO^ J02 COj to + 3CO2 MnOz + CO— > MnO Fe203 + CO->2Fe CuO + CO— > Cu + NiO + C0— >Ni + C CoO + CO— >Co + M0203 + 3CO— >2*/ Metals reduction, pet: Co Cu Matte temperature ° C Granulation temperature, initial ° C Output partde size mesh Pulp density pet Leaching Temperature ° C Pressure psig Time hr Leachate Recovery, pet Co. Cu. Fe. Mn. Ni.. Washing effiaeney pet Filtration stages 1.325 95 Ball mill -325 9.5 110 150 2 Fe Mn Mo Ni Temperature. ° C Solids H2S04 99 99 99 80 Gas 99 SMELTING (FIG. E-4) | 98 2 °C *1,325 hr 2 Silica Co) (0 ;'> iir r> Ly r\ n ly 85 86 87 90 91 % I 1 1 92 ?5 = 94 \rsi 97 '^"^3:^ 98 f -! ^ i §t ' I (0 o 1 %> '•''/"V-^^^-"."' ' x'-^^-'.y' ^o**,. '.'^IP^: .*^°-' •. V* ^•l^iLr* o. ^K" .^.. •"-. ^^ 0* 0> _,; ;->*'\ •; V J'-'-^ /% •■ „ *'..V^\/.. v^v V^y %-Wv' V^ 'T^V \*^^\/ V^^'/ \/^^\/ V^V \^^ ^^SEP 83^\ %^^ :^^% %/" :^g. %,<.* ,-^fe\ \/ ,*^^^. %^^