TN295 .U4 No. 9176 LIBRARY OF CONGRESS 0DQ01044345 ■> ^ -♦*$£<♦ a,* % ***** 0* .MV.V ^ ^3 V x* v *v oV vv ;- -w • *bv v :Sf^ w V^\/ -^^, -ee-o* 'o.o' ^> . , t * A. U * o « o v/||V#\/» V.-. ■*♦ ^*.«fc.V" x.iv->*"" ^a^S'"'V -rA 4X <> *'..«* <& x * AT "O* ,* ... o -o . » « A 3 *•_ ■» o /.•^i-.\. c°*..i^.% f.-£H..%, ^ v 4 0, ^ i W #: W '»- \/ .^'- W V <#»*""*'.* «-' C /i v "x ^^s^\^Ws^x #A# ^x : - :\/' S°* .. -+„ ° h ° ^ ... e». " , ' ^ ; X .X <^^mkX <*V <► ■5* A°« ^ ** % C° * * J>' \ •\/ ♦♦ ^ :J\ \Wws J*x l S§g^s J\ i^w) /x \»f •- /C liBfc .^ ^ *-o^. 4 a X ^?r^ , o^ ^°^ Bureau of Mines Information Circular/1988 Processing Technologies for Extracting Cobalt From Domestic Resources By C. E. Jordan UNITED STATES DEPARTMENT OF THE INTERIOR Information Circular 9176 Processing Technologies for Extracting Cobalt From Domestic Resources By C. E. Jordan UNITED STATES DEPARTMENT OF THE INTERIOR Donald Paul Hodel, Secretary BUREAU OF MINES David S. Brown, Acting Director TV Library of Congress Cataloging in Publication Data: Jordan, C. E. Processing technologies for extracting cobalt from domestic resources. (Bureau of Mines information circular; 9176) Bibliography: p. 23-24. Supt. of Docs, no.: I 28.27:9176. 1. Cobalt— Metallurgy. 2. Cobalt ores— United States. I. Title. II. Series: Information circular (United States. Bureau of Mines) ; 9176. TN296.U4 [TN799.C6] 622 s [669'.733] 87-600331 CONTENTS Page Abstract Introduction Sulfide ores Blackbird deposit Duluth Gabbro deposits Yakobi Island deposit Madison Mine Missouri lead belt deposits Copper deposits Pyrite concentrates Spent copper leach solutions Iron deposits Oxides Laterite deposits Manganese sea nodules Extraction methods. Gas reduction and ammoniacal leach process Cuprion ammoniacal leach process High -temperature, high-pressure H 2 S0 4 leach process Reduction and HC1 leach process Smelting and H 2 S0 4 leach process Refining process Ammoniacal solution refining Acid sulfate refining Acid chloride refining Manganese sea crusts Discussion Conclusions References ILLUSTRATIONS 1. Domestic cobalt resources 3 2. Blackbird Mine cobalt process 5 3. Madison Mine cobalt process 8 4. Cobalt from spent Cu leach solution 10 5. Cobalt from Pennsylvania iron ore 11 6. Reduction roast -ammoniacal leach process for laterites 13 7. Sulfuric acid leach process for laterites 14 8. Gas reduction and ammoniacal leach process for sea nodules 16 9. Cuprion ammoniacal leach process for sea nodules 16 10. High-temperature, high-pressure H2SO4 leach process for sea nodules 17 11. Reduction and HC1 leach process for sea nodules 17 12. Smelting and H2SO4 leach process for sea nodules 18 TABLE 1. U.S. cobalt resource summary 22 1 2 4 4 6 7 7 7 9 9 9 11 12 12 15 15 15 15 15 15 18 18 18 19 19 20 20 23 23 UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT cm centimeter pet percent °C degree Celsius ppm part per million g/L gram per liter psi pound per square inch h hour st/d short ton per day lb pound um micrometer lb/h pound per hour yr year mm Hg millimeter of mercury PROCESSING TECHNOLOGIES FOR EXTRACTING COBALT FROM DOMESTIC RESOURCES by C. E. Jordan 1 ABSTRACT Domestic cobalt resources are relatively large, but low grade. The full potential for cobalt production from domestic sources is at least 19.4 million lb of cobalt per year exclusive of offshore resources. A summary of the cobalt processing technologies for the major domestic resources is presented in this Bureau of Mines report. The processing technologies for the Blackbird, Madison Mine, Duluth Gabbro, iron ore pyrite, laterites, and manganese sea nodules are nearly complete, but the economics are not favorable. Research on these resources should be limited to approaches that promise to cut the total processing costs by at least 50 pet. The most promising sources of cobalt are the spent copper leach solutions and siegenite from the Missouri lead ores. Research on cobalt processing from these two sources needs to be completed. 'Metallurgist, Tuscaloosa Research Center, Tuscaloosa, AL. INTRODUCTION The United States is the largest consumer of cobalt in the world, but has no domestic cobalt production. Except for some scrap recycling, the United States depends entirely on foreign nations to supply over 15 million lb of cobalt annually. More than half of the cobalt metal comes from Zaire and Zambia (19). 2 With most of the cobalt coming from a small number of foreign producers, the United States is vulnerable to supply disruptions. Cobalt is a critical element in many industrial and military products such as jet engine parts, high-strength tool steels, heat- and corrosion-resistant alloys, magnets, catalysts, drying addi- tives in paints, and other chemicals. About two-thirds of the domestic cobalt consumption requires cobalt metal, either as a powder or as a high-purity cathode. The remaining third is as oxides or salts used in chemicals and paint drying addi- tives (19). About half of the 15 million lb the United States consumes annually is considered essential (22). In recogni- tion of the strategic importance of cobalt, research has been directed in three areas, recycling technology, cobalt — - Underlined numbers in parentheses re- fer to items in the list of references at the end of this report. substitutes, and recovery technology from domestic resources. Currently, recycling accounts for about 8 pet of domestic consumption and has a potential for over 25 pet (19). Substitution technology has potential for an additional 25 pet, but research and development is moving slowly owing to the current low price of cobalt (22). Extensive research on primary cobalt production has been conducted on virtually every major cobalt-bearing deposit. In an effort to optimize future research and establish research priori- ties, the Bureau of Mines investigated the status of processing technologies for extracting cobalt from primary domestic resources. Because domestic cobalt resources are low grade, cobalt production is likely only if metals such as nickel, copper, iron, lead, and zinc are recovered. There are two basic types of cobalt deposits, sulfide and oxide. The benefi- ciation and processing largely depends upon the type of ore and the associated metals. Each resource with over 15 million lb of Co is presented in this report (fig. 1). Along with the resource description, the benef iciation and processing technologies are presented with an assessment of the technical progress and the environmental problems. CA OR ID NV MT WY ND MINI SD NE IA Wl IL CO KS *> OK NM rx jyio_ AR NH- ME J// Ml IN OH PA KY TN, MS LA AL ,GA WV. SC VA NC -MA Rl NJ -DE MD If o \ \ Cb \ \ ~v \ X sHI N s ..^a:*? 1 ^? LEGEND • Cobalt FL FIGURE 1 .—Domestic cobalt resources. SULFIDE ORES BLACKBIRD DEPOSIT The Blackbird deposit near Salmon, ID, contains 0.65 pet Co with nearly all of the cobalt found in cobaltite (CoAsS) and a minor amount associated with chalcopy- rite (CuFeS 2 ). The deposit's proven reserves are over 60 million lb of Co. Noranda Exploration Inc., Cobalt, ID is waiting for favorable economic conditions to build a benef iciation and cobalt processing plant with a planned capacity of 4 million lb of Co per year. The arsenic content of the ore has focussed some attention on the need for safe min- ing and tailings disposal. High cobalt and copper levels were found in streams draining from the deposit's historical mining site. An ion exchange process was successfully field tested to remove both the cobalt and copper from these runoff streams. The environmental costs associ- ated with this deposit have been estimated at $3.00/lb of Co produced. Benef iciation of the Blackbird ore began with crushing and grinding the ore to 70 pet minus 200-mesh size. Using a sequential sulfide flotation process, chalcopyrite was floated first. Only 5 pet of Co was lost in the chalcopyrite concentrate containing 26 pet Cu and 0.65 pet Co. The remaining sulfides were floated, producing a bulk sulfide concen- trate containing 5 pet Co, 0.1 pet Ni , and 0.4 pet Cu. A pilot plant was oper- ated to demonstrate this process and 80 pet of the cobalt was recovered in the concentrate (10). Historically, extraction of cobalt from the Blackbird concentrate began with a controlled oxidizing roast. However, because of the environmental problems associated with arsenic fumes, this technique is no longer considered appro- priate. Fortunately, there are three hydrometallurgical alternative methods that do not require the oxidation roast. Cobalt can be leached as a sulfate, chloride, or ammine complex. Because a single company owns the Blackbird deposit, only the process proposed by Noranda Exploration Inc. will be discussed here. The cobalt was dissolved by pressure leaching with sodium sulfate at 200° C and 150 psi oxygen pressure (fig. 2). Actually, the pyrite in the cobalt concentrate was oxidized, produc- ing H 2 S0 4 and Fe 2 (S0 4 ) 3 , which leached the cobalt from CoAsS. Sodium sulfate helped to suppress iron and arsenic dissolution in the autoclave. Ninty- seven percent of cobalt was extracted in a solution containing 30 g/L Co and 100 g/L H 2 S0 4 . The leach slurry was cooled to 95° C and neutralized to pH 1.5, precipitating jarosite and ferric arsenate. The pregnant solution also contained nickel and other impurities such as iron, arsenic, copper, and zinc. A semicontinuous pilot plant using 22-lb batches was operated to demonstrate the extraction process. Commercial equipment for this high-pressure oxidation-H 2 S04 leach process is available. After extraction of the cobalt into solution, the refining process used largely depends upon the impurities and the final commercial cobalt product desired. Noranda chose to market a high- purity cathode cobalt. First, the hot pregnant solution was filtered from the leach residue, jarosite and ferric arsenate. At 50° C and 1.5 pH, H 2 S was added to precipitate the copper and arsenic as sulfides. The filtered solu- tion was oxidized and neutralized to pH 5.0 at 75° C to precipitate the residual iron. Zinc was removed from the filtered and cooled solution by ion exchange at pH 4.0 and ambient temperature. The solu- tion pH was then adjusted to 2.5 and the nickel was removed by ion exchange. Finally, the cobalt was precipitated as cobaltous hydroxide with a lime slurry. After filtering the solution, the cobalt- ous hydroxide was leached with spent cobalt electrolyte followed by cobalt electrowinning at 50° C. The cobalt metal was stripped from the cathode and traces of hydrogen were removed by vacuum degassing at 820° C and 200 mm Hg. All of the unit operations, except the zinc and nickel ion exchange steps, are estab- lished commercial technologies. ^ Bulk con cent rate Wash 1 iquor ^ i \ \ Oxygen Pressure leach Ma CD %M««k w ^ nd23U4 ■ I CIOII Lime ' i \ r ^r V w Do<.i Fe, Zn, Al Stripping NH4OH -► Co, Ni Stripping -►Cu, Ni SX Stripping -* Fe, Zn, Al SX Cobalt stripping Nickel precipitation Cobalt electrowinning NiC0 3 Co FIGURE 4.— Cobalt from spent Cu leach solution. 11 copper leach solution (18). An economic evaluation of the process from spent copper solution to cathode cobalt metal showed that, with credits for the copper and nickel values, cobalt could be produced for about $7.00/lb. The research on cobalt recovery from spent copper leach solutions is nearly complete. A long-term pilot program may be needed to establish resin life and the steady-state solution concentrations. An economic evaluation indicated that this cobalt source has the highest potential for domestic production. IRON DEPOSITS Cobalt is also found in some iron deposits, associated with sulfides. In the iron deposits of Cornwall, PA, 56 million lb of Co is present at grades ranging from 0.02 to 0.056 pet Co. About 40 to 60 pet of the ore is magnetite and 3.5 pet pyrite (29). Historical process- ing of these ores recovered byproduct Cu, Co, Ag, and Au. In the 2 yr prior to closure of these iron mines, the cobalt grade dropped by 30 pet. However, it was not clear from that report whether the grade of cobalt in the sulfides dropped or if the sulfide content of the ore dropped. Before the iron mines closed, 1.5 million lb of Co was produced annually. Benef iciation of these ores began with crushing and grinding followed by magnet- ic separation to recover a high-grade magnetite concentrate (fig. 5). The non- magnetic product was floated to recover a bulk sulfide concentrate, mostly pyrite, containing 0.7 to 1.4 pet Co. Cobalt extraction from the pyrite con- centrate began with roasting and shipping the calcine product to Pyrites Co. Inc. in Wilmington, DE. The calcine was per- colation leached with sulfuric acid for an average of 250 h. The Co, along with some of the Fe, Cu, and Mn, was dissolved into the solution. The extraction percentage was not reported. Ore Magnetic separation -+• Magnetite Flotation Tailings Pyrite Roast H 2 S0 4 Na 2 C0 3 Leach -► Fe Purification Na 2 C0 3 ► Cl 2 — ► -+■ Fe, Cu, Mn Precipitation _£ Calcined — J~~ Co oxide l-± Reduction T Co FIGURE 5.— Cobalt from Pennsylvania iron ore. The pregnant solution was purified with sodium carbonate, which precipitated the Fe, Cu, and Mn. After filtering, chlor- ine and more sodium carbonate were used to precipitate the cobalt. The filtered cobalt cake was either calcined to the oxide containing 70 pet Co or reduced to the Co metal with charcoal. The final metal powder contained 98 to 99 pet Co. Over 99 pet of the cobalt in the solution was recovered (33). This process was used commercially up until 1971 when the mining companies closed their iron mines, cutting off the supply of byproduct cobalt-bearing pyrite. If iron ore markets improve, this source of cobalt may resume production. 12 OXIDES LATERITE DEPOSITS The laterite deposits of northwestern California and southwestern Oregon are the result of extensive weathering of ultramafic serpentine rocks and contain over 96 million lb of Co. However, the deposits are located in a wilderness preserve. The average grade is 0.08 pet Co, but varies throughout the deposits from 0.06 to 0.25 pet Co. There is also 0.5 to 1.2 pet Ni, 2 pet Cr, and 0.3 pet Mn in these deposits. Nearly all of the Co is found in the manganese mineral, which makes up only 1 pet of the ore. Well over a third of the ore is goethite, which contains most of the nickel. The remainder of the ore is quartz, hematite, magnetite, and chromite (5). Ore benef iciation of the laterite deposits has been limited to crushing and screening out some of the coarse pieces of quartz (30). The grain size of the goethite and manganese mineral is very small, which would require fine grinding of the ore to obtain liberation. No successful benef iciation techniques have been developed for this fine-grain mineral separation. Even with a perfect separation, the Co and Ni grades would only be 2.5 times larger than in the ore. Considering the limited benef iciation potential and the lack of technology, these laterite ores will not be signifi- cantly beneficiated before the extraction process. The extraction of cobalt from laterites has received considerable research atten- tion. The two major techniques for cobalt extraction are reduction roast- ammoniacal leach (fig. 6) and H2SO4 leach (fig. 7). The basic reduction roast- ammoniacal leach process shown in figure 6 began with drying and crushing the ore to minus 2-cm size. Selective reduction of the nickel and cobalt with H2 and CO was conducted in multiple hearth furnaces at 700° to 760° C. The reduced ore was leached with an aerated solution of ammonium hydroxide and ammonium carbon- ate. The nickel and cobalt were leached out while the iron remained insoluble. This process, commonly called the Caron process, was used in Australia and the Philippines. Only about 50 pet of the cobalt and 75 to 80 pet of the nickel were recovered. The Bureau's reduction roast-ammoniacal leach process was tested on laterite ores (30). Crushed pyrite was added to the ore followed by multiple-hearth roasting at 525° C with pure CO gas. The nickel and iron combined to form ferronickel and the copper and cobalt were reduced to the individual metals. After cooling, the calcine was finely ground and leached with ammonium hydroxide, ammonium sul- fate, and oxygen to dissolve Co and Ni as ammine complexes. The process was demonstrated in a 500 lb/d integrated, continuous operation using an ore sample containing 0.2 pet Co and 0.97 pet Ni. About 85 pet of the Co and 90 pet of the Ni were recovered in the pregnant leach solution. However, with ore samples containing 0.05 to 0.1 pet Co (average, 0.08 pet), a 4. 5-st/d pilot plant test of this process only extracted 67 pet of the Co and 83 pet of the Ni (3_1> Laterites were also processed with a H2SO4 leach process (fig. 7) (_3). The ore containing 0. 08 pet Co and 1 pet Ni was slurried and pumped to leaching towers where it was leached with H2SO4 at 200° to 250° C under more than 500-psi pressure. The process dissolved cobalt, nickel, and magnesium, while the iron was hydrolyzed and precipitated. In labora- tory tests, about 85 to 90 pet of the cobalt and 90 to 95 pet of the nickel was extracted. California Nickel Corp. , San Francisco, CA completed pilot-plant work on this process and AMAX proposed to employ this process for the New Caledonia C0FREMI laterite operations to recover nickel and cobalt (2^). To minimize acid consumption, the process is generally restricted to low-magnesia ores (less than 5 pet MgO). The California and Oregon laterites contain 4. 3 to 7. 5 pet MgO as serpentine and chlorite. The solution refining process depends upon the extraction method. For the ammoniacal solutions, NH4H2PO4 was added to precipitate Mg and Mn as (Mg, Mn) NH4PO4, a fertilizer byproduct. After Laterite Crush 13 FeS 3 NH 4 H 2 P0 4 Dry Calcine - 525° C reduction Cool Grind Leach Liquid - solids separation Product liquor Impurity removal CO (fuel) Coke C0 2 CO CO producer Ash Makeup NH 4 OH, (NH 4 )2S0 4 , H 2 2 Leach solution NH 3 L (NH 4 )2S0 4 solution Recovery NH 3 and (NH 4 hS0 4 solution Solution H,0 Tailings wash and filter NH 4 (Mg, Mn)P0 4 NH 3 stripping H 2 Gangue NH, Ni-Cu solvent extraction Spent electrolyte recycles Absorber Raffinate Co solvent extraction Electrolytes + i i Electrowinning Spent electro- 11 I ■ eiectro- I I iyte ▼ ▼ recycle ▼ Raffinate recycle Ni Cu Co ZnS0 4 FIGURE 6.— Reduction roast-ammoniacal leach process for laterites. 14 H 2 S0 4 ► Ore Pressure leach H 2 S Purification Mn concentrate Acid <- Co+Ni precipitation Mn precipitation MgS0 4 H 2 crystallization MgO recovery MgO -> Zn, Cu Co+Ni Ammonia leach Ammonia ± L SX Stripping Ni electrowinning Ammonia stripping Ni Co leach Co electrowinning T Co FIGURE 7.— Sulfuric acid leach process for laterites. stripping excess ammonia from the solu- tion with steam, solvent extraction was used to remove both the Ni and the Cu. The nickel was stripped from the organic with spent nickel electrolyte followed by a electrowinning the nickel. The copper was then stripped from the organic with spent copper electrolyte followed by copper electrowinning. The Ni-Cu raffin- ate was neutralized precipitating the residual Cu, Mg, Zn, Mn, and Ni. The cobalt in the solution was reduced from Co 3+ to Co 2+ with cobalt shot and solvent extraction used to recover the cobalt. The organic was stripped cobalt electrolyte and the electrowon. This process strated in a pilot plant, metal only 65 pet of the with spent cobalt was was demon- From ore to cobalt and 83 pet of the nickel were recovered in the cathode products (31 ). For the H2SO4 extraction refining process, H 2 S was added to precipitate Cu and Zn as sulfides. Next the Ni and Co were precipitated with lime. Then the Ni-Co precipitate was leached with ammonia solution, followed by solvent extraction. The organic phase was stripped with spent nickel electrolyte and the nickel was electrowon. The raffinate was treated with steam to strip the ammonia and then the cobalt was precipitated. The cobalt precipitate was leached with spent cobalt electrolyte followed by cobalt electrowinning. A 5-st/d pilot plant was operated to demon- strate this technique. From ore to metal over 90 pet of the Co and Ni were 15 recovered (2^). Research on laterite processing appears to be complete, how- ever, the economics are not favorable. The residues have been evaluated and proved to be nontoxic to the environment (25), but development of these deposits will face significant opposition from environmentalists. MANGANESE SEA NODULES Manganese nodules are found over large areas of the ocean floor at depths of around 2,000 to 18,000 ft. They range from 0.25 to 3 in. in size and contain 25 to 30 pet Mn, 1.0 to 1.5 pet Ni, 0.5 to 1.0 pet Cu, and 0.25 pet Co (26). Large concentrations are found in the east central Pacific area. Only a small por- tion of nodule samples were taken from the U.S. Exclusive Economic Zone (EEZ ) located off the Hawaiian shore and other islands of the Trust Territory of the Pacific Islands and affiliated terri- tories. Estimates as high as 1 billion lb of Co have been made for this resource. The mineralogy is basically fine grained oxides mixed with layers of silicious gangue minerals (13). Although a number of problems remain to be solved, the mining technology proposed for deep- sea nodules has been a hydraulic air suction system or a continuous-line bucket system. Much of the prototype equipment has already been designed, patented, built, and tested (15). How- ever, international politics and many legal problems appear to have stalled the development of this resource. The fine grain mineral structure of the nodules precludes any effective benefici- ation except for physically separating the nodules from the bottom sediment by screening Extraction Methods A great deal of research has been conducted to develop methods for extract- ing the Mn, Cu, Ni, and Co from the nodules (12). Five of the most promising techniques for cobalt extraction are (1) gas reduction and ammoniacal leach, (2) cuprion ammoniacal leach, (3) high- temperature and high-pressure H 2 S0 4 leach, (4) reduction and HC1 leach, and (5) smelting and H2SO4 leach. Gas Reduction and Ammoniacal Leach Process First, the nodules were ground to 65- mesh size and dried followed by high- temperature (625° C) reduction of mangan- ese dioxide to manganese oxide by CO gas (fig. 8). After cooling to 40° C, the Cu, Ni, and Co were dissolved with an oxidizing ammoniacal ammonium carbonate leach (Caron process). About 90 pet of the Cu, 90 pet of the Ni, and 70 pet of the Co were dissolved into the solution. Cuprion Ammoniacal Leach Process The nodules were ground to 65-mesh size and reduced with cuprous ions (Cu + ) in a ammoniacal solution at 50° C (fig. 9). The manganese dioxide was reduced to manganese oxide. The cobalt was leached with a strong solution of ammonia and carbon monoxide. Next, the slurry was oxidized with air to oxidize the soluble ions and precipitate the iron. Only 50 pet of the Co was recovered, but 90 pet of the Cu and Ni was recovered (1_). This extraction technique is not appropriate for cobalt recovery. High-Temperature, High-Pressure H2SO4 Leach Process The nodules were ground to 65-mesh size and leached with H 2 S0 4 at 245° C and 500-psi pressure (fig. 10). The recovery of Co, Cu, and Ni was 90, 95, and 95 pet, respectively. Small amounts of Fe, Mn, and Zn were also dissolved. Reduction and HC1 Leach Process The nodules were ground to 65-mesh size and dried, followed by high-temperature (500° C) gaseous HC1 reduction of mangan- ese dioxide to manganese chloride (fig. 11). This reaction also produced chlor- ine gas that reacted with the other metal oxides, forming metal chloride salts. 16 Ni t Cu t Co ▲ Nodules 1 Carbon monoxide i i-> Electro- winning •4- r* Electro- winning Grinding Chemical reduction Ni stripping Cu stripping «- i i IS«« V SolUuun recycle Leaching - reduced pulp preparation Ni- ion Cu liquid exchange Co recovery i ' Waste containment Leaching - liquid-solid separation Metal so ► w - bearing ution 1 i k 1 L ■ •*. Ammonia recovery Makeup water T Mak amm i eu on P a i FIGURE 8.— Gas reduction and ammoniacal leach process for sea nodules. Ni t Cu t Co ▲ Nodules co 2 -co -► Electro- winning ■4- r> Electro- winning Grinding and drying Chemical reduction -► Ni stripping - Cu stripping i L i ' Leach Ni- ion Cu liquid exchange Co recovery ' r Waste containment Liquid - solid separation ■ w w Metal-bearing solution T H,S i i j . Ammonia recovery Makeup water T" Mak amm i eu on P a l FIGURE 9.— Cuprion ammoniacal leach process for sea nodules. FIGURE 10.— High-temperature, high-pressure H 2 S0 4 leach process for sea nodules. Grinding and drying Nodules Hydrochloric acid Hydro- chlorination - reduction H 2 Hydrochloric acid-chlorine recovery Hydrolysis Water storage Cu liquid ion exchange Leach liquid - solid separation Waste containment Electro- winning Co liquid ion exchange H 2 S _i_ Co recovery Ni liquid ion exchange Electro- winning Evaporation - crystalization Fused salt electrolysis FIGURE 11.— Reduction and HC1 leach process for sea nodules. 17 Cu Ni Grinding 4 Nodules k t t Electro- winning h. Electro- winning f H 2 J Leaching 4 H2SO4 Cu liquid ion exchange Ni liquid ion exchange PH adjustment ^ — l Neutralization fe ^ — ' t »- W i Liquid - solid separation T i i 1 ■ i 1 Ammonia recovery Co recovery 4 Y Waste containment k n n ^ 4 Lime ~ \*o t t t Makeup Makeup H 2 S ammonia water -> Cu Co -►Mn Water was sprayed to cool the product and caused the iron to form insoluble ferric hydroxide. After cooling, the mixture was leached with HC1 acid. Over 99 pet of the Co was extracted along with 96, 99, and 94 pet of the Cu, Ni, and Mn, respectively. 18 Smelting and H 2 S0 4 Leach Process Refining Process The nodules were dried, roasted with coke and CO gas at 725° C, and smelted with silica flux at 1,425° C (fig. 12). The slag, containing Mn, Fe, and Si0 2 , was removed and subsequently used for f erromanganese production. The manganese alloy containing the Fe, Cu, Ni, and Co was converted by adding sulfur and heat- ing to form a matte and an iron alloy. More silica flux was added and the matte- iron alloy was blown with air to lower the iron content of the matte to 5 pet. The finished matte was granulated, ground to 325-mesh size, and pressure leached with H2SO4 at 150 psi and 110° C. About 90 pet of the Co was dissolved in the solution along with 90 pet of the Cu and Ni. The Co recovery was uncommonly high for a matte smelting process. Refining of the leach solutions depends upon the solution type such as ammonia- cal, acid chloride, or acid sulfate. Ammoniacal Solution Refining For the gas reduction and ammoniacal leach process and cuprion process, the pregnant liquor passed through a solvent extraction circuit where the copper, nickel, and some ammonia were removed. Most of the ammonia was removed from the organic phase by washing with a weak aqueous ammonia solution. The organic phase was cleansed of the residual ammonia with a slightly acidic ammonium sulfate solution. The nickel was care- fully stripped from the organic phase using spent nickel electrolyte containing Reducing gas Chemical Grindii ■tg and < Nodules reduction drying 1 ' Electric furnace smelting Ferro- manganese reduction Ferromanganese 3 Gyr. r i »sum 2 > 1 3 r Cu t Ni t Oxidizing sulfiding Waste treatment Electro- winning Electro- winning •w i — w «— ' r pH adjustment Cu liquid ion exchange Neutralization Ni liquid ion exchange 1 1 Leaching- liquid - solid separation 4- z 1 r 1 Ammonia recovery Co recovery ^ Waste -* CO ntai nment T Co H 2 S FIGURE 12.— Smelting and H 2 S0 4 leach process for sea nodules. 19 40 g/L H 2 S0 4 . Then the nickel was electrowon from the nickel-rich electro- lyte. The copper was stripped from the nickel-free organic phase using a spent copper electrolyte containing 160 g/L H 2 S0 4 . The copper was electrowon from the copper-rich electrolyte and the metal-free organic was recycled to the solvent extraction circuit. The LIX raffinate was treated with ammonium hydrosulf ide, which precipitated the Co along with residual Cu, Ni, and Zn. The sulfide mixture was pressure leached with air to preferentially dissolve the Ni and Co, leaving the Cu and Zn sulfides behind in the residue. After filtering off the residue, hydrogen sulfide was used to remove the residual Zn and Cu impurities. Then the solution was heated in an autoclave with hydrogen to reduce the Ni. The nickel powder was washed, dried, and briquetted. The cobalt sulfate solution was concentrated in an evaporator-crystallizer to precipi- tate the Co and residual Ni with ammonium sulfate. The salts were redissolved with a strong ammonia solution and oxidized with air, converting the Co 2+ to Co 3+ . The remaining nickel was precipitated with acid and removed from the solution. The nickel-free solution was reduced with hydrogen in an autoclave, producing cobalt powder, which was dried and briquetted. About 98 pet of the cobalt was recovered in the refining process. The overall cobalt recovery from nodule to cobalt powder was 89 pet. Acid Sulfate Refining The pregnant solution from the high- temperature, high-pressure H 2 S0 4 leach process and the smelting and H2SO4 leach process were refined by the same method. First, the excess acid was neutralized with limestone to lower the H2SO4 content to 0.5 g/L. After filtering off the precipitated gypsum, the solution passed through solvent extraction with LIX to transfer the copper to the organic phase. The copper was stripped with spent copper electrolyte containing 160 g/L H 2 S0 4 . Then the copper was electrowon from the copper electrolyte. The copper raffinate was adjusted with ammonia to a pH of 4. During neutralization the tank was also aerated, causing the Fe, Mn, Mg, and Al impurities to precipitate. After centri- fuging out the solids, solvent extraction collected the nickel in the organic phase. The nickel was stripped from the organic phase with spent nickel electro- lyte. Then the nickel was electrowon from the nickel-rich electrolyte. The nickel raffinate was purified by the method described earlier for the LIX raffinate in the ammoniacal refining method. The final product was briquetted cobalt metal powder. Again, 89 pet of the cobalt was recovered from the nodules. Acid Chloride Refining The pregnant liquor was passed through a solvent extraction circuit to remove the Cu. The organic phase was stripped with spent copper electrolyte containing 160 g/L H 2 S0 4 . Then the copper was electrowon from the copper-rich electro- lyte. The pH of the raffinate from the copper extraction was adjusted to 4 with NaOH. The cobalt and some of the mangan- ese were recovered by solvent extraction with TNOA. Both the cobalt and manganese were stripped from the organic and then the cobalt was precipitated with H 2 S. This precipitate was treated by the same method reported earlier for the ammonia- cal refining method, beginning with the pressure leaching of the mixed sulfide precipitates. The cobalt was recovered as a briquetted metal powder. About 99 pet of the cobalt was recovered from the nodules. The cobalt-free manganese solution went to the manganese recovery circuit. The cobalt-free raffinate was treated by solvent extraction to recover the nickel. The nickel was stripped from the organic phase with spent nickel electrolyte and recovered by electro- winning. The nickel-free raffinate was 20 combined with the cobalt-free raffinate and treated with H 2 S to precipitate any residual Co, Ni , or Cu as sulfides. The remaining solution was evaporated, crystallizing manganese chloride salt. The salt was dried and fed to a high- temperature (1,300° C) f used-salt elec- trolysis furnace where molten manganese metal was formed and periodically tapped. About 96, 99, and 94 pet of the Cu, Ni, and Mn were recovered from the nodules. Research on manganese sea nodule processing is complete. The residues were environmentally acceptable and the process waste waters blended safely with seawater (9_). At least five process options are available, but the economics are not favorable. At a price of $17/lb, the calculated rate of return on invest- ment was 4 to 6 pet. However, at least a 30-pct rate of return on investment is required to attract the necessary venture capital for sea nodule production. MANGANESE SEA CRUSTS Cobalt is also found in offshore manga- nese crust deposits located in the EEZ of Hawaii and Trust Territory of the Pacific Islands and affiliated territories. Cobalt-rich deposits on seamounts and slopes have been estimated to contain up to 25 billion lb of Co. The average cobalt content of these crusts is 1.0 pet and they also contain 0.5 pet Ni, and 15 to 25 pet Mn ((>)• These deposits, 1-in thick layers, present a problem for selective mining. Equipment is being developed to break up the crust layer and leave most of the substrate behind, but this technology is still being tested. The mineralogy of the Pacific sea crusts is very similar to the Pacific nodules. Some places contain both crust and nodules. Even with a good selective mining system, benef iciation will be needed to separate the crust minerals from the substrate. Preliminary flota- tion tests have recovered 92 pet of the cobalt in a crust mineral concentrate. This research is not complete. The extraction technology is expected to be identical to the sea nodule extrac- tion methods. This resource still needs exploration, because very few samples have been taken from U.S. EEZ. The mining and benef iciation technology needs to be developed. The extraction technol- ogy is well developed, but the overall costs from mining through cobalt process- ing will be high. DISCUSSION The full potential for cobalt produc- tion from domestic sources is at least 19.4 million lb of Co per year. With the offshore manganese deposits, the amount doubles. A summary of each deposit's characteristics, benef iciation, extrac- tion, and refining results is presented in table 1. Also, included is a six- level technological assessment of the research and development status for the benef iciation and cobalt processing of each ore. The technology for Blackbird, Madison Mine, iron ore pyrite, laterites, and manganese sea nodules is nearly complete- ly developed, but the economics are not favorable. Small technological improve- ments will have virtually no effect upon the economics of these operations. The most promising source of cobalt is the spent copper leach solution. With all the mining and benef iciating costs paid by the primary copper production, cobalt can be extracted at a relatively low cost. The Missouri lead belt ores are also promising. Siegenite benef iciation from the copper concentrate is well developed, however the technology for extracting and refining the cobalt from the siegenite concentrate has not been fully explored. A secure domestic processing plant is needed to produce cobalt from the siegenite concentrate. It appears that benef iciation can not increase the bulk sulfide grade from the lead mill tailings without losing recovery. Because of the dolomite gangue, cobalt processing technology for this concentrate will probably be limited to an ammoniacal leach. The 21 applicability of this technology needs to be established so that the technical feasibility and economics of cobalt recovery from this source can be evaluated. The pyrite from some primary copper operations also shows potential for sig- nificant cobalt recovery. The technology for benef iciation and cobalt processing still needs research, but without the mining cost burden, this cobalt source should be one step closer to favorable economic production of cobalt. Both the Duluth Gabbro and Yakobi Island deposits show potential for cobalt. The benef iciation technology for the Duluth Gabbro deposit is complete, but the cobalt processing technology needs to be established on the Co-Ni concentrate. Cobalt production from this deposit appears to be many years into the future. For each pound of Co produced from this deposit, 95 lb of Cu and 14 lb of Ni are also produced. The present market conditions for copper offer no incentive for exploiting this deposit for copper. The deposit on Yakobi Island is in a similar predicament. The benef iciation and processing technologies need to be explored. With 6 lb of Cu and 10 lb of Ni produced for each pound of cobalt, there is also little hope for development in the near future. Economic evaluation of cobalt deposits should help to focus research on the deposits with the best future for econom- ical production. A Bureau cobalt avail- ability study on cobalt from world resources indicated that the mining and benef iciation costs are a substantial portion of the operating costs (20). That is why byproduct recovery from present copper, lead, and zinc opera- tions, where the mining and benef iciation costs are already paid, have the highest potential for immediate cobalt produc- tion. Over 5 million lb of Co per year could be produced from these operations. Research on these sources should be given the highest priority. Even with virtual- ly unlimited offshore cobalt resources, the high mining and ore transportation costs associated with the offshore mining will postpone the development of these resources for many years (14). Under the present economic conditions cobalt could be made available within 2 to 3 yr only as a byproduct from present operations. 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