Weak metal markets last year saw many operations unashamedly following the owner's desperate plea to reduce costs - increase throughput - and do it now!". Such conditions are not the ideal background for metallurgists to approach management for large research and development budgets, nor for costly technical evaluations to be performed.
This is reflected in the. interest in top new technology as monitored annually by Mining Journal. In 1996/97, 3 of the top 12 positions went to extractive metallurgy developments (enhanced gold leaching, copper leaching and bioprocessing), with 4 other mineral processing technologies featuring. In the equivalent survey for 1998, only one technology (improved grinding efficiency) in either of these categories made the top 12, the list featuring several worthy, but less imaginative, products such as even larger trucks and new drills and loaders.
Of the newer technologies in operation, copper flash converting at Kenecott continued to show its promise and a second unit is planned as part of Southern Peru's $1.8 billion expansion. Cominco's Kivcet lead plant at Trail has shown that it can perform, after a succession of mechanical problems during ramp-up.
Most recent interest has centred around the hydrometallurgical processes being installed at the new Australian nickel operations. Acid leaching/SX is also being piloted for the proposed Skorpion oxide zinc deposit in Namibia. Unfortunately progress made in flowsheet development was minimal in 1998 for the large modernisations planned at plants in the FSU and central Africa, where political considerations have added to price difficulties in delaying investment.
Despite these cyclic problems, metallurgists have remained traditionally optimistic and much background work has continued. If conference attendances have been a little lower, the output of good published papers has remained high.
The future of extractive metallurgy has been reviewed by Habashi(1) who has also discussed the role of hydrometallurgy in solving the environmental problems of smelters(2). The main point made in the first paper is a prediction that major change in the next century will be a swing to pressure leaching by the copper industry, thereby alleviating S[O.sub.2] pollution. Hot S[O.sub.2] rich gases can also be reduced to elemental sulphur by injecting a reducing agent into the gases. In certain leaching processes elemental sulphur can also be formed. While the zinc industry has been the first to adopt such technology, a similar approach has been proposed for nickel.
Sohn and Ramachandran(3) have discussed developments and innovations in sulphide smelting of copper, nickel, lead, zinc and PGMs, and have reviewed current research in the field. The environmental issues are also briefly discussed along with future challenges for the industry. Further environmental matters associated with S[O.sub.2] off-gases have been discussed by Puricelli et al(4). The key process variables for evaluation and selection of pyroprocessing systems have also been discussed(5).
Bioleaching continues to command much attention and a useful review has been presented by Boseker(6) covering both sulphidic and non-sulphidic ores and concentrates. The author notes also that bioleaching has some potential for metal recovery and detoxification of industrial waste products, sewage sludge and soil contaminated with heavy metals.
The future directions for gold and base metal hydrometallurgical processing in Australia have been discussed by Houchin(7). Strategies for reducing ore processing costs include better in- field ore characterisation and selection, improved reagent measurement and management, greater emphasis on process control and modelling and better design of plant equipment. Technological development is also driven by environmental considerations. Potential growth areas for the treatment of complex ores are the use of in-situ leaching, biooxidation and resin in pulp.
An overview of the solvent extraction of some major metals has been given by Mackenzie(8). Common features of metals recovered by solvent extraction are identified, reagents and processes are discussed, together with operational and equipment issues including contactor design and choice, crud issues, choice of coalescers, mixer efficiency and removal of organics from the strip aqueous.
Recoflo ion-exchange technology has also been reviewed(9). This technology is particularly suited for hydrometallurgical applications because of its ability to handle more concentrated feed solutions, its smaller resin volume requirement and the recovery of concentrated metal product streams. Recoflo systems are used to separate strong acids from dissolved metal salts and have been tested or are in use to remove As, Bi and Sb at Bingham Canyon, Utah, Ni and Cu sulphates at Kidd Creek, Ontario, Mg and Mn from zinc electrolyte at Valleyfield, Quebec, Ni from cobalt electrolyte at Port Colborne, Ontario and Ni and Co from ammonium sulphate by-product streams.
1. Habashi, F. The Future of Extractive Metallurgy. Montreal'98. CIMM, 1998, Paper 028.pdf on CD.
2. Habashi, F. How can Hydrometallurgy Solve the Environmental Problems of Smelters? Proc. 4 th, Conf. Environmental and Mineral processing, held in Ostrava, Czech Republic. 25-27 June. 1998, ed. P. Fecko, Ostrava, Czech Republic: University of Ostrava Faculty of Geology, 1998, p. 119-124.
3. Sohn, H. Y. and Ramachandran, V. Advances in Sulphide Smelting: Technology, R & D and Education. Sulphide Smelting'98. eds., J. A. Asteljoki and R. L. Stephens, The Minerals. Metals and Materials Society, Warrendale. Pa., 1998, p. 3-37.
4. Puricelli, S. M.; Fries, R. M.; Watson, G. A. and Grendel. R. W. Treating those Dirty Dastardly, Stray S[O.sub.2] Offgases. Prepr. Soc. Min. Metall. Explor. No. 98-229, 1998.
5. Field, B. T.; Lindquist, W. E. and Euston, C. R. Selection of Pyroprocessing Systems: An Overview. Prepr. Soc. Min. Metall. Explor. No. 98-213, 1998
6. Boseker, K. Bioleaching. Proc. 4 th. Conf. Environmental and Mineral Processing, held in Ostrava, Czech Republic. 25-27 June, 1998. ed. P. Fecko, Ostrava, Czech Republic: University of Ostrava Faculty of Geology, 1998, p. 351-356.
7. Houchin, M. Future Directions for Gold and Base Metal Hydrometallurgical Processing in Australia. Australasian metals Symposium. Surfers Paradise. Queensland, 17-18 August, 1998. North Sydney, NSW: IIR Conferences. 1998, 13 pp.
8. Mackenzie. M. The Solvent Extraction of Some Major Metals: An Overview. Ibid, 37 pp.
9. Sheedy, M. Case Studies in Applying Recoflo Ion-Exchange Technology. J. Metals, October 1998, Vol. 50, no. 10, p. 66-69.
Prominent amongst recent conferences was that on sulphide smelting at the TMS annual meeting in San Antonio in February 1998. A range of papers from fundamental research to applied practice were given, the former including new insights into the comparisons between copper, lead and nickel thermodynamics(1), general sulphide smelting thermodynamics (2), the viscosity of iron silicate slags(3) and the reaction and dissolution behaviour of silica flux in copper smelting(4).
Copper smelter operations papers featured plant performance at Huelva(5), Ilo(6), El Paso(7), Hidalgo(8) and Magma(9) and the Saganoseki flash smelter(10) was featured at the TMS Orlando meeting (but is included in the San Antonio volume). The Atlantic Copper paper describes the smelter modernisation and expansion. Using an Outokumpu flash smelter and four Peirce Smith converters the smelter now produces some 300,000 t/y of copper. Improvements at the Southern Peru plant, where two reverberatory furnaces and seven PS converters, plus one Teniente converter are employed are also illustrated, current capacity being now 285,000 t/y of copper.
The Asarco plant has the first commercial CONTOP furnace, retrofitting of which has allowed for a wide variety of concentrates to be smelted while maintaining a consistent matte grade of around 58% Cu. The cyclone furnace has demonstrated low dust carry over and low copper loss in discard slag, all achieved at low capital cost and with a fast construction timetable. Phelps Dodge's focus at Hidalgo has been on plant maintenance and their successful programme of reliability-centred maintenance is described. The Kennecott Utah plant, which ramped-up to full production in 1997, is of particular interest as it covers the development of the Outokumpu flash converter furnace, the first such operating unit. At Nippon Mining and Metal's smelter, the paper focuses on the integration of two flash furnaces into one, with the consequent substantial cost reduction.
Copper converter practice at Tamano(11) and operational performance at Kosaka(12), Glogow(13), Guixi(14), Ronnskar(15) and Kidd Creek(16) were also covered, along with new developments at Refimet(17), Chino(18), Naoshima(19) and Horne(20) plus a number of papers on gas handling and sulphur fixation.
Hibi Kyodo Smelting at tamano has recently made a number of expansions, including extending the length of the PS converters. Modernisation at Kosaka includes modification of the flash furnace concentrate burner and control of the dust sulphurising reaction in the boiler to improve metallurgical performance. At KGHM, Glogow, impurity deportment across the shaft furnace and flash furnace is described, as it is at the Jiangxi plant at Guixi, where soda ash and lime are injected into the converters to enhance impurity removal. Impurity handling is also a key feature at Boliden's Ronnskar smelter where a wide range of primary and secondary sources are treated.
Converter upgrades at Falconbridge's Kidd Creek smelter system include roof modifications with a novel brick/cooler suspension system. Refimet at La Negra, Chile, has now completed plans for a third stage expansion to take output beyond the present 160,000 t/y level. At Phelps Dodge Chino smelter, efforts have been concentrated on increasing productivity and environmental performance, a policy also followed by Mitsubishi with their process at Naoshima. Noranda at the Horne is similarly reducing emissions with the construction of a new Noranda converter.
A general review of Doe Run's La Oroya developments has been published(21). Eventually total base metal output will approach 300,000 t/y, with lead, zinc, silver and gold production and a range of minor metal by-products as well as copper.
An equally significant number of papers on copper pyrometallurgy also appeared in the Environment and Innovation in Mining and Mineral technology conference held in Chile in May(22). Slag cleaning was a particular topic of interest with a general paper on slag cleaning with special reference to new smelting methods and more specific papers were presented on such topics as reduction in copper losses in slags from new copper smelting processes, removal of As and Sb from molten copper by sodium-based slags and their process implications, the cleaning of slags after autogeneous copper smelting by injection of reduction gases near the electrodes of the electric furnace, the use of reducing-sulphidising gas mixes for slag cleaning and the optimisation of fuel combustion in a Teniente slag cleaning furnace. The development of the Vanyukov furnace at Norilsk was described as was the development of autogeneous smelting of copper concentrates at the Jinchuan smelter in China. Environmental and productivity improvements at the Tamano smelter in Japan are discussed as is a real-time dynamic simulation model for the flash smelting operation at Chagres.
Two papers have described the renewal of the Mt. Isa copper smelter and the operation of the Isasmelt unit(23,24). Upgrading of the smelter will allow the production of 250,000 t/y of copper. This expansion in throughput has been achieved by using an additional 525 t/d of oxygen in the primary smelting unit. To meet this new capacity the fluosolids roaster/reverberatory furnace has been removed and a fourth converter added. The anode furnaces have also been extended and the casting wheel changed to a double pour wheel to increase casting rate. As a result of the adoption of the Isasmelt-only flowsheet, higher levels of arsenic elimination from matte have been achieved and As levels in the anodes are expected to decrease by 25%.
The combination of flash smelting and flash converting has been discussed by Hanniala(25). The flash converting process controls the sulphur content of the blister copper by means of the oxygen:matte ratio and the heat balance. Because of low gas volumes leaving both the smelting and converting furnaces, cleaning costs are low and the degree of sulphur fixation is very high, in line with the 95% minimum now in force in many countries and should exceed 99% with the newest flash technology. A model for simulating the oxidation reactions of chalcopyrite particles in the Outokumpu flash smelting furnace shaft has been developed(26).
In the Watanabe Medal Award lecture the changeover from two flash furnaces to only one at the Saganoseki smelter in Japan has been described(27). The modifications to the pneumatic concentrate drying system, concentrate feeding system, concentrate burner and furnace cooling system to double the throughput are discussed. The new system, which started in March 1996 has achieved a considerable improvement in productivity and reduction in costs.
New anti-pollution legislation in Chile has resulted in the redesigning of the Ventanas smelter. The reverberatory furnace has been replaced by a Teniente converter and a circular electric slag cleaning furnace designed to treat all the slags generated by the converter. Performance is reported to have exceeded all expectations and a grade of [less than]0.8% Cu in the final slag has been achieved at a specific energy consumption of 150 kWh/t(28).
In copper hydrometallurgy, bioleaching continues to be of interest. The application of the silver catalysed IBES process to a Spanish copper-zinc sulphide concentrate has been further studied(29, 30). In the first paper the biooxidation of the ferrous iron generated in the chemical leaching stage and the recovery of silver used as a catalyst have been examined. Effective biooxidation of the ferrous iron has been demonstrated both in static batch and packed bed reactors. It is possible to recover 100% of the silver added as catalyst plus 93% of the silver present in the concentrate by leaching with a 200 g/l NaCl-0.5 M [H.sub.2]S[O.sub.4] solution at 90 [degrees] C for 2 hours.
The operating conditions for a chalcopyrite-sphalerite concentrate from Rio Tinto were then selected using the data generated in the batch tests. From these a flowsheet has been devised for a pilot plant for the treatment of 50 kg/h of the concentrate, including the solvent extraction and electrowinning steps for zinc and copper, elemental sulphur recovery, silver recovery and effluent purification. A preliminary economic evaluation of the process indicates that for an industrial plant of 100,000 t/y, the direct operating cost would amount to US$141/t; the products, copper and zinc, would be worth US$450/t and the value of the by-products (mainly silver and elemental sulphur) was estimated at US$29/t.
The Chinese have also reported work on silver catalysed bacterial leaching of copper ore. In work with a mainly chalcopyrite ore from Dexing in Jiangxi it was found possible to increase copper recovery in 21 days from 28.2% to 77.5% by the addition of 0.16 g Ag/Kg ore. The catalytic action of the silver is discussed(31).
In other work, the dissolution of copper from ores and concentrates with T. ferrooxidans in cultures grown from mine waters or from imported dry inoculum has been studied(32) using Turkish chalcopyrite samples. A study has also been made on the biohydrometallurgical processing of Chilean ores in order to prevent As volatilisation and the associated environmental problems. The kinetics of chemical and bacterial leaching at 30 and 70 [degrees] C with t. ferrooxidans and Sulpholobus BC have been determined. Both organisms had a significant effect on the kinetics but the enargite was found to be very refractory to leaching. Silver was found to be an effective catalyst and slightly increased the dissolution rate at 30 [degrees] C. The precipitation of ferric arsenate and arsenite was also studied. A flowsheet for the process has been proposed(33).
The development of additives that improve heap leaching has been discussed in two papers. Thus the use of fluorosurfactants to give better wetting of the copper-bearing minerals has been described by Pavez and Olave(34) while Pincheira et al(35) discuss the development of Dearcodox, a reagent for use in the leaching, solvent extraction and electrowinning processes. Both papers claim increased copper recovery while the use of Dearcodox is also claimed to reduce chlorine levels prior to electrowinning.
Copper leaching at Gunpowder, Queensland has been described(36). The operation has been recently purchased by Aberfoyle Ltd and expansion is underway. An innovative high recovery tank leaching process, including a low temperature, low pressure autoclave, has been adopted and the copper output is being expanded to 45,000 t/y. The combination of expanded throughput and high metal recoveries will make Gunpowder a cost competitive producer.
The effect of ore character on the management of copper heap leaching has been described by Miller(37). The success and failure of heap leaching operations has been reviewed by Pyper et al(38) who conclude that the largest source of risk is often in the area of management and operations with inadequate care over percolation issues or quality control testing.
Pressure leaching of copper sulphide concentrates is being tested on a continuous, fully integrated, 1/500 scale demonstration plant by Cominco Engineering Services Ltd (CESL)(39). Preceding leaching the concentrate is reground to about 95% -325 mesh. Sulphur oxidation, with the inclusion of pyrite is 15% overall. Copper recoveries are 97-98% for high grade (40%) and 96% for low grade (28%) concentrates. For a full-scale plant treating 40% concentrates and producing 200,000 t/y copper via solvent extraction and electrowinning the capital cost is estimated to be US$225 million with an operating cost of US$0.09/lb Cu.
Alternative processes to the existing smelting refining routes for treatment of chalcopyrite have been reviewed(40). These include hydrometallurgy in sulphate, chloride and ammoniacal media. Roasting as a pretreatment prior to leaching is also considered. Ferric chloride leaching of mechanically activated chalcopyrite has been studied by Maurice and Hawk(41). Thus chalcopyrite was autogeneously milled in a horizontal mill and leaching was carried out using a 5 M chloride leach solution. Leaching of as-received concentrate resulted in extractions of 75% in 5 hours while over 95% extraction was achieved in 3 hours with mechanically activated concentrate. The contributions of the increased surface area and the deformed structure were correlated with leaching kinetics.
Solvent extraction remains a strong feature. Electrolyte filters, coalescers and a crud treatment plant are featured in a paper on the technological contributions of Chuquicamata division to the development of solvent extraction in the copper industry(42). Chloride control in the SX-EW plants of Mantos Blancos and Mantoverde has been discussed by Tapia and Kelley(43). The main form of chloride contamination is by entrainment in the organic phase. The use of fixed-bed coalescers gives an aqueous recovery between 500-1,000 ppm in summer and up to 5,000 ppm in winter leaving a final entrainment level of 50-100 ppm. This allows the Mantos Blancos plant to work with only a small bleed while no bleed is required at Mantoverde.
The use of solvent extraction to control Sb and Bi levels in copper refinery electrolytes has been described by Kim et al(44) with special reference to the San Manuel refinery in Arizona. Good results were achieved using the mono alkyl phosphoric acid reagent Acorga SBX-50 in conjunction with a patented high chloride strip.
Use of the Diphonix ion exchange resin for iron control in copper electrowinning circuits has been discussed by Shaw et al(45). This resin is very selective for ferric iron over copper and cobalt. Thus bleeding requirements for iron control are reduced by 90%. Stripping of iron from the resin is achieved reductively using sulphurous acid. The first commercial plant using this technology is at Cananea, Mexico where iron removal rates of [less than]1 t/d have been achieved.
New lead anode materials for metal electrowinning have been described(46). Advances in cell design for base metal electrowinning have been described by Gunn(47). This includes the recently developed EMEW high flowrate, high current density cell. The development of high current electrowinning for copper has been described by Anastasijevic et al(48). The process is based on bipolar electrodes and activated titanium electrodes. The technology has been verified by pilot tests under industrial conditions. The potential use of sulphur dioxide as an anodic depolariser in copper electrowinning has been studied by Dawson et al(49). Glassy carbon is the preferred substrate for S[O.sub.2] oxidation. Further information can be found in another paper by the same authors(50).
A risk-free electrowinning process has been developed in Chile which avoids arsine generation and which claims optimum yield of both copper and arsenic(51). The process has been installed at the Ventanas copper refinery. The arsenic product is a soft solid, easily removable and suitable for recycling when melted.
The presence of nickel in electrorefinery electrolytes has received attention. Thus the removal of nickel by precipitation from a bleed using aqueous ammonia has been studied as has partial evaporation, the latter being found, not surprisingly, to be very effective(52). In a study of the influence of nickel on the electrorefining of copper in a secondary smelter it was found that passivation effects occurred during electrorefining with high Ni concentrations(53). Together with Sb, O and Cu, nickel forms copper glimmer, a floating slime which creates low current efficiencies because of short circuiting and contamination of the copper cathodes. Ni ions in solution also decrease the solution conductivity. Further work by this team has recorded a number observed phenomena during the copper refining process at Brixlegg in Austria(54).
Thus the weight of copper deposited depended on whether the current flow was from the air side or the mould side of the anode. Passivation effects were due to nickel and copper sulphates and secondary oxides or sometimes lead sulphate or oxide if the anode lead content was particularly high. Nickel at concentrations in the copper matrix of below 0.25-0.3% went into solution during electrolysis, influencing the copper content and electrical conductivity of the electrolyte while at greater concentrations it formed NiO octahedra in the anodes and slimes.
Increasing production in copper electrorefining by means of the Isa process has been discussed by Sako et al(55). Simulation of the electrorefining process has shown that it reduces metal inventory, required no steam to heat the electrolyte at current densities higher than 320 A/[m.sup.2] and improved productivity with the elimination of the complex starter sheet manufacturing process.
1. Yazawa, A. and Nakazawa, S. Comparisons between Copper, Lead and Nickel Smelting Processes from Thermodynamic Viewpoints. Sulphide Smelting'98, eds., J. A. Asteljoki and R. L. Stephens, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.39-48.
2. Mindin, V.; Kiknadze, N. and Mindin, Y. Thermodynamic Investigations of Pyrometallurgy of Sulfide Concentrates. Ibid, p, 125-134.
3. Vartainen, A. Viscosity of Iron-Silicate Slags at Copper Smelting Conditions. Ibid, p. 363-372
4. Fagerlund, K.; Palmu, L. and Jalkanen, H. Experimental Study on the Reaction and Dissolution Behaviour of Silica Flux in Copper Smelting. Ibid, p.375-386.
5. Barrios, P.; Contreras, J, and Palacios, P. Recent Operations at the Atlantic Copper Smelter in Huelva. Ibid, p.135-146.
6. Torres, W. E. Current Teniente Converter Practice at the SPL llo Smelter. Ibid, p. 147-158.
7. Brueggermann, M. and Caba, E. Operation of the CONTOP Process at the Asarco El Paso Smelter. Ibid, p.158-166.
8. Marquez, R. B. and Burgess, B. Introduction of Reliability Centered Maintenance in a Copper Smelter. Ibid, p.197-204.
9. Newman, C. J.; Probert, T. I. And Weddick, A. J. Kennecott Utah Copper Smelter Modernisation. Ibid, p.201-218.
10. Suzuki, et al. Productivity Increase in Flash Smelting Furnace Operation at Saganoseki Smelter and Refinery. Ibid, p.587-596.
11. Maruyama, T.; Saito, T. and Kato, M. Improvements of the Converter's Operation at Tamano Smelter. Ibid, p.219-226.
12. Maeda, Y.: Inoue, H. Hoshikawa, Y and Shirasawa, T. Current Operation of Kosaka Smelter. Ibid, p.305-314.
13. Czernecki, J.; Smieszek, Z.; Gizicki, S.; Dobrzanski, J. and Warmuz, M. Problems with Elimination of the Main Impurities in the KGHM Polska Miedz. S.A. Copper Concentrates from the copper Production Cycle (Shaft Furnace) process, Direct Blister Smelting in a Flash Furnace). Ibid, p.315-344.
14. Zeping, Y. Impurity Distribution and Removal practice in the Copper Smelting process at Guixi Smelter. Ibid, p.345-352.
15. Lehner, T. and Vikdahl, A. Integrated Recycling of NonFerrous Metals at Boliden LTD Ronnskar Smelter. Ibid, p.353-362.
16. Macrae, A.; Wallgren, M.; Wasmund, B.; Lenz, J.; Majumdar, A.; Zuliani, P. and Elvestad, P. Converting Furnace Upgrades at the Kidd Creek Metallurgical Division Copper Smelter. Ibid, p.387-400.
17. Campos, R.; Miranda, S. and Smith, T. J. A. Refimet Smelter Operation and Development. Ibid, p.519-534.
18. King, M. J. and Phipps, R. D. Process improvements at the Phelps Dodge Chino Smelter. Ibid, p.535-548.
19. Oshima, E.; Igarashi, T.; Hasegawa, N. and Kiyotani, K. Naoshima Smelter Operation Present and Future. Ibid, p.549-558.
20. Boisvert, M.; Janneteau, G.; Landry, J. P.; Levac, C. A.; Perron, D. McGlynn, F.; Zamalloa, M. and Porretta, F. Design and Construction of the Noranda Converter at the Horne Smelter. Ibid, p.569-586.
21. Chadwick, J. La Oroya. Mining Magazine, February 1998, Vol. 178, No. 2, p.106-111.
22. Environment and Innovation in Mining and Mineral Technology. Proceedings IVth International Conference Clean Technologies for the Mining Industry, eds. M. A. Sanchez, F. Vergara, and S. H Castro, University of Concepcion, Chile, 1998, Vol. 11.
23. Player, R. Renewal of the Copper Smelter at Mount Isa. The Mining Cycle. Proc. AusIMM 1998 Annual Conference, Australasian IMM Publication Series no. 2/98, 1998, p. 391-393.
24. Edwards, J. S. Isasmelt: A 250,000 t/y Copper Smelting Furnace. Ibid, p.395-400.
25. Hanniala, P. Flash Converting Copper. Mining Environmental Management, September 1998, Vol.6, no. 5, p.13-16.
26. Ahokainen, T. and Jokilaakso, A. Numerical Simulation of the Outokumpu Flash Smelting Furnace Reaction Shaft. Can. Metall Q., July-October, 1998, Vol. 37, no. 3-4, p. 275-283.
27. Ishikawa, M. High-intensity Operation and Productivity Increase in Flash Smelting Furnace Operation at Saganoseki Smelter and Refinery. Metall. Rev. MMIJ, December 1998, Vol. 15 no. 2, p.139-158.
28 Schwarze Dintrans, H.; Moreno, A. and Sanchez, G. Cleaning Teniente Converter Slags in an Electric Furnace at the Ventanas Smelter. Minerales, July-September 1998, Vol. 53, no. 223, p.17-22.
29. Palencia, I.; Romero, R. and Carranza, F. Silver Catalysed IBES Process: Application to a Spanish Copper-Zinc Sulphide Concentrate. Part 2. Biooxidation of the Ferrous Iron and Catalyst Recovery. Hydrometallurgy, March 1998, Vol. 48, no 1, p.101-112.
30. Romero, R.; Palencia, I. and Carranza, F. Silver Catalysed IBES Process: Application to a Spanish Copper-Zinc Sulphide Concentrate. Part 3. Selection of the Operational parameters for a Continuous Pilot Plant. Ibid, June 1998, Vol. 49, no. 1-2. p.75-86.
31. Qiu, G.; Wang, J.; Zhong, K. and Wang, D. Application of silver to Catalysing of Bacterial Leaching of Copper Ore. Min. Metall. Engng., September 1998, Vol. 18, p.22-26.
32. Dogan, M. Z. and Yuce, A. E. Dissolution of Copper Ore and Concentrate by the Application of Thiobacillus Ferrooxidans. Proceedings 4th Conference Environmental and Mineral processing, held in Ostrava, Czech Republic, 25-27 June, 1998, ed. P. Fecko, Ostrava, Czech Republic: University of Ostrava Faculty of Geology, 1998, p.685-689.
33. Weirtz, J.; Espinoza, S.; Mendoza, C.; Ruiz, M. A.; Huepni, E.; Casas,. J. M. and Escobar, B. Preliminary Study of (Bio)leaching Process for Enargite Bearing Ores and Concentrates. Environment and Innovation in Mining and Mineral Technology. Proceedings IVth International Conference Clean Technologies for the Mining Industry, eds. M. A. Sanchez, F. Vergara. and S. H. Castro, University of Concepcion, Chile, 1998, Vol. 1, p.141-153.
34. Pavez, E. A. and Olave, X. G. The Use of Fluorosurfactants in copper and Gold leaching Applications. Environment and Innovation in Mining and Mineral Technology. Proc. IV th. Int. Conf. Clean Technologies for the Mining Industry, eds. M. A. Sanchez, F. Vergara, and S. H. Castro, University of Concepcion, Chile, 1998, Vol. 11, p.515-521.
35. Pincheira, A.; Reghezza, A.; Vergara, J.; Arcos, F.; Cifuentes, R. and Vergara, C. Dearcodox: Development of a Reagent that Favours the Production, Quality and the Environment in Hydrometallurgical Systems. Ibid, p.501-514.
36. Brock, J. G.; Chomley, J. C. and Richmond, G. D. Copper Leaching at Gunpowder. The Mining Cycle. Proc. AusIMM 1998 Ann. Conf., Australasian IMM Publication Series no. 2/98, 1998, p.281-284.
37. Miller, G. Ore Character Effects on Copper Heap Leach Management. Mine to Mill 1998, Australasian Institute of Mining and Metallurgy, 1998, Australian IMM Publication Series, no. 4/98, p.121-125.
38. Pyper, R.; Morrissey, C. and Middleditch, L. Heap Leaching: Simple Why Not Successful? Australian Metals Syrup., IIR Conferences, North Sydney, NSW, 1998, 11 pp.
39. Anon. Hydrometallurgical Copper Processing from Cominco. EMJ, June 1998, p.52-WW.
40. Prasad, S. and Pandey, B.D. Alternative Processes for Treatment of Chalcopyrite: A Review. Mineral Engineering, August 1998, Vol. 11, no. 8. p.763-781.
41. Maurice, D. and Hawk, J. A. Ferric Chloride leaching of Mechanically Activated Chalcopyrite. Hydrometallurgy, June 1998, Vol. 49, no. 1-2, p.103-123.
42. Pincheira, A. A.; Reghezza, I. A., Vergara, C. J. and Matta, V. J. Latin American Perspectives: Exploration, Mining and Processing. Ed. O. A. Bascur, Society for Mining, Metallurgy and Exploration, Littleton, Colorado, 1998, p.229-235.
43. Tapia, G. and Kelley, R. J. Chloride Control in SX-EW Plants. Ibid, p.259-265.
44. Kim, D. K.; Leese, T. A.; Neild, M. P.; Saito, B. R.; Young, S. K. and Weidner, C. J. Use of Solvent Extraction to Remove Bismuth and Antimony from Copper Electrolyte at the San Manuel Refinery. EPD 1998, ed. B. Mishra, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.301-315.
45. Shaw, D. R.; Wallace, S.; Dreisinger, D. B. and Gula, M. J. iron Control in Copper Electrowinning Streams Utilising Diphonix, a Selective Ion Exchange Resin. Environment and Innovation in Mining and Mineral Technology. Proceedings Ivth International Conference Clean Technologies for the Mining Industry, eds. M. A. Sanchez, F. Vergara, and S. H. Castro, University of Concepcion, Chile, 1998, Vol. II, p.555-567.
46. Stelter, M.; Hein, K. and Bauer, I. New Lead Anode Materials for Metal Electrowinning. Erzmetall, 1998, Vol. 51, no. 4, p.281-289.
47. Gunn, M. J. Electrowinning of Base Metals: Advances in Cell Design. Australian Metals Syrup., IIR Conferences, North Sydney, NSW, 1998, 18 pp.
48. Anastasijevic, N.; Laibach, S.; Nepper, J. P. and Werner, D. High-Current Density Electrolysis: Stages in the Development of a Metal Electrowinning Process. Erzmetall, 1988, Vol. 51, no. 11, p.733-742.
49. Dawson, J. N; Singh, P. and Hefter, G. T. The Potential Use of Sulphur Dioxide as an Anodic Depolariser in Copper Electrowinning. Proceedings AUSIMM Annual Meeting, Australasian IMM Publication series no. 2/98, 1998, p.319-325.
50. Dawson, J. N; Singh, P. and Hefter, G. T. The Effect of Run-Of-Mine Constituents on the Anodic Depolarisation of Sulphur Dioxide during Copper Electrowinning. Mine to Mill 1998, Australasian Institute of Mining and Metallurgy, 1998, Australian IMM Publication Series, no. 4/98, p.69-73.
51. Gallegos, A.; Martinez, I. and Olmos, C. New Electrolytic Cleaning Process Developed by CIMM and ENAMI. Environment and innovation in Mining and Mineral Technology. Proceedings IVth International Conference Clean Technologies for the Mining Industry, eds. M. A. Sanchez, F. Vergara, and S.H. Castro, University of Concepcion, Chile, 1998, Vol. II, p.817-825.
52. Nyirenda, R. L. and Phiri, W. S. The Removal of Nickel from Copper Electrorefining Bleed-Off Electrolyte. Mineral Engineering, January 1998, Vol.11, no. 1, p.23-37.
53. Anzinger, A.; Wallner, J, and Wobking. Influence of Nickel on Electrolytical Refining of Copper in a Secondary Smelter, BHM Berg-u. Huttenm. Mh., 1998, Vol. 143, no, 3, p.82-85.
54. Anzinger, A. and Wobking, H. Phenomena Observed during Electrolytical Copper Refining at Montanwerke Brixlegg AG. Erzmetall, 1998, Vol. 51, no. 11, p.743-749.
55. Sako, Y.; Nishimura, Y.; Kitahara; K. and Yukimasa, T. An Examination of the Effective Method on Increasing Production Capacity of Copper Electrorefining. Metall. Rev. MMIJ. December 1998, Vol. 15, no. 2, p. 175-183.
In the EU there are ten zinc electrolysis plants and four lead-zinc shaft furnaces for reduction of mixed concentrates by the Imperial Smelting process, plus five attendant plants for zinc production by fractional distillation. The strengths and weaknesses of the processes and the future demands of the industry have been discussed by Kruger(1).
The major conference concerned with zinc during the year under review was Zinc and Lead Processing(2), held in Calgary, Canada in August. Highlights of this excellent conference include discussions of plant operations at Cominco's Trail operation, IMMSA's zinc refinery at San Luis Potosi in Mexico, the Cinkur zinc and lead metal production operations, the sixteen year operation at the Cajamarquilla zinc refinery, a description of improvements at Kokkola Zinc, recent operation of the Haematite Process at the Iijima Zinc Refinery, improvements in zinc recovery in the leaching circuit at the Met-Mex Penoles Electrolytic Zinc plant, the conversion of the Pasminco Hobart smelter to the paragoethite process, the automation of the zinc stripping process at Kidd Creek zinc electrowinning plant and the operation of the HBM&S zinc pressure leach plant.
Emerging pyrometallurgical processes for zinc and lead recovery from zinc-bearing waste materials and the use of the ISP for direct zinc recovery from secondaries are discussed and there are several papers on off-gas treatment from primary and secondary zinc operations, including a paper on the Topsoe process for desulphurisation of off-gas from Zn and Pb smelters. Under the banner of emerging technologies the commercialisation of the EZINEX process is describes as is the flash roasting of zinc concentrates and leach residues using a Torbed reactor.
Other papers of interest under this topic head include fluoride control in the Dynatec zinc pressure leach process, the use of coal as an additive in the Dynatec process and the hydrometallurgical recovery of zinc from sulphide ores and concentrates by bacterially assisted leaching with ferric sulphate followed by zinc solvent extraction with the Zeneca reagent DS6010 (formerly DS5869) and electrowinning. The current zinc cellhouse practice at Union Miniere and the state-of-the-art and feasibility for upgrading the electrowinning cellhouse of Asturiana de Zinc in Spain are also described. Other papers on secondary zinc recovery and environmental matters associated with zinc production are also included.
Van Os(3) has given an update on Pasminco and the upgrades to its smelters at Hobart, Tasmania, Port Pirie, South Australia and Cockle Creek, New South Wales. The company also owns the Budel zinc smelter in Holland. Through these upgrades significant environmental improvements have been made as well as efficiency gains yielding productivity improvements of 70%.
The Warner process has been presented and discussed yet again with particular reference to its possible use for exploiting ores arising in remote locations such as the McArthur River mine. It is proposed that bulk flotation concentrates are smelted in the melt circulation reactor using oxygen. It is claimed that by smelting the high energy pyrite and reacting the evolved S[O.sub.2] with dolomite available nearby a large amount of electricity can be generated, enough to supply all the requirements of the smelter itself, an air separation plant and a facility for compressing and liquefying C[O.sub.2] which is disposed of for off-shore use in enhanced oil recovery or by other means. The reacted dolomite fixes the sulphur as an inert and environmentally acceptable solid residue. It is concluded that treatment by these means leads to virtually no gas emission(4).
The BUKA zinc process has been discussed by Buckett et al(2, 5, 6). This process, which has been developed to produce high quality zinc from a wide range of concentrates but principally for BUKA Minerals Ltd Lady Loretta project in Queensland, Australia, consists of three stages. These are fuming, fume separation and zinc refining. Fuming of concentrates or high-grade ores can be carried out in a variety of furnaces including the Sirosmelt top-submerged lance furnace.
Fume is separated by dissolution into a leach solution thus leaving the lead and silver as a residue which is easily smelted. The pregnant leach liquor is purified and a high-grade zinc oxide produced. This is dissolved and high-grade zinc electrowon. High metal recoveries to concentrate, lower costs, increased electrical efficiency, increased metal recovery and reduced environmental disposal costs are claimed. This approach would certainly avoid the problem of iron disposal from the roast-leach-electrowin zinc plants. Test rig and laboratory bench testing of fuming and fume separation was carried out during 1997 and further work is planned.
It has been found that, in the processing of Waelz process slags, an increase in the CaO:Si[O.sub.2] silicate ratio from 0.3 to 1.0 improved zinc extraction by 2.5 to 4.1 times, during the start of process operations(7). The addition of CaO also had a positive effect on the activity of iron in the slags which contributed to zinc removal. This in turn reduced consumption of energy resources such as coke and oxygen.
Production of zinc from secondary material is increasing and the current situation and available technology has been discussed by Schneider and Schwab(8). Spouted bed electrowinning for the recovery of zinc from scrap galvanised steel has been described by Roy et al(2) and Jiricny et al(9). Laboratory scale results suggest that the recycling of galvanised scrap could be significantly improved by the use of this technology and pilot scale studies are being carried out.
Biohydrometallurgical processing of zinc sulphide ores and concentrates continued to be a feature during the period of this review. Thus the kinetics of sphalerite leaching with thermophilic bacteria Acidianus brierleyi has been studied by Konishi et al(10) at 65 [degrees] C and pH 2.0. The presence of iron was shown to be detrimental to the process probably because of precipitation of jarosites In iron-free solution the rate of bioleaching was shown to be about seven times that with the common leaching mesophile, Thiobacillus ferrooxidans.
The application of the IBES process to the treatment of zinc sulphide concentrates has been described by Carranza and Iglesias(11). In this process the chemical oxidation of the sulphide with ferric iron is separated from the bacterial regeneration of the lixiviant. In this process copper ions have been shown to have a negative effect on the leaching rate which was stronger at 70 than at 80 [degrees] C. Because of this it is recommended that for zinc concentrates with low copper sulphides the IBES process flowsheet should be modified so that copper is removed by cementation after the biooxidation stage.
A selective pressure oxidation leaching process, capable of extracting zinc from concentrates containing both sphalerite and chalcopyrite has been reported(12). While zinc was extracted selectively from several Cu/Zn concentrates using various combinations of temperature and oxygen concentration, it was found that the presence of only 5% galena retarded the dissolution of sphalerite by up to 13%. On the other hand a 10% addition by weight of pyrite increased zinc extraction while decreasing copper extraction. Increased concentrations of pyrite, meanwhile, gave increased copper extraction and thus decreased zinc selectivity.
A new process for the production of a zinc concentrate from a zinc leach residue while simultaneously recovering lead and silver has been reported (13). The process involves the treatment of a silver concentrate produced by flotation of the zinc leach residue by a sulphuric acid leach at moderate temperatures to partially dissolve the zinc and iron. The residue from this leach is treated with a brine leach to recover the silver and lead which are then cemented out. The brine leach residue represents a high grade zinc concentrate which can be recycled to the roaster for zinc recovery.
Kvaerner Metals has been awarded a US$148 million contract by CalEnergy Minerals LLC for the design and construction of a zinc recovery project in the Imperial Valley of California. In this project, brine that has been used for power generation will be treated by ion exchange before passing to the re-injection wells.
Through reverse osmosis this plant will produce a concentrated zinc solution which will be treated by a standard solvent extraction and electrowinning plant to produce zinc cathodes. The operation is expected to yield 30,000 t/y zinc and commissioning should begin in mid 2000(14).
The recovery of zinc from jarosite residues by chlorination has been studied by Tailoka and Fray(15). The jarosite was initially dried and heat-treated on its own or with zinc ferrite at 450 or 600 [degrees] C and leached in hot water to remove almost all the sodium as sodium sulphate. The filtered residue was chlorinated in air either with pure chlorine or indirectly with HCl liberated by the combustion of scrap polyvinyl chloride. More than 90% of the zinc was removed with iron carryover as ferrous chloride of 1-3 wt%. A flowsheet for this process has been proposed.
The use of solvent extraction in zinc hydrometallurgy remains of interest. Thus solvent extraction will be used in Reunion Mining's Skorpion project in Namibia. Here the zinc silicate ore will be leached with sulphuric acid and, after solid/liquid separation, zinc will be extracted from the leach liquor with di(2-ethylhexyl)phosphoric acid (DEHPA). Zinc metal will be produced by electrowinning from the acid strip liquor. This will be the first plant to use direct solvent extraction on a leach liquor in primary zinc processing. The process as designed is reported to become one of the lowest cost zinc producers(16).
The use of solvent extraction for iron removal and control in primary zinc processing has often been advocated. A review of work done in Canada on the use of DEHPA and the mono analogue, MEHPA, has been published(17). The aim of the work was to integrate solvent extraction into a zinc processing flowsheet as a means of removing Fe in a concentrated chloride strip liquor from which saleable [Fe.sub.2][O.sub.3] could be produced. Problems arise in the transfer of sulphate and/or chloride ions from one phase to the other. Flowsheet designs and downstream processing to address these and other problems are presented and discussed.
Other workers have also studied the solvent extraction of ferric iron with DEHPA from zinc sulphate solutions(18). In this case nitric acid was used as the stripping agent. The nitric acid stripping efficiency decreased with increasing DEHPA/Fe ratio in the organic phase and the strip liquors contained no more than 5-10 g/l Fe. It was concluded that DEHPA was not suitable for this process and another reagent resistant to nitric acid but based on weaker compound formation was needed.
Solvent extraction has also been studied for the removal of copper from zinc sulphate leach liquors(19). The reagent used was LIX 622. Up to 97% of the copper was extracted from a solution containing 1.9 g/l Cu; 2 g/l Fe, 173 g/l Zn and 7.9 g/l [H.sub.2]S[O.sub.4]. Less than 45 ppm residual organics in the raffinate had little effect on zinc electrowinning. With higher contamination, the current efficiency decreased and the cathode had a rough surface morphology which increased with increasing organic concentration.
The anodic behaviour of a lead-silver alloy anode in zinc electrowinning has been investigated(20). The results of the work showed that the anode potential decreased when the Ag content of the alloy increased in the range 0.7-1.4 wt%. The anode potential also decreased markedly with increasing manganese concentration in the electrolyte or with increasing temperature.
1. Kruger, J. Future Demands to the Industry of Lead and Zinc. Erzmetall, 1998, Vol. 51, no. 4. p.259-265.
2. Zinc and Lead Processing, eds. J. E. Dutrizac, J. A. Gonzalez, G. L. Bolton and P. Hancock. The Metallurgical Society of CIM, 1998, 886p.
3. Van Os, J. Metal Bulletin, January 1988, no. 325, p.39-41. 4. Warner, N. A. Refined Zinc Metal Production at the Minesite. The Mining Cycle. Proceedings AusIMM 1998 Annual Conference, Australasian IMM Publication Series no. 2/98, 1998, p. 417.-425.
5. Buckett, G. A.; Fountain, C. R. and Sinclair, R. J. The BUKA Zinc Process: a Major Step Forward in Zinc Processing? The Mining Cycle, Proceedings AusIMM 1998 Annual Conference, Australasian IMM Publication Series no. 2/98, 1998, p.411-416.
6. Buckett, G. A. and Sinclair, R. J. The BUKA Zinc Process A 21st Century Standard. Zinc and Lead Processing, eds. J. E. Dutrizac, J. A. Gonzalez, G. L. Bolton and P. Hancock. The Metallurgical Society of CIM, 1998, p.579-596.
7. Kozlov, P. A.; Kazanbaev, L. A. and Kolesnikov, A. V. Effect of Calcium Oxide on Zinc, Extraction during Waelz Process. Tsvetnye Metally, August 1998, no. 8, p.25-27.
8. Schneider, W. D. and Schwab, B. Production of zinc from secondary material. Erzmetall, 1998, Vol. 51, no. 4, p.266-272.
9. Jiricny, V.; Roy, A. and Evans, J. W. Spouted Bed Electrowinning in the Recovery of Zinc from Scrap Galvanised Steel. EPD 1998, ed. B. Mishra, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.411-426.
10. Konishi, Y.; Nishimura, H. and Asai, S. Bioleaching of Sphalerite by the Acidophilic Thermophile Acidianus brierleyi. Hydrometallurgy, January 1998, Vol. 47, p.339-352.
11. Carranza, F. and Iglesias, N. Application of IBES Process to a Zn Sulphide Concentrate: Effect of Cu2+ Ion. Mineral Engineering, April 1998, Vol. 11, no. 4, p.385-390.
12. Harvey, T. J. and Yen, W. T. The Influence of Chalcopyrite, Galena and Pyrite on the Selective Extraction of Zinc from Base Metal Sulphide Concentrates. Mineral Engineering, January 1998, Vol. 11, no. 1, p.1-21.
13. Raghavan, R.; Mohanan, P. K. and Patnaik, S. C. Innovative processing Technique to produce Zinc Concentrate from Zinc Leach Residue with Simultaneous Recovery of Lead and Silver. Hydrometallurgy, April 1998, Vol. 48, p. 225-237.
14. Anon. Kvaerner wins Novel Zinc Project Contract. Mining Journal, October 30 1998, p.347.
15. Tailoka, F. and Fray, D. J. Recovery of Zinc from Jarosite Residues by Chlorination with both Chlorine and Scrap Polyvinyl Chloride. Trans. Instn. Min. Metall. C, May-August 1998, Vol.107, p. C60-C64.
16. Meyer, J. and Davey P. Reunion Mining, Societe General Euro Research, 5 February 1998.
17. Principe, F. T. and Demopoulos, G. P. Solvent Extraction Removal of iron from Zinc Process Solutions Using Organophosphorus Extractants. EPD 1998, ed. B. Mishra, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.267-287.
18. Van Weert, G.; van Sandwijk, T. and Hogeweg, P. Solvent Extraction of Ferric Iron from Zinc Sulphate Solutions with DEHPA, Investigation of Nitric Acid as Stripping Agent. EPD 1998, ed. B. Mishra, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p. 245-266.
19. Owosu, G. Solvent Extraction of Copper from Zinc Sulphate Leach Solutions Using LIX 622 and the Effect of Organic Contamination on Zinc Electrowinning. EPD 1998, ed. B. Mishra, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.289-300.
20. Mei, G.; Zhong, Z.; Liu, Y. and Peng, X. Investigation of the Electrochemical Behaviour of Pb-Ag Anodes. J. Cent. South Univ. Technol. China. August 1998. Vol. 29, no. 4, p. 337-340.
Lead production in the EU is from four lead-zinc shaft furnaces via the Imperial Smelting process and five other plants using new technologies i.e. Boliden, Kivcet, Mount Isa, NA and QSL. There are also another ten shaft furnaces and twenty short-drum rotary furnaces, lead-zinc slag blowing plants etc. The strengths and weaknesses of the various processes are discussed in the context of the future demands of the lead industry(1). Also concerned with the development of lead, Broad(2) notes that existing and pending legislation could hit lead markets and also therefore affect mining, processing and recycling. For example the Swedish Ministry of the Environment is trying to phase out the use of lead in all products by 2007.
The major conference concerned with lead during the year under review was Zinc and Lead Processing(3) held in Calgary, Canada in August. Highlights of this conference include a description of Asarco's Glover lead plant, lead and zinc processing at Boliden's Ronnskar smelter, lead smelting operations at Noranda's Belldune smelter, slag fuming at the Hachinohe smelter, recent changes to Pasminco's lead smelter at Port Pirie, the current state of the Kivcet process with specific papers on the Cominco Kivcet smelter at Trail, an update on the QSL process and on the Ausmelt process for lead and zinc processing, recent improvements in lead and zinc production at KCM-SA and the processing of lead-zinc secondary materials at Kabwe. There are also two papers on the Betts process, one on the current operation at Sumitomo Metal Mining's lead electrorefinery and the other on the lead electrolysis operation at the Chigirishima refinery, both in Japan.
Stack particle emissions and reductions in S[O.sub.2] emissions at the lead-zinc smelter at Plovdiv in Bulgaria have been reported(4). The lead plant treats sulphide concentrates by sintering, shaft furnace reduction, two stage liquation of copper and alkaline oxidation of impurities. The zinc plant is of the roast-leach-electrowin type.
The Kivcet process has received some attention during the period under review. In addition to the papers in reference 3, a further description of the Kivcet smelter at Trail has been given by Walker(5). The smelter came on-stream in 1997 and it was expected to reach its design capacity of 120,000 tpa in the second quarter of 1998. This furnace and a new slag fumer replaced the original sinter plant, two blast furnaces and two slag fuming furnaces. The smelter treats a wide range of sulphide concentrates, zinc leach residues, battery scrap and other material. The start-up is reported to have gone reasonably well. Two metallurgical process models have been developed namely a training simulator and a model to predict all furnace variables. Particulate air emissions from the plant could be reduced by more than 90% and metals and S[O.sub.2] emissions by 75%.
Further information on this process is given by Kulenov et al(6) who have reviewed the state- of-the-art and note that the evolution of the process, which was first implemented at Ust-Kamenogorsk in 1986, has been characterised by a change from lead concentrates as source materials to difficult-to-process lead-containing residues from zinc production such as jarosite and zinc ash leaching residues. The future of the process is discussed.
A physical and chemical description of the Kivcet lead flash smelting process has been given by Sannikov et al(7). The Kivcet process is fundamentally simple in concept and operation. Charge materials should satisfy two main requirements namely they should have a heat-generating capability for melting the flame products and the chemical composition should be such that a free-flowing, fluid slag is produced at 1200-1400 [degrees] C at all stages of the process. Zinc leach residues, Ag-Pb residues, high zinc slags and Waelz process slags have all been treated successfully at pilot and industrial scale.
A description of Boliden's Ronnskar smelter has been given(3, 8). The plant includes a lead flash smelter and refinery, a copper smelter and refinery, a sulphuric acid plant and S[O.sub.2] plant, a slag fuming plant and a precious metals plant. Raw materials for the smelter include concentrates, ashes, slags, metallic scrap, electronic scrap and electric arc steel-making dusts. The current operations are discussed together with measures to improve productivity. The deportment of critical elements is also considered along with environmental issues.
Improvements at the Herculaneum smelter in Missouri have been reported by Rosin(9). This smelter, which processes 225,000 t/y of lead concentrates from Doe Run's polymetallic mines, has been in continuous operation since 1892. Atmospheric emissions have been drastically reduced over the years and in 199697 a new 165 m stack has been built. Slag output has been reduced and slag recycling increased as a result of improvements in concentrate grades by the mining and milling divisions. Slag fuming is under investigation to reduce land releases of Zn and Pb by a further 90%. Waste and storm waters are treated by the addition of lime slurry before discharge to the Mississippi while recycling has halved water consumption over 10 years.
At Mt. Isa Mines, Pb-Zn sinter products have been examined over the past 25 years to provide an understanding of the phase chemistry and structural characteristics and their significance in the smelting process. A full description of these sinter products has been provided(10). Macroscopically the most significant feature of plant sinter is the formation of sulphate rings. Return sinter is characterised by an abundance of Pb metal and sulphide. Plant sinter is more prone to degradation than return sinter.
The use of the Isasmelt process for both the lead and copper smelters at Mt. Isa Mines has improved energy efficiency(11). This has caused C[O.sub.2] emissions per ton of metal to be reduced by 20%.
The conversion of Metaleurop's Nordenham smelter to bath smelting technology has been reported(12). After a long and difficult commissioning period it is now showing significant economic and technical benefits. The 90,000 t/y Ausmelt furnace is expected to reduce the cost of lead production by around 20%, reduce emissions by 83% for Pb, 97% for Cd, 35% for Sb, 76% for Th, 65% for Hg and 90% for S[O.sub.2], reduce slag by-product from 74,000 to 9,500 t/y and improve slag flexibility by handling up to 100% scrap and lead residues if required.
A study has been made of the ElectroSlurry process for the treatment of galena(13) wherein lead dissolution takes place electrolytically into a chloride solution. This has shown that there are three leaching mechanisms which may occur simultaneously namely chemical dissolution of PbS, oxidation of PbS by Cu(II) and Fe(III) and anodic oxidation. Elemental sulphur can form on the surface of the particles from the oxidation of PbS and also from oxidation of [H.sub.2]S formed in the first reaction. The coexistence of galena and pyrite is beneficial because of their galvanic interaction.
Increasing environmental pressure over the last 20 years on both primary and secondary lead facilities has stimulated research for new, more environmentally friendly technologies such as hydrometallurgy and electrochemistry as an alternative to pyrometallurgy. Over the last ten years Engitec has developed the use of fluoborate-based technology for the treatment of different lead bearing materials such as primary concentrates, high impurity lead bullion and lead paste from scrap batteries. A summary of such applications has been given by Olper(14). Thus ferric fluoborate can be used to dissolve lead concentrates and in the electrowinning step oxygen evolution at the anode is replaced by oxidation of ferrous fluoborate to ferric fluoborate. Elemental sulphur is produced in the leaching reaction.
1. Kruger, J. Future Demands to the Industry of Lead and Zinc. Erzmetall, 1998, Vol. 51, no. 4, p.259-265.
2. Broad, A. Legislation: a Lead Weight for the Industry? Metal Bull. Mon. April 1998, no.328, p.14-17.
3. Zinc and Lead Processing, eds. J. E. Dutrizac, J. A. Gonzalez, G. L. Bolton and P. Hancock. The Metallurgical Society of CIM, 1998, 886p,
4. Clifford, D. KCM takes the Initiative. Mining Magazine, January 1998, Vol. 178, no. 1, p. 19-25.
5. Walker, M. Kivcet Smelter On-stream at Trail. Mining Magazine, April 1998, Vol.178, no.4, p. 256-263.
6. Kulenov, A. S.; Sannikov, Y. I.; Slobodkin, L. V. and Ushakov, N. N. The KIVCET Process: A Unique Technology for the Smelting of Raw Materials Containing Lead, Zinc and/or Copper. Erzmetall, 1998, Vol. 51, no. 4. p.273-279.
7. Sannikov, Y. I.; Liamina, M. A.; Shumskij, V. A.; Grinin, Y. A. and Radashin, M. V. A Physical and Chemical Description of the Kivcet Lead Flash Smelting Process. CIM Bulletin, July-August 1998, Vol. 91, no. 1022, p.76-81.
8. Lehner, T. and Vikdahl, A. integrated Recycling of NonFerrous Metals at Boliden Ltd Ronnskar Smelter. Ibid p.353-362.
9. Rosin, N. Herculaneum Smelter. Mining Environmental Management, September 1998, Vol. 6, no. 5, p.10-12.
10. Riley, J. F. Mineral Chemistry and Structure of Lead Sinter products from Mount Isa Mines Limited Lead Smelter. Trans. Instn. Min. Metall. C, January-April 1998, Vol.107, p. C11-C17.
11. Budd, B. Reducing Carbon Dioxide Emissions from Mount Isa Operations. The Mining Cycle. Proceedings AusIMM 1998 Annual Conference, Australasian IMM Publication Series no. 2/98, 1998, p.441.-445.
12. Karpel, S. Greater Flexibility and Lower Costs at Nordenham. Metal Bulletin, December 1998, no. 336, p. 40-45.
13. Zhang, Y.; Yang, X. and Qiu, D. Thermodynamics for Electro-Slurry Process of Galena. Nonferrous Metallurgy. Chinese Society of Metallurgy, August 1998, Vol. 50, no. 3, p.71-75.
14. Olper, M. Fluoborate Technology A New Challenging Way for Primary and Secondary Lead Processing. Zinc and Lead Processing, eds. J. E. Dutrizac, J. A. Gonzalez, G. L. Bolton and P. Hancock. The Metallurgical Society of CIM, 1998, p.185-198.
Nickel and cobalt
Subjects covered by nickel papers presented at Sulphide Smelting'98 included the physical chemistry of direct nickel matte smelting(1) and nickel sulphide smelting reaction mechanisms(2). The reaction mechanism was found to depend primarily on the materials composition as well as on the oxygen content of the reaction gas. A review of environmental developments at the WMC smelter at Kalgoorlie(3) was also given, demonstrating how availability has now increased to over 90%.
In a DON smelter(1), Ni concentrate is smelted in one step into low Fe-Ni matte without conventional smelting. The slag is treated in an electric furnace to matte and waste slag. Compared with conventional Ni smelting, this process needs a lower number of smelting vessels resulting in considerable economic and environmental benefits. Conditions prevailing in DON mattes at high degrees of oxidation are considered and the distribution of elements in direct smelting is presented and compared with the conventional process.
The oxidation of a violarite and pyrite-containing nickel concentrate from the Forrestania mine in Western Australia has been studied(4). Total desulphurisation of the nickel concentrate was observed at 900 [degrees] C with 50% [O.sub.2] in the reaction gas. Oxidation was seen to start with the formation of a porous oxide rim, after which it proceeded either via grain growth of a two-phase structure or via preferential oxidation of the outer layer.
In other work by the same team the behaviour of synthetic nickel mattes under suspension smelting conditions has been studied(5). The best sulphur removal was achieved under the most oxidising conditions and with the finest size fraction. The reactivity of the mattes was dependent on the particle size as well as on the chemical composition of the material.
Several low-cost improvements have been made to Stillwater's electric furnace at the smelter in Columbus, Montana which treats a Ni-Cu-PGM concentrate along with secondary PGM-bearing materials(6). These improvements have led to substantial increases in furnace availability and capacity as well as reducing unit costs. Improvements include installation of water-cooled copper cooling elements to stabilise refractory wear and increase furnace life, improved furnace bindings, adjustment of the electrode regulator to allow higher furnace levels, installation of an automated air-slide feed system to distribute the furnace feed in an insulating layer on the bath surface and electrode seals to reduce ingress of air.
A study of refractory wear in the tuyere region of a Peirce-Smith nickel converter has been carried out by Liow et al(7). Chrome-spinel was found to be more resistant to attack than periclase which was highly susceptible to attack by S[O.sub.2] and formed MgS[O.sub.4] in the cooler regions of the brick. The pore structure in the refractory brick was identified as the key variable in resisting sulphate formation.
Iron removal from cobalt-nickel matte has been studied by Mwema et al(8). Shinkolobwe cobalt hydrates are treated in electric furnaces at Panda in the Democratic Republic of Congo under a reducing atmosphere. Slag and scraps from fire refining of broken cobalt cathodes from the Shituru plant are also added to the feed for cobalt recovery. Due to the variation in the feed composition and operating conditions of the smelter a large quantity of unsaleable matte with unacceptable Fe content is often produced. Tests have been carried out on the removal of iron by controlled oxidation, taking into account the thermodynamic distribution of Co, Ni and Cu between matte and slag. The influences of slag-matte time of contact and the time of air blowing at a constant flow on the distribution of Fe, Ni and Cu have also been considered.
Cobalt can be recovered from a variety of slags by treatment with a carbonaceous reducing agent in a direct-current arc furnace at around 1,500 [degrees] C(9). The principal material suitable for treatment by this technology are primary smelter slags originating from the processing of sulphide concentrates. In all cases studied, cobalt was recovered as a valuable by-product thus helping to improve overall plant profitability. Pilot plant tests have demonstrated a cobalt recovery of [greater than]80% at power levels up to 600 kW. Very high recoveries of other valuable elements such as nickel and copper, have also been achieved.
The use of Ausmelt technology for the recovery of cobalt from smelter slags has been reported by Matusewicz and Mounsey(10). This technology is already successfully established in eleven commercial plants for the recovery of base metals. Studies of specific parameters indicated that a commercial unit could produce mattes containing around 8% Co from a single furnace using pyrite or other sulphidising material, limestone flux, low lance-combustion stoichiometry of 80-90%, lump coal reductant and an operating temperature of up to 1,400 [degrees] C. A second, batch furnace could upgrade the product to 20% Co. Alternatively, a single furnace at higher temperatures could produce a sulphur-deficient matte/alloy.
In hydrometallurgy interest continues to be mainly focused on the processing of nickel laterites, particularly by acid pressure leaching. The current position with regard to the hydrometallurgical laterite projects has been reviewed(11). In Australia, production of three of them namely Murrin Murrin, Cawse and Bulong are due to come on-stream in 1999. The pressure acid leaching (PAL) technology they will employ, although used successfully in Cuba for 35 years has yet to be demonstrated on a commercial scale for this new generation of projects. A major advantage of almost all the new PAL projects is that they will recover substantial quantities of cobalt and there appears to be the potential for these projects to produce nickel at cash costs below US$1.00/lb.
Papers presented at the Alta 1998 Nickel/Cobalt Pressure Leaching and Hydrometallurgical Forum reflect this interest(12). Papers were presented on the Murrin Murrin, Marlborough and Weda Bay projects while a whole session was devoted to pressure leaching technology. There was also a session on the treatment of sulphide ores which included a paper from BacTech on pilot plant experience in bacterial leaching of Ni/Co concentrates and a Billiton/QNI paper on the BioNIC process. Pressure leaching of NI/Co concentrates via the CESL process was the subject of a paper from Cominco while the Activox process was further discussed by Western Minerals Technology. A further paper outlined the commissioning of the Hartley Platinum Base Metals Refinery in Zimbabwe.
There was also a session on cobalt recovery projects which included an update paper on the Kasese project in Uganda, and a joint paper from MINTEK and Bateman in South Africa, and International Panorama Resource Corp. and the University of British Columbia on the pilot plant work that has been done on the Kakanda Cu-Co tailings project in the Democratic Republic of Congo. A session on technology developments included a paper from Henkel on the solvent extraction of nickel from ammoniacal leach solutions and a joint paper by Impala Platinum and Bateman comparing the performance of mixer settlers and pulsed columns in the separation of cobalt from nickel using CYANEX 272 as the extractant.
A comprehensive review of cobalt extractive metallurgy has been provided by Flett and Anthony(13). This review covers the recovery processes from both pyrometallurgical and hydrometallurgical processes. In particular the hydrometallurgy of cobalt production is dealt with in detail. The current methods for cobalt-nickel separation are highlighted and the rising importance of solvent extraction in this area noted.
Another review has been provided by Hawkins(14) that covers all aspects of cobalt production. Hawkins notes that copper-cobalt flowsheets do not generally include solvent extraction and comments that selective precipitation may be used. The authors of the current review comment that this is however an inefficient process which results in cobalt losses. Adoption of solvent extraction in such flowsheets could bring dividends of higher cobalt recoveries and purer cobalt products.
Bioleaching of cobaltiferous pyrite using mixed cultures of T. ferrooxidans, T. thiooxidans and I. ferrooxidans has been described by Battaglia-Brunet et al(15). The use of this mixed culture proved very successful with cobalt solubilisation increasing when L. ferrooxidans was present but was unaffected by its initial concentration. It was only improved by adding T. thiooxidans when the initial concentration of the latter was higher than that of T. ferrooxidans.
The bacterial leaching of a Chilean arsenide ore from the Atacama has been studied by Chamorro and Frenay(16). The cobalt is present mainly as safflorite (Co[As.sub.2]) and in minor quantities as skutterudite (Co, Ni)[As.sub.3]. The results show that the microorganisms keep their activity up to 5 g/l As and 10 g/l Co. Optimal conditions gave a yield of 91% Co extraction in 12 days.
Higher grade limonite acid pressure leaching with particular reference to the Calliope project in New Caledonia has been discussed by Russel(17) and contrasted to the lower grade nontronite clays at Murrin Murrin. Acid leaching of limonites benefits from lower energy consumption, fewer environmental problems and higher Ni and Co recoveries. High grading the Western Australian ores from 1% to around 1.8% Ni in the early years, by mining limited high-grade pods and by screening, will accelerate cash flows and payback periods ahead of commissioning of Voisey's Bay. A comparison with QNI shows how undervalued Calliope is claimed to be.
Ammoniacal leaching of cobalt crusts using S[O.sub.2] as the reductant has been reported(18). Leaching efficiencies of 90% Ni, 97% Co, 3% Cu, 1.8% Fe and 6% Mn were attained. The metals were separated from the ammonium carbonate solution by solvent extraction with LIX 84-I.
A review of the precipitation of nickel from salt solutions by hydrogen reduction has been presented by Saarinen et al(19). In this review a general presentation is made of the principal phenomena, procedures, catalysts and problems involved.
Solvent extraction of both nickel and cobalt remains of interest also. Thus the process parameters for the recovery of nickel from solutions containing ammonium sulphate by solvent extraction with LIX 84-I and production of nickel sulphate from the strip solution have been determined(20). The extraction of nickel, cobalt and copper using LIX 84-I in a solvent impregnated resin has been reported(21). The solvent extraction of nickel with LIX 54 has also been described(22) as has the extraction of nickel using Acorga M5640(23). The separation and recovery of cobalt and nickel from sulphate solutions using the sodium salts of DEHPA, PC 88A and CYANEX 272 has been reported(24).
The selective removal of cadmium impurity from cobalt-nickel solutions by precipitation with sodium diisobutyldithiophosphinate prior to Co/Ni separation by solvent extraction has been studied(25). Such an approach may be required to prevent cadmium contamination of the cobalt product by co-extraction in the Co/Ni separation step.
The results of mini-plant testing of the direct solvent extraction of cobalt and nickel from the acid pressure leach liquors from the Bulong deposit in Western Australia have been reported(26). Two solvent extraction processes are involved namely firstly CYANEX 272 to extract cobalt and reject nickel and secondly Versatic 10 to extract and purify nickel. The results of the testwork shows that the first process achieves 97.5% Co extraction and over 99% co-extraction of Mn and Zn. Very good separation of cobalt and nickel is achieved. In the second step Mg and Ca are rejected to the raffinate and a nickel recovery of [greater than]99% is achieved. Any cobalt not extracted in the first step is recovered here and reports to the nickel strip liquor and thence to the nickel cathode.
A process has been developed to directly electrowin zinc from the strip liquor from its removal by solvent extraction from a cobalt-bearing solution(27). The reagent used was CYANEX 272 which is selective for zinc at a controlled pH of 3.0. The zinc loaded organic phase is scrubbed with a zinc rich electrolyte solution and the scrubbed organic is stripped to yield a rich electrolyte in which all contaminants are [less than]2 mg/l. zinc is subsequently electrowon using conventional technology. The process forms part of the Boleo project currently under development by International Curator Resources Ltd. for its property in Baja California, Mexico.
A process using Caro's acid to precipitate cobalt from a Zn-Cd-Co-Ni sulphate solution has been investigated(28). About 98-99% of the cobalt was precipitated at pH 3.5-4.0 while only about 8% was precipitated at pH 3.5. The cobalt content of the Co[(OH).sub.3] precipitate was 47- 49% which is fairly close to the theoretical value of 53.6%. The impurities consisted of 34% Zn, approximately 1% Cd and less than 0.005% Ni.
In nickel electrowinning from chloride electrolytes the use of an ion exchange membrane between an impure anolyte and catholyte offers the opportunity to produce a nickel product low in impurities which form anionic complexes. This has been demonstrated in work done by Liao et al(29) wherein lead-contaminated nickel electrolyte was electrolysed in a cell with a cation exchange membrane separating the catholyte from the impure anolyte. A relatively pure nickel was produced due to the preferential migration of nickel cations through the membrane and the rejection of the anionic lead chloride species.
Finally, with the growth in production of lithium ion and nickel-metal hydride batteries it was inevitable that work would be carried out to develop possible ways of recycling them(30, 31). For the lithium ion batteries(30) HCl was found to be the most suitable leachant giving a leaching efficiency of [greater than]99% for both Co and Li. The cobalt in the leach liquor was effectively separated from the lithium by solvent extraction with PC 88A at an equilibrium pH of 6-7. The lithium in the raffinate was concentrated and recovered as lithium carbonate while the cobalt was stripped with sulphuric acid from the organic phase and recovered as cobalt sulphate.
For the nickel hydride batteries(31) again HCl was found to be the most suitable lixiviant giving 100% recovery of cobalt, [greater than]96% Ni and 99% rare earths. The rare earths were recovered by solvent extraction with DEHPA and a mixed rare earth oxide of [greater than]99% purity was obtained after oxalic acid precipitation from the HCI strip liquor. The cobalt and the nickel in the raffinate were effectively separated by solvent extraction of the cobalt with 25v/o trioctylamine in kerosine. The cobalt in the strip liquor and the nickel in the raffinate were subsequently recovered as oxalates at purifies of close to 99.9%.
Nickel and cobalt references
1. Makinen, T. and Taskinen, P. Physical Chemistry of Direct Nickel Matte Smelting, Sulphide Smelting '98, eds., J. A. Asteljoki and R. L. Stephens, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.59-68.
2. Jyrkonen, S.; Stromberg, S.; Jokilaakso, A. and Sjoblom, J. On the Reaction Conditions of Sulfidic Nickel particles in Suspension Melting Conditions. Ibid. p.77-92.
3. Tjerkstra. M. and Philips, P. Kalgoorlie Nickel Smelter Environmental Project. Ibid, p.93-112.
4. Stromberg, S. K.; Jokilaakso, A. T. and Jyrkonen, S. K. Oxidation Behaviour of Violarite-based Nickel Concentrate in Simulated Suspension Smelting Conditions. Trans. Instn. Min. Metall. C. January-April 1998, Vol. 107, p.C18-C29.
5. Jyrkonen, S. K.; Jokilaakso, A. T.; Stromberg, S. K. and Taskinen, P. A. Behaviour of Synthetic Nicked mattes under Suspension Smelting Conditions. Trans. Instn. Min. Metall. C, January-April 1998, Vol. 107, p.C30-C36.
6. Voermann, N.; Vaculik, V.; Ma, T.; Nichols, C.; Roset, G. and Thurman, W. Improvements to Stillwater Mining Company's Smelting Furnace Yielding Increased Capacity and Productivity. Sulphide Smelting'98, eds., J. A. Asteljoki and R. L. Stephens, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.503-518.
7. Liow, J. L.; Tsirikis, P. and Gray, N. B. Study of Refractory Wear in the Tuyere Region of a Peirce-Smith Nickel Converter. Canadian Metallurgy Quarterly, April 1998, Vol. 37, no. 2, p.99-117.
8. Mwema, M. D. and Kazadi, G. M. Iron Removal from Cobalt-Nickel Matte. Montreal'98, CIMM, 1998. Paper 052.pdf on CD.
9. Jones, R. T. and Deneys, A. C. Using a Direct-Current Arc Furnace to Recover Cobalt from Slags. Journal of Metals, October 1998, Vol.50, no. 10, p.57-61.
10. Matusewicz, R. and Mounsey, E. Using Ausmelt technology for the Recovery of Cobalt from Smelter Slags. Ibid, October 1998, Vol. 50, no. 10, p.53-56.
11. Anon. Testing Times for Nickel. Mining Journal, London, December 4, 1998, p.454-455.
12. Alta 1998 Nickel/Cobalt Pressure Leaching and Hydrometallurgical Forum, held 25-27 May, 1998, Alta Metallurgical Services, 1998.
13. Flett, D. S. and Anthony, M. T. Cobalt Extractive Metallurgy: A Review. International Minerals and Metals, November 1998, Vol. 1, no. 11, p.299-312.
14. Hawkins, M. J. Recovering Cobalt from Primary and Secondary Sources. Journal of Metals, October 1998, Vol. 50, no. 10, p.46-50.
15. Battaglia-Brunet, F.; d'Hugues, P.; Cabral, T.; Cezac, P.; Garcia, J. L. and Morin, D. Mineral Engineering., February 1998, Vol. 11, no. 2, p.195-205.
16. Chamorro, J. and Frenay, J. Bacterial Leaching of a Chilean Cobalt Arsenide. Environmental and Innovation in Mining and Mineral Technology. Eds. M. A. Sanchez, F. Vergara and S. H. Castro, University of Concepcion, Chile, 1998, Vol. 1, p.115-129.
17. Russel, D. S. The Case for Higher-Grade Laterite Nickel Ore Processing. Chamber Mines J. Zimbabwe. February 1998. Vol. 40, no. 2, p. 33-41.
18. Inoue, A. and Kawahara, M. Ammoniacal Leaching and Solvent Extraction of Cobalt Crusts using Sulphur Dioxide. J. Min. Mater. Process. Inst. Jpn., March 1998, Vol. 114, no. 3, p.195-199.
19. Saarinen, T.; Lindfors, L-E. and Fugleberg, S. A Review of the Precipitation of Nickel from Salt Solutions by Hydrogen Reduction. Hydro-metallurgy, January 1998, Vol. 47, no.2-3, p.309-324.
20. Parija, C.; Reddy, B. R. and Bhaskara Sarma, P. V. R. Recovery of Nickel from Solutions Containing Ammonium Sulphate using LIX 84-1. Ibid, August 1998, Vol. 49, no. 3, p. 255-261.
21. Inoue, A.; Hatae, S. and Kawahara, M. The Extraction of Nickel, Cobalt and Copper Using Solvent-impregnated Resin-Adsorbed LIX 84-1. J. Min. Mater. Process. Inst. Jpn., October 1998, Vol. 114, no. 11, p.819-824.
22. Alguacil, F. J. and Cobo, A. Solvent Extraction Equilibrium of Nickel with LIX 54. Hydrometallurgy, May 1998, Vol. 48, no. 3, p. 291-299.
23. Alguacil, F. J. and Cobo, A. Extraction of Nickel from Ammoniacal/Ammonium Carbonate Solutions using Acorga M5640 in Iberfluid. Ibid, October 1998, Vol. 50, no. 2, p. 143-151.
24. Devi, N. B.; Nathsarma. K. C. and Chakravortty, V. Separation and Recovery of Cobalt(II) and Nickel(II) from Sulphate Solutions using Sodium Salts of D2EHPA, PC 88A and Cyanex 272. Ibid, June 1998, Vol. 49, no. 1-2, p.47-61.
25. Rickelton, W. A. The Removal of Cadmium Impurities from Cobalt-Nickel Solutions by Precipitation with Sodium Diisobutyldithio-phosphinate. Ibid, November 1998, Vol. 50, no. 3, p.341-346.
26. Soldenhoff, D.; Hayward, N. and Wilkins, D. Direct Solvent Extraction of Cobalt and Nickel from Laterite-Acid Pressure Leach Liquors EPD 1998, ed. B. Mishra, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p.153-165.
27. Tomlinson, M.; Lommen, J. and Molnar, R. Solvent Extraction and Direct Electrowinning of Zinc from a Cobalt-Bearing Solution. Zinc and Lead Processing, eds. J. E. Dutrizac, J. A. Gonzalez, G. L. Bolton and P. Hancock, The Metallurgical Society of CIM, 1998, p.733-747.
28. Owusu, G. Oxidation-Precipitation of Co from Zn-Cd-Co-Ni Sulphate Solution using Caro's Acid. Hydrometallurgy, March 1998, Vol. 48, no. 1, p.91-99.
29. Lioa, L.; van Sandwijk, A. and Van Weed, G. Low-Lead Nickel Production by Electrowinning of Lead-Contaminated Nickel Chloride Solution in a Membrane Equipped Electrolytic Cell. EPD 1998, ed. B. Mishra, The Minerals, Metals and Materials Society, Warrendale, Pa., 1998, p. 381-391.
30. Zhang, P.; Yokoyama, T.; Itabashi, 0.; Suzuki, T. M. and Inoue, K. Hydrometallurgical Process for Recovery of Metal Values from Spent Lithium-Ion Secondary Batteries. Hydrometallurgy, January 1998, Vol. 47, no. 2-3, p. 259-271
31. Zhang, P.; Yokoyama, T.; Itabashi, O.; Wakui, Y.; Suzuki, T. M. and Inoue, K. Hydrometallurgical Process for Recovery of Metal Values from Spent Nickel-Metal Hydride Secondary Batteries. Ibid, September 1998, Vol. 50, no. 1, p.61-75.
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|Author:||Flett, D.S.; Anthony, M.T.|
|Date:||Jul 1, 1999|
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