The selective leaching of zinc from chalcopyrites-phalerite concentrates using slurry electrodes

Hydrometallurgy, 14 (1985) 317--329

317

Elsevier Science Publishers B.V., Amsterdam -- Printed in The Netherlands

THE SELECTIVE LEACHING OF ZINC F R O M C H A L C O P Y R I T E - S P H A L E R I T E C O N C E N T R A T E S USING S L U R R Y E L E C T R O D E S

M.N. MAHMOOD and A.K. TURNER

B.P. Research Centre, Sunbury-on-Thames, Middlesex (Great Britain) (Received September 21, 1984; accepted in revised form March 1, 1985)

ABSTRACT M a h m o o d , M.N. and Turner, A.K., 1985. The selective leaching of zinc from chalcopyrite--sphalerite concentrates using slurry electrodes. Hydrometallurgy, 14: 317--329. A n electrohydrometallurgical method has been identified which facilitatesthe selective separation of zinc from a chalcopyrite--sphalerite concentrate. Over 80 w t % zinc extraction has been demonstrated with less than 0.2 w t % copper dissolution. This was achieved by the application of a controlled potential to a concentrate slurry in a strongly acidic solution. A suitable redox potential (Eh) was identified which both suppressed copper dissolution and enhanced zinc extraction.

INTRODUCTION

A n investigation of the removal of zinc from a copper concentrate was prompted by the need to reduce penalty payments imposed by the smelter when the zinc content of the concentrate rose above 2.5 wt%. The concentrate consisted of a finely intergrown mixture of chalcopyrite and sphalerite which were not readily separable by flotation. A n electrochemical method has been demonstrated iwhich enables the selective separation of zinc from the copper concentrate. Although the technique is not considered to be economically viable for this application, it clearly indicates the capability of achieving the selective leaching of complex sulphides by controlling the redox potential (Eh) during leaching. The anodic oxidation of chalcopyrite has been shown to lead to the dissolution of copper and iron, and the formation of elemental sulphur or sulphate ions. Early workers (Linge [I], Bauer et al. [2], Jones and Peters [3], and Warren [4] ) observed the preferential dissolution of iron during the oxidation of chalcopyrite, and postulated the formation of a metal-deficient sulphide layer (richer in Cu than Fe) on the electrode surface. Alternatively, the formation of covellite (CuS) was proposed (Jones and Peters [3], Amrnou-Chokroum et al. [5 ]) to explain copper enrichment at the surface and a decrease in the rate of oxidation of chalcopyrite. Parker et al. [6,7] observed that in the potential range 0.2--0.6 V vs. SCE the main oxidation pro-

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© 1985 Elsevier Science Publishers B.V.

318 ducts were Cu 2÷ (Cu ÷ in chloride media), Fe :÷ and S °. At potentials higher than 0.6 V vs. SCE, Cu :÷, Fe 3÷ and S O ~ - w e r e formed, b u t the reaction was retarded b y the formation of a passive layer, thought to be a copper polysulphide (CuSx where x > 2) on the electrode surface. The nature of the surface layer was f o u n d b y McMillan et al. [8] to be different in sulphate and chloride media. Several electrohydrometallurgical processes have been developed for copper extraction from sulphide concentrates including the Dextec Process (Everett [9] ), the Cymet Process (Kruesi et al. [10] ), and the Electroslurry Process (Dahlstrom et al. [11] ). In these systems copper is anodically dissolved from a sulphide slurry by direct electro-oxidation and/or by the in-situ generation of an oxidant such as ferric ions. Electrolytic oxidation methods have also been used to treat complex sulphide concentrates. Lead, zinc, copper and silver were extracted (Scheiner et al. [14]) from lead--zinc concentrates by anodic dissolution in sodium chloride solution, using a membrane-separated slurry electrolysis cell. More recently, the Dextec Lead Process has been described (Everett [15] ) for the anodic dissolution of lead from a mixture of P b / Z n / C u / F e sulphides, with simultaneous cathodic lead deposition. Lead dissolution was considered to occur by the acid decomposition of PbS, producing H2S which was oxidised to elemental sulphur at the anode. Slurry electrolysis has also been used (Oki and Kammel [12] ) to anodically leach zinc from sphalerite concentrates, with elemental sulphur as the major b y p r o d u c t . The anodic dissolution of ZnS (a mixture o f sphalerite (~-ZnS) and wurtzite (~-ZnS)) in sulphuric acid was thought (Narasagoudar et al. [13] ) to occur via both electrochemical and chemical steps. A corrosion-type reaction was proposed in which ZnS was oxidised to Zn 2÷ and SOat the anodic sites, and So was reduced to H~S at the cathodic sites. Recent reports in the literature (Scott [16] and Nicol [17] ) suggest that the non-oxidative dissolution of metal sulphides proceeds b y steps involving varying degrees of electron transfer. The rate of iron dissolution from pyrrhotite was f o u n d to be strongly dependent on the potential of the mineral and a volcano-type relationship was observed between potential and rate of leaching. They also observed that the non-oxidative leaching of nickel and lead as well as iron sulphides was initiated b y short cathodic pulses. In the case of pyrrhotite it was shown that either a cathodic shift imposed on the mineral, or the addition of sodium sulphide (which effected a cathodic shift of potential) initiated dissolution. In the present work, an electrohydrometaUurgical m e t h o d has been used for the selective leaching of zinc from a copper/zinc sulphide concentrate using slurry electrodes. Selectivity has been achieved b y electrochemical control of the redox conditions (E h ) during leaching.

319

EXPERIMENTAL

Samples of copper concentrate from Les Mines Selbaie, Quebec, Canada were used, which consisted mainly of chalcopyrite (CuFeS2) and sphalerite (ZnS), with some pyrite (FeS2) and quartz (SiO2). The concentrate assayed 27.9 w t % copper, 25.3 w t % iron and 9.7 w t % zinc. Mineralogical examination revealed that the chalcopyrite occurred as jagged grains under 100 p m in size,nearly always liberated from the gangue, and that the pyrite occurred as more or less rounded grains. The sphalerite was found to be intergrown with chalcopyrite; approximate values for the form of the sphalerite were determined by measuring areas under the microscope. The following distributions were calculated: 11% sphalerite was locked to individual grains (>15 pro) of chalcopyrite, 4 7 % sphalerite contained small (I--15 pro) disseminations of chalcopyrite, 32% sphalerite contained minute (
"

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Fig. I. Slurry electrolysis cell. I: P T F E cell head; 2: graphite feeder electrode; 3: graphite counter electrode; 4: stirrer; 5: reference electrode; 6: screw thread connection; 7: Luggin capillary; 8 : anion exchange membrane; 9: flanged glass cell; 10: slurry.

320 saturated calomel electrode (SCE) was used as the reference electrode; it was housed in a separate c o m p a r t m e n t with the Luggin capillary positioned close to one of the graphite feeder electrodes. The graphite counter electrode was separated from the slurry b y an anion exchange membrane. This ensured that metal ions liberated from the slurry did n o t react electrochemically at t h e counter electrode. During constant-potential electrolysis experiments the current was continuously monitored using a chart recorder. Standard electrode potential (E °) values referred to below are q u o t e d with respect to the SCE, assuming that 0.0 V vs. SCE is equivalent to +0.245 V vs. the standard hydrogen electrode. It should be noted that this cell design was n o t optimised to give uniform potential distribution within the slurry. In particular, the counter electrode was n o t symmetrically positioned with respect to the feeder electrodes. Nevertheless, appreciable changes in the selectivity of leaching were observed for quite small (150 mV) changes in the potential applied to the feeder electrodes. This implies that distribution of potential in the slurry was not very broad. The cell was partially immersed in a thermostatically controlled water bath which was positioned within a fume cupboard, in order to remove any hydrogen sulphide emitted from the cell. At the end of each experiment the slurry was filtered using a Buchner funnel and the filtrate analysed b y atomic absorption spectrometry. The amounts of copper, iron and zinc dissolved into the electrolyte were then calculated as percentages recovered from the concentrate. In some experiments, slurry samples were withdrawn using a syringe at intervals during the run, then filtered through a Millipore filter. When compared, these t w o techniques were f o u n d to give close, but not identical, results. Step scanned X-ray diffraction data of filtered residues were collected with a Philips PW1050 diffractometer using copper K-alpha radiation, and a graphite crystal m o n o c h r o m a t o r in the diffracted beam. RESULTS AND DISCUSSIONS The redox conditions during the electroleaching of the concentrate slurries were controlled b y a constant potential applied to the carbon feeder electrodes immersed in the slurry. At the end of each batch experiment, the extent of metal dissolution was determined b y analysis of the filtered leach liquor. The effect of applied potential on the dissolution o f copper, iron and zinc from the concentrate is reported below. The variation of the current with time during controlled potential electrolysis is also discussed. Direct comparison is made between leaching under controlled potential conditions and results obtained with no applied potential.

Copper Copper dissolution from the concentrate was suppressed to less than 0.2 wt% when leaching in 500 g/1 H=SO4 at 60°C was carried o u t at applied po-

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tentials below 0.15 V vs. SCE (Fig. 2). Appreciable copper dissolution occurred only at more positive potentials, probably caused by either oxidative or acid attack of the chalcopyrite (reactions (1) and (2), respectively). At 0.0 V vs. SCE copper dissolution was suppressed in solutions containing from 100 to 500 g/1 H2SO4 (see Fig. 3). For comparison, similar experiments were carried out w i t h o u t potential control, where the redox potential during leaching was approximately 0.3 V vs. SCE. In these cases, copper dissolution was significant (up to 6 wt%) which increased with acid strength in the range 100 to 500 g/1 (see Fig. 3). This was probably due to the acid decomposition of chalcopyrite according to reaction (2). CuFeS: -* Cu 2÷ + Fe 2÷ + 2 So + 4 eCuFeS2 + 4 I-r ~ Cu 2÷ + Fe 2÷ + 2 H2S

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After controlled potential electrolysis at 0.0 V vs. SCE in 500 g/1 H2SO4 at 60°C, the electroleached residue was shown by X-ray diffraction (XRD) to contain a small a m o u n t of covellite (CuS), in addition to unreacted chalcopyrite, sphalerite, pyrite and quartz. The covellite m a y have been formed by the partial oxidation of chalcopyrite according to reaction (3). Other possible mechanisms of coveUite formation are considered below. C u F e S 2 - ~ C u S + F e 2+ + S O + 2 e -

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Iron

When the concentrate was electroleached in 500 g/1 H2SO4 (60 ° C) at potentials where chalcopyrite dissolution was suppressed, i.e., below 0.15 V vs. SCE, the dissolution of iron increased from 4 to 19 wt% towards more negative applied potentials (Fig. 2). This implies t h a t iron dissolution was due to a reductive process, presumably reduction of either pyrite or chalcopyrite. Indeed, XRD of the residue leached at 0.0 V vs. SCE revealed a decrease in the pyrite content. This was probably due to pyrite reduction according to reaction (4). However, the absence of any appreciable dependence of iron dissolution u p o n acid strength, b o t h at 0.0 V vs. SCE and

323 with no applied potential (Fig. 3), implies a significant contribution from the dissolution of iron from surface layers of oxidised material on the concentrate particles. FeS2 + 4 I-l* + 2 e- -~ Fe 2÷ + 2 H2S

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(4)

At potentials approaching - 0 . 3 V vs. SCE, the increase in iron dissolution to 19 wt% was probably due to the reduction of chalcopyrite, the major iron-bearing phase in the concentrate, to produce chalcocite (Cu2S) according to reaction (5). It should be noted t h a t the electrode potential for this reaction will be more positive than the E ° value at low concentrations of Fe 2÷ and H2S, increasing by 0.06 V and 0.09 V, respectively, for an order of magnitude decrease in concentration. No Cu2S was detected in residues after electroleaching at 0.0 V vs. SCE. 2 CuFeS2 + 6 I-l* + 2 e- -* Cu2S + 2 Fe 2÷ + 3 H~S

E ° = - 0 . 4 6 V vs. SCE

(5) Zinc Controlled potential electrolysis of concentrate slurries in 500 g/1 H2SO4 at 60°C revealed t h a t zinc dissolution was markedly dependent on the applied potential in the range - 0 . 3 to +0.3 V vs. SCE. Figure 2 shows that zinc dissolution was highest after electroleaching at 0.0 V vs. SCE, and that it decreased sharply at both more cathodic and more anodic potentials. With no applied potential, acid decomposition o f ZnS in the concentrate was small with only 4--5 wt% zinc dissolution in 4 hours at 60 ° C. Furthermore, under these conditions, zinc recovery was independent of acid strength in the range 100 to 500 g/1 H2SO4 (see Fig. 3). This implies that, in this instance, the leaching of zinc may have been due mostly to the dissolution of oxidised material from the concentrate. In contrast, when the concentrate was electroleached at a constant potential of 0.0 V vs. SCE, zinc dissolution increased dramatically with acid strength, reaching 16 wt% after 4 hours in 500 g/1 H2SO4 at 60°C (Fig. 3). This strong dependence on acid strength would n o t be expected if zinc dissolution was occurring by the simple oxidation of ZnS (reaction 6). The results can be satisfactorily explained if the acid decomposition of ZnS is considered to be a "corrosion-type" reaction (Narasagoudar et al. [13]) in which the anodic oxidation of ZnS (reaction 6) is coupled with the cathodic reduction of elemental sulphur (reaction 7), to give the overall reaction (8) ZnS-~ Zn 2÷ + SO + 2 e-

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So + 2 I-l* + 2 e- -~ H2S

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ZnS + 2 I-l* -~ Zn 2÷ + H2S

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Assuming the following initial conditions: [Zn 2÷] = 1 pmol/1, [H2S] = 1

324

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E6 = 0.06 + 0.03 log [Zn 2÷] = - 0 . 1 2 V vs. SCE and E7 = --0.11 -- 0.03 log [H2S] - 0.06 pH = 0.07 V vs. SCE Thus a corrosion potential would exist b e t w e e n --0.12 and +0.07 V vs. SCE, which would provide a negative overpotential for the cathode reaction (7), and a positive overpotential for the anode reaction (6). Consequently, when electroleaching was carried o u t at 0.0 V vs. SCE, close to the likely corrosion potential, the anodic and cathodic reactions would have occurred at similar rates. This would have prevented the build-up of a passivating sulphur layer, allowing the zinc dissolution reaction to proceed. The increase in zinc extraction with acid strength in experiments at 0.0 V vs. SCE could then be due to more effective sulphur reduction according to reaction (7). This is consistent with the absence of elemental sulphur in concentrate residues electroleached at 0.0 V vs. SCE, as revealed b y XRD. The formation of a surface coating o f elemental sulphur could account for the observed decrease in zinc dissolution at more positive potentials (see Fig. 2), due to passivation. Indeed, when the concentrate was leached with no applied potential, the redox conditions were probably sufficiently anodic (approximately 0.3 V vs. SCE) to favour the oxidation of ZnS; the reaction could then have become passivated b y sulphur deposition. Similar dependence of non-oxidative dissolution of iron, lead and nickel sulphides on potential were also reported b y Scott and Nicol [16, 17]. They observed that Fe, Ni and Pb sulphides have a c o m m o n broad Eh region over which the metals are leached. Since the region was found to be cathodic of the potential required for the reduction of elemental sulphur, the authors concluded that non-oxidative dissolution o f these metal sulphides may proceed through a "corrosion-type" mechanism, involving simultaneous anodic dissolution o f the metal sulphide and the cathodic reduction of sulphur. The current flowing through the cell during constant potential electrolysis was recorded, and a typical current response is shown in Fig. 4. At the start of the controlled potential electrolysis experiments at 0.0 V vs. SCE a large cathodic current of up to 1 A (equivalent to 250 A / m : ) was observed initially, which gradually decreased over the following 40 min to 1 h. This high cathodic current would have led to an uncompensated ohmic potential drop between the tip of the Luggin capillary and the graphite/slurry interface facing the counter electrode (dependent on the cell configuration). During this period the copper concentration in the electrolyte solution was quite high (up to 1000 mg/1 Cu). It is likely that this initial current was due to the deposition of copper which had dissolved from surface layers of oxidised

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material on the concentrate particles. However, the deposition o f either metallic copper or chalcocite was n o t confirmed b y X R D measurements of the electroleached residue, although conceivably t h e y could have been present as amorphous phases. It could be that the dissolved copper was precipitated during this period b y its reaction with electrogenerated H2S, with the formation of covellite according to reaction (9). Indeed, X R D did reveal the presence of covellite in residues leached at 0.0 V vs. SCE. Cu2+ + H2S ~ CuS + 2 H+

(9)

Following the cathodic current peak, the current reached a fairly steady cathodic value less than - 5 0 mA (equivalent to approximately 10 A/m2), and the copper concentration in solution remained at a low level. In some experiments, a small net anodic current was observed at stages during the runs. In either case, the current density was sufficiently low at this stage to make any uncompensated ohmic potential drop negligible. The observation of a net current close to a value of zero, usually cathodic, but sometimes anodic, is indicative of the balance between the anodic and cathodic components of ZnS dissolution during electroleaching at 0.0 V vs. SCE.

326

This provides further evidence to support a "corrosion-type" mechanism for zinc dissolution. A further aspect of the reaction mechanism is the possible role of H2S, liberated during controlled potential electrolysis, in aiding redox potential control in the slurry. Dissolved H:S m a y serve to enhance a uniform distribution of potential in the slurry under these mildly reducing conditions (0.0 V vs. SCE) by acting as an electrogenerated reducing agent. The dissolution of zinc from the concentrate by controlled potential electrolysis at 0.0 V vs. SCE in 500 g/1 H2SO4 for 4.2 h increased with temperature from 16 wt% at 60°C to 40 wt% at 90°C, as shown in Fig. 5. After 8.2 h at 80°C the zinc removal was 50 wt%, with less than 0.2 wt% of the copper dissolved. For comparison, leaching for 6 h in 500 g/1 H2SO4 with no applied potential produced 3.5 wt% zinc dissolution at 60°C (Fig. 6). Copper 50

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dissolution with no applied potential reached 5.5 wt% at 90°C (Fig. 6), but was suppressed to less than 0.2 wt% after controlled potential electrolysis at 0.0 V vs. SCE at temperatures from 60 to 90°C (Fig. 5). The extent o f zinc dissolution on electroleaching at 0.0 V vs. SCE using 500 g/1 H2SO4 at 70°C was found to increase with time (Fig. 7). After electroleaching for 72 h, 84 wt% of the zinc had been removed from the concentrate, with copper extraction at less than 0.2 wt%. This produced a leach liquor containing 40 g/1 zinc, 13 g/1 iron, and less than 300 rag/1 copper. CONCLUSION The selective dissolution of zinc from a chalcopyrite--sphalerite concentrate has been achieved with no significant loss of copper. Controlled potential electrolysis at 0.0 V vs. SCE of concentrate slurries in 500 g/1 H2SO4 at 70°C led to over 80% zinc dissolution in 72 h with less than 0.2% copper dissolution. The rate of zinc removal increased with acid strength and temperature. Controlling the redox potential at 0.0 V vs. SCE serves to "cathodically p r o t e c t " the chalcopyrite, preventing copper dissolution. It is also thought to enhance zinc dissolution by reducing the passivating film of sulphur, f o r m e d on the particles by anodic decomposition o f the sphalerite, to H2S.

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ACKNOWLEDGEMENTS The authors gratefully acknowledge the assistance of Mr. S. Hickman for carrying out part of the experimental work, Dr. R.A. Schedler for mineralogical examination of the concentrate, Mr. D. Wood for X-ray diffraction measurements, Miss P. Franks for the diagrams and Miss S. Ellis and Mrs. G. Tindall for typing the manuscript. Permission to publish this paper has been granted by the British Petroleum Company p.l.c. REFERENCES 1 Linge,H.G., Hydrometallurgy, 2 (1976) 51. 2 Bauer, J.P., Gibbs, H.L. and Wadsworth, M.E., Annual Meeting of the A m e r i c a n Institute of Mechanical Engineers,San Francisco, 1972, Paper 72-B-96.

329 3 Jones, D.L. and Peters, E., International Corrosion Conference Series 1973, NACE-4, 1977, pp. 433--458. 4 Warren, G.W., Ph.D. Thesis, University of Utah, 1978. 5 Ammou-Chokroum, M., Sen, P.K. and Fouques, F., In: Laskowski, J. (Ed.), Developments in Mineral Processing, Proceedings of the 13th International Minerals Processing Congress, Warsaw, Poland, June 1979, Elsevier, Amsterdam, 1981, pp. 759--809. 6 Parker, A.J., Paul, R.L. and Power, G.P., Aust. Chem., 34 (1981) 13. 7 Parker, A.J., Paul, R.L. and Power, G.P., J. Electroanal. Chem., 118 (1981) 305. 8 McMillan, R.S., Mackinnon, D.J. and Dutrizac, J.E., J. Appl. Electrochem., 12 (1982) 743--757. 9 Everett, P.K., The Dextec Copper Process, Extractive Metallurgy '81, Institute of Mining and Metallurgy, 1981, p. 149. 10 Kruesi, C.R., Allen, E.S. and Lake, J.L., Can. Min. Metall. Bull., (June 1973) 81. 11 Dahlstrom, D.A., Bacjek, F.A., Wojcik, B.C. and Emmett, R.C., Joint Meeting of the Mining and Metallurgy Institute of Japan and the American Institute of Mining and Metallurgical Engineers, Tokyo, 1980, pp. 23--40. 12 Oki, T. and Kammel, R., Annual Meeting of the American Institute of Mechanical Engineers New York, February 16--20, 1975, Preprint No. 75-B-55. 13 Narasagoudar, R.A., Johnson, J.W. and O'Keefe, T.J., Hydrometallurgy, 9 (1982) 37--55. 14 Scheiner, B.J., Lei, K.P.V. and Lindstrom, R.E., U.S. Bur. Mines, Rep. Invest. No. 8092 (1975). 15 Everett, P.K., The Dextec Lead Process, in: Osseo-Assare, K. and Miller, J.D. (Eds.), Hydrometallurgy -- Research, Development and Plant Practice, Metallurgical Society of AIME, New York, 1983, p. 165. 16 Scott, P.D. and Nicol, M.H., in: Bockris, J.O'M., Rand, D.A.J. and Welch, B.J. (Eds.), Trends in Electrochemistry, Plenum Press, New York, 1977, pp. 303--316. 17 Nicol, M.J. and Scott, P.D., J. South Aft. Inst. Min. Metall., (May 1979) 298--305.