Some Studies on Sulfuric Acid Leaching of Anode Slime with Additives

The most effective hydrometallurgical processing of anode slime is based on a chlorination route. However, it requires special attention with respect ...
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Ind. Eng. Chem. Res. 2002, 41, 6593-6599

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Some Studies on Sulfuric Acid Leaching of Anode Slime with Additives Jhumki Hait,*,† R. K. Jana,† Vinay Kumar,† and S. K. Sanyal‡ National Metallurgical Laboratory, Jamshedpur 831007, India, and Chemical Engineering Department, Jadavpur University, Jadavpur 700037, India

The most effective hydrometallurgical processing of anode slime is based on a chlorination route. However, it requires special attention with respect to pollution problems and construction materials. To minimize the problems, a new hydrometallurgical approach was pursued for the recovery of valuable metals and metalloids from copper electrorefining anode slime. The process involved the leaching of anode slime in sulfuric acid medium. At room temperature, sulfuric acid leaching without any additive resulted in poor recoveries of metals other than copper. With the addition of MnO2 to the leaching system at room temperature, an increase in the recovery values of copper, selenium, and tellurium was found. The recovery was further increased to 90% Cu, 37% Se, and 66% Te with an increase in temperature to 80 °C. However, the recoveries of Ni (13%), Au (trace), and Ag (9.5%) were poor. Satisfactory amounts of copper, selenium, tellurium, and gold were leached when both manganese dioxide and sodium chloride were added during leaching. The maximum recoveries were 90% Cu, 80% Se, 79% Te, and 77% Au at 50 °C after 40 min of leaching. The effects of various parameters such as time, temperature, solid/ liquid ratio, amount of additives, and acid concentration on the recoveries of metals were studied. The kinetics of leaching was also examined at different temperatures. The dissolution of metals was observed to be very fast with both manganese dioxide and sodium chloride addition even at room temperature. The leaching kinetics of selenium, tellurium, and gold with manganese dioxide and sodium chloride additives followed the ash diffusion control model, whereas copper leaching followed the mixed control (chemical and ash diffusion control) model. These leaching mechanisms were further corroborated with SEM findings. 1. Introduction Anode slime is the insoluble product deposited at the bottom of an electrorefining tank during the electrorefining of copper. The composition of the slime depends on the host ore that produces the anodes and also on the dissolution mechanism and the physicochemical characteristics of the impurities present in the anodes. The various metals present in the anode slimes are Cu, Ni, Se, Te, Ag, Au, platinum group metals (PGMs), Pb, Fe, Ba, etc.,1,2 and the common mineral phases in the anode slimes include CuSO4‚5H2O, NiO, Cu2Te, Cu2Se, Ag2Se, and CuAgSe, among others. The slimes are collected periodically and sent to a byproduct metals recovery plant for processing. Depending on the composition and morphology of the anode slime, a number of processes following pyro-, pyrohydro-, hydropyro-, and hydrometallurgical routes have been developed for the extraction of valuable metals from anode slime. One of the well-known pyrometallurgical processes is slime smelting in a Dore furnace for the recovery of selenium from the flue gases, followed by sodium nitrate treatment for the recovery of precious metals.3,4 The pyrohydrometallurgical method of processing the anode slime includes oxidizing roasting with KNO3 at 400450 °C, followed by leaching with sulfuric acid or dilute NaOH at 80-100 °C.5 Some of the hydropyrometallurgical processes are sulfuric acid leaching followed by * Corresponding author address: Non-ferrous Process Division, National Metallurgical Laboratory, Jamshedpur 831007, India. Fax no. 091-0657-270527, phone No. 091-0657-271806, e-mail [email protected]. † National Metallurgical Laboratory. ‡ Jadavpur University.

sulfation roasting,6-8 sulfuric acid leaching followed by oxidation roasting,9,10 and sulfuric acid leaching followed by soda roasting.9,11-13 Wet chlorination is the only hydrometallurgical process developed for anode slime.14,15 In the present investigation, an attempt has been made to develop a suitable hydrometallurgical process for the recovery of the valuable metals present in the anode slime obtained from Indian Copper Complex, Ghatsila, India. The process involves leaching of anode slime in sulfuric acid medium with MnO2 and NaCl additives for the recovery of valuable metals. This new approach avoids problems such as the emission of corrosive gases in pyrometallurgical and pyrohydrometallurgical processes and chlorine evolution and corrosion problems encountered in the chlorine leaching process. 2. Materials and Methods Anode slime from the copper electrorefining plant of Indian Copper Complex, Ghatsila, India, was used for the experiments. It contained about 12% Cu, 37% Ni, 10.5% Se, 3.38% Te, 1.54% Ag, and 0.1% Au. The slime was ground to -200 mesh (BS) and dried in an air oven to remove free moisture. Samples for chemical analysis and leaching experiments were prepared by thoroughly mixing the dried powder and then using the cone and quartering method to prepare representative samples. Leaching experiments were carried out in conical flasks fitted in the clamps of a thermostatic water bath. Each conical flask had a separate stirrer controller that maintained the speed of the stirrer at 500 rpm. Each flask was charged with 2.5 g of solids, and the liquid/ solid ratio was maintained at 10:1 unless otherwise stated. AR-grade chemical reagents were used in all experiments. Chemical analyses of the as-received

10.1021/ie020239j CCC: $22.00 © 2002 American Chemical Society Published on Web 11/13/2002

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Table 1. Recoveries of Metals in Sulfuric Acid Medium without Any Additive pressure atmospheric pressure

time (min)

temp (°C)

acid conc (vol %)

Cu

240 240 240

30 60 80

20 20 20

45 50 58

% recovery Te Se 10.0 9.5 9.5

poor poor poor

Ni poor poor poor

sample and leach solutions were carried out by AAS, ICP-AES, and conventional methods. 3. Results and Discussion 3.1. Sulfuric Acid Leaching without Additive. Sulfuric acid is a good lixiviant, and therefore, it was used for leaching of the anode slime without any additive. Table 1 shows the recoveries of metals under various conditions. It can be observed from the table that, at ambient pressure, for an increase in leaching temperature from 30 to 80 °C the extraction of Cu increased from 45 to 58%. The Te recovery, though poor, remained almost constant at about 10% with the increase in leaching temperature. However, only negligible amounts of Se and Ni were extracted at all temperatures. 3.2. Leaching with Additives. 3.2.1. Sulfuric Acid Leaching with Manganese Dioxide Addition. Effect of MnO2 Addition. Leaching experiments with additives were carried out in the next stage, as the acid leaching without additive resulted in poor recoveries of all metals other than copper. Referring to the Eh-pH diagrams of selenium and tellurium in water,16 it can be seen that these metals can be leached at low pH and high oxidation potential. Therefore, the addition of an oxidizing agent was examined, as it might be helpful in the leaching of selenium and tellurium present in the anode slime. Manganese dioxide, a good oxidizing agent, was added to the leaching system. Table 2 shows the effect of MnO2 addition on the recoveries of Cu, Se, and Te at a temperature of 80 °C. It was found that, with an increase in MnO2 addition, the recoveries of all of the metals increased considerably. Because the recoveries of Ni, Ag, and Au were not appreciable, these metals were not analyzed in most of the leaching tests. The improvement in the recoveries with MnO2 addition might be due to quadrivalent manganese in MnO2 being reduced to divalent manganese and oxidizing Cu, Se, and Te, as follows

Cu2Se + 4Mn4+ f 2Cu2+ + Se4+ + 4Mn2+

(1)

Cu2Te + 4Mn4+ f 2Cu2+ + Te4+ + 4Mn2+

(2)

The recoveries of the metals with the addition of 1.25 g of MnO2 in the sulfuric acid leaching system were found to be about 90% Cu, 37% Se, and 66% Te (Table 2). Further increases in MnO2 addition did not improve the recoveries of the metals appreciably. Effect of Acid Concentration. Table 2 shows the effect of the sulfuric acid concentration on the extractions of Cu, Se, and Te in the leaching of anode slime with 1.25 g of MnO2 as an additive at 80 °C. It can be observed that, with an increase in acid concentration from 10 to 30 vol %, the Se recovery improved from 22 to 46%, and the Te recovery increased from 42 to 75%. There was a slight improvement in the recovery of Cu. However, it

was found that, with a 30 vol % acid concentration, the extraction of nickel also increased considerably along with those of Cu, Se, and Te. Nickel extraction at this stage is undesirable, as it unnecessarily complicates the separation process at a later stage. Therefore, 20 vol % sulfuric acid was used as the leachant in other leaching experiments. Effect of Temperature. The results of leaching experiments carried out at different temperatures are also shown in Table 2. It can be observed that, in 60 min of leaching at 30 °C, only 47% Cu and 10% Te were extracted. During this period, a negligible amount of Se (1.5%) was recovered. At 60 °C, the recovery of Cu increased to 65% in 60 min, whereas the recoveries of Se and Te were only 9 and 17%, respectively. On a further increase in temperature to 80 °C, the Cu recovery increased to about 82%, with the corresponding values for Se and Te being 25 and 34%, respectively. Thus, an increase in temperature was marked by an improvement in the recoveries of Cu, Se, and Te. Effect of Time. The leaching curves in Figure 1 show that the extractions of the metals increased with increasing time of leaching at 80 °C. For an increase in leaching time from 30 to 330 min, the recovery of copper improved from 74 to 90%. Similarly, the Se and Te recoveries also increased to 42 and 74%, respectively, in 330 min of leaching. It was seen that only the Te recovery depended strongly on the leaching time. However, no appreciable amounts of Ni, Ag, and Au were recovered at 80 °C even after 330 min. Kinetics of Leaching. An attempt was made to study the dissolution kinetics of Cu, Se, and Te in the leaching of anode slime with MnO2 addition using the shrinkingcore model. This is the most widespread model describing fluid-solid reaction kinetics of dense particles. The standard equations of this model are

x ∝ t: (i) film diffusion control, dense, constant-size small particles, all geometries (ii) chemical reaction control, dense, flat plate particles 1 - (1 - x)2/3 ∝ t: (i) film diffusion control, dense, shrinking spheres 1 - (1 - x)1/2 ∝ t: (i) film diffusion control, dense, large shrinking spheres (ii) chemical reaction control, dense, constant-size cylindrical particles 1 - (1 - x)1/3 ∝ t: (i) chemical reaction control, dense, constant-size shrinking spheres 1 - 3(1 - x)2/3 + 2(1 - x) ∝ t: (i) ash diffusion control, dense constant-size spherical particles x in the above equations is the fraction of metals extracted at time t. After examining the experimental results in comparison with the above equations, it was found that, at 80 °C, the kinetics of Se and Te (Figure 2) followed the ash diffusion control model, i.e., 1 - 3(1 - x)2/3 + 2(1 - x) ∝ t. The proportionality constant (Kp) for the above equation is known as the chemical rate constant, and it is

Ind. Eng. Chem. Res., Vol. 41, No. 25, 2002 6595 Table 2. Effect of Various Parameters on Recoveries of Metals effect of MnO2 additiona

effect of H2SO4 concentration

MnO2 (g)

Cu

% recovery Se

Te

[H2SO4] (vol %)

Cu

0.25 0.5 0.75 1.0 1.25

68 80 84 87 90

2.17 14 25.6 34 37.3

14 32 51 60 66

10 20 30 -

89 90 91 -

% recovery Ni Se 4.7 13 45 -

22.2 37.3 45 -

effect of temperature Te

temp (°C)

Cu

% recovery Se

Te

42 66 75 -

30 60 80 -

47.4 65 82.5 -

1.5 8.6 25 -

10.8 17 34.1 -

a Conditions: temperature ) 80 °C, time ) 240 min, [H SO ] ) 20 vol %. b Conditions: temperature ) 80 °C, time ) 240 min, MnO 2 4 2 ) 1.25 g. c Conditions: time ) 60 min, [H2SO4] ) 20 vol %, MnO2 ) 1.25 g.

Figure 1. Effect of time on recoveries of metals at 80 °C.

Figure 3. Scanning electron micrograph of untreated anode slime: (1) BaSO4, (2) NiO, (3) selenides and tellurides of Cu and Ag, (4) copper selenide.

Figure 2. Leaching kinetics with MnO2 addition at 80 °C.

related to the other parameters as follows

Kp ) 2bMDC/Fr2 min-1 where D is the diffusion coefficient, F is the particle density, r is the radius of the unreacted particle, M is the molecular weight of the particle, C is the concentration of the reactants, and b is the stoichiometry factor. Because Cu was leached appreciably during the initial stage of leaching, its kinetics could not be determined. The ash diffusion model was further confirmed by examining the leached residue under scanning electron microscope with EDX studies. The porous structure of the Se-rich material provided an indication that elemental Se deposited during the early stage of leaching. This deposited Se acted as a porous layer favoring ashdiffusion-controlled reaction kinetics. The SEM micrographs of the unleached anode slime and residue of the leached anode slime (at 80 °C) are shown in Figures 3 and 4.

Figure 4. Scanning electron micrograph of the residue after leaching the anode slime at 80 °C: (1) Mn and selenides and tellurides of Ag, (2) selenides and tellurides of Cu and Ag.

3.2.2. Sulfuric Acid Leaching with Manganese Dioxide and Sodium Chloride Addition. In the preceding sections, it was shown that leaching with manganese dioxide addition both at room temperature and at high temperature did not yield high recoveries for any metal other than copper. Hence, it was decided to use sodium chloride as another additive in the leaching system. Figure 5 shows the effect of leaching duration with only sodium chloride addition at 80 °C on the recovery of metals from anode slime. It can be seen that the Cu and Te recoveries increased with time and that, in 180 min, 80% Cu and 31% Te could be recovered. However, the recoveries of the other metals were negligible. Effect of MnO2 and NaCl Addition. The next set of experiments was carried out with both manganese dioxide and sodium chloride as additives in the leaching

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Figure 5. Effect on leaching time with 3 g of NaCl on recovery of metals.

Figure 7. Effect of H2SO4 on recoveries of metals.

the leaching system can be described as follows

MnO2 + 2NaCl + H2SO4 f MnSO4 + Na2SO4 + Cl2 + H2O (3) Ag2Se + Cl2 + H2O f AgClV + H2SeO3 + HCl

(4)

Cu2Se + Cl2 + H2O f CuCl2 + H2SeO3 + HCl (5) Cu2Te + Cl2 + H2O f CuCl2 + H2TeO3 + HCl (6)

Figure 6. Effect of MnO2 on recoveries of metals.

system. The effects of various amounts of MnO2 addition with 3 g of NaCl in the leaching system are shown in Figure 6. The addition of NaCl was fixed to 3 g, because preliminary experiments had shown that this was the optimum amount. It can be seen from the figure that, at 80 °C, the Se and Te recoveries increased significantly with the increase in MnO2 addition from 0.25 to 2.5 g. It was also noted that, in 240 min of leaching with 1.25 g of MnO2, metal recoveries on the order of 76.4% Cu, 60% Se, and 62% Te were achieved. The recovery of Se was much higher than that (35% Se) obtained by leaching without sodium chloride addition in 240 min (Figure 1). One of the reasons for the increasing metal recoveries might be the increase in H+ activity with sodium chloride addition in the sulfuric acid medium.17 It is known from the Eh-pH diagrams that increases in both the acidity and the oxidation potential are beneficial for the leaching of Se and Te.16 Thus, the increase in H+ activity upon sodium chloride addition might have contributed to the enhanced recoveries of these metals. Another reason was that, upon addition of sodium chloride and manganese dioxide to dilute sulfuric acid, chlorine was generated. Highly oxidizing chlorine, perhaps, took part in the reaction by disspoling Se and Te more effectively. However, no chlorine evolution was noted during leaching with MnO2 and NaCl additives. Perhaps, whatever chlorine was generated was instantly utilized in the leaching reaction. The dissolution reactions of Se and Te with the addition of sodium chloride and manganese dioxide in

NaCl + H2SO4 f Na2SO4 + HCl

(7)

MnO2 + HCl f MnCl2 + Cl2 + H2O

(8)

It might also be noted that the nickel recovery did not increase to an appreciable level upon the addition of sodium chloride along with manganese dioxide in the leaching system. This might be beneficial for the present system, as the decreased concentration of recovered nickel would reduce the complications for metal separation by solvent extraction at a later stage. However, nickel could be extracted from the residue in a second stage of leaching with a higher acid concentration at elevated temperature.18 There was no silver recovery, which perhaps, precipitated as silver chloride with the addition of sodium chloride in the leaching system. Gold recovery was evaluated at a later stage when the effect of time was studied. Effect of Acid Concentration. Figure 7 shows the effect of the sulfuric acid concentration on the extractions of Cu, Se, and Te after 90 min of leaching at 80 °C. There was no significant change in the recoveries of Cu and Te with increasing sulfuric acid concentration. However, the Se recovery was a maximum (74%) at 20 vol % acid. The maximum recoveries of Cu and Te were 88 and 78%, respectively, at 20 vol % sulfuric acid. Effect of Time. It was found that the recoveries of Cu, Se, Te, and Au almost reached their maxima in 30 min of leaching at 80 °C. However, the faster rate of recovery of the metals resulted in difficulty in determining the reaction kinetics at 80 °C. Therefore, leaching experiments in the presence of manganese dioxide and sodium chloride were performed at 30, 40, and 50 °C for different durations. Table 3 shows the leaching recoveries of the metals at 30, 40, and 50 °C. The extraction of the metals increased with increasing time of leaching at different temperatures. It can be observed that the maximum recoveries of Se, Te, and Au were 70, 70, and 71%, respectively, at 30 °C. At 40 °C, the recoveries of the metals increased to 75% Se, 74% Te, and 77% Au.

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Figure 8. Leaching kinetics for Se.

Figure 9. Leaching kinetics for Te.

Table 3. Effect of Time at Various Temperatures on Recoveries of Metals temp ) 30 °C

temp ) 40 °C

% recovery

% recovery

time (min) Cu 5 15 25 35 40

70 78 83 85 86

Se

Te

30.2 50.8 63.3 67 70

25.4 46.5 59.5 66.8 70.4

Au Cu 29 52 63 69 71

72.4 79.3 86.7 87.7 88.9

Se

Te

39.2 56 66 73 75

36.3 53.4 64 72.7 73.8

temp ) 50 °C % recovery Au Cu 37 56 68 74 77

81.6 86 88.8 91.2 91

Se

Te

Au

44 63.1 71.7 79 80

53.7 68.9 76 79 78

49 70 77 77 77

There were further improvements in the recoveries of these metals at 50 °C. An increase in extraction of Se from 44% in 5 min to 80% in 40 min of leaching was observed at this temperature. However, the recoveries of Te and Au became constant at 50 °C after 30 and 20 min of leaching, respectively. They were 79% Te and 77% Au. It was also found that significant amounts of copper were extracted in leaching at 30, 40, and 50 °C (Table 3). At 30 °C, in 5 min of leaching, 70% of the copper was leached out, and the extraction increased with time to 86% in 40 min of leaching. The recovery increased to 89% in 40 min when the leaching was performed at 40 °C. At 50 °C, the maximum recovery of 91% Cu was obtained after 35 min of leaching, and it remained constant upon further increases in the time of leaching. Kinetics of Leaching. Attempts were made to determine the dissolution kinetics of Se, Te, and Au in the leaching of anode slime with both manganese dioxide and sodium chloride addition in terms of the shrinkingcore model. From the experimental results (Table 3) for Se, Te, and Au, it was found that the data could be best fitted according to the ash diffusion control model, i.e., 1 - 3(1 - x)2/3 + 2(1 - x) ∝ t. However, the experimental data for copper (Table 3) revealed that the mixed control model, i.e., chemical reaction control [1 - (1 - x)1/3 ∝ t] with ash diffusion control [1 - 3(1 - x)2/3 + 2(1 - x) ∝ t], is fitted well. The kinetics plots for Se, Te, Au, and Cu at all the three temperatures (30, 40, and 50 °C) are shown in Figures 8-11. During the leaching process insoluble silver chloride was formed, and this product layer might have contributed to the ash diffusion reaction kinetics of Se, Te, and Au. This was further confirmed by examining the leached residue under scanning electron microscope with EDX studies. The SEM micrographs of leached anode slimes are shown in Figures 12 and 13. To calculate the activation energy, the values of ln Kp, where Kp ) [1 - 3(1 - x)2/3 + 2(1 - x)]/t, were plotted

Figure 10. Leaching kinetics for Au.

Figure 11. Leaching kinetics for Cu.

against 1/T (Arrhenius plot) in Figure 14 for Se, Te, and Au. The activation energies of the overall reactions were calculated to be about 1.95 kcal/mol for Se, 3.65 kcal/ mol for Te, and 3.89 kcal/mol for Au. The activation energy value of a diffusion-controlled process is generally in the range of 1-3 kcal/mol.19 Therefore, this further confirms that the sulfuric acid leaching of Se, Te, and Au from anode slime in the presence of MnO2 and NaCl follows ash diffusion control reaction kinetics. 4. Conclusions 1. Sulfuric acid leaching of anode slime without any additive results in poor recoveries of metals (Ni, Se, Te, Au, and Ag) other than Cu.

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79% Te, and 77% Au were observed after 40 min of leaching at 50 °C. 5. The reaction kinetics of Se, Te, and Au follow the ash diffusion control model in sulfuric acid leaching using both MnO2 and NaCl additives. SEM-EDX studies confirm that the porous silver chloride product layer formed during leaching contributes the ash diffusion control kinetics. The reaction kinetics of Cu follows the mixed control model. 6. The activation energies of the overall reactions are found to be about 1.95 kcal/mol for Se, 3.65 kcal/mol for Te, and 3.89 kcal/mol for Au. Acknowledgment Figure 12. SEM microstructure for the morphology of the residue after leaching of anode slime with MnO2 and NaCl: (1) silver selenide, (2) AgCl.

The authors are thankful to Shri Premchand, Head, NFP Division, NML, Jamshedpur, India, for his advice and help in various ways in carrying out this work. Anode slime samples received from M/S ICC, Ghatsila, India, are also thankfully acknowledged. Literature Cited

Figure 13. SEM microstructure for the morphology of the residue after leaching of anode slime with MnO2 and NaCl: (1) porous structure of AgCl.

Figure 14. Arrhenius plots for Se, Te, and Au.

2. Leaching in sulfuric acid in the presence of MnO2 does not improve the recoveries of the metals significantly at room temperature, but at 80 °C, maxima of 90% Cu, 42% Se, and 74% Te were recovered in 330 min of leaching. The reaction kinetics for Se and Te at 80 °C follow the ash diffusion control model. 3. Leaching of anode slime in sulfuric acid in the presence of only NaCl does not improve the recovery of the metals significantly. 4. With both MnO2 and NaCl addition in the sulfuric acid leaching system, the recoveries of Cu, Se, Te, and Au improve. Maximum recoveries of 91% Cu, 80% Se,

(1) Hoffmann, J. E. Processing slimes: The base case and opportunities for improvement. JOM 1990, 42 (8), 38. (2) Newton, Joseph; Wilson, C. L. Metallurgy of Copper; Chapman & Hall Ltd.: London, 1942. (3) Dixon, C. P. Gold and silver refining at the electrolytic refining and smelting company of Australia Ltd., Port Kembla, N.S.W. In Mining and Metallurgical Practices in Australasia (The Sir Maurice Mawley Memorial Volume); Woodcock, J. T., Ed.; Monograph Series No. 10; The Australasian Institute of Mining and Metallurgy: Parkvile, Victoria, Australia, 1980; pp 519-521. (4) Dixon, C. P. Selinium recovery at the electrolytic refining and smelting company of Australia Ltd., Port Kembla, N.S.W. In Mining and Metallurgical Practices in Australasia (The Sir Maurice Mawley Memorial Volume); Woodcock, J. T., Ed.; Monograph Series No. 10; The Australasian Institute of Mining and Metallurgy: Parkvile, Victoria, Australia, 1980; pp 628-629. (5) Jung, V.; Deierling, B.; Karl, R. Recovery of precious metals from materials containing selenium, tellurium, arsenic and antimony. Chem. Abstr. 1983, 99, 198566z. (6) Monahan, R. K.; Loewen, F. Treatment of anode slimes at the Inco Copper Refinery. Presented at the Canadian Institute of Mining and Metallurgy Annual Conference, Halifax, Nova Scotia, Aug 1972. (7) Tishchenko, A. A. Extraction of selenium and tellurium from copper electrolytic slimes. Chem. Abstr. 1964, 61, 3955. (8) Kunev, D.; Vasilev, Kh.; Chimbulev, M.; Karagozov, L. Study of the sulfation roasting of electrolytic copper slimes. Metallurgiya (sofia) 1977, 32 (8), 14-16; Chem. Abstr. 1978, 88, 64561. (9) Cooper, W. C. The treatment of copper refinery anode slime. JOM 1990, 42 (8), 45-49. (10) Morrison, B. H. Slimes Treatment Process. U.S. Patent 4,047,939, Sep 13, 1977. (11) Buketov, E. A.; et al. Shaft furnace sintering of electrolytic copper slimes. Tsvetn. Met. 1965, 38 (4), 28-31; Chem. Abstr. 1966, 63, 12722. (12) Shkodin, V. G.; Malyshev, V. P.; Evtyukhova, O. V.; Malkov, V. E.; Galimova, S. A. Roasting-sintering of anode slimes with soda in shaft furnaces. Tr. Ural. Nauch. Issled. Proekt. Inst. Mednoi Prom 1969, 12, 204-208; Chem. Abstr. 1970, 72, 114078. (13) Victorovich, G. S.; Bell, M. C.; Sridhar, R.; Raskauskas, J. Novel soda ash process for the recovery of selenium from anode slimes. Presented at the 109th AIME Annual Meeting, Las Vegas, NV, Feb 24-28, 1980. (14) Hoffmann, J. E. The wet chlorination of electrolytic refinery slimes. JOM 1990, 42 (8), 50-54. (15) Toraiwa, A.; Abe, Y. New hydrometallurgical process of copper anode slimes at Saganoseki Smelter and Refinery. Presented at the Second International Conference on Processing Materials for Properties, San Francisco, CA, Nov 5-8, 2000.

Ind. Eng. Chem. Res., Vol. 41, No. 25, 2002 6599 (16) Pourbaix, M. Atlas of Electrochemical Equilibrium in Aqueous Solution; Pergamon Press: Oxford, U.K., 1966 (Translated by James A. Franklin). (17) Majima, H.; Awakura, Y.; Yazaki, T.; Chikamori, Y. Acid dissolution of cupric oxide. Met. Trans. 1980, 11B, 209-214. (18) Morrison, B. H. Recovery of Silver and Gold from Refinery Slimes at Canadian Copper Refiners. In Extractive Metallurgy 1985; The Institute of Mining and Metallurgy: London, 1985; pp 249-269.

(19) Habashi, F. Extractive Metallurgy, General Principles; Gordon & Breach: New York, 1969; Vol. I.

Received for review April 1, 2002 Revised manuscript received September 11, 2002 Accepted September 13, 2002 IE020239J