- Email: [email protected]
Processing of copper anodic-slimes for extraction of valuable metals
Waste Management 23 (2003) 763–770 www.elsevier.com/locate/wasman
Processing of copper anodic-slimes for extraction of valuable metals A.M. Amer* Environmental Science Department, Faculty of Science, Alexandria University, Mohrrem bek, Alexandria, Egypt Accepted 2 April 2003
Abstract This work focuses on processing of anodic slimes obtained from an Egyptian copper electrorefining plant. The anodic slimes are characterized by high concentrations of copper, lead, tin and silver. The proposed hydrometallurgical process consists of two leaching stages for the extraction of copper (H2SO4–O2) and silver (thiourea–Fe3+), and pyrometallurgical treatment of the remaining slimes for production of Pb–Sn soldering alloy. Factors affecting both the leaching and smelting stages were studied. # 2003 Elsevier Ltd. All rights reserved.
1. Introduction Considering the depleting reserves of primary mineral resources for the extraction of valuable metals, every effort should be made to process secondary sources such as slag, slimes etc. Electrolytic refining of copper is carried out in Jumbo-cells where a series of cathodes made of pure copper and insoluble lead anodes were placed in steel tanks. Metallic copper is deposited at the cathodes, while the water molecules are decomposed at the anode to produce oxygen. Sulphuric acid is regenerated in the solution, which is recirculated (Landsberg, 1977; Subbiah et al., 1986; Mahdavi and Ditze, 1996). The electrolysis of copper may be described as follows: Cu2þ þ SO2 CuSO4! 4 At the anode :
At the cathode :
Cu2þ þ 2e ¼ Cu H2 O þ 2e ¼ 0:1=2O2 þ 2Hþ
CuSO4 þ H2 O ¼ Cu þ H2 SO4 þ 1=2O2
ð1Þ
ð2Þ
ð3Þ
ð4Þ
* Tel.: +20-3-0733161; fax: +20-3-3911794. E-mail address: [email protected] (A.M. Amer). 0956-053X/03/$ - see front matter # 2003 Elsevier Ltd. All rights reserved. doi:10.1016/S0956-053X(03)00066-7
Anodic slimes are collected from the bottom of the electrolytic cells during the refining of copper. There are two different kinds of slimes depending on the sources from which they are obtained. The first one is produced during the processing of copper concentrate and has a relatively high concentration of gold, silver, tellurium and selenium. The second slime is produced during the processing of recycled scrap and has higher concentrations of lead, copper, tin and silver than the first one. This work focuses on the processing of anodic slimes obtained from an Egyptian copper electrorefining plant. The anodic slimes are characterized by high concentrations of copper, lead, tin and silver. Several methods have been attempted to recover copper from slags and anodic slimes. These methods can be differentiated into: (a) pyrometallurgical processing which includes roasting in presence of an oxidizing agent, suflat-roasting and soda-ash processing (Ying, 1983; Swayn et al. 1993; Hughes, 2000; Filipov et al. 2000) and (b) hydrometallurgical processing using different leachants namely chlorine, nitric acid and sulphuric acid (Gill, 1980; Holmes, 1981; Everett, 1994; Petrov et al., 1999). From the previously mentioned processes which deal with processing of copper anodicslimes, two of them have been successful on an industrial scale, namely roasting and pressure leaching (Gaylarde and Videla, 1995; Belen’kii et al., 1999). The production of sulphuric acid is the troublesome aspect of a roasting process, as the marketability of
764
A.M. Amer / Waste Management 23 (2003) 763–770
sulphuric acid is highly dependent on site location. Furthermore, even if gas emission regulations are complied with, harmful SO2 may be released. The advantages of the roasting process are its simplicity and its attractive costs compared with other processes, as well as the fact that the technology is available worldwide. On the other hand pressure leaching results generally in a high degree of extraction, and produces wastes that can be ultimately confined as stable solids. Hydrometallurgical processing of copper anodic slimes has the following advantages: 1. Capital costs are low compared to those for a smelter. 2. It may be applied to small and large operations. 3. Air pollution by sulfur dioxide is eliminated. 4. The equipment may be designed and installed in modules. In this investigation, a sulphuric acid pressure leaching process was proposed among the other hydrometallurgical processes to treat the copper-anodic slimes for extraction of copper. The slimes without the copper are then enriched in lead, tin and silver. Many hydrometallurgical processes have been reported for extraction of valuable metals including cyanide leaching (McClincy, 1990) and thiourea leaching. The latter seems to be an effective reagent for silver extraction due to its high complexing ability. Thiourea leaching has been used in hydrochloric sulphuric and nitric acid media. Dissolution of gold and silver of up to 97% was reported in a solution containing 20 g/l HCl, 0.20 g/l H2O2 and 100 g/l thiourea (Schulze, 1984). Soviet researchers (Kusnierov et al., 1993; Kucha and Cichowsa, 2001) have reported that thiourea is very effective in the recovery of valuable metals from copper concentrates that are especially refractory to cyanidation. An industrial application of thiourea leaching has been reported in New South Wales (Deschenes and Ghali, 1988), where thiourea leaching shows a good potential application to treat sulphide concentrates. The low toxicity of thiourea compared to cyanide is always mentioned as one of the motivations for developing the research in the field.
absorption spectrophotometer. The results of the chemical analysis are given in Table 1.
3. Apparatus and methodology A schematic diagram of the experimental apparatus in Fig. 1 shows the 1-l glass reactor with baffles used for the leaching tests. The reactor cover had four groundglass parts, which contained a Friedrich’s condenser, a sampling device, an oxygen dispersion tube and a glass impeller. The impeller was inserted through the center by means of a Chesapeake stirrer connection and was driven by a Fisher mixer. The entire reactor assembly was clamped into place in a water bath which was heated with a temperature controller and maintained at within 0.1 C of the desired temperature. In each test run, 100 g of ground anodic slime (0.075 mm) were treated with a 1-l solution containing a certain amount of sulphuric acid of known strength. The leaching tests were conducted at different temperatures for various time intervals with the admission of variable
2. Materials and apparatus Slime samples were collected from the bottom of the electrolytic cells during the electro-refining of copper. The slimes were sieved (2.0 mm) to remove the large inclusions of copper and other extraneous materials present in the slime. A representative sample was analysed for Cu, Pb, Sn, Zn, Ni, Fe, As, Sb, SiO2, MgO and Ag using an atomic
Fig. 1. Leaching apparatus.
A.M. Amer / Waste Management 23 (2003) 763–770
rates of oxygen as an oxidizing agent for copper dissolution. Samples of 5 ml of liquid were taken for subsequent analysis using atomic absorption for the determination of dissolved copper. At the conclusion of the leaching experiment, the slurry was filtered using a water suction pump through Table 1 Chemical analysis of representative sample of anodic slime studied Chemical constituents
Weight (%)
Cu Pb Sn Ag Ni Fe Sb As Zn S SiO2 CaO MgO Au Se Te
18.57 15.3 9.8 0.46 0.94 1.72 1.30 0.49 0.74 9.22 8.51 1.54 0.82 – – –
765
A-No-42 Whatman filter paper and washed thoroughly with hot distilled water, dried at 110 C, and then ground to 0.075 mm. The slime without the copper was analysed for Pb, Sn and Ag. Decopperised slime (40 g) was treated with a 1-l solution containing different amounts of acidified thiourea and ferric ion as oxidant. The slurry was agitated with a magnetic stirrer at various agitation speeds. Factors such as temperature, leaching time, thiourea concentration were taken into consideration to determine the optimum operating conditions for silver dissolution. A liquor sample was analysed using atomic absorption for determination of dissolved silver. The decopperised, silver-free residue was enriched in lead and tin, and was considered a base for the production of Pb–Sn soldering alloys. Decopperised, silver-free residue (50 g) was mixed with proper proportions of sodium carbonate and carbon, smelted in a graphite available at different temperatures and time for production Pb–Sn soldering alloy. The purpose of the investigation, as shown in Fig. 2, was to investigate and optimize the parameters influencing the dissolution of both copper and silver, as well as those affecting the production of soldering alloys.
Fig. 2. Proposed technique for processing of Egyptian Cu anodic slimes.
766
A.M. Amer / Waste Management 23 (2003) 763–770
where: A=constant, EA=activation energy in kJ/mole, R=gas constant, and T=temperature in k. The activation energy is determined from the Arrhenius plot given in Fig. 5 as follows:
4. Results and discussions 4.1. Extraction of copper 4.1.1. Effects of temperature Fig. 3 shows that the copper extraction increased with an increase in temperature in the range 65–85 C. Further increase in temperature had little effect on copper extraction. The following equation can be applied: 1 ð1 Þ1=3 ¼ kt
ð5Þ
where: =fraction reacted, k=a constant for the reaction rate, and t=duration of the reaction. Fig. 4 shows that the plot of 1(1a)1/3 versus t over the studied temperature range consists of a series of straight lines indicating that the sulphuric acid leaching of anodic slimes is llimited by the chemical concentrations. From the slopes of linearised isotherms in Fig. 4, certain isotherms were selected for rate constants at given temperatures. Using the Arrhenius equation: K ¼ A:eE=RT
ð6Þ
Slope ¼ EA =R
ð7Þ
The value calculated for the activation energy is 69 kJ/mole, which indicates surface chemical control. 4.1.2. Effect of acid concentration Fig. 6 shows the effect of charging the initial solution with acid. The rate of copper extraction increases when the acid concentration increases from 0.02 to 0.5 M. Further increase in acid concentration resulted in no significant increase in copper extraction. This may be due to the fact that the solubility of oxygen in high acid concentrations is low in comparison with its solubility in low acid concentrations. Fig. 7 illustrates the relation of pH value of the leaching solution and extraction, which indicates that the rate of metal uptake increases with increase of pH up to pH 4–5. Thereafter, it is more or less constant in the pH range investigated.
Fig. 3. Effect of temperature on copper extraction.
Fig. 5. Arrhenius plot for determination of activation energy.
Fig. 4. A plot of 1(1a)1/3 versus time for various temperatures.
Fig. 6. The effect of sulphuric acid concentration on copper extraction.
A.M. Amer / Waste Management 23 (2003) 763–770
4.1.3. Effect of oxygen partial pressure Fig. 8 shows the results of leaching 75 mm diameter anodic slime in 0.5 M H2SO4 under various oxygen partial pressures. It is concluded that, the higher the oxygen partial pressure, the higher the extraction of copper under the experimental conditions used. The maximum dissolution of anodic slime in sulphuric acid under oxygen pressure led to the extraction of copper (97%) under the following leaching conditions: Temperature: 85 C Sulphuric acid concentration: 0.5 M Oxygen partial pressure: 1 bar Leaching time: 90 min Particle size: 75 mm
767
4.2.1. Effect of thiourea concentration In an acidic solution of thiourea, the dissolution of silver proceeds according to the following reaction: Au CSðNH2 Þ2 þ þe Ag þ 2 CSðNH2 Þ2! ð8Þ 2 This is a reversible reaction with an electrode potential of 0.352 V at 25 C on a fresh silver surface. As shown in Fig. 9, the silver extraction increased with the increase of thiourea concentration and reached its maximum (90%) at 10 g/l thiourea concentration. An increase in thiourea concentration up to 10 g/l, produced an increase in the silver extraction, especially in the initial leaching period (0.5 h). There is, however, a decrease in silver extraction when thethiourea concentration exceeded 20 g/l, possibly because the thiourea is less stable at higher concentrations and decomposes.
4.2. Extraction of silver This part represents the experiments conducted for the extraction of silver from decopperized slime residue. The variables influencing the potential of using acidified thiourea solution as leaching agent are: thiourea concentration, temperature, leaching time, and ferric-ion concentration as oxidant. The decopperized slime residue is ground to 0.075 mm and then leached by acidified thiourea solution.
4.2.2. Effect of temperature The effect of temperature on extraction rate is studied, using 10 g/l thiourea and 5.0 g/l ferric ion. Extraction of more than 90% silver is reached at both 40 and 60 C after 30 min leaching time. The silver extraction is improved with the increase of temperature to 40 C where it reaches 93.5% after 4 h leaching time. 4.2.3. Effect of addition of ferric ion The oxidating effect of ferric ion in the presence of thiourea can be illustrated in the following reaction. þ Fe3þ ¼ CSðNH2Þ2 ð9Þ 2 ! Fe CSðNHÞ2 2 In contact with oxidants such as Fe+3, thiourea may be oxidized in successive stages to form several products. The first is the formation of formamidine disulphide (Schulze, 1984). ½ðNH2 ÞðNHÞCSSCðNHÞðNH2 Þ 2CSðNH2 Þ2! þ 2Hþ þ 2e; Eo ¼ 0:42 V:
ð10Þ
Fig. 7. Effect of pH on copper extraction.
In a second, irreversible reaction, the formamide disulphide yields thiourea and an unidentified sulphinic
Fig. 8. The effect of oxygen flow rate on copper extraction.
Fig. 9. Influence of thiourea concentration.
768
A.M. Amer / Waste Management 23 (2003) 763–770
compound, which then decompose to yield cyanomide and elemental sulphur. The elemental sulphur created at the end of this decomposition stage has a detrimental effect by passivating the leachable silver surface and precipitating some silver from solution. The study of the effect of oxidizing agent shows that there is an equilibrium between thiourea and formamidine disulfide. The leaching efficiency of silver depends on this equilibrium. The results indicate that the optimum extraction of silver was 93% using 10 g/l thiourea and 5.0 g/l Fe3+ in one leaching run of 4 h. As seen in Fig. 10, an addition of 2 g/l of Fe+3 slightly improves the initial leaching kinetics, but results in no real improvement in the rate of silver extraction. Further increase in Fe3+ to 5.0 g/l increases the leaching kinetics and also increases the extraction rate of silver to 93.4% in 7 h. The rate of silver dissolution is very rapid in the first 30 min of leaching. For the 5.0 g/l concentration of the oxidant, 80% of silver was leached in this period. Adding 8.0 g/l Fe3+ in the solution caused a significant decrease in silver extraction. When the concentration of oxidant is too large, oxidation of thiourea occurs and the remaining leachant is less effective in extracting silver.
Fig. 11 illustrates the relationship between the amount of carbon added and the metal content in leadtin soldering alloys. The figure shows that as the concentration of carbon increases, the lead content decreases and the tin content increases. The addition of > 20% carbon led to an increase of the tin content at the expense of a decrease in lead content. The results demonstrate that a concentration of 30% carbon is the optimum value. Fig. 12 represents the relationship between the percentage of sodium carbonate added and metal content (lead–tin) in soldering alloys. With the increase of sodium carbonate, a slight decrease of Pb content and slight increase of tin content was observed. Therefore, a concentration of 40% sodium carbonate is preferred. Fig. 13 reveals the relationship between temperature and percentage of metal content in soldering alloys. The figure shows that the content of lead in the alloy increases over the entire temperature range (900– 1300 C), while the tin content increases up to 1100 C and then decreases. Therefore, the optimum temperature is 1100 C.
4.3. Production of lead–tin soldering alloy After the removal of the copper and silver, the slimes were then enriched in lead (41.6%) and tin (29.0%). Residues were dried and ground to < 0.075 mm and then mixed with different proportions of sodium carbonate and carbon. The mixtures were smelted in graphite crucibles at different temperature levels and time intervals. The effects of the following parameters were studied: Carbon as reducing agent. Sodium carbonate as flux. Temperature. Reduction time.
Fig. 10. Influence of Fe3+ concentration on silver extraction.
Fig. 11. Effect of added carbon on lead–tin content.
Fig. 12. Effect of added sodium carbonate on lead–tin content.
A.M. Amer / Waste Management 23 (2003) 763–770
769
The effect of time on lead and tin content is shown in Fig. 14. The increase in reduction time from 1.0 to 3.0 h resulted in no significant increase in lead and tin content. Consequently, a reduction time of 1.0 h was chosen as a favourable value. Based on the experimental results, the more favourable operating conditions for production of soldering alloy can be summarized as follows:
However, a wide range of Pb–Sn soldering alloy is known depending on the Pb/Sn ratio. Therefore, by addition of the required quantity of tin before smelting, the tin content of the soldering alloy could be increased and adjusted with respect to the required lead content.
Carbon content: 30% (by weight of slime residue). Sodium carbonate content: 40% (by weight of slime residue). Temperature: 1100 C Reduction time: 1.0 h
About 90% of copper initially present in anodic slimes with 75 mm diameter particles dissolved in 0.25 M H2SO4 solutions at 85 C under 1 bar oxygen pressure within 90 min leaching time. The apparent activation energy is 69.0 kJ/mole characteristic of a controlled chemical reaction. Thiourea leaching of silver appears to be a good potential application to treat slimes containing silver to recover silver. The low toxicity of thiourea compared to cyanide is one of the motivations for developing the research in this field. Using concentration of thiourea of 10 g/l, under temperature 60 C in 30 min leaching time, 92.2% of silver is extracted. Smelting of the slimes after removal of copper and silver at 1100 C in the presence of sodium carbonate (40%) and carbon (30%) for 1 h led to production of Pb–Sn soldering alloy with the following characteristics: Pb: 53.0%–Sn: 33%.
Under the above operating conditions, the lead and tin in the soldering alloy had the following concentrations: Pb 53.0% Sn 33.0%
5. Conclusions
References
Fig. 13. Effect of temperature on lead–tin content.
Fig. 14. Effect of reduction time on lead–tin content.
Belen’kii, A.M., Petrov, G.V., Plekhanov, K.A., Kozlovskaya, A.E., Greiver, T.N., 1999. Hydrometallurgische Technologie zur verarbeitung von kupferelektrolyseschlammen. Cvetnye metally 74 (1), 43–45. Deschenes, G., Ghali, E., 1988. Leaching of gold from chalcopyrite concentrates by thiourea. Hydrometallurgy 20, 179–202. Everett, P.K. (1994) Development of copper process by an international consortium. Proceedings of Hydrometallurgy 94, England. Filipov, Ju.A., Anisimova, N.N.l., Kotuchova, A.K., Ter-Oganesjane, A.K., Tichov, I.V., 2000. Rekonstruktion der anodenschlammverabeitung. Cventnye metally 75 (6), 61–63. Gaylarde, C.C. and Videla, H.A. (1995) Bioextraction and biodeterioration of metals. Cambridge University Press, pp. 51–55. Holmes, J.A. (1981) Recent experience in the application of hydrometallurgical techniques. Proceedings of Hydrometallurgy 81, England. Gill, C.B., 1980. Non Ferrous Extractive Metallurgy. Wiley-Interscience, New York. Hughes, S., 2000. Applying ausmelt technology to recover Cu, Ni and Co from slags (overview). JOM 30–33. Kusnierov, M., Sepelk, V., Briancen, J., 1993. Effects of biodegradation and mechanical activation on gold recovery by thiourea leaching. JOM 54–56. Kucha, H., Cichowska, K., 2001. Precious metals in copper smelting products. Physicochemical Problems of Mineral Processing Journal 35, 91–101. Landsberg, R., 1977. Elektrochemische Reaktionen und Prozesse. VEB Deutscher Verlag der Wissenschaften. Mahdavi, M., Ditze, A., 1996. Basic investigations of the dissolution of copper and the formation of anodic slime during copper electrolysis with pure copper-oxygen anodes. Erzmetall 49 (6), 358–365. McClincy, R.J., 1990. Unlocking refractory gold ores and concentrates. JOM 10–11.
770
A.M. Amer / Waste Management 23 (2003) 763–770
Petrov, G.V., Plekhanov, K.A., Kozlovskoya, A.E., Greiver, T.N., 1999. Hydrometallurgische Technologie zur verarbeitung von kupferelektroly schlammen. Cventnye metally 74 (1), 43–45. Schulze, R.G., 1984. New aspects in thiourea leaching of precious metals. Int. Precious Metals Symposium, Los Angeles 27–29. Schulze, R.G., 1984. New aspects in thiourea leaching of precious metals. Int. Precious Metals Symposium, Los Angeles 27–29.
Subbaiah, T., Das, S.C., Das, R.P., 1986. Electrowinning of copper under forced convection. Erzmetall 34 (10), 501–506. Swayn, G.P., Robilliard, K.R., Floyd, J.M., 1993. Applying ausmelt processing to complex copper smelter dusts. JOM 35–38. Ying, C.H. (1983) Copper refinery anodic slime. Proceedings of the Third Int. Symp. of Hydrometallurgy. Annual Meeting, Atlanta, Georgia, March 6–10.
Processing of copper anodic-slimes for extraction of valuable metals A.M. Amer* Environmental Science Department, Faculty of Science, Alexandria University, Mohrrem bek, Alexandria, Egypt Accepted 2 April 2003
Abstract This work focuses on processing of anodic slimes obtained from an Egyptian copper electrorefining plant. The anodic slimes are characterized by high concentrations of copper, lead, tin and silver. The proposed hydrometallurgical process consists of two leaching stages for the extraction of copper (H2SO4–O2) and silver (thiourea–Fe3+), and pyrometallurgical treatment of the remaining slimes for production of Pb–Sn soldering alloy. Factors affecting both the leaching and smelting stages were studied. # 2003 Elsevier Ltd. All rights reserved.
1. Introduction Considering the depleting reserves of primary mineral resources for the extraction of valuable metals, every effort should be made to process secondary sources such as slag, slimes etc. Electrolytic refining of copper is carried out in Jumbo-cells where a series of cathodes made of pure copper and insoluble lead anodes were placed in steel tanks. Metallic copper is deposited at the cathodes, while the water molecules are decomposed at the anode to produce oxygen. Sulphuric acid is regenerated in the solution, which is recirculated (Landsberg, 1977; Subbiah et al., 1986; Mahdavi and Ditze, 1996). The electrolysis of copper may be described as follows: Cu2þ þ SO2 CuSO4! 4 At the anode :
At the cathode :
Cu2þ þ 2e ¼ Cu H2 O þ 2e ¼ 0:1=2O2 þ 2Hþ
CuSO4 þ H2 O ¼ Cu þ H2 SO4 þ 1=2O2
ð1Þ
ð2Þ
ð3Þ
ð4Þ
* Tel.: +20-3-0733161; fax: +20-3-3911794. E-mail address: [email protected] (A.M. Amer). 0956-053X/03/$ - see front matter # 2003 Elsevier Ltd. All rights reserved. doi:10.1016/S0956-053X(03)00066-7
Anodic slimes are collected from the bottom of the electrolytic cells during the refining of copper. There are two different kinds of slimes depending on the sources from which they are obtained. The first one is produced during the processing of copper concentrate and has a relatively high concentration of gold, silver, tellurium and selenium. The second slime is produced during the processing of recycled scrap and has higher concentrations of lead, copper, tin and silver than the first one. This work focuses on the processing of anodic slimes obtained from an Egyptian copper electrorefining plant. The anodic slimes are characterized by high concentrations of copper, lead, tin and silver. Several methods have been attempted to recover copper from slags and anodic slimes. These methods can be differentiated into: (a) pyrometallurgical processing which includes roasting in presence of an oxidizing agent, suflat-roasting and soda-ash processing (Ying, 1983; Swayn et al. 1993; Hughes, 2000; Filipov et al. 2000) and (b) hydrometallurgical processing using different leachants namely chlorine, nitric acid and sulphuric acid (Gill, 1980; Holmes, 1981; Everett, 1994; Petrov et al., 1999). From the previously mentioned processes which deal with processing of copper anodicslimes, two of them have been successful on an industrial scale, namely roasting and pressure leaching (Gaylarde and Videla, 1995; Belen’kii et al., 1999). The production of sulphuric acid is the troublesome aspect of a roasting process, as the marketability of
764
A.M. Amer / Waste Management 23 (2003) 763–770
sulphuric acid is highly dependent on site location. Furthermore, even if gas emission regulations are complied with, harmful SO2 may be released. The advantages of the roasting process are its simplicity and its attractive costs compared with other processes, as well as the fact that the technology is available worldwide. On the other hand pressure leaching results generally in a high degree of extraction, and produces wastes that can be ultimately confined as stable solids. Hydrometallurgical processing of copper anodic slimes has the following advantages: 1. Capital costs are low compared to those for a smelter. 2. It may be applied to small and large operations. 3. Air pollution by sulfur dioxide is eliminated. 4. The equipment may be designed and installed in modules. In this investigation, a sulphuric acid pressure leaching process was proposed among the other hydrometallurgical processes to treat the copper-anodic slimes for extraction of copper. The slimes without the copper are then enriched in lead, tin and silver. Many hydrometallurgical processes have been reported for extraction of valuable metals including cyanide leaching (McClincy, 1990) and thiourea leaching. The latter seems to be an effective reagent for silver extraction due to its high complexing ability. Thiourea leaching has been used in hydrochloric sulphuric and nitric acid media. Dissolution of gold and silver of up to 97% was reported in a solution containing 20 g/l HCl, 0.20 g/l H2O2 and 100 g/l thiourea (Schulze, 1984). Soviet researchers (Kusnierov et al., 1993; Kucha and Cichowsa, 2001) have reported that thiourea is very effective in the recovery of valuable metals from copper concentrates that are especially refractory to cyanidation. An industrial application of thiourea leaching has been reported in New South Wales (Deschenes and Ghali, 1988), where thiourea leaching shows a good potential application to treat sulphide concentrates. The low toxicity of thiourea compared to cyanide is always mentioned as one of the motivations for developing the research in the field.
absorption spectrophotometer. The results of the chemical analysis are given in Table 1.
3. Apparatus and methodology A schematic diagram of the experimental apparatus in Fig. 1 shows the 1-l glass reactor with baffles used for the leaching tests. The reactor cover had four groundglass parts, which contained a Friedrich’s condenser, a sampling device, an oxygen dispersion tube and a glass impeller. The impeller was inserted through the center by means of a Chesapeake stirrer connection and was driven by a Fisher mixer. The entire reactor assembly was clamped into place in a water bath which was heated with a temperature controller and maintained at within 0.1 C of the desired temperature. In each test run, 100 g of ground anodic slime (0.075 mm) were treated with a 1-l solution containing a certain amount of sulphuric acid of known strength. The leaching tests were conducted at different temperatures for various time intervals with the admission of variable
2. Materials and apparatus Slime samples were collected from the bottom of the electrolytic cells during the electro-refining of copper. The slimes were sieved (2.0 mm) to remove the large inclusions of copper and other extraneous materials present in the slime. A representative sample was analysed for Cu, Pb, Sn, Zn, Ni, Fe, As, Sb, SiO2, MgO and Ag using an atomic
Fig. 1. Leaching apparatus.
A.M. Amer / Waste Management 23 (2003) 763–770
rates of oxygen as an oxidizing agent for copper dissolution. Samples of 5 ml of liquid were taken for subsequent analysis using atomic absorption for the determination of dissolved copper. At the conclusion of the leaching experiment, the slurry was filtered using a water suction pump through Table 1 Chemical analysis of representative sample of anodic slime studied Chemical constituents
Weight (%)
Cu Pb Sn Ag Ni Fe Sb As Zn S SiO2 CaO MgO Au Se Te
18.57 15.3 9.8 0.46 0.94 1.72 1.30 0.49 0.74 9.22 8.51 1.54 0.82 – – –
765
A-No-42 Whatman filter paper and washed thoroughly with hot distilled water, dried at 110 C, and then ground to 0.075 mm. The slime without the copper was analysed for Pb, Sn and Ag. Decopperised slime (40 g) was treated with a 1-l solution containing different amounts of acidified thiourea and ferric ion as oxidant. The slurry was agitated with a magnetic stirrer at various agitation speeds. Factors such as temperature, leaching time, thiourea concentration were taken into consideration to determine the optimum operating conditions for silver dissolution. A liquor sample was analysed using atomic absorption for determination of dissolved silver. The decopperised, silver-free residue was enriched in lead and tin, and was considered a base for the production of Pb–Sn soldering alloys. Decopperised, silver-free residue (50 g) was mixed with proper proportions of sodium carbonate and carbon, smelted in a graphite available at different temperatures and time for production Pb–Sn soldering alloy. The purpose of the investigation, as shown in Fig. 2, was to investigate and optimize the parameters influencing the dissolution of both copper and silver, as well as those affecting the production of soldering alloys.
Fig. 2. Proposed technique for processing of Egyptian Cu anodic slimes.
766
A.M. Amer / Waste Management 23 (2003) 763–770
where: A=constant, EA=activation energy in kJ/mole, R=gas constant, and T=temperature in k. The activation energy is determined from the Arrhenius plot given in Fig. 5 as follows:
4. Results and discussions 4.1. Extraction of copper 4.1.1. Effects of temperature Fig. 3 shows that the copper extraction increased with an increase in temperature in the range 65–85 C. Further increase in temperature had little effect on copper extraction. The following equation can be applied: 1 ð1 Þ1=3 ¼ kt
ð5Þ
where: =fraction reacted, k=a constant for the reaction rate, and t=duration of the reaction. Fig. 4 shows that the plot of 1(1a)1/3 versus t over the studied temperature range consists of a series of straight lines indicating that the sulphuric acid leaching of anodic slimes is llimited by the chemical concentrations. From the slopes of linearised isotherms in Fig. 4, certain isotherms were selected for rate constants at given temperatures. Using the Arrhenius equation: K ¼ A:eE=RT
ð6Þ
Slope ¼ EA =R
ð7Þ
The value calculated for the activation energy is 69 kJ/mole, which indicates surface chemical control. 4.1.2. Effect of acid concentration Fig. 6 shows the effect of charging the initial solution with acid. The rate of copper extraction increases when the acid concentration increases from 0.02 to 0.5 M. Further increase in acid concentration resulted in no significant increase in copper extraction. This may be due to the fact that the solubility of oxygen in high acid concentrations is low in comparison with its solubility in low acid concentrations. Fig. 7 illustrates the relation of pH value of the leaching solution and extraction, which indicates that the rate of metal uptake increases with increase of pH up to pH 4–5. Thereafter, it is more or less constant in the pH range investigated.
Fig. 3. Effect of temperature on copper extraction.
Fig. 5. Arrhenius plot for determination of activation energy.
Fig. 4. A plot of 1(1a)1/3 versus time for various temperatures.
Fig. 6. The effect of sulphuric acid concentration on copper extraction.
A.M. Amer / Waste Management 23 (2003) 763–770
4.1.3. Effect of oxygen partial pressure Fig. 8 shows the results of leaching 75 mm diameter anodic slime in 0.5 M H2SO4 under various oxygen partial pressures. It is concluded that, the higher the oxygen partial pressure, the higher the extraction of copper under the experimental conditions used. The maximum dissolution of anodic slime in sulphuric acid under oxygen pressure led to the extraction of copper (97%) under the following leaching conditions: Temperature: 85 C Sulphuric acid concentration: 0.5 M Oxygen partial pressure: 1 bar Leaching time: 90 min Particle size: 75 mm
767
4.2.1. Effect of thiourea concentration In an acidic solution of thiourea, the dissolution of silver proceeds according to the following reaction: Au CSðNH2 Þ2 þ þe Ag þ 2 CSðNH2 Þ2! ð8Þ 2 This is a reversible reaction with an electrode potential of 0.352 V at 25 C on a fresh silver surface. As shown in Fig. 9, the silver extraction increased with the increase of thiourea concentration and reached its maximum (90%) at 10 g/l thiourea concentration. An increase in thiourea concentration up to 10 g/l, produced an increase in the silver extraction, especially in the initial leaching period (0.5 h). There is, however, a decrease in silver extraction when thethiourea concentration exceeded 20 g/l, possibly because the thiourea is less stable at higher concentrations and decomposes.
4.2. Extraction of silver This part represents the experiments conducted for the extraction of silver from decopperized slime residue. The variables influencing the potential of using acidified thiourea solution as leaching agent are: thiourea concentration, temperature, leaching time, and ferric-ion concentration as oxidant. The decopperized slime residue is ground to 0.075 mm and then leached by acidified thiourea solution.
4.2.2. Effect of temperature The effect of temperature on extraction rate is studied, using 10 g/l thiourea and 5.0 g/l ferric ion. Extraction of more than 90% silver is reached at both 40 and 60 C after 30 min leaching time. The silver extraction is improved with the increase of temperature to 40 C where it reaches 93.5% after 4 h leaching time. 4.2.3. Effect of addition of ferric ion The oxidating effect of ferric ion in the presence of thiourea can be illustrated in the following reaction. þ Fe3þ ¼ CSðNH2Þ2 ð9Þ 2 ! Fe CSðNHÞ2 2 In contact with oxidants such as Fe+3, thiourea may be oxidized in successive stages to form several products. The first is the formation of formamidine disulphide (Schulze, 1984). ½ðNH2 ÞðNHÞCSSCðNHÞðNH2 Þ 2CSðNH2 Þ2! þ 2Hþ þ 2e; Eo ¼ 0:42 V:
ð10Þ
Fig. 7. Effect of pH on copper extraction.
In a second, irreversible reaction, the formamide disulphide yields thiourea and an unidentified sulphinic
Fig. 8. The effect of oxygen flow rate on copper extraction.
Fig. 9. Influence of thiourea concentration.
768
A.M. Amer / Waste Management 23 (2003) 763–770
compound, which then decompose to yield cyanomide and elemental sulphur. The elemental sulphur created at the end of this decomposition stage has a detrimental effect by passivating the leachable silver surface and precipitating some silver from solution. The study of the effect of oxidizing agent shows that there is an equilibrium between thiourea and formamidine disulfide. The leaching efficiency of silver depends on this equilibrium. The results indicate that the optimum extraction of silver was 93% using 10 g/l thiourea and 5.0 g/l Fe3+ in one leaching run of 4 h. As seen in Fig. 10, an addition of 2 g/l of Fe+3 slightly improves the initial leaching kinetics, but results in no real improvement in the rate of silver extraction. Further increase in Fe3+ to 5.0 g/l increases the leaching kinetics and also increases the extraction rate of silver to 93.4% in 7 h. The rate of silver dissolution is very rapid in the first 30 min of leaching. For the 5.0 g/l concentration of the oxidant, 80% of silver was leached in this period. Adding 8.0 g/l Fe3+ in the solution caused a significant decrease in silver extraction. When the concentration of oxidant is too large, oxidation of thiourea occurs and the remaining leachant is less effective in extracting silver.
Fig. 11 illustrates the relationship between the amount of carbon added and the metal content in leadtin soldering alloys. The figure shows that as the concentration of carbon increases, the lead content decreases and the tin content increases. The addition of > 20% carbon led to an increase of the tin content at the expense of a decrease in lead content. The results demonstrate that a concentration of 30% carbon is the optimum value. Fig. 12 represents the relationship between the percentage of sodium carbonate added and metal content (lead–tin) in soldering alloys. With the increase of sodium carbonate, a slight decrease of Pb content and slight increase of tin content was observed. Therefore, a concentration of 40% sodium carbonate is preferred. Fig. 13 reveals the relationship between temperature and percentage of metal content in soldering alloys. The figure shows that the content of lead in the alloy increases over the entire temperature range (900– 1300 C), while the tin content increases up to 1100 C and then decreases. Therefore, the optimum temperature is 1100 C.
4.3. Production of lead–tin soldering alloy After the removal of the copper and silver, the slimes were then enriched in lead (41.6%) and tin (29.0%). Residues were dried and ground to < 0.075 mm and then mixed with different proportions of sodium carbonate and carbon. The mixtures were smelted in graphite crucibles at different temperature levels and time intervals. The effects of the following parameters were studied: Carbon as reducing agent. Sodium carbonate as flux. Temperature. Reduction time.
Fig. 10. Influence of Fe3+ concentration on silver extraction.
Fig. 11. Effect of added carbon on lead–tin content.
Fig. 12. Effect of added sodium carbonate on lead–tin content.
A.M. Amer / Waste Management 23 (2003) 763–770
769
The effect of time on lead and tin content is shown in Fig. 14. The increase in reduction time from 1.0 to 3.0 h resulted in no significant increase in lead and tin content. Consequently, a reduction time of 1.0 h was chosen as a favourable value. Based on the experimental results, the more favourable operating conditions for production of soldering alloy can be summarized as follows:
However, a wide range of Pb–Sn soldering alloy is known depending on the Pb/Sn ratio. Therefore, by addition of the required quantity of tin before smelting, the tin content of the soldering alloy could be increased and adjusted with respect to the required lead content.
Carbon content: 30% (by weight of slime residue). Sodium carbonate content: 40% (by weight of slime residue). Temperature: 1100 C Reduction time: 1.0 h
About 90% of copper initially present in anodic slimes with 75 mm diameter particles dissolved in 0.25 M H2SO4 solutions at 85 C under 1 bar oxygen pressure within 90 min leaching time. The apparent activation energy is 69.0 kJ/mole characteristic of a controlled chemical reaction. Thiourea leaching of silver appears to be a good potential application to treat slimes containing silver to recover silver. The low toxicity of thiourea compared to cyanide is one of the motivations for developing the research in this field. Using concentration of thiourea of 10 g/l, under temperature 60 C in 30 min leaching time, 92.2% of silver is extracted. Smelting of the slimes after removal of copper and silver at 1100 C in the presence of sodium carbonate (40%) and carbon (30%) for 1 h led to production of Pb–Sn soldering alloy with the following characteristics: Pb: 53.0%–Sn: 33%.
Under the above operating conditions, the lead and tin in the soldering alloy had the following concentrations: Pb 53.0% Sn 33.0%
5. Conclusions
References
Fig. 13. Effect of temperature on lead–tin content.
Fig. 14. Effect of reduction time on lead–tin content.
Belen’kii, A.M., Petrov, G.V., Plekhanov, K.A., Kozlovskaya, A.E., Greiver, T.N., 1999. Hydrometallurgische Technologie zur verarbeitung von kupferelektrolyseschlammen. Cvetnye metally 74 (1), 43–45. Deschenes, G., Ghali, E., 1988. Leaching of gold from chalcopyrite concentrates by thiourea. Hydrometallurgy 20, 179–202. Everett, P.K. (1994) Development of copper process by an international consortium. Proceedings of Hydrometallurgy 94, England. Filipov, Ju.A., Anisimova, N.N.l., Kotuchova, A.K., Ter-Oganesjane, A.K., Tichov, I.V., 2000. Rekonstruktion der anodenschlammverabeitung. Cventnye metally 75 (6), 61–63. Gaylarde, C.C. and Videla, H.A. (1995) Bioextraction and biodeterioration of metals. Cambridge University Press, pp. 51–55. Holmes, J.A. (1981) Recent experience in the application of hydrometallurgical techniques. Proceedings of Hydrometallurgy 81, England. Gill, C.B., 1980. Non Ferrous Extractive Metallurgy. Wiley-Interscience, New York. Hughes, S., 2000. Applying ausmelt technology to recover Cu, Ni and Co from slags (overview). JOM 30–33. Kusnierov, M., Sepelk, V., Briancen, J., 1993. Effects of biodegradation and mechanical activation on gold recovery by thiourea leaching. JOM 54–56. Kucha, H., Cichowska, K., 2001. Precious metals in copper smelting products. Physicochemical Problems of Mineral Processing Journal 35, 91–101. Landsberg, R., 1977. Elektrochemische Reaktionen und Prozesse. VEB Deutscher Verlag der Wissenschaften. Mahdavi, M., Ditze, A., 1996. Basic investigations of the dissolution of copper and the formation of anodic slime during copper electrolysis with pure copper-oxygen anodes. Erzmetall 49 (6), 358–365. McClincy, R.J., 1990. Unlocking refractory gold ores and concentrates. JOM 10–11.
770
A.M. Amer / Waste Management 23 (2003) 763–770
Petrov, G.V., Plekhanov, K.A., Kozlovskoya, A.E., Greiver, T.N., 1999. Hydrometallurgische Technologie zur verarbeitung von kupferelektroly schlammen. Cventnye metally 74 (1), 43–45. Schulze, R.G., 1984. New aspects in thiourea leaching of precious metals. Int. Precious Metals Symposium, Los Angeles 27–29. Schulze, R.G., 1984. New aspects in thiourea leaching of precious metals. Int. Precious Metals Symposium, Los Angeles 27–29.
Subbaiah, T., Das, S.C., Das, R.P., 1986. Electrowinning of copper under forced convection. Erzmetall 34 (10), 501–506. Swayn, G.P., Robilliard, K.R., Floyd, J.M., 1993. Applying ausmelt processing to complex copper smelter dusts. JOM 35–38. Ying, C.H. (1983) Copper refinery anodic slime. Proceedings of the Third Int. Symp. of Hydrometallurgy. Annual Meeting, Atlanta, Georgia, March 6–10.