ublished in German in METALL (Berlin) 28 (11), 1051-54 (1974) THE RECOVERY OF COER, IRON, AND SULFUR FROM CHALCOYRITE CONCENTRATE BY REDUCTION Fathi Habashi Department of Mining & Metallurgy, Laval University Québec 10, Québec, Canada Raymond Dugdale Magma Copper Company, San Manuel, Arizona, USA Meguru Nagamori Centre de recherche industrielle Complexe scientifique, Sainte-Foy, Québec, Canada ABSTRACT Calcium oxide favors greatly the reduction of chalcopyrite by H 2 at 800ºC. The reduction takes place according to any of the following reactions depending on the amount of CaO added: Cu 2 S.Fe 2 S 3 + H 2 + CaO Cu 2 S.2FeS + CaS + H 2 O Cu 2 S.Fe 2 S 3 + 3 H 2 + 3 CaO Cu 2 S + 2 Fe + 3 CaS + 3 H 2 O Cu 2 S.Fe 2 S 3 + 4 H 2 + 4 CaO 2 Cu + 2 Fe + 4 CaS + 4 H 2 O Gravity and magnetic methods were not successful in separating the reaction products. However, leaching by dilute HCl was successful in eliminating CaS and iron (as FeS as well as metal) leaving behind Cu 2 S concentrate or metallic copper, respectively: CaS + 2 HCl CaCl 2 + H 2 S Cu 2 S.2FeS + 4 HCl Cu 2 S + 2 FeCl 2 + 2 H 2 S Fe + 2HCl FeCl 2 + H 2 Hydrogen sulfide generated during leaching can be oxidized by known methods to yield elemental sulfur: H 2 S + ½ O 2 S + H 2 O while H 2 can be recycled in the reduction step. Ferrous chloride can be crystallized from solution and reduced to metallic iron FeCl 2 + H 2 Fe + 2 HCl and the HCl generated can be recycled in the leaching step. INTRODUCTION The present treatment of copper sulfide concentrates is based exclusively on oxidation steps. First, a partial removal of sulfur by oxidation roasting, then formation of matte in the reverberatory furnace, then slagging and controlled oxidation of the white metal to blister copper in the converter. All these operations involve the formation of SO 2 which causes either pollution problems when emitted in the atmosphere, or economic problems if its recovery is considered. A different approach to 1
copper smelting would be the reduction of sulfides, e. g., by H 2 to metallic copper. The reduction of metal sulfides by hydrogen is thermodynamically unfavorable. For the reaction MS + H 2 M + H 2 S the equilibrium constant at 800 C is typically 2 x 10-3 to 6 x 10-3 for Cu, Ni, Co, and Fe sulfides. One way of increasing the yield of the metal is by shifting the equilibrium continuously to the right. This can be achieved by removing H 2 S as soon as it is formed. A convenient way of doing this would be to add to the sulfide a certain reagent that has a strong affinity for H 2 S. Calcium oxide fulfills such requirements. It reacts readily with H 2 S to form CaS according to the equation: CaO + H 2 S CaS + H 2 O The equilibrium constant for this reaction at 800 C is 1.7 x 10 3. It would therefore be expected that when CaO is added in stoichiometric amounts to the sulfide undergoing reduction by H 2, it would remove H 2 S continuously from the system, shifting the equilibrium favorably to the right. The overall reaction would therefore be MS + H 2 + CaO M + CaS + H 2 O Table 1 Chalcopyrite concentrate from Twin Buttes, Arizona % Cu 30.2 Fe 24.0 S 31.4 Zn 1.25 b 0.44 Mo 0.13 Insol 3.6 Gorin et al 1 have proposed the de-sulfurization of carbonaceous solid fuels by H 2 in presence of CaO. Barker 2 heated pyrite with lime in a reducing atmosphere to get a product, presumably metallic iron and calcium sulfide, suitable for precipitating copper from H 2 SO 4 leach solution. Kay 3 has proposed the de-sulfurization of molybdenite by reduction with H 2 in presence of CaO. Cech and Tiemann 4 were also interested in the reduction of sulfides by H 2 in the presence of CaO as a laboratory method for preparing metals in the form of filaments, since this is the form in which the metal is obtained by the reduction of the metal sulfides. They studied briefly the reduction of Cu 2 S, Ni 3 S 2, Co 9 S 8, and Fe sulfides by H 2 in the presence of CaO, and were able to confirm the thermodynamic predictions. REDUCTION OF CALCOYRITE In the present report the recovery of copper from its ores by reduction with H 2 in the presence of CaO as an additive is proposed. This proposal is based on detailed research data recently published by the authors 5. Chalcopyrite flotation concentrate from Twin Buttes, Arizona; -400 mesh, was used to demonstrate this possibility. The concentrate studied had the analysis given in Table 1. It was mixed with different amounts of CaO and heated at different temperatures in a hydrogen gas flow for one hour. The product was then leached at room temperature with 5% HCl, filtered, dried, weighed and analyzed. X-ray diffraction analysis, microscopic analysis of polished section of products, chemical analysis, and magnetic tests have shown that depending on the amount of CaO added, reduction at 800 C follows the following equations: Cu 2 S.Fe 2 S 3 + H 2 +CaO Cu 2 S.2FeS + CaS + H 2 O (1) Cu 2 S.FeS 3 + 3 H 2 +3 CaO Cu 2 S + 2 Fe + 3 CaS + 3 H 2 O (2) Cu 2 S.Fe 2 S 3 + 4 H 2 +4 CaO 2 Cu + 2 Fe + 4 CaS + 4 H 2 O (3) Magnetic methods failed to separate iron from the product because of the close association of the products. Gravity method was considered as a possible means for separating CaS from the copper-iron components. This was also not successful, apparently because of its fineness. Leaching by HCl was found to be the most effective method of separating the products. The product of reactions 1 and 2 was pure Cu 2 S, and that of reaction 3 was metallic copper. The leach solutions were free from copper. Leaching the product of reactions 2 and 3 takes place in two distinct steps: the first step is very rapid accompanied by H 2 S evolution. In the second step, when no H 2 S smell can be detected, leaching proceeds smoothly with the evolution of small bubbles of hydrogen. A portion of the product of reaction 3 was 2
leached then melted and metallic copper having the analysis shown in Table 2 was obtained. Table 2 Semi-quantitative spectrographic analysis of copper obtained from chalcopyrite by reduction % 0.1-1.0 Zn 0.05-0.5 As, Si 0.01-0.1 Mo, Sb 0.005-0.05 Ag, B, Bi, Fe, Sn, Mg, Mn 0.001-0.01 Nb, Ni, b < 0.001 Al, Ti Nil Au, t, Co, Se, Te, Ce reduction, E, can therefore be expressed in percentage as follows: Since E = K = H 2 O H 2 K Therefore E = 100 1 + K H 2 O H 2 O + H 2 x100 Calculations have shown that the reduction of chlacopyrite by H 2 is an endothermic reaction while the absorption of H 2 S by CaO is exothermic. The overall reduction of chalcopyrite in presence of CaO is slightly exothermic. Large batches containing 100g chalcopyrite were mixed with certain amounts of CaO corresponding to equations 1, 2, and 3. The mixture was pressed in a hydraulic press to pellets about ¼ in diameter. The charged in a vertical Vycor tube heated electrically at 800ºC as shown in Figure 1. Hydrogen at the rate of 1 Cu ft/hr was allowed to flow downwards and the water of reaction was collected in a graduated tube. In about 4 hrs the reaction was complete. The batch was discharged from the tube, weighed, leached with dilute HCl, filtered, washed, dried, weighed then analyzed. Results of these tests (Table 3) indicated that the amount of water collected and the loss in weight due to leaching were in good agreement with the calculated values according to equations (1), (2), and (3). Efficiency of the reaction An important criteria in evaluating the reduction process is the efficiency of hydrogen utilization, i. e., the amount of hydrogen that is utilized when the hot reducing gas is passed through a bed of ore heated at a certain temperature. If all the hydrogen is consumed in reduction, the exit gases will be hydrogen-free. This is an ideal case because under such conditions no hydrogen will be wasted, i. e., the efficiency is 100%. If, on the other hand, the reducing gas is not efficiently utilized, the exit gases will contain an appreciable amount of H 2 that warrants its recovery in some way. This is usually an undesirable operation. The hydrogen efficiency of Figure 1 Apparatus used for reduction of pellets of chalcopyrite concentrate To calculate the efficiency of reduction of chalcopyrite by H 2 in presence of CaO, one can approximate the calculations on Cu 2 S or FeS. The reason is that chalcopyrite concentrate liberates sulfur on heating to form a high-temperature phase of chalcopyrite which may be in coexistence with either iron sulfide or copper sulfide depending on the copper/iron ratio in the original concentrate. The efficiency of reduction is then based on the sulfide component that is most difficult to reduce. Table 4 shows the calculated efficiencies for the reduction of Cu 2 S and FeS according to the equations: Cu 2 S + CaO + H 2 2 Cu + CaS + H 2 O FeS + CaO + H 2 Fe + CaS + H 2 O 3
Table 4 Calculated values of hydrogen utilization efficiency for reduction of Cu 2 S and FeS in presence of CaO Sulfide Effiency at 500ºC 600ºC 700ºC 800ºC 900ºC Cu 2 S 81 72 64 56 50 FeS 76 74 72 70 69 It can be seen that the efficiency decreases with increasing temperature since the process is exothermic. However, the rate of reduction increases with increasing temperature. Therefore, there must be a compromise between the two factors and a temperature should be selected at which the rate of reduction is reasonably fast and at the same time the efficiency of hydrogen utilization is not below acceptable limits. At 800ºC the rate was reasonably high and a calculated efficiency of about 56% is acceptable, i. e., about 56% of the hydrogen introduced will be utilized in reducing the sulfide and only about 44% has to be recovered. catalytically oxidized to elemental sulfur by Claus process: H 2 S + ½ O 2 S + H 2 O The leach solution obtained, which is mainly ferrous chloride can be processed by known methods, e. g., evaporated to the proper concentration to crystallize FeCl 2.2H 2 0 and this is reduced by H 2 to metallic iron and regenerating HCl for leaching: FeCl 2 + H 2 Fe + 2 HCl ROOSED ROCESSES On the basis of these data, the reduction of copper sulfide ores, concentrates, or precipitates from leach solution, by hydrogen in presence of lime, seems to be technically attractive. The sulfide raw material is mixed with the proper amount of lime and heated in a rotary kiln or a fluidized bed reactor in a reducing atmosphere, e.g., hydrogen at about 800ºC. The only gaseous reaction product would be water vapor (or CO 2 if CO was used as a reducing gas). Three possible processes are available for the treatment of chalcopyrite concentrates. 1. Low-Lime rocess In this process (Fig. 2), the concentrate is mixed with about 0.25 ton CaO per ton chalcopyrite so that reaction according to equation 1 takes place. When the product is leached with dilute HCl, CaS and ferrous sulfide will go into solution liberating H 2 S as follows: CaS + 2HCl CaCl 2 + H 2 S Cu 2 S.2FeS + 4 HCl Cu 2 S + 2FeCl 2 + 2H 2 S while cuprous sulfide will not be attacked. The product obtained analyzes typically 57-65% Cu, 1-8% Fe, 16-20%, and the balance gangue. Such material can be melted and charged in a converter then blown directly by conventional methods to metallic copper. Hydrogen sulfide liberated during the leaching step can be Figure 2 roposed process for reducing chalcopyrite concentrate with small amount of CaO 2. Intermediate Lime rocess In this process (Fig. 3), chalcopyrite concentrate is mixed with about 0.75 ton CaO per ton chalcopyrite so that reaction according to equation 2 takes place. The product which is mixture of Cu 2 S, Fe and CaS is leached with dilute HCl to dissolve firstly CaS liberating H 2 S (first stage leaching) then metallic iron (second stage leaching) as follows: Fe + 2 HCl FeCl 2 + H 2 Hydrogen evolved during this step would be used for the reduction process. The product obtained after leaching has nearly the same analysis as that obtained in the low lime process. 4
3. High Lime rocess In this process (Fig. 4), chalcopyrite concentrate is mixed with about 1 ton CaO per ton chalcopyrite so that reaction according to equation 3 takes place. The product which is a mixture of Cu, Fe, and CaS is leached with dilute HCl to dissolve firstly CaS liberating H 2 S, then metallic iron as described earlier. The residue from this operation is rich in metallic copper and has the following typical analysis: 70-77% Cu, 0.5 3% Fe, 0.1 4% S, and the balance gangue. Such a product can be melted and slagged, then cast for electrolytic refining. Eventually, the copper-iron mixture obtained after removing CaS, can be used to replace sponge iron or iron scrap in cementation plants. The High Lime rocess is an original method by which metallic copper can be obtained directly from a chalcopyrite concentrate. Figure 4 roduction of copper from chalcopyrite concentrates by reduction REFERENCES Figure 3 Reduction of chalcopyrite concentrate with medium amount of CaO 1 E. Gorin, G.. Curran, and J. D. Batchelor, Desulfurization of Solid Fuels, U. S. atent 2 824 047 (1958). CA 52, p. 8512b 2 L. M. Barker, Recovery of Mineral Values from Ore, U. S. atent 3 168 396 (1965) 3 H. Kay, Hydrogen Reduction of Molybdenite using Sulfur Acceptors, pp. 33-44 in High Temperature Refractory Metals, AIME Met. Soc. Conference, February 1965, Ed. W. A. Krisvsky, ublished by Gordon & Breach, New York, 1968. 4 R. E. Cech and T. D. Tiemann, The Hydrogen Reduction of Cu, Ni, Co, and Fe Sulfides and the Formation of Filamentary Metals, Trans. Met. Soc. AIME 245, 1727-33 (1969) 5 F. Habashi and R. Dugdale, The Reduction of Sulfide Minerals by Hydrogen in resence of Lime, Met. Trans. 4, 1865-71 (1973). 5
Table 3 Reduction of large batches of chalcopyrite CaO mixture by H 2 at 800ºC. The mixture was palletized to about ¼ in diam. pellets and reduced in a static bed by a downward gas flow at 1 Cuft/hr. Reaction Studied Cu 2 S.Fe 2 S 3 + CaO + H 2 Cu 2 S.2 Fe + CaS + H 2 O Cu 2 S.Fe 2 S 3 + 3 CaO + 3 H 2 Cu 2 S + 2 Fe + 3 CaS + 3 H 2 O Cu 2 S.Fe 2 S 3 + 4 CaO + 4 H 2 2 Cu + 2 Fe + 4 CaS + 4 H 2 O Feed Water collected Weight Loss in weight Analysis of leach residue Weight of of CaO product Expected 1 Found Expected residue 2 Found Cu Fe S Gangue 3 g g g ml ml g % % % % % % Concentrate 100 14.5 101.3 6.6 7.5 49.5 58.2 57.0 57.2 8.0 19.9 14.9 100 42.7 124.4 16.8 15.3 38.3 66.5 74.0 65.9 0.17 14.6 19.3 100 58.0 138.2 21.7 18.0 34.5 75.0 78.5 76.7 0.5 0.1 22.7 1 Moisture in feed material were as follows : chalcopyrite concentrate 1.5%, lime 5% 2 Based on feed to furnace 3 By difference 1