Copper chloride leaching from chalcopyrite and bornite concentrates containing high levels of impurities and minor elements

Hydrometallurgy 138 (2013) 40–47

Contents lists available at SciVerse ScienceDirect

Hydrometallurgy journal homepage: www.elsevier.com/locate/hydromet

Copper chloride leaching from chalcopyrite and bornite concentrates containing high levels of impurities and minor elements Jianming Lu ⁎, David Dreisinger Department of Materials Engineering, University of British Columbia, 309–6350 Stores Road, Vancouver, BC, Canada V6T 1Z4

a r t i c l e

i n f o

Article history: Received 1 November 2012 Received in revised form 1 March 2013 Accepted 1 June 2013 Available online 7 June 2013 Keywords: Copper Chalcopyrite Bornite Chloride leaching

a b s t r a c t This study was conducted as part of the development of a novel process for copper recovery from copper sulfide concentrates by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper chloride leaching from chalcopyrite and bornite concentrates containing high levels of impurities and minor elements has been studied using a two-stage countercurrent leach circuit at 95 °C to maintain a high copper extraction and a low cupric concentration in the pregnant leach solution for subsequent copper solvent extraction. A high calcium chloride concentration (165 g/L or higher) was used to maintain a high cuprous solubility and enhance copper leaching. With 3 h of leaching time for each stage, at a particle size of P80 of 14 μm, the copper extractions were 99.1 and 98.5 % for chalcopyrite and bornite concentrates respectively while the iron extractions were 73.3 and 99% respectively. The extractions of copper and iron from chalcopyrite concentrate decreased very slightly with increasing particle size to P80 of 30 μm while the extractions of copper and iron from bornite concentrate decreased by ~3.6% with increasing particle size to P80 of 40 μm. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was 0.1 to 0.2 M, indicating the pregnant leach solution was suitable for subsequent copper solvent extraction. Sulfur oxidation to sulfate was 1.2–1.7% for chalcopyrite concentrate, and 2.1% for bornite concentrate. The extractions of Bi, Cr, Pb, Ag and Zn were very high. The extractions of Sb, As and Ni were high for chalcopyrite concentrate while they were low for bornite concentrate. © 2013 Elsevier B.V. All rights reserved.

1. Introduction Ferric and/or cupric chloride leaching of copper from sulfide concentrates has some advantages over sulfate systems in supporting high metal solubility, enhanced redox behaviour, and increased rates of leaching (Dutrizac, 1990, 1992; O’Malley and Liddell, 1987; Peters, 1977). Both cuprous and cupric ions are stabilized through complexation with chloride ions. Another advantage is the minimization of sulfate formation and the formation of predominately elemental sulfur. Pyrite is generally not attacked. A series of two-stage counter-current chloride leaching tests have been conducted for the Falconbridge Copper Chloride Leach Process. The copper extraction from chalcopyrite concentrates can exceed 99% (Lu and Dreisinger, 2013). However, limited studies have been performed on the behaviour of deleterious impurities in the leach system. The behaviour of Bi, Sb, As, Hg, Zn, and Cr require further investigation of leach chemistry. A series of copper concentrates with high levels of impurities should be tested for amenability to the Falconbridge Copper Chloride Leach Process.

⁎ Corresponding author. Tel.: +1 604 822 1357; fax: +1 604 822 3619. E-mail address: [email protected] (J. Lu). 0304-386X/$ – see front matter © 2013 Elsevier B.V. All rights reserved. http://dx.doi.org/10.1016/j.hydromet.2013.06.001

The following reactions are generally accepted as representing the leaching of chalcopyrite in the solution of FeCl3 and CuCl2: 3þ

¼ Cu

2þ

¼ 4Cu þ Fe

CuFeS2 þ 4Fe

CuFeS2 þ 3Cu

2þ

2þ

þ 2S

2þ

þ 2S

þ 5Fe

þ

0

ð1Þ

0

ð2Þ

The ferric/cupric chloride leaching of bornite is a complicated process (Dutrizac et al., 1985; Pesic and Olson, 1983). The oxidation of bornite by ferric/cupric ions may be divided into two stages. At the first stage, bornite (Cu5FeS4) is oxidized to Cu3FeS4. At the second stage, Cu3FeS4 is further oxidized to Cu+, Fe2+ and S. The overall reaction may be written as: 3þ

Cu5 FeS4 þ 12Fe

2þ

Cu5 FeS4 þ 7Cu

2þ

¼ 5Cu

2þ

þ 13Fe

þ

¼ 12Cu þ Fe

2þ

0

þ 4S

ð3Þ

0

ð4Þ

þ 4S

Less than 5% of sulfide (Dutrizac, 1990, 1992) may be oxidized to sulfate as a side reaction, which may be expressed as: 0

3þ

S þ 6Fe

2þ

þ 4H2 O ¼ 6Fe

2

þ

þ SO4 þ 8H

ð5Þ

J. Lu, D. Dreisinger / Hydrometallurgy 138 (2013) 40–47 0

2þ

S þ 6Cu

þ

2

þ

þ 4H2 O ¼ 6Cu þ SO4 þ 8H

Leach solution

ð6Þ Chalcopyrite concentrate

The reduction of cupric ion by chalcopyrite or bornite is also limited by the following thermodynamic equilibrium in the presence of elemental sulfur (McDonald and Langer, 1983; McDonald et al., 1984): 2Cuþ þ S0 ¼ CuS þ Cu2þ

k ¼ 3:5  1011

41


105• C

Pregnant leachate

ð7Þ Slurry Leach Stage 1

The extent of the above reaction is dependent on temperature and Cu+ /Cu2+ ratio. The presence of chloride can significantly reduce free Cu+ concentration through complexation. A higher chloride concentration results in a lower free Cu+ concentration, a higher Cu2+/Cu+ redox potential and therefore suppression of Reaction 7. The following reactions may also occur (Peters, 1977): 2þ

Cu

þ

2þ

þ 2CuS

ð8Þ

2þ

þ CuS þ Cu2 S

ð9Þ

þ CuFeS2 ⇔Fe

2Cu þ CuFeS2 ⇔Fe

Fresh feed solution

As discussed by Lu and Dreisinger (2013), when the concentration of Cu(II) + Fe(III) is below a limit, only iron is leached and copper gradually precipitates due to Reactions 7 to 9. To suppress the above side reactions, a low temperature should be used. However, a lower temperature also results in a slower leaching rate. With respect to copper recovery and iron deportment, there are four typical processes: Outokumpu HydroCopper, Intec, Sumitomo and Falconbridge copper chloride. In the HydroCopper process, copper concentrates are leached in strong sodium chloride solution using Cu(II) as an oxidant (Hyvarinen and Hamalainen, 2005). The iron leached is oxidized to Fe(III) by oxygen and re-precipitated as hematite. The pregnant leach solution is highly purified using multi-steps. In the first step, the impurities (Fe, Ni, Zn, Co, Mg and Cu(II)) are removed by raising the pH to 4–5 using CaCO3 and NaOH. In the second step, silver is removed as amalgam by cementation on mercury coated copper particles. In the third step, the majority of the remaining impurities (Zn, Pb, Ni and etc.) are removed by raising pH to 6–7 using Na2CO3. In the last step, the impurities are removed to trace levels using a chelating resin. The above multi-step purification contributes greatly to the process operating cost. Cu2O is produced from the purified solution by the addition of NaOH and then reduced to metallic copper by H2. NaOH, H2 and Cl2 are generated by chlor-alkali electrolysis. Cl2 is used to oxidize Cu(I) and regenerate Cu(II). In the Intec copper process, copper concentrates are leached in a NaCl-NaBr solution at 80–85 °C during four-stage countercurrent leaching using Cu2+, O2 and BrCl-2 as oxidants (Moyes et al., 2000). Iron is leached and re-precipitated as goethite during the first three stage leaching. Copper is recovered by electrowinning from copper chloride/bromide solution in a diaphragm cell. The anode product is NaBrCl2. Silver and mercury are removed as an amalgam by cementation with a copper-mercury liquid alloy. The other impurities are removed by increasing pH. The operation cost is relatively low. However, the copper extraction can be as low as 94%. Sulfur and goethite are rejected to the leach residue, causing a potential environmental issue. Even with a step for the removal of silver and mercury, the copper deposit may still be contaminated by silver and mercury. In the Sumitomo copper process, copper concentrates are leached in chloride solution using Cl2 as an oxidant (Asano et al., 2007). CuCl is extracted into an organic phase from the copper pregnant solution. CuCl is stripped from CuCl-loaded organic phase to CuCl aqueous solution. Copper metal is produced in a diaphragm cell while Cl2 is produced at the anode for leaching. The Fe-containing raffinate is electrolyzed to produce metallic iron that is saleable as scrap metal. The cost for production of metallic iron is high. In the Falconbridge copper chloride process, with respect to iron deportment (as hematite or goethite), two processes were proposed

Solids Thickener

Pressure Filtration

Slurry Leach Stage 2

Residue Fig. 1. Schematic diagram for two-stage countercurrent leach circuit.

featuring chloride leaching of chalcopyrite combined with solvent extraction and conventional copper electrowinning from sulfate media to recover a pure copper product (Liddicoat and Dreisinger, 2007). The hematite process, in which no oxygen is introduced to the leach, allows separate sulfur and iron residues to be produced, thus making waste treatment and disposal easier. The hematite powder can be a saleable product such as pigment. In the hematite process, the separate solvent extraction and hematite precipitation are expressed as the following reactions: 4CuCl þ 4HRðorgÞ þ O2 ¼ 2CuR2 ðorgÞ þ 2CuCl2 þ 2H2 O 2þ

6Fe

ð10Þ

3þ

þ 1:5O2 ¼ Fe2 O3 þ 4Fe

ð11Þ

Copper is stripped from the copper-loaded organic solution to copper sulfate solution for conventional copper electrowinning with minimum impurity transfer to recover a pure copper product. The objective of this study was to investigate the ferric-cupric chloride leaching of chalcopyrite and bornite concentrates containing high levels of impurities and minor elements under well-controlled conditions to produce a pregnant leach solution with a minimum concentration of Cu(II) + Fe(III) for subsequent cuprous oxidation and cupric solvent extraction.

2. Experimental 2.1. Apparatus The schematic diagram of the two-stage countercurrent leach circuit is shown in Fig. 1. One mini-thickener and pressure filter were used for solid–liquid separation of slurries for Stages 1 and 2 respectively. Chloride leaching was conducted in two 2-L glass reactors with three built-in baffles. The temperature was controlled at 95 ± 1 °C using two heating mantles with temperature controllers. The reactors were sealed with removable lids with several openings for various purposes. Agitation was provided by single 45° pitched impellers with a diameter of 5.7 cm. The rotational speed was controlled at 850 rpm. To protect cuprous and ferrous oxidation, all operations were conducted under a nitrogen atmosphere.

Table 1 Assays of chalcopyrite and bornite concentrates. Concentrate

Particle size (P80) / μm

Cu /%

Fe /%

STotal /%

S(SO2− 4 ) /%

S0 /%

S2− /%

Chalcopyrite Chalcopyrite Bornite Bornite

14 30 14 40

22.4 22.2 40.3 40.1

32.5 32.3 18.8 18.9

34.9 36.0 27.4 28.3

0.61 1.96 0.43 1.12

0.57 0.56 0.06 0.05

33.7 33.5 26.9 27.1

42

J. Lu, D. Dreisinger / Hydrometallurgy 138 (2013) 40–47

800

Ideal Formula Chalcopyrite Sphalerite Bornite Tennantite Pyrite Pyrrhotite Anglesite Galena Gypsum Andradite Wroewolfeite Molybdenite Quartz Total

Chalcopyrite concentrate

CuFeS2 (Zn,Fe)S Cu5FeS4 (Cu,Ag,Fe,Zn)12As4S13 FeS2 Fe1-xS PbSO4 PbS CaSO4 · 2H2O Ca3Fe3+ 2 (SiO4)3 Cu2+ 4 (SO4)(OH)6 · 2H2O MoS2 SiO2

Bornite concentrate

65.0 4.0

37.3 11.6 37.7 5.9

19.3 4.2 4.4 2.1 1.6 4.2 0.3 0.5 0.9 100.0

1.0 100.0

Potential vs. Ag/AgCl / mV

Table 2 Results of quantitative phase analysis (%).

Stage 1

700

Stage 2

600

500

400

Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6 300 0

3

6

9

12

15

18

Time / h Fig. 3. Redox potential vs. time (Test 3: P80 of 14 μm and bornite concentrate).

2.2. Experimental procedures The continuous countercurrent leach tests began with a required amount of feed concentrate and 1 L of feed solution in each reactor. At the end of the designated residence time countercurrent transfer would occur with new feed chalcopyrite concentrate added to the reactor for Stage 1 and fresh feed solution added to the reactor for Stage 2. This leach/transfer was repeated for 6 cycles to reach a steady state. The procedures are: (1) 1 L of feed solution was added into each reactor. Once the temperature had reached a target value, the redox potential was recorded, and then the concentrate was added into the reactors. 15 minutes prior to transferring Stage 1 slurry to the thickener, 10 mL of 1000 ppm Superfloc® A110 anionic flocculant was added to enable quick and complete settling of solids. (2) At the end of the leach residence time, Stage 1 slurry was pumped to the thickener while Stage 2 slurry was pumped to the pressure filter. The clarified pregnant leach solution was first removed, and then the underflow (solids) was pressure-transferred to the reactor for Stage 2. After pressure filtration of Stage 2 slurries, the filtrate was sampled and pressure-transferred to the reactor for Stage 1. When the leach solution temperature reached the target value, fresh chalcopyrite was added to the reactor for Stage 1. 2.3. pH and redox potential measurements Due to problems of quick probe fouling in hot concentrated chloride solutions, the pH was measured at room temperature. An epoxy body

platinum-Ag/AgCl combination electrode (VWR Symphony brand) with a free flow junction was used to measure redox potential. 2.4. Preparation of feed concentrate and feed solution The chalcopyrite concentrate tested (BMS) was provided from Brunswick Mines, Canada while the bornite concentrate was obtained from Antamina, Peru. The concentrates were ground to P80 of 30/40 to 14 μm using a rod mill. The assays of the concentrates are given in Table 1. During grinding, metal sulfates were dissolved into water. Therefore the contents of sulfur as sulfate in finer concentrates were lower than those in coarse ones. Feed solutions were prepared using deionized water, technical grade ferric chloride, reagent grade calcium chloride dihydrate, reagent grade cupric chloride dihydrate and reagent grade hydrochloric acid. 2.5. Chemical analysis The slurry samples were taken and cooled down under the protection of argon gas. After the solids settled and the solution samples were taken, the solids were filtered and washed with 5% HCl solution until the filtrate became colorless. Finally the solid was further washed with de-ionized water. The solid was dried in an oven at 40 °C for ICP and sulfur group analyses. The free acid titration was conducted with oxalate masking under an inert environment. 2 M potassium oxalate was used to complex metal 0

Stage 1

700

Stage 2

-0.2

Stage 1 Stage 2

-0.4 600

pH

Potential vs. Ag/AgCl / mV

800

-0.6

500 -0.8 400

-1

Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6

Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6 -1.2

300 0

3

6

9

12

15

18

Time / h Fig. 2. Redox potential vs. time (Test 2: P80 of 30 μm and chalcopyrite concentrate).

0

3

6

9

12

15

18

Time / h Fig. 4. pH of leach solution of each cycle (Test 2: P80 of 30 μm and chalcopyrite concentrate).

J. Lu, D. Dreisinger / Hydrometallurgy 138 (2013) 40–47

0

43

10

pH

-0.2

HCl concentration / g L-1

9

Stage 1 Stage 2

-0.4

-0.6

8 7 6 5 4

Stage 1 Stage 2 Average calculated

3 2 1

Cycle 1 Cycle 2 Cycle 3 Cycle 1 Cycle 5 Cycle 6

Cycle 1 Cycle 2 Cycle 3Cycle 4 Cycle 5 Cycle 6

0

-0.8 0

3

6

9

12

15

18

0

3

6

Time / h

9

12

15

18

Time / h

Fig. 5. pH of leach solution of each cycle (Test 3: P80 of 14 μm and bornite concentrate).

Fig. 6. HCl concentration for each cycle (Test 2: P80 of 30 μm and chalcopyrite concentrate).

ions that would normally contribute protons, via reaction with hydroxide ions in water. The titration end point was determined using standard solutions with different concentrations of iron and copper. The titration of reduced species was conducted using certified acidic cerium (Ce4+) sulfate under an inert environment.

an autoclave (Reaction 11). At 165 g/L CaCl2 and 120 g/L Fe as FeCl2, cuprous concentration can reach 80 g/L. However, the zinc and lead dissolved compete for chloride to form complexes. Cuprous concentration was set at 72 g/L. The calcium chloride concentration for Tests 3 and 4 was set at 200 g/L to maintain cuprous chloride solubility above 72 g/L since ferrous concentration was around 60 g/L (ferrous chloride contributes chloride ions for complexation of other metals). A retenion time of 3 h was sufficiently long for each stage leaching. Copper extraction was not improved further much in a longer retention time. A shorter retention time resulted in a lower copper extraction. Therefore a retention time of 3 h was used in this study (Table 3). The results for Tests 1 to 4 were consistently good. Only the results for Tests 2 and 3 are discussed here.

3.1. Copper concentrate mineralogy and experimental conditions The mineralogical information of two copper concentrates is summarized in Table 2. The chalcopyrite concentrate mainly consisted of 65% of chalcopyrite, 4.0% of sphalerite, 19.3% of pyrite, 4.2% of pyrrhotite, 4.4% of anglesite, 2.1% of galena while the bornite concentrate contained 37.3% of chalcopyrite, 11.6% of sphalerite, 37.7% of bornite and 5.9% of tenantite. The ferric and cupric concentrations in the feed solution were determined according to the concentrate mineralogy, target cuprous concentration and the excess amount of Cu(II). The target cuprous concentration was set at 72 g/L assuming all the iron in the pregnant leachate was presented as ferrous. All the copper and iron except that in pyrite were extracted into solution. Since the excess amount of Cu(II) has to be reduced to Cu(I) using metallic copper for solvent extraction and purification stages, and the reduction of cupric to cuprous is costly (significant recycle of metallic copper to reduction step), the concentration of excess Cu(II) in the final leach solution should be as low as possible. However, if the concentration of excess Cu(II) is too low, copper can be precipitated according to Reactions 7 to 9, resulting in a low copper extraction. According to the previous study (Lu and Dreisinger, 2013), 8 g/L of excess Cu(II) was set for Tests 1 to 3, and 12 g/L of excess Cu(II) for Test 4. Additional ferric chloride was considered to oxidize minor element sulfides (ZnS and PbS). The amount of the copper extracted during leaching should be equal to that removed during solvent extraction (Reaction 10) while the amount of the iron extracted should be the same as that precipitated in

Table 3 Test conditions for chloride leaching (95 °C and 3 h of retention time for each stage). Test No. Cu concentrate Particle size Pulp density Cu Fe CaCl2 HCl (μm (P80)) (g/L) (g/L) (g/L) (g/L) (g/L) 1 2 3 4

Chalcopyrite Chalcopyrite Bornite Bornite

14 30 14 40

160.7 160.7 89.3 89.3

44 44 44 48

84.1 84.1 45.4 45.4

165 165 200 200

3 3 3 3

3.2. Potential, free acid and pH The slurry potential vs. time for Tests 2 and 3 is shown in Figs. 2 and 3 respectively. The potential decreased sharply in Cycle 1 since the reduction of ferric and cupric ions by chalcopyrite was rapid at the beginning of the test. The Cu2+/Cu+ and Fe3+/Fe2+ ratios decreased quickly, leading to a rapid decrease in the solution potential according to the Nernst equation. For Cycles 2 to 6, the ratios of Cu2+/Cu+ and Fe3+/Fe2+ did not decrease as greatly as those in Cycle 1. The residue, together with a small amount of leachate, was transferred from Stage 1 to Stage 2 after solid/liquid separation in the thickener. A certain amount of cuprous and ferrous ions were introduced into Stage 2 reactor. Therefore the 9 8

HCl concentration / g L-1

3. Results and discussion

7 6 5

Stage 1 Stage 2 Average calculated

4 3 2 1

Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6

0 0

3

6

9

12

15

18

Time / h Fig. 7. HCl concentration for each cycle (Test 3: P80 of 14 μm and bornite concentrate).

44

J. Lu, D. Dreisinger / Hydrometallurgy 138 (2013) 40–47

100

Table 4 Oxidation of sulfur to sulfate at different conditions. Test No.

1

2

3

4

Concentrate Type Particle size (P80) / μm HCl in feed solution / g L−1 Oxidation of S to SO24 / % HCl in pregnant leachate / g L−1

Chalcopyrite 14 3.3 1.7 11.8

Chalcopyrite 30 3.3 1.2 8.7

Borinite 14 3.3 2.1 7.7

Bornite 40 3.3 2.1 7.2

þ

þ

2þ

þ 1=2H2 O

ð12Þ

þ

3þ

þ 1=2H2 O

ð13Þ

þ 1=4O2 þ H ¼ Fe

The oxidation of cuprous or ferrous ions occurred during leaching and solid/liquid separation since the equipment was not perfectly sealed against oxygen ingress. The oxidation of cuprous and ferrous ions during the free acid titration probably contributed greatly to a lower analyzed free acid concentration if there was not perfectly sealed.

Extraction / %

74.1 67.8

60

37.3

40

14.6 Fd 2 11.8 Fd 3

Fd 1

0

Fd 4

Fd 5

Cycle1 Cycle 2 Cycle 3 Cycle 4 0

3

6

37.0

26.5 25.9

9

Fd 6

Cycle 5 Cycle 6

12

15

18

Time / h Fig. 9. Copper and iron extractions of 6 chalcopyrite concentration feeds for Test 3 (P80 of 14 μm and bornite concentrate). Solid markers for Cu and open markers for Fe.

The average oxidation of S to SO24 and HCl concentration in pregnant leachate in Cycle 6 are summarized in Table 4. The sulfur oxidation to sulfate was 1.2 to 1.7% for chalcopyrite concentrate while 2.1% for bornite concentrate. Finer grinding of chalcopyrite resulted in oxidation of slightly more S to sulfate, leading to a higher acid concentration. Finer grinding of bornite concentrate practically had no effect on sulfur oxidation. 3.3. Copper and iron extractions The extractions of copper and iron vs. time for Tests 2 and 3 are shown in Figs. 8 and 9 respectively. The final (Stage 2) extractions of copper and iron fluctuated slightly while Stage 1 extractions fluctuated significantly. The extractions of copper and iron at Stage 1 were related to the concentrations of ferric and cupric ions in the leach solution from Stage 2 in the previous cycle. In Cycle 1, the concentrate (Feed 1) was contacted with fresh feed solution and therefore Reactions 1 and 2 were fast, resulting in higher extractions of copper and iron in Stage 1. At Stage 2 (Cycle 2), the residue from Stage 1 (Cycle 1) was contacted with fresh feed solution. Since the copper content of the residue was low, only a small amount of cupric and ferric ions were consumed and so the concentrations of cupric and ferric ions were maintained at a high level during Cycle 2. This resulted in higher final extractions of copper and iron for Feed 1 at Stage 2 (Cycle 2). 140

100

99.4

97.0

92.2

98.9 97.8

98.7 120

87.1

80

73.4 Extraction / %

99.0 98.2 95.9 98.5

98.7 97.9

77.2 68.2

20

71.6

70.2

73.2

79.773.1 74.8

62.9

60

57.2

52.0

45.6 40

41.1 36.5 19.6 18.8

20

Fd 1 0

Cycle 1 Cycle 2 0

3

Fd 4

Fd 5

Cycle 3 Cycle 4 Cycle 5 6

100 80 60 40

Stage 1 Copper Stage 2 Copper

20

Fd 3

Fd 2

Concentration / g L-1

2þ

Fe

97.9 95.7

80

initial potential at Stage 2 was much lower than that of fresh feed solution. The initial potential at Stage 1 was nearly the same as the final potential of the previous cycle at Stage 2 since the leachate from Stage 2 was transferred to Stage 1. The pH values of feed and pregnant leach solution are given in Figs. 4 and 5 for Tests 2 and 3 respectively. In Test 2, the pH decreased from − 0.50 (for feed) to around − 1.0 (for leachate) while in Test 3, it decreased from − 0.45 to around − 0.60. The decrease in the pH indicates that the proton (H+) concentration increased or the composition of the other species such as free chloride, iron and copper chloride concentrations changed since these species affect the activity coefficient of H+. The free acid (HCl) concentrations for feed solution and pregnant leach solution for each cycle are shown in Figs. 6 and 7 for Tests 2 and 3 respectively. In Test 2, the free acid concentration increased from 3.3 to 7.2 g/L after one cycle while in Test 3, it increased from 3.3 to 5.2 g/L. The acid was generated due to Reactions 5 and 6. The amount of sulfate generated during leaching was calculated based on the analysis of sulfate in feed solution, chalcopyrite concentrate, residue, leach solution and wash solution. In the presence of 165–200 g/L CaCl2, nearly all the sulfate was precipitated as gypsum. The concentration of sulfate was below 0.1 g/L. Based on the amount of sulfate generated in leaching, the average HCl concentration was calculated according to Reactions 5 and 6 and is shown in Figs. 6 and 7. The average HCl concentration calculated was higher than those analyzed due to the following reactions: Cu þ 1=4O2 þ H ¼ Cu

99.2 98.6

93.3 93.2

9

12

Fd 6

Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6

Cycle 6 15

Stage 1 Iron Stage 2 Iron

0 18

Time / h Fig. 8. Copper and iron extractions of 6 chalcopyrite concentration feeds for Test 2 (P80 of 30 μm and chalcopyrite concentrate). Solid markers for Cu and open markers for Fe.

0

3

6

9

12

15

18

Time / h Fig. 10. Concentrations of copper and iron in the leach solution for each cycle for Test 2 (P80 of 30 μm and chalcopyrite concentrate).

J. Lu, D. Dreisinger / Hydrometallurgy 138 (2013) 40–47

Table 6 Mass of the residues and their contents of copper, iron and sulfur species from Stage 2 in Cycle 6.

60

40

Stage 1 Copper

Stage 1 Iron

Stage 2 Copper

Stage 2 Iron

20

Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6 0 0

3

6

9

12

15

18

Time / h Fig. 11. Concentrations of copper and iron in the leach solution for each cycle for Test 3 (P80 of 14 μm and bornite concentrate).

At Stage 1 (Cycle 2), the fresh chalcopyrite concentrate (Feed 2) was contacted with the leach solution (filtrate) from Stage 2 (Cycle 1). This Stage 2 leach solution composition was nearly the same as that from Stage 1 (Cycle 1), in which the ferric and cupric concentrations were low. Therefore the extractions of copper and iron were low due to the slow reaction. The copper extraction was lower than that for iron due to the side reactions (Reaction 7, Reactions 8 and 9 for chalcopyrite and the reactions similar to Reactions 8 and 9 for bornite). At Stage 2 (Cycle 3), the residue from Stage 1 (Cycle 2) was contacted with fresh feed solution and the final copper extraction for Feed 2 was slightly lower than that for Feed 1. Since the residue contained a larger amount of chalcopyrite and most of cupric and ferric ions were consumed in this stage, ferric and cupric concentrations were not maintained at a high level, resulting in a slower reaction and then a lower extraction of copper. At Stage 1 (Cycle 3), the fresh chalcopyrite concentrate (Feed 3) was contacted with the leach solution from Stage 2 (Cycle 2). Since the ferric and cupric concentrations were high in this leach solution, the extractions of copper and iron at Stage 1 were higher. These extraction values were still smaller than those for Feed 1 since Feed 1 was contacted with fresh feed solution in which the ferric and cupric concentrations were the highest. At Stage 2 (Cycle 4), the residue from Stage 1 (Cycle 3) was contacted with fresh feed solution. Since most of iron and copper were leached at Stage 1 (Cycle 3), higher ferric and cupric concentrations were maintained, resulting in a higher extraction of copper for Feed 3. The similar analysis can be applied to Feeds 4 to 6. The iron extraction for Test 2 was 73–75% since the chalcopyrite concentrate (BMS) used for Test 2 contained pyrite, which is practically not attacked. The iron-containing species in the bornite concentrate (ANT) were attacked during chloride leaching. The extractions of both copper and iron reached 99%. The concentrations of copper and iron in the leach solution for each cycle are shown in Figs. 10 and 11 for Tests 2 and 3 respectively. The concentrations of copper and iron in the leach solution from Stage 1 fluctuated slightly since they were mainly limited by ferric and cupric concentrations in feed solutions. The concentrations of

Test No.

Mass / g

Cu / %

Fe / %

Stotal / %

S (SO24 )/%

S2- / %

S0 / %

1 2 3 4

72.9 72.6 26.9 29.6

0.43 0.62 2.07 6.33

20.2 20.4 0.64 2.62

73.4 75.3 85.4 78.6

0.33 0.02 b0.01 0.02

20.2 20.1 5.4 11.6

52.9 55.2 80.0 66.9

copper and iron in the leach solution from Stage 2 fluctuated significantly since they were not only affected by the contents of copper and iron in the residues from Stage 1, but also the ferric and cupric concentrations in feed solutions. For Test 1, in Cycle 2 the copper extraction at Stage 1 was only 3.9% while the iron extraction was 12.6%. This fine chalcopyrite concentrate with lead and zinc sulfides was very reactive. During Cycle 2 (Stage 1), after the free cuprous concentration or solution potential reached a certain value, Reactions 7 to 9 became significant and the copper reprecipitation rate became larger than the leaching rate, resulting in a lower copper extraction. Therefore a sufficiently high cupric or ferric concentration must be maintained to avoid the re-precipitation of copper. A short leaching time (probably 2 h) may be used in this fine chalcopyrite concentrate. The extractions of copper and iron in Cycle 6 are summarized in Table 5. With decreasing particle size from P80 of 30 to 15 μm, the extraction of copper from chalcopyrite increased from 98.7 to 99.1% while the iron extraction practically did changed. With decreasing particle size of bornite concentrate from p80 of 40 to 15 μm, the extractions of copper and iron increased by 3.7 and 3.6% respectively although the cupric concentration in the feed for the finer (p80 of15 μm) bornite concentrate was 4 g/L less than that for the coarse (p80 of 40 μm) one. The above phenomena are related to the mineralogy of the two concentrates. The chalcopyrite concentrate contained pyrite, which is not readily attacked. The unleached iron was present as pyrite. Finer grinding did not improve leaching of iron from pyrite. The contents of copper, iron and sulfur species in the residues from Stage 2 in Cycle 6 are given in Table 6. There was virtually no sulfate in the residues since they were well washed and sulfate was detected in wash solutions. The residue mainly consisted of elemental sulfur. The calculated head contents of copper, iron and total sulfur are very close to the assayed values. For Tests 3 and 4, the mass loss was 70–73% with the bulk of the residue consisting of elemental sulfur. For Test 3, at 80% of element sulfur, the residue mass could be reduced by 5 times after the complete removal of elemental sulfur. This could correspond to an overall 4

Concentration / M

Concentration / g L-1

80

45

3

Stage 1 Total [Cu+ Fe] Stage 1 [Cu(I)+Fe(II)] Stage 1 [Cu(II)+Fe(III)]

2

1

Table 5 Summary of four chloride leaching tests.

Cycle 1 Cycle 2

Test No.

1

2

3

4

Concentrate Type Particle size (P80) / μm Cu extraction / % Fe extraction / %

Chalcopyrite 14 99.1 73.3

Chalcopyrite 30 98.7 73.1

Bornite 14 98.5 99.0

Bornite 40 94.8 95.4

Cycle 3 Cycle 4 Cycle 5 Cycle 6

0 0

3

6

9

12

15

18

Time / h Fig. 12. Concentrations of total copper and iron, Cu(I) + Fe(II) and Cu(II) + Fe(III) at Stage 1 for each cycle in Test 2 (P80 of 30 μm and chalcopyrite concentrate).

46

J. Lu, D. Dreisinger / Hydrometallurgy 138 (2013) 40–47

2.5

Stage 2 Total [Cu+ Fe] Stage 2 [Cu(I)+Fe(II)] Stage 2 [Cu(II)+Fe(III)]

3

2.0

Concentration / M

Concentration / M

4

2

1

1.5

1.0

Stage 2 Total [Cu+ Fe] Stage 2 [Cu(I)+Fe(II)] Stage 2 [Cu(II)+Fe(III )] Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6

0.5

Cycle 1 Cycle 1 Cycle 3

Cycle 4 Cycle 5 Cycle 6

0

0.0 0

3

6

9

12

15

18

0

3

6

Time / h

9

12

15

18

Time / h

Fig. 13. Concentrations of total copper and iron, Cu(I) + Fe(II) and Cu(II) + Fe(III) at Stage 2 for each cycle in Test 2 (P80 of 30 μm and chalcopyrite concentrate).

Fig. 15. Concentrations of total copper and iron, Cu(I) + Fe(II) and Cu(II) + Fe(III) at Stage 2 for each cycle in Test 3 (P80 of 14 μm and bornite concentrate).

remaining mass of 5.4 g or an overall mass loss of 94.6%. This is significant in terms of ability to recover gold from the concentrate in a concentrated residue.

chromium, nickel and zinc were practically completely extracted into solution, and the particle size virtually had no effect. With decreasing particle size from 30 to 15 μm, the extractions of antimony and silver increased from 88 to 96%, and 90 to 97% respectively. The extractions of both arsenic and bismuth were around 94% for the particle sizes of P80 of 14 and 30 μm. For bornite concentrate (Tables 9 and 10), lead was virtually completely extracted into solution for both the particle sizes (P80 of 14 and 40 μm). With decreasing particle size from 40 to 14 μm, the extractions of antimony, arsenic, silver, nickel and zinc increased from 32 to 43%, 46 to 80%, 95 to 99%, 80 to 83%, and 93 to 99% respectively. The extractions of bismuth and chromium were 98 and 96- 97% respectively and insensitive to the particle size (P80 of from14 to 40 μm).

3.4. Cupric/cuprous and ferric/ferrous species The concentrations of total Cu and Fe, Cu(I) + Fe(II) and Cu(II) + Fe(III) are given in Figs. 12 and 13 for Test 2. For Stage 1, the concentrations of total Cu and Fe and Cu(I) + Fe(II) increased from 2.2 to 3.3 M and from 0 to 3.1 M respectively and then fluctuated after Cycle 2. The concentration of Cu(II) + Fe(III) decreased from the feed concentration (2.2 M) to 0.2 M. For Stage 2, the concentrations of total Cu and Fe, Cu(I) + Fe(II) and Cu(II) + Fe(III) fluctuated greatly since the contents of copper and iron in the residue from Stage 1 fluctuated. The fluctuation decreased with increasing number of cycles. The concentration of Cu(II) + Fe(III) was stabilized around 0.2 M. This value was slightly higher than the target value (0.13 M) mainly because there were some ferric and cupric salts or oxides in the concentrate, and cuprous and ferrous ions were oxidized to cupric and ferric ions by oxygen. The similar leaching behaviour was observed for Test 3 (Figs. 14 and 15). 3.5. Extractions of minor and trace metals The leaching results of minor and trace metals are summarized in Tables 7–10. For chalcopyrite concentrate (Tables 7 and 8), lead, 2.5

Concentration / M

2.0

Stage 1 Total [Cu+ Fe] Stage 1 [Cu(I)+Fe(II)] Stage 1 [Cu(II)+Fe(III)]

1.5

1.0

0.5

0.0

Cycle 1 Cycle 2 Cycle 3 Cycle 4 Cycle 5 Cycle 6 0

3

6

9

12

15

18

Time / h Fig. 14. Concentrations of total copper and iron, Cu(I) + Fe(II) and Cu(II) + Fe(III) at Stage 1 for each cycle in Test 3 (P80 of 14 μm and bornite concentrate).

4. Conclusions In the presence of 165–200 g/L calcium chloride, ferric and cupric chloride leaching of copper from chalcopyrite and bornite concentrates was very effective under the atmospheric conditions using a two-stage countercurrent circuit with a retention time of 3 h. At a particle size of 14 μm, copper extractions were 99.1 and 98.5% for chalcopyrite and bornite concentrates respectively while the iron extractions were 73.3 and 99% respectively. The extraction of copper from the chalcopyrite concentrate decreased by 0.4% with increasing particle size to 30 μm while that of iron practically did not changed. The extractions of copper and iron from the bornite concentrate with a particle size of 40 μm were about 3.5% lower than those with 14 μm. The sulfur oxidation to sulfate was 1.2% to 1.7% for the leaching of chalcopyrite concentrate, and 2.1% for the leaching of bornite concentrate. The oxidation of sulfide to sulfate resulted in an increase in the HCl concentration. The extractions of Bi, Cr, Pb, Ag and Zn were very high for both chalcopyrite and bornite concentrates. The extractions of Sb, As and Ni were high for chalcopyrite concentrate while they low for bornite concentrate. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was 0.1 to 0.2 M. These cupric and ferric ions can easily be removed

Table 7 Extraction results of minor/trace metals for Cycle 6 of Test 1 (chalcopyrite concentrate and P80 of 14 μm). Elements

Sb

As

Bi

Cr

Pb

Ni

Ag

Zn

Head / % 0.57 0.18 0.041 0.074 6.78 0.035 0.269 2.63 Residue / % 0.040 0.020 0.0055 0.0032 0.024 b0.0001 0.017 0.018 Extraction / % 96 95 93 98 99.8 100 97.1 99.7 Leachate / g L −1 0.73 0.20 0.061 0.092 11.4 0.046 0.393 3.95

J. Lu, D. Dreisinger / Hydrometallurgy 138 (2013) 40–47

47

Table 8 Extraction results of minor/trace metals for Cycle 6 of Test 2 (chalcopyrite concentrate and P80 of 30 μm). Elements

Sb

As

Bi

Cr

Pb

Ni

Ag

Zn

Head / % Residue / % Extraction / % Leachate / g L −1

0.57 0.15 88 0.80

0.18 0.022 94 0.21

0.041 0.0050 95 0.062

0.074 0.0003 99.8 –

6.78 0.027 99.8 9.67

0.035 0.0005 99.3 0.043

0.269 0.060 89.9 0.362

2.63 0.048 99.2 3.94

Table 9 Extraction results of minor/trace metals for Cycle 6 of Test 3 (bornite concentrate and P80 of 14 μm). Elements

Sb

As

Bi

Cr

Pb

Ni

Ag

Zn

Head / % Residue / % Extraction / % Leachate / g L−1

0.077 0.095 43 0.023

0.62 0.44 80 0.43

0.50 0.024 98 0.41

0.019 0.0022 97 0.017

0.352 0.0018 99.9 0.26

0.0082 0.0049 83.1 0.0071

0.0950 0.0042 98.9 0.075

6.23 0.126 99.4 4.95

Table 10 Extraction results of minor/trace metals for Cycle 6 of Test 4 (bornite concentrate and P80 of 40 μm). Elements

Sb

As

Bi

Cr

Pb

Ni

Ag

Zn

Head / % Residue / % Extraction / % Leachate / g L −1

0.077 0.17 32 0.012

0.62 1.08 46 0.29

0.50 0.035 98 0.39

0.019 0.0023 96 0.015

0.352 0.0041 99.6 0.345

0.0082 0.0053 80 0.0050

0.0950 0.015 95.1 0.060

6.23 1.33 93.4 4.37

from the leach solution by adjusting pH for subsequent simultaneous cuprous oxidation and cupric solvent extraction to transfer copper from the chloride system to the conventional sulfate system. The impurities are separated from copper during solvent extraction. The advantages of chloride leaching (rapid leaching kinetics) can be utilized while the disadvantages of chloride systems (inefficient purification and poor quality copper cathodes) can be avoided by the transfer of copper to a conventional sulfate electrowinning circuit. Acknowledgements The authors would like to thank Falconbridge Limited (now XSTRATA) for funding this project and allowing for publication. References Asano, S., Imamura, M., Takeda, K., Ando, K., Nagase, N., 2007. Chloride processing of chalcopyrite copper concentrate, Copper 2007 Short Course. Dutrizac, J.E., 1990. Elemental sulfur formation during the ferric chloride leaching of chalcopyrite. Hydrometallurgy 23, 532–564.

Dutrizac, J.E., 1992. The leaching of sulfide minerals in chloride media. Hydrometallurgy 29 (1992), 299–304. Dutrizac, J.E., Chen, T.T., Jambor, J.L., 1985. Mineralogy changes occurring during the ferric ion leaching of bornite. Metall. Trans. B 16, 679–693. Hyvarinen, O., Hamalainen, M., 2005. HydroCopper™ – a new technology producing copper directly from concentrate. Hydrometallurgy 77, 61–65. Liddicoat, J., Dreisinger, D., 2007. Chloride leaching for chalcopyrite. Hydrometallurgy 89, 323–331. Lu, J., Dreisinger, D., 2013. Copper chloride leaching from chalcopyrite concentrate. Miner. Eng. 45, 185–190. McDonald, G.W., Langer, S.H., 1983. Cupric chloride leaching of model sulfur compounds for simple copper ore concentrations. Metall. Trans. B 14, 559–569. McDonald, G.W., Udovic, T.J., Dumesic, J.A., Langer, S.H., 1984. Equilibria associated with cupric chloride leaching of chalcopyrite concentrate. Hydrometallurgy 13, 125–135. Moyes, J., Houllis, F., Bhappu, R.R., 2000. The Intec Copper process demonstration plant. 5th Annual Copper Hydromet Roundtable'99 International Conference; Phoenix, AZ; USA; 10 Oct. 1999. Randol International, pp. 65–72. O’Malley, M.L., Liddell, K.C., 1987. Leaching of CuFeS2 by aqueous FeCl3, HCl and NaCl: effects of solution composition and limited oxidant. Metall. Trans. B 18, 505–510. Pesic, B., Olson, F.A., 1983. Leaching of bornite in acidified ferric chloride solutions. Metall. Trans. B 14, 577–588. Peters, E., 1977. Applications of chloride hydrometallurgy to treatment of sulfide minerals. Chloride Hydrometallurgy, Benelux Metallurgie, Brussels, pp. 1–36.