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Recovery of metal values from copper slags by flotation and roasting with pyriteG. Bulut, K.T. Perek, A. Gl, F. Arslan and G. nalAssistant professor, research assistant, assistant professor, professor and professor, respectively, Mining Engineering Department, Mineral and Coal Processing Division, Istanbul Technical University, Istanbul, Turkey

AbstractThe recovery of copper and cobalt from ancient copper slags from the Kre region of Turkey was investigated. A fayalitic-type of ancient copper slag containing 1.24% Cu, 0.53% Co and 53.16% Fe were subjected to an experimental study. Two different routes were followed to recover the metal values. The first route was roasting the slag with pyrite followed by leaching. The second route was flotation of the slags, roasting of the flotation tailings with pyrite and leaching. The second route was found to be suitable for the treatment of the copper slag studied. In the flotation step, a copper concentrate containing approximately 11% Cu was produced with a 77% recovery while 93% of the cobalt stayed in the tailings. In the roasting experiments, the effects of roasting time, pyrite:slag ratio and roasting temperature on the dissolution efficiencies of Cu and Co were investigated, and the optimum conditions were determined. In the tests, 87% of the cobalt dissolution was observed after one hour of roasting at 500C with a pyrite:slag ratio of 3:1. The copper dissolution in the tests was 31%. Based on the results of this study, a process flowsheet is proposed for the evaluation of these slags.

Keywords: Copper slag, Cobalt, Flotation, Roasting

IntroductionVarious slags are produced as byproducts in metallurgical processes or as residues in incineration processes. Slags usually contain a quantity of valuable metals, and such slags can actually be a secondary resource of metals rather than an end waste. They have been applied as a resource material in many areas. In addition, for some applications, slags have comparable or even better properties than their competitive materials.

There are a variety of nonferrous slags produced from nonferrous smelters. Extensive studies on metal recovery from nonferrous slags have been carried out over the last several decades. Most of the studies were on metal recoveries from copper slags. The chemical compositions of copper slags from different origins are quite different depending on factors such as the type of ores processed, the type of furnaces used and the cooling methods. When copper slag is crystalline, the major phases are usually fayalite along with other silicates. In most cases Ni and Co are in the form of oxides. Some studies show that Co distribution in the slags is very homogeneous. However, the forms of Cu minerals are different in different copper slags. They may be in the form of oxides or sulfides or a mixture of them. Copper sulfide minerals are also different in different copper slags.

A number of methods for metal recovery from copper slag are briefly reviewed (Ycel et al., 1992; Bipra and Jana, 2003; Shen and Forssberg, 2003). Basically, they can be classified into the following three categories: flotation, leaching and roasting. In principle, copper slag flotation is the same as sulfide ore flotation (Osborn et al., 1986; Giray et al., 1996; Arslan et al., 2002; Sarrafi et al., 2003; Bruckard et al. 2004). This means that only metallic Cu and sulfide minerals can be floated effectively. Because in some slags Cu occurs in the oxide form (Co is in homogenous distribution) and is disseminated in small amounts in the slag matrix, the application of flotation to copper slag processing could be limited, and the flotation method will not be effective for the recovery of Co, Ni and oxide Cu. In spite of some positive attributes of hy-droxametes, which are known as collectors for nonsulfide minerals (Lee et al., 1998), they may not be efficient in the flotation of oxide copper in the ores and slags.

Leaching is another method for the extraction of met-als from copper slags. Leaching with nitric, perchloric, hydrochloric and sulfuric acids; ferric chloride; ferric sulfate; cyanide; SO2; and ammonia solutions and the use of pressure leaching were among the hydrometallurgical processes studied (Herreros, et al., 1998; Banza et al., 2002). The combination of the pyrometallurgical process

Paper number MMP-06-016. Original manuscript submitted online February 2006. Revised manuscript accepted for publica-tion June 2006. Discussion of this peer-reviewed and approved paper is invited and must be submitted to SME Publications Dept. prior to Aug. 31, 2007. Copyright 2007, Society for Mining, Metallurgy, and Exploration, Inc.

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with leaching was also investigated. These methods consist of roasting with ferrous sulfate, ferric sulfate, pyrite, am-monium sulfate and sulfuric acid followed by water leaching (Arslan, 1982, Sukla et al., 1986; Tmen and Bailey, 1990; Altundoan and Tmen, 1997; Ziyadanoullar, 2000; Arslan and Arslan, 2002; Gl et al., 2003; Bulut et al., 2004; Giray et al., 2005).

There are 2.5 Mt (2.75 million st) of ancient copper slags located in the Kre region of Turkey, and they contain con-siderable amounts of Cu (1.24%) and Co (0.53%). The aim of this experimental study was to investigate the possibilities of recovering Cu and Co from these ancient copper slags by applying flotation and roasting of the flotation tailings with pyrite followed by leaching.

Material and methodsThe ancient copper slags taken from the Kre region of Turkey contain considerable amounts of copper and cobalt. Repre-sentative samples were taken from stockpiles of these ancient slags. Physical, chemical and mineralogical properties of the slag sample were determined prior to the concentration tests. The chemical analysis results are given in Table 1.

Mineralogical studies of the slag sample indicated the existence of 60% to 65% fayalite (Fe2SiO4), 15% to 17% wustite (FeO), 5% to 6% leucite (K(AlSi2O6)), 5% to 6% hercynite (FeAl2O4), 4% to 5% vitreous phase and 4% to 5% other minerals and metals such as magnetite (Fe3O4), pyrrhotite (FeS), limonite (FeO-OH), chalcopyrite (CuFeS2), digenite (Cu9S5), chalcocite (Cu2S), covellite (CuS), metallic copper, metallic iron, hematite (Fe2O3), pyrite (FeS2), wurtz-ite ((Zn,Fe)S) and cuprite (Cu2O). The maximum size of the sulfide minerals in the slag was measured to be 50 to 60 m using an optical microscope. Although Co minerals were not observed during the microscopic examinations, some copper and cobalt oxide minerals were determined by XRD analyses (Bulut, 2006). On the other hand, it is known that this type of slag, which is usually fayalitic based, contains Cu and Co in the form of silicates and ferrites (Arslan and Arslan, 2002; Shen and Forssberg, 2003).

In this study, flotation and roasting were applied for re-covering the main metallic values, such as the copper and cobalt in the slag. Flotation experiments using a conventional flotation unit (Denver flotation machine) were performed under the following conditions: 15% solids in pulp and an agitation speed of 1,300 rpm. Hostaflot M-91, Potassium Amyl Xanthate (KAX), Hostaflot X-231, Hostaflot LSB and Aero 211 were used as collectors; Na2S (300 g/t) was used as sulfidizing control reagent; and methyl isobutyl carbinol (MIBC) was used as frother. The effects of particle size, pH, type of collector and amount of collector on flotation were systematically investigated. Two-stage cleaning was applied in all flotation tests.

Two different routes were followed in the roasting experi-ments. The first route was roasting of the slags with pyrite.

The second route was flotation of the slag followed by roast-ing of the flotation tailings with pyrite. The pyrite used in the roasting experiments contained 0.22% Cu, 0.034% Co and 44.73% Fe and was in the size range of -0.1 mm. In the pyrite roasting tests with slag, the effects of roasting time, the pyrite:slag ratio and the roasting temperature on the dis-solution efficiencies of copper and cobalt were investigated. The pyrite roasting was followed by hot water or diluted acid leaching and filtration.

The roasting experiments with pyrite were carried out as follows: the slag/pyrite mixtures were placed in a porcelain dish and roasted in a muffle furnace preheated to the required temperature. After the samples were cooled to room tempera-ture, leaching was carried out at the following conditions: 300 mL water, 70C, a stirring speed of 450 rpm and a leaching time of one hour.

In the diluted acid leaching experiments, the effect of acid concentration was investigated. Standard glassware, a heater plus magnetic stirrer and a contact thermometer were used in the leaching experiments. After leaching, the leach slurry was filtered and all chemical analyses were made on the filtrates using an atomic absorption spectrometer.

Results and discussionsFlotation experiments. In the flotation tests the optimum conditions were determined to be as follows: pH 7 to 7.2, a particle size of -0.1-mm, 300 g/t Na2S and 200 g/t collec-tor. The results of the flotation tests performed at optimum conditions with different collectors are given in Table 2. The results after distribution of middlings (using Aero 211 collec-tor) are shown in Table 3. As can be seen from Tables 2 and 3, a copper concentrate containing approximately 10.7% Cu can be produced with a recovery of 76.7% (after middlings was distributed), while cobalt is distributed evenly among the products. A low flotation recovery (92.6% cobalt remained in the tailings) of cobalt was probably due to the formation of nonfloatable cobalt spinels and silicates in the slag. Therefore, the pyrite roasting method was conducted with the tailings to recover the remaining cobalt.

Pyrite roasting experiments. Pyrite roasting experiments were carried out on slags and flotation tailings that were discarded from the flotation experiment with Aero 211 collector. The flota-tion tailings contain 0.29% Cu, 0.57% Co and 51.43% Fe.

Pyrite roasting of slags: In the roasting experiments on the slags, the effects of roasting time, pyrite:slag ratio and roasting temperature were investigated. In the leaching tests, the effect of sulfuric acid concentration on the dissolution efficiencies of Cu and Co was investigated.

The effect of temperature was investigated at pyrite:slag ratios of 1:2 and 1:1 for a roasting time of one hour. The results are shown in Figs. 1 and 2, respectively. Increasing the roast-ing temperature above 450C caused the metal dissolutions to decrease.

The effect of pyrite:slag ratio was investigated at roasting temperatures of 450 and 500C for one hour. The results are illustrated in Figs. 3 and 4, respectively. Increasing the pyrite:slag ratio resulted in a gradual increase in Co dissolution, while Cu and Fe dissolution efficiencies started to decrease at a 1:1 pyrite:slag ratio.

The effect of roasting time was investigated at a pyrite:slag ratio of 1:1 at roasting temperatures of 450 and 500C. The results are given in Figs. 5 and 6, respectively. In both cases, after one hour of roasting (even 30 minutes at 500C), dissolution efficiencies stayed almost the same.

Cu 1.24 Pb Nil

Co 0.53 S 1.67

Ni 0.0035 SiO2 21.27

Zn 0.232 Fe 53.16

Element Content, % Element Content, %

Table 1 Chemical composition of the Kre copper slag.

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Aero 211 Concentrate 8.2 10.68 76.7 0.51 7.4

Tailings 91.8 0.29 23.3 0.57 92.6

Total 100.0 1.14 100.0 0.56 100.0

Weight, Content, Recovery, Content, Recovery, Collector type Product % % % % %

Cobalt Copper

Table 3 Distribution of middlings.

M-91+KAX Concentrate 5.4 9.57 44.7 0.54 5.9

Middlings 21.1 1.78 32.4 0.46 19.6

Tailings 73.5 0.36 22.9 0.50 74.5

Total 100.0 1.16 100.0 0.49 100.0

M-91+X-231 Concentrate 11.2 5.31 57.0 0.53 12.1

Middlings 23.3 1.14 25.4 0.50 23.5

Tailings 65.5 0.28 17.6 0.48 64.4

Total 100.0 1.04 100.0 0.49 100.0

M-91+LSB Concentrate 3.0 10.35 28.7 0.43 2.6

Middlings 24.7 1.90 42.5 0.48 24.3

Tailings 72.3 0.44 28.8 0.50 73.1

Total 100.0 1.10 100.0 0.49 100.0

Aero 211 Concentrate 7.0 10.68 65.6 0.51 6.3

Middlings 8.8 1.66 12.9 0.49 7.7

Tailings 84.2 0.29 21.5 0.57 86.0

Total 100.0 1.14 100.0 0.55 100.0

Weight, Content, Recovery, Content, Recovery, Collector type Product % % % % %

Cobalt Copper

Table 2 The effect of collector type on the flotation of slag sample.

Figure 1 Effect of roasting temperature on metal dis-solutions for a pyrite:slag ratio of 1:2.

Figure 2 Effect of roasting temperature on metal dis-solutions for a pyrite:slag ratio of 1:1.

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Figure 6 Effect of roasting time on metal dissolutions at a roasting temperature of 500C.

Figure 5 Effect of roasting time on metal dissolutions at roasting temperature of 450C.

In the hot water leaching step following the roasting step, the effect of acid concentration was investigated at a pyrite:slag ratio of 3:1, a roasting temperature of 500C and a roast-ing time of one hour. The results are shown in Fig. 7. The best result was obtained at an acid concentration of 10 g/L, where cobalt dissolution was 83.5% and the copper dissolution was 38%. The direct application of pyrite roasting on slag resulted in high cobalt dissolution efficiencies but very poor copper dissolution efficiencies. Therefore, pyrite roasting would be a more efficient method for recovering cobalt after copper is recovered by flotation.

Pyrite roasting of flotation tailings: Pyrite roasting experi-ments were also carried out on flotation tailings to recover cobalt. The pyrite:slag ratio and acid concentration were found to be the most effective parameters for cobalt dissolution in

the earlier tests. Therefore, only the effects of these parameters were investigated on the dissolution of metals from the flota-tion tailings. The effect of pyrite/tailings ratio is shown in Fig. 8, while the effect of acid concentration in hot water leaching after roasting is given in Fig. 9. When the acid concentration increases, the dissolution efficiency of cobalt increases up to 86.5%, while copper dissolution efficiency is 31.4% and iron dissolution efficiency is 22.4%.

The leach residue after roasting pyrite and leaching was also analyzed to determine the possibility of being used in the iron and steel industry as a raw material. Table 4 shows the chemical composition of the leach residue compared to the charge material used in the iron industry. All analyses fit within the range for of the charge material used in the iron industry except for sulfur.

Figure 4 Effect of pyrite:slag ratio on metal dissolutions at a roasting temperature of 500C.

Figure 3 Effect of pyrite:slag ratio on metal dissolutions at a roasting temperature of 450C.

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ConclusionsBased on the results of this experimental study, the flotation of slags followed by roasting of the flotation tailings with pyrite prior to its leaching was found to be the most efficient process for the treatment of the ancient copper slag studied. In the flotation of slags, a copper concentrate containing ap-proximately 10.68% Cu can be produced with a 77% recovery, while 93% of the cobalt stayed in tailings. Therefore, pyrite roasting was conducted on the flotation tailings to recover the remaining cobalt. Cobalt was leached with an 86.5% dissolution efficiency at a roasting temperature of 500C, a roasting time of one hour and a pyrite:slag ratio of 3:1. An attempt will be made to recover the metals, especially cobalt, by roasting the material together with the pyrite concentrate now produced in the copper ore beneficiation plant in Kre-Turkey.

Fe 50.0-65.0 61.04

Mn 0.1-2.0 0.015

P (0.05-0.3) 0.021

S 0.05-2.0 2.36

SiO2 1.0-15.0 7.66

Al2O3 1.0-10.0 1.96

CaO 0.3-3.0 0.09

MgO 0.05-3.0 0.16

Sn

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and environmental viewpoints. The results of this research may also help in evaluating similar types of slags located in the Samsun and Ergani regions of Turkey, as well as slags that exist in other parts of the world.

AcknowledgmentsThe authors thank Istanbul Technical University-Scien-tific Research Project Center (BAP) for the support provided throughout this research.

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Figure 10 Proposed flowsheet for the evaluation of Kre ancient copper slag.

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