Extracting antimony from high arsenic and gold-containing stibnite ore using slurry electrolysis

Hydrometallurgy 186 (2019) 284–291

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Extracting antimony from high arsenic and gold-containing stibnite ore using slurry electrolysis ⁎

T

⁎

Yonglu Zhanga,b, Chengyan Wanga, , Baozhong Maa, , Xiaowu Jiea,b, Peng Xinga a b

School of Metallurgical and Ecological Engineering, University of Science and Technology Beijing, Beijing 100083, China BGRIMM Technology Group, Beijing 100160, China

A R T I C LE I N FO

A B S T R A C T

Keywords: Antimony Arsenic Gold Stibnite Slurry electrolysis

This study investigated the leaching and electrowinning of antimony from a complex, high arsenic and goldcontaining stibnite concentrate ore (As-Au-Sb ore) using slurry electrolysis (SE). The results indicated that the leaching efficiency of antimony was higher than 98%, and those of iron and arsenic were about 5.50% and 1.14%, respectively. An energy consumption (Ah) of 1.1 times the ‘benchmark’ value, a duration of 6 h, temperature of 60 °C, 30 g/L HCl, 5 g/L Fe, 30 g/L antimony, and an anodic current density of 80 A/m2 were found to be the optimal experimental conditions. The SE selectively leached antimony, while gold and arsenic were retained in the residue as a raw material for further gold extraction. The antimony content in the cathode antimony was > 98%, and arsenic content was less than or equal to 0.21%. The As-Au-Sb ore consisted of stibnite (Sb2S3), pyrite (FeS2), arsenopyrite (FeAsS), and quartz (SiO2), along with a small amount of muscovite (KAl2(AlSi3O10)(OH)2) and other gangue phases. Antimony mainly occurred in stibnite. During the SE, stibnite disappeared, leaving behind elemental sulfur. We proposed a comprehensive procedure for the separation and recovery of antimony from the As-Au-Sb ore. In this process, antimony was recovered in a single step in the form of cathode antimony (> 98%), thereby resolving the technical problem of antimony and arsenic separation and hence, preventing the contamination of plenty of arsenic soda slag and low-concentration SO2 that exists in the traditional pyrometallurgical process. The process was environmentally friendly and had advantages such as short flow and high resource utilization rate.

1. Introduction Antimony is a lustrous, silver-white, non-malleable metal with poor electrical and thermal conductivity. At room temperature, antimony cannot be oxidized by humid air or pure water. Metallic antimony is too brittle to be used alone and is always used as an alloy or compound (Anderson, 2012; Zhao, 1987). Antimony is used as a hardening agent in lead alloy, which is consequently used in lead storage batteries, munitions, printing industries, wear-resisting bearings etc. Antimony compounds are utilized in flame retardants, semiconductors, and pharmaceuticals (Awe and Sandström, 2013; Multani et al., 2016). Among these, the major usage of the metal is in flame-retardants, accounting for > 60% of the total antimony consumption. The most important antimony resource is the sulfide mineral, stibnite (Sb2S3) (Multani et al., 2016; Zhao, 1987). The complex As-Au-Sb ore is also an important antimony resource that is predominantly mined in Hunan and other places in China. It is essential to develop an economical and environmentally friendly process for the extraction of

⁎

antimony from this ore. Antimony production involves hydrometallurgy and pyrometallurgy, the latter being still dominant. The conventional extraction of antimony from stibnite and jamesonite involves a two-step pyrometallurgical process comprising volatilization roasting and carbothermal smelting (Ouyang et al., 2019; Tian et al., 2016). The sulfide antimony ore is first volatilized and oxidized into antimony trioxide (Sb2O3) in a blast furnace, and then Sb2O3 is reduced to antimony metal in a reverberatory furnace (Hua et al., 2003; Qin et al., 2015). However, the traditional pyrometallurgical process requires a high-grade concentrated ore and high energy and causes serious environmental contamination by the low-concentration SO2, various heavy metals, and arsenic-alkali residue. Hence, it is unsuitable for treating low-grade, complex antimony ores. (Mahlangu et al., 2006; Sun, 2012; Wang and Lei, 2000; Yang et al., 2017). In order to reduce the energy consumption and eliminate the SO2 contamination, many researchers have studied the pyrometallurgy of antimony. Liao et al. (2010) used an oxygen-enriched blast furnace

Corresponding authors. E-mail addresses: [email protected] (C. Wang), [email protected] (B. Ma).

https://doi.org/10.1016/j.hydromet.2019.04.026 Received 30 October 2018; Received in revised form 19 March 2019; Accepted 21 April 2019 Available online 22 April 2019 0304-386X/ © 2019 Elsevier B.V. All rights reserved.

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decreasing the electricity consumption and cell voltage for electrolysis. SE can transform the sulfide (S2−) in sulfide minerals into elemental sulfur (S0). This not only prevents the generation of environmentally hazardous SO2 but also avoids the inconvenience of transport and storage of the sulfuric acid produced during pyrometallurgical processing. SE has been developing for almost forty years. It has been used to treat single sulfide minerals of Cu, Pb, and Bi (Everett, 1981, 1983; Wang et al., 1995); complex multi-metal sulfide minerals (Wang et al., 2002b; Wang et al., 2006a; Zhang et al., 2002); and ocean nodules (Wang et al., 2010) and MneCo containing oxides (Wang et al., 2006b). Wang et al. (2002b) used SE to treat refractory jamesonite in a medium of HCl + NH4Cl, wherein antimony and lead were separated in a single step, and metal antimony was directly obtained at the cathode. Meanwhile, the lead in jamesonite remained in the residue as PbCl2. Zeng et al. (1991) and Zhang et al. (1998) used SE to extract antimony from a sulfide antimony ore of stibnite in 2 mol/L HCl system. Although the antimony leaching rate was found to reach > 94% and 98% in the respective studies, some of the problems were still unsolved. In the study by Zeng et al. (1991), the newly formed elemental sulfur adhered to the surface of the anode and hindered the electrolysis, while in the study by Zhang et al. (1998), the cell voltage was high because of the generation of chlorine gas (Cl2) on the anode. Until recently, only a little information about treating complex high arsenic and gold-containing stibnite concentrate ore (As-Au-Sb ore) was available in the literature. In this study, we introduced a comprehensive process involving SE for the separation of antimony from the As-Au-Sb ore (Fig. 14). Antimony was selectively extracted in the anode, and metal antimony was formed in the cathode; gold and arsenic were retained in the residue. We also investigated the factors affecting the antimony extraction，such as energy consumption, duration, temperature, HCl concentration, initial iron and antimony concentrations, and anodic current density.

volatilization roasting process to extract antimony from stibnite (Sb2S3). The oxygen-enriched air was pumped into the blast furnace, which improved the production efficiency and productivity and reduced the energy consumption (Luo and Liu, 2015). However, the SO2 concentration in flue gas was still low, and the treatment of the flue gas was difficult. Some other researchers investigated the oxygen-enriched molten pool smelting process, such as top blowing, side blowing, and bottom blowing, to smelt the antimony sulfide ore (Duan, 2010; Wang and Wang, 2015; Lei and Wang, 2001; Liu et al., 2014; Chen, 2015). Compared with the process using the traditional blast furnace, this process has obvious advantages such as reduced energy consumption and increased concentration of SO2 in the flue gas that could be suitable for sulfuric acid production. Despite this, technical problems such as high Sb content in slag, relatively low purity of antimony trioxide (Sb2O3), and large fluctuation in the concentration of SO2 still exist. Ye et al. (2015) proposed a low-temperature molten salt smelting process that could directly produce antimony from its ore. Zinc oxide (ZnO) was added as a desulfurization agent, and sulfur was fixed in the form of ZnS. Under the optimum conditions, the average direct recovery rate of antimony reached 92.88%. In the study by Xu et al. (2017), FeO was used as desulfurization agent, and under the optimum conditions, the direct recovery rate of antimony was 91.48%. The antimony content in crude antimony was 94.31%. Although this process with short flow could eliminate the low-concentration SO2 pollution, the antimony recovery rate and the antimony content in crude antimony were low. In the treatment of high arsenic and gold-containing stibnite concentrate, the problem of arsenic alkali residue is still prominent, and the loss by gold dispersion is high. Compared with pyrometallurgy, hydrometallurgy, which involves alkaline (Na2S) and acidic (HCl) systems, is an environmentally friendly technique that has attracted significant attention because of its potential to deal with low-grade complex stibnite without emitting SO2 (Multani et al., 2016; Yang and Wu, 2014). China, Australia, and the United States have been employing the alkaline system at an industrial scale. (Anderson, 2012). In the alkaline system, the lixiviant is a mixture of sodium sulfide and sodium hydroxide. Antimony dissolves in the form of sodium thioantimonite (Na3SbS3) and is then extracted by electrowinning. However, both sodium polysulfide (Na2SX) and sodium thiosulfate (Na2S2O3) are severely accumulated in the system (Anderson, 2012; Wang et al., 2002a). Moreover, the use of alkaline system causes the leaching of gold during the treatment of the goldcontaining antimony ore (Jin et al., 2014). Although the alkaline system is widely used, investigations for employing a chloride-based technology are still being carried out. In the acidic chloride system, HCl is used as the lixiviant, and stibnite (Sb2S3) is dissolved by adding oxidizing agents such as Cl2, FeCl3, SbCl5, H2O2, or O3 (Guo et al., 2017; Liu et al., 1991; Tian et al., 2016; Yang and Wu, 2014). The sulfide (S2−) in Sb2S3 is oxidized into elemental sulfur (S0), and antimony gets converted into SbCl3 in the solution. The antimony in pregnant leach solution is mainly extracted in the form of antimony white and metal antimony, by hydrolysis and electrowinning, respectively. The differences in oxidizing agents lead to different flow-sheets, but all these processes need additional oxidizing agents in the leaching stage. Slurry electrolysis (SE) is a new hydrometallurgical technology that realizes the three main stages (leaching, partial solution purifying, and electrowinning) of traditional hydrometallurgy in an electrolytic bath (Qiu, 1999a, 1999b; Zhang et al., 2017). It can simultaneously leach the sulfide ore on the anode by oxidation and accumulate the metal on the cathode. Therefore, it lets the high energy-consuming anode reaction to efficiently leach the sulfide minerals by oxidation, while simultaneously

2. Materials and methods 2.1. Material characterization The raw material of As-Au-Sb ore used in this study was collected from the western region of Hunan province (China). It was a flotation concentrate, with particle size of 100% below 250 μm and 73.3% below 74 μm. Before the chemical composition analysis and characterization, the raw material and leaching residue sample were dried in an oven at 70 °C until a constant dry mass was attained. The metals were determined by inductively coupled–plasma atomic emissive spectrometry (IRIS Intrepid II, XRS), and the higher antimony contents (> 10% (wt %)) were determined by titration with ceric sulfate. The gold contents in the solid samples were determined by fire assay. The “silicon molybdenum blue spectrophotometry method” and “ combustion-neutralization titrimetric method” were adopted for the determining the silicon and sulfur contents, respectively. The raw material and leaching residue were further characterized by X-ray diffraction (XRD) with a Cu-Kα X-ray tube operated at 40 kV and 40 mA (Rigaku, Japan). The micro-morphologies of the solid samples were investigated with a scanning electron microscope (SEM, HITACHI S-3500 N), in combination with an energy dispersive spectroscope (EDS, Inca Oxford). Chemical analysis of the As-Au-Sb ore (Table 1) indicates that the main valuable metals are antimony and gold. It is clear from the XRD pattern (Fig. 1) that the ore consists of stibnite (Sb2S3), pyrite (FeS2), arsenopyrite (FeAsS), and quartz (SiO2), along with a small amount of muscovite (KAl2(AlSi3O10)(OH)2) and other gangue phases.

Table 1 Elemental composition of the As-Au-Sb ore. Component Percentage (wt%)

Sb 31.62

Fe 13.67

S 25.1

Cu 0.12

Mg 0.16

285

Ca 0.15

Si 7.49

As 4.4

Al 1.84

Au 32.88(g/t)

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cathode chamber. A continuous stirring speed was maintained using a Teflon-coated agitator blade. When the temperature increased to the desired value, a quantitative ore, which was pulped with the lixiviant, was added into the anode chamber, and the electric current was switched on to leach the ore. After leaching for the desired duration, the slurry in anode chamber was filtered by a vacuum pump. The volume of the filtered pregnant solution was measured by a measuring cylinder, and the liquid sample was collected. The filter cake was washed first with HCl (1 mol/L) and then with distilled water and dried at 70 °C for 12 h. The dried filter cake was weighed, and consequently, elemental and mineralogical analyses were conducted.

5000 T

Intensity (counts)

T

4000

T

M

3000

Q

2000 M T

1000 0

T. Stibnite P. Pyrite A. Arsenopyrite Q. Quartz M. Muscovite

T Q M

P PT T T A P A M A

Q

P Q

T

P T

T A Q T

P M

Q

5 10 15 20 25 30 35 40 45 50 55 60 65 70 75 2 (degree)

2.3. Chemical reaction at anode and cathode during slurry electrolysis

Fig. 1. XRD pattern of the As-Au-Sb ore.

The HCl–NaCl system, in which SbCl3 is highly soluble, was used as the electrolyte during the SE. Furthermore, a given mass of iron ions was added into the electrolyte to facilitate the anode reaction. The main possible anode and cathode reactions are discussed in the next two sections.

Based on the results of optical microscopy, semi-quantitative analysis of XRD, SEM-EDS detection, and chemical analysis, it was concluded that antimony was mainly present in stibnite, which was easily oxidized and leached during the SE. Gold and arsenic were present in pyrite and arsenopyrite, which were stable during the SE and were retained in the leaching residue. This laid the theoretical basis for separating antimony from gold and arsenic in the As-Au-Sb ore by SE. All the reagents were of analytical grade, and all the aqueous solutions were prepared using distilled water. The antimony extraction rate was calculated using the relation

α=

m 0 × x 0 − m1 × x1 × 100% m0 × x 0

2.3.1. Reactions at the anode The possible chemical reactions occurring at the anode during SE are discussed below. (1) Direct anodic oxidation In this case, the graphite anode is equivalent to a conductor, while the As-Au-Sb ore is equivalent to a soluble anode. When the As-Au-Sb ore collides and contacts the graphite anode, electrons are transferred from the ore to the graphite anode because of the electric field. The ore itself is oxidized, releasing Sb ions into the solution. The reaction can be represented by Eq. (2):

(1)

where α is the extraction rate (%) of antimony, m0 is the initial weight (g) of the ore, x0 is the Sb content (wt%) in the ore, m1 is the weight (g) of the leaching residue, and x1 is the Sb content (wt%) in the leaching residue.

Sb2 S3 = 2Sb3 + + 3S0 + 6e− E0 ox = −0.510 V

(2)

E0ox is the standard oxidation potential of the oxidation half reaction.

2.2. Experimental equipment and procedure

(2) Chemical dissolution

All the experiments were conducted in a cylindrical home-made electrolyzer (Fig. 2). Titanium strip and graphite stick were used as the cathode and anode, respectively. A polypropylene permeable membrane was used to divide the electrolyzer into separate concentric anode and cathode chambers. In the anode chamber, the slurry was agitated by an electric agitator. The electrolyzer was kept in a thermostatic water bath to maintain a constant temperature. During the experiment, the prepared solution was added first into the anode and then into the

The As-Au-Sb ore first reacts with acid to produce hydrogen sulfide (H2S) (Eq. (3)), which then oxidizes to elemental sulfur (Eq. (4)).

Sb2 S3 + 6HCl = 2Sb3 + + 6Cl− + 3H2 S

(3)

2Fe3 + + H2 S = 2Fe2 + + S0 + 2H+

(4)

(3) Chemical oxidation In the HCl–NaCl system, chlorine gas (Cl2) may be evolved on the graphite anode. The As-Au-Sb ore can get oxidized by Cl2, releasing the Sb ions into the solution. This can be represented by the following reactions:

2Cl‐ = Cl2 + 2e− E0 ox = −1.358 V

(5)

Sb2 S3 + 3Cl2 = 2Sb3 + + 6Cl− + 3S0

(6)

Iron ions in the solution are involved in leaching. At the anode, ferrous ions (Fe2+) are oxidized to ferric (Fe3+) ions (Eq. (7)), which can oxidize the As-Au-Sb ore, and then reduce to Fe2+ (Eq. (8)). The Fe2+ ions are reoxidized to the Fe3+ ions at the anode. Repeated oxidation and reduction results in complete leaching of antimony from the As-Au-Sb ore. Fig. 2. Schematic diagram of home-made slurry electrolyzer (1 Agitator blade, 2 Titanium cathode, 3 Polypropylene diaphragm bag, 4 Graphite anode, 5 Glass beaker). 286

Fe2 + = Fe3 + + e− E0 ox = −0.771 V

(7)

Sb2 S3 + 6Fe3 + = 2Sb3 + + 6Fe2 + + 3S0

(8)

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2.3.2. Reactions at the cathode The electrodeposition of antimony (Eq. (9)) is the main reaction at the cathode, while the reduction of Fe3+ to Fe2+ (Eq. (10)) occurs as a side reaction.

Sb3 + + 3e‐ = Sb

Fe3 +

+

e‐

=

E0 red = 0.24 V

Fe2 +

E0

red

T

Q

T

T

T T

Q T

(9)

(a)

S

(10)

= 0.771 V

T. Stibnite P. Pyrite Q. Quartz

T

Q

E0red is the standard reduction potential of the reduction half reaction. The above reaction mechanism reveals that the SE comprises three processes: (a) oxidization at anode leading to the leaching of the As-AuSb ore, (b) electrodeposition of antimony at the cathode, and (c) simultaneous oxidative regeneration of the leaching agent. Therefore, the oxidation reaction at the anode and the reduction reaction at the cathode are completely utilized.

S

C

Q

S

S

P C

15

20

25

P A

P A

A

S P A

PA

C

C

S

(c)

10

T

PA

P S

S

(b)

A. Arsenopyrite S. Sulfur C. Stibiconite

30

C

35

40

2 (degree)

Fig. 4. The XRD patterns comparison of raw material and leaching residues: (a) Raw ore (b) leaching residue at energy consumption of 1.1 times (c) leaching residue at energy consumption of 1.4 times.

3. Results and discussion the stibnite phase in the residue. However, an excessive energy consumption would oxidize the trivalent antimony (Sb3+) in the solution to the pentavalent antimony (Sb5+), which would easily get hydrolyzed and precipitated into stibiconite. The stibiconite will then be absorbed by the leaching residue, resulting in low Sb extraction. Thus, the optimum energy consumption (Ah) was 1.1 times the benchmark value.

3.1. Slurry electrolysis experiments In this study, we adopted the HCl–NaCl system, because the high solubility of NaCl can provide enough chloride ions to chelate with the antimony ions in the solution. In addition, NaCl is cheap and easily available.

3.1.2. Effect of duration The results for antimony extraction with increasing duration of leaching (from 3 to 8 h) are given in Fig. 5. From the initial point up to 6 h of leaching, the antimony extraction increased from 91.59% to 97.13%. However, the leaching time did not affect the antimony extraction after 6 h. Essentially，the dissolution of the As-Au-Sb ore in the anode chamber is a continuous oxidative leaching process of the antimony-containing mineral; the reaction duration of 6 h is sufficient for the dissolution of the soluble antimony mineral under the optimum energy consumption. The results suggested that 6 h was the optimal leaching duration for the selective extraction of antimony.

3.1.1. Effect of energy consumption In the SE, the oxidative leaching of sulfide minerals in the anode chamber depended on the energy consumption (Eqs. (7), (8)). We assumed that all the antimony in the ore was present in stibnite. The theoretical energy consumption (Ah) for the complete oxidative leaching of stibnite was considered as the benchmark. The relation between Sb extraction and the electrolytic energy consumption was investigated in terms of the energy consumption multiples, which were 0.8–1.4 times that of the benchmark value (Fig. 3). Fig. 3 shows that the Sb extraction increased from 89.61% to 97.78% in the energy consumption multiple range from 0.8 to 1.1 times. However, when the energy consumption multiple was > 1.1 times, the extraction rate of Sb decreased rapidly, and it was only 83.7% when the energy consumption was 1.4 times. The XRD patterns (Fig. 4) show a comparison of the antimony mineral phases in raw material and leaching residues at 1.1 times and 1.4 times energy consumption. Although stibnite disappeared in the two residues, a new mineral phase of stibiconite (Sb3O6(OH)) was found in the residue of 1.4 times energy consumption. We inferred that at an energy consumption below 1.1 times, stibnite was gradually oxidized and dissolved in the solution, which resulted in the disappearance of

3.1.3. Effect of temperature The temperature not only influences the speed of the ions in the solution and the rate of the chemical reaction leading to mineral leaching, but also influences the rate of the electrochemical reaction on the surface of the electrode (Wang et al., 2010). The leaching of antimony was investigated from 30 to 70 °C (Fig. 6). An increase in temperature from 30 to 50 °C led to a sharp increase in antimony extraction. However, increasing the temperature beyond 50 °C did not have any prominent influence on the final antimony extraction. In addition, an increase in temperature will increase the energy consumption and

100

90

Sb extraction (%)

Sb extraction (%)

100

80

70

0.8

0.9

1.0

1.1

1.2

1.3

95

90

85

1.4

Energy consumption (times) Fig. 3. Effect of energy consumption on antimony extraction (experimental conditions: duration of 6 h, temperature of 60 °C, 30 g/L HCl, 5 g/L Fe and anodic current density of 80 A/m2).

3

4

5 6 Duration (h)

7

8

Fig. 5. Effect of duration on antimony extraction (experimental conditions: energy consumption of 1.1 times, duration of 6 h, temperature of 60 °C, 30 g/L HCl, 5 g/L Fe, and anodic current density of 80 A/m2). 287

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100

100

Sb extraction (%)

Sb extraction (%)

98 90

80

96 94 92

70 25

30

35

40

45

50

55

60

65

70

90

75

60

80

100

120

140

Fig. 9. Effect of anodic current density on antimony extraction (experimental conditions: energy consumption of 1.1 times, duration of 6 h, temperature of 60 °C, 30 g/L HCl, and 5 g/L Fe).

Fig. 6. Effect of temperature on antimony extraction (experimental conditions: energy consumption of 1.1 times, duration of 6 h, 30 g/L HCl, 5 g/L Fe and, anodic current density of 80 A/m2).

2.0

100

Bath voltage (V)

Sb extraction (%)

95

90

1.6

300

1.4 1.2

200

1.0

Bath voltage Electric consumtion 100

0.8 0.6 40

80

100

120

140

160

Anodic current density (A/m2)

85

10

20

30

40

50 Fig. 10. The variation of bath voltage and electric consumption with current density(experimental conditions: energy consumption of 1.1 times, duration of 6 h, temperature of 60 °C, 30 g/L HCl, and 5 g/L Fe).

HCl concentration (g/L) Fig. 7. Effect of HCl concentration on antimony extraction (experimental conditions: energy consumption of 1.1 times, duration of 6 h, temperature of 60 °C, 5 g/L Fe, and anodic current density of 80 A/m2).

100

Sb extraction (%)

2.0

100 95

1.5 90 1.0

85 80 Sb extraction Bath voltage

75 0

2

4

6

8

10

0.5

Bath voltage (V)

Sb extraction (%)

60

Electric consumtion (kWh/t-ore)

400

1.8

70

160

Anodic current density (A/m2)

Temperature ( C)

95

90

85 0.0 12

0

5

10

15

20

25

30

Initial Sb concentration (g/L)

Initial iron ions concentration (g/L)

Fig. 11. Effect of initial Sb concentration on antimony extraction (experimental conditions: energy consumption of 1.1 times, duration of 6 h, temperature of 60 °C, 30 g/L HCl, 5 g/L Fe, and anodic current density of 80 A/m2).

Fig. 8. Effect of initial iron ions concentration on antimony extraction and bath voltage (experimental conditions: energy consumption of 1.1 times, duration of 6 h, temperature of 60 °C, 30 g/L HCl, and anodic current density of 80 A/m2).

is necessary to prevent the hydrolysis in a solution leached from ores. In the studies by Zeng et al. (1991) and Zhang et al. (1998), stibnite was extracted with 2–3 mol/L HCl. In this experiment, adding a certain amount of NaCl in the electrolyte provided sufficient Cl− ions for chelating the antimony ions. This increased the antimony solubility and significantly reduced the HCl concentration of the electrolyte system. The extraction of antimony was investigated with the HCl concentrations ranging from 10 to 50 g/L. Fig. 7 shows that the leaching efficiency of antimony reached 92.35% with 10 g/L HCl and then gradually increased to 97.63% with 30 g/L HCl. Further increase in the HCl concentration did not result in any significant increase in the antimony

the loss of electrolyte volatilization. Thus, temperatures between 50 and 60 °C were considered as the optimum temperatures. Other tests in this study were carried out at 60 °C. 3.1.4. Effect of HCl concentration Liu et al. (1991) reported that the antimony pregnant leach liquor with 1.3 mol/L HCl hydrolyzed to precipitate Sb4O5Cl2 upon diluting the leachate with water to 0.55–0.7 mol/L HCl. The study by Zheng et al. (1991) showed that the antimony concentration was about 0.1 g/L when HCl was diluted to 0.7 mol/L. Since antimony chloride is easily hydrolyzed, maintaining a certain HCl concentration in the electrolyte 288

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Table 2 Comprehensive tests of extracting antimony from the As-Au-Sb ore using SE. Leaching residue

Test 1

chemical composition(%) extraction(%) chemical composition(%) extraction(%) chemical composition(%) extraction(%) chemical composition(%)

Test 2 Test 3 Average

T

T

M Q Q

M T

M

T

Sb

As

Fe

Au

Sb

As

Fe

Au

1.01 98.08 0.89 98.31 0.98 98.14 0.96

7.22 1.55 7.27 0.86 7.26 1.00 7.25

21.36 6.25 21.75 4.54 21.48 5.72 21.53

54.20 g/t 1.09 54.76 g/t 0.07 54.40 g/t 0.73 54.45 g/t

98.78

0.21

0.02

0.12 g/t

98.95

0.15

0.03

0.11 g/t

99.01

0.18

0.03

0.14 g/t

98.91

0.18

0.03

0.12 g/t

T. Stibnite P. Pyrite Q. Quartz

T

PT

T

M A

P A

T A P Q QP

P

T

Region T

A Q

S

S

M

T

T P

1 2 3 4

Raw ore

S Q

Table 3 SEM-EDS analysis results of the SE leaching residue of comprehensive test (wt %).

A. Arsenopyrite S. Sulfur M. Muscovite

Q

M

Cathode antimony

P A M S Sp A S S

A P

SM Q

P Q

S

Q

P

P

A Q

10

15

20

25

30 35 2 (degree)

40

45

50

55

Mg-K

Al-K

46.92 0.35

1.9

Si-K

53.08 6.48

S-K

Fe-K

As-K

19.31 50.64

37.09 49.36

43.6

63.54

6.99

Sb-L

20.73

extraction was observed beyond an initial irons concentration of 2 g/L. Under the selected test conditions, adding iron ions can change the bath voltage. The bath voltage was 1.21 V in the absence of iron ions. However, when iron ions were added, the bath voltages were found to be 1.02, 0.93, 0.91, and 0.89 V, when the initial iron ions concentrations were 2, 5, 8, and 11 g/L, respectively.. Hence, a moderate amount of iron ions in the electrolyte reduces the bath voltage and the power consumed. However, this effect was no longer evident when the concentration of iron ions exceeded 5 g/L. In addition, a higher concentration of iron ions had an adverse effect on the SE. It increased the viscosity of the solution and reduced the current efficiency of the anode and cathode (Wang, 2002). Partial iron in the As-Au-Sb ore leached into the solution during the SE will get accumulated during the recycling of the electrolyte. In practical engineering applications, it is necessary to remove the iron regularly.

SE leaching residue 5

O-K

60

Fig. 12. The XRD patterns comparison of raw ore and SE leaching residue of comprehensive test.

extractions. Thus, 20–30 g/L HCl was chosen for further experiments. 3.1.5. Effect of initial iron ions concentration The presence of iron ions (Fe) in the electrolyte is important in SE. It directly participates in the anode and cathode reaction, as well as the oxidative leaching of the As-Au-Sb ore. It can not only transfer electrons but also change the anode reaction, which can reduce the bath voltage. The anodic reactions in the absence of iron ions are mainly expressed by Eqs. (5) and (6), while those in the presence of iron ions are expressed by Eqs. (7) and (8). The Sb extraction from ores and the corresponding variation of bath voltage were investigated with the initial iron ion concentration varying from 0 to 11 g/L. It is clear from Fig. 8 that when the initial solution was free of iron ions, the terminal Sb extraction was merely 79.94%, while it reached 96.10% when the initial concentration of iron ions was 2 g/L. However, no obvious increase in the terminal Sb

3.1.6. Effect of anodic current density During the SE, the antimony extraction from the As-Au-Sb ore is realized by the oxidation of anode. Therefore, the current density of the anode certainly influences the antimony leaching from the ore. This effect is manifested in two ways. First, as the anode current density increases, the amount of oxidizer available per unit time in the anode increases, and the amount of ore processed increases accordingly.

Fig. 13. The SEM image of (a) SE leaching residue of comprehensive test, (b) the region 4 in (a). 289

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antimony: 30 g/L, and anodic current density: 80 A/m2. The results are listed in Table 2. Antimony extraction reached over 98%, the average extraction of Fe, As, and Au were 5.50%, 1.14%, and 0.63%, respectively. Gold and aresenic remained in the leaching residue, which was further used as a raw material for gold extraction. Thus, the separation of antimony from gold and arsenic was achieved. The cathode metal antimony contained > 98% antimony and less than or equal to 0.21%, 0.03%, and 0.14% of As, Fe, and Au, respectively. A comparison of the XRD spectra of the leaching residue and raw ore is shown in Fig. 12. It indicates that the diffraction peak of stibnite almost disappeared after the SE, and plenty of elemental sulfur appeared in the leaching residue. Other minerals in the residue were pyrite, quartz, arsenopyrite, and a small amount of muscovite. The SEM images of SE leaching residue are shown in Fig. 13, and SEM-EDS analysis of the mineral phases in Fig. 13(a) are presented in Table 3. Combined with the XRD results in Fig. 12, it can be concluded that regions 1, 2, and 3 in Fig. 13(a) corresponded to arsenopyrite, pyrite, and quartz, respectively. Region 4 corresponds mainly to the elemental sulfur cluster produced during SE. This elemental sulfur had a porous structure (Fig. 13(b)) that could accommodate other mineral particles with smaller grain size, such as muscovite, residual antimony minerals, or leaching products, resulting in a complex elemental composition. Based on the above experiments, the process flow during SE for the extraction of antimony from the As-Au-Sb ore was proposed (Fig. 14). The As-Au-Sb ore is first pre-leached by a leaching agent, and then the solid–liquid is separated. The pre-leaching solution is used as a replenishing liquor, which is injected into the cathode chamber and subjected to electrodeposition on the cathode plate. Following this, the electrolyte permeates through the diaphragm bag into the anode chamber. The pre-leaching residue is slurried with the leaching liquid, added with HCl, and pumped quantitatively into the anode area to extract antimony through SE. During the SE, antimony is electrodeposited in the cathode area, and the As-Au-Sb ore is leached by electro-oxidation in the anode area. After leaching, the anode slurry is filtered to obtain the gold-containing leaching residue, the filtrate is added back to the pre-leaching and slurrying stage for recycling. The cathode plate is periodically removed from the electrolytic cell and stripped to obtain metal antimony (Sb > 98%). The SE treatment of the As-Au-Sb ore had the advantages of short flow, high recycling rate of resource, and environmental friendliness. In this process, antimony and arsenic were effectively separated to produce cathode metal antimony (> 98%) in a single step. The sulfur present in stibnite was transformed to elemental sulfur remaining in the residue. Therefore, the contamination by large amount of arsenic soda slag and low-concentration SO2 were avoided.

High arsenic gold-bearing stibnite ore

Pre-leaching

Pre-leaching solution

Pre-leaching residue

Slurrying

Anode area

HCl

Cathode area

Slurry electrolysis

Leaching Solution

Solid-liquid separation

Cathode antimony

Leaching residue Fig. 14. The flowsheet of SE for the extraction of antimony from the high arsenic and gold-containing stibnite ore. (For interpretation of the references to colour in this figure legend, the reader is referred to the web version of this article.)

Second, with the increase in anode current density, the bath voltage increases, which results in the increasing power consumption of the treating ore. We studied the influence of anode current density in the range 50–160 A/m2 on the antimony leaching efficiency as well as the bath voltage and power consumption. Fig. 9 and Fig. 10 show that the current density had little impact on the antimony extraction in the range of current density studied, although it appreciably affected the bath voltage and power consumption while treating the As-Au-Sb ore. The bath voltage and power consumption increased from 0.917 V and 230 kWh/ t-ore to 1.712 V and 429 kWh/t-ore respectively, in the anode current density range from 50 to 160 A/m2. In practical applications, the power consumption and production efficiency should be considered for determining the appropriate anode current density. In this study, the anode current density was determined to be 80 A/m2. 3.1.7. Effect of initial antimony concentration The influence of initial antimony concentration in the electrolyte on the antimony extraction from the As-Au-Sb ore was investigated under the optimized conditions. The experiments were performed at initial antimony concentrations ranging from 0 to 30 g/L. The results (Fig. 11) showed that the first few initial antimony concentrations almost had no influence on the antimony extraction but affected the morphology of cathode metal antimony. When the concentration of antimony was < 20 g/L, the loose, spongy, powdery antimony that did not adhere to the cathode plate was electrodeposited. By contrast, the dense metal antimony was electrodeposited when the concentration was above 20 g/L. Consistent with this experiment, the study by Wang (2002) showed that dense metal antimony could be obtained with 20–60 g/L antimony. Therefore, for further experiments, the initial electrolyte was free of antimony. In the comprehensive test below, the initial antimony concentration for producing dense cathode metal antimony was 30 g/L.

4. Conclusion Mineralogical characterization of the As-Au-Sb ore showed that the major minerals were stibnite (Sb2S3), pyrite (FeS2), arsenopyrite (FeAsS), and quartz (SiO2), and the minor components included muscovite (KAl2(AlSi3O10)(OH)2) and other gangue minerals. Antimony was mainly present in stibnite. In this study, the basic SE parameters such as energy consumption; duration; temperature; initial concentrations of HCl, Fe, and Sb; and anodic current density were investigated. The results showed that Sb extraction was obviously influenced by energy consumption, duration, temperature, and the initial concentrations of HCl and Fe. An energy consumption of 1.1 times the benchmark value, a duration of 6 h, temperature of 60 °C, 30 g/L HCl, 5 g/L Fe, 30 g/L antimony, and anodic current density of 80 A/m2 were found to be the optimum conditions for SE. Under these conditions, the antimony extraction reached over 98%; the average extraction of Fe, As, and Au were about 5.5%, 1.1%, and 0.6%, respectively. Gold and aresenic remained in the leaching residue, thereby facilitating their separation from antimony.

3.2. Comprehensive SE experiments Three comprehensive tests were carried out using the optimized conditions, namely, energy consumption: 1.1 times the benchmark value, duration: 6 h, temperature: 60 °C, HCl: 30 g/L, Fe: 5 g/L, 290

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The antimony in cathode metal antimony was > 98% and As, Fe, and Au were less than or equal to 0.21%, 0.03%, and 0.14%, respectively. In the SE treatment of the As-Au-Sb ore, antimony and arsenic were effectively separated to produce cathode metal antimony in a single step. The sulfur present in stibnite was transformed into elemental sulfur remaining in the residue. Thus, the contamination of arsenic soda slag and low-concentration SO2 were avoided. The removal of impurities accumulated in the electrolyte and the extraction of gold in the SE leaching residue will be studied in the future.

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