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The solvent extraction separation of bismuth and molybdenum from a low grade bismuth glance flotation concentrate
Hydrometallurgy 96 (2009) 342–348
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Hydrometallurgy j o u r n a l h o m e p a g e : w w w. e l s e v i e r. c o m / l o c a t e / h y d r o m e t
The solvent extraction separation of bismuth and molybdenum from a low grade bismuth glance flotation concentrate Jian-guang Yang a,b,⁎, Jian-ying Yang c, Mo-tang Tang a, Chao-bo Tang a, Wei Liu a a b c
Department of Metallurgical Science and Engineering, Central South University, 410083, China Institute of Powder Metallurgy Research, Central South University, 410083, China Environmental Engineering Vocational College, JiangXi 342000, China
a r t i c l e
i n f o
Article history: Received 24 September 2008 Received in revised form 9 December 2008 Accepted 19 December 2008 Available online 31 December 2008 Keywords: Bismuth Molybdenum Solvent extraction Extractive metallurgy
a b s t r a c t The main purpose of this study was to characterize and to extract bismuth and molybdenum from a low grade bismuth glance concentrate. Selective leaching of bismuth could be obtained at a temperature range 60 to 85 °C for a leaching duration of 2 h with hydrochloric acid concentration of 150 gpl, lignin calcium concentration of 0.02 gpl and using a solid–liquid ratio 1/4 g/cc. Treatment of leach liquor for the solvent extraction of bismuth with N235 showed that 8.0 × 10− 2 M N235 in kerosene, a 3 min period of equilibration and a pH 0.2 were sufficient for the extraction of Bi(III). This bismuth-loaded organic phase was almost completely stripped using 0.5 M EDTA solution. Treatment of leached residue was dealt with by roasting in the presence of slaked lime, and followed by hydrometallurgical treatment of the roasted products. In the lime roasting process, molybdenum recoveries of around 99% were achieved when an excess of 50% lime over stoichiometric requirement was roasted at 700 °C for 2 h and the calcine was leached with 4 M HCl, at 70–80 °C for 2 h. Molybdenum then was effectively extracted from the leached residual solution with N235. An optimum pH of 0.5 was determined for molybdenum extraction. From loaded solvent, this metal was easily stripped with ammonia solutions to give a pregnant solution suitable for final recovery of metal by salt precipitation. Under the optimized conditions, the ultimate recovery rate of bismuth and molybdenum was more than 99% and 98% respectively. © 2008 Elsevier B.V. All rights reserved.
1. Introduction Very little bismuth and molybdenum come from the processing of minerals where bismuth and molybdenum are the main metals. Most bismuth and molybdenum are obtained as the sub-products in other metal metallurgical processes. During the processing of these minerals, leaching with H2SO4 or HCl is always involved, and highly acidic solutions with base metals and bismuth are obtained. Generally, bismuth can be recovered from these solutions by hydrolyzation from spent solution, and few authors have reported the separation of bismuth from leaching mineral solutions by solvent extraction method. In the literature (Reyes-Aguilera et al., 2008), a number of other methods have been mentioned about bismuth recovery. Szymanowski (Fisher et al., 1975); (Szymanowski, 1998), Kim (Szymanowski, 1996), Dreinsiger (Kim et al., 1998); (Dreinsiger et al., 1993a) and Wang (Dreinsiger et al., 1993b) have proposed the separation of bismuth from copper through ionic exchange and solvating extractants, using the Acorga SBX-50 extractant in chloride media and organic phosphorus extractants, respectively. Velea and co-workers (Wang et al., ⁎ Corresponding author. Department of Metallurgical Science and Engineering, Central South University, 410083, China. E-mail address: [email protected] (J.-g. Yang). 0304-386X/$ – see front matter © 2008 Elsevier B.V. All rights reserved. doi:10.1016/j.hydromet.2008.12.006
2002) have studied the extraction of As, Bi and Sb by LIX 1104 SM, Cyanex 923 and Acorga SBX-50, from H2SO4 in the presence of chlorides, and best results are obtained when using Cyanex 923 as extractant. Other authors (Cox et al., 2002; Iyer and Dhadke, 2003; Ali and Vanjara, 2001; Moriya et al., 2001) have proposed the extraction of Bi(III) from acid or highly acidic solutions of HCl, HBr, HNO3 and/or H2SO4 using Cyanex 925, Cyanex 921, 2-bromoalkanoic acid and Cyanex 302 as extractant, and Reyes-Aguilera et al. (2008) had ever proposed the idea about supported liquid membranes (SLM) for recovery of bismuth from aqueous solutions in literature. As for the molybdenum, it is produced mainly from its high grade sulfide ore, the commercial route for the extraction of molybdenum from its sulfide mineral molybdenite involves roasting of the concentrate, purification of the resultant calcine, either by distillation of MoO3, or by a hydrometallurgical route, and, finally, hydrogen reduction of the trioxide to the met al. During roasting, much generated SO2 is a source of environmental pollution. Further, such a pyrochemical process requires a high-grade concentrate and is not suitable for the treatment of low grade molybdenite because it would result in the formation of stable molybdates of impurity elements which render subsequent processing difficult. Processing of low grade molybdenite concentrates for the extraction of molybdenum without entailing SO2 emission has, therefore, received considerable attention
J.-g. Yang et al. / Hydrometallurgy 96 (2009) 342–348 Table 1 Chemical composition of ShiZhuYuan bismuth glance floatation concentrate (%) Bi
Mo
Fe
Cu
S
SiO2
6.52
5.17
28.26
1.77
37.56
12.10
in recent years. Different processing routes have been studied, for instance, roasting with lime or soda ash followed by sulphuric acid leaching and selective adsorption of molybdenum with activated charcoal (Juneja et al., 1996). Other processing methods for different types of molybdenum-containing materials are well introduced by Gupta (1992), and literature (Yun et al., 2006) introduces the investigations on the extraction of molybdenum and vanadium from ammonia leaching residue of spent catalyst. In the present study, a process for separation and recovery of bismuth and molybdenum from a low grade bismuth glance flotation concentrate by solvent extraction was studied. N235 (R3N, R = C8–C10, trialkyl amine, a widely-used amic extractant for extracting Zn (Qiong et al., 2006), rare earth (Chunhua et al., 2006), molybdenum and vanadium (Yun et al., 2006), commercialized in China) was used as extraction agent. Under the optimized conditions, the ultimate recovery ratios of bismuth and molybdenum were more than 99% and 98% respectively. The study proposed in this paper at laboratory-scale revealed that the recovery of bismuth and molybdenum from the low grade bismuth glance can be carried out on an industrial scale by applying a continuous hydrometallurgical process, which would produce an enriched bismuth and molybdenum solution for bismuth and ammonium paramolybdate manufacture. 2. Experimental 2.1. Materials The low grade bismuth glance flotation concentrate obtained from ShiZhuYuan Rare Metal Co., China, in the ground form was characterized physically, chemically and mineralogically. The dry screen analysis of this bismuth glance shows that almost all concentrate samples were around 100 mesh (147 µm) and all samples were
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leached in this form without further grinding. A representative sample of the bismuth glance was taken and ground before chemical analysis. Most of the chemical analysis of the bismuth glance was done by AAS except SiO2 which was determined by gravimetric analysis. The results of chemical analysis of this bismuth glance are given in Table 1. For the mineralogical characterization of the bismuth glance X-ray diffraction analysis and optical microscopy investigations were undertaken. The X-ray diffraction analysis was performed by using the Regaku Geigerflex X-ray difractometer. The X-ray diffraction pattern obtained indicated the presence of bismuth sulfide (Bi2S3), a small quantity of bismuth oxide (Bi2O3), molybdenum sulfide (MoS2), a small quantity of molybdenum oxide (MoO3). The mineralogical analysis using the Leitz-Orthoplan optical microscope indicated the presence of small amounts of quartz (SiO2), limonite (2Fe2O3·3H2O), copper sulfide (Cu2S), pyrite (FeS2), wustite (FeO), etc. 2.2. Method The main purpose of this study was to characterize and to extract bismuth and molybdenum from a low grade bismuth glance concentrate. A procedure for separation and recovery of metals from this low grade bismuth glance flotation concentrate by solvent extraction was studied (Fig. 1). The process for the recovery of metals was carried out in five stages: hydrochloric acid leaching; deoxidization; solvent extraction and stripping of bismuth; lime roasting; solvent extraction and stripping of molybdenum. In the leaching experiments a hot plate with contact thermometer and a magnetic stirrer were used. The leaching was done in 250 cc glass balloons using Teflon coated magnets. A condenser attached to the balloon prevented the vaporization losses and with the aid of the contact thermometer, the temperature of the system was controlled within 2 °C. At the end of each leaching experiment, the insoluble leach residue was separated from the pregnant leach solution by passing it through a filter paper and the solids were washed with distilled water. In the leaching tests the procedure followed was as follows: A measured amount of distilled water and calculated amount of reagent grade hydrochloric acid were put into the glass balloon and heated to the
Fig. 1. Schematic diagram of the leach-solvent extraction process.
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desired temperature while magnetically stirring. Then, the weighted amount of bismuth glance floatation concentrate powder, sodium chlorate and lignin calcium was added to the dilute hydrochloric acid solution and the timing was initiated. The stirring speed (250 r/min) which gave sufficient mixing was kept constant in all the experiments. At the end of leaching duration, solid–liquid separation by filtration followed as explained above. The leach recovery of bismuth was calculated using the analysis of the bismuth glance flotation concentrate and the leach residue and its mass. Thereinto, sodium chlorate was used as oxidization agent and lignin calcium was used to selectively enwrap the other sulfide excepted bismuth sulfide. Bi2 S3 þ NaClO3 þ 6HCl ¼ 2BiCl3 þ NaCl þ 3H2 O þ 3S In the deoxidization tests the procedure followed was as follows: The measured leached residual solution was poured into a 250 cc glass balloons using Teflon coated magnets to stir it. Then, the weighted amount of bismuth powder was added to the solution and the timing was initiated. The stirring speed which gave sufficient mixing was kept constant in all the experiments (250 r/min). At the end of deoxidization duration, solid–liquid separation by filtration followed as explained above. The deoxidization degree was calculated using the analysis of the Fe3+, Cu2+ and their concentrations. 3FeCl3 þ Bi ¼ 3FeCl2 þ BiCl3 3CuCl2 þ 2Bi ¼ 2BiCl3 þ 3Cu In the solvent extraction tests the procedure followed was as follows: a measured amount of purified solution and extraction agents was put into the separatory funnel and shook rapidly. The shaking speed which gave sufficient mixing was kept constant in all the experiments (120 r/min). At the end of solvent extraction duration, organic–aqueous phases were separated by separatory funnel. The extraction rate of bismuth was calculated by measuring the volume of the purified solution and the extraction residual solution and metals concentration. R3N þ Hþ þ Cl− ¼ ½R3NHCl − 3½R3NHCl þ BiCl3− 6 ¼ ½R3NH3 BiCl6 þ 3Cl
In the stripping tests, bismuth or molybdenum was stripped by ethylenediaminetetraacetic acid (EDTA) or ammonia respectively to give an enriched aqueous phase: a measured amount of pregnant organic solution and EDTA or ammonia solution was put into the separatory funnel and shook rapidly. The shaking speed which gave sufficient mixing was kept constant in all the experiments. At the end of stripping duration, organic–aqueous phases separation by that separatory funnel. The stripping rate of bismuth or molybdenum was calculated using the measure the volume of the pregnant organic solution and the pregnant solution and metals concentration. 3. Results and discussions 3.1. Leaching results and discussions ShiZhuYuan bismuth glance floatation concentrate samples were leached with hydrochloric acid solution. The samples used in the leaching experiments were previously ground to −100 mesh (147 µm).
Fig. 2. The percentage extractions for each metal ion as a function of the equilibrium pH.
Throughout the leaching experiments duration of leaching, acid concentration, temperature of leaching, solid–liquid ratio and effect of addition of chemical oxidants have been examined in order to determine the technical optimum conditions. Stirring speed was kept constant in each test. The optimum leaching conditions of bismuth were obtained at a temperature range 60 to 85 °C for a leaching duration of 2 h with hydrochloric acid concentration of 150 gpl, lignin calcium concentration of 0.02 gpl and using a solid–liquid ratio 1/4 g/cc. Under these optimized conditions, the leaching experiment results showed in Table 2. As given in Table 2, the leach recovery of bismuth was substantially higher while the other metals were substantially low. The leaching rate of bismuth was 99.7% while hardly any other metals were leached. This leaching procedure yielded a solid phase constitute mainly of 5.87% MoS2, 33.31% FeS2, 17.87% SiO2, and a residual solution constituted mainly of 9.78 gpl Bi3+, 0.81 gpl Fe2+, 0.2 gpl Fe3+, 0.12 gpl Cu2+. In this procedure, lignin calcium was used as enwrapping agent, which can be enwrapped on the surface of copper sulfide (Cu2S), pyrite (FeS2), wustite (FeO), etc to prevent more sulfide be leached. However, the mechanism of this enwrapping agent in this process has not been investigated in detail. The extraction characteristics of individual metal ions were examined by conducting a series of shake-out tests with equal volumes of organic solvent of 8.0 × 10− 2 M N235 (R3N, R = C8–C10, trialkyl amine, a widely-used amic extractant, commercialized in China) dissolved in kerosene and the aqueous hydrochloric acid solutions containing 0.3 g/dm3 of each metal at the various pH values (pH 0 to pH 2.0) at 313 K. The percentage extractions for each metal ion involved were shown in Fig. 2 as a function of the equilibrium pH. It was found that the extraction ability for bismuth (III) was nearly the same as Fe(III) and Mo(VI) at pH b 0.8, whereas the other metal ions, such as Fe(II), Cu(II), were hardly extracted by N235 under these conditions. An increase in pH value from 0.6 to 2.0 weakened the extraction of bismuth (III), Fe(III), and Cu(II), Fe(II) irons were still hardly extracted. Such results indicate that it should be possible to deoxidize Fe(III) into Fe(II) before bismuth extraction in order to avoid collective extraction of Bi(III) and Fe(III). Therefore, the following deoxidization procedure was introduced. 3.2. Deoxidization results and discussions
Table 2 Leaching experiment results under optimized conditions 3+
Elements in leached residual solution
Bi
Concentration (gpl) Leaching ratio (%)
9.78 99.7
6+
2+
Mo
Cu
Trace –
0.12 10.2
2+
3+
Fe
Fe
0.81 1.21
0.2 2.10
The major aim of this study was production of pre-solvent extraction solution in pure form as possible. In the deoxidization tests the procedure was followed as follows: The measured leached residual solution was poured into a 250 cc glass balloon using Teflon coated magnets to stir it. Then, the weighted
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Fig. 3. Effect of duration of deoxidization on the Fe3+ and Cu2+ concentration in deoxidized solution.
amount of bismuth powder was added to the solution and the timing was initiated. The stirring speed which gave sufficient mixing was kept constant in all the experiments. At the end of deoxidization duration, solid–liquid separation by filtration followed as explained above. The deoxidization degree was calculated using the analysis of the Fe3+, Cu2+ and their concentrations. 3.2.1. Effect of duration of purification on the Fe3+ and Cu2+ concentration in purified solution The experimental durations were varied between 10 min and 60 min. As shown in Fig. 3, about half an hour duration of purification was sufficient for the deoxidization of Cu2+, whereas the total deoxidization of Fe3+ to Fe2+ needed 1 h. Therefore, if collective deoxidization (deoxidization of Fe3+ to Fe2+ and Cu2+ to Cu) was taken into consideration, 1 h duration of deoxidization would be sufficient for the optimum deoxidization duration. 3.2.2. Effect of mass of bismuth powder addition on the Fe3+ and Cu2+ concentration in purified solution Experiments of the this group were made to examine the effect of mass of bismuth powder addition on the Fe3+ and Cu2+ concentration in deoxidized solution. These experiments were carried out at bismuth powder addition changing from 1.0 to 1.5 multiple of stoichiometric quantum. The results obtained were given in Fig. 3. By looking at Fig. 4, it was decided that the addition of 1.4 multiple of stoichiometric quantum of bismuth powder could be the best choice
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Fig. 5. Extraction rate of Bi(III) as a function of N235 concentration.
available because this addition was the best result with more than 99.0% Cu2+ deoxidization and more than 99.0% Fe3+ be reduced into Fe2+. 3.2.3. Bismuth recovery 3.2.3.1. Extraction as a function of N235 concentration. In order to optimize the conditions for extraction of Bi(III), kerosene solutions of N235 with varying molar concentrations were employed and TBP (1% v/v) was added as phase modifier. It was found that 8.0 × 10− 2 M N235 was sufficient for quantitative extraction of Bi(III), and further increase in the ratio of N235 was found to be unnecessary (Fig. 5). 3.2.3.2. Extraction with various diluents. Bi(III) was extracted with 8.0 × 10− 2 M N235 in various solvents. The extraction of Bi(III) was quantitative with kerosene, toluene, xylene, cyclohexane and benzene, whereas toluene, hexane, xylene, cyclohexane and benzene, chloroform, carbon tetrachloride and dichloromethane were found to be less efficient than kerosene as the diluent, and only kerosene shows a practically quantitative extraction. Hence, kerosene was preferred as the suitable diluent in this study (Table 3). 3.2.3.3. Effect of time of equilibrium. The solutions were shaken for a varying period ranging from 1.0 to 15.0 min. The extraction was quantitative over the entire period range after 3 min. Therefore a 3 min period of equilibration was used for the extraction of Bi(III). 3.2.3.4. Effect of pH value. The solutions were shaken for a varying pH values for 3 min, using hydrochloric acid and ammonia solution to adjust the pH value. It was determined that the extraction was highly efficient around pH 0.2. Therefore pH 0.2 was used for the extraction of Bi(III) (Fig. 6).
Table 3 Extraction for various diluents
Fig. 4. Effect of mass of bismuth powder addition on the Fe3+ and Cu2+ concentration in deoxidized solution.
Diluent
E (%)
Kerosene Toluene Xylene Cyclohexane Benzene Hexane Dichloromethane Carbon tetrachloride Chloroform
99.2 97.4 89.2 51.1 91.8 89.7 90.2 89.2 92.4
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Fig. 6. Extraction rate of Bi(III) as a function of pH value.
Fig. 7. Influence of roasting temperature on the extraction of molybdenum values from molybdenite concentrate by roasting in the presence of lime.
3.2.3.5. Effect of various stripping agents. Bi(III) was stripped with different strengths of ligands after their extraction. The stripping of Bi (III) was quantitative with 0.1 M to 1.0 M EDTA, DTPA, EGTA, DCTA. It is found that only 0.5 M EDTA showed the maximal stripping ratio of Bi (III). Hence, for all further investigations, 0.5 M EDTA was used as stripping agent for Bi(III) (Table 4). Under the optimum condition, a pregnant aqueous solution was obtained (Table 5). Based on the above mentioned investigation, it is clear that from this purified solution, bismuth could be extracted with N235, and it was determined that the extraction was highly efficient around pH 0.2 and no co-extraction of other metals was observed at the pH of bismuth recovery. This metal was stripped by ethylenediaminetetraacetic acid (EDTA) to give an enriched aqueous phase containing 9.93 gpl, and this pregnant solution was suitable for final recovery of bismuth by electrolysis, replacement or salt precipitation. 3.3. Molybdenum recovery The commercial route for the extraction of molybdenum from its sulfide mineral molybdenite involves roasting of the concentrate, purification of the resultant calcine, either by distillation of MoO3, or by a hydrometallurgical route, and, finally, hydrogen reduction of the trioxide to the metal. During roasting, much valuable metals are lost due to volatilization and the SO2 generated is a source of environmental pollution. Further, such a pyrochemical process requires a high-grade concentrate and is not suitable for the treatment of low grade molybdenite because it would result in the formation of stable molybdates of impurity elements which render subsequent processing difficult. Processing of low grade molybdenite concentrates for the extraction of molybdenum without entailing SO2, emission has, therefore,
received considerable attention in recent years. The processes reported are based on selective extraction of either molybdenum or the impurity elements. Selective extraction of MoO3 from low grade concentrate has been carried out by leaching with sodium hypochlorite solution, nitric acid, sodium hydroxide or by chlorination in the presence of oxygen. Leaching with ferric chloride or cyanidecontaining solutions has been used for preferential removal of impurity elements from the low grade concentrate. As far as the application of the roasting process to low grade concentrate is concerned, the process can be modified if roasting is carried out in the presence of lime. Roasting of molybdenite in the presence of soda ash has been reported for extraction of molybdenum by Wlodyka and Mehra et al. However, the process conditions reported by these investigators differ widely. A patent dealing with roasting of molybdenite in the presence of lime is available but complete process details are not known. The feasibility of applying the modified roasting processes to low grade molybdenite for extraction of MoO3 and Re values has been investigated by Juneja et al. (1996). The processes studied involve roasting of molybdenite in the presence of lime and of soda ash, and examines results of both processes for the extraction of MoO3 and Re values by a carbon adsorption technique involving selective adsorption of Mo on activated charcoal from the leach solution followed by desorption with ammonia. In this study, the bismuth-leached residue was treated with by roasting in the presence of slaked lime, and followed by hydrometallurgical treatment of the roasted products. The calcine was
Table 4 Effect of stripping agents for Bi(III) Stripping agent
Bi(III) recovery (%) at different concentration (M) 0.1
0.2
0.3
0.5
0.7
1.0
EDTA DPTA EGTA DCTA
69.2 32.1 18.2 23.1
86.5 34.0 – –
94.9 38.1 – 34.7
99.1 40.1 21.0 –
99.0 44.0 – –
– 47.5 32.2 40.1
Table 5 Chemical component of stripped solution Elements
Bi3+
Fe3+
Fe2+
Cu2+
Concentration/gpl
9.93
Trace
Trace
Trace
Fig. 8. Influence of charge composition on the extraction of molybdenum values from molybdenite concentrate by roasting in the presence of lime.
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Table 6 Results of leaching calcines Mode of leaching
Temperature of leaching (°C)
Period of leaching (h)
Molybdenum recovery (%)
Water 1 M HCl 2 M H2SO4 3 M HCl 4 M HCl 6 M HCl
80–90 80–90 80–90 80–90 80–90 80–90
3 3 3 3 3 3
0.3 89.2 56.3 94.8 99.1 99.2
Table 7 Main chemical component of the leached residual solution Elements
Mo6+
Fe3+
Fe2+
Bi3+
Concentration/gpl
5.98
10.88
0.01
0.02
leached with HCl and molybdenum can then be extracted from the leached residual solution with N235. From loaded solvent, this metal was stripped with ammonia solutions to give a pregnant solution suitable for final recovery of molybdenum by salt precipitation. 3.3.1. Lime roasting process The roasting of bismuth-leached residue, molybdenite, in the presence of lime/slaked lime (Ca(OH)2) results in the oxidation of molybdenum into calcium molybdate in accordance with: 2MoS2 þ 6CaðOHÞ2 þ 9O2 ¼ 2CaMoO4 þ 4CaSO4 þ 6H2 O The results of the influence of the temperature of roasting on the recovery of molybdenum were presented in Fig. 7. The molybdenum recoveries improved from 42.1% to 98.7% as the roasting temperature was increased from 400 °C to 800 °C. Further increase in the temperature of roasting was found to be unnecessary. Studies on the effect of charge composition indicated that the molybdenum yield increased from 36.4% to 98.9% as excess of lime over stoichiometric requirement was raised from 0%to 50%, and further increase in the excess ratio was found to be unnecessary (Fig. 8). Thus, a maximum recovery of 98.9% was obtained when an excess of 50% lime over stoichiometry requirement was roasted at 700 °C for 2 h. Results on the influence of roasting time indicated that 2 h of roasting was adequate for the roasting operation at 700 °C.
Fig. 10. Extraction rates of Mo(VI) and Fe(III) as a function of pH value.
calcined mass. The results of leaching studies were given in Table 6. It can be observed from the table that it was possible to extract more than 99% Mo in the leaching stage. Table 7 showed the main chemical components of the leached residual solution. 3.3.3. Solvent extraction of molybdenum The residual solution generated in the acid leaching step contains variable but significant amounts of molybdenum (VI) and iron (III) as shown in Table 7. In order to recover its contents of molybdenum and iron, this solution was treated by means of solvent extraction by contacting with an organic solvent formed by solutions of N235 in kerosene. TBP (1% v/v) was added as phase modifier in all experiments. Two variables were investigated for studying the extraction of molybdenum: the pH of aqueous solution and the concentration of N235 in the organic phase. The results are shown in Figs. 8 and 9. 3.3.3.1. Extraction as a function of N235 concentration. In order to optimize the conditions for extraction of Mo(IV) and Fe(III), kerosene solutions of N235 with varying molar concentrations were employed. It was found that 15 × 10− 2 M N235 was sufficient for quantitative extraction of Mo(IV) and Fe(III) (Fig. 9).
3.3.2. Solvent leaching of molybdenum Investigations on the leaching of the calcine with HCl indicated that a leaching with 4 M HCl at 80–90 °C for 3 h with constant stirring was necessary to extract the maximum amount of molybdenum from the
3.3.3.2. Effect of pH value. The solutions were shaken for varying pH values for 3 min, using hydrochloric acid and ammonia solution to adjust the pH value. It was determined that the extraction is highly efficient around pH 0.4, Therefore pH 0.4 was used for the extraction of Mo(IV) and Fe(III) (Fig. 10).
Fig. 9. Extraction rate E as a function of N235 concentration.
Fig. 11. Stripped rate of Fe(III) as a function of pH value.
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then optimized for selective and collective recovery and the optimum conditions for bismuth and molybdenum were determined. (2) Solvent extraction of bismuth and molybdenum by N235 from leach liquors mentioned in the processing of ShiZhuYuan bismuth glance flotation concentrate was studied. Under the optimized conditions, the ultimate recovery ratios of bismuth and molybdenum were more than 99% and 98% respectively. Acknowledgements The authors would like to express their gratitude to the Department of Science and Technology of China for supplying financial support, and ShiZhuYuan Rare Metal Co., China for supplying the material as well as the experiment equipments for this study. References Fig. 12. Stripped rate of Mo(VI) as a function of ammonia volume percentage.
From Figs. 9 and 10 the following information was observed: at pH 0.4, over 99% of Mo(VI) and over 98% of Fe(III) were extracted. Both molybdenum (VI) and Fe(III) were the metal values to be recovered. In order to obtain purified products, the two metals loaded in the organic phase must be further separated. Thus a stripping treatment was employed to unload the Fe(III) co-extracted with molybdenum (VI). According to the data shown in Fig. 2, the extraction of Fe(III) depends greatly on the equilibrium acidity, whereas that of molybdenum (VI) is essentially unaffected by acidity when pH value lies in the range of pH 0 to pH 2.0. It was, therefore, to be expected that the Fe(III) in the loaded solvent can be unloaded by shifting the equilibrium acidity by using hydrochloric acid solution as a scrubbing reagent. The results, given in Fig. 11, showed that the stripping ratio of Fe(III) was markedly increased with the pH value increase from pH 0.5 to pH 1.5, and more than 99% of the Fe(III) can be stripped into the aqueous phase. It was expected that molybdenum (VI) can be stripped by contacting the pregnant organic phase with an alkaline solution such as ammonia or ammonium carbonate. The strip tests were carried out with a variety of concentrations of ammonia solutions to examine the stripping efficiency for molybdenum (VI) in this work. The results were presented in Fig. 12 and showed that the complete stripping of molybdenum (VI) took place when the content of the ammonia solution was higher than 10 vol.%. Moreover, good phase separation was observed in all cases. After stripping, N235 is recycled to the extraction by regeneration with a 2 M hydrochloric acid solution. 4. Conclusions (1) A comprehensive procedure was introduced to separation and collective extraction of bismuth and molybdenum from low grade bismuth glance concentrate. All detailed processes, include: the hydrochloric acid leaching, purification and deoxidization, solvent extraction of bismuth, lime roasting process, leaching and solvent extraction of molybdenum, were
Ali, K.A., Vanjara, A.K., 2001. Solvent extraction and separation of Bi(III) and Sb(III) from HCl and HBr media using tri-n-octyl phosphine oxide (TOPO). Ind. J. Chem. Technol. 8, 239–243. Chunhua, Y., Jiangtao, J., Chunsheng, L., Sheng, W., Guangxian, X., 2006. Rare earth separation in China. Tsinghua Sci. Technol. 11, 241–247. Cox, M., Flett, D.S., Velea, T., Vasiliu, C., 2002. Impurity Removal from Copper Tankhouse Liquors by Solvent Extraction. International Solvent Extraction Conference, Cape Town, pp. 995–1000. Dreinsiger, D.B., Leong, B.J.Y., Saito, B.R., West-Sells, P.J., 1993a. Hydrometallurgy Fundamentals. In: Hiskey, J.B., Warren, G.W. (Eds.), Technology and Innovation. SME, Litleton, Colorado, pp. 801–815. Dreinsiger, D.B., Leong, B.J.Y., Bailint, B.J., Beyad, M.H., 1993b. The solvent extraction of As, Sb and Bi from copper refining electrolytes using organophosphorus reagents. Proceedings of International Solvent Extraction Conference, pp. 1271–1278. Fisher, B.M., Hills, R.C., Touro, F.J., 1975. Process for the manufacture of electrolytic copper. US Patent 3,917,519. Gupta, C.K., 1992. Extractive Metallurgy of Molybdenum. CRC Press, Boca Raton, USA. Iyer, J.N., Dhadke, P.M., 2003. Solvent extraction and separation studies of antimony (III) and bismuth(III) by using Cyanex-925. Ind. J. Chem. Technol. 10, 665–669. Juneja, J.M., Singh, S., Bose, D.K., 1996. Investigations on the extraction of molybdenum and rhenium values from low grade molybdenite concentrate. Hydrometallurgy 41, 201–209. Kim, D.K., Leese, T.A., Neild, M.N., Saito, B.R., Young, S.K., Weidner, C.J., 1998. Use of solvent extraction to remove bismuth and antimony from copper electrolyte at the San Manuel refinery; EPD Congress 1998. Proceedings of Sessions and Symposia held at the TMS Annual Meeting, San Antonio, February 16–19, pp. 301–315. Moriya, Y., Sugai, M., Nakata, S., Ogawa, N., 2001. Extraction of bismuth (III) with 2-bromoalkanoic acid non-donating solvent from highly acidic aqueous solution. Anal. Sci. 17, 297–300. Qiong, J., Jie, W., Ting-Ting, L., Wei, H.Z.H., 2006. Synergistic extraction of Zn(II) by mixtures of tri-butyl-phosphate and trialkyl amine extractant. Chin. J. Anal. Chem. 36, 619–622. Reyes-Aguilera, J.A., Gonzale, z, M.P., Navarro, R., Saucedo, T.I., Avila-Rodriguez, M., 2008. Supported liquid membranes (SLM) for recovery of bismuth from aqueous solutions. J. Membr. Sci. 310, 13–19. Szymanowski, J., 1996. Industrial processes for copper extraction. Part II Removal of arsenic, antimony and bismuth from technological sulfuric acid solutions. Pol. J. Appl. Chem. 40, 3–15. Szymanowski, J., 1998. Removal of toxic elements from copper electrolyte by solvent extraction. Miner. Process. Extr. Metall. Rev. 18, 389–418. Wang, C., Jiang, K., Liu, D., Wang, H., 2002. Purification of Copper Electrolyte with Cyanex 923. International Solvent Extraction Conference, Cape Town, pp. 1039–1044. Yun, Ch., Qiming, F., Yanhai, Sh., Guofan, Zh., Leming, O., Yiping, L., 2006. Investigations on the extraction of molybdenum and vanadium from ammonia leaching residue of spent catalyst. Inter. J. Mineral Processing 79, 42–48.
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Hydrometallurgy j o u r n a l h o m e p a g e : w w w. e l s e v i e r. c o m / l o c a t e / h y d r o m e t
The solvent extraction separation of bismuth and molybdenum from a low grade bismuth glance flotation concentrate Jian-guang Yang a,b,⁎, Jian-ying Yang c, Mo-tang Tang a, Chao-bo Tang a, Wei Liu a a b c
Department of Metallurgical Science and Engineering, Central South University, 410083, China Institute of Powder Metallurgy Research, Central South University, 410083, China Environmental Engineering Vocational College, JiangXi 342000, China
a r t i c l e
i n f o
Article history: Received 24 September 2008 Received in revised form 9 December 2008 Accepted 19 December 2008 Available online 31 December 2008 Keywords: Bismuth Molybdenum Solvent extraction Extractive metallurgy
a b s t r a c t The main purpose of this study was to characterize and to extract bismuth and molybdenum from a low grade bismuth glance concentrate. Selective leaching of bismuth could be obtained at a temperature range 60 to 85 °C for a leaching duration of 2 h with hydrochloric acid concentration of 150 gpl, lignin calcium concentration of 0.02 gpl and using a solid–liquid ratio 1/4 g/cc. Treatment of leach liquor for the solvent extraction of bismuth with N235 showed that 8.0 × 10− 2 M N235 in kerosene, a 3 min period of equilibration and a pH 0.2 were sufficient for the extraction of Bi(III). This bismuth-loaded organic phase was almost completely stripped using 0.5 M EDTA solution. Treatment of leached residue was dealt with by roasting in the presence of slaked lime, and followed by hydrometallurgical treatment of the roasted products. In the lime roasting process, molybdenum recoveries of around 99% were achieved when an excess of 50% lime over stoichiometric requirement was roasted at 700 °C for 2 h and the calcine was leached with 4 M HCl, at 70–80 °C for 2 h. Molybdenum then was effectively extracted from the leached residual solution with N235. An optimum pH of 0.5 was determined for molybdenum extraction. From loaded solvent, this metal was easily stripped with ammonia solutions to give a pregnant solution suitable for final recovery of metal by salt precipitation. Under the optimized conditions, the ultimate recovery rate of bismuth and molybdenum was more than 99% and 98% respectively. © 2008 Elsevier B.V. All rights reserved.
1. Introduction Very little bismuth and molybdenum come from the processing of minerals where bismuth and molybdenum are the main metals. Most bismuth and molybdenum are obtained as the sub-products in other metal metallurgical processes. During the processing of these minerals, leaching with H2SO4 or HCl is always involved, and highly acidic solutions with base metals and bismuth are obtained. Generally, bismuth can be recovered from these solutions by hydrolyzation from spent solution, and few authors have reported the separation of bismuth from leaching mineral solutions by solvent extraction method. In the literature (Reyes-Aguilera et al., 2008), a number of other methods have been mentioned about bismuth recovery. Szymanowski (Fisher et al., 1975); (Szymanowski, 1998), Kim (Szymanowski, 1996), Dreinsiger (Kim et al., 1998); (Dreinsiger et al., 1993a) and Wang (Dreinsiger et al., 1993b) have proposed the separation of bismuth from copper through ionic exchange and solvating extractants, using the Acorga SBX-50 extractant in chloride media and organic phosphorus extractants, respectively. Velea and co-workers (Wang et al., ⁎ Corresponding author. Department of Metallurgical Science and Engineering, Central South University, 410083, China. E-mail address: [email protected] (J.-g. Yang). 0304-386X/$ – see front matter © 2008 Elsevier B.V. All rights reserved. doi:10.1016/j.hydromet.2008.12.006
2002) have studied the extraction of As, Bi and Sb by LIX 1104 SM, Cyanex 923 and Acorga SBX-50, from H2SO4 in the presence of chlorides, and best results are obtained when using Cyanex 923 as extractant. Other authors (Cox et al., 2002; Iyer and Dhadke, 2003; Ali and Vanjara, 2001; Moriya et al., 2001) have proposed the extraction of Bi(III) from acid or highly acidic solutions of HCl, HBr, HNO3 and/or H2SO4 using Cyanex 925, Cyanex 921, 2-bromoalkanoic acid and Cyanex 302 as extractant, and Reyes-Aguilera et al. (2008) had ever proposed the idea about supported liquid membranes (SLM) for recovery of bismuth from aqueous solutions in literature. As for the molybdenum, it is produced mainly from its high grade sulfide ore, the commercial route for the extraction of molybdenum from its sulfide mineral molybdenite involves roasting of the concentrate, purification of the resultant calcine, either by distillation of MoO3, or by a hydrometallurgical route, and, finally, hydrogen reduction of the trioxide to the met al. During roasting, much generated SO2 is a source of environmental pollution. Further, such a pyrochemical process requires a high-grade concentrate and is not suitable for the treatment of low grade molybdenite because it would result in the formation of stable molybdates of impurity elements which render subsequent processing difficult. Processing of low grade molybdenite concentrates for the extraction of molybdenum without entailing SO2 emission has, therefore, received considerable attention
J.-g. Yang et al. / Hydrometallurgy 96 (2009) 342–348 Table 1 Chemical composition of ShiZhuYuan bismuth glance floatation concentrate (%) Bi
Mo
Fe
Cu
S
SiO2
6.52
5.17
28.26
1.77
37.56
12.10
in recent years. Different processing routes have been studied, for instance, roasting with lime or soda ash followed by sulphuric acid leaching and selective adsorption of molybdenum with activated charcoal (Juneja et al., 1996). Other processing methods for different types of molybdenum-containing materials are well introduced by Gupta (1992), and literature (Yun et al., 2006) introduces the investigations on the extraction of molybdenum and vanadium from ammonia leaching residue of spent catalyst. In the present study, a process for separation and recovery of bismuth and molybdenum from a low grade bismuth glance flotation concentrate by solvent extraction was studied. N235 (R3N, R = C8–C10, trialkyl amine, a widely-used amic extractant for extracting Zn (Qiong et al., 2006), rare earth (Chunhua et al., 2006), molybdenum and vanadium (Yun et al., 2006), commercialized in China) was used as extraction agent. Under the optimized conditions, the ultimate recovery ratios of bismuth and molybdenum were more than 99% and 98% respectively. The study proposed in this paper at laboratory-scale revealed that the recovery of bismuth and molybdenum from the low grade bismuth glance can be carried out on an industrial scale by applying a continuous hydrometallurgical process, which would produce an enriched bismuth and molybdenum solution for bismuth and ammonium paramolybdate manufacture. 2. Experimental 2.1. Materials The low grade bismuth glance flotation concentrate obtained from ShiZhuYuan Rare Metal Co., China, in the ground form was characterized physically, chemically and mineralogically. The dry screen analysis of this bismuth glance shows that almost all concentrate samples were around 100 mesh (147 µm) and all samples were
343
leached in this form without further grinding. A representative sample of the bismuth glance was taken and ground before chemical analysis. Most of the chemical analysis of the bismuth glance was done by AAS except SiO2 which was determined by gravimetric analysis. The results of chemical analysis of this bismuth glance are given in Table 1. For the mineralogical characterization of the bismuth glance X-ray diffraction analysis and optical microscopy investigations were undertaken. The X-ray diffraction analysis was performed by using the Regaku Geigerflex X-ray difractometer. The X-ray diffraction pattern obtained indicated the presence of bismuth sulfide (Bi2S3), a small quantity of bismuth oxide (Bi2O3), molybdenum sulfide (MoS2), a small quantity of molybdenum oxide (MoO3). The mineralogical analysis using the Leitz-Orthoplan optical microscope indicated the presence of small amounts of quartz (SiO2), limonite (2Fe2O3·3H2O), copper sulfide (Cu2S), pyrite (FeS2), wustite (FeO), etc. 2.2. Method The main purpose of this study was to characterize and to extract bismuth and molybdenum from a low grade bismuth glance concentrate. A procedure for separation and recovery of metals from this low grade bismuth glance flotation concentrate by solvent extraction was studied (Fig. 1). The process for the recovery of metals was carried out in five stages: hydrochloric acid leaching; deoxidization; solvent extraction and stripping of bismuth; lime roasting; solvent extraction and stripping of molybdenum. In the leaching experiments a hot plate with contact thermometer and a magnetic stirrer were used. The leaching was done in 250 cc glass balloons using Teflon coated magnets. A condenser attached to the balloon prevented the vaporization losses and with the aid of the contact thermometer, the temperature of the system was controlled within 2 °C. At the end of each leaching experiment, the insoluble leach residue was separated from the pregnant leach solution by passing it through a filter paper and the solids were washed with distilled water. In the leaching tests the procedure followed was as follows: A measured amount of distilled water and calculated amount of reagent grade hydrochloric acid were put into the glass balloon and heated to the
Fig. 1. Schematic diagram of the leach-solvent extraction process.
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desired temperature while magnetically stirring. Then, the weighted amount of bismuth glance floatation concentrate powder, sodium chlorate and lignin calcium was added to the dilute hydrochloric acid solution and the timing was initiated. The stirring speed (250 r/min) which gave sufficient mixing was kept constant in all the experiments. At the end of leaching duration, solid–liquid separation by filtration followed as explained above. The leach recovery of bismuth was calculated using the analysis of the bismuth glance flotation concentrate and the leach residue and its mass. Thereinto, sodium chlorate was used as oxidization agent and lignin calcium was used to selectively enwrap the other sulfide excepted bismuth sulfide. Bi2 S3 þ NaClO3 þ 6HCl ¼ 2BiCl3 þ NaCl þ 3H2 O þ 3S In the deoxidization tests the procedure followed was as follows: The measured leached residual solution was poured into a 250 cc glass balloons using Teflon coated magnets to stir it. Then, the weighted amount of bismuth powder was added to the solution and the timing was initiated. The stirring speed which gave sufficient mixing was kept constant in all the experiments (250 r/min). At the end of deoxidization duration, solid–liquid separation by filtration followed as explained above. The deoxidization degree was calculated using the analysis of the Fe3+, Cu2+ and their concentrations. 3FeCl3 þ Bi ¼ 3FeCl2 þ BiCl3 3CuCl2 þ 2Bi ¼ 2BiCl3 þ 3Cu In the solvent extraction tests the procedure followed was as follows: a measured amount of purified solution and extraction agents was put into the separatory funnel and shook rapidly. The shaking speed which gave sufficient mixing was kept constant in all the experiments (120 r/min). At the end of solvent extraction duration, organic–aqueous phases were separated by separatory funnel. The extraction rate of bismuth was calculated by measuring the volume of the purified solution and the extraction residual solution and metals concentration. R3N þ Hþ þ Cl− ¼ ½R3NHCl − 3½R3NHCl þ BiCl3− 6 ¼ ½R3NH3 BiCl6 þ 3Cl
In the stripping tests, bismuth or molybdenum was stripped by ethylenediaminetetraacetic acid (EDTA) or ammonia respectively to give an enriched aqueous phase: a measured amount of pregnant organic solution and EDTA or ammonia solution was put into the separatory funnel and shook rapidly. The shaking speed which gave sufficient mixing was kept constant in all the experiments. At the end of stripping duration, organic–aqueous phases separation by that separatory funnel. The stripping rate of bismuth or molybdenum was calculated using the measure the volume of the pregnant organic solution and the pregnant solution and metals concentration. 3. Results and discussions 3.1. Leaching results and discussions ShiZhuYuan bismuth glance floatation concentrate samples were leached with hydrochloric acid solution. The samples used in the leaching experiments were previously ground to −100 mesh (147 µm).
Fig. 2. The percentage extractions for each metal ion as a function of the equilibrium pH.
Throughout the leaching experiments duration of leaching, acid concentration, temperature of leaching, solid–liquid ratio and effect of addition of chemical oxidants have been examined in order to determine the technical optimum conditions. Stirring speed was kept constant in each test. The optimum leaching conditions of bismuth were obtained at a temperature range 60 to 85 °C for a leaching duration of 2 h with hydrochloric acid concentration of 150 gpl, lignin calcium concentration of 0.02 gpl and using a solid–liquid ratio 1/4 g/cc. Under these optimized conditions, the leaching experiment results showed in Table 2. As given in Table 2, the leach recovery of bismuth was substantially higher while the other metals were substantially low. The leaching rate of bismuth was 99.7% while hardly any other metals were leached. This leaching procedure yielded a solid phase constitute mainly of 5.87% MoS2, 33.31% FeS2, 17.87% SiO2, and a residual solution constituted mainly of 9.78 gpl Bi3+, 0.81 gpl Fe2+, 0.2 gpl Fe3+, 0.12 gpl Cu2+. In this procedure, lignin calcium was used as enwrapping agent, which can be enwrapped on the surface of copper sulfide (Cu2S), pyrite (FeS2), wustite (FeO), etc to prevent more sulfide be leached. However, the mechanism of this enwrapping agent in this process has not been investigated in detail. The extraction characteristics of individual metal ions were examined by conducting a series of shake-out tests with equal volumes of organic solvent of 8.0 × 10− 2 M N235 (R3N, R = C8–C10, trialkyl amine, a widely-used amic extractant, commercialized in China) dissolved in kerosene and the aqueous hydrochloric acid solutions containing 0.3 g/dm3 of each metal at the various pH values (pH 0 to pH 2.0) at 313 K. The percentage extractions for each metal ion involved were shown in Fig. 2 as a function of the equilibrium pH. It was found that the extraction ability for bismuth (III) was nearly the same as Fe(III) and Mo(VI) at pH b 0.8, whereas the other metal ions, such as Fe(II), Cu(II), were hardly extracted by N235 under these conditions. An increase in pH value from 0.6 to 2.0 weakened the extraction of bismuth (III), Fe(III), and Cu(II), Fe(II) irons were still hardly extracted. Such results indicate that it should be possible to deoxidize Fe(III) into Fe(II) before bismuth extraction in order to avoid collective extraction of Bi(III) and Fe(III). Therefore, the following deoxidization procedure was introduced. 3.2. Deoxidization results and discussions
Table 2 Leaching experiment results under optimized conditions 3+
Elements in leached residual solution
Bi
Concentration (gpl) Leaching ratio (%)
9.78 99.7
6+
2+
Mo
Cu
Trace –
0.12 10.2
2+
3+
Fe
Fe
0.81 1.21
0.2 2.10
The major aim of this study was production of pre-solvent extraction solution in pure form as possible. In the deoxidization tests the procedure was followed as follows: The measured leached residual solution was poured into a 250 cc glass balloon using Teflon coated magnets to stir it. Then, the weighted
J.-g. Yang et al. / Hydrometallurgy 96 (2009) 342–348
Fig. 3. Effect of duration of deoxidization on the Fe3+ and Cu2+ concentration in deoxidized solution.
amount of bismuth powder was added to the solution and the timing was initiated. The stirring speed which gave sufficient mixing was kept constant in all the experiments. At the end of deoxidization duration, solid–liquid separation by filtration followed as explained above. The deoxidization degree was calculated using the analysis of the Fe3+, Cu2+ and their concentrations. 3.2.1. Effect of duration of purification on the Fe3+ and Cu2+ concentration in purified solution The experimental durations were varied between 10 min and 60 min. As shown in Fig. 3, about half an hour duration of purification was sufficient for the deoxidization of Cu2+, whereas the total deoxidization of Fe3+ to Fe2+ needed 1 h. Therefore, if collective deoxidization (deoxidization of Fe3+ to Fe2+ and Cu2+ to Cu) was taken into consideration, 1 h duration of deoxidization would be sufficient for the optimum deoxidization duration. 3.2.2. Effect of mass of bismuth powder addition on the Fe3+ and Cu2+ concentration in purified solution Experiments of the this group were made to examine the effect of mass of bismuth powder addition on the Fe3+ and Cu2+ concentration in deoxidized solution. These experiments were carried out at bismuth powder addition changing from 1.0 to 1.5 multiple of stoichiometric quantum. The results obtained were given in Fig. 3. By looking at Fig. 4, it was decided that the addition of 1.4 multiple of stoichiometric quantum of bismuth powder could be the best choice
345
Fig. 5. Extraction rate of Bi(III) as a function of N235 concentration.
available because this addition was the best result with more than 99.0% Cu2+ deoxidization and more than 99.0% Fe3+ be reduced into Fe2+. 3.2.3. Bismuth recovery 3.2.3.1. Extraction as a function of N235 concentration. In order to optimize the conditions for extraction of Bi(III), kerosene solutions of N235 with varying molar concentrations were employed and TBP (1% v/v) was added as phase modifier. It was found that 8.0 × 10− 2 M N235 was sufficient for quantitative extraction of Bi(III), and further increase in the ratio of N235 was found to be unnecessary (Fig. 5). 3.2.3.2. Extraction with various diluents. Bi(III) was extracted with 8.0 × 10− 2 M N235 in various solvents. The extraction of Bi(III) was quantitative with kerosene, toluene, xylene, cyclohexane and benzene, whereas toluene, hexane, xylene, cyclohexane and benzene, chloroform, carbon tetrachloride and dichloromethane were found to be less efficient than kerosene as the diluent, and only kerosene shows a practically quantitative extraction. Hence, kerosene was preferred as the suitable diluent in this study (Table 3). 3.2.3.3. Effect of time of equilibrium. The solutions were shaken for a varying period ranging from 1.0 to 15.0 min. The extraction was quantitative over the entire period range after 3 min. Therefore a 3 min period of equilibration was used for the extraction of Bi(III). 3.2.3.4. Effect of pH value. The solutions were shaken for a varying pH values for 3 min, using hydrochloric acid and ammonia solution to adjust the pH value. It was determined that the extraction was highly efficient around pH 0.2. Therefore pH 0.2 was used for the extraction of Bi(III) (Fig. 6).
Table 3 Extraction for various diluents
Fig. 4. Effect of mass of bismuth powder addition on the Fe3+ and Cu2+ concentration in deoxidized solution.
Diluent
E (%)
Kerosene Toluene Xylene Cyclohexane Benzene Hexane Dichloromethane Carbon tetrachloride Chloroform
99.2 97.4 89.2 51.1 91.8 89.7 90.2 89.2 92.4
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Fig. 6. Extraction rate of Bi(III) as a function of pH value.
Fig. 7. Influence of roasting temperature on the extraction of molybdenum values from molybdenite concentrate by roasting in the presence of lime.
3.2.3.5. Effect of various stripping agents. Bi(III) was stripped with different strengths of ligands after their extraction. The stripping of Bi (III) was quantitative with 0.1 M to 1.0 M EDTA, DTPA, EGTA, DCTA. It is found that only 0.5 M EDTA showed the maximal stripping ratio of Bi (III). Hence, for all further investigations, 0.5 M EDTA was used as stripping agent for Bi(III) (Table 4). Under the optimum condition, a pregnant aqueous solution was obtained (Table 5). Based on the above mentioned investigation, it is clear that from this purified solution, bismuth could be extracted with N235, and it was determined that the extraction was highly efficient around pH 0.2 and no co-extraction of other metals was observed at the pH of bismuth recovery. This metal was stripped by ethylenediaminetetraacetic acid (EDTA) to give an enriched aqueous phase containing 9.93 gpl, and this pregnant solution was suitable for final recovery of bismuth by electrolysis, replacement or salt precipitation. 3.3. Molybdenum recovery The commercial route for the extraction of molybdenum from its sulfide mineral molybdenite involves roasting of the concentrate, purification of the resultant calcine, either by distillation of MoO3, or by a hydrometallurgical route, and, finally, hydrogen reduction of the trioxide to the metal. During roasting, much valuable metals are lost due to volatilization and the SO2 generated is a source of environmental pollution. Further, such a pyrochemical process requires a high-grade concentrate and is not suitable for the treatment of low grade molybdenite because it would result in the formation of stable molybdates of impurity elements which render subsequent processing difficult. Processing of low grade molybdenite concentrates for the extraction of molybdenum without entailing SO2, emission has, therefore,
received considerable attention in recent years. The processes reported are based on selective extraction of either molybdenum or the impurity elements. Selective extraction of MoO3 from low grade concentrate has been carried out by leaching with sodium hypochlorite solution, nitric acid, sodium hydroxide or by chlorination in the presence of oxygen. Leaching with ferric chloride or cyanidecontaining solutions has been used for preferential removal of impurity elements from the low grade concentrate. As far as the application of the roasting process to low grade concentrate is concerned, the process can be modified if roasting is carried out in the presence of lime. Roasting of molybdenite in the presence of soda ash has been reported for extraction of molybdenum by Wlodyka and Mehra et al. However, the process conditions reported by these investigators differ widely. A patent dealing with roasting of molybdenite in the presence of lime is available but complete process details are not known. The feasibility of applying the modified roasting processes to low grade molybdenite for extraction of MoO3 and Re values has been investigated by Juneja et al. (1996). The processes studied involve roasting of molybdenite in the presence of lime and of soda ash, and examines results of both processes for the extraction of MoO3 and Re values by a carbon adsorption technique involving selective adsorption of Mo on activated charcoal from the leach solution followed by desorption with ammonia. In this study, the bismuth-leached residue was treated with by roasting in the presence of slaked lime, and followed by hydrometallurgical treatment of the roasted products. The calcine was
Table 4 Effect of stripping agents for Bi(III) Stripping agent
Bi(III) recovery (%) at different concentration (M) 0.1
0.2
0.3
0.5
0.7
1.0
EDTA DPTA EGTA DCTA
69.2 32.1 18.2 23.1
86.5 34.0 – –
94.9 38.1 – 34.7
99.1 40.1 21.0 –
99.0 44.0 – –
– 47.5 32.2 40.1
Table 5 Chemical component of stripped solution Elements
Bi3+
Fe3+
Fe2+
Cu2+
Concentration/gpl
9.93
Trace
Trace
Trace
Fig. 8. Influence of charge composition on the extraction of molybdenum values from molybdenite concentrate by roasting in the presence of lime.
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Table 6 Results of leaching calcines Mode of leaching
Temperature of leaching (°C)
Period of leaching (h)
Molybdenum recovery (%)
Water 1 M HCl 2 M H2SO4 3 M HCl 4 M HCl 6 M HCl
80–90 80–90 80–90 80–90 80–90 80–90
3 3 3 3 3 3
0.3 89.2 56.3 94.8 99.1 99.2
Table 7 Main chemical component of the leached residual solution Elements
Mo6+
Fe3+
Fe2+
Bi3+
Concentration/gpl
5.98
10.88
0.01
0.02
leached with HCl and molybdenum can then be extracted from the leached residual solution with N235. From loaded solvent, this metal was stripped with ammonia solutions to give a pregnant solution suitable for final recovery of molybdenum by salt precipitation. 3.3.1. Lime roasting process The roasting of bismuth-leached residue, molybdenite, in the presence of lime/slaked lime (Ca(OH)2) results in the oxidation of molybdenum into calcium molybdate in accordance with: 2MoS2 þ 6CaðOHÞ2 þ 9O2 ¼ 2CaMoO4 þ 4CaSO4 þ 6H2 O The results of the influence of the temperature of roasting on the recovery of molybdenum were presented in Fig. 7. The molybdenum recoveries improved from 42.1% to 98.7% as the roasting temperature was increased from 400 °C to 800 °C. Further increase in the temperature of roasting was found to be unnecessary. Studies on the effect of charge composition indicated that the molybdenum yield increased from 36.4% to 98.9% as excess of lime over stoichiometric requirement was raised from 0%to 50%, and further increase in the excess ratio was found to be unnecessary (Fig. 8). Thus, a maximum recovery of 98.9% was obtained when an excess of 50% lime over stoichiometry requirement was roasted at 700 °C for 2 h. Results on the influence of roasting time indicated that 2 h of roasting was adequate for the roasting operation at 700 °C.
Fig. 10. Extraction rates of Mo(VI) and Fe(III) as a function of pH value.
calcined mass. The results of leaching studies were given in Table 6. It can be observed from the table that it was possible to extract more than 99% Mo in the leaching stage. Table 7 showed the main chemical components of the leached residual solution. 3.3.3. Solvent extraction of molybdenum The residual solution generated in the acid leaching step contains variable but significant amounts of molybdenum (VI) and iron (III) as shown in Table 7. In order to recover its contents of molybdenum and iron, this solution was treated by means of solvent extraction by contacting with an organic solvent formed by solutions of N235 in kerosene. TBP (1% v/v) was added as phase modifier in all experiments. Two variables were investigated for studying the extraction of molybdenum: the pH of aqueous solution and the concentration of N235 in the organic phase. The results are shown in Figs. 8 and 9. 3.3.3.1. Extraction as a function of N235 concentration. In order to optimize the conditions for extraction of Mo(IV) and Fe(III), kerosene solutions of N235 with varying molar concentrations were employed. It was found that 15 × 10− 2 M N235 was sufficient for quantitative extraction of Mo(IV) and Fe(III) (Fig. 9).
3.3.2. Solvent leaching of molybdenum Investigations on the leaching of the calcine with HCl indicated that a leaching with 4 M HCl at 80–90 °C for 3 h with constant stirring was necessary to extract the maximum amount of molybdenum from the
3.3.3.2. Effect of pH value. The solutions were shaken for varying pH values for 3 min, using hydrochloric acid and ammonia solution to adjust the pH value. It was determined that the extraction is highly efficient around pH 0.4, Therefore pH 0.4 was used for the extraction of Mo(IV) and Fe(III) (Fig. 10).
Fig. 9. Extraction rate E as a function of N235 concentration.
Fig. 11. Stripped rate of Fe(III) as a function of pH value.
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then optimized for selective and collective recovery and the optimum conditions for bismuth and molybdenum were determined. (2) Solvent extraction of bismuth and molybdenum by N235 from leach liquors mentioned in the processing of ShiZhuYuan bismuth glance flotation concentrate was studied. Under the optimized conditions, the ultimate recovery ratios of bismuth and molybdenum were more than 99% and 98% respectively. Acknowledgements The authors would like to express their gratitude to the Department of Science and Technology of China for supplying financial support, and ShiZhuYuan Rare Metal Co., China for supplying the material as well as the experiment equipments for this study. References Fig. 12. Stripped rate of Mo(VI) as a function of ammonia volume percentage.
From Figs. 9 and 10 the following information was observed: at pH 0.4, over 99% of Mo(VI) and over 98% of Fe(III) were extracted. Both molybdenum (VI) and Fe(III) were the metal values to be recovered. In order to obtain purified products, the two metals loaded in the organic phase must be further separated. Thus a stripping treatment was employed to unload the Fe(III) co-extracted with molybdenum (VI). According to the data shown in Fig. 2, the extraction of Fe(III) depends greatly on the equilibrium acidity, whereas that of molybdenum (VI) is essentially unaffected by acidity when pH value lies in the range of pH 0 to pH 2.0. It was, therefore, to be expected that the Fe(III) in the loaded solvent can be unloaded by shifting the equilibrium acidity by using hydrochloric acid solution as a scrubbing reagent. The results, given in Fig. 11, showed that the stripping ratio of Fe(III) was markedly increased with the pH value increase from pH 0.5 to pH 1.5, and more than 99% of the Fe(III) can be stripped into the aqueous phase. It was expected that molybdenum (VI) can be stripped by contacting the pregnant organic phase with an alkaline solution such as ammonia or ammonium carbonate. The strip tests were carried out with a variety of concentrations of ammonia solutions to examine the stripping efficiency for molybdenum (VI) in this work. The results were presented in Fig. 12 and showed that the complete stripping of molybdenum (VI) took place when the content of the ammonia solution was higher than 10 vol.%. Moreover, good phase separation was observed in all cases. After stripping, N235 is recycled to the extraction by regeneration with a 2 M hydrochloric acid solution. 4. Conclusions (1) A comprehensive procedure was introduced to separation and collective extraction of bismuth and molybdenum from low grade bismuth glance concentrate. All detailed processes, include: the hydrochloric acid leaching, purification and deoxidization, solvent extraction of bismuth, lime roasting process, leaching and solvent extraction of molybdenum, were
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