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Metals Recovery From Multimetal Sulphide Concentrates (CuFeS2–PbS–ZnS): Combination of Thermal Process and Pressure Leaching
Abstract
A laboratory-scale method for treating a bulk concentrate (CuFeS2–PbS–ZnS) for metals recovery was developed utilising a combination of thermal process (roasting) and pressure leaching as an alternative to conventional pyrometallurgical processing. Pyrometallurgy is becoming less acceptable from environmental standpoints for the treatment of bulk concentrates. Additionally, high capital costs make modern facilities cost prohibitive. The leaching agents employed, namely, sulphuric acid (lixivant) and ferric sulphate, are selective for metal sulphides, this coupled with the fact that they create fewer environmental problems and are economical makes this new process highly favourable. In the laboratory evaluation of this process, the metal values in the flotation concentrates were selectively recovered by combining roasting and pressure leaching. The experimental parameters studied included roasting temperature, and pressure leaching pulp density, temperature, and retention time. Laboratory results indicate that roasting followed by pressure leaching is an efficient and cost effective method of treating base metal sulphide concentrates.
... NOx and SO2 emissions can potentially occur during the hydrometallurgical processing of raw materials/intermediates to obtain useful metals. These may be intermediate products resulting from the processing of various metalcontaining ores, e.g., sulphide concentrates of metals [6]. They also occur, for example, in the recycling of electrical and electronic waste or in the recycling of lead-acid batteries, where leaching of stone and slag (lead melting intermediates) in nitric and sulfuric acid produces NOx or SO2 emissions [6,7]. ...
... These may be intermediate products resulting from the processing of various metalcontaining ores, e.g., sulphide concentrates of metals [6]. They also occur, for example, in the recycling of electrical and electronic waste or in the recycling of lead-acid batteries, where leaching of stone and slag (lead melting intermediates) in nitric and sulfuric acid produces NOx or SO2 emissions [6,7]. The lead smelting slag and matte from Pb batteries contain Pb in metallic form, as well as in the form of galena (PbS). ...
... In addition to these phases, impurities, such as Ca and Si are present in the rock and slag. The authors of the study [6] focused on the selective leaching of Pb from stone and slag. ...
With the increasing demand for electricity due to the increasing economic boom, there is an excessive production of emissions that natural processes cannot cope with. Fortunately, there are various technological solutions for capturing harmful substances from produced emissions. However, the European Union aims to prevent the formation of emissions in the process of industrial production itself. In order to achieve this, it is necessary to reconcile the interests of individual European Union member states by the implementation of regulations into laws, to monitor short-term and long-term changes in air quality, and also to put into practice increasingly effective methods of capturing emissions from the air. This paper offers an example of thermodynamic calculation of quantities by appropriate software in the process of recycling waste from the metallurgical industry and possibilities of technological improvement in emission reduction.
... The oxidants' presence is necessary for the successful leaching of sulfide minerals with an acidic solution [4][5][6]. Ferric [7][8][9][10][11][12][13] and cupric [14,15] ions, oxygen [16][17][18][19], other oxidants such as nitrate and nitrite ions [20][21][22][23][24][25], the manganese ions [26][27][28], dichromate ions [29], ozone [30], and bacteria [31][32][33][34] have been used as oxidative leaching agents of sulfide minerals in sulfate and chloride media, under atmospheric or pressure conditions. Hydrogen peroxide is a very strong oxidant with a standard redox potential of 1.77 V in an acidic solution [35]. ...
... -(19)) and per 1 mol of Fe 2 (SO 4 )3 (Eqs.(20)-(27)) are plotted.It is evident fromFigure 3that reactions Eqs. (7), (8), and(18) are thermodynamically the most dominant in the leaching process within the entire temperature range (0 to 100°C). Also, reactions with negative values of Δ r G θ , but less favourable are reactions Eqs. ...
... During the smelting of galena ores, gaseous sulfur dioxide (SO 2 ) and lead dust are released into the atmosphere, which requires the installation of expensive capturing systems. 2 Hydrometallurgical metal extraction is a fast developing and greener technique with a bright future 3,4 as it enables the processing of the mineralogically complex low-grade ores. 5,6 This is particularly important for heap-leaching, which eliminates the use of reactors and in situ leaching where metals are recovered directly from ore deposits without open-pit or underground mining. Lead extraction by hydrometallurgical leaching is being developed, 7,8 to meet the ever-increasingly stringent environmental regulations. ...
... Hydrogen peroxide is a very strong oxidant with a standard redox potential of 1.77 V against the standard hydrogen electrode under acidic conditions. 5 It is commonly used in leaching of metal-bearing sulfides as it is adsorbed on the mineral surface, receives electrons from the mineral, and decomposes to water (reaction 1). In the meantime, the mineral is oxidized and dissolved. ...
Organic solutions are promising lixiviants for fast and environmentally sustainable Pb extraction from ore minerals at low temperatures. However, engineering of novel leaching flowsheets has been hindered by poor understanding of the leaching mechanism, particularly the formation of surface-passivating phases. Here, we studied leaching of galena (PbS), the most abundant Pb ore mineral, in citrate and acetate solutions at 25–50 °C. The results show faster and higher Pb extraction in citrate solutions than in acetate solutions. For example, leaching of 53–106 μm galena particles at 35 °C for 2 h achieved 69.3% Pb extraction in a pH 7 citrate solution but only 30.1% in a pH 3 acetate solution. Investigation of solid residues by SEM, EDS, quantitative powder X-ray diffraction, and Raman spectroscopy proved the formation of a porous yet poorly permeable layer of anglesite during acetate leaching by pseudomorphic replacement of galena and subsequent overgrowth, hindering further leaching after 55.6% Pb extraction. In contrast, in citrate solutions, no anglesite was observed, but the formation of an impermeable thin Pb-oxide layer caused surface passivation after 87.1% Pb extraction. Our experimental results and thermodynamic calculations suggest that Pb-citrate complexes [e.g., Pb2(C6H5O7)22–, Pb(C6H5O7)24–, and Pb(C6H5O7)−] are far more effective than Pb-acetate complexes [e.g., Pb(CH3COO)+ and Pb(CH3COO)2] in suppressing the precipitation of anglesite because of the high solubility of Pb-citrate complexes in sulfate-rich solutions. This work provides a scientific basis for developing greener approaches such as in situ leaching and heap leaching for recovering Pb from galena-bearing ores.
... [5][6][7] Roasting + hydrometallurgy followed by mineral processing are one of the most applicable processes for the extraction of copper and cobalt from sulfidesbearing ores. After obtaining a sulfide concentrate containing a sufficient grade of cobalt and copper by flotation, a thermal process (roasting) is applied to the concentrate to transform most of the cobalt and copper sulfide to a soluble sulfate, while iron originated from other sulfides is converted to an oxide form, such as Fe 2 O 3 and Fe 3 O 4. Thus, the iron remains practically insoluble and can be discarded as tailings from the liquor containing the copper and cobalt metals [8,9]. As well known, higher Fe concentration in the solution influences the recovery of copper and cobalt from solution negatively in the following process [10,11]. ...
... MeS(s) + Na 2 SO 4 (s) → MeSO 4 (s) +Na 2 S(s) (8) Na 2 S(S) + 2O 2 (g) → Na 2 SO 4 (s) (9) It was explained that the catalytic action was due to the formation of sodium pyro-sulfate, which makes the reactant surface porous. This porous surface enhances the diffusion of SO 3 gas [30]. ...
The ancient flotation tailings from Lefke, Cyprus, have a potential for non-ferrous metals such as cobalt and copper from more than 9.5 million tons of reserves containing 0.38% Cu, 0.032% Co, and 22.6% Fe. Recovery of cobalt and copper from these tailings can provide great benefits from economic and environmental perspectives. While the ancient tailings were kept long time in storage in dumps, the characteristics of the material has become different from the common Co and Cu bearing ores. In order to extract these valuable metals, a process involving combination of roasting and leaching was applied in this study. Since this process responded to recovery to some extent, an innovative technique of using Na2SO4 as a promoter during roasting was proposed. Utilizing of Na2SO4 did not provide only higher metal extractions, but also resulted in high selectivity. In the scope of the study, following the determination of the mineralogy and chemical composition of the tailings, certain processes such as direct leaching and sulfation roasting with/without additives before leaching were applied. Since low metal extractions were obtained from direct leaching, a selective sulfation process was applied on the tailings before leaching in order to produce a pregnant solution containing higher amounts of Co and Cu metals by reducing the iron concentration. When sulfation roasting was performed without any additive, the desired selectivity could not be provided, since the sample was considerably oxidized under the atmospheric conditions in the past. Therefore, the usage of Na2SO4 as an additive with the amount of 25% at a roasting temperature of 700єC was tested. Consequently, Na2SO4 improved the sulfation and resulted in higher cobalt (90.1%) and copper (71.2%) recovery during leaching. Beside this, an improvement with a selective sulfation was obtained by combined effect of Na2SO4 addition and temperature, and the iron extraction decreased from 26.1% to 3.9%.
... Alternatively, pyrite or chalcopyrite addition increased zinc extractions. Many authors have reported that galvanic interactions are known to occur between conducting minerals and play a significant role in flotation (Rao and Finch, 1988;Kelebek et al., 1996;Zhang et al., 1997;Ekmekçi and Demirel, 1997;Huang and Grano, 2005), leaching (Mehta and Murr, 1983;Abraitis et al., 2003;Akcil and Ciftci, 2003), supergene enrichment of sulfide ore deposits (Thornber, 1975;Sato, 1992), environment governance (Alpers and Blowes, 1994), and geochemical processes (Sikka et al., 1991;Banfield, 1997). In addition, other researchers (Koleini et al., 2010;Koleini et al., 2011;Dixon and Tshilombo, 2005;Mehta, and Murr 1983;Holmes and Crundwell, 1995) reported when the amount of pyrite in contact with chalcopyrite increases, the leaching rate of chalcopyrite mineral increases. ...
... This is in agreement with other studies where it was noticed when pyrite was mixed with a second sulphide mineral, the second mineral oxidized more rapidly (Buehler and Gottschalk, 1910). The formation of greater amounts of H2O2 when pyrite was mixed with a second sulphide mineral may explain the effect of interaction between pyrite and the second sulphide mineral on flotation, leaching, environment governance and geochemical processes while the entire literature describes so far of galvanic interaction between two contacting sulphide minerals and electron transfer from one to the other (Rao and Finch, 1988;Kelebek et al., 1996;Zhang et al., 1997;Ekmekçi and Demirel, 1997;Huang and Grano, 2005;Mehta and Murr, 1983;Abraitis et al., 2003;Akcil and Ciftci, 2003;Thornber, 1975;Sato, 1992 , 1994;Sikka et al., 1991;Banfield, 1997). ...
Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by sulphide minerals during grinding was examined. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite ((Zn, Fe) S), and galena (PbS) generated H2O2 in pulp liquid during wet grinding and also when the freshly ground solids are placed in water immediately after dry grinding. Pyrite produced more H2O2 than other minerals and the order of H2O2 production by the minerals was found to be pyrite > chalcopyrite > sphalerite > galena. The pH of the water influenced the extent of hydrogen peroxide formation with greater amounts of H2O2 produced at highly acidic pH. Furthermore, the effect of mixed sulphide minerals, i.e., pyrite–chalcopyrite, pyrite–galena, chalcopyrite–galena and sphalerite–pyrite, sphalerite–chalcopyrite and sphalerite-galena on the formation of H2O2 showed increasing H2O2 formation with increasing pyrite fraction. There is clear correlation of the amount of H2O2 production with the rest potential of the sulphide minerals; the greater the rest potential of a mineral the greater the formation of H2O2. This study highlights the necessity of revisiting the electrochemical and/or galvanic interactions between sulphide minerals, and interaction mechanisms between pyrite and other sulphide minerals in terms of their flotation behaviour in the context of inevitable H2O2 existence in the pulp liquid.
... Alternatively, a pyrite or chalcopyrite addition increased zinc extractions. Many authors have reported that galvanic interactions are known to occur between conducting minerals and play a significant role in flotation (Rao and Finch, 1988;Kelebek et al., 1996;Zhang et al., 1997;Ekmekçi and Demirel, 1997;Huang and Grano, 2005), leaching (Mehta and Murr, 1983;Abraitis et al., 2003;Akcil and Ciftci, 2003), supergene enrichment of sulfide ore deposits (Thornber, 1975;Sato, 1992), environment governance (Alpers and Blowes, 1994), and geochemical processes (Sikka et al., 1991;Banfield and Nealson, 1997). In addition, other researchers (Koleini et al., 2010;Koleini et al., 2011;Dixon and Tshilombo, 2005;Mehta and Murr, 1983;Holmes and Crundwell, 1995) reported when the amount of pyrite in contact with chalcopyrite increases, the leaching rate of chalcopyrite mineral increases. ...
... This is in agreement with other studies where it was noticed when pyrite was mixed with a second sulphide mineral, the second mineral oxidized more rapidly (Buehler and Gottschalk, 1910). The formation of higher amounts of H 2 O 2 when pyrite was mixed with a second sulphide mineral may explain the effect of interaction between pyrite and second sulphide mineral on flotation, leaching, environment governance and geochemical processes while the entire literature describes so far of galvanic interaction between two contacting sulphide minerals and electron transfer from one to the other (Rao and Finch, 1988;Kelebek et al. 1996;Zhang et al. 1997;Ekmekçi and Demirel, 1997;Huang and Grano, 2005;Mehta and Murr, 1983;Abraitis et al. 2003;Akcil and Ciftci, 2003;Thornber, 1975;Sato, 1992;Alpers and Blowes, 1994;Sikka et al. 1991;Banfield and Nealson, 1997). an increasing sphalerite percent in the mixture, the concentration of H 2 O 2 is decreased. ...
The formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by sulphide minerals during grinding was investigated. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite (ZnS), and galena (PbS), which are the most abundant sulphide minerals on Earth, generated H2O2 in pulp liquid during wet grinding in the presence and absence of dissolved oxygen in water and also when the freshly ground solids were placed in water immediately after dry grinding. Pyrite generated more H2O2 than the other sulphide minerals and the order of H2O2 production by the minerals was found to be pyrite > chalcopyrite >sphalerite> galena. The pH of water influenced the extent of hydrogen peroxide formation where higher amounts of H2O2 were produced at highly acidic pH. The amount of H2O2 formed also increased with increasing sulphide mineral loading and grinding time due to increased surface area and its interaction with water.
The sulphide surfaces are highly catalytically active due to surface defect sites and unsaturation because of broken bonds and capable of breaking down the water molecule leading to hydroxyl free radicals. The type of grinding medium on formation of hydrogen peroxide by pyrite revealed that the mild steel produced more H2O2 than stainless steel grinding medium, where Fe2+ and/or Fe3+ ions played a key role in producing higher amounts of H2O2.
Furthermore, the effect of mixed sulphide minerals, i.e., pyrite–chalcopyrite, pyrite–galena, chalcopyrite–galena and sphalerite–pyrite, sphalerite–chalcopyrite and sphalerite–galena on the formation of H2O2 showed increasing H2O2 formation with increasing pyrite fraction in chalcopyrite–pyrite composition. In pyrite–sphalerite, chalcopyrite–sphalerite or galena– sphalerite mixed compositions, with the increase in pyrite or chalcopyrite proportion, the concentration of H2O2 increased but with increase in galena proportion, the concentration of H2O2 decreased. By increasing the pyrite proportion in pyrite–galena mixture, the concentration of H2O2 increased. Similarly, in the mixture of chalcopyrite–galena, the concentration of H2O2 increased with increasing chalcopyrite fraction. The results of H2O2formation in pulp liquid of individual sulphide minerals and in combination at different experimental conditions have been explained by Eh–pH diagrams of these minerals and the existence of free metal ions that are equally responsible for H2O2 formation besides the catalytic activity of surfaces. The results of the amount of H2O2 production also corroborate with the rest potential of the sulphide minerals; higher the rest potential more is the formation of H2O2. Most likely H2O2 is responsible for the oxidation of sulphide minerals and dissolution of non-ferrous metal sulphides in the presence of ferrous sulphide besides the galvanic interactions.
This study highlights the necessity of revisiting the electrochemical and/or galvanic interactions between the grinding medium and sulphide minerals, and interaction mechanisms between pyrite and other sulphide minerals in terms of their flotation behaviour in the context of the inevitable existence of H2O2 in the pulp liquid.
... Alternatively, a pyrite or chalcopyrite addition increased zinc extractions. Many authors have been reported that galvanic interactions are known to occur between conducting minerals and play a significant role in flotation (Ekmekçi and Demirel, 1997;Huang and Grano, 2005;Kelebek et al., 1996;Rao and Finch, 1988;Zhang et al., 1997), leaching ( Abraitis et al., 2003;Akcil and Ciftci, 2003;Mehta and Murr, 1983), supergene enrichment of sulfide ore deposits (Sato, 1992;Thornber, 1975), environment governance ( Alpers and Blowes, 1994), and geochemical processes (Banfield and Nealson, 1997;Sikka et al., 1991). In addition, other researchers ( Dixon and Tshilombo, 2005;Holmes and Crundwell, 1995;Koleini et al., 2010Koleini et al., , 2011Mehta, and Murr, 1983) reported when the amount of pyrite in contact with chalcopyrite increases, the leaching rate of chalcopyrite increases. ...
... It can be seen from this figure that with an increase in rest potential, the concentration of H 2 Fig. 7 shows the effect of single or mixture of minerals on formation of hydrogen peroxide at pH 4.5. It can be seen that pyrite and a mixture of pyrite and other sulphide mineral generated more H 2 O 2. This is in agreement with other studies where it was noticed when pyrite was mixed with a second sulphide mineral, the second mineral oxidized more rapidly (Buehler and Gottschalk, 1910 Mixture of minerals and other sulphide minerals in the context of flotation, leaching, environmental control and geochemical processes, which the literature so far has described as galvanic interactions between two contacting sulphide minerals and electron transfer from one to the other ( Abraitis et al., 2003;Akcil and Ciftci, 2003;Alpers and Blowes, 1994;Banfield and Nealson, 1997;Ekmekçi and Demirel, 1997;Huang and Grano, 2005;Kelebek et al., 1996;Mehta and Murr, 1983;Rao and Finch, 1988;Sato, 1992;Sikka et al., 1991;Thornber, 1975;Zhang et al., 1997). ...
Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by sulphide minerals during grinding was investigated. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite ((Zn,Fe)S), and galena (PbS), which are the most abundant sulphide minerals on Earth, generated H2O2 in pulp liquid during wet grinding in the presence of dissolved oxygen in water and also when the solids are placed in water immediately after dry grinding. Pyrite generated more H2O2 than other minerals and the order of H2O2 production by the minerals found to be pyrite > chalcopyrite > sphalerite > galena. The pH of water influenced the extent of hydrogen peroxide formation where higher amounts of H2O2 are produced at highly acidic pH. Furthermore, the effect of mixed sulphide minerals, i.e., pyrite–chalcopyrite, pyrite–galena, chalcopyrite–galena and sphalerite–pyrite, sphalerite–chalcopyrite and sphalerite–galena on the formation of H2O2 showed increasing H2O2 formation with increasing pyrite fraction in chalcopyrite–pyrite, galena–pyrite and sphalerite–pyrite compositions. The results also corroborate the amount of H2O2 production with the rest potential of the sulphide minerals; higher rest potential of a sulphide mineral results in more formation of H2O2. Most likely H2O2 is responsible for the oxidation of sulphide minerals and dissolution of non-ferrous metal sulphides in the presence of ferrous sulphide in addition to galvanic interactions. This study highlights the necessity of revisiting the electrochemical and/or galvanic interactions between pyrite and other sulphide minerals in terms of their flotation and leaching behaviour in the context of inevitable H2O2 existence in the pulp liquid.
... Various mechanisms have been proposed to explain the influence of pyrite on the flotation of chalcopyrite. It has been reported by many authors that galvanic interactions occur between conducting minerals and play a significant role in flotation (Rao and Finch, 1988; Kelebek et al., 1996; Zhang et al., 1997; Ekmekçi and Demirel, 1997; Huang and Grano, 2005), leaching (Mehta and Murr, 1983; Abraitis et al., 2003; Akcil and Ciftci, 2003), supergene enrichment of sulfide ore deposits (Thornber, 1975; Sato, 1992 ), environment governance (Alpers and Blowes, 1994) and geochemical processes (Sikka et al., 1991; Banfield and Nealson, 1997). Recently, it was revealed that formation of H 2 O 2 takes place during wet grinding of complex sulfide ore (Ikumapayi et al., 2012). ...
... 14b and 14c, while the entire literature describes a galvanic interaction between two contacting sulfide minerals and electron transfer from one to the other as shown inFig. 14a (Rao and Finch, 1988; Kelebek et al., 1996; Zhang et al., 1997; Ekmekçi and Demirel, 1997; Huang and Grano, 2005; Mehta and Murr, 1983; Abraitis et al., 2003; Akcil and Ciftci, 2003; Thornber, 1975; Sato, 1992; Alpers and Blowes, 1994; Sikka et al., 1991; Banfield, 1997).Mikhin et al., 2004; Misra and Fuerstenau, 2005; Watling, 2006; Liu et al., 2007) The ferric ions would generate ferrous ions according to Eq. (16), and ferrous ions can generate H 2 O 2 (Eqs. (17) and (18)), as shown inTable 1. H 2 O 2 oxidizes chalcopyrite (Eq. ...
Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by chalcopyrite (CuFeS2), which is a copper iron sulfide mineral, during grinding, was investigated. It was observed that chalcopyrite generated H2O2 in pulp liquid during wet grinding and also the solids when placed in water immediately after dry grinding. The generation of H2O2 in either wet or dry grinding was thought to be due to a reaction between chalcopyrite and water where the mineral surface is catalytically active in producing •OH free radicals by breaking down the water molecule. Effect of pH in grinding medium or water pH in which solids are added immediately after dry grinding showed lower the pH value more was the H2O2 generation. When chalcopyrite and pyrite are mixed in different proportions, the formation of H2O2 was seen to increase with increasing pyrite fraction in the mixed composition. The results of H2O2 formation in pulp liquid of chalcopyrite and together with pyrite at different experimental conditions have been explained by Eh-pH diagrams of these minerals. This study highlights the necessity of revisiting the electrochemical and/or galvanic interaction mechanisms between the chalcopyrite and pyrite in terms of their flotation behaviour.
... According to the Θ-pH diagram of metal sulfides, it is known that the properties of FeS 2 are quite stable and theoretically require a pH of less than −2 to achieve optimal leaching, while FeS 2 is more reactive and would be leached at a pH greater than 1 [17][18][19]. Through screening the literature [20], metal sulfides (e.g., FeS 2 and MoS 2 ) with a metal-tosulfur atomic ratio of 1:2 are classified as sulfate ion-forming sulfides, and oxygen pressure leaching only produces SO 4 2− from these sulfides, while metal sulfides (e.g., CuFeS 2 , ZnS, PbS and Fe 7 S 8 ) with a metal-to-sulfur atomic ratio close to or greater than 1:1 are classified as elemental sulfur-forming sulfides [21], and oxygen pressure leaching of these sulfides could produce elemental sulfur with a yield approaching 100% [22]. This suggests that if pyrite concentrate can be pyrolyzed to remove parts of sulfur, then sulfur products and pyrolytic slag pyrrhotite can be obtained, and the slag can be further processed through oxygen pressure acid leaching to obtain sulfur, and the concurrently obtained hematite can be further processed into iron concentrate [23]. ...
The preparation of high-purity sulfur and pyrrhotite by pyrolysis holds great potential to realize the high-value utilization of pyrite concentrate (FeS2), i.e., a by-product during the flotation of sulfide ore. In this study, the pyrrhotite obtained from the pyrolysis of pyrite concentrate was taken as the study object, and the effects of acid types, initial acidity, leaching time, leaching temperature, oxygen pressure, and liquid-to-solid ratio on the leaching behavior of pyrrhotite under oxygen pressure, were explored. The results show that elemental sulfur and hematite-based iron residue can be obtained by oxygen pressure leaching of pyrrhotite. It is found that the optimal experimental conditions for pyrrhotite oxygen pressure leaching are hydrochloric acid with 0.8 mol/L of initial acidity, 5 h of leaching time, 0.8 MPa of oxygen partial pressure, and 9:1 of liquid to solid ratio at 150 °C; moreover, the yield of sulfur reached 88.37%. Under optimal conditions, the leaching ratios of Fe, Pb, and Zn were 19.8%, 92.25%, and 99.11%, respectively. The sieved leaching residue was roasted at a low temperature of 500 °C, where the grade of Fe in the obtained hematite iron powder was 61.46%, and the grades of Pb, Zn, and S were 0.082%, 0.024%, and 0.1%. Clearly, the results meet well with the standard of the first grade of pyrite cinder, and this process realizes the comprehensive recovery of Fe and S resources in pyrolysis slag, which provides a superb technical route for the high-value utilization of pyrite concentrate.
... The influences of temperature and time on the leaching degrees of the zinc, copper and iron are presented in Figure 6. It is well known that the oxidative dis-solution of galena produces insoluble PbSO4 in the sulphuric acid leaching medium [51].Low leaching degree is obvious for all three metals, but for copper it is extremly low (Fig. 6.a). The similar situation occures for the influence of time on the leaching degrees at the temperature of 40 o C (Fig. 6.b). ...
The results of polymetallic sulphide Cu-Zn-Pb concentrate leaching with sulphuric acid in the presence of sodium nitrate as an oxidizing agent, at atmospheric pressure, are presented and discussed. Chemical composition and phase ratio of the starting concentrate and solid residuals after the leaching process are shown. Chemical reactions of leaching and their thermodynamic probabilities are predicted based on the calculated Gibbs energies and analysis of E-pH diagrams. The influence of temperature and time on the leaching degree of the concentrate's components is experimentally determined. It is shown that it is possible to obtain copper, zinc and iron in a solute form, while lead in the anglesite (PbSO4) form remains in the solid residual after the leaching process. The iron is being oxidized to Fe(III)-sulphate, which takes part in a sulphide leached minerals and turns into Fe(II)-sulphate.
... The demand for Pb continuously increases with industrial and economic development. Galena (PbS) is an important and abundant source of Pb (Hu et al., 2020b) and mining and processing of galena increase yearly with increasing Pb demand (Akcil and Ciftci, 2003;Yang et al., 2013). Galena often coexists with other sulphide minerals such as sphalerite (ZnS), pyrite (FeS 2 ) and oxidised minerals (Luo et al., 2016;Sun et al., 2012). ...
Galena (PbS) is an important and abundant source of lead (Pb) metal. It often coexists with other sulphide minerals such as sphalerite (ZnS). Flotation has been extensively used to separate galena and sphalerite. Herein, a reagent scheme comprising aerofloat collectors and Zn 2+ and SO 3 2− depressants is developed for the flotation separation of galena from sphalerite-rich sulphide ore. The flotation results demonstrate that Zn 2+ and SO 3 2− provide a synergistic depression effect on sphalerite and the aerofloat collectors exhibit more selectivity towards galena. Static calculations and ab initio molecular dynamics simulations were employed to investigate the mechanism of the adopted reagent scheme on galena/sphalerite flotation separation at the atomic scale. The reagent scheme used here has great potential for beneficiating other galena-containing sulfide ores. The calculation and simulation results can help in the design of novel collectors and depressants exhibiting excellent selectivity for the flotation separation of galena/sphalerite.
... The primary objectives of this study are, firstly, to assess leaching efficiency of ferric methanesulfonate; secondly, to study the effect of Fe content in sphalerite on leaching because sphalerite often contains various amount of Fe in its structure [30,31]; thirdly, to study the effect of Pb 2+ concentration in lixiviant on leaching, because lixiviant used in zinc leaching may contain Pb 2+ due to the common presence of Pb minerals (e.g. galena) in sphalerite in many ore deposits [32][33][34], and because the presence of Pb 2+ may lead to anglesite (PbSO 4 ) precipitation. Particularly, this work aims to understand the formation mechanism of surface product phases, and the impact of porosity of the produced surface phases on leaching kinetics. ...
Zinc is currently extracted from sphalerite (ZnS) by either sintering-smelting or roasting-leaching. These high-temperature processes produce environmentally hazardous sulfur dioxide gas. Hence, sustainable zinc extraction calls for direct leaching at low temperatures. Here, we show that ferric methanesulfonate is a high performing lixiviant for this purpose. Using 0.8 M ferric methanesulfonate, 99.3% Zn was extracted from 106-150 μm sphalerite particles after leaching at 70 °C for 96 h. Elemental sulfur, rather than sulfur dioxide, was produced as a by-product. When compared to common inorganic lixiviants such as ferric sulfate or ferric chloride, ferric methanesulfonate demonstrates greater extraction efficiency and is less corrosive, less toxic, and does not release harmful gases. Mineralogical, microscopical, and compositional characterization of reaction products confirmed the formation of a core-shell structure consisting of a sphalerite core and a sulfur shell, and the manifestation of a coupled dissolution-reprecipitation mineral replacement mechanism. Efficient zinc extraction was facilitated by the interconnected pores in sulfur, which provided pathways for mass transfer between the lixiviant and sphalerite core. The precipitation of anglesite (PbSO4), which in some instances caused surface passivation through the infilling of sulfur shells, suggests that Pb²⁺ concentration should be controlled when applying ferric methanesulfonate for Zn extraction.
... In this field, several authors have studied different (bio)-hydrometallurgical processes to recover copper and zinc from polymetallic sulphide ores, paying special attention to ferric leaching and bioleaching (Lorenzo-Tallafigo et al. 2018;Fomchenko and Muravyov, 2018;Fomchenko et al. 2019;Fomchenko and Muravyov 2018;Carranza et al. 1997;Tipre and Dave 2004). Besides, other processes have been also proposed to recover target metals from polymetallic ores, such as pressure leaching (with oxygen or sulphur dioxide), roasting, chlorination, nitrate-sulphuric acid leaching, glycine leaching, sodium meta-bisulphate leaching, or hydrogen peroxide leaching (Xu et al. 2011, Akcil andCiftci 2003;Tarasov and Timoshenko 2006;Cui et al. 2020;Sokic et al., 2017;Hara et al. 2020;Sokic et al., 2019;Shin et al. 2019). ...
Nowadays sulphide ores exploitation is undergoing some troubles, which are hindering the treatment through traditional routes. Bulk flotation followed by a novel hydrometallurgical process can dodge these difficulties. In this work, an integral hydrometallurgical process consists of two ferric leaching steps, followed by a hot brine leaching stage, is proposed to recover target metals from a bulk sulphide concentrate (2.9% Cu, 7.4% Zn, 2.5% Pb, 67 ppm Ag and 37.2% Fe). In the first ferric leaching step, sphalerite, galena and copper secondary sulphides are dissolved and, in the second leaching step, a silver salt is added in order to catalyse chalcopyrite oxidation. If silver salt is added at the beginning of the process, sphalerite passivation is observed, and therefore zinc recovery is not possible. However, when catalytic leaching is performed after a previous ferric leaching, copper and zinc recoveries higher than 95% are achieved. The leached concentrate (0.3% Cu, 0.8% Zn, 3.3% Pb, 1438 ppm Ag, 38.0% Fe and 6.6% S0), is treated by a hot brine leaching. When hot brine leaching is performed at high pulp density, elemental sulphur removal is necessary to recover all silver added as a catalyst. Extractions higher than 95% for Zn, Cu and Pb are achieved as well as the total recovery of catalyst. The proposed process is silver surplus; therefore, this agent can be recirculated.
... Therefore, it is convenient to treat these ores using some of the hydrometallurgical processes. Leaching of copper from chalcopyrite requires the presence of the oxidants in an acidic environment; some of the frequently present oxidants are: the ferric ions [3][4][5][6][7][8][9], the cupric ions [10][11][12], some acidophilic bacteria [13][14][15][16][17][18], the nitrate ions [19,20], the nitrite ions [21,22], the dichromate ions [23,24], the manganese ions [25,26], the permanganate ions [27], the chlorate ions [28], the oxygen ions [29][30][31][32][33][34], ozone [35], the silver ions [36,37], and the use of microwaves [38,39]. ...
In ores, chalcopyrite is usually associated with other sulfide minerals, such as sphalerite, galena, and pyrite, in a dispersed form, with complex mineralogical structures. Concentrates obtained by flotation of such ores are unsuitable for pyrometallurgical processing owing to their poor quality and low metal recovery. This paper presents the leaching of chalcopyrite concentrate from the location “Rudnik, Serbia”. The samples from the flotation plant were treated with hydrogen peroxide in sulfuric acid. The influences of temperature, particle size, stirring speed, as well as the concentrations of hydrogen peroxide and sulfuric acid were followed and discussed. Hence, the main objective was to optimize the relevant conditions and to determine the reaction kinetics. It was remarked that the increase in temperature, hydrogen peroxide content, and sulfuric acid concentration, as well as the decrease in particle size and stirring speed, contribute to the dissolution of chalcopyrite. The dissolution kinetics follow a model controlled by diffusion, and the lixiviant diffusion controls the rate of reaction through the sulfur layer. Finally, the main characterization methods used to corroborate the obtained results were X-ray diffraction (XRD) as well as qualitative and quantitative light microscopy of the chalcopyrite concentrate samples and the leach residue.
... The most straightforward is the increase in cyanide concentration, since according to the principles of chemical equilibrium the reaction is pushed forward as higher amounts of reactants are used. It is also possible to increase temperature and pressure to force the reaction to take place, [13]- [15] but these methods are often energy-intensive and/or require expensive designs of reactors. There is another family of methods in which leaching is favored by applying energy from different sources, such as microwaves and ultrasound. ...
Improving the efficiency of noble metal extraction without increasing the consumption of the leaching agent is one of the goals in the development of novel hydrometallurgical routes. This is particularly important when dangerous compounds (such as cyanide) are involved in mineral processing. Ultrasound-assisted leaching is presented here as an advantageous process able to increase the amount of silver leached from a polymetallic sulfide-based mineral by ~300%, without increasing cyanide consumption. Our results suggest that sonication removes solid by-products that hinder the leaching process. While in the absence of sonication cyanide consumption rises steadily even though silver leaching has reached a maximum, under sonication both cyanide consumption and silver extraction follow similar trends, indicating that side reactions are decreased. The enhanced efficiency in both silver extraction and cyanide consumption indicate that ultrasound-assisted leaching can be introduced as a greener method in mineral processing.
... Nevertheless, direct leaching of chalcopyrite has some problems due to the low solubility of chalcopyrite without the presence of an oxidant, the formation of precipitation, and the disposal of a large amount of iron that dissolves along with the copper [16][17][18]. Therefore, many studies have been investigated using various leaching/oxidizing reagents such as ferric and cupric ions, bacteria, oxygen in sulfuric and chloride media under the atmospheric or pressure leaching conditions at elevated temperatures [19][20][21][22][23][24][25][26][27][28][29][30][31][32][33][34][35]. Unfortunately, the kinetics of copper dissolution from chalcopyrite are very slow due to the formation of the passivation layer on the surface of chalcopyrite particles by oxidation of sulfide to form elemental sulfur under oxidative conditions. ...
The leaching of copper from chalcopyrite in H2SO4 solution under pressure-oxidative conditions and its kinetics were investigated in this study. Leaching variables that affect the rate of copper dissolution from chalcopyrite are agitation speed (300–900 rpm), total pressure (0.8–2.0 MPa), temperature (160–180 °C), and sulfuric acid concentration (0.1–2.0 M). Results showed that dissolution of chalcopyrite increases with increasing agitation speed, total pressure, and temperature, whereas it decreases with the increasing sulfuric acid concentration. Under the optimal conditions, copper extraction of 94.5% was achieved after 90-min leaching, while a dissolution of iron at 4.2% was obtained. The kinetic study showed that the dissolution of chalcopyrite is represented by a shrinking core model with chemical reaction controlling mechanism given as (1 − (1 − α)1/3). The activation energy (Ea) for the leaching reaction was calculated to be 42.4 kJ/mol. The reaction order with respect to total pressure was about 8.0, which indicates that total pressure, i.e., oxygen partial pressure, in an autoclave is the most important factor in controlling the dissolution of chalcopyrite in H2SO4 solution under pressure-oxidative leaching conditions. The effect of Fe/Cu mole ratio (1–20 mol/mol, adjusted by addition of pyrite) on chalcopyrite leaching from a copper ore was investigated. The results show that the sulfuric acid produced during pyrite oxidation promotes the chalcopyrite dissolution.
... Sphalerite (ZnS) is the main phase in sulfide zinc concentrates, which will oxidize during oxidative roasting. However, the formation of ZnFe 2 O 4 decreases the rate of leaching of zinc (Akcil and Ciftci, 2003 ...
... Ferric and cupric ions, oxygen and bacteria have been used as oxidative leaching agents of chalcopyrite in sulphate and chloride media, under atmospheric or pressure conditions. Leaching investigations of chalcopyrite show that this mineral is hardly reactive at moderate temperatures around 60°C (Dutrizac et al., 1969;Dutrizac, 1981;Dutrizac and MacDonald, 1974;Charles and Han, 1997;Beckstead and Miller, 1977;Ammou-chokroom et al., 1977a,b;Bonan et al., 1981;Rath et al., 1988;Ngoc et al., 1990;Saxena and Mandre, 1992;Akcil and Ciftci, 2003). ...
In this study, the leaching conditions of chalcopyrite (CuFeS2) concentrate enriched through flotation of Menka Corporation Cu-Pb-Zn complex ore in sulphuric acid medium under effect of potassium dichromate were investigated. Effects of stirring speed in the range of 100-600 rpm, sulphuric acid (H2SO4) concentration in the range of 0.1-0.5 M, potassium dichromate (K2Cr2O7) concentration in the range of 0.02-0.2 M, leaching temperature in the range of 20-90°C, particle size fractions of -212 +106, -106 +75, -75 +45, -45 +38 to minus 38 μm and solid/liquid ratio in the range of 10-200 g/L on copper dissolution were investigated in the experimental study. According to obtained results, it was determined that the copper dissolution from chalcopyrite under the effect of potassium dichromate ions in sulphuric acid medium was directly proportional to stirring speed, sulphuric acid and potassium dichromate concentrations and leaching temperature. On the other hand, it increased with decreasing particle size and solid/liquid ratio. The copper dissolution of 94.94% was obtained with following conditions: leaching time of 120 minutes, stirring speed of 400 rpm, sulphuric acid concentration of 0.5 M, potassium dichromate concentration of 0.1 M, leaching temperature of 50°C, particle size of minus 38 μm and solid/liquid ratio of 10 g/L. At the same leaching condition, the copper dissolution of 96.82% was obtained for 90°C.
... [1] By the process of zinc production, the sphalerite (ZnS) present in the concentrates is firstly converted into soluble oxide structure by an oxidative roasting and, then, the zinc calcine produced is subjected to a neutral and low acid leaching to generate a zinc sulfate solution for purification and subsequent electrolysis. [2][3][4] In general, approximately 10 pct of iron impurity accompanies with the concentrates, which tends to react with the zinc oxide to form zinc ferrite (ZnFe 2 O 4 ) in the oxidative roasting process. [5][6][7] Zinc ferrite is stable and insoluble in dilute sulfuric acid solution, and thus its formation results in a generation of considerable amount of leach residues in the following leaching process. ...
The selective leaching of zinc from high iron-bearing zinc calcine after reduction roasting was optimized by Taguchi experimental design method. The experimental parameters and their ranges were 303 to 343 K (30 to 70 °C) for leaching temperature (T), 7 to 15 mL/g for liquid/solid ratio (L/S), 70 to 150 g/L for H2SO4 concentration (C), 5 to 25 minutes for time (t), and 100 to 500 rpm for stirring speed (R). The results show that the optimum conditions were 303 K (30 °C), 9 mL/g, 110 g/L, 20 minutes, and 400 rpm, respectively. Under these conditions, about 92.81 pct Zn was extracted and more than 86 pct Fe was reported into the leach residue. L/S and C had significant effects on the extractions of zinc and iron, while t and R had no significant effects, and T had significant effect on iron extraction but negligible effect on zinc extraction. This indicates that diffusion was not a major control step of the leaching process, and the dissolution of iron was controlled by chemical reaction. The interactive effects of parameters were negligible. The leach residue was mainly composed of Fe3O4 and ZnS, and its particle size was very fine. © 2015 The Minerals, Metals & Materials Society and ASM International
... 16 Recently, a number of new pyrohydrometallurgical processes have been proposed to recover zinc or iron resources from zinc leaching residue. [17][18][19] Transformational roasting of zinc ferrite with additives has been studied, 10,16 showing that despite of much more zinc extracted, considerable iron was simultaneously dissolved. Other researchers have studied the process of recovering iron from the residue by reduction roasting and sequent magnetic separation, the results of which indicated that the experimental conditions required further optimising, as the grade of the iron concentrate recovered was unqualified for steelmaking. ...
High iron bearing zinc sulphide ore is an important resource for zinc, and the reserve is very great in the world. It is very difficult to effectively obtain zinc and iron from the source by traditional technology. In this study, a novel method was proposed for recovery of zinc and iron from high iron bearing zinc calcine, and the key procedure, reduction roasting, was investigated. The effects of CO concentration, CO2 concentration, temperature and time on reduction roasting were studied respectively. The experimental results show that the content of soluble zinc and magnetic susceptibility reached 91.15% and 4.1 x 10(-4) m(3) kg(-1) under the optimum conditions respectively. About 90% Zn and 9.5% Fe were dissolved from the reduction roasted zinc calcine by low acid leaching, and simultaneously, 84.3% Fe was recycled to the iron concentrate containing iron of 53.2% from zinc leaching residue by low intensity magnetic separation.
... Hydrometallurgical processes can be typically categorized as chloride, sulfate, nitrate and ammonia leaching and bioleaching. Many studies on the dissolution of chalcopyrite using different solutions have been performed by many researchers (Prasad and Pandey, 1998; Majima et al., 1985; Dutrizac,1981 Dutrizac, ,1989 Dutrizac, ,1990 Arslan et al., 2004; Havlik et al.,1995 Havlik et al., , 2005 Mikhlin et al., 2004; Lu et al., 2000; Saxena and Mandre, 1992; Tchoumou and Roynette, 2007; McDonald and Muir, 2007; Akcil and Ciftci, 2003; Dreisinger, 2006; Hackl et al., 1995; Aydogan et al., 2006; Misra and Fuerstenau, 2005; Mahajan et al., 2007; Antonijevi et al., 1994 Antonijevi et al., , 2004 Padilla et al., 2007; Al-Harahsheh et al., 2005). In this study, the leaching conditions of chalcopyrite (CuFeS 2 ) concentrate taken from Koyulhisar (Sivas, Turkey) in a nitric acid medium were investigated by studying the effects of its leaching parameters, such as stirring speed, temperature and nitric acid concentration on Cu extraction. ...
The leaching conditions of chalcopyrite (CuFeS 2) concentrate taken from Koyulhisar (Sivas, Turkey) in a nitric acid medium were investigated by studying the effects of its leaching parameters, such as stirring speed, temperature and nitric acid concentration on Cu extraction. It was found that stirring speed has a little effect on the leaching. Copper extraction from chalcopyrite is directly proportional to nitric acid concentration. As the temperature increases, copper extraction also increases. The maximum copper extraction was obtained in 180 min of leaching time with the following conditions; 400 rpm stirring speed, 4.0 M nitric acid concentration and 80°C leaching temperature.
... To control the level of inorganic parameter contaminations in Obuasi area, the mine introduced hydrometallurgical (bio-oxidation) pre-treatment (Marsden and House 2006) of sulphide ores in 1992 and eventually phased out the pyrometallurgical pre-treatment of ores (Akcil and Ciftci 2003) in 2000. The hydrometallurgical process involves the use of natural bacteria such as thio-and ferrooxidan Thiobacilli to release the undesired constituent elements which are often discharged as harmless precipitates. ...
This study assessed the levels of selected inorganic contaminants in streams and stream sediments in the effluent areas relating to the pyrometallurgical and hydrometallurgical treatment of gold ores in the Obuasi gold mine, Ghana. Water and stream sediment samples were taken from specific locations during the consecutive rainy and dry seasons, and concentrations of phosphate (PO43-), nitrate (NO3-), chloride (Cl-), sulphate (SO42-), sodium (Na+), potassium (K+), calcium (Ca2+), magnesium (Mg2+), arsenic (As), copper (Cu), iron (Fe), zinc (Zn) and lead (Pb), were determined. Alkalinity, pH, temperature and specific electrical conductivity were also measured. In the water samples, the average pH range for both the seasons is 6.9-7.4, most anions and metals have relatively higher concentrations in the wet season than in the dry season at both the metallurgical sites. Trace metals concentrations were comparatively low (\0.01-5.00 mg/l), higher in the dry season at the pyrometallurgical sites. Irrespective of seasons, SO42- (0.80-949.50 mg/l) and PO43- (\0.01-6.30 mg/l) were pronounced at the pyrometallurgical sites, while NO3- (0.01-98.45 mg/l) and Cl- (1.88-49.05 mg/l) were higher at the hydrometallurgical sites. In water samples, Ca2+ and SO42+ were the dominant cation and anion, respectively. In the stream sediments, except pH, NO3-, Cl-, Na+ and Mg2+, all other parameter values were relatively higher at the hydrometallurgical areas. The average concentrations of Ca2+, Mg2+, As and Fe are remarkably high at both metallurgical sites (3,217-46,026 mg/kg). Overall, the level of parameters in the water samples are pronounced at pyrometallurgical sites, whereas the levels in sediments are higher at the hydrometallurgical sites.
... This mechanism may be the explanation for the behavior of increasing pyrite floatability in the presence of chalcopyrite as reported by Peng et al. (2003) of increased pyrite flotation after addition of chalcopyrite. The formation of higher amounts of H 2 O 2 when chalcopyrite is in contact with pyrite may explain the effect of the interaction between pyrite and chalcopyrite on flotation, leaching, environment governance and geochemical processes while the entire literature describes so far of galvanic interaction between two contacting sulphide minerals and electron transfer from one to the other (Rao and Finch, 1988;Zhang et al., 1997;Ekmekçi and Demirel, 1997;Huang and Grano, 2005;Mehta and Murr, 1983;Abraitis et al., 2003;Akcil and Ciftci, 2003;;Sato, 1992;Alpers and Blowes, 1994;Banfield, 1997). ...
Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by chalcopyrite (CuFeS2), which is a copper iron sulfide mineral, during grinding, was investigated. It was observed that chalcopyrite and pyrite generated H2O2 in pulp liquid during wet grinding and also the solids when placed in water immediately after dry grinding. The generation of H2O2 in either wet or dry grinding was thought to be due to a reaction between chalcopyrite and water where the mineral surface is catalytically active in producing •OH free radicals by breaking down the water molecule. When chalcopyrite and pyrite are mixed in different proportions, the formation of H2O2 was seen to increase with increasing pyrite fraction in the mixed composition. The results of H2O2 formation in pulp liquid of chalcopyrite and together with pyrite at different experimental conditions have been explained by Eh-pH diagrams of these minerals. This study highlights the necessity of revisiting the electrochemical and/or galvanic interaction mechanisms between the chalcopyrite and pyrite.
... The sulphide minerals in the ores are mainly dominated by the primary As bearing arsenopyrite (Osae et al., 1995). Ore treatment using diversity of metallurgical methods (Marsden and House, 2006;Akcil and Ciftci, 2003) has been going on at the mine for over a century. Pyrometallurgic and hydrometallurgical pre-treatment methods as well as tailing retreatment were used to recover the gold over the years . ...
Arsenic (As) sorption characteristics in decommissioned tailings dam environment at the Obuasi mine, Ghana was studied. The aim was to outline effective remediation strategy for As, hence the objectives were to establish: (1) Arsenic degradation capacity of the decommissioned tailings dam environment and empirical model(s) to describe the degradation pattern; (2) relevant equilibrium concentration range and time frame for As degradation by natural attenuation. Eighteen water sampling events from monitoring boreholes were spread over 24 months, while leachate from dynamic leaching experiment of soil/sediment were sampled during 14 sampling events, spread over 30 weeks. The samples were filtered through 0.45 μm cellulose membrane and filtrates analyzed for As concentrations. Results followed reducing mass pattern for both sets of observations. The Freundlich and Langmiur isotherms were applied to both data, with the Freundlich isotherm indicating a strong sorption capacity, with regression gradients of 0.528 and 0.615 and high predictive model (R2) values of 0.842 and 0.912 for both experimental and field data respectively. A t-test statistic established that, equilibrium existed between sorption processes within the media between compliance value (C), ranges of 0.50-0.01 mg/L and time period (t) of 14.5-45.5 months. The As degradation pattern within both investigated media is governed by the model equation: [t_n= - 8ln (c_n) + 9] or; [t_n = 8ln (1/C_n ) +9]; where n is a variable. The model may be used to evaluate soil sample leaching results from contaminated sites ear-marked for remediation through setting targets and objectives for environmental management plans.
... Alternatively, a pyrite or chalcopyrite addition increased zinc extractions. Many authors have been reported that galvanic interactions are known to occur between conducting minerals and play a significant role in flotation (Rao and Finch, 1988;Kelebek et al., 1996;Zhang et al., 1997;Ekmekçi and Demirel, 1997;Huang and Grano, 2005), leaching (Mehta and Murr, 1983;Abraitis et al., 2003;Akcil and Ciftci, 2003), supergene enrichment of sulfide ore deposits (Thornber, 1975;Sato, 1992), environment governance (Alpers and Blowes, 1994), and geochemical processes (Sikka et al., 1991;Banfield and Nealson, 1997). However, participation of H 2 O 2 and @BULLET OH, if any, in oxidation of the sulfide ore pulp components and hence in deteriorating of the concentrate grade and recovery of metal-sulfides has not yet been explored. ...
... Krysa (1995) described that in the zinc pressure leach process, zinc concentrate is acid leached in an autoclave at 145-150°C in an oxygen enriched atmosphere with a total pressure of 1100 kPa gauge. Amongst the more decisive researchers on this subject were Farbenindustrie (1927), Forward and Halpern (1956), Forward and Veltman (1959), Kunda et al. (1965), Boldt (1967), Veltman and O'Kane (1968), Jan et al.(1976), Doyle et al. (1978), Parker and Romanchuk (1979), Veltman and Bolton (1980), Parker (1981), Corriou et al. (1988), Dreisinger and Peters (1989), Doyle et al. (1989), Ashman and Jankola (1990), Dreisinger et al.(1990), Collins et al.(1990), Mason (1992), Harvey et al. (1992), Chalkley et al. (1993), Harvey et al. (1993), Krysa (1995), Ozberk et al. (1995), Jankola (1995), Baldwin and Demopoulos (1995), Hofirek and Nofal (1995), Owusu et al. (1995), Malone et al. (1995), Boissoneault et al. (1995), Bolorunduro et al. (2003), Akcil and Ciftci (2003), Duoqiang et al. (2008), Duoqiang et al. (2009), Li et al. (2010b, Langová and Matý sek (2010). ...
Mixed sulphide–oxide lead and zinc ores are most often found in the transition, and occasionally in the oxidised, zones of lead–zinc ore-bodies. They are of great importance because there are numerous unexploited or abandoned reserves of these ores in the world.However they present difficulties for conventional mineral processing due to complex mineralogy. In this paper, the specific problems associated with these types of ores are described and methods for solving these problems, combining economic and technical considerations, are discussed.The results of experiments carried out at laboratory scale are presented, in which the dissolution of mixed ore in sulphuric acid without oxidising agents was investigated. The results show the feasibility of zinc recovery from mixed sulphide–oxide lead and zinc ores, which underlines the potential of this approach. We also propose a conceptual flow diagram for the hydrometallurgical processing of these ores.
... Galvanic interactions between metal sulfide minerals have significant influences not only on hydrometallurgy (Chandraprabha et al., 2004;Cruz et al., 2005;Seke and Pistorius, 2006), flotation (Kelebek et al., 1996;Zhang et al., 1997;Ekmekc ßi and Demirel, 1997;Treviño et al., 2003) and leaching (Metha and Murr, 1983;Abraitis et al., 2003;Akcil and Ciftci, 2003), but also on geochemical processes (Sikka et al., 1991). ...
Managing mine water that has been contaminated with metal sulfide minerals due to galvanic corrosion is becoming an increasingly important environmental problem. Here, galvanic corrosion was investigated by studying galvanic interactions between pyrite–chalcopyrite and pyrite–galena in flowing mediums such as mine discharge water and flowing rainwater. The results showed that the corrosion current density of pyrite–galena is greater than that of pyrite–chalcopyrite under identical conditions. The corrosion current density of the galvanic cell tends to increase with increasing concentrations of strongly oxidizing ions (e.g., Fe3+) in the flowing medium, whereas the existence of non-oxidizing and non-reducing ions (e.g., Na+) have no obvious influence on the galvanic cell. In addition, the corrosion current density increases with increasing flow rate. Using the galvanic model, mixed potential theory and Butler–Volmer equation, the experimental results were explained theoretically. Because these experiments were performed under conditions very similar to those seen in mine discharge water and flowing rainwater, these results have direct implications for the future management and control of environmental pollution from mining operations.
Direct leaching of sphalerite has been considered as a cost effective and an environmentally benign alternative to the traditional two-step roasting-leaching approach. Yet the slow leaching rate due to the formation of surface passivating phases remains the main challenge. Here, we studied the mechanism and kinetics of sphalerite leaching in the temperature range of 35–130 °C using three oxidants, Fe2(SO4)3, FeCl3, and Fe(NO3)3, and observed distinctly different surface passivation and significantly different leaching rate. Leaching using Fe2(SO4)3 was the slowest, due to the formation of passivating surface layers of sulfur at the early stage and hydrated (Fe,Zn)-sulfates at the later stage. The formation of hydrated (Fe,Zn)-sulfates reduced Zn extraction by up to 20%, leading to incomplete Zn extraction. Using FeCl3, leaching was faster than in Fe2(SO4)3. However, the formation of surface sulfur also caused passivation. Leaching in Fe(NO3)3 was the fastest, as the initially formed surface sulfur was quickly oxidized to sulfuric acid and hence passivation was negligible. Using Fe(NO3)3, complete Zn extraction from 106 to 150 μm sphalerite particles took 7 days at 35 °C, 2 days at 70 °C, 5 h at 90 °C, and 1 h at 130 °C. This is about one order of magnitude and two orders of magnitude faster than leaching in FeCl3 and Fe2(SO4)3, respectively. The observed leaching behaviors were in agreement with changing activation energy as a function of leaching extent, analyzed by the modified ‘time-to-a-given-fraction’ method. In the case of Fe2(SO4)3 and FeCl3, leaching was controlled by phase-boundary reactions at the early-to-middle stages, but changed to diffusion control at the later stage after the formation of passivating surface layers; in the case of Fe(NO3)3, leaching was controlled by phase boundary reactions over the entire leaching process. This work demonstrates that Fe(NO3)3 is the more efficient oxidant than FeCl3 and Fe2(SO4)3 for fast leaching of sphalerite at low temperatures with minimum surface passivation.
Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by chalcopyrite (CuFeS2), which is a copper iron sulfide mineral, during grinding, was investigated. It was observed that chalcopyrite generated H2O2 in pulp liquid during wet grinding and also in the solids when placed in water immediately after dry grinding. The generation of H2O2 during either wet or dry grinding was thought to be due to a reaction between chalcopyrite and water where the mineral surface is catalytically active in producing ●OH free radicals by breaking down the water molecule. The effect of pH in the grinding medium or in the water in which solids are added immediately after dry grinding showed that the lower the pH value, the higher the H2O2 generation. When chalcopyrite and pyrite are mixed in different proportions, the formation of H2O2 was seen to increase with increasing pyrite fraction in the mixed composition. The results of H2O2 formation in pulp liquid of chalcopyrite and together with pyrite at different experimental conditions have been explained by Eh-pH diagrams of these minerals. This study highlights the necessity of revisiting the electrochemical and/or galvanic interaction mechanisms between the chalcopyrite and pyrite in terms of their flotation behavior.
As an important treatment method of processing the sphalerite, the technology of pressure acid leaching has many benefits, such as process flow short, no waste pollution, low production costs, high leaching rate of Zn, and S output in elemental form, which can be sold as a product. This paper takes the high indium sphalerite provided by Kunming Metallurgical Research Institute as the research object. A metallurgical behavior of sulfur conversion process is studied under high temperature and high pressure. The results indicate that the conversion rate of sulfur increases with the increasing liquid-solid ratio, leaching temperature, initial acid concentration and partial oxygen pressure in a certain range. With increasing leaching time, the conversion rate of sulfur increases first and then decreases. Considering all the aspects, the optimum parameters in the condition of conventional electric heating and microwave heating of the process are as follows: liquid-solid ratio 8:1, leaching temperature 423 K, initial acid concentration 120 g/L, partial oxygen pressure 1.0 Mpa, leaching time 90 min. The conversion rate of sulfur reaches 72.00% under the condition of conventional electric heating, 65.74% under the condition of microwave heating.
A process of the high temperature ferric leaching of the copper-zinc concentrate and pyrite product during a two-step biohydrometallurgical technology was studied. It was shown that the use of the galvanic effect during the leaching of the copper-zinc concentrate enabled to obtain a copper concentrate with low zinc content in the relatively short time (approximately 5.7 h), while zinc concentration in the liquid phase was sufficient for zinc and copper recovery by extraction and cementation, respectively. Zinc content in the concentrate decreased from 7.36 to 0.5% in the process of leaching. The recovery of zinc and copper into the liquid phase was 96.1 and 40.3%, respectively. The leaching of copper-zinc pyrite product and galvanic interactions of minerals made it possible to recover nonferrous metals from it almost completely and to leave the main amount of sulfur and iron in leach residues. Operation of the laboratory unit with the use of bioregeneration of the liquid phase showed the principal possibility of functioning of a two-step biohydrometallurgical technology under semi-continuous conditions with the closed cycle of technological flows. Zinc content decreased from 15.25 to 1.3% during the first step of the leaching process, whereas during the second (biological) step it declined to 1.03% (within 24 h). The recovery of zinc and copper into the liquid phase was 92.6 and 54.6%, respectively. Flow sheet of the copper-zinc concentrate treatment, which can serve as a basis for modernization of the treatment of sulfidic raw materials, is proposed.
The improvement of an evaporation-condensation method allows for successful recovery of elemental sulfur from sulfide concentrates from the zinc industry. Elemental sulfur can be obtained with this method in samples with a low (60%) sulfur content. The effects of heating temperature between 150 °C and 250 °C and heating time up to 120 minutes on the recovery of sulfur are also studied. Elemental sulfur obtained in this way is of high purity and therefore, there is no need for further purification. The treatment of these industrial residue would help removing sulfur from the environment.
The catalytic and galvanic effects of pyrite on the dissolution kinetics of a zinc sulfide concentrate in an acidified ferric sulfate medium were assessed. The effects of particle size, oxidant concentration (ferric ion), and amount of pyrite added with respect to the original amount of zinc concentrate was explored, while temperature (70 °C), hydrogen potential (pH = 1.0) and agitation speed (600 rpm) of the leach solution were held constant. A process based on the galvanic coupling between pyrite and sphalerite at solution potentials above 500 mV to ensure rapid and complete sphalerite dissolution in a ferric sulfate medium is proposed. As a result, the highest zinc extraction in the shortest possible time was obtained through minimizing the formation of a passivating film, which inhibits further reaction, and by a synergetic effect between pyrite and sphalerite. The rate of zinc leaching was considerably improved by doping with pyrite in the presence of Fe3 +. The results have demonstrated that it was possible to achieve 98.9% zinc extraction with doping pyrite in 6 h and 85% zinc extraction without pyrite in 7 h, both results being for a fine grain size. Ferric leaching reached 75% Zn extraction with doped pyrite and 70% without it, both in 7 h using a coarse grain size. Scanning electron microscopy (SEM) images show a passivating layer on the sphalerite surface after ferric leaching without pyrite, while in presence of pyrite this layer was not observed and the pyrite surface was not attacked. The galvanic interaction between pyrite and sphalerite particles can be explained by the difference between their Fermi energies and the electrochemical potential of the solution. During this process pyrite remains mostly unleached, acting as a redox catalyst (electron-accepting) for sphalerite (electron-donating) dissolution. This diminishes the typical passive behavior of sphalerite in a ferric solution enhancing considerably the kinetics of sphalerite. This confirms that the difference between the rest potentials of pyrite and sphalerite is the driving force that improves zinc dissolution. For an industrial hydrometallurgical operation, using an appropriate sphalerite-pyrite concentrate does not necessarily represent a significant cost because the ore contains both sulfides.
The adsorptive characteristics of Mycobacterium phlei on the surfaces of pyrite and galena were reported in this paper. The influence of environmental conditions, such as adsorptive time, pH, temperature, concentration of microbial cells and minerals, on the adsorptive effect was discussed in order to look into the probability of using the microorganism as depressant in the flotation of sulfide minerals. The experimental results showed that the adsorptive performances of Mycobacterium phlei on the surfaces of these minerals are distinctively different. Mycobacterium phlei showed a tendency to adsorb onto the surface of pyrite and the lower adsorption rate of the microorganism on the surface of galena is measured. SEM results indicated that the capsule on the surface of Mycobacterium phlei cells was the significant attachable section. According to the experiments, it can be concluded that pyrite will be depressed and galena will be floated preferentially in flotation process when the suitable amount of Mycobacterium phlei cells exists in flotation pulp.
Taking zinc sulfide concentrate as the research object, the method of oxygen-rich acid leaching was used to extract zinc element of the mineral and sulfur element remained in the leached residue. Effects of particle size, temperature, initial acid concentration, and oxygen partial pressure on kinetics of acid leaching were studied. The results show that the leaching process of zinc sulfide concentrate is controlled by the surface chemical reaction and follows the kinetic law of "shrinking of unreacted core", and the reaction activation energy is 73.58 kJ/mol. ©, 2014, Central South University of Technology. All right reserved.
Metal and energy extractive industries play a strategic role in the economic development of Sweden. At the same time these industries present a major threat to the environment due to multidimensional environmental pollution produced in the course of ageing of ore processing tailings and waste rocks. In the context of valuable sulphide mineral recovery from sulphide ore, the complex chemistry of the sulphide surface reactions in a pulp, coupled with surface oxidation and instability of the adsorbed species, makes the adsorption processes and selective flotation of a given sulphide mineral from other sulphides have always been problematic and scientifically a great challenge. Invariably, the problems associated with acid mine drainage and selectivity in flotation are explained to be associated with the oxidation of metal sulphides. Although metal sulphides oxidation and galvanic effects were well known in flotation and leaching of sulphides, recent studies reveal the formation of reactive, oxidizing oxygen species and H2O2 by sulphides due to the catalytic activity of sulphide surfaces. The inherent formation of H2O2 by single and mixture of sulphide minerals during wet and dry grinding systems and in open and closed environments have been investigated. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite ((Zn,Fe)S), and galena (PbS) generated H2O2 in pulp liquid during wet grinding in the presence and absence of dissolved oxygen in water and also when the freshly dry ground solids are placed in water immediately after grinding. Pyrite generated more H2O2 than other sulphide minerals and the order of H2O2 production by the minerals found to be pyrite > chalcopyrite > sphalerite > galena. The pH of water influenced the extent of hydrogen peroxide formation where higher amounts of H2O2 are produced at highly acidic pH. The amount of H2O2 formed also increased with increasing sulphide mineral loading and grinding time due to increased surface area and its interaction with water. The sulphide surfaces are highly catalytically active and capable of breaking down the water molecule leading to hydroxyl free radicals. Type of grinding medium on formation of hydrogen peroxide by pyrite and galena revealed that the mild steel produced more H2O2 than stainless steel grinding medium, where Fe2+ and/or Fe3+ ions played a key role in producing higher amounts of H2O2. In addition, the effect of mixed sulphide minerals, i.e., pyrite–chalcopyrite, pyrite–galena, chalcopyrite–galena and sphalerite–pyrite, sphalerite–chalcopyrite and sphalerite–galena on the formation of H2O2 showed increasing H2O2 formation with increasing the content of a nobler mineral or higher rest potential mineral in a mixed composition. The results of H2O2 formation in pulp liquid of sulphide minerals and mixed minerals at different experimental conditions have been explained by Eh–pH diagrams of these minerals and the existence of free metal ions that are equally responsible for H2O2 formation besides surfaces catalytic activity. The results also corroborate the amount of H2O2 production with the rest potential of the sulphide minerals; higher is the rest potential more is the formation of H2O2. Most likely H2O2 is answerable for the oxidation of sulphide minerals and dissolution of non-ferrous metal sulphides in the presence of ferrous sulphide besides the galvanic interactions. Studies have also been carried out to build correlation between percentage of pyrite in the concentrate, grinding conditions and concentration of OH•/H2O2 in the pulp and as well of controlling the formation of these species through known chemical means for depressing the generation of the oxidant. Flotation tests using a complex sulphide ore with the same reagent scheme that is being used at Boliden concentrator but with the addition of collector and depressant during grinding stage have been performed to judge the beneficial or detrimental role of H2O2 on the selective flotation of sulphides. The results demonstrate that the selectivity of metal sulphides against pyrite increases with increasing generation of H2O2 in the pulp liquid. This study highlights the necessity of revisiting into the electrochemical and/or galvanic interactions between the grinding medium and sulphide minerals, and interaction mechanisms between pyrite and other sulphide minerals in terms of their flotation behaviour, leaching and environmental degradation in the context of inevitable H2O2 existence in the pulp liquid.
The multimetallic sulfide copper ore containing large amounts of lead and iron was roasted in air atmosphere in the presence of calcium oxide, and the calcinate obtained was leached in sulfuric acid solution to extract copper. Characterization of the raw materials, calcinates and leach residues was conducted by using XRD and SEM/EDS analysis. The calcification mechanism of the complex ore was studied. The effects of temperature, stirring speed, liquid/solid ratio and sulfuric acid concentration on the kinetics and mechanism of copper dissolution from the calcinate were also investigated. Results of experiments show that sulfur retention efficiency in the calcinate achieves 97.77%, and that increasing both reaction temperature and acid concentration are capable of resulting in the increase of dissolution rate of copper. Leaching kinetics follows the un-reacted shrinking core model with a rate controlling step by diffusion through the solid product layer and the corresponding apparent activation energy is calculated as 19.21 kJ/mol, and consequently the rate of the dissolution of copper on aspect of H2SO4 concentration, liquid/solid ratio, and stirring speed can be expressed as .
The leaching of lead from galena in acidic hydrogen peroxide in presence of sodium chloride solution has been investigated with respect to the effects of hydrochloric acid and hydrogen peroxide concentrations, by changing the stirring speed, leaching temperature and the particle size. It was observed that leaching rate increases with increasing hydrochloric acid concentration, hydrogen peroxide concentration and the temperature. However, it decreases with increase in the particle size. The kinetic study showed that the leaching process is represented by shrinking core model with mixed kinetic. The activation energy (Ea) for the leaching reaction was calculated as 14·60 kJ mol-1, which is suggestive of the mixed controlled kinetics for the leaching reaction. © 2015 Institute of Materials, Minerals and Mining and The AusIMM.
A laboratory-scale method for treating bulk concentrate for reclaiming lead and silver was developed utilizing new hydrometallurgical technology as an alternative to the traditional pyrometallurgical processing. The condition experiments of every chief segment in the whole flowsheet have been systematically investigated, and then the whole hydrometallurgical processing flowsheet was determined. The main contents are followed as: Bulk concentrate was treated using pressure leaching in autoclave, the optimal leaching conditions were determined. The elemental sulphur was deprived from the pressure leaching residue using flotation-distillation. Carbonate conversion -silicofluoric acid leaching on flotation gangue containing lead sulfate using hydrometallurgy was carried. And Leaching silver using thiourea from the residue was carried after extracting lead. Through the whole hydrometallurgy flowsheet, the reclaiming of lead and silver was actualized.
The purpose of this work was to study the feasibility at laboratory-scale of a hydrometallurgical process for the selective recovery of valuable metals from partial silicated sphalerite in an oxygen pressure acid leaching system. The factors influencing dissolution efficiency of the ore were investigated and optimized. Under optimum conditions (i.e., temperature of 433 K, sulfuric acid concentration of 41.2 g/L, leaching time of 2.5 h, liquid/solid ratio of 6 mL/g, and pressure of 1.6 MPa) over 97% Zn was extracted into the leach liquor together with 0.3% SiO2 and 2.9% Pb. The leaching slurry had good solid–liquid separation characteristics, and the filtration rate could be as high as 716 L/m2 h. About 96% oxidation of sulfide sulfur to sulfate was achieved under these conditions. Analysis of the content of elemental sulfur in the leaching residues indicated that the fraction of sulfide sulfur determined as elemental sulfur was about 10% at 393 K, and that it decreased with temperature down to 0.5% at 453 K. Ultimate solid residues, which have been concentrated in silica and lead, can be oriented toward the lead smelter after alkali roasting-water washing pretreatment for metal recovery.
Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by sulphide minerals during grinding was investigated. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite ((Zn,Fe)S), and galena (PbS), which are the most abundant sulphide minerals on Earth, generated H2O2 in pulp liquid during wet grinding in the presence or devoid of dissolved oxygen in water and also when the freshly ground solids are placed in water immediately after dry grinding. Pyrite generated more H2O2 than other sulphide minerals and the order of H2O2 production by the minerals found to be pyrite > chalcopyrite > sphalerite > galena. The pH of water influenced the extent of hydrogen peroxide formation where higher amounts of H2O2 are produced at highly acidic pH. The amount of H2O2 formed also increased with increasing sulphide mineral loading and grinding time due to increased surface area and its interaction with water. The sulphide surfaces are highly catalytically active due to surface defect sites and unsaturation because of broken bonds and capable of breaking down the water molecule leading to hydroxyl free radicals. Type of grinding medium on formation of hydrogen peroxide by pyrite revealed that the mild steel produced more H2O2 than stainless steel grinding medium, where Fe2+ and/or Fe3+ ions played a key role in producing higher amounts of H2O2.
Furthermore, the effect of mixed sulphide minerals, i.e., pyrite–chalcopyrite, pyrite–galena, chalcopyrite–galena and sphalerite–pyrite, sphalerite–chalcopyrite and sphalerite–galena on the formation of H2O2 showed increasing H2O2 formation with increasing pyrite fraction in chalcopyrite–pyrite composition. In pyrite–sphalerite, chalcopyrite–sphalerite or galena–sphalerite mixed compositions, the increase in pyrite or chalcopyrite proportion, the concentration of H2O2 increased but with increase in galena proportion, the concentration of H2O2 decreased. Increasing pyrite proportion in pyrite–galena mixture, the concentration of H2O2 increased and also in the mixture of chalcopyrite–galena, the concentration of H2O2 increased with increasing chalcopyrite fraction. The results of H2O2 formation in pulp liquid of sulphide minerals and mixed minerals at different experimental conditions have been explained by Eh–pH diagrams of these minerals and the existence of free metal ions that are equally responsible for H2O2 formation besides surfaces catalytic activity. The results also corroborate the amount of H2O2 production with the rest potential of the sulphide minerals; higher is the rest potential more is the formation of H2O2. Most likely H2O2 is answerable for the oxidation of sulphide minerals and dissolution of non-ferrous metal sulphides in the presence of ferrous sulphide besides the galvanic interactions.
This study highlights the necessity of revisiting into the electrochemical and/or galvanic interactions between the grinding medium and sulphide minerals, and interaction mechanisms between pyrite and other sulphide minerals in terms of their flotation behaviour in the context of inevitable H2O2 existence in the pulp liquid.
This paper reports an acid leaching process study on the selective extraction of base metals from complex sulfidic, silicate-containing zinc ore containing sphalerite, hemimorphite, smithsonite, quartz, galena, pyrite, troilite, muscovite, calcium aluminate, calcite, gypsum and cerussite with traces of other minerals under oxygen pressure. The effects of influential factors (leaching temperature, sulfuric acid concentration, leaching time, ore particle size distribution, liquid/solid ratio and pressure) on the dissolution efficiencies of zinc, iron and silica from said ore were investigated. The optimum conditions, which were sought for reducing downstream purification requirements by limiting iron and silica dissolution, were found to be a leaching temperature of 160°C, a liquid/solid ratio of 6mL/g, a sulfuric acid concentration of 0.42mol/L, a leaching time of 150min, a particle size of −98+74μm and 1.6MPa partial pressure. Under these conditions, the zinc extraction was above 97%, with an iron and silica dissolution of less than 25% and 0.3%, respectively. The reduction of silica dissolution during leaching also promotes easier solid–liquid separation.
Effects of particle size of the zinc sulfide concentrate, leaching temperature, solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied. The results indicate that the additive can improve the reaction kinetics and the conversion rate. And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur. The reasonable experiment parameters are obtained as follows: the leaching temperature 150 °C, oxygen partial pressure 1 MPa, additive amount 1%, solid-to-liquid ratio 1:4, leaching time 2 h, initial sulfuric acid concentration 15%, and particle size less than 44 μm. Under the optimum conditions, the leaching rate of the zinc can reach 95% and the reduction rate of the sulfur can reach 90%.
Indium was recovered from zinc oxide flue dust (ZOFD) with sulfuric acid by oxidative pressure leaching in an autoclave, and the effects of different technological conditions on indium leaching were studied. Potassium permanganate and hydrogen peroxide were used as oxidants. The atmospheric pressure leaching experiments were also carried out. The experimental results show that the leaching rate of indium can be effectively improved by oxidative pressure leaching. The optimum conditions of pressure leaching are determined as sulfuric 5.10 mol/L acid, leaching time 150 min, temperature 90 °C, and the H2O2 dosage of 0.5 mL/g or 2.5% KMnO4. The leaching rate of indium is more than 90%, which is increased by 13% compared with that of atmospheric pressure leaching process without oxidant under the optimum conditions.
An attempt was made to investigate the catalytic role of manganese in enhancing the reaction rate with respect to autoclave oxidation of germanium-rich sphalerite concentrates. A series of batch experiments was performed in different conditions to investigate the variables such as temperature and oxygen pressure. Experimental results obtained show that as a catalyst, aqueous divalent manganese can accelerate the leaching of sphalerite concentrates significantly.On a tenté d'étudier le rôle catalytique du manganèse pour améliorer la vitesse de réaction de l'oxydation en autoclave de concentrés de sphalérite riche en germanium. Une série d'expériences discontinues ont été menées dans différentes conditions afin d'étudier les variables telles que la température et la pression d'oxygène. Les résultats expérimentaux obtenus montrent qu'en tant que catalyseur, le manganèse divalent aqueux peut significativement accélérer la lixiviation des concentrés de sphalérite.
Currently, low-grade and complex ores and mining wastes can be processed economically by using bacteria in heap and agitation leaching processes. Bacterial leaching tests are performed on the run-of-mine ore which is a mixture of two different massive and dissemine copper ores, fed to Küre Copper Plant. In this leaching process, using "Acidithiobacillus ferrooxidans" culture, bacteria count, pH, copper and iron recoveries are monitored during the 576 hours of test period. By increasing the solid ratio (1 %→5 %) the oxidation ability of bacteria decreases, thus the leaching rate. Therefore copper and iron recoveries decreased from 68 %, 35 % and 45 %, 20 %, respectively. As a result of laboratory tests, it is found that as the pulp density increased, the efficiency of copper recovery decreased using this bacterial culture.
Bacterial leaching of minerals is a simple, effective and environmental by benign technology in the treatment of sulphidic ores. This method has been successfully applied for the recovery of copper, gold and uranium in commercial scale for the past 25 years. Efficiency and cost-effectiveness of the bacterial leaching process depend mainly on the activity of bacteria and mineralogical and chemical composition of the ores. Bacterial leaching is based on the activity of mesophilic iron- and/or sulphur-oxidizing bacteria, notably Acidithiobacillus ferrooxidans, Acidithiobacillus thiooxidans and Leptospirillum ferrooxidans. These bacteria oxidize metal compounds to water soluble metal sulphates by a series of biological and chemical oxidation reactions occurring in leaching medium. After the isolation of above bacteria from acidic mine drainage waters, two oxidation mechanisms (direct and indirect bacterial leaching) have been discussed as related to oxidation/leaching of sulphidic ores in leaching systems. Fully understanding the bacterial leaching mechanisms of sulphidic ores improves the design and operation of bacterial leaching plants. In this article, the importance of various leaching mechanisms employed for metal recovery and their application aspects are critically reviewed with emphasis on copper, lead, zinc and nickel minerals.
A three-electrode system was adopted to investigate the corrosion current density and mixed potential of unstrained pyrite–galena and strained pyrite–galena galvanic cells in a flowing system. The results showed that when present in the same solution, strained pyrite produces a lower electrode potential than that of the galena electrode because of its strain energy; moreover, increasing the sodium sulfate solution concentration causes only slight fluctuations in the corrosion current density and mixed potential, while these values clearly increased with increasing ferric sulfate solution concentrations. In addition, for the sodium sulfate solution or ferric sulfate electrolyte, the faster the flow rate, the bigger the corrosion current density and the more positive the mixed potential of the galvanic cell. The experimental results are significant for hydrometallurgy and for controlling environmental pollution in mining activities. By using the galvanic model, mixed potential theory and the Butler–Volmer equation, the experimental results were explained theoretically.
The direct sulphation of individual and mixed sulphides of copper and iron has been studied using a steam–oxygen mixture as the oxidant. The investigation showed that copper sulphide formed copper sulphate with this gaseous mixture at 773 K, whereas iron sulphide converted mostly to hematite at this temperature. It was also observed that the mixture of copper–iron sulphides yielded a higher amount of copper sulphate (92.7%) than that obtained (40.0%) from copper sulphide alone in the presence of 10 wt.% ferric oxide. This was mainly attributed to an improved sulphatising environment during the roasting of mixed sulphides. The kinetics study of the CuS–FeS system with a steam–oxygen mixture showed that the copper sulphate formation followed the topochemical model. An activation energy value for this conversion was found to be 30.36 kJ/mol in the temperature range 673 to 773 K. The sulphides and different calcined products obtained were characterised by X-ray diffraction (XRD), Scanning Electron Microscopy (SEM) and the metallographic studies to examine the path of reaction.
The book covers the general principles of solution chemistry, engineering aspects, and detailed studies of hydrometallurgical processes in 750 pages fully illustrated with drawings and photographs. It contains a selected list of over 250 review articles, proceedings volumes, and books directly related to the subject. Emphasis is laid on chemical reactions, equipment used, and flowsheets.
An integrated pilot plant for the development of the S-C Copper Process was constructed at Fort Saskatchewan in 1975 and operated in 1976. The pilot plant was designed to treat 9 tonnes per day of a pyritic copper concentrate. A demonstration run conducted during the last quarter of 1976 averaged 90% on-stream time, and confirmed that the process is operable as a stable and integrated unit. Part 2 of this article describes the parameters and equipment developed for thermal activation of chalcopyrite copper concentrates to make them amenable for the dissolution of iron and zinc, and their subsequent separation and recovery. The resulting high-grade copper sulphide was leached with return electrolyte and oxygen to extract 98. 5% of the copper into solution. The copper was electrowon at 65 amps per square meter from purified solution, to produce cathodes analyzing 0. 1 ppm Se, 0. 1 ppm Bi, 4 ppm Pb and 10 ppm S.
Thermal analysis (TA) studies have been made to understand changes in mineralogy that occur when pure sulphide minerals, single concentrates and different Cu-Zn-Pb sulphide concentrates are heated in air. The observed sequential oxidation behaviour of bath the synthetic and natural sulphide minerals is characterised by TA traces that are reasonably reproducible. It is shown that oxidation mechanisms of synthetic analogues during roasting are similar to those of bulk concentrates. The oxidation data thus obtained on sulphide roast products are also found to be useful to characterise some partial leach residues. During ammonia leaching of bulk concentrates, the pyrite grains are practically inert and remain in the leach residue along with oxidated lead compounds and goethite. The TA data for roasting of partially leached samples are compared with those for roasting of untreated feed concentrates. The oxidation behaviour of copper, zinc and lead mineral phases of leach residues is then characterised. The observations are corroborated by chemical phase analysis, X-ray diffraction (XRD) and optical microscopy.
Oxidation in an autoclave with sulfuric acid has been shown to extract 95 to 99 pct of the zinc present, produce a quantitative amount of elemental sulfur, and avoid loss of zinc due to ferrite formation.
Part 1 of this article describes salient features of the Sherritt-Cominco copper process, a hydrometallurgical process developed for refining copper from sulfide concentrates. For the most abundant and relatively refractory chalcopyrite concentrates, the first step of the process is thermal activation with hydrogen to decompose chalcopyrite and pyrite, commonly associated with it, into more reactive sulfides more amenable to leaching. The process was shown to be suitable for handling relatively low-grade concentrates and, in addition, it was capable of high recoveries of associated zinc, molybdenum, nickel and cobalt in saleable form.
The first ammonia leaching plants, applied to copper carbonate and native copper tailings in 1915, were followed more recently by research and development of flowsheets for ammonia leaching of sulfide concentrates. These were applied to two commercial plants. Anaconda's Arbiter Plant started up in 1974 with a design capacity of 36,000 tons/year of cathodes, to be produced by ammonia leaching with oxygen, followed by solvent extraction and electrowinning. The plant shutdown in late 1977 as a result of high maintenance and operating costs, partly due to harsh winters; to complications associated with sulfate disposal; and to changes in mineralogy. BHP's Coloso plant in Chile was designed to produce 80,000 tons/year of cathodes by leaching part of Escondida's concentrate production. Using a similar flowsheet but with air and low temperatures to avoid sulfate production, it started up in late 1994 and shutdown in mid-1998, after failing to reach cathode design capacity, and experiencing problems with its technology. The paper reviews the technologies and also alternative methods for overcoming the problems.
Sherritt's zinc pressure leaching process converts zinc sulphide concentrates directly to zinc sulphate solutions and elemental sulphur. This paper describes the process principles, results obtained from the pilot plant and integration of the process into Cominco's operations at Trail. Application to a grass-roots plant is compared with conventional roast-leaching.
First developed in the 1950s for treating base metal sulfides and refractory gold ores, pressure leaching technology underwent rapid commercial expansion during the 1980s. Initially used to treat nickel concentrates and mattes, pressure leaching has since been extended to the processing of zinc sulfide concentrates and sulfide-containing refractory gold ores. The adoption of pressure leaching in the zinc industry and of pressure oxidation in the gold industry is expected to continue throughout this decade.
The opportunities for pressure hydrometallurgy remain bright. New processes remain to be developed in the areas of sulfide pressure leaching, laterite/oxide treatment and in the environmental area of metallurgical dust treatment. Careful fundamental studies combined with technical development along a variety of fronts should result in the anticipated breakthroughs.
Electric arc furnace (EAF) dust is produced when iron and steel scrap is remelted in an electric arc furnace. There are still significant problems associated with the pyrometallurgical and/or hydrometallurgical processes for the treatment of this dust. In the present research, the dust was roasted with caustic soda at low temperatures. It was found that the zinc ferrite (ZnFe2O4) in the dust was converted into sodium zincate (Na2ZnO2) and iron oxide (Fe2O3). In the subsequent dilute caustic leaching process both the zincite (ZnO) and the sodium zincate were soluble and the hematite was relatively insoluble. After roasting and leaching, the zinc recovery was found to be about 95%, while the majority of the iron oxide remained in the leach residue. The lead, cadmium and chromium recoveries were approximately 85, 89 and 37%, respectively, for roasting tests in which moisture was added. Without a moisture addition in roasting, the lead recovery decreased to 63% while the chromium recovery increased to 81%. Based on the experimental results, a hybrid low temperature roasting and dilute caustic leaching process followed by zinc cementation and electrowinning is proposed and discussed. This proposed hybrid EAF dust treatment process provides some potential advantages in comparison to the other hybrid processes which have been described in the literature.
The newly developed toroidal fluidized bed reactor has potential for improving sulfide roasting efficiency due to its unique and good mass/heat transfer characteristics. To achieve effective roasting at high temperature using the toroidal fluidized bed reactor, engineering issues associated with sintering and ferrite formation are investigated in this study by roasting industrial sulfide concentrates of zinc and copper. Laboratory tests are conducted at 800–1100°C by using an electric tube furnace under controlled roasting conditions. The test program is based on statistical design, and the products are characterized by XRD, SEM and EDX. It is revealed that the oxygen concentration plays an important role in promoting roasting. Roasting temperatures higher than 950°C do not favor roasting conversion due to the sintering of sulfide particles which inhibits oxygen diffusion into the particle core. Close contact of the oxidized particles favors the atomic interdiffusion between particles, leading to the formation of undesired zinc ferrite. Roasting in toroidal fluidized bed reactor results in zinc calcines with increased surface area and reduced zinc ferrite formation, thus improving zinc recovery.
Recovery of copper, lead and zinc from complex sulphide concentrates with hydrometallurgical processes has been used as an alternative due to the technological and environmental impacts. In laboratory evaluation of the selective leaching, the metal values in the flotation concentrates were selectively recovered by sulphuric acid (H2SO4) and ferric sulphate (Fe2SO4)3. The experimental parameters studied were pulp density, temperature and time for leaching. From the experimental results, it is concluded that recovery of Cu and Zn from sulphide concentrates can be as high as 89% and 97%, respectively under laboratory conditions.
The effect of the mixture of surfactants nonylphenolpolyethylene glycol with molecular weight 900 (D1), dinaphthylmethane-4,4′-disulphonic acid (D2) and polyethylene glycol with molecular weight 400 (D3) on the processes taking place during the acid and neutral leaching of zinc calcine was investigated: dissolution of zinc and metal impurities from calcine; precipitation of Fe(OH)3 and AL(OH)3 and coagulation of colloidal silicic acid solutions. The influence of the surfactant mixture (D1+D2+D3) was determined by comparing the results obtained with and without surfactants. The following effects of surfactants were found:•The presence of surfactants causes a small decrease of zinc, copper and cadmium dissolution during the acid leaching of zinc calcine;•The surfactant mixture improves deposition of impurities during the neutral leaching stage;•The volume of Fe(OH)3 or Al(OH)3 precipitated in solutions increases in the presence of surfactants which improves the effect of both hydroxides as scavengers;•Surfactant mixture initiates silicic acid coagulation at lower pH values compared to experiments without surfactants providing higher purification of zinc sulphate solutions and better separation from the insoluble residue.
A study of the recovery of copper, cobalt, nickel and zinc from copper converter slag by roasting with ferric sulphate is reported. Roasting of converter slag with ferric sulphate, followed by leaching with water, was carried out in order to bring the metal values into solution. For 500°C roasting temperature, 120 min roasting time and Fe2(SO4)3 · xH2O/slag = 1 ratio, recoveries of copper, cobalt, nickel and zinc were about 93%; 38%; 13% and 59%, respectively. Higher extraction yields could be achieved with a higher ratio of Fe2(SO4)3 · xH2O/slag for copper and zinc, whereas cobalt and nickel could not be extracted in acceptable yields. Using H2SO4 in the leaching process markedly improves the metal recoveries.
The zinc-recovery results obtained in leaching tests run on a complex sulphide ore of Caribbean origin have been statistically processed. The raw material contained 6.95% Zn, 2.41% Pb and traces (0.05%) of Cu. The study was conducted by means of consecutive factorial plans which permitted determination of the manner in which the leaching was influenced by the following factors: (1) oxidizing agent (Fe2(SO4)3) concentration; (2) temperature; (3) ore particle size characteristics; and (4) solid/liquid ratio. A regression equation was calculated by the results obtained from the tests; this permitted the derivation of a response surface which provides zinc recovery levels in relation to the levels of the factors adopted in the experiments. It is important to understand the parameters which exert most influence on zinc recoveries during leaching, especially when the presence of finely distributed polluting agents in the useful phase, pyrite in this case, prevents zinc enrichment by the usual physical means.
Recent studies on the ferric chloride and cupric chloride leaching of chalcopyrite, galena and sphalerite are reviewed. Although the chloride leaching of chalcopyrite concentrates has been proven at the demonstration plant scale, the leaching reaction is difficult. In contrast, galena dissolves rapidly in FeCl3 media, and both the leaching rate and lead solubility increase significantly with increasing chloride concentration. The hydrometallurgical treatment of lead concentrates seems to be technically feasible. The sphalerite leaching rate is strongly affected by its solid solution iron content, and leaching processes for iron-rich zinc concentrates could probably be developed. The importance of intermediate sulphide phases and insoluble reaction products on the leaching of sulphides is discussed. In addition, it is postulated that at least part of the elemental sulphur reaction product is formed via the oxidation of dissolved H2S. The ability of chloride leaching processes to generate elemental sulphur while leaving pyrite largely unaffected makes them especially useful for the treatment of pyritic complex sulphides. Thus, recent chloride leaching technologies for complex sulphides are also reviewed.
Natural monoclinic pyrrhotite particles (Fe1−xS) were subjected to pressure leaching by oxygen in sulphuric acid solutions at temperatures ranging between 353 and 453 K (80–180°C). For temperatures below the melting point of sulphur (392 K), the rate of pyrrhotite oxidation shows a moderate dependence on temperature, while it is totally independent of sulphuric acid concentration. Nonetheless, in the absence of oxygen, as much as 30% of the mineral can be dissolved in 0.5 mol/l H2SO4. The conversion data were found to fit well to a shrinking-core model with mixed control by half-order surface reaction and oxidant diffusion though a product layer. Despite the high initial reactivity of pyrrhotite, complete oxidation of the mineral was never achieved at temperatures below 393 K, apparently due to an impervious sulphur product layer covering the particles. Complete pyrrhotite oxidation was achieved at temperatures above the melting point of sulphur and only with the use of lignin sulphonate as dispersant of molten sulphur. By analysing the conversion data with the shinking-core model, pyrrhotite oxidation in the high temperature range (403–453 K) was found to be surface-reaction controlled and of first order with respect to oxygen pressure.
Multi-mineral complex sulphide ores would, typically, not respond well to conventional single mineral flotation due to their fine grained nature. Thus, many complex ore deposits remain undeveloped due to a lack of alternative economic treatment scenarios. The selective leaching process that has been explored is capable of selectively extracting zinc from concentrates containing both sphalerite and chalcopyrite. Unfortunately, many complex sulphide concentrates may also contain appreciable levels of pyrite and galena. Understanding the influence of these mineral on the leaching system is critical to developing a selective leaching process that is easily adapted to a wide variety of concentrate materials.Based on these considerations the selective extraction of zinc using pressure oxidation from multi-mineral concentrates was investigated. Concentrates were artificially produced from pure minerals and ranged in composition of: 25–100% by weight sphalerite, 5–75% chalcopyrite, 0–75% pyrite and 0–75% galena. Zinc was selectively extracted from several Cu/Zn concentrates by utilizing various combinations of temperature and oxygen concentration in the pressure leaching process and the influence of the additions of pyrite and galena was defined. It was discovered that addition of only 5% by weight galena to the selective leaching system retarded the dissolution of sphalerite by up to 13%. Alternatively, a 10% pyrite addition increased zinc extractions while simultaneously decreasing the copper extractions, however, increased concentrations resulted in increased copper extraction and thus, decreased zinc selectivity.
The combination of roasting and pressure leaching is an alternative process that offers advantages over conventional processes because of the shorter leaching time and higher metal recovery. The copper and iron sulphide minerals examined in this study were chalcopyrite (CuFeS2) and pyrite (FeS2). The best results obtained were with a pre-treatment by roasting followed by acid pressure leaching in an autoclave system. The extraction of copper achieved was over 85%. Copper dissolution in this system is affected by particle size, leaching time and oxygen pressure. This paper presents the preliminary research on acid leaching of pyritic copper ore in an autoclave system under laboratory conditions.
Leaching of a copper residue, produced by selective oxidative leaching of a nickel matte, in oxygenated sulfuric acid solution in the presence of chloride ions was investigated. The leaching behaviors of copper and nickel in the copper residue were determined. The effects of chloride addition, oxygen flowrate, sulfuric acid concentration, and temperature were studied as leaching variables. Addition of chloride in small amounts into the leach slurry was found to enhance copper leaching from the residue.
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