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Of the four sulphurica acid leaching methods studied - atmospheric, atmospheric with KMnO4, roasting-atmospheric and oxygen pressure - O2 pressure leaching proved most effective for In extraction, 94.1% of In being leached at an optimum 300 g/l H2SO4, 0.4 MPa O2, 10 ml/g liquid-solid ratio and 100 degrees C for 5 h. X-ray diffraction and scanning electron microscopy indicated that Zn, Fe and intermetallic CuZn compounds dissolved to leave PbSO4 and Pb as the major components of the leaching residue. Of the In in solution, 98.5% was selectively extracted with 30% bis(2-ethylhexyl) phosphate and 70% kerosene by 3-stage countercurrent extraction at an organic/aqueous phase ratio of 1/5, 99.5% of In in the organic phase then being stripped by 6 mol/l HCl through 4-stage countercurrent stripping at a 6/1 organic/aqueous phase ratio to give overall In recovery yield of c.92%. The process has the advantages of a short leaching time, high recovery rate and avoidance of harmful gas and solid discharges., Of the four sulphurica acid leaching methods studied - atmospheric, atmospheric with KMnO4, roasting-atmospheric and oxygen pressure - O2 pressure leaching proved most effective for In extraction, 94.1% of In being leached at an optimum 300 g/l H2SO4, 0.4 MPa O2, 10 ml/g liquid-solid ratio and 100 degrees C for 5 h. X-ray diffraction and scanning electron microscopy indicated that Zn, Fe and intermetallic CuZn compounds dissolved to leave PbSO4 and Pb as the major components of the leaching residue. Of the In in solution, 98.5% was selectively extracted with 30% bis(2-ethylhexyl) phosphate and 70% kerosene by 3-stage countercurrent extraction at an organic/aqueous phase ratio of 1/5, 99.5% of In in the organic phase then being stripped by 6 mol/l HCl through 4-stage countercurrent stripping at a 6/1 organic/aqueous phase ratio to give overall In recovery yield of c.92%. The process has the advantages of a short leaching time, high recovery rate and avoidance of harmful gas and solid discharges.

The ore contained 11.27% Zn, 80.14% of which was present in disseminated fine particles of carbonate minerals, and 3.45% Pb of which 75.63% was in carbonates; the main ore minerals included calamine, hemimorphite, zincit, sphalerite, hydrozincite, cerussite, galena and lead oxide along with haematite, limonite, dolomite, calcite and quartz. Autoclaving with elemental S on a laboratory scale hydrothermally sulphidised up to 75% of Zn and 82% of Pb, giving subsequent flotation recoveries of more than 90% into a concentrate grading 52% Zn and 37% Pb., The ore contained 11.27% Zn, 80.14% of which was present in disseminated fine particles of carbonate minerals, and 3.45% Pb of which 75.63% was in carbonates; the main ore minerals included calamine, hemimorphite, zincit, sphalerite, hydrozincite, cerussite, galena and lead oxide along with haematite, limonite, dolomite, calcite and quartz. Autoclaving with elemental S on a laboratory scale hydrothermally sulphidised up to 75% of Zn and 82% of Pb, giving subsequent flotation recoveries of more than 90% into a concentrate grading 52% Zn and 37% Pb.

Zinc metal was produced in the laboratory by the carbothermic reduction of sphalerite under vacuum in the presence of lime. Zinc recovery was almost 100% with a reaction temperature of 1 273-1 373 degrees K, initial C/ZnS and CaO/ZnS molar ratios of above 2.0 and 1.5 respectively, and a reaction time of 60 minutes. Base on XRD analyses, the presumed reaction sequence was ZnS(s, beta) to ZnS(s, alpha) to Zn(g)., Zinc metal was produced in the laboratory by the carbothermic reduction of sphalerite under vacuum in the presence of lime. Zinc recovery was almost 100% with a reaction temperature of 1 273-1 373 degrees K, initial C/ZnS and CaO/ZnS molar ratios of above 2.0 and 1.5 respectively, and a reaction time of 60 minutes. Base on XRD analyses, the presumed reaction sequence was ZnS(s, beta) to ZnS(s, alpha) to Zn(g).

The optimum conditions for the extraction of zinc from a zinc silicate ore by acid pressure leaching were determined in a laboratory study. The ore came from China's Yunnan province and contained willemite, hemimorphite, zincite, franklinite, as well as calcite, quartz and other minerals. The influences of particle size, sulphuric acid concentration, and leaching time, temperature and solid-liquid ratio were examined. The results showed that pressure acid leaching was effective for the recovery of zinc, with temperature and acidity being the most important variables. Raising the temperature from 80 degrees C to 160 degrees C reduced the dissolution of iron and silica but also reduced zinc extraction slightly. Increasing acid additions increased zinc recovery but also increased iron and silica dissolution. The particle size had little effect on both zinc recovery and iron and silica dissolution. 99% of the zinc could be extracted under the optimum conditions of 140 degrees C, 1.5 hr leach time, 0.46 mole/l acidity, liquid-solid ratio 6:1 and ore size less than 98 microns., The optimum conditions for the extraction of zinc from a zinc silicate ore by acid pressure leaching were determined in a laboratory study. The ore came from China's Yunnan province and contained willemite, hemimorphite, zincite, franklinite, as well as calcite, quartz and other minerals. The influences of particle size, sulphuric acid concentration, and leaching time, temperature and solid-liquid ratio were examined. The results showed that pressure acid leaching was effective for the recovery of zinc, with temperature and acidity being the most important variables. Raising the temperature from 80 degrees C to 160 degrees C reduced the dissolution of iron and silica but also reduced zinc extraction slightly. Increasing acid additions increased zinc recovery but also increased iron and silica dissolution. The particle size had little effect on both zinc recovery and iron and silica dissolution. 99% of the zinc could be extracted under the optimum conditions of 140 degrees C, 1.5 hr leach time, 0.46 mole/l acidity, liquid-solid ratio 6:1 and ore size less than 98 microns.

The separation of zinc and iron is essential in hydrometallurgical processes, especially for treating high-iron sphalerite. The haematite precipitation process for removing iron is an effective way to achieve the high-efficiency separation of zinc and iron. The authors studied the effect of temperature and time on the precipitation behaviour and characterised the precipitation products through X-ray diffraction, X-ray photoelectron spectroscopy, scanning electron microscopy and chemical analysis. The haematite precipitate contained more than 50% iron, less than 0.5% zinc, 0.1% arsenic and 5% sulphur; more than 95% K, 50% Na and 50-60% F were co-precipitated, and less than 1% Zn, Mg, Mn and Cl remained in the residue. Because of the uncontrolled supersaturation conditions, jarosite and goethite were generated. Extension of the reaction time and increasing the reaction temperature enabled conversion of most of the goethite and sodium jarosite to haematite during the haematite precipitation process. (Authors.), The separation of zinc and iron is essential in hydrometallurgical processes, especially for treating high-iron sphalerite. The haematite precipitation process for removing iron is an effective way to achieve the high-efficiency separation of zinc and iron. The authors studied the effect of temperature and time on the precipitation behaviour and characterised the precipitation products through X-ray diffraction, X-ray photoelectron spectroscopy, scanning electron microscopy and chemical analysis. The haematite precipitate contained more than 50% iron, less than 0.5% zinc, 0.1% arsenic and 5% sulphur; more than 95% K, 50% Na and 50-60% F were co-precipitated, and less than 1% Zn, Mg, Mn and Cl remained in the residue. Because of the uncontrolled supersaturation conditions, jarosite and goethite were generated. Extension of the reaction time and increasing the reaction temperature enabled conversion of most of the goethite and sodium jarosite to haematite during the haematite precipitation process. (Authors.)

Nickel was separated and precipitated with potassium nickel sulphate hexahydrate [K2Ni(SO4)2·6H2O] from acidic sulphate solution, a leach solution from molybdenum-nickel black shale. The effects of the potassium sulphate (K2SO4) concentration, crystallisation temperature, solution pH, and crystallisation time on nickel(II) recovery and iron(III) precipitation were investigated, revealing that nickel and iron were separated effectively. The optimum parameters were K2SO4 concentration of 200 g/l, crystallisation temperature of 10 degrees C, solution pH of 0.5, and crystallisation time of 24 h. Under these conditions, 97.6% nickel(II) was recovered as K2Ni(SO4)2·6H2O crystals while only 2.0% of the total iron(III) was precipitated. After recrystallisation, 98.4% pure K2Ni(SO4)2·6H2O crystals were obtained in the solids. The mother liquor was purified by hydrolysis-precipitation followed by cooling, and more than 99.0% K2SO4 could be crystallised. A process flowsheet was developed to separate iron(III) and nickel(II) from acidic sulphate solution. (Authors.), Nickel was separated and precipitated with potassium nickel sulphate hexahydrate [K2Ni(SO4)2·6H2O] from acidic sulphate solution, a leach solution from molybdenum-nickel black shale. The effects of the potassium sulphate (K2SO4) concentration, crystallisation temperature, solution pH, and crystallisation time on nickel(II) recovery and iron(III) precipitation were investigated, revealing that nickel and iron were separated effectively. The optimum parameters were K2SO4 concentration of 200 g/l, crystallisation temperature of 10 degrees C, solution pH of 0.5, and crystallisation time of 24 h. Under these conditions, 97.6% nickel(II) was recovered as K2Ni(SO4)2·6H2O crystals while only 2.0% of the total iron(III) was precipitated. After recrystallisation, 98.4% pure K2Ni(SO4)2·6H2O crystals were obtained in the solids. The mother liquor was purified by hydrolysis-precipitation followed by cooling, and more than 99.0% K2SO4 could be crystallised. A process flowsheet was developed to separate iron(III) and nickel(II) from acidic sulphate solution. (Authors.)

A feed solution was prepared from the leaching of stone coal with H2SO4+HF at atmospheric pressure and 90 degrees C for 6 h. The solution was pre-treated with 13.5 g/l Na2SO3 to reduce Fe(3+) to Fe(2+) and VO2(+) to VO(2+). The number of stages required for the extraction and stripping of V was determined from a McCabe-Thiele plot and confirmed in countercurrent simulation studies. V extraction increased with increasing initial pH and D2EHPA concentration in the organic phase. A six-stage countercurrent extraction simulation test conducted over 40 h at an initial pH of 2.48 with 10% v/v D2EHPA and 5% v/v TBP resulted in V extraction of 97% from a feed solution containing 5.78 g/l V2O5 and 10.86 g/l total Fe. The loaded organic phase containing 5.34 g/l V2O5 and 1.0 g/l Fe could be completely stripped by three-stage countercurrent stripping using 1.5 mol/l H2SO4 at an organic/aqueous phase flow ratio of 5:1 to produce a strip solution containing 26.3 g/l V2O5 and 0.72 g/l Fe., A feed solution was prepared from the leaching of stone coal with H2SO4+HF at atmospheric pressure and 90 degrees C for 6 h. The solution was pre-treated with 13.5 g/l Na2SO3 to reduce Fe(3+) to Fe(2+) and VO2(+) to VO(2+). The number of stages required for the extraction and stripping of V was determined from a McCabe-Thiele plot and confirmed in countercurrent simulation studies. V extraction increased with increasing initial pH and D2EHPA concentration in the organic phase. A six-stage countercurrent extraction simulation test conducted over 40 h at an initial pH of 2.48 with 10% v/v D2EHPA and 5% v/v TBP resulted in V extraction of 97% from a feed solution containing 5.78 g/l V2O5 and 10.86 g/l total Fe. The loaded organic phase containing 5.34 g/l V2O5 and 1.0 g/l Fe could be completely stripped by three-stage countercurrent stripping using 1.5 mol/l H2SO4 at an organic/aqueous phase flow ratio of 5:1 to produce a strip solution containing 26.3 g/l V2O5 and 0.72 g/l Fe.

The thermodynamics of the V-H2O and V-S-H2O systems at 298.15 K are presented in the form of activity-pH and potential-pH diagrams calculated from published standard Gibbs energies of the formation of the various species and phases involved. Revised Pourbaix diagrams of the V-H2O system show that V mainly exists as mononuclear ions at low concentrations in natural waters, while polyvanadates replace the mononuclear oxyanions as the predominant species in more concentrated solutions. Comparison of the potential-pH diagrams for the two systems indicates that SO4(2-) and HSO4(-) ions enlarge the regions of stability of V(3+), VO(2+) and VO2(+) ions, forming VSO4(+), VOSO4(aq) and VO2SO4(- ) in the pH range 0-5. The predominance areas for the V sulphates increase with a decrease in V activities. The activity-pH diagrams for each of the stable V oxidation states indicate that an increase in the activity of sulphuric acid from 0.1 to 1.0 results in an increase in the predominance areas for V sulphates and a decrease in the stability areas for solid phases such as V2O3, V2O4 and V2O5. The use is discussed of the predominance diagrams for predicting the optimum conditions and the parameters affecting V extraction processes, and reasonable explanations are provided for V precipitation and adsorption based on the thermodynamic values. Particular reference is made to the results of a previous study on V extraction using an ion exchange resin., The thermodynamics of the V-H2O and V-S-H2O systems at 298.15 K are presented in the form of activity-pH and potential-pH diagrams calculated from published standard Gibbs energies of the formation of the various species and phases involved. Revised Pourbaix diagrams of the V-H2O system show that V mainly exists as mononuclear ions at low concentrations in natural waters, while polyvanadates replace the mononuclear oxyanions as the predominant species in more concentrated solutions. Comparison of the potential-pH diagrams for the two systems indicates that SO4(2-) and HSO4(-) ions enlarge the regions of stability of V(3+), VO(2+) and VO2(+) ions, forming VSO4(+), VOSO4(aq) and VO2SO4(- ) in the pH range 0-5. The predominance areas for the V sulphates increase with a decrease in V activities. The activity-pH diagrams for each of the stable V oxidation states indicate that an increase in the activity of sulphuric acid from 0.1 to 1.0 results in an increase in the predominance areas for V sulphates and a decrease in the stability areas for solid phases such as V2O3, V2O4 and V2O5. The use is discussed of the predominance diagrams for predicting the optimum conditions and the parameters affecting V extraction processes, and reasonable explanations are provided for V precipitation and adsorption based on the thermodynamic values. Particular reference is made to the results of a previous study on V extraction using an ion exchange resin.

Experiments were carried out on a sphalerite concentrate in which the main In-bearing minerals were determined to be pyrite, zinckenite, galena, cassiterite and jamesonite. The kinetics of In dissolution from the concentrate in a sulphuric acid-oxygen system were studied. The results showed that reducing the particle size enhanced In dissolution but that the effect was minor at particle sizes below 58 micrometres. Increasing the temperature from 120 to 150 degrees C and the acid concentration from 1 to 2.5 mol/l both improved the leaching rate. Under optimum conditions, up to 92% In extraction was achieved. An empirical first-order kinetic model used to describe the dissolution kinetics showed that the process was controlled by chemical reaction with a calculated activation energy of 47 kJ/mol and a reaction order with respect to H2SO4 concentration close to 1., Experiments were carried out on a sphalerite concentrate in which the main In-bearing minerals were determined to be pyrite, zinckenite, galena, cassiterite and jamesonite. The kinetics of In dissolution from the concentrate in a sulphuric acid-oxygen system were studied. The results showed that reducing the particle size enhanced In dissolution but that the effect was minor at particle sizes below 58 micrometres. Increasing the temperature from 120 to 150 degrees C and the acid concentration from 1 to 2.5 mol/l both improved the leaching rate. Under optimum conditions, up to 92% In extraction was achieved. An empirical first-order kinetic model used to describe the dissolution kinetics showed that the process was controlled by chemical reaction with a calculated activation energy of 47 kJ/mol and a reaction order with respect to H2SO4 concentration close to 1.

The optimum conditions were investigated for the dissolution of zinc silicate concentrates in sulphuric acid pressure leaching. The results showed that the process is more efficient than other hydrometallurgical methods in relation to reduced dissolution of SiO2 and Fe and increased Zn extraction. Zn extraction of about 97% and silica and Fe dissolution rates of less than 0.77% and 4%, respectively, were obtained at a temperature of 140 degrees C, partial pressure of 1.4 MPa, solid to liquid ratio of 1:6 g/ml, sulphuric acid concentration of 0.44 mol/l, reaction time of 120 min, ore particle size of 98 micrometres and control of the final slurry pH in the range 1.3-1.5., The optimum conditions were investigated for the dissolution of zinc silicate concentrates in sulphuric acid pressure leaching. The results showed that the process is more efficient than other hydrometallurgical methods in relation to reduced dissolution of SiO2 and Fe and increased Zn extraction. Zn extraction of about 97% and silica and Fe dissolution rates of less than 0.77% and 4%, respectively, were obtained at a temperature of 140 degrees C, partial pressure of 1.4 MPa, solid to liquid ratio of 1:6 g/ml, sulphuric acid concentration of 0.44 mol/l, reaction time of 120 min, ore particle size of 98 micrometres and control of the final slurry pH in the range 1.3-1.5.

The recovery of V from black shale samples from Guizhou in China was investigated using sulphuric and hydrofluoric acid solutions. Chemical characterisation of the shale showed that V mainly existed as free oxide and in mica. Analysis of the potential-pH diagram for the V-H2O system at 25 degrees C showed that the dissolved V species present in aqueous solution were VO(2+) and VO2(+) at high potential and low pH values. The effects of process variables on the leaching rate were studied and the optimum conditions determined to be a reaction time of 6 h at 95 degrees C, sulphuric acid concentration of 87.5 g/l, hydrofluoric acid concentration of 15 g/l, sodium hypochlorite concentration of 1 g/l, added after a reaction time of 3 h, and a liquid/solid ratio of 4 ml/g. V recovery of up to 86% was obtained, only 4% less than for a two-step countercurrent leaching process which required a leaching temperature of 150 degrees C., The recovery of V from black shale samples from Guizhou in China was investigated using sulphuric and hydrofluoric acid solutions. Chemical characterisation of the shale showed that V mainly existed as free oxide and in mica. Analysis of the potential-pH diagram for the V-H2O system at 25 degrees C showed that the dissolved V species present in aqueous solution were VO(2+) and VO2(+) at high potential and low pH values. The effects of process variables on the leaching rate were studied and the optimum conditions determined to be a reaction time of 6 h at 95 degrees C, sulphuric acid concentration of 87.5 g/l, hydrofluoric acid concentration of 15 g/l, sodium hypochlorite concentration of 1 g/l, added after a reaction time of 3 h, and a liquid/solid ratio of 4 ml/g. V recovery of up to 86% was obtained, only 4% less than for a two-step countercurrent leaching process which required a leaching temperature of 150 degrees C.