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In this study, hydrometallurgical processes involving pressure acid leaching and solvent extraction were developed to aid recovery of indium from zinc slag, which is produced in the imperial smelting process. Four different acid leaching methods were studied, namely atmospheric leaching, atmospheric leaching with KMnO4, roasting-atmospheric leaching, and oxygen pressure leaching in a sulfuric acid medium. Oxygen pressure acid leaching is the most effective method for indium extraction, and 94.1% of indium was leached under the optimum conditions, i.e., 300 g/L H2SO4,oxygen pressure 0.4 MPa, liquid/solid ratio 10 mL/g, and temperature 100°C for 5 h. X-ray diffraction and scanning electron microscopy examination of the raw material and leaching residue samples indicated that the intermetallic compounds Cu5Zn8and Cu2Zn, metallic zinc, and iron in the raw material dissolved, leaving the insoluble components PbSO4 and Pb as the major compounds in the leaching residue. A 98.5% proportion of the indium in the leaching solution was selectively extracted with 30% bis(2-ethylhexyl) phosphate and 70% kerosene by three-stage counter-current extraction, and 99.5% of the indium in the loaded organic phase was stripped by 6 mol/L HCl through four-stage counter-current stripping. The overall recovery yield of indium through all processes was approximately 92%.

Hematite produced by the zinc industry exhibits a high sulfur content, being mainly contaminated by jarosite as the main sulfur-containing phase, and is thus of limited use for the steelmaking industry. Herein, we investigate the effects of parameters such as temperature (170–200 °C) and initial concentrations of sulfuric acid (0–19.6 g/L), sodium sulfate (0–35.5 g/L), hematite seeds (0–30 g/L), and Zn2+ (80–120 g/L) on the formation and conversion of jarosite during hematite precipitation from a ferrous sulfate solution, obtaining high-quality hematite with a low sulfur content. We reveal that increased temperature not only benefits jarosite conversion but also reduces the sulfur content of precipitates and enriches them in iron. Sulfuric acid concentrations of 4.9–19.6 g/L greatly decrease the fraction of precipitated iron and produce sulfur-rich precipitates containing significant amounts of both hematite and jarosite, whereas sodium sulfate increases the iron removal efficiency and the sodium content in hematite precipitates. Hematite seeds prompt the precipitation of hematite and affect the composition of the obtained precipitates, promoting the precipitation of hematite in preference to jarosite at relatively low temperatures. Finally, zinc sulfate concentrations of 80–120 g/L have a positive effect on iron removal as hematite.

Nickel was separated and precipitated with potassium nickel sulfate hexahydrate [K2Ni(SO4)2·6H2O] from acidic sulfate solution, a leach solution from molybdenum-nickel black shale. The effects of the potassium sulfate (K2SO4) concentration, crystallization temperature, solution pH, and crystallization 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, crystallization temperature of 10°C, solution pH of 0.5, and crystallization 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 recrystallization, 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 crystallized. A process flowsheet was developed to separate iron(III) and nickel(II) from acidic-sulfate solution.

The leaching residue in zinc smelting system is taken as the research object, and the experiment of leaching residue in reduction leaching under sulphuric acid–sulfur dioxide system is carried out. The effects of reaction temperature, liquid–solid ratio, initial sulfuric acid concentration, partial pressure of SO2, and stirring speed on leaching rate of medium leaching residue under sulfuric acid-sulfur dioxide reduction conditions were studied. The leaching rates of zinc 95.24%, iron 98.66%, indium 95.04%, and copper 0.036% in medium leaching residue can be obtained at the conditions of temperature of 105 °C, time of 4 h, liquid–solid ratio of 8 ml/g, initial sulfuric acid concentration of 120 g/L, SO2 partial pressure of 200 kPa, and stirring speed of 500 r/min.

The “CO2–NaOH” brine purification method and flocculation performance of polyacrylamide (PAM) flocculants in vacuum salt making brine purification process are reported in this paper. The scale-forming trends of the brine were assessed by the Langlier index (LSI), the Ryznar stability index (RSI) and the Larson-Skold index (LR). The thermodynamic calculation results and PAM flocculant coagulation mechanism show that removal of Ca2+ and Mg2+ ions and suspended solids (SS) from brine by the “CO2–NaOH” method and PAM flocculants is feasible. The total concentration of Ca2+ and Mg2+ ions was decreased from 486.1 mg/L to 3.41 mg/L, and SS was decreased from 20 mg/L to 5 mg/L at the optimal experimental condition of CO2 flow velocity of 0.04 m3/h, injection time of 30 min, equilibrium pH of 11.5, and the dosage of the FL2949 and AN945SH PAM flocculants of 1 mg/L. LSI, RSI and LR show that the scale-forming was slowed down by brine purification. PAM flocculants were not retained in the brine as confirmed by the analysis of total organic carbon and chemical oxygen demand, and the X-ray diffraction, Fourier transform infrared, and scanning electron microscopy electron dispersive spectroscopy measurements for the flocs sludge during the brine purification process, results shown that the flocs sludge is calcite CaCO3. PAM flocculants purification was applied in industrial testing and achieved deep brine purification to remove scale-forming ions and SS. Additionally, it was found that the “CO2–NaOH” brine purification method promotes the utilization of CO2.

Among the currently available processes, iron precipitation as hematite from hydrometallurgical process solutions is the most suitable from an environmental point of view. The crystallization and dissolution of ferrous sulfate play an essential role in the hematite process. In this work, the effects of the Zn2+ concentration, temperature and oxygen partial pressure on the ferrous ion concentration, solution pH, and the composition of filter residues and crystallization sediments were studied. The influence of the dissolution behavior of ferrous sulfate on the hematite process was explained based on crystallization theory. The results indicated that the ferrous ion concentration in the solution showed a “wave-like” change in the presence of Zn2+ due to the common ion effect and the isomorphism effect. The sulfur content in filter residues increased as the concentration of SO42− was increased by the addition of zinc sulfate and due to the increasing concentration of ferrous ions. Additionally, the maximum variation was obtained for the Fe content decreased from 11.52% to 7.3% and Zn content increased from 20.86% to 24.54% in the sediments.

The solvent extraction of zinc from ammoniacal/ammonium sulfate solutions using a mixture of β-diketone, 1-benzoyl-2 nonyl ketone (Mextral54-100, HA), and 2-hydroxy-5-nonylacetophenone oxime (Mextral 84H, HR) in sulphonated kerosene has been investigated. The effects of different ratio of extractants, extractants concentration, aqueous pH, total ammonia concentration and temperature on zinc extraction were examined. The results indicate that the mixture of Mextral54-100 and Mextral 84H shows evident synergistic effects on zinc extraction, and zinc is extracted as ZnAR by the mixed extractants. As the molar fraction of HR rises, the co-extraction ammonia content increases constantly and reaches a maximum value at χ(HR) 0.5. The extraction efficiency decreases dramatically with the increase of aqueous pH and total ammonia concentration, indicating that the formation of zinc amine complexes depresses the ligand exchange reaction. The thermodynamic parameter ΔH shows that the extractions of zinc by both extraction systems are exothermic driven. FT-IR studies confirm that the ammonia molecules can be co-extracted into synergistic adducts.

Indium and zinc extraction from an indium-bearing zinc residue were investigated using sphalerite concentrate as a reductant in sulfuric acid medium. The effects of amount of sphalerite concentrate, sulfuric acid concentration, particle size, leaching time as well as temperature were discussed. The results showed that high indium and zinc extraction yield as well as high Fe 2 + /Fe 3 + molar ratio could be obtained by reduction leaching of zinc residue with addition of sphalerite concentrate as a reductant. The optimal leaching condition was determined as 150 g/L H 2 SO 4 , 0.95 times of theoretic amount of sphalerite concentrate for 4 h at 90 °C while using particles in the range of 74–58 μm. The leaching efficiencies were 94.8% of indium, 96.1% of zinc and 92.8% of iron, respectively, and a Fe 2 + /Fe 3 + molar ratio of 7.5 in the leach solution was also obtained. The process above was a viable method that effectively extracted zinc and indium and converted Fe 3 + into Fe 2 + at the same time.

In this study, the enhanced extraction of zinc from a blast furnace dust (BFD) ammonia leaching solution via carbon dioxide (CO2) was investigated. 1-Benzoyl-2-nonyl ketone (commercially known as Mextral54–100) and tri-alkyl phosphine oxide (TRPO) were used as the extractant and synergistic extractant, respectively. The effects of different operational factors, such as the CO2 pressure, extractant concentration, phase ratio, and extraction temperature, were examined to optimize selective and quantitative zinc extraction. The loaded organic phase was stripped using sulfuric acid and waste electrolyte. Quantitative zinc extraction was achieved in two stages using 1.2 mol/L Mextral54–100 and 0.2 mol/L TRPO at an initial pH of 8.0, CO2 at 0.2 MPa, and an aqueous:organic (A:O) ratio of 2:1. The loaded organic phase, which contained 0.08 mol/L zinc, could be stripped in a single stage using 1.6 mol/L H2SO4 at a phase flow ratio of A/O = 1:1 to create a loaded strip solution that contained 5.5 g/L zinc. Zinc extraction from BFD with the ammonia leaching solution was enhanced by CO2 addition. Fourier-transform infrared spectroscopy confirmed that CO2 could prevent the co-extraction of ammonia into the synergistic system of the Mextral54–100/TRPO mixture. Based on the experimental results, a flowsheet for the recovery of zinc from BFD was developed and tested. The extents of zinc leaching, extraction, and stripping were ~ 91%, 97%, and 99%, respectively.

The extracted vanadium residues of vanadium titanomagnetite is one of the most hazardous residues that have a large annual output. In the paper, vanadium was leached by composed leaching reagent of sulphuric acid, hydrofluoric acid and potassium permanganate from the residues. The process parameters such as reagent concentration, ratio of liquid to solid, temperature, leaching time, which impact on the vanadium leaching percent, were investigated. The results show that HF can destroy the silicate phase which surrounds the vanadium-iron spinel. The addition of KMnO4 can enhance the oxidation transformation of insoluble vanadium (trivalence) to soluble vanadium (tetravalence) or vanadium (pentavalence), thus improve the vanadium leaching percent to 85%, under the reaction conditions: granule size of 0.15–0.25 mm, 125 g L−1 H2SO4, 30 g L−1 HF, 3 wt% KMnO4, liquid-to-solid ratio of 5 mL g−1, at 90 °C for 4 h.

Solvent extraction and separation of indium and copper from zinc residue reductive leach liquor containing a high concentration of zinc (ca. 85 g/L) and iron (ca. 35 g/L) was studied. Di-(2-ethylhexyl) phosphoric acid (D2EHPA) and Acorga M5640 were used as extractants for indium and copper, respectively. Indium was preferentially extracted by 15%(v/v) D2EHPA with three-stage extraction using an aqueous/organic (A/O) phase ratio of 6/1. Indium was effectively stripped from the D2EHPA organic phase by 4 M HCl. Copper was subsequently extracted selectively from the indium-extracted solution using 15%(v/v) Acorga M5640 with three-stage extraction using an A/O phase ratio of 2/1. Copper in the loaded Acorga M5640 organic phase was stripped by a synthetic spent electrolyte solution containing 25 g/L Cu and 180 g/L H 2 SO 4 . Based on the experimental results, a flowsheet for the separation and recovery of indium and copper from the reductive leach solution was developed and tested. The results showed that 95.4% of indium and 95.5% of copper were extracted, and the measured separation factors β In/Zn , β In/Cu , β In/Fe , β CuZn , and β CuFe were 4979, 1975, 3719, 1719, and 2446, respectively.

A direct solvent-extraction process was developed to selectively recover indium from solutions generated by reductive leaching of zinc residues. This approach avoids the traditional steps of intermediate precipitation, solid–liquid separation, and re-leaching. Copper in the leach solution is easily removed by cementation with iron powder and the remaining Fe(III) is reduced to the ferrous state, which avoids the possibility of its co-extraction with indium. Indium is effectively extracted from the zinc sulfate solution by 20% (v/v) di(2-ethylhexyl)phosphoric acid (D2EHPA) dissolved in kerosene at an initial pH of 0.5 and aqueous-to-organic phase ratio (A:O) of 6:1 using three countercurrent stages. Indium extraction is 96.1%, zinc and iron are barely extracted, and the separation factors of indium with respect to zinc and iron are 3640 and 4809, respectively. The complete stripping of indium from the loaded organic phase is achieved using 4 mol/L HCl at an A:O of 1:6. A scheme for direct solvent extraction of indium in zinc hydrometallurgical processing is suggested, by which indium can be concentrated into a small volume of strip solution containing 11 g/L of indium, which is 85 times its concentration in the feed solution.

This paper reports on continuous testing of a process for the recovery of vanadium from a H2SO4-HF solution, generated from the leaching of stone coal, by solvent extraction using a D2EHPA/TBP mixture as the extractant. Three unit processes are included: fully continuous countercurrent solvent extraction and stripping; oxidation and precipitation; and calcination. The solvent-extraction circuit was operated for 96 h. With six stages of countercurrent extraction, 97.7% vanadium, 98.3% molybdenum, and 97.9% titanium were extracted, while the sodium, potassium, fluorine, and nickel impurities were barely extracted. Scrubbing removed co-extracted and aqueous-entrained zinc, magnesium, silicon, and arsenic impurities. Most co-extracted impurities were scrubbed in two stages. Using five stages of stripping, more than 99.8% of the vanadium was stripped, yielding a vanadium concentration in the loaded strip solution of 34.1 g/L. Co-extracted molybdenum(VI), titanium(IV), and iron(III) remaining in the stripped organic phase were removed by a saturated sodium carbonate solution and the regenerated organic was reused for vanadium extraction. Vanadium was selectively precipitated with 98.5% efficiency from the loaded strip solution by oxidation and precipitation with ammonium sulfate. A high-purity V2O5 (99.61%) product was successfully produced by this flowsheet.

Cooling crystallization was used to separate iron (III) and nickel (II) from acidic sulfate solution produced by leaching of molybdenum-nickel black shale. The effect of K2SO4 concentration, crystallization temperature, solution pH and crystallization time on recovery of nickel (II) and precipitation of iron (III) was investigated, in order to obtain effective separation of nickel and iron. The optimum parameters determined were as follows: 200 g/L K2SO4, 10 °C crystallization temperature, solution pH of 0.5 and 24 h crystallization. Under these conditions, 97.6% nickel (II) was recovered as K2Ni(SO4)2·6H2O crystals and only 2.0% of the total iron (III) was precipitated. Recrystallization yielded K2Ni(SO4)2·6H2O crystals with a purity of 98.4%.

Ni–Mo ore is a black shale containing amorphous colloidal sulfides which have highly active. The oxygen pressure leaching behavior of nickel from black shale in aqueous media is presented. The effects of agitation speed, temperature, oxygen partial pressure and particle size on the rate of nickel leaching were determined. The results indicate that at low temperature hydrothermal and excess oxygen presence, the amorphous sulfides are easily oxidized to sulfuric acid and sulfate, and leaching nickel. The mathematical analyses of the experimental data for various experimental conditions indicated that the dissolution process was controlled by the chemical reaction during the early stage of dissolution, and was then controlled by liquid-film diffusion. In the initial stage of leaching, the reaction depends on oxygen partial pressure.

Iron precipitation occupies a vital position in the metallurgy industry, especially for treating iron-bearing sphalerite in hydrometallurgical processes. This paper emphasizes techniques for removing iron from high-iron sphalerite via hematite precipitation, and describes the results of research performed to examine ferrous oxydrolysis and precipitation. The behavior and mechanism of conversion between the iron phase and hematite residue at elevated temperature under pressure is ascertained by investigating the dissolving characteristic and thermodynamics stable area of hematite at sulfate system. The results show hematite precipitation went through the process of crystallization, dissolve, oxidation and precipitation of ferrous sulfate, and the overall iron precipitation was determined to be controlled by the rate of ferric sulphate hydrolysis rather than the oxidation of ferrous sulphate or the re-dissolution of ferrous sulphate crystals, or the transformation of basic ferric sulphate.

An improved analytical procedure has been described for the spectrophotometric determination of colloidal silica in leaching solution by means of adding fluoride to convert colloidal silica into reactive silica. The method is a variation of the classical molybdenum blue method, and the optimal wavelength, temperature, ratio of H3BO3/NH4F, pH and ammonium molybdate in the colorimetric process has been determined. The pH especially needs to be fairly strictly controlled. The linear equation is established as: C = 3.15956A − 0.51912, and R is more than 0.9992. Its feasibility and reproducibility have been confirmed by ICP-OES; the average relative error is 1.5%. The detection scope of silicon concentration extends almost 1000 times higher in a leaching solution, so the method in this paper can be used to detect leaching samples with a high silica concentration.

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.

Iron hydroxysulphate phases, such as hydronium jarosite and basic ferric sulphate, precipitate with hematite, which reduces the hematite quality and its potential for marketing. The factors of temperature, time, initial ferric and Zn2+ concentrations, and amount of hematite seed that have either positive or negative effects on metastable iron phase formation or conversion were studied. The temperature and initial ferric concentration have significant influence on hydronium jarosite formation in the hematite process. A large initial ferric concentration of 20 g/L can lead to the precipitation of basic ferric sulfate instead of hematite. The amounts of hematite, hydronium jarosite and basic ferric sulfate in the precipitates depend on the level of the initial ferric concentration and the presence/absence of Zn2+; purified hematite formation is favored at low initial ferric concentrations and in the presence of Zn2+. Metastable iron phases formed in the precipitation process convert to hematite to some extent with increasing temperature and reaction time.

The sample with smithsonite and cerussite as the main valuable metal minerals is a carbonate-hosted lead-zinc ore in the current study. Hydrothermal sulfidation of oxidized zinc-lead ore was carried out with a particle size of 74 to 58 μm and the mole ratio of sulfur and zinc of 2.0 at 453 K (180 °C) for 240 minutes, and 73 pct zinc and 86 pct lead sulfidation fraction were achieved. Flotation of the unsulfided sample was ineffective, with less than 55 pct recovery of zinc and lead. After sulfidation, the recoveries of zinc and lead in flotation concentrate were over 92 pct. A flotation concentrate was obtained with zinc and lead content of 41.2 pct and 22.1 pct, respectively.

Molybdenum and nickel in black shale ore exists as amorphous sulfides which has highly active, and easily oxide to sulfuric acid and sulfate at low temperature in the presence of excess oxygen. Take advantage of the characteristic to leach molybdenum and nickel from black shale. Oxidation, conversion and dissolution behavior of amorphous colloidal molybdenum and nickel sulfide in pressure leaching process was studied. The effects of stirring speed, temperature, concentration of sulphuric acid, oxygen process and mineral granularity on molybdenum and nickel leaching were investigated. The results showed that dissolution rate increased with increasing temperature and stirring speed, but decreased with increasing concentration of sulphuric acid. And dissolution of molybdenum and nickel from ore only need oxygen without any other reagent.

The recovery of vanadium from sulfuric and hydrofluoric mixed acid solutions generated by the direct leaching of black shale was investigated using solvent extraction and precipitation methods. The process consisted of reduction, solvent extraction, and stripping, followed by precipitation and calcination to yield vanadium pentoxide. The influence of various operating parameters on the extraction and recovery of vanadium was studied. Vanadium (IV) was selectively extracted using a mixture of 10% (v/v) di(2-ethylhexyl)phosphoric acid and 5% (v/v) tri-n-butylphosphate in sulfonated kerosene. Using six extraction and five stripping stages, the extraction efficiency for vanadium was 96.7% and the stripping efficiency was 99.7%. V2O5 with a purity of 99.52% was obtained by oxidation of the loaded strip solution and precipitation of ammonium polyvanadate at pH 1.8 to 2.2, followed by calcination of the dried precipitate at 550 °C for 2 h. It was concluded that the combination of solvent extraction and precipitation is an efficient method for the recovery of vanadium from a multi-element leach solution generated from black shale.

Selective extraction of Mo by Cyanex-272 and tributyl phosphate (TBP) was investigated from the leach liquor of a low grade Ni–Mo ore, which contains mainly Mo(VI), Fe(III), Ni(II) and a small quantity of other impurities. The elements in the leach liquor were divided into four groups based on the extraction mechanism of acid phosphorus extractants, and the behavior of four elements (Mo, Fe, As and V) was investigated. The separation factors of Mo to Fe, As and V were calculated separately using di-2-ethylhexylphosphoric acid, 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester and bis(2,4,4-trimethylpentyl)phosphine acid (Cyanex-272) under the same conditions. Cyanex-272 was found to be the optimum extractant. The synergistic coefficients of Cyanex-272 and co-solvents (secondary caprylic alcohol and TBP) were calculated separately for the extraction of Mo, Fe, As and V and TBP was found to be the better choice. Several operational factors such as initial leach liquor pH, extractant concentration, co-solvent concentration and stripping reagent concentration were examined to establish suitable conditions for the effective extraction and stripping of Mo. The separation factors, β Mo/Fe = 1803.61, β Mo/As = 480.09 and β Mo/V = 103.22, were obtained under the optimum experimental conditions of 15%(v/v) Cyanex-272, 15%(v/v) TBP, initial leach liquor pH of 0.0 and organic/aqueous (O/A) phase ratio of 1.0. The theoretical number of stages required for extraction and stripping was determined from the McCabe–Thiele diagrams. Molybdenum (11.62 g/L) was extracted in five stages with an O/A ratio of 1.0, yielding a raffinate with only 0.30 g/L Mo. Three stripping stages yielded a 98.0% Mo stripping efficiency from 11.23 g/L Mo loaded organic phase, producing an enriched 55.05 g/L Mo solution that contained 0.058 g/L As and 0.023 g/L V. A high purity final MoO 3 product (99%) was obtained by re-crystallization from the stripped Mo solution.

A D2EHPA and TBP mixed solvent system diluted with kerosene were used for the selective extraction of vanadium(IV) from iron(II) from an acidic stone coal leach solution that was reduced using sodium sulfite. Extraction studies were carried out under different pH and solvent concentrations, and optimized conditions were determined. The loaded organic was stripped with sulfuric acid. The number of stages required for the extraction and stripping of vanadium were determined from a McCabe-Thiele plot and confirmed by counter-current simulation studies. Results demonstrate that the extraction of vanadium increases as the initial pH and the D2EHPA concentration increases in the organic phase. A six-stage counter-current extraction simulation test was conducted over a period of 40 h at a initial pH of 2.48 with 10% (v/v) D2EHPA and 5% (v/v) TBP mixed extractant and resulted in a vanadium extraction of 97% for the feed solution containing 5.78 g/L V2O5 and 10.86 g/L total Fe. The loaded organic phase that contained 5.34 g/L V2O5 and 1.0 g/L Fe can be completely stripped by three-stage counter-current stripping with 1.5 mol/L H2SO4 at a phase flow ratio of O/A = 5:1 to give a strip solution containing 26.3 g/L V2O5 and 0.72 g/L Fe.

Pressure acid leaching of zinc silicate ores selectively separates valuable zinc from silica and iron, replacing conventional pyrometallurgical and hydrometallurgical methods. This study determines the optimum conditions for the dissolution of zinc silicate concentrates and achieves a higher zinc extraction (> 97%), lower iron (

The feasibility of vanadium recovery from black shale was investigated using sulfuric acid and hydrofluoric acid solutions. The chemical characterization of black shale from Guizhou, China showed that vanadium mainly existed in free oxide and mica. According to E –pH diagram for V–H 2 O at 25 °C, the dissolved species of vanadium in aqueous solution mainly present in form of VO 2+ and VO 2 + at high potential and low pH value. The effects of leaching time, sulfuric acid concentration, hydrofluoric acid concentration, sodium hypochlorite concentration, temperature and liquid to solid ratio on the vanadium leaching were studied. The optimum parameters obtained were a reaction time of 6 h, sulfuric 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), leaching temperature of 95 °C and a liquid to solid ratio of 4 mL/g. Under these conditions, up to 86% vanadium was recovered. HF has an important role in the leaching vanadium process. Compared to our previous two-step countercurrent leaching, the vanadium recovery only decreased 4% despite the leaching temperature was decreased from 150 °C for the previous tests to 95 °C for the present work.

The leaching of a sphalerite concentrate with high indium and iron content in sulfuric acid–oxygen media has been studied in the present paper. The variables considered in the study were particle size, concentration of sulfuric acid, partial pressure of oxygen, leaching temperature, and leaching time. The experimental data indicated that under the conditions employed up to 99% zinc and 90% indium extraction were achieved. The mineralogical analysis of the residue showed that the main minerals are elemental sulphur, unreacted pyrite and quartz, the amount of sulphide sulphur oxidized to sulphate during leaching is 81%.

The extraction of vanadium from black shale was attempted using pressure acid leaching. The effects of the several parameters which included reaction time, concentration of sulfuric acid, leaching temperature, liquid to solid ratio and concentration of additive (FeSO 4 ) upon leaching efficiency of vanadium were investigated and a two-step counter-current leaching approach was developed. The results showed that the leaching efficiency of vanadium in the two-step process could reach above 90%. Vanadium was effectively separated and enriched by solvent extraction after leachate pretreatments, including the reduction of Fe 3+ and adjustment of pH value. The extraction and stripping yields of vanadium were both > 98%. Ammonia was added to a stripping liquor to precipitate vanadium and then the ammonium poly-vanadate produced was calcined at 550 °C for 3 h to produce the high purity V 2 O 5 powder. The overall yield of vanadium through all process stages was about 85%.

Marmatite is an important resource of zinc, which exist in forms of high iron and high indium. In order to recycle the valuable metal from these ores, the testwork of simultaneous leaching on neutral residue mixed with zinc concentrate of high iron and high indium has been conducted. The primary objectives are to leach the zinc and indium remaining in the neutral leach residue as ferrite using acid, and to reduce the ferric iron in solution to the ferrous by zinc concentrate, thereby extracting the zinc and indium in the zinc concentrate. The results indicated that in the following conditions, mass ratio of neutral residue to zinc concentrate of 1:0.25, particle size of 74–58 µm, initial sulfuric acid concentration of 150 g/L, liquid to solid ratio of 8 L/kg, temperature of 90°C, reaction time of 4h, the leaching efficiency of zinc and indium were over 95%, nearly 90% Fe was Ferrous in the leaching solution. This process effectively extracted zinc and indium and converted Fe3+ into Fe2+ in solution, and simplified the traditional process of reduction after leaching.

Diffusion dialysis with a series of anion exchange membranes was used to recover H(2)SO(4) from an acid leach solution produced during the vanadium manufacturing process. The effects of sulfuric acid, FeSO(4) and VOSO(4) concentration, flow rate and flow rate ratio on the recovery of H(2)SO(4) were investigated. The results showed that sulfuric acid permeated well through the membranes used, while metal ions were efficiently rejected. The recovery of H(2)SO(4) increased as the sulfate concentration of the feed increased and the flow rate ratio of water to feed increased. More than 80% of the H(2)SO(4) could be recovered from the leach solution which contained 61.7 g/L free H(2)SO(4), 11.2 g/L Fe and 4.60 g/L V at a flow rate of 0.19x10(-3) m(3)/h m(2). V and Fe ion rejection were within 93-95 and 92-94%, respectively. A preliminary economic evaluation revealed that an investment in this process could be recovered within 27 months.