Extraction of vanadium from black shale using pressure acid leaching

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 (FeSO₄) 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 Fe³⁺ 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 V₂O₅ powder. The overall yield of vanadium through all process stages was about 85%.     

Reductive leaching of manganese from low-grade manganese ore in H2SO4 using cane molasses as reductant

Manganese extraction from a low-grade ore was investigated using cane molasses as a reducing agent in dilute sulfuric acid medium. The effects of concentrations of cane molasses and sulfuric acid, leaching temperature as well as reaction time were discussed. The results showed that high manganese recovery with low Fe and Al extraction yield could be obtained by analyzing the leaching efficiencies of Mn, Fe and Al during the leaching process. The optimal leaching condition was determined as 1.9 mol/L H₂SO₄ and 60.0 g/L cane molasses for 120 min at 90 °C while using particles smaller than 0.147 mm. The leaching efficiencies were 97.0% for Mn, whereas 21.5% for Al and 32.4% for Fe, respectively.     

Solvent extraction of cadmium from sulfate solution with di-(2-ethylhexyl) phosphoric acid diluted in kerosene

In the present paper, solvent extraction process has been used for extraction of cadmium from sulfate solution using di-(2-ethylhexyl) phosphoric acid (D2EHPA) with 1% isodecanol in kerosene diluent expected from industrial effluents or leaching of ores/secondary materials. Different process parameters such as pH, contact time, extractant concentration, O/A ratio etc. were investigated. Results demonstrated that quantitative extraction of cadmium was feasible from 4.45 mM cadmium feed solution in single stage at equilibrium pH 4.5, time 2 min and O/A ratio 1:1 with 0.15 mM D2EHPA. The extraction mechanism of cadmium from sulfate solution by D2EHPA in kerosene could be represented at equilibrium by Cd²⁺ + 3/2 (H₂A₂)_org ⇔ CdA₂(HA)_org + 2H⁺. The loading capacity of 0.15 mM D2EHPA in sulfate solution was determined to be ∼ 8.9 mM cadmium. The loaded cadmium was effectively stripped using 180 g/L sulfuric acid. The metal or salt could be produced by electrolysis or crystallization from the stripped solution.     

Kinetics of silver leaching from manganese-silver associated ores in sulfuric acid solution in the presence of hydrogen peroxide

Kinetics of silver leaching from a manganese-silver associated ore in sulfuric acid solution in the presence of H₂O₂ has been investigated in this article. It is found that sulfuric acid and hydrogen peroxide have significant effects on the leaching rate of silver. The reaction orders of H₂SO₄ and H₂O₂ were determined as 0.80 and 0.68, respectively. It is found that the effects of temperature on the leaching rate are not marked, the apparent activation energy is attained to be 8.05 kJ/mol within the temperature range of 30 °C to 60 °C in the presence of H₂O₂. Silver leaching is found to be diffusion-controlled and follows the kinetic model: 1−2x/3−(1−x)^2/3=Kt. It is also found that particle size presents a clear effect on silver leaching rate, and the rate constant (k) is proportional to d ₋₂ ⁰ .     

Recovery of yttrium from deep-sea mud

Deep-sea mud rich in rare earth yttrium has received lots of attention from the international community as a new resource for Y. A novel process, which mainly includes acid leaching, solvent extraction, and oxalic acid precipitation-roasting, is proposed for recovery of Y from deep-sea mud. A series of experiments were conducted to inspect the impacts of various factors during the process and the optimum conditions were determined. The results show that the Y of deep-sea mud totally exists in apatite minerals which can be decomposed by hydrochloric acid and sulfuric acid solution. The highest leaching efficiency of Y is 94.53% using hydrochloric acid and 84.38% using sulfuric acid under the conditions of H~+concentration 2.0 mol/L, leaching time 60 min, liquid-solid ratio 4:1 and room temperature 25 ℃(only in case of sulfuric acid, when using hydrochloric acid, the leaching temperature should be 60 ℃). Because of the much lower leaching temperature, sulfuric acid leaching is preferred. The counter current extraction and stripping tests were simulated by a cascade centrifugal extraction tank device. Using 10 vol% P204,15 vol% TBP and 75 vol% sulfonated kerosene as extractant, 98.79% Y~(3+) and 42.60% Fe~(3+) are extracted from sulfuric acid leaching liquor(adjusted to pH = 2.0) after seven-stage counter current extraction with O/A ratio of 1:1 at room temperature, while other metals ions such as Al~(3+), Ca~(2+), Mg~(2+)and Mn~(2+) are almost not extracted. The Y~(3+) in loaded organic can be selectively stripped using 50 g/L sulfuric acid solution and the stripping efficiency reaches 99.86% after seven-stage counter current stripping with O/A ratio of 10:1 at room temperature, while only 2.26% co-extracted Fe~(3+) is stripped. The Y~(3+) of loaded strip liquor can be precipitated by oxalic acid to further separate Y~(3+) and Fe~(3+). The precipitation efficiency of Y~(3+) in loaded strip liquor can be 98.56% while Fe~(3+) is not precipitated under the conditions of oxalic acid solution concentration 200 g/L, quality ratio of oxalic acid to Y of 2, and precipitation time 0.5 h. And the precipitate was roasted at 850 ℃ for 3 h to obtain the oxide of Y in which the purity of Y_2 O_3/REO is 79.02% and the contents of major non-rare earth impurities are less than 0.21%.Over the whole process included acid leaching, solvent extraction, and oxalic acid precipitation-roasting,the yttrium yield is 82.04%.     

Leaching of silica from vanadium-bearing steel slag in sodium hydroxide solution

The work aims to selectively extract silica from vanadium-bearing steel slag by a leaching process. The effects of the particle size, the ratio of solid to liquid, the concentration of sodium hydroxide solution and the leaching temperature on the leaching behavior of silica from vanadium-bearing steel slag were investigated. The leaching kinetics of silica from vanadium-bearing steel slag in 30-50% w/w NaOH solutions was studied at 240 °C and the shrinking-core model was established to express the leaching kinetics of silica. The data showed that the leaching rate was controlled by the chemical reaction on the system interface and the activation energy for the process was found to be 36.4 kJ mol^− 1. By the leaching process, the majority of silica could be removed effectively from the vanadium-bearing steel slag and a residue with a low SiO₂ content of 4.28% and a high V₂O₅ content of 11.15% was obtained. Under these conditions there was partial dissolution of Al and slight dissolution of Cr, Mn and Ti.     

Extraction of selenium from copper anode slimes in a sealed leaching system

A new method was proposed for extracting selenium from copper anode slimes with a low concentration of nitric acid in a sealed sulfuric acid leaching system. It is performed under an atmosphere of oxygen which allowed for a cyclic utilization of nitric acid. The effects of main parameters on selenium leaching were studied. The mineralogical characterizations of the typical samples were investigated by XRD and SEM. The results showed that the optimal conditions of the process are considered to be total gas pressure of 0.1 MPa, leaching temperature of 388 K, solid-liquid ratio of 0.20 g mL^–1, H₂SO₄ concentration of 2 mol L^–1, HNO₃ concentration of 0.07 mol L^–1 and leaching time of 2 h. The high selenium leaching efficiency of 99.23% was obtained under these conditions. According to the results of XRD and SEM-EDS, Cu–Ag selenide in the raw anode slimes is difficult to be leached with sulfuric acid alone; copper can be leached more easily from Cu-Ag selenide than silver; selenide is oxidized into the solution, undergoing the intermediate product of elemental selenium.     

Kinetics of leaching selenium from Ni-Mo ore smelter dust using sodium chlorate in a mixture of hydrochloric and sulfuric acids

The kinetics of leaching selenium from Ni-Mo ore smelter dust in H₂SO₄-HCl-H₂O system was investigated. The effects including leaching temperature and time, particle size of the smelting dust, stirring speed, acid concentration and the coefficient β (the molar ratio of sodium chlorate to selenium in the smelter dust) on leaching of selenium were studied. The results indicated that the leaching of selenium increased sharply with the increase of temperature. The leaching of selenium reached 98% at 95 °C and stirring speed of 350 rpm for 150 min with the particle size of − 0.15 mm, initial [H⁺] concentration of 8 mol/L, the solid/liquid ratio of 1:5 g/mL and the coefficient β of 3.33. The leaching process was controlled by the surface chemical reaction and the kinetics of leaching selenium from Ni-Mo ore smelter dust followed the model of “shrinking core”. The apparent activation energy of leaching selenium was determined to be 44.4 kJ/mol, which was consistent with the values of activation energy reported for the surface chemical reaction control. The kinetics equation of leaching selenium from Ni-Mo ore smelter dust was expressed as , which coincided with the experimental results.     

Study of extraction and purification of Ni, Co and Mn from spent battery material

In spent battery material, there are plenty of valuable metals, such as copper, nickel, cobalt, manganese. Recovery of valuable metals from spent battery material not only protects the environment but also improves the utilization of resources and decreases the cost of battery material. In this study, hydrochloric acid is used as lixivant with characteristics of faster leaching rate and being recycled easily. The optimal conditions are that hydrochloric acid concentration is 6 mol/L, reaction temperature is exactly 60 °C, liquid/solid ratio is 8:1, (H₂O₂)_mol/(MeS)_mol = 2, and the leaching time is 2 h, the results show that the dissolution yields of Ni, Co and Mn can be 95 wt.% at least. The basic purification concept of the leaching solution includes that copper is removed through replacement by iron powder followed by iron precipitation in goethite method. The results show that Cu and Fe can be removed 99 wt.% at the least. At the same time, the loss of Ni, Co and Mn is not beyond 2 wt.%, 3 wt.％ and 2 wt.%, respectively. This method makes the preparation of pure Ni_xCo_yMn_z ternary system precursor economical. The process seems to be able to claim base metals from waste in a reliable and feasible way.     

Leaching of gold and palladium with aqueous ozone in dilute chloride media

The recycling of gold and palladium from metallic scraps can be carried out by ozone-leaching at ambient temperature and low (∼0.1 M) H⁺ and Cl⁻ concentrations. Rh and Pt remain un-reacted, whereas metals such as Cu, Ni, Ag, can be previously eliminated through O₂/H⁺ and O₂/O₃/H⁺ leaching pretreatments. Gold and palladium are dissolved in O₃/Cl⁻/H⁺ with formation of AuCl₄⁻ and PdCl₄²⁻. Leaching studies showed a passive region, basically located at < 0.01 and < 0.05 M Cl⁻ for Au and Pd, respectively. In the non-passive region, rates were only slightly dependent on either H⁺ and Cl⁻. Secondary formation of chlorine or hypochlorous acid was negligible at ≤ 0.1 M Cl⁻. Kinetics appeared to be controlled by mass transfer of O_3(aq) to the solid–liquid interface, showing first order dependency with respect to [O₃]_aq. Rates increased with temperature up to about 40 °C, but decreased at higher temperatures due to the fall in the O₃ solubility. The ozone mass transfer coefficients showed an activation energy < 20 kJ/mol. Gold leaching rate gradually diminished for pH > 2, as consequence of the influence of the [H⁺] on transfer control. The electric power consumption associated with O₃ generation was in the range 4–8 kWh/kg metal leached.     

Processing copper–vanadium precipitate formed from crude TiCl4 in titania and titanium sponge production

Copper wire is used to remove vanadium from crude TiCl₄ in titania and titanium sponge production which produces a copper–vanadium precipitate. The recovery of copper and vanadium from this precipitate was studied. Experiments found that the precipitate can be naturally oxidized by stacking for one month in air, converting > 90% metallic copper contained in the original precipitate into CuCl₂·2H₂O, Cu₂Cl(OH)₃ and Cu₂(OH)₃Cl. The copper oxy-chlorides were easily converted to Cu(OH)₂ by stirring in dilute NaOH at pH 11 and 80 °C under a liquid-to-solid ratio of 4:1. When the pH was lowered to about pH 2.5 by sulfuric acid, iron, titanium and vanadium oxides remained in the first acid leach residue and copper was selectively leached into solution. By evaporating and cooling the leach solution, a product of CuSO₄·5H₂O with 99.7% purity was obtained.To recover the vanadium, the filter cake was roasted with Na₂CO₃ at 700 °C for 3 h under the stoichiometric proportion of 2.5 for V. The calcine was then leached with water at 70 °C and NH₄VO₃ was precipitated by the addition of NH₄Cl. Calcination of NH₄VO₃ at 550 °C for 2 h produced V₂O₅ with a purity of 98.6%. After vanadium recovery, the residue was leached once again with sulfuric acid and the total recoveries of copper and vanadium were 98.6% and 95.7% respectively.     

Leaching behavior of ilmenite with sulfuric acid

A study of the rate of dissolution of ilmenite in sulfuric acid solutions has been carried out. The effects of temperature, particle size, stirring speed, and concentration of sulfuric acid on the rate of dissolution of ilmenite has been investigated. Temperature range studied in this investigation was 88° to 115°C, and the Arrhenius activation energy was found to be 64.4 kJ (15.4 kcal) per mole. The rate of dissolution increased with concentration of sulfuric acid up to about 14 M sulfuric acid and decreased beyond this concentration. The maximum recovery at 14 M H₂SO₄ can be explained partially by the fact that H⁺ ion concentration peaks at about this concentration. Furthermore, reaction products, TiOSO₄ and FeSO₄, cover the surface of ilmenite when high concentrations of sulfuric acid are used, while these products are dissolved in water and removed from the surface when diluted sulfuric acid is involved. Based on the results obtained in this study, it can be concluded that the overall leaching of ilmenite with sulfuric acid at 88° to 115°C is described best by surface chemical reaction limiting with an order of 0.55 with respect to sulfuric acid concentration.     

Leaching of vanadium from carbonaceous shale

The leaching of vanadium from carbonaceous shale using dilute H₂SO₄ was investigated, and the mechanism of leaching determined. The results showed that higher leaching efficiency of vanadium was obtained by increasing initial concentration of H₂SO₄, raising leaching temperature and prolonging leaching time. Addition of ammonium fluoride also enhanced recovery. A recovery of 92% was obtained using a liquid to solid ratio 4:1, initial H₂SO₄ concentration 18%, NH₄F addition 4.8 wt.% of carbonaceous shale, leaching temperature 95 °C and contact time 8 h; the recovery was only 56% without NH₄F. The presence of NH₄F enhanced the leaching of vanadiferous mica.     

Leaching of Chalcopyrite Concentrate with Hydrogen Peroxide and Sulfuric Acid in an Autoclave System

In this study, leaching of chalcopyrite concentrate was investigated in an autoclave system using hydrogen peroxide and sulfuric acid. By decomposition of hydrogen peroxide, the active oxygen formed can provide both high oxidation potential and high pressure in a closed vessel for leaching. Preliminary studies showed that hydrogen peroxide can be used as an oxidant instead of oxygen gas in the autoclave. Central composite design (CCD) was used to examine the effects of the experimental parameters on the copper and iron extraction as a response. The proposed model equation using CCD showed good agreement with experimental data, the correlation coefficients R ² for copper and iron being 0.84 and 0.86, respectively. The optimum conditions to obtain the main goal of maximum copper and minimum iron extraction from chalcopyrite were determined as to be sulfuric acid concentration of 2.5 M, hydrogen peroxide concentration of 2.3 M, leaching time of 24 minutes, chalcopyrite amount of 3.17 g (in 50-mL solution), stirring speed of 630 rpm, and leaching temperature of 351 K (78 °C). Under the optimum condition, 76 pct of copper and 9 pct of iron were extracted from chalcopyrite concentrate. Extraction yield results of metals indicate that selective leaching of chalcopyrite can be achieved using hydrogen peroxide and sulfuric acid in an autoclave system.     

Leaching of oxidized zinc ores in various media

The leaching of various oxidized zinc ores in different solvents has been studied in laboratory tests by agitation leaching and in some cases by percolation leaching. Some tests have also been carried out on synthetic zinc compounds.The comparison of the leaching of four different Belgian ores in sulfuric acid, sulfurous acid, ammonium hydroxide and sodium hydroxide, shows that the better results are obtained with sulfuric acid and with caustic soda when the concentration of the latter is high enough. All the ores respond very well to leaching except one which is very rich in iron and can be processed only with concentrated caustic soda. Mineralogical studies have shown that smithsonite is completely leached but that hemimorphite is refractory to leaching in any solvent studied.The leaching in alkaline media (ammonia, caustic soda, diethylenetriamine) has been studied on five different ores. It is confirmed that hemimorphite is difficult to dissolve. Good results are obtained only with caustic soda at high temperature. In ammoniacal solutions, the presence of ammonium carbonate enhances the leaching of hemimorphite but decreases the leaching of smithsonite. The optimum concentration of diethylenetriamine is 250 g/l.Caustic soda leaching tests were carried out on pure compounds and showed that ZnO, Zn(OH)₂, PbO, PbCO₃ and 2PbCO₃βPb(OH)₂ dissolve very rapidly and completely, that the leaching of Zn₂SiO₄ is controlled by kinetics and that the leaching of ZnS, PbS and PbSiO₄ is very poor, their solubility product being very quickly attained.The percolation leaching by sulfuric acid has been tested on Belgian ores and the good results obtained support the idea that this kind of ore could be successfully treated in this way.     

Rare earth double sulfates from pre-concentrated bastnasite

A mixture of rare earth double sulfates was produced from a Turkish bastnasite-containing pre-concentrate (low grade concentrate) by sulfuric acid baking, subsequent water leaching and precipitation with sodium sulfate. The results of acid baking and leaching indicated that recoveries of rare earth elements up to 90% were readily obtained and the recovery of hydrofluoric acid as a by-product was also possible. Reasonable decontamination of the rare earth double sulfate salt from impurities such as Th, Fe, Al and Mg was possible by rapid precipitation at 50 °C using 1.25 times the stoichiometric amount of Na₂SO_4. The total rare earth double sulfate content (TREDS) was > 90% and analysed 17.3% La, 15.6% Ce, 3.2% Nd, 1.1% Pr, 0.3% Sm, 0.03% Eu, 0.01% Yb and 0.02% Y together with about 0.7% Ca, Fe, Al and other impurities.     

The cyanidation of silver metal: Review of kinetics and reaction mechanism

Published rate data are analysed for the chemical and electrochemical dissolution of silver metal from rotating discs in aerated/oxygenated cyanide solutions at ≈25 °C, pH 11 and different partial pressures of oxygen. The current status of the reaction mechanism is also reviewed. Speciation analysis of 0.01 mM silver(I) in 1–100 mM cyanide solutions shows that Ag(CN)₂⁻ is the predominant complex (50%) at cyanide concentrations < 20 mM. However, at higher cyanide concentrations, Ag(CN)₃²⁻ (up to 60%) and Ag(CN)₄³⁻ (up to 10%) can be formed. Thus, it is important to consider a silver(I) : cyanide ion ratio of 2 or 3 in the Levich equation to calculate the diffusion coefficient of cyanide ion. Likewise, it is important to consider a silver(I) : oxygen ratio of 1 : 0.5 to calculate the diffusion coefficient of oxygen. This indicates the reduction of oxygen to hydrogen peroxide in the surface reaction. Analysis of exchange current density data for silver oxidation as a function of cyanide concentration shows the involvement of between 1 and 2 cyanide ions in the surface reaction. The limiting rate of silver dissolution at high cyanide concentrations (2.5 × 10^− 5 mol m^− 2 s^− 1 at ≈21 kPa oxygen pressure) represents the maximum surface coverage by cyanide. This value is in close agreement with the rate constant of the surface reaction 4 × 10^− 5 mol m^− 2 s^− 1 based on the pure kinetic current of the mixed “charge transfer plus diffusion” model proposed by Li and Wadsworth [Li, J., Wadsworth, M.E., 1993. Electrochemical study of silver dissolution in cyanide solutions. J. Electrochem. Soc. 140, 1921–1927].     

Kinetics of copper dissolution during pressure oxidative leaching of lead-containing copper matte

The kinetics of pressure oxidative leaching of lead-containing copper matte with sulfuric acid was investigated. The effects of particle size, leaching temperature, oxygen partial pressure and sulfuric acid concentration on the kinetics and mechanism of copper extraction were studied. It was found that the reaction kinetic model follows the shrinking core model of chemical reaction control and the apparent activation energy was determined as 39.1 kJ/mol. The order of the reaction with respect to total pressure was found to be 0.64. The kinetic equations for the effect of particle size, leaching temperature, total pressure and sulfuric acid concentration were obtained and a mathematical model of copper extraction from lead-containing copper matte was developed as:

This equation estimates the extraction of copper with very good agreement (r = 0.99) between the experimental and calculated values.     

Efficient removal of K2O and Fe2O3 impurities from bauxite tailings through active calcination combined with acid leaching

In the present study, direct acid and active calcination–acid leaching processes were employed to remove K₂O and Fe₂O₃ impurities from bauxite tailings. For direct acid leaching, the results indicated that Fe₂O₃ can be removed with the addition of 3?mol?L^?1 HCL, whereas K₂O cannot be removed even when the concentration of HCl is higher than 7?mol?L^?1. As the bauxite tailings were calcined at 400–700°C, both K₂O and Fe₂O₃ were removed at low HCl concentration (3?mol?L^?1), which was ascribed to the removal of hydroxyls and increased K activity, derived from X-ray diffraction, Fourier transform infrared spectroscopy and X-ray photoelectron spectroscopy.