Li Minting

Extracting vanadium from stone-coal by oxygen pressure acid leaching and solvent extraction

Abstract Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction. The mineralogy of the stone-coal from Tongren City of Guizhou Province, China, was investigated by various determination methods. The effects of leaching time, leaching temperature, leaching agent concentration, leaching L/S ratio, granularity of material, additive consumption were investigated based on the mineralogy. The results show that under the conditions of leaching time of 3–4 h, temperature of 150 °C, sulfuric acid consumption of 25%–30%, ratio of liquid to solid of 1.2:1, the granularity less than 0.074 mm, additive consumption of 3%–5%, and oxygen pressure of 1.2 MPa, and the vanadium leaching rate can be more than 92% by the method of two-step pressurized acid leaching. The powdery V2O5 product with 99.52% in V2O5 content is obtained by the flowsheet of acid recovery, removing iron by reduction process, solvent extraction, precipitating vanadium with ammonium water, and pyrolysis from the stone-coal oxygen pressure acid-leaching solution. The total recovery efficiency of vanadium is above 85%, which is more than 20% higher than that obtained in the conventional process. Furthermore, the new process does not cause air pollution since no HCl or Cl2 is released by calcination of the raw material.

Thermodynamics and mechanism of vanadium(IV) extraction from sulphate medium with D2EHPA, EHEHPA and CYANEX 272 in kerosene

Extraction of vanadium(IV) from sulphate acid solution was studied using organophosphorous-based extractants D2EHPA, EHEHPA and CYANEX 272 in kerosene. The different parameters affecting the extraction of vanadium(IV) under equilibrium conditions were separately investigated to elucidate the stoichiometry of the extracted species. The distribution ratio of vanadium increased with increasing equilibrium pH of the aqueous phase, concentration of the extractants and temperature. D2EHPA was found to be a stronger extractant, having greater pH functionality than EHEHPA and CYANEX 272. Extraction of vanadium(IV) by these organophosphorous-based extractants involved cation exchange mechanism, and the extracted species appear to be VOR2(HR)2 in the low equilibrium pH and VOR2 in the higher equilibrium pH, where HR refers to the three acidic extractants.

Recovery of vanadium from black shale

Abstract The recovery of vanadium from a black shale from Guizhou Province of China was performed using a three-step process, which consists of a leaching step in the H2SO4-HF-NaClO system under atmospheric pressure, the vanadium separation from leachate by solvent extraction and stripping, followed by precipitation of ammonium vanadate and thermal decomposition. Under the optimum leaching conditions of 100 g/L sulfuric, 15 g/L hydrofluoric acid, 1.5 g/L NaClO, 6 h leaching at 90 °C and oxygen partial pressure of 1 200 kPa, and the liquid-to-solid ratio of 2, about 91% vanadium extraction is achieved. The vanadium extraction yields of solvent and stripping are 99.83% and 97.66% when using 10% (volume fraction) P204, 5% TBP and 85% sulfonated berosene as organic phase and 15% (mass fraction) sulphuric acid as stripping agent. After thermal decomposition the purity of powder vanadium pentoxide products is 99.18% and the overall vanadium recovery is over 81% in the whole process.

Pressure acid leaching of black shale for extraction of vanadium

Abstract Extraction of vanadium from black shale was attempted in pressure acid leaching. The chemical components of the sample obtained from Guizhou Province of China show that it contains 3.258% V 2 O 5 , 52.880% SiO 2 and 16.140% Al 2 O 3 . Phase analyses of vanadium indicates vanadium mainly exists in the free oxide and mica. Vanadium contents in the two phases are 18% and 53%, respectively. The contents of V 3+ , V 4+ and V 5+ are almost equal. Under the optimum parameters of one-step leaching (reaction time of 3 h, sulfuric addition of 25%, temperature of 150 °C, liquid to solid ratio of 1.2 mL/g, catalyst (FeSO 4 ) addition of 5% and size of 85% particle 0.074 mm), about 77% of vanadium is recovered. After two-step countercurrent leaching, the leach recovery of vanadium can reach above 90%. Air replacing oxygen is completely feasible. The impurity metals can dissolve into solution in different degrees.

Kinetics of hydrothermal sulfidation of synthetic hemimorphite with elemental sulfur

Abstract The kinetics of hydrothermal sulfidation of synthetic hemimorphite with elemental sulfur was investigated. The effects of stirring speed (100-600 r/min), temperature (120-220 °C), sulfur dosage (mole ratio of 0.6-1.6), and particle size (48-246 μm) on the hemimorphite conversion rate were studied. The results show that the conversion rate of hemimorphite increases with increasing temperature and decreasing particle size. A stirring speed above 400 r/min and a sulfur dosage above 1 have very little effect on the conversion rate. The kinetics analyses of the experimental data under various conditions indicate that the sulfidation process is controlled by the chemical reaction at the early stage of process with activation energy of 50.23 kJ/mol, and then controlled by diffusion through the product layer with activation energy of 11.12 kJ/mol.

Reductive leaching of zinc and indium from industrial zinc ferrite particulates in sulphuric acid media

Abstract Zinc ferrite is the principal constituent in zinc neutral-leach residue (NLR) which is commonly treated by hot-acid leaching in electrolytic zinc plants. Reductive leaching of zinc ferrite with sphalerite concentrate as a reducing agent was performed. It was found that leaching of zinc ferrite in the presence of sphalerit concentrate was a viable process that effectively extracted zinc and indium and converted Fe 3+ into Fe 2+ at the same time. Reflux leaching tests by two stages were performed to achieve extractions of 98.1% for zinc and 97.5% for indium, and a Fe 2+ /Fe 3+ molar ratio of 9.6 in leach solution was also obtained. The leaching behaviors of other elements, such as iron, copper and tin were also studied. The results showed that iron and copper were completely leached, whereas tin presented lower extraction values.

Pressure-Leaching Behavior of Nickel from Ni–Mo Ore in Aqueous Oxygenated Media

Abstract Ni–Mo ore is a multi-metal complex and unique mineral resource that exists as a black shale. This ore contains more than 4 wt% Mo and at least 2 wt% Ni. Economic value is associated with deriving products from Ni–Mo ore that contains amorphous colloidal sulfides and that is highly active. Under low-temperature hydrothermal conditions (water at 120 °C or more) and in excess oxygen, amorphous sulfides are oxidized easily to metal-sulfate forms, which leads to nickel and molybdenum dissolution. In this study, the effects of agitation speed, temperature, oxygen partial pressure and particle size on the rate of nickel leaching were investigated. The leaching rate was nearly independent of agitation speed above 400 rpm and increased with a decrease in particle size. A temperature increase to 150 °C contributed significantly to the nickel leaching rate. Oxygen pressure influences nickel leaching, with an oxygen pressure of between 0.5 MPa and 0.7 MPa providing the greatest effect. The pressure-leaching behavior of nickel was fitted to a shrinking-core kinetic model. The mathematical analyses indicates that the dissolution process is chemical-reaction controlled during early dissolution, with an activation energy of 42.68 kJ/mol, and the reaction order with respect to the oxygen partial pressure was 0.79. Up to nearly 70 % nickel was dissolved into solution, and product layers of molybdenum and iron oxide were formed on the surface of the Ni–Mo ore particles, which prevented further nickel dissolution. Thereafter, nickel leaching was controlled by liquid-film diffusion, with an activation energy of 11.01 kJ/mol.

Hydrothermal Sulfidation of Cerussite with Elemental Sulfur

Hydrothermal sulfidation of pure cerussite with elemental sulfur has been reported in the present paper. The variables considered in the study were temperature, time, particle size and elemental sulfur fraction in the reacting mixture. Temperature was the most important factor, at temperatures above 180°C, PbS and PbSO4 were the only stable reaction products in the sulfidation of cerussite. The experimental data indicated that under the hydrothermal conditions with a particle size of -74+58 m and sulfur fraction in reacting mixture of 15% at 200°C for 120 min, and 92% of lead sulfidation rate was achieved.