The preparation of nickel, cobalt and manganese ternary precursors by coprecipitation method was studied. The effects of ammonia concentration, system pH, reaction time and reaction temperature on the properties of the precursors prepared by coprecipitation method were investigated. The results show that under the optimal conditions of ammonia concentration of 1 mol/L, pH=12, reaction temperature of 55 ℃ and reaction temperature of 28 h, the prepared precursor products have good particle size distribution, uniform spheroid shape and dense crystallization. The first discharge capacity of the cathode material after calcination is 169.14 mAh/g and the first efficiency is 88.32% under the charge and discharge system of 2.8~4.4 V and 0.2 C.

The separation and enrichment of precious metal gold in chloride system with synergistic extraction system composed of DIBK (HA) and TBP (B) was stuied. The effects of extraction system and composition, extraction time, extraction temperature, aqueous chloride ion concentration,and extraction ratio (V_O/V_A) on the separation performance of gold extraction was investigated, and the extraction mechanism was preliminarily explored by slope method. The results indicate that under the conditions of synergistic extraction system of TBP-DIBK, total extractant concentration of 1.5 mol/L, n(TBP)∶n(DIBK)=1∶4, extraction time of 20 min,extraction temperature of 20 ℃, V_O/V_A= 2/1, aqueous chloride ion concentration of 6 mol/L, the extraction rate of gold can reach 98.82%.The maximum gold/copper extraction separation coefficient is 1 189.05. Gradient method suggests that the composition of the extract may be [AuCl₆·3A·B]. The extraction chemical reaction formula can be rewritten as: Au³⁺+6Cl^-+3HA+B→[AuCl₆·3A·B]+3H⁺.

The influence of fluorine on the production wet-process phosphoric acid is described. The principle, application status, advantages and disadvantages of the main defluorination process of wet phosphoric acid are reviewed. The problems existing in the current defluorination process of wet phosphoric acid are pointed out, and the future development direction is put forward in order to provide reference for the comprehensive utilization of fluorine resources in wet phosphoric acid in our country.

The leaching of manganese from electrolytic manganese residue with sulfuric acid was studied, and the process parameters were optimized by single factor test and orthogonal test. The phase, structure, morphology and leaching kinetics of electrolytic manganese residue before and after leaching were analyzed. The leaching model was established and the leaching mechanism was explored. The results show that under the optimal conditions of particle size of 200 mesh, stirring speed of 400 r/min, leaching temperature of 95 ℃, sulfuric acid concentration of 1.5 mol/L, leaching time of 120 min and liquid volume/solid mass ratio of 11/1, the leaching rate of Mn can reach 97.90%. Manganese leaching kinetics accords with the liquid-solid shrinkage model and is controlled by chemical reaction. In the process of manganese leaching, the leaching rate of Mn can be improved by controlling the liquid volume/solid mass ratio and temperature.

The leaching of molybdenum from molybdenite by oxygen pressure water leaching method was studied.The effects of ore size, oxygen partial pressure, temperature and stirring speed on oxidation of molybdenite were investigated, and the distribution of oxidation products in solid and liquid phases was discussed. The results show that the oxidation of molybdenite can be promoted by the decrease of ore particle size and the increase of oxygen partial pressure, temperature and stirring speed. The oxidation products first enter the liquid phase. When the liquid phase is saturated with $\mathrm{MoO}_{4}^{2-}$, the oxidation products will enter the slag phase in the form of MoO₃. The oxidation of molybdenum during oxygen pressure water leaching of molybdenite can be described by the unreacted nuclear shrinkage model.The reaction rate is controlled by the mixed control model, and the apparent activation energy is 40.55 kJ/mol.

The comprehensive recovery of copper and arsenic from black copper slime by oxidation leaching method was studied. The effects of acidity, hydrogen peroxide content, temperature, leaching time and liquid volume to solid mass ratio on the leaching rate of copper and arsenic were investigated. The results show that under the conditions of acidity of 200 g/L, hydrogen peroxide content of 30%, temperature of 80 ℃, leaching time of 2.0 h, liquid volume to solid mass ratio of 9∶1, the leaching rates of copper and arsenic can reach 99.53% and 98.24%, respectively. After oxidative acid leaching, sodium hydrosulfide is added into the solution at the rate of n(NaHS)/n(Cu)=1.1 to precipitate copper. The mass concentration of copper in the solution after copper and arsenic separation is lower than 0.01 g/L, and the mass concentration of arsenic is greater than 40 g/L. The copper and arsenic are effectively separated. Arsenic-rich liquid can be reduced with sulfur dioxide, and arsenic trioxide and reduced liquid can be returned to the oxidation leaching process.

The change of iron grade in the leaching residue and the leaching kinetics under normal pressure and pressure were examined. The pressure leaching of nickel and cobalt from a high iron and low grade laterite nickel ore in Indonesia with sulfuric acid was studied. The results show that under the conditions of acid/ore ratio of 260 kg/t, leaching temperature of 250 ℃ (corresponding to water vapor pressure of 4.0 MPa), liquid volume/solid mass ratio of 3/1, stirring speed of 300 r/min, particle size of 100 mesh and reaction time of 1 h, the leaching rates of nickel, cobalt and iron are 98.1%, 98.3% and 4.7%, respectively, and the iron grade can reach 51.3%. The leaching processes of nickel and cobalt under pressure and atmospheric pressure are in line with the shrinkage kernel model of interfacial chemical reaction control, and the activation energies of the reactions are 116 kJ/mol and 91 kJ/mol under pressure, respectively, and the activation energies of the reaction under atmospheric pressure are 41 kJ/mol and 53 kJ/mol, respectively. The leaching is mainly goethite under normal pressure, and chromite, magnetite and other mineral phases under high pressure.

The desorption of tungsten and molybdenum from ZGA351 porous strong basic anion exchange resin using NaOH+NaNO₃ as desorption agent and H₂O₂ as strong oxidant was studied. The effects of combined adsorbent composition,concentration of NaOH, concentration of NaNO₃, concentration and dosage of H₂O₂, flow rate ratio of NaOH+NaNO₃/H₂O₂ and desorption times on the desorption of tungsten and molybdenum were investigated. The results show that under the conditions of NaOH concentration of 1.5 mol/L, NaNO₃ concentration of 5%, H₂O₂ addition of 1.2 times the theoretical doasge, H₂O₂ concentration of 30%, the NaOH+NaNO₃/H₂O₂ flow rate of 40 for 6 times, the tungsten and molybdenum desorption rate are 99.48% and 99.38%, respectively. The molybdenum penetration adsorption capacity of the resin is 107.10 g/L, which is similar to that of the new resin 107.77 g/L, indicating that the adsorption properties of ZGA351 resin remained basically unchanged after one desorption and cyclic adsorption. The mass concentration of tungsten in the primary desorption solution is 3.82 g/L, the mass concentration of molybdenum is 10.71 g/L, and the ratio of molybdenum to tungsten is 2.80. After adjusting the acid, it can be directly put into the subsequent industrial production section to make ammonium molybdate products. The results of 30 desorption and adsorption cycles show that the service life of the resin can meet the requirements of industrial production cost. The method synchronizes the process of liquid absorption oxygenation and resin regeneration.The process is simplified, the production cost is reduced, and a feasible process route is explored for the desorption process of large pore strong basic anion exchange resin.

In order to realize the high value utilization of solid and liquid wastes, extraction of vanadium from vanadium tailings with titanium dioxide waste acid was studied. The effects of leaching temperature, leaching time, solid mass to liquid volume ratio, waste acid concentration and other technological parameters on vanadium leaching rate were investigated by orthogonal experiments, and the optimal leaching conditions were determined. In order to further improve the vanadium leaching rate, the appropriate amount of leaching aid (fluorite) and oxidant (hydrogen peroxide) was discussed. The results show that the order of influence of the process parameters on vanadium leaching rate is leaching temperature> leaching time> solid mass to liquid volume ratio> titanium dioxide waste acid concentration. Under the optimum conditions of temperature of 80 ℃, leaching time of 2 h, solid mass to liquid volume ratio of 1∶5 and waste acid concentration of 5%, the vanadium leaching rate is about 53%. After adding 7% fluorite and 5% hydrogen peroxide, the vanadium leaching rate can be increased to 71.09%.

The removal of Mo(Ⅵ) from wastewater was studied by using jarosite after iron deposition in zinc leaching solution as adsorption material. The effects of the initial pH of simulated wastewater, the initial mass concentration of Mo(Ⅵ), the additon amount of jarosite and the adsorption time on the adsorption of Mo(Ⅵ) by jarosite were investigated, and the kinetic mechanism of the adsorption process was discussed. The results show that under the conditions of initial mass concentration of Mo(Ⅵ) of 40 mg/L, pH=3.0, additon amount of jarosite of 1.0 g/L and adsorption time of 240 min, the adsorption capacity of Mo(Ⅵ) can reach 13.21 mg/g. The adsorption process of Mo(Ⅵ) in wastewater by jarosite is more suitable to be described by quasi-second-order kinetic model, which belongs to chemical adsorption. The adsorption process of molybdenum by jarosite follows the Langmuir isothermal adsorption model and has higher fitting accuracy. The adsorption process belongs to monolayer molecular adsorption. This method has the advantages of simple process and low cost, and is expected to realize high-value recycling of waste residue.