The extraction of lithium and defluorination from lithium aluminum waste electrolyte by sulfuric acid leaching—lime precipitation was studied. The effects of various process parameters on the leaching rate and defluorination rate of lithium were investigated. The results show that under the optimal conditions of sulfuric acid concentration of 1.2 mol/L, leaching temperature of 95 ℃, leaching time of 20 min, liquid volume to solid mass ratio of 4/1, pH=7, precipitation time of 1 h, precipitation temperature of 95 ℃, the comprehensive leaching rate of lithium is 85%, and the defluorination rate can reach 99.91%. The addition of calcium oxide can promote the conversion of fluoride ions in the solution to calcium fluoride precipitation, and CaF₂ products with high crystallinity are obtained. The pH of the treated solution can meet the requirements for the preparation of lithium carbonate, and the separation and recovery of fluoride and lithium containing solution can be realized.

The g-C₃N₄/Bi₂WO₆ composite photocatalyst was prepared by hydrothermal method, and its properties and mechanism for Cr(Ⅵ) reduction were studied under visible light. The results show that g-C₃N₄/Bi₂WO₆ is a petaloid structure composed of nanosheets with larger surface area than the single g-C₃N₄ and Bi₂WO₆, which can provide more active sites for photocatalytic reaction. The adsorption efficiency of g-C₃N₄/Bi₂WO₆ for Cr(Ⅵ) is 43.2% under 40 min dark condition, and the adsorption behavior is consistent with the quasi-second-order kinetic model. After 100 min of visible light irradiation, the photocatalytic reduction rate of g-C₃N₄/Bi₂WO₆ for Cr(Ⅵ) is 81.3%, and the photocatalytic reduction process is consistent with the quasi-first-order kinetic model. After the combination of g-C₃N₄ and Bi₂WO₆, forms a Z-scheme heterojunction, which broadens the light absorption range and promots the separation of photogenerated electron-hole, thus showing excellent visible light catalytic activity.

The recovery of calcium and silicon from steel slag by acid-base synergistic separation method was studied. The effects of leaching agent concentration, temperature, leaching time and liquid volume to solid mass ratio on the leaching rate of calcium and silicon in steel slag were investigated. The leaching rules under different leaching conditions were investigated. The results show that the leaching concentration, time, and liquid volume to solid mass ratio significantly influence the leaching rate of calcium and silicon, and temperature has no significant effect. Under the conditions of hydrochloric acid concentration of 0.75 mol/L, leaching time of 30 min, temperature of 25 ℃, and liquid volume to solid mass ratio of 10∶1, the calcium leaching rate is 53.74%, and the resulting calcium oxalate purity can reach 98.94%. Under the conditions of sodium hydroxide concentration of 1.5 mol/L, temperature of 95 ℃, leaching time of 120 min, and liquid volume to solid mass ratio of 20∶1, the silicon dioxide leaching rate is 43.22%, and the purity of the resulting silicon dioxide can reach 80.85%. The method can effectively separate and recover calcium and silicon from steel slag, and realize resource utilization.

The separation and recovery of magnesium and nickel from serpentine and the co-mineralization of CO₂ using green vitriol roasting—water leaching—ammonium carbonate precipitation process was investigated. The effects of roasting temperature, green vitriol to serpentine mass ratio and roasting time on the leaching rates of magnesium and nickel during sulfuric acid roasting were examined. Additionally, the influence of ammonium carbonate concentration and mineralization temperature on the precipitation rates of magnesium and nickel in the leachate during the mineralization reaction(resulting in the formation of magnesium carbonate precipitate) was evaluated. The effects of ammonium carbonate concentration and mineralization temperature on the precipitation behavior of metal ions and the microstructure of the precipitates were also explored. The results demonstrate that under the optimal roasting conditions of temperature of 670 ℃, green vitriol to serpentine mass ratio of 5/1, and roasting time of 90 min, the leaching rates of magnesium and nickel can reach 89% and 85%, respectively. During the mineralization reaction, with an ammonium carbonate mass concentration of 150 g/L and mineralization temperature of 80 ℃, the precipitation rate of Ni²⁺ is less than 5%, while Mg²⁺ and Fe³⁺ precipitate at rates of 97% and 100%, respectively. With the increase of the mass concentration of ammonium carbonate, the precipitate transforme from Ni₆Fe₂(CO₃)(OH)₁₆•4H₂O to MgCO₃•3H₂O, and then to (NH₄)₂Mg(CO₃)₂•3H₂O. With the increase of mineralization temperature, the precipitate gradually shifted from (NH₄)₂Mg(CO₃)₂•3H₂O to Mg₅(CO₃)₄(OH)₂•4H₂O. The morphology of the magnesium carbonate precipitate is rod-like at low temperatures and plate-like at high temperatures. The experimental results can provide a novel approach for nickel extraction from serpentine, CO₂ reduction processes, and the efficient utilization of waste ferrous sulfate.

Aiming at the problems of high content of fluorine and chloride in zinc oxide smelting dust, corrosion of anode and cathode plate and poor quality of zinc electrodeposition caused by direct return system, a step process of "acid leaching—defluorination of leaching liquid lime—neutralization of zinc after defluorination" was studied to recover zinc from zinc oxide smelting dust. The results show that under the optimal conditions of leaching temperature of 80 ℃, leaching termination point pH=1.0, liquid volume to solid mass ratio of 4/1 and leaching time of 1 h, the leaching rates of Zn, F and Cl are 96.15%, 95.26% and 97.44%, respectively. When the pH of lime and fluorine-containing soot acid leaching solution is 5.2, the fluorine precipitation rate is 95.17%, while the zinc loss rate is only 1.73%. After fluoride removal, when the liquid is further neutralized with lime to pH=8.0, all zinc can be precipitated, and the mass fraction of zinc in the obtained neutralized slag is about 24%. The removal rates of fluorine and chlorine in the whole process are about 92% and 97%, respectively, the recovery rate of zinc is more than 90%, and the lead and silver can be recovered into the acid leaching residue. The process can realize the efficient removal of harmful elements fluorine and chlorine and the effective recovery of zinc, lead and silver, and has a certain popularization value.

The synergistic decomposition of scheelite by HCl-H₂SO₄ was studied. The influence of various process parameters on the decomposition effect was investigated, and the kinetics of decomposition process was discussed. The results show that under the optimal conditions of hydrochloric acid concentration of 22%, liquid volume of concentrated sulfuric acid of 0.5%, liquid volume to solid mass ratio of 2.5∶1, decomposition temperature of 85 ℃, decomposition time of 2 h, stirring rate of 360 r/min, the decomposition rate of calcium tungstate in scheelite is 99.6%, and the decomposition rate is high. The process of HCl-H₂SO₄ synergistic decomposition of scheelite is controlled by chemical reaction and solid film mixing, and the apparent activation energy is 45.52 kJ/mol. The CaWO₄ in scheelite can be efficiently converted into tungstic acid, thus effectively improving the tungsten extraction rate in scheelite.

The efficient utilization of secondary silver-containing resources is of great significance for making up for the shortage of silver ore resources in China, alleviating the contradiction between supply and demand, and ensuring the supply security of national strategic metals. For waste Ag-Cu filler metal, the leaching and separation of copper and silver in the HCl-H₂O₂ system was studied, and high-purity AgCl was prepared. The influence of various factors on the separation effect of copper and silver was investigated. The results show that under the optimized conditions of H₂O₂ excessive coefficient of 1.3, HCl excessive coefficient of 1.4, liquid-solid mass ratio of 11∶1, leaching temperature of 40 ℃, and leaching time of 3 h, the leaching rate of copper can reach 99.38%. The leaching residue is irregular spherical particle with particle size of 2~4 μm. The main phase is AgCl, and the main components are Ag and Cl, with total mass fraction of more than 99%.

The direct leaching of Cu and Co from a copper-cobalt mixed ore in Congo(Kinshasa) by oxidation leaching method and reduction leaching method was studied. The leaching behavior of Cu and Co by the two methods was compared and analyzed. The effect of flotation—oxidation leaching on the leaching rate of copper and cobalt was further discussed. The results show that the leaching rates of Cu for oxidation and reduction leaching are about 80% and 70% respectively, and the leaching rates of Co are about 63% and 57% respectively, and the oxidation leaching effect is better than that of reduction leaching. By using the combined flotation—oxidation leaching process, the Cu and Co grades in closed-circuit flotation concentrates are 16.84% and 6.53%, respectively and the combined recovery rates of Cu and Co are 91.71% and 61.87%, respectively. The recovery effect of Cu and Co can be significantly improved.

Aiming at the dilute acid of copper smelting as raw material, the selective precipitation of copper and removal of arsenic with oxalic acid was studied. The effects of oxalic acid addition dosage, solution pH, stirring speed, reaction temperature and time on the separation of copper and arsenic were investigated. The results show that under the conditions of H₂C₂O₄/Cu molar ratio of 1.1, solution pH=0.4, stirring speed of 500 r/min, reaction temperature of 25 ℃ and reaction time of 30 min, the precipitation rate of copper is more than 98.5%, and the precipitation rate of arsenic is less than 0.20%, copper and arsenic can be separated efficiently.

Chlorine is a harmful element in the process of zinc hydrometallurgy. During zinc electrodeposition, it will corrode plates and equipment and release toxic gases such as chlorine, resulting in increased production cost and environmental pollution. The sources of chlorine elements in the process of zinc hydrometallurgy and its hazard to the process of zinc hydrometallurgy are briefly introduced. The research and development status, advantages and disadvantages of the main chlorine removal processes such as chemical precipitation, extraction and ion exchange are summarized, and the main research and development directions in the future are pointed out.