For zinc acid leaching residue, the reduction leaching of zinc, iron and copper and the enrichment of lead and silver were studied by using zinc concentrate as the reducing agent and zinc smelting waste electrolyte as the leaching agent under micro-pressure conditions. The micro-pressure reduction leaching residue was then leached with hot acid. The effects of various factors on the two-stage leaching were investigated. The results show that the optimal conditions for micro-pressure reduction leaching are pressure of 0.2 MPa, final acid concentration of 35 g/L, reaction temperature of 100 ℃, liquid volume to solid mass ratio of 8/1, zinc concentrate addition of 1.1 times the theoretical amount, stirring speed of 400 r/min, and reaction time of 2 h. The micro-pressure reduction leaching residue obtained under these conditions was leached with hot acid at 90 ℃ and a final acid concentration of 110 g/L for 2 h. The total leaching rates of zinc, iron and copper after the two-stage leaching can reach 95.81%, 92.67% and 92.63%, respectively, and the residue rate is 24.61% of the zinc acid leaching residue, which can achieve the reduction of the residue. Lead and silver are enriched in the hot acid leaching residue, with a sulfur mass fraction of 23.42%. The leaching residue has a high content of valuable metals and strong potential for self-heating. The method utilizes the reducing property of zinc concentrate and the oxidizing property of Fe³⁺ in the zinc acid leaching residue leaching solution to achieve efficient and comprehensive recovery of zinc acid leaching residue, and has certain application and promotion value.

High efficient and selective recovery of lithium from spent LiFePO₄ batteries by hydrogen peroxide oxidative leaching method was studied. The effects of hydrogen peroxide concentration, solid mass to liquid volume ratio, leaching temperature and time on the leaching behavior of lithium were examined. Kinetic analysis, XRD and SEM were used to investigate the leaching mechanism. The results show that under the optimal conditions of hydrogen peroxide concentration of 20%, solid mass to liquid volume ratio of 3 mg/1 mL, leaching temperature of 40 ℃ and leaching time of 40 min, the leaching rate of lithium can reach 99%, while the leaching rate of iron is only 0.53%, and the leaching process is mainly controlled by internal diffusion. The method can provide technical support for the efficient selection and recovery of lithium in spent LiFePO₄ batteries.

Iron and fluorine co-doped manganese ion sieve HMFeO-F was prepared by high temperature solid phase method. The effects of calcination temperature, calcination time and doping ratio on the structure of the ion sieve were investigated by XRD, SEM, FT-IR and BET, and its adsorption performance was studied. The results show that the adsorption capacity for lithium of HMFeO-F is 31.57 mg/g under the conditions of calcination temperature of 800 ℃, calcination time of 10 h, molar ratio of lithium to manganese of 1/2, and doping mole fractions of Fe and F of 5% and 10%, respectively. The Mn dissolution loss rate during the adsorption process is only 4.52%, which is 61% lower than that of the undoped. At the same time, the introduction of magnetic Fe₃O₄ makes the magnetic recovery rate of the material can reach 90.2%. After five cycles of tests, the equilibrium adsorption capacity for lithium of HMFeO-F can still reach 25.87 mg/g, and the manganese dissolution loss rate remains within 5%. For a coal gangue acid leaching solution with a lithium mass concentration of 12.98 mg/L, lithium is adsorbed by HMFeO-F, and the lithium adsorption rate can reach 99%. The use of ion sieve to extract lithium from coal gangue can not only develop a new way to extract lithium, but also is expected to achieve the high-value utilization of coal gangue.

The original biochar (BC) was prepared by pyrolysis using chestnut shells as raw materials, and then the phosphate-modified nitrogen-doped biochar (NP-BC) was prepared by nitrogen doping and phosphate modification methods. The adsorption performance of BC and NP-BC for uranium in wastewater was investigated. The surface morphology and elements of BC and NP-BC were analyzed by means of SEM, EDS, BET, FT-IR and XPS, and the adsorption mechanism of BC and NP-BC for uranium was discussed. The results show that for the wastewater with an initial mass concentration of uranium of 10 mg/L and initial pH of 4, under the conditions of NP-BC dosage of 0.2 g/L and reaction temperature of 30 ℃, the adsorption removal rate of U(Ⅵ) can reach 99.32%. The influence of coexisting anions and cations on the division of U(Ⅵ) in NP-BC increases with the increase of ion concentration. The adsorption process is more suitable to be described by the pseudo-second-order kinetic model and the Langmuir isothermal adsorption model. It is concluded through FT-IR and XPS analysis that the adsorption mechanism of uranium by NP-BC is the coordination effect and reduction effect of oxygen-containing functional groups, nitrogen-containing functional groups and phosphate groups.

In the process of extracting antimony by alkaline electro-deposition, problems such as cracking of the cathode deposited antimony, causing warping and shedding, occur due to insufficient interfacial bonding force between Sb and Fe, surface sandblasting pretreatment technology was proposed. The effects of the sandblasting pretreatment on the cathodic reduction and nucleation mechanism of antimony ions were systematically analyzed by combining cyclic voltammetry, steady-state polarization curves, and time-current transient curves. The results show that for the alkaline electro-deposition system of antimony, after sandblasting pretreatment of the iron-based cathode surface, the initial potential of the cathodic reduction reaction shifts to the left and positive. At the same overpotential, the current density is significantly increased compared to that without sandblasting treatment, and transfer coefficient and exchange current density are also enhanced. This indicates that the sandblasting treatment of the cathode surface can reduce the energy barrier of the cathodic reduction reaction of antimony ions, significantly increase the reduction reaction rate, and promote the three-dimensional nucleation process of antimony on the cathode surface, thereby obtaining a dense. well-shaped cathode antimony deposit layer. Further in-depth research was conducted on the influence of sandblasting particle size on the antimony electro-deposition process, and the optimal sandblasting particle size was determined to be 80 mesh. Under the condition, the cathode current efficiency is increased to 80.1%, the cell voltage is 2.031 V, the direct current power consumption per ton of antimony is reduced to 1674.76 kW·h/t, and the purity of cathode antimony reaches 98.81%.

To address the issue of low extraction efficiency of aluminum from fly ash in pulverized coal furnaces due to its mullite structure, a process combining high—temperature enhanced decalcification with medium recycling to remove calcium and extract alumina from activated fly ash clinker was proposed. The optimal process parameters for decalcification and alumina extraction were determined through single-factor experiments. The results indicate that the optimal decalcification conditions are reaction temperature of 160 ℃, hydrochloric acid concentration of 12%, and reaction time of 4 h. Under the conditions, the CaO removal rate exceeds 98%, while the Al₂O₃ dissolution rate is effectively suppressed below 5%. The optimal alumina extraction conditions are temperature of 60 ℃, hydrochloric acid concentration of 26%, and reaction time of 0.5 hours, achieving an Al₂O₃ dissolution rate of over 93% and controlling the Ca²⁺ concentration in the leachate below 0.39 g/L. Innovatively, a closed-loop calcium chloride recycling system is constructed, enabling recovery rate of over 90% for the activator in filtrate reuse tests. The process can achieve efficient calcium-aluminum separation and recycling of additives, offering a low-carbon solution for alumina extraction from fly ash.

The phosphorus content in sludge incineration ash is relatively high. Extracting and recovering it can effectively alleviate the contradiction between the supply and demand of phosphorus. Leaching of phosphorus from sludge incineration ash with nitric acid was studied. The effects of nitric acid concentration, leaching temperature, leaching time and the ratio of liquid volume to solid mass on the leaching rate of phosphorus were investigated. The restrictive links of phosphorus leaching were determined and the kinetic equation was established. The results show that under the conditions of nitric acid concentration of 1 mol/L, temperature of 298 K, leaching time of 60 min, stirring speed of 500 r/min, and liquid volume to solid mass ratio of 40 mL/1 g, the leaching rate of phosphorus reaches 91.36%. Internal diffusion control is the limiting link in the phosphorus leaching process, and the activation energy of the leaching reaction is 2.11 kJ/mol. The method has a relatively high phosphorus extraction efficiency. It is an efficient and feasible technology for recovering phosphorus from sludge incineration ash and can provide an important technical approach to alleviate the contradiction between supply and demand of phosphorus resources.

The composition and physicochemical properties of waste lead paste are complex, and its recycling and utilization are difficult. Among them, the reduction of high-valent PbO₂ is a key problem in the recycling and treatment process of waste lead paste. The reduction sulfation reaction of converting PbO₂ in waste lead paste to PbSO₄ using the glucose-sulfuric acid system at room temperature was studied. The effects of sulfuric acid concentration, glucose excess coefficient, liquid volume to solid mass ratio and reaction time on the reduction sulfation process of waste lead paste were investigated. The results show that under the optimal reaction conditions of waste lead paste of 5 g, sulfuric acid concentration of 85%, excess coefficient of glucose of 3, liquid volume to solid mass ratio of 3/1, and reaction time of 3 h, the reduction rate of PbO₂ is 91.88% and the mass fraction of PbSO₄ is 92.52%. The system can reduce and sulfate various complex components in waste lead paste to obtain PbSO₄ products of higher purity. It is green and efficient, and convenient for the subsequent recycling and treatment of waste lead paste.

Using high-tellurium copper crude selenium as raw material, the process of preparing 4N selenium powder by high temperature oxidatio—water dissolution-low temperature reduction—temperature rise transformation was studied. The effects of key process parameters in each process on the purification of crude selenium to prepare 4N selenium powder were investigated, and the reduction reaction of selenium and the competitive behavior of selenium and tellurium were analyzed based on thermodynamic theory. The results show that under the optimal process conditions of oxidation temperature of 550 ℃, oxidation time of 3.0 h, oxygen flow rate of 2.5 m³/h, reduction temperature of 25 ℃, hydrazine hydrate concentration of 15%, stirring speed of 300 r/min, reaction endpoint solution pH=6.5, and transformation temperature of 55 ℃, the selenium product can meet the standard of 4N selenium powder. The method can provide a new technical route for the production of high-purity 4N grade selenium.

Removal of impurities (arsenic, antimony, bismuth) from copper electrolytic waste liquid by hydrogen sulfide was studied. The effects of the excess coefficient of hydrogen sulfide, reaction temperature, reaction time and stirring speed on the impurity removal of copper electrolytic waste liquid were investigated. The experimental conditions were further optimized by the response surface methodology, and the experiments were verified under these conditions. The results show that the optimal test conditions are as follows：the excess coefficient of hydrogen sulfide is 1.9, the reaction temperature is 68.1 ℃, the reaction time is 2.1 h, and the stirring speed is 524 r/min. Under the optimal conditions, the average impurity removal rate of copper electrolysis waste liquid can reach 94.2%. The optimized process parameters determined by the method are reliable and applicable, and will not affect the subsequent recovery of nickel elements.