To address the problems such as insufficient real-time data analysis, lack of intelligent decision-making support, and limited dynamic optimization capabilities of in-situ leaching dynamic mining system in practical application, the advantages of integrating Digital Twin technology with the in-situ leaching process for uranium extraction were analyzed. By incorporating advanced technologies such as muon imaging, fiber-optic water level monitoring, advanced sensing technologies, artificial intelligence, and big data analysis, an efficient digital-intelligent uranium mining platform was constructed. The platform includes geological structure model, groundwater seepage model, reactive transport model, and intelligent agent model. The core algorithms cover deep learning, data assimilation, multi-objective optimization, and uncertainty analysis. The establishment of the dynamic in-situ leaching mining system can improve the efficiency of uranium resource development and can provide valuable reference for the extraction of other mineral resources, thus having certain theoretical significance and practical value.

In order to solve the issue of low automation in the hydrometallurgical process in in-situ leaching of uranium, especially in the preparation of yellowcake from precipitation slurry using plate and frame filter presses, a horizontal belt vacuum filter was selected for the first time in an in-situ leaching of uranium mine to conduct a filtration dewatering experiment on sodium diuranate precipitation slurry. The influence of different operating conditions on moisture content of yellow cake products was investigated. The results show that under the conditions of a filtration vacuum of -0.06~-0.08 MPa, a filter cloth precision of 38 μm, a slurry residence time of 640 s, a feed flow rate of 353~589 kg/h, and a backwash and sealing circulation water flowrate of 10 m³/h, the entire process from cloth feeding to filtration, dewatering, cake discharge, and cloth backwashing can be automatically cycled and run continuously and stably. The product moisture content meets the first-grade standard, satisfying the dewatering capacity requirements for a mine producing about 500 t of metallic uranium per year. The technical feasibility of using belt vacuum filter to dehydrate the precipitated slurry has been preliminatively confirmed. The technical feasibility of using belt vacuum filter for dewatering the precipitation slurry has been preliminatively confirmed. The filtration dewatering process can reduce the labor intensity of operators, reduce the radiation hazard, and improve the automation of the entire hydrometallurgy process, it has value for promotion and application.

Extracting vanadium from stone coal vanadium ore with sericite as the main material and vanadium in the form of isomorphic vanadium by barium salt roasting—acid leaching was studied. The effects of grinding fineness, dosage of barite, roasting temperature, roasting time, liquid volume to solid mass ratio, sulfuric acid dosage, leaching temperature and leaching time on vanadium leaching rate were investigated. The results show that the vanadium leaching rate can reach 89% under the conditions of grinding fineness of -200 meshes of 70%, barite content of 5%, roasting temperature of 850 ℃, roasting time of 16 h, liquid volume to solid mass ratio of 2.0, sulfuric acid dosage of 8%, leaching temperature of 25 ℃ and leaching time of 2 h. The leaching effect is good.

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.

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%.

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.

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 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.

Lead leaching residue of acid mud contains a large number of complex and scattered precious metals, which has high recovery value. The leaching of selenium, bismuth, copper and mercury from lead leaching residue using a mixture of hydrochloric acid and sulfuric acid as the leaching agent and sodium chlorate as the oxidizing agent was studied. The effects of temperature, dosage of sodium chlorate, leaching time, liquid volume to solid mass ratio and hydrochloric acid concentration on the leaching rates of selenium, bismuth, copper and mercury were investigated. The results show that under the conditions of temperature of 40 ℃, dosage of sodium chlorate as 40% of raw material mass, leaching time of 1 h, liquid volume to solid mass ratio of 4 mL/1 g, hydrochloric acid concentration of 1 mol/L and sulfuric acid concentration of 1.5 mol/L, the leaching rates of selenium, bismuth, copper and mercury are 98.78%, 98.13%, 99.38% and 99.42%, respectively. The method can achieve high efficiency leaching of selenium, bismuth, copper and mercury, and enrich the tin, lead and silver in the residue.

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 challenge of achieving profound separation between rare earth elements(REEs) and non-rare earth impurities during the purification process via extractive resin adsorption has consistently posed a formidable obstacle. The static method was employed to analyze the adsorption behavior of La³⁺ and Zn²⁺ ions onto CL-P507 resin. The influences of adsorption duration, solution pH, and temperature on the adsorption efficiency of La³⁺ and Zn²⁺ by the resin were investigated. The adsorption kinetics and thermodynamics of La³⁺ and Zn²⁺ were analyzed. Additionally, the separation and migration mechanisms of La³⁺ and Zn²⁺ during the adsorption process was discussed. The results indicate that H⁺ participates in the reaction involving the phosphoric acid functional group of CL-P507 resin. At a feed liquid pH of 4.0, the adsorption equilibrium values for La³⁺ and Zn²⁺ are achieved at 0.098 9 mmol/g and 0.153 4 mmol/g, respectively within 8 min and 15 min. The adsorption of La³⁺ and Zn²⁺ onto CL-P507 resin is an endothermic process adhering to the Langmuir isothermal model and quasi-second-order kinetics, with chemical reactions governing the process. The findings offer theoretical insights into the efficient separation of rare earths from non-rare earth impurities.

A hydrangea-like γ-AlOOH adsorbent with a layered structure was prepared by hydrothermal method using sodium aluminate solution obtained from acid leaching of red mud as raw material and used for the adsorption of Congo red. The effects of initial pH and initial mass concentration of Congo red on the adsorption were investigated, and the adsorption mechanism of γ-AlOOH on Congo red was discussed. The results show that the saturated adsorption amount of Congo red can reach 1 748.15 mg/g under suitable conditions. The adsorption process is more appropriately described by the quasi-second-order kinetic model and the Langmuir isothermal adsorption model, and the adsorption behaviours are mainly dominated by monolayer adsorption. The adsorption process is spontaneous, heat-absorbing and chaotic with an increasing degree of confusion. The process can provide an effective solution to solve the problem of Congo red pollution and red mud accumulation in the water body, which is in line with the green development concept of treating waste with waste and turning waste into treasure, and has certain value of popularisation and application.

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 extraction of zinc by P204(Di-(2-ethylhexyl) phosphate) from sulfuric acid leaching purification liquid of flotation lead tailings was studied. The extraction and stripping conditions were optimized. ZnSO₄•7H₂O was prepared by evaporation and crystallization of zinc-rich liquid. The results show that under the conditions of organic phase composition of 50%P204+50%sulfonated kerosene, saponification degree of P204 of 65%, extract pH of 3.5, extraction phase ratio V_A/V_O=1/1, extraction time of 10 min and two stages of extraction, the extraction rate of zinc can reach 99.11%. Under the conditions of H₂SO₄ concentration of 7%, stripping phase ratio V_O/V_A=2/1, stripping time of 10 min and three stages of stripping, the stripping rate of zinc can reach 99.32%. The prepared ZnSO₄•7H₂O Can meet the requirements of HG/T 2326—2015, and its microstructure is uniformly distributed in granular and layered structures.

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.

In order to further improve the recovery rate of hydrometallurgical resources and solve the problem that the intelligent and automatic control degree of resource recovery process control is not high, a hydrometallurgical process control method is proposed, which uses Transformer model to predict metal leaching rate and then uses Distributional Q-function to improve DQN model to maximize gold leaching rate. The results show that the system control method can effectively improve the prediction accuracy of metal leaching rate in hydrometallurgy process. Improving the DQN model based on Distributional Q-function can effectively reduce the iterative calculation time of the model with maximum resource recovery rate. The method can effectively improve the recovery rate of hydrometallurgical resources in a certain plant.