The leaching of magnesium from lightly fired powders obtained from low temperature calcination of magnesite by hydrochloric acid acidification was studied. The effects of addition method of hydrochloric acid, concentration of hydrochloric acid, liquid volume/solid mass ratio, external force way and reaction time on magnesium leaching rate were investigated. The leaching kinetics of lightly fired powders under ultrasonic condition was analyzed by using shrinkage core model. And the particle size distribution, mineral composition and mineral distribution characteristics of leaching residue were further analyzed. The results show that under the conditions of slowly adding hydrochloric acid, hydrochloric acid concentration of 14.8%, liquid volume/solid mass ratio of 9.8/1, stirring speed of 400 r/min and leaching time of 12.5 min, the leaching rate of magnesium can reach 70.48%. Using ultrasound can cause hard agglomeration to produce cracks or disintegration, which is beneficial to solid-liquid mass transfer, and the magnesium leaching rate is increased to 90.56%. The leaching process of lightly fired powders conforms to the mixed control model of surface chemical reaction and product layer diffusion. The ratio of magnesium to oxygen in the hard agglomeration is 1.33, the edge is 1.65, and there is a large amount of unreacted magnesium in the hard agglomeration. The particle size of the leaching slag is reduced from 42.61 μm to 24.39 μm by ultrasonic action. The ultrasonic bubble vibration force and the shock wave formed by the cavitation collapse can effectively prevent the formation of hard agglomeration or make the hard agglomeration re-disaggregate.

Recovering and reusing ruthenium from platinum group metal waste has great influence on sustainable development, resource conservation and environmental protection. Ruthenium waste generally includes ruthenium-containing waste catalysts, ruthenium-containing alloy materials, ruthenium-containing nuclear waste and other ruthenium-containing waste. The comprehensive recovery and treatment methods of different ruthenium wastes are summarized. The processes of transforming soluble salts by oxidative distillation, melting-reduction-oxidation, melting-oxidation-distillation, microwave leaching-cloud point extraction, distillation-melting-reduction, melting-reduction-oxidation, ion exchange, oxidation volatilization, electrolysis, biosorption and physical adsorption are introduced respectively. The advantages and disadvantages of existing ruthenium waste recovery technologies are summarized. And the development direction of ruthenium waste recovery technology in the future is prospected.

After contaminated acid in zinc hydrometallurgy was vulcanized, neutralized by zinc calcine and calcium carbonate, arsenic and fluorine can be removed from the solution, and the amount of gypsum residue decreased. But the solution can not return to zinc hydrometallurgy system because it has a high chlorine content. The chlorine in the solution was removed with copper slag. The results show that under the conditions of hydrogen peroxide addition of 2.3 mL/L, copper slag coefficient of 6.8, initial acid mass concentration of 15 g/L, temperature of 35 ℃ and reaction time of 2 h, the chlorine removal rate is 72.58% and the chlorine mass concentration decreases from 0.62 g/L to 0.17 g/L. The copper mass concentration in solution is 2.69 g/L copper, and it can be returned to the zinc hydrometallurgy after subsequen copper removal by zinc powder or iron powder.

A set of aluminum dross hydrolysis hydrogen production equipment was experimentally designed and installed. The effects of reaction temperature, liquid volume to solid mass ratio, stirring speed, and aluminum dross particle size on the aluminum dross hydrolysis hydrogen production process were studied, and the kinetics of the aluminum dross hydrolysis hydrogen production process were explored. The results show that the optimal process conditions for aluminum dross hydrolysis hydrogen production are reaction temperature of 85 ℃, liquid volume to solid mass ratio of 10 mL/1 g, and stirring speed of 130 r/min, aluminum dross particle size of >80 mesh. The main phases of the hydrolysis residue obtained under optimal conditions are MgAl₂O₄, Al(OH)₃ and Al₂O₃. The hydrogen production process from aluminum dross hydrolysis is controlled by chemical reactions, with a chemical apparent activation energy of 67.01 kJ/mol. The leaching process follows the Avrami-Erofeyev model. The experimental results can provide certain technical references for the design of aluminum dross hydrolysis hydrogen production process.

In the process of producing ammonium molybdate by roasting and ammonia leaching, the leaching rule of potassium in ammonia leaching process of molybdenum calcine was studied by ICP, XPS, SEM, etc., and the change of occurrence form of potassium in ammonia leaching process was investigated by MLA and Factsage thermodynamic software. The kinetic law of potassium release was investigated by using the unreacted kernel model, parabolic diffusion equation, double constant equation, Elovich equation and first-order kinetic equation. The results show that the leaching of potassium in the ammonia leaching process of washed molybdenum calcination can be divided into two stages:Preleaching stage is the leaching process of ionic potassium, which mainly occurred the ion exchange reaction of KCl and K₂SiF₆, and the leaching activation energy is 4.79 kJ/mol. The Elovich model is the best fit for this stage. Late leaching stage is mainly the leaching process of potassium from mineral potassium, the leaching activation energy is 34.55 kJ/mol, and the optimal kinetic model is a double constant model. At the late stage of leaching, the potassium release rate is slow, and the potassium release ability of the four potassium-containing minerals is from weak to strong in the order of illite<barium ferromica<mica<orthoclase.

Polyaniline (PANI) coated iron tetroxide was used to prepare of Fe₃O₄@PANI magnetic adsorption material for the adsorption of iodine in water. The morphology and properties of the composite before and after adsorption were characterized by SEM, TEM, FT-IR, XRD, VSM, XPS and Raman, and the adsorption mechanism of the composite was explored. The effects of adsorption time, iodine solution concentration, adsorption temperature and regeneration on the adsorption properties were also investigated. The results show that iodine is bound to benzene ring, quinone ring and nitrogen atom of quinone ring structure unit of polyaniline when coated with ferric tetroxide adsorbed I₂. The adsorption process of I₂ by Fe₃O₄@PANI is endothermic and spontaneous, which accords with the quasi-second-order kinetic model and the Langmuir isothermal adsorption model. At 303.15 K, the theoretical maximum adsorption capacity is 1 777.13 mg/g. After the adsorbent is desorbed with ethanol and recycled for 3 times, the adsorption rate reached 44.22% of the first adsorption rate.

Leaching of vanadium from stone coal vanadium ore by blank roasting—acid leaching process was studied. The effects of grinding fineness, roasting time, roasting temperature, sulfuric acid addition dosage, leaching aid CaF₂ addition dosage, leaching temperature, leaching time and liquid volume to solid mass ratio on vanadium leaching rate were investigated. The results show that the optimum blank roasting conditions are grinding fineness 70% of -74 μm, roasting temperature of 775 ℃ and roasting time of 5 h. The optimum acid leaching conditions are sulfuric acid addition dosage of 10%, leaching aid CaF₂ addition dosage of 3%, liquid volume to solid mass ratio of 1.5∶1, leaching temperature of 25 ℃ and leaching time of 1.5 h. Under suitable conditions vanadium leaching rate can reach 81.5%.

Aiming at the comprehensive utilization of niobium minerals in Bayan Obo tailings, the optimum parameters of niobium leaching process were determined by using HF as leaching agent, and the leaching kinetics was studied. The results show that the leaching rate of niobium-containing minerals can reach 90.91% under the conditions of reaction temperature of 90 ℃, liquid volume to solid mass ratio of 7/1, HF concentration of 20 mol/L, reaction time of 2 h and stirring rate of 300 r/min. The leaching process of niobium is consistent with the nuclear shrinkage model, and the leaching process is controlled by chemical reaction and diffusion mixture, and the apparent activation energy of the leaching reaction is 35.459 kJ/mol. This method can effectively extract niobium from tailings and is beneficial to subsequent separation and purification.

A new process for preparing high-purity copper by electrorefining in a nitric acid system was developed using a 4N copper cathode as raw material. The effects of electrolyte composition on the content of S, Cl, typical positively charged and negatively charged impurities in copper cathodes were investigated. Moreover, the influence of Cu(NO₃)₂·3H₂O raw material purity, electrolyte components(Cu²⁺, HNO₃ and Cl^- mass concentrations), and electrolysis period on copper cathode purity, impurity content, current efficiency, and electrical energy consumption was analyzed. The results show that the optimum process conditions are to prepare the electrolyte using 99% purity of Cu(NO₃)₂·3H₂O, with Cu²⁺ mass concentration of 50 g/L, HNO₃ mass concentration of 31.5 g/L, Cl^- mass concentration of 0.1 g/L, and electrolysis period of 72~120 h. Under the optimal conditions, high-purity copper cathode with purity up to 6N5 and main control impurity elements can meet the high purity cathode copper required by HPCu-6N in the national standard (GB/T 26017—2020).

The removal of aluminum nitride is one of the keys processes of the harmless treatment of secondary aluminum ash. The removal of aluminum nitride from secondary aluminum ash in calcium oxide-glucose hydrolysis system was studied. The effects of calcium oxide addition amount, glucose addition amount, hydrolysis temperature, hydrolysis time and liquid volume to solid mass ratio on the removal rate of aluminum nitride were examined. The mechanism of aluminum nitride hydrolysis removal promoted by additives was investigated. The results show that under the optimum hydrolysis conditions of hydrolysis temperature of 90 ℃, hydrolysis time of 2 h, liquid volume to solid mass ratio of 10 mL/1 g, calcium oxide addition amount of 10%, glucose addition amount of 0.75%, the removal rate of aluminum nitride reaches 95.42%. Through analysis and characterization, it is determined that in the calcium oxide-glucose hydrolysis system, glucose acts as a retarder, which can inhibit the formation of calcium aluminate coating layer on the surface of aluminum nitride particles, make the hydrolysis of aluminum nitride thoroughly, and improve its removal rate.