Vanadium-containing steel slag is an industrial waste rich in high-value elements and has significant secondary utilization value. Therefore, how to recover vanadium resources from steel slag at low cost and high efficiency is one of the key challenges that China's vanadium industry needs to overcome. The current research status of wet vanadium extraction technology from vanadium-containing steel slag is reviewed. The progress of calcination-leaching and direct leaching processes is introduced. Other new vanadium extraction processes and enhanced auxiliary methods for vanadium extraction are analyzed. Finally, the advantages and disadvantages of various processes are compared, and the future development trends are prospected.

The treatment and resource utilization of titanium-bearing blast furnace slag are the focus and difficult point of research on environmental protection for the iron and steel industry. The research progress on extraction of titanium from titanium-bearing blast furnace slag is reviewed, and the principle, process characteristics, treatment effects, existing problems, industrial prospects and environmental impact of sulfuric acid leaching, hydrochloric acid leaching, organic acid leaching, ammonium sulfate melting, ammonium chloride roasting, alkali melting methods for treatment of titanium-bearing blast furnace slag are analyzed. The technical requirements and development trends of extraction of titanium from titanium-bearing blast furnace slag by hydrometallurgy are prospected.

Using lactic acid solution as the leaching agent, selective leaching of zinc from metallurgical dust collected from a steel plant in Shanxi was investigated. The effects of leaching time, leaching temperature, liquid volume to solid mass ratio and lactic acid concentration on the leaching effect were investigated, and the leaching mechanism was explored. The results show that under the optimal leaching conditions of lactic acid concentration of 3%, leaching temperature of 20 ℃, leaching time of 5 min, and liquid volume to solid mass ratio of 10 mL/1 g, the leaching rates of zinc and iron are 91.56% and 4.69%, respectively, achieving selective leaching of zinc. In the leaching solution, zinc primarily exist in the forms of Zn［CH₃CH(OH)COO］⁺ and Zn²⁺, while in the leach residue, zinc is mainly present as zinc ferrite. The weak acidic environment of the leaching system can effectively suppress the dissolution of iron.

Recovery of active materials from the cathode materials of waste NCM ternary lithium batteries using NaOH-directed dissolution method of oxide films was investigated. The effects of solid mass to liquid volume ratio, temperature, and NaOH concentration on the recovery rate of active materials were examined. The results indicate that under the conditions of solid mass to liquid volume ratio, of 4 g/10 mL, temperature of 65 ℃, and NaOH concentration of 3 mol/L, the complete separation of the active materials from the aluminum foil can be achieved in just 1 minute. The shape of the recovered aluminum foil is intact, and the recovery rate of the active materials can reach 98.6%. The method can achieve complete and efficient separation of aluminum foil and active materials from the cathode material with PAA-NH₄ as the binder, and has certain promotion and application value.

The leaching of zinc and germanium from germanium-rich zinc concentrate by oxygen pressure leaching process was studied. The effects of key factors such as the initial acidity of sulfuric acid, reaction temperature, leaching time, oxygen partial pressure, sodium lignosulfonate addition amount, liquid volume to solid mass ratio, and stirring speed on the leaching rates of zinc and germanium were systematically investigated. The optimal process conditions for oxygen pressure leaching are determined through single-factor experiments as follows:leaching temperature is 160 ℃, leaching time is 150 min, initial acidity of sulfuric acid is 160 g/L, partial pressure of oxygen is 1.2 MPa, addition amount of sodium lignosulfonate is 1 g, liquid volume to solid mass ratio is 8/1, and stirring speed is 600 r/min. Under the conditions, the germanium leaching rate reaches 94.86% and the zinc leaching rate reaches 99.98%. Compared with the traditional atmospheric pressure leaching, the optimized oxygen pressure process can increase the germanium leaching rate by more than 15%, shorten the leaching time by 40%, and also avoid the emission of high-concentration H₂S gas, reducing environmental pollution and providing a technical reference for the comprehensive utilization of valuable elements in zinc concentrate.

Due to its late discovery, wide distribution, large reserves, complex composition, and low grade, the comprehensive development and utilization of clay type lithium mines still lack systematic research. A systematic study was conducted on the optimization of enhanced leaching process parameters for clay type lithium ore in Liuzhi Special District, Liupanshui City, Guizhou Province. The effects of roasting process conditions such as roasting temperature and time, as well as the effects of leaching process conditions such as leaching agent hydrochloric acid concentration, leaching temperature and time, stirring speed, and liquid volume to solid mass ratio on the experimental results were investigated. The results show that roasting can destroy the crystal lattice of minerals such as kaolinite, superphosphate, diopside, and lithium chlorite in clay type lithium ores, which is beneficial for the sufficient contact between lithium containing phases and leaching agents during subsequent leaching processes, thereby improving their leaching efficiency;After high-temperature roasting at 600 ℃ for 90 min, the clay type lithium ore was subjected to optimal leaching conditions of 60 min, leaching temperature of 80 ℃, hydrochloric acid concentration of 4 mol/L, stirring speed of 400 r/min, and liquid volume to solid mass ratio of 6∶1. The lithium leaching rate can reach 93.48% and the leaching effect is good. The method does not require any additives during the calcination process, which can reduce pollution and lower costs. It has certain promotion and application value.

Zircon sand often contains minerals such as kyanite, thorite and thorianite, which contain trace amounts of radioactive elements such as uranium and thorium. During the production of zirconium oxychloride, these radioactive elements are concentrated in the waste residue, which has certain recycling value. To meet the demands of comprehensive utilization and radioactive pollution control of such waste residues, the process conditions for extracting valuable elements such as uranium and thorium from the waste residues were studied, and the effects of different leaching factors and parameters on the leaching efficiency of uranium and thorium were investigated. The results show that under the conditions of sulfuric acid concentration of 4 mol/L, leaching time of 4 h, liquid volume to solid mass ratio of 4∶1, and leaching temperature of 90 ℃, the leaching rates of uranium and thorium reach 99.9% and 88.6%, respectively. The result is better.

In order to efficiently enrich indium (In) in zinc oxide dust and realize the directional separation and recovery of associated elements, Zn, Fe, Pb and In were leached from the dust by choline chloride-oxalate dihydrate low eutectic solvent, and the “hydrolytic removal of zinc and lead-ultraviolet light reduction iron removal” process was used to remove impurities step by step. The results show that the leaching rates of Zn, In, Pb and Fe can reach 95.52%, 88.91%, 80.82% and 89.13%, respectively, under the optimal leaching conditions of liquid volume to solid mass ratio of 12 mL/1 g, temperature of 70 ℃ and time of 6 h. Zn/Pb is removed by hydrolytic precipitation, and then Fe is removed by ultraviolet reduction precipitation. After the step-by-step precipitation and impurity removal treatment, the mass fraction of indium in the solution has increased from 0.03% to 1.5%, successfully achieving efficient enrichment of indium and directional separation and recovery of associated elements such as Zn, Pb, and Fe.

The leaching of valuable metals from low-grade lead matte using hydrochloric acid and hydrogen peroxide mixed solution was studied. The effects of initial mass concentration of hydrochloric acid, hydrogen peroxide dosage, ratio of liquid volume to solid mass, leaching temperature and time on the leaching rates of copper, iron and lead were investigated. The results show that under the optimum conditions of hydrochloric acid mass concentration of 150 g/L, hydrogen peroxide to lead matte mass ratio of 3∶1, liquid volume to solid mass ratio of 9∶1, leaching temperature of 50 ℃ and leaching time of 2 h, the leaching rates of Cu, Fe and Pb can reach 90.21%, 33.49%, 31.19%, respectively, and Au and Ag are enriched in the resiclue. Efficient separation from copper is achieved.

For the raw material of quartzite-type quartz ore in Gansu Province, the pretreatment process of “magnetic separation-flotation-roasting and water quenching” was adopted, and then the purification process was carried out by atmospheric pressure acid leaching. The effects of leaching agent type, leaching temperature, liquid volume to solid mass ratio and leaching time on the impurity removal rate of quartz sand were investigated. The research results show that in the mixed acid system of 2 mol/L HF+2 mol/L HCl+2 mol/L HNO₃, under the conditions of leaching temperature of 80 ℃, liquid volume to solid mass ratio of 3 mL/1 g, and leaching time of 7 h, the purity of SiO₂ in the product can be increased from 99.80% of the raw ore to 99.98% and the total impurity removal rate can reach 90%. The results of leaching kinetics analysis indicate that the leaching processes of Al and Fe are controlled by internal diffusion, with apparent activation energies of 7.2 and 21.1 kJ/mol, respectively. The method can provide a certain technology reference for the preparation of high-purity quartz from quartzite-type quartz ore.

The selective separation of manganese and cobalt from the leaching solution of spent ternary lithium batteries by nickel saponification extraction was studied. The effects of extraction agent concentration, pH, phase ratio, saponification rate, and acid concentration on the extraction and stripping performance were systematically investigated. The results show that after adjusting the pH to 4.8 with Ni(OH)₂ and significantly reducing the contents of Fe, Al, and Cu, nickel saponification P204 can achieve efficient extraction of manganese under two-stage cross-flow conditions, with an extraction rate exceeding 99%. Meanwhile, the co-extraction rates of cobalt and lithium are controlled at 18.67% and 9.89%, respectively. After low-acidity washing and sulfuric acid stripping, the stripping rate of manganese reaches 99.96%. Further, nickel saponification P507 is used for two-stage cross-flow extraction of cobalt, achieving a cobalt extraction rate of 99.92% and a co-extraction rate of lithium of 13.68%. After washing and stripping, the stripping rate of cobalt can reach 99.87%. The process significantly reduces the introduction of Na⁺ impurities while achieving efficient separation of manganese and cobalt, leaving nickel and lithium in the system. It provides experimental technical references for the high-purity recovery of multi-metal resources and has potential for further industrial application.

To address the difficulty of separating copper and cobalt coexisting in the leaching solution of low-grade sulfide concentrates after roasting, a combined process of synergistic extraction-stripping-chemical precipitation was investigated to recover copper and cobalt. The results show that using P204+P507 as the extractant achieves copper extraction rate of 99.31% under the conditions of original station pH = 3, V_o∶V_a = 1∶1, and reaction time of 10 min. When N235 + C272 was used as the extractant at copper extract residue pH=4, V_o∶V_a = 2∶1 and reaction time of 10 min, the cobalt extraction rate reaches 99.63%. Using 0.2 mol/L H₂SO₄ as the stripping agent results in stripping efficiencies of 99.65% for copper and 98.65% for cobalt. With NaOH and Na₂CO₃ as precipitants, Cu(OH)₂ with a purity of 99.90% and CoCO₃ with a purity of 99.99% are obtained. The combined separation process can achieve efficient separation of copper and cobalt and produce high-purity products, providing a feasible technical approach for the comprehensive utilization of low-grade sulfur concentrate resources and having certain significance in alleviating the shortage of copper and cobalt resources in China.

Aluminum ammonium reaction cooling crystallization technology was studied to extract aluminum from purification slag of typical ion-type rare earth ore and prepare ammonium alum. The effects of ammonium sulfate dosage, reaction temperature, reaction time, and cooling crystallization temperature on the aluminum crystallization rate were investigated. The results show that under the conditions of molar ratio of ammonium to aluminum of 1.3/1, mass concentration of ammonium sulfate solution of 760 g/L, reaction temperature of 60 ℃, reaction time of 120 min, and cooling temperature of 5 ℃, the aluminum crystallization rate is 94.8%, and the obtained ammonium alum product meets the national standard. The process can effectively achieve the efficient extraction and recovery of aluminum elements from the impurity slag, and has good economic benefits.

Preparation of sulfonic acid-based functionalized biochar (PPBC) by modifying discarded pomelo peels through low-temperature sulfonation method and its application in adsorbing iron ions from high-concentration sulfuric acid solutions was investigated. The effects of adsorption time, temperature, stirring speed, initial mass concentration of iron ions and acidity on the adsorption performance of PPBC were examined, and the adsorption and separation mechanism as well as the recycling and regeneration performance were discussed. The results show that under the optimized conditions of sulfuric acid concentration of 0.1 mol/L, temperature of 55 ℃, stirring speed of 200 r/min, initial mass concentration of iron of 1 000 mg/L, and adsorption time of 120 min, the adsorption rate of PPBC for iron ions reaches 98.2%, and the adsorption capacity is 98.2 mg/g. When the initial mass concentration of iron ions is 5 000 mg/L, the adsorption capacity of PPBC for iron ions reaches a maximum of 324.6 mg/g. The adsorption behavior conforms to the Langmuir isothermal adsorption model and the pseudo-second-order kinetic model, and belongs to monolayer chemical adsorption. After five adsorption-desorption cycles, the regeneration rate of PPBC remains at 87.8%. PPBC has excellent structural stability and reusability, which can provide technical references for the treatment of high-acid and high-iron wastewater and the high-value utilization of agricultural waste.

To enhance the reaction rate of V₂O₃ preparation by hydrogen reduction from V-rich solution, Raney Ni was synthesized by selectively leaching aluminum from a Ni-Al alloy using an alkaline leaching method. This Raney Ni was employed as a catalyst to reduce the activation energy for the hydrogen reduction vanadium precipitation reaction. Characterization results from XRD, SEM, and BET indicate that the leaching of aluminum creates numerous pores and fissures, significantly increasing the specific surface area of the Raney Ni and providing abundant active sites for the hydrogen reduction vanadium precipitation reaction. Under the optimal conditions of Raney Ni dosage of 10 g/L, hydrogen partial pressure of 3 MPa, reaction temperature of 250 ℃, solution pH of 5, and reaction time of 100 min, the vanadium precipitation rate reaches 99.84%. The Raney Ni demonstrates excellent cycling stability, after magnetic separation and five reuse cycles, the purity of the obtained V₂O₃ product remains stable at 99.27%. The apparent activation energy for the Raney Ni-catalyzed hydrogen reduction reaction is determined to be 96.254 4 kJ/mol, indicating that the hydrogen reduction process is primarily controlled by chemical reaction. Raney Ni catalyst exhibits effective vanadium precipitation performance, ease of magnetic separation, and high stability. The V₂O₃ product prepared by catalyzed hydrogen reduction with it have high purity, and certain potential for promotion and application.

The leaching of arsenic with air oxidation method in an alkaline system was studied. Based on the results of thermodynamic analysis, the influence of different factors on the leaching process was investigated. The leaching residue was characterized by XRD and SEM, and the leaching kinetics was discussed. The results show that under the optimized conditions of initial NaOH concentration of 2 mol/L, leaching temperature of 60 ℃, liquid volume to solid mass ratio of 10 mL/1 g, and leaching time of 168 h, the arsenic leaching rate reaches 91.9%. The leaching rate of As is jointly controlled by diffusion and chemical reactions, with apparent activation energy of 83.41 kJ/mol. The method can achieve the purpose of arsenic removal under normal pressure in an alkaline system and has certain value for promotion and application.

Dealkalization of vanadium extraction tailings using calcium oxide was investigated. The effects of calcium oxide dosage, reaction temperature, reaction time, liquid volume to solid mass ratio, and stirring speed on the dealkalization rate were investigated, and the removal behavior of major metals during the dealkalization process was explored. The results indicate that under the optimal conditions of calcium oxide dosage of 19%, reaction temperature of 200 ℃, reaction time of 60 min, liquid volume to solid mass ratio of 2∶1, and stirring speed of 800 r/min, the dealkalization rate of the tailings reaches about 85%, with the Na₂O content in the residue reduces to less than 0.6%. Meanwhile, the process ensures relatively low removal rates for vanadium and chromium. The findings of the study can serve as a technical reference for the comprehensive resource utilization of vanadium extraction tailings.