Modified biochar (BC) technology can significantly enhance the adsorption performance and selectivity for U(Ⅵ) in uranium-containing wastewater by optimizing pore structures, introducing specific functional groups, and constructing a multi-functional synergistic system. This review summarizes the research status of modified biochar in the treatment of uranium-containing wastewater in recent years, with a focus on modification strategies such as activation, surface functionalization (e.g., oxidation, phosphorylation, amination), as well as the loading of novel composite materials including nano-zero-valent iron, layered double hydroxides, and metal-organic frameworks. The adsorption mechanisms for U(Ⅵ), including synergistic effects such as ion exchange, surface complexation, and reductive immobilization, are analyzed. Meanwhile, the advantages, disadvantages, and applicable scenarios of various modification methods are summarized. Future research directions are proposed, aiming to provide theoretical reference and technical support for the engineering application of modified biochar in the treatment of uranium-containing wastewater.

Conventional processes for extracting and processing both mineral resources and secondary resources are generally plagued by slow reaction rates, low efficiency, high energy consumption, and environmental pollution, making it difficult to meet the industrial demands of green and sustainable development. As an efficient physical field intensification method, ultrasonic technology has demonstrated significant advantages in the field of resource extraction, leveraging the synergistic effects of its unique cavitation, mechanical, and thermal effects. The research progress of ultrasonic technology in the flotation and leaching of metallic ores, the processing of non-metallic ores, and the recycling of secondary resources is systematically reviewed, and the effectiveness and mechanisms of its application in flotation and leaching processes are systematically summarized. Finally, the current bottlenecks of ultrasound-assisted extraction technology are summarized, and future research directions are proposed. This review aims to provide references for theoretical innovation and engineering application of ultrasonic technology in the comprehensive utilization of mineral resources and secondary resources.

Low-nickel matte is an important intermediate product in the pyrometallurgical processing of nickel sulfide ores. In order to achieve efficient recovery of valuable metals and reduce the losses of cobalt and nickel associated with conventional converter blowing, this paper systematically reviews the research progress on leaching technologies for low-nickel matte. Emphasis is placed on four representative processing routes: roasting pretreatment-leaching processes (including ammonium chloride roasting, calcification roasting, and sulfation roasting), chlorination leaching (FeCl₃ and FeCl₃-HCl systems), pressure hydrometallurgical leaching (oxygen pressure acid leaching and oxygen pressure water leaching), and atmospheric pressure intensified leaching (mixed-acid systems and persulfate-based oxidative systems). The reaction mechanisms, key advantages, and inherent limitations of each technique are comparatively analyzed. Furthermore, under the framework of green metallurgy, future development directions are discussed, including the construction of advanced oxidation systems, the introduction of physical intensification methods such as ultrasound, and the integration of machine learning for process modeling and optimization. This review aims to provide a valuable reference for the development of efficient and environmentally benign hydrometallurgical processes for low-nickel matte.

Spent vanadium catalysts, mainly generated from the petrochemical industry, are classified as hazardous waste. However, they also serve as important secondary resources due to their higher contents of valuable metals such as vanadium, tungsten, molybdenum, and nickel compared to natural ores. This paper reviews the recent research progress in hydrometallurgical recovery processes of spent vanadium catalysts. It focuses on analyzing the principles, experimental conditions, advantages, and disadvantages of acid leaching, alkali leaching, and roasting-leaching processes for vanadium extraction. Furthermore, the technical routes for preparing high value-added products from recovered vanadium are summarized, including vanadium electrolyte for vanadium redox flow batteries, VO₂(B) and K_0.51V₂O₅ as cathode materials for aqueous zinc/potassium-ion batteries, and Ag₃VO₄-based photocatalysts. Finally, future research directions for the resource utilization and recycling of spent vanadium catalysts are proposed.

Aiming at a carbon-in-leach tailings with relatively high gold grade in Hebei Province, based on the analysis of physicochemical properties and gold chemical phase, the process of ultrafine grinding-chemical pre-oxidation-cyanide leaching was adopted to extract gold from the tailings. The effects of ultrafine grinding fineness, pulp pH, sodium cyanide dosage, leaching temperature and time on gold extraction were investigated, and the kinetics of the chemical pre-oxidation-cyanide leaching process was discussed. The results show that under the conditions of ultrafine grinding fineness of -10 μm accounting of 80%, pre-oxidation of 14 h at an air flow rate of 0.3 m³/h, followed by cyanide leaching with sodium cyanide dosage of 1.5 kg/t, leaching time of 24 h, pulp pH of 12.5, leaching temperature of 303 K, and air flow rate of 0.1 m³/h, the gold leaching rate can reach 70.23%. The kinetic results indicate that the gold leaching process conforms to the classical unreacted shrinking core model, and the leaching reaction rate is controlled by the mixed control of interfacial chemical reaction and diffusion. The SEM and gold chemical phase analysis of the leaching residue show that there exist extremely finely disseminated gold or lattice gold in the carbon-in-leach tailings, which are difficult to leach. The research results can provide certain theoretical guidance and technical reference for further gold extraction from these tailings.

Aiming at the complex process, severe co-precipitation of impurities, and environmental pollution problems existing in the recovery process of platinum and palladium from silver anode slime, extracting platinum and palladium from electrolytic silver anode slime by selective nitric acid leaching-complexation directional precipitation was studied. The test results show that under the optimized conditions of nitric acid concentration of 40%, temperature of 90 ℃, and liquid volume to solid mass ratio of 4∶1, almost all impurities such as silver, copper, and bismuth are completely leached, and the leaching rates of platinum and palladium reach 89.97% and 99.17%, respectively. Subsequently, under conditions of temperature of 70 ℃, pH=0.5, and dimethylglyoxime addition amount of 2.5 times the theoretical addition amount, directional precipitation is carried out, and the precipitation rates of platinum and palladium are 88.45% and 99.78%, respectively, achieving high selectivity enrichment of platinum and palladium and effectively inhibiting the co-precipitation of impurities such as silver and copper. The expanded test results show that the direct recovery rates of platinum and palladium by this process are 78.32% and 99.05%, respectively, and a concentrate with platinum and palladium contents of 11.43% and 28.56% can be obtained. This method achieves efficient enrichment of platinum and palladium through the high selectivity interaction of the complexing agent, and there is no wastewater discharge, providing a new strategy for the clean and efficient recovery of platinum and palladium resources from silver anode slime.

Gold cyanidation barren solution is rich in copper/zinc cyano-complexes and possesses high recovery value. However, existing processes typically require the conversion of copper cyano-complexes into compound precipitates prior to smelting recovery, resulting in a relatively complex flowsheet and high energy consumption. To address this issue, a coupled process integrating ferrous precipitation and zero-valent aluminum reduction is proposed to achieve barren solution purification and valuable resource recovery. In this process, zinc cyano-complexes are first selectively precipitated by ferrous ions for zinc recovery; then, copper cyano-complexes are reduced by zero-valent aluminum assisted by ferrous ions to recover elemental copper; finally, ferrocyanide is precipitated by ferrous mass ions to recover Prussian blue. The results indicate that after treatment by this method, the total zinc concentration in the solution decreases from 1.95 g/L to 0.01 g/L, total copper mass concentration from 2.25 g/L to 0.03 g/L, and total cyanide mass concentration from 10.30 g/L to 0.20 g/L. The obtained zinc sludge contains over 22% zinc, the copper powder contains over 90% elemental copper, and the iron sludge contains over 50% Prussian blue. Economic analysis shows that the comprehensive reagent cost is approximately 78.4 CNY/m³ of barren solution, while the direct revenue from the produced elemental copper sludge is about 200 CNY/m³. This method provides a technical reference for the resource recycling and utilization of cyanide gold leaching barren solution.

In order to realize the further recovery of valuable metals in the alkali-washed copper smelting dust residue, the leaching of copper and zinc from alkali-washed residue by sulfuric acid was studied. The effects of leaching time, sulfuric acid initial mass concentration, liquid volume to solid mass ratio, leaching temperature and stirring rate on the leaching rate of copper and zinc were investigated. The phase composition of acid leaching slag was analyzed by XRD. The results show that the mass fractions of Cu and Zn in the alkali-washed residue are 13.86% and 12.50%, respectively, mainly in the form of oxides that are easily soluble in acid. Under the optimal conditions of leaching time of 30 min, sulfuric acid initial mass concentration of 100 g/L, liquid volume to solid mass ratio of 8∶1, leaching temperature of 80 ℃ and stirring rate of 400 r/min, the leaching rates of Cu and Zn are 95.27% and 92.02%, respectively. The acid leaching residue is mainly composed of SiO₂, PbSO₄ and CaSO₄, which meets the requirements of subsequent oxygen-enriched side-blown lead smelting and reduces the amount of lead matte disposal. This method provides an alternative method for the resource recovery of fluorine and chlorine-containing copper smelting dust.

To address the problems of high energy consumption in pyrometallurgical recovery and low germanium leaching rate in atmospheric hydrometallurgical recovery of germanium from oxygen pressure leaching residues of zinc concentrates, the leaching of germanium in a high-temperature and high-pressure sulfuric acid system, using oxygen as the oxidant and sodium lignosulfonate as the activating agent was investigated. The effects of key parameters, including leaching temperature and time, mass concentration of sulfuric acid, oxygen pressure, liquid volume to solid mass ratio, and sodium lignosulfonate dosage on the leaching rate of germanium were examined. The results show that the oxygen pressure leaching process can enhance the oxidative decomposition of refractory sulfides such as sphalerite in the oxygen pressure leaching residue of zinc concentrates, destroy their crystal lattice structures, and enable the dissociation and leaching of isomorphous germanium. Under the optimized conditions of leaching temperature of 150 ℃, leaching time of 3 h, mass concentration of sulfuric acid of 180 g/L, oxygen pressure of 1.0 MPa, and liquid volume to solid mass ratio of 8 mL/1 g, and sodium lignosulfonate dosage of 3% by mass of the oxygen pressure leaching residue of zinc concentrates, the leaching rate of germanium reaches 72.17%, which is significantly improved compared with the atmospheric acid leaching method, and the leaching rate of zinc reaches 89.39%. This study provides technical reference for the green and efficient recovery of germanium from oxygen pressure leaching residue and is of great significance for ensuring the sustainable supply of germanium resources.

Waste aluminum electrolyte and desulfurization gypsum residue are typical hazardous solid wastes generated during the aluminum electrolysis process. To achieve the harmless disposal and resource recovery of these two hazardous solid wastes, a roasting-water leaching-flotation process was developed for the synergistic treatment of waste aluminum electrolyte and desulfurization gypsum residue to regenerate fluorite. The results demonstrate that under the optimal roasting conditions (roasting temperature of 800 ℃, mass ratio of waste aluminum electrolyte to desulfurization gypsum residue of 1∶2, and roasting duration of 120 min), calcium, lithium, and fluorine in the feedstocks are directionally converted into calcium fluoride and lithium fluoride, while sodium, aluminum, and sulfur are transformed into sodium sulfate and aluminum oxide. Subsequently, sodium sulfate and lithium salts are separated via water leaching. The water leaching residue is further subjected to a single-stage flotation to separate aluminum oxide from calcium fluoride. The conversion rate of fluorine reaches 99.17%, and the grade of regenerated calcium fluoride is 80.92%, which meets the requirements for "Lump Fluorite Ore, Third Grade (KK-3)" as specified in the national standard "Technical Specification for Fluorspar" (GB/T 44539—2024). The mass concentration of fluorine in the effluent is 5.37 mg/L (less than 10 mg/L), complying with the wastewater discharge standard. This process exhibits high fluorine fixation rate, a simple flow sheet, no secondary pollution, and high treatment efficiency, thereby demonstrating promising prospects for industrial application.