Bauxite resources in China are predominantly sedimentary diasporic ores, characterized by well-developed clay minerals and the fine-grained intergrowth of aluminum-bearing minerals with siliceous gangue. These ores typically have high silica content and a low Al/Si ratio, making them unsuitable for direct processing by the Bayer method. With the depletion of high-grade resources, efficient desilication and beneficiation of low-grade bauxite have become essential for ensuring the security of the alumina industry. This paper systematically reviews desilication technologies for low-grade bauxite. It analyzes the distribution of bauxite resources and the characteristics of the ores, and further clarifies the constraints imposed by mineralogical features and mineral intergrowth on beneficiation performance. The review focuses on physical beneficiation methods, including gravity separation and flotation, as well as chemical methods, microwave activation, and the calcification-carbonation process. Their desilication efficiency, economic cost, and environmental impact are comparatively evaluated. Flotation remains the dominant technology, but its selectivity is significantly reduced by fine-grained mineral intergrowth. Chemical methods and energy-assisted techniques can help overcome the limitations of mineral liberation, but they are often associated with high energy consumption and complex process flows. The calcification-carbonation process integrates desilication with red mud valorization, representing a promising pathway for green metallurgy. Finally, desilication technologies should be tailored to differences in mineralogical characteristics. Process selection should move beyond a single-indicator criterion toward a comprehensive evaluation of energy efficiency, silica transformation pathways, and solid-waste properties. This approach will promote the efficient, green, and high-value utilization of high-silica bauxite.

In view of the complex state of vanadium occurrence in vanadium resources, the high difficulty in extraction and separation, and the increasingly strict environmental protection requirements, this paper systematically summarizes the research progress of vanadium extraction processes at home and abroad, and focuses on discussing the technical characteristics and existing problems of representative vanadium extraction processes such as roasting-leaching, hydrometallurgical separation, and others. According to the properties of different raw materials such as vanadium-titanium magnetite and vanadium-cotaining shale, the suitable routes of the listed vanadium extraction processes are compared and analyzed. Proposing that future vanadium extraction processes should develop towards the direction of low pollution, high recovery rate, intelligence, short process, and efficient comprehensive utilization of resources, is an important way to promote the improvement and innovation of vanadium metallurgy technology.

Recovery of titanium from coal gangue is significant for both high-value utilization of solid waste and resource security. This review systematically examines the occurrence of titanium in coal gangue and its influence on enrichment and extraction processes. It focuses on the applicable conditions and technical bottlenecks of mainstream hydrometallurgical techniques, such as acid leaching, alkali fusion, and roasting-leaching. Comparisons are made with physical sorting and emerging green technologies, including bioleaching and ionic liquid extraction. Future research directions are also discussed, aiming to provide a reference for the economical, environmentally friendly, and efficient recovery of titanium resources from coal gangue.

Aiming at the problems of low molybdenum grade, complex ore properties and poor enrichment effect by conventional mineral processing methods for a certain molybdenum deposit, on the basis of ore property research, acid leaching, alkaline leaching and enhanced leaching tests were carried out respectively. The high-acid enhanced leaching process was optimally selected. The effects of sulfuric acid dosage, ore particle size, leaching temperature, oxidant dosage, liquid volume to solid mass ratio and leaching time on molybdenum leaching efficiency were investigated, and the optimal leaching process parameters were determined. The results show that under the optimal conditions of ore particle size of -100 mesh, sulfuric acid dosage of 20%, pyrolusite dosage of 3%, leaching temperature of 80 ℃, liquid volume to solid mass ratio of 1.5∶1 and leaching time of 3 h, the molybdenum leaching rate calculated by leaching residue reaches 71.18% via the high-acid enhanced leaching process. To solve the problem of excessive residual acid in high-acid enhanced leaching solution, a two-stage countercurrent leaching process is adopted. The average molybdenum leaching rate calculated by leaching residue increases to 72.16%, which is about 1 percentage point higher than that of single-stage leaching. Meanwhile, the mass concentration of residual acid decreases by more than 40%, and the total consumption of sulfuric acid reduces by 25%. The ton-scale pilot test results of the two-stage countercurrent leaching process indicate that under suitable conditions, the total sulfuric acid dosage is 15%, the residual acid mass concentration of leaching solution is 48.8 g/L, and the molybdenum leaching rate calculated by leaching residue remains above 71%. This process proves stable and reliable with good reproducibility of leaching indexes, which serves as the technical basis for the industrial design of this molybdenum deposit.

Aiming at the problems of poor heap leaching permeability, low molybdenum leaching rate and difficult solid-liquid separation in stirring leaching of low-grade uranium-molybdenum ore in Guyuan, a process path of "mixing acid curing-low temperature roasting" synergistic enhanced acid heap leaching was proposed, and the key parameters such as acid concentration, curing temperature and time, roasting temperature and time, and spray intensity were optimized under column leaching conditions. Considering the energy consumption and process effect, the recommended conditions are as follows：ore particle size is -1 mm, acid mixing mass fraction is 12%, curing at 90 ℃ for 3 h, roasting at 300 ℃ for 30 min, acidizing water mass concentration is 10 g/L, spray intensity is 76 L/(m²·h). Under these conditions, the leaching rate of molybdenum slag can reach 82.89%, the molybdenum grade of tailings is reduced to 0.038%, and the uranium grade is reduced to 100.33×10^-6. The results of multi-batch ore verification test show that the process flow has adaptability to uranium-molybdenum ore with fluctuating grade and property. The molybdenum leaching rate of different low-grade uranium-molybdenum ore is stable at 72.51%–77.37%, which can achieve the process goal. The research results can provide technical support for the efficient extraction of low-grade complex uranium-molybdenum ore resources.

Aiming at the characteristics of complex intergrowth and intimate association of various mineral phases in titanium-iron tailings, which lead to difficult mineral separation, this study adopts benzohydroxamic acid as collector, sodium hexametaphosphate as depressant and 2^# oil as frother to realize flotation separation of valuable components from titanium-iron tailings. The influences of pulp pH, benzohydroxamic acid dosage, sodium hexametaphosphate dosage, 2^# oil dosage and flotation stages on the mass proportion of flotation products, TiO₂ recovery rate, TiO₂ grade and total iron mass fraction were systematically investigated, and the corresponding reaction mechanism was further discussed. The results show that under the optimal conditions of sodium hexametaphosphate dosage of 500 g/t, benzohydroxamic acid dosage of 35 kg/t, 2^# oil dosage of 1 mL and pulp pH=9, the mass proportion of floated product reaches 25.85%, the TiO₂ recovery rate is up to 42.23%, the TiO₂ grade is 14.93%, and the total iron mass fraction is 15.22%. After six stages of flotation, qualified flotation concentrate is prepared, with the floated product mass proportion of 71.82%, TiO₂ recovery rate of 75.81%, TiO₂ grade of 27.84% and total iron mass fraction of 24.49%. XRD and FT-IR analysis results demonstrate that the depressant sodium hexametaphosphate can exert an obvious inhibitory effect on pyroxene phase and silica. Benzohydroxamic acid is capable of generating adsorption behavior on the surface of titanium-iron tailings particles, which remarkably improves the overall flotation separation effect.

An antimony diaphragm electrorefining technology suitable for tartrate system was developed using refined antimony as raw material. The effects of different diaphragm types and electrolysis process parameters including current density, electrolysis temperature, electrolyte circulation speed and electrode spacing on cathode current efficiency, cell voltage, direct current consumption, cathode antimony purity and typical impurity removal rate are systematically investigated. The optimum process conditions are determined as follows:tartrate ion mass concentration is 250 g/L, antimony ion mass concentration is 45 g/L, current density is 150 A/m², electrolysis temperature is 35 ℃, electrolyte circulation speed is 300 mL/min, electrode spacing is 70 mm. The research results show that according to the single point impurity content requirements of 5N grade high purity antimony in the national standard' high purity antimony' (GB/T 10117—2021), the impurities that are easy to exceed the standard in the antimony tartrate electrolysis system are mainly arsenic, bismuth, lead and copper. Only regulating the electrolysis process conditions can not achieve the deep removal of the above impurities. Further electrolyte purification measures are needed to ensure the batch stability of high purity antimony products, while the remaining 16 main control impurities can meet the requirements of the national standard Sb-05 by electrolytic refining. Under the optimal electrolysis conditions, the electrolyte is deeply purified with antimony-based adsorbent (polyhydroxyantimonic acid with the main phase of H₁₄Sb₁₄O₂₁(OH)₄₂) to obtain cathode antimony with a purity of 99.995% and a dense and bright surface. The average cell voltage is 0.808 V, the current efficiency is 96.27%, and the direct current consumption is 485.85 kWh/t.

Extraction of aluminum from fly ash by ammonium sulfate roasting-leaching process was studied. The effects of roasting temperature, roasting time and the molar ratio of ammonium sulfate to Al₂O₃ in fly ash on the extraction of aluminum from fly ash were investigated. The interaction of various factors on the extraction rate of aluminum from fly ash was discussed by response surface method, and the optimum roasting parameters were determined. The phase and microstructure of fly ash raw materials and calcined products were characterized by X-ray diffractometer (XRD) and scanning electron microscope (SEM). The results show that the main product of fly ash activated by ammonium sulfate is NH₄Al(SO₄)₂, and the surface of the roasted product is corroded. The optimum process parameters of activation roasting of fly ash are as follows：roasting temperature is 438 ℃, roasting time is 63.8 min, molar ratio of ammonium sulfate to Al₂O₃ in fly ash is 7.31. Under the optimum process conditions, the leaching rate of Al is 95.30%, and the leaching effect is better.

In view of the recovery and utilization of associated thorium resources and the control of radioactive pollution in the process of bastnaesite smelting, the migration and distribution characteristics of thorium elements in each link of smelting separation were systematically studied. After the pretreatment of bastnaesite by alkali decomposition, the leaching acidity was adjusted to 3-5 mol/L, and the thorium leaching rate was increased to 98%. Fe³⁺ in the feed solution was selectively removed by N235 extractant in advance, and then thorium was directional extracted by P507 extractant. The results showed that the saturated loading of thorium oxide in the diluted P507-sulfonated kerosene extraction system could reach 25 g/L. After oxalic acid back extraction, washing and burning at 1 000 ℃, the purity of thorium oxide in the final product reached 98.88%. The total yield of thorium is more than 96% for the raw material of bastnaesite containing 0.2% -0.3% thorium. The method is efficient and environmentally friendly, and has certain application potential in the release treatment of thorium-containing solid waste.

The iron oxide loaded dolomite composite (DFC) was prepared by ball milling-calcination method and used to degrade malachite green (MG) in water. The degradation performance and mechanism of DFC were investigated. The results indicate that DFC has good degradation ability for MG, and the degradation rate is affected by the adding amount of DFC, solution initial pH, temperature, and coexisting anions. Under the optimal conditions of DFC adding amount of 30 mg, initial pH=7, and temperature of 40 ℃, the degradation rate of MG can reach 96% within 10 min. Free radical trapping experiments show that superoxide radical () is the dominant active species, and hydroxyl radical (·OH) plays a synergistic role. Intermediate analysis reveals that MG mainly achieves deep degradation through demethylation, chromophore destruction, and aromatic ring breakage. This study can provide a technical reference for the application of natural mineral-based iron oxide composites in dye wastewater treatment.

Addressing the technical challenges of poor settling performance and low solid-liquid separation efficiency of the collapse deposits in the MSESA orebody of the Democratic Republic of the Congo, the process mineralogy and settling characteristics of these deposits was systematically analyzed. Furthermore, it investigates the mineralogical features and settling behaviors of both the collapse deposits and their mixed systems with the MSESA run-of-mine ore. The results indicate that the collapse deposits are low-grade copper-cobalt oxide materials, with copper mainly existing as free copper oxide and cobalt primarily hosted in gangue minerals. Settling tests determined the optimal parameters for different systems:the optimal flocculant dosage for both raw collapse deposits and the mixed ore is 20 g/t, with final settling times of 40 and 25 min, respectively. For the leaching slurries of the collapse deposits and the mixed ore, the optimal flocculant dosage is 120 g/t, with final settling times of 20 and 40 min, respectively. These findings provide a reliable basis for the resource utilization and process design of collapse deposits in the MSESA orebody.

Lead-zinc ore is the main raw material for zinc hydrometallurgy, and cadmium is one of the major associated metal elements in lead-zinc ore. During the neutral leaching process, cadmium enters the leachate, producing a high-cadmium zinc sulfate solution. Aiming at the common issue in the zinc hydrometallurgy industry of the lengthy process for cadmium separation and recovery via zinc powder cementation, the two-stage countercurrent purification of cadmium from high-cadmium zinc sulfate solution using zinc powder was investigated. The appropriate zinc powder addition rate and residue return enrichment system were explored to increase the cadmium content in the cementation residue. The results show that the suitable zinc powder addition for the first-stage purification is 0.9 times the theoretical amount, and for the second-stage purification, it is 1.2 times the theoretical amount. Under optimal conditions, the cadmium concentration in the solution after the first-stage purification is approximately 70 mg/L, and after the second-stage purification, it is below 5 mg/L. The main phase of the first-stage purification residue is cadmium, with a cadmium mass fraction of 85.11% at the purification endpoint and a zinc mass fraction below 5%. The main phases of the second-stage purification residue are metallic cadmium and zinc. SEM-EDS analysis shows that the particles of the first-stage cementation residue are spherical with an encapsulation phenomenon. The encapsulation layer consists of basic zinc sulfate, while the inner layer is mainly unreacted zinc powder. The findings of this study can provide a technical reference for the separation and recovery of cadmium from high-cadmium leach solutions.

In order to accurately control the reduction, nucleation and growth process of platinum atoms in the preparation of high purity and high dispersion platinum powder, high purity and high dispersion platinum powder was prepared by chemical liquid phase reduction method with Na₂PtCl₆ aqueous solution as precursor, N₂H₄·H₂O as reducing agent and PVP as dispersant. The effects of temperature, reaction system and feeding method on the morphology, particle size distribution and dispersion of platinum powder were investigated. The results show that the nucleation and growth process of platinum atoms can be effectively separated by using PVP as the dispersion system and Na₂PtCl₆ and N₂H₄·H₂O as the synchronous feeding method at 80 ℃. The high purity and high dispersion micron platinum powder with good dispersion and narrow particle size distribution (D₉₀-D₁₀=1.18 μm) is prepared. This method provides a reliable theoretical basis and technical path for the stable and mass production of high purity and high dispersion micron platinum powder.

In view of the complex technical problems such as "running muddy" in the thickening system, excessive Cu²⁺ in the solution and poor open circuit of Fe³⁺ in the atmospheric leaching process of high copper nickel matte, the migration behavior and inhibition mechanism of Cu and Fe were systematically revealed by mineralogy, chemistry and reaction kinetics. The mineral analysis system (MLA) was used to clarify the occurrence state and evolution path of raw materials and intermediate slag nickel sulfide (Ni₃S₂), copper nickel sulfide and iron oxide (FeOOH, etc.). The coupling effects of pH, temperature, liquid volume to solid mass ratio, reaction time and initial Cu²⁺ mass concentration on the leaching kinetics and hydrolysis precipitation behavior of Ni, Cu and Fe were investigated by designing systematic single factor and L₉(3⁴) orthogonal experiments. The results show that Fe³⁺ is significantly hydrolyzed at pH>3.0, and the generated Fe(OH)₃ colloid encapsulates the microparticles through electrical neutralization and bridging, which is the direct cause of system stability damage and "running muddy". In addition to direct leaching, Cu²⁺ has a hidden cycle path of "adsorption-coprecipitation", which is the core reason why it is difficult to completely open the circuit. Based on this, a "Fe/Cu synergistic regulation" strategy is proposed. The core of this strategy is to accurately control the initial pH in a narrow window of 2.5–3.0：in this interval, the catalytic effect of Cu²⁺can be used to promote nickel leaching, and the hydrolysis of Fe³⁺ into gel at the initial stage of the reaction can be inhibited to the greatest extent. Combined with orthogonal test optimization, the optimum process conditions are determined as follows：initial pH=2.5-3.0, temperature is 80-85 ℃, liquid mass solid mass ratio is 8/1-9/1, reaction time is 4-5 h. Compared to the traditional process, after applying this collaborative regulation strategy, the total nickel leaching rate of the system increases steadily from approximately 90% to over 93%. The Ni mass concentration in the externally supplied nickel sulfate solution reaches over 105 g/L, while the Cu mass concentration decreases from over 15 g/L to below 3 g/L, and the Fe mass concentration drops from over 5 g/L to below 1 g/L. This method solves the problems of "slurry leakage" and impurity control in industrial production effectively, and provides a solid theoretical basis and engineering practice solution for the precise regulation of the high-nickel concentrate hydrometallurgical process.

In recent years, several CO₂+O₂ in-situ leaching uranium mines in China have been facing final mining closure and decommissioning. Natural attenuation technology is an economically effective solution, and its feasibility has been demonstrated by multiple remediation cases of in-situ leaching uranium mines abroad. However, a systematic understanding of the natural attenuation patterns of groundwater uranium at in-situ leaching uranium mines has not yet been established. This study focuses on a typical post-mining area of a CO₂+O₂ in-situ leaching uranium mine in China. Based on field investigation and sampling, hydrochemical analysis and multivariate statistical methods were employed to determine the distribution characteristics and natural attenuation patterns of U in groundwater. The results show that the hydrochemical types of groundwater in the post-mining area has shifted to HCO₃·SO₄-Na type and multi-anion-Na type. The concentrations of groundwater U in the post-mining area are 0.11-7.99 mg/L and exhibit weak correlations with other major components. Natural attenuation of U concentrations is observed in 69% of the groundwater samples, with attenuation ratios ranging from 4.82% to 77.14% and attenuation rates ranging from 0.112 and 10.227 mg/(L·a). The research results demonstrate that this CO₂+O₂ in-situ leaching uranium mine possesses natural attenuation capacity, and natural attenuation technology can serve as a remediation scheme for groundwater in post-mining areas of in-situ leaching sites.

For high acid and high chlorine uranium-containing wastewater ([H⁺] = 1.05 mol/L, ρ(Cl^-)=120 g/L, uranium existing as ), the adsorption behavior and mechanism of H6, T1 acid- and chlorine- tolerant resin and 201×7 resin on uranium were compared and studied. The results show that the equilibrium adsorption capacity (24.9 and 23.2 mg/g) and adsorption rate of uranium for H6 and T1 resins are significantly higher than those of 201×7 resin. All three resins follow the pseudo-second-order kinetic model (R²=0.999), indicating that chemical adsorption dominates the process. The adsorption process of uranium on the three resins is more consistent with the Langmuir isothermal adsorption model, and the adsorption is mainly monolayer adsorption. However, the surface active sites of T1 resin are more uniform, and the adsorption selectivity is stronger (b=0.32-0.52 L/mg). The maximum adsorption capacity of H6 resin is the highest (27.9-31.0 mg/g), and its Freundlich fitting is also good (R²=0.971-0.982, n=3.4-4.0), showing characteristics of multi-molecular layer adsorption. The results of dynamic breakthrough tests show that the adsorption performance of H6 and T1 resins for uranium is better than that of 201×7 resin. In contrast, H6 resin has a faster adsorption rate, higher effective capacity utilization, and more engineering potential in the early stage. The results of leaching and desorption experiments show that, under the premise of similar leaching rates, the leaching efficiency at 2 BV/h (contact time 0.2 h) is higher and yields a qualified solution with higher uranium concentration. This study can provide some technical reference for resin screening and process optimization under extreme water quality.

A high-clay low-grade sandstone-type uranium ore, predominantly composed of clay fine sandstone, exhibits significant thickness variations in both the aquifer and aquiclude within the mining area, resulting in poor sand permeability and high groundwater salinity, which render in-situ leaching infeasible. To comprehensively evaluate the leachability of this ore, the applicability of acid/alkaline agitation leaching and column leaching processes was systematically investigated. The results show that under the acid agitation leaching conditions of sulfuric acid dosage of 16%, MnO₂ dosage of 1%, liquid volume to solid mass ratio of 1.6∶1-1.8∶1, leaching temperature of 60 ℃, and leaching time of 4 h, the uranium leaching rate reaches 95%. However, when the acid column leaching process is applied, the high carbonate content in the ore causes severe argillization and gas blockage, making the process unstable and difficult to operate. In contrast, under the alkaline agitation leaching conditions with a Na₂CO₃/NaHCO₃dosage of 2.5%, liquid volume to solid mass ratio of 1.0∶1-1.2∶1, temperature of 80 ℃, and leaching time of 16 h, the uranium leaching rate attains 88%. For the alkaline column leaching process, alkali-mixing and curing can maintain the permeability of the ore heap. Nevertheless, the overall leaching rate remains low owing to the poor leaching efficiency of coarse particles. Based on a comprehensive comparison, the alkaline agitation leaching process is recommended for the treatment of this ore, and enhanced pretreatment of coarse particles should be implemented to further improve the uranium leaching rate.

In order to solve the problems of low detection efficiency of phosphate in boiler water and cooling water of nuclear power plant, the automation operation has not been realized, and the operation risk of radioactive samples is high, a flow injection method for the determination of phosphate was investigated. The method is based on the reaction of phosphate with molybdate and metavanadate to form yellow phosphovanadium molybdic acid complex in acid medium, and the detection is carried out at the wavelength of 420 nm. The results show that the method exhibits good linearity in the phosphate mass concentration range of 0-500 mg/L (correlation coefficient R=0.999), with a method detection limit of 2.28 mg/L. The relative standard deviations at low, medium, and high concentration levels are 2.41%, 1.61%, and 1.33%, respectively. Compared with the conventional spectrophotometric method, the relative error ranges from -0.81% to 1.27%. This method meets the requirements for phosphate determination in nuclear power plant boiler feed water and cooling water. The method enables fully automated sample analysis, with an analysis cycle of only 100 seconds per sample, significantly improving detection efficiency and automation level while greatly reducing the risk of radioactive exposure to operators. It provides an efficient, accurate, and safe automated analytical technique for rapid and precise phosphate detection in the field of nuclear power chemistry, demonstrating important engineering application value.