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.

Uranium serves as the "grain" for the nuclear industry and constitutes a strategic resource essential for both national and energy security. Currently, the global uranium supply relies primarily on conventional extraction technologies, such as open-pit and underground mining combined with heap leaching, as well as in-situ leaching. These well-established technologies, characterized by high maturity and relatively low costs, represent the primary methods for uranium extraction at present. However, conventional uranium mining is facing mounting challenges, including the gradual depletion of high-grade resources and increasing difficulty in developing low-grade deposits. Therefore, it has become imperative to expand uranium extraction technologies toward unconventional resources. Unconventional uranium resource extraction technologies aim to recover uranium from low-grade polymetallic associated ores, seawater, and salt lake brines. Although various technical pathways are still in the research and development stage, they have demonstrated significant potential for broadening resource boundaries and achieving comprehensive resource utilization. This paper systematically reviews the fundamental principles, research progress, and future development trends of uranium extraction technologies, aiming to provide a reference for technological innovation and strategic direction in the field of uranium extraction.

Niobium and tantalum are strategic rare metals, and their efficient separation and purification are crucial for the sustainable utilization of resources. The resource characteristics and industrial demand of niobium and tantalum are summarized, and the research progress of solvent extraction method in niobium and tantalum separation is focused on. The formation mechanism of niobium-tantalum fluoride complexes and their coordination behavior with extractants are discussed. The extraction performance of ketones, esters, amines, crown ethers, organic phosphonates, and novel functionalized ionic liquid extractants is summarized, and current challenges and development trends are outlined.

In order to improve the leaching efficiency of chalcopyrite, on the basis of the conventional heating leaching system, the pulse microbubble method is introduced to reduce the influence of the sulfide layer by improving the mass transfer efficiency in the leaching process. The kinetics is discussed, the morphology of the leaching residue is analyzed, and the mechanism of pulse microbubble-enhanced leaching is revealed. The results show that the pulsed microbubbles can significantly improve the leaching rate of chalcopyrite, and the increase rate rises with the decrease in microbubble size. Under the optimum conditions of conventional leaching, the copper leaching rate is only 53.6%, while under the same conditions, the copper leaching rate can increase to 62.48% and 80.40% when the pore size of the orifice plate is 94 and 75 μm, respectively. The SEM test results show that the pulsed microbubbles can effectively improve the aggregation degree and coverage area of the sulfide layer on the surface of chalcopyrite. The mechanism may be that the pulsed microbubbles have higher mass transfer efficiency and can accelerate the reaction rate. At the same time, the local high temperature and high pressure released during the burst can weaken the passivation of the sulfide layer during the leaching process, thereby improving the leaching effect of chalcopyrite.

In order to expand the development boundary of overseas low-grade copper-cobalt oxide ore, the leaching of copper and cobalt with sulfuric acid was studied for the Colluvium of MSESA orebody in Congo (DRC). The effects of various key factors on the leaching were investigated, and the economy of the process was evaluated and analyzed. The results show that the contents of copper and cobalt in the colluvium are 1.1% and 0.15%, respectively. The oxidation rate of the ore is high and the argillization is serious. The main gangue phases are talc, biotite, quartz, chlorite and dolomite. Copper mainly exists in the form of oxide, and cobalt exists in the form of cobaltite, cobalt carbonate and cobalt in gangue. Under the conditions of grinding fineness of 70% (－200 mesh), sulfuric acid addition of 7%, leaching temperature of 25 ℃, leaching time of 2 h, liquid volume to solid mass ratio of 2.5∶1, and stirring rate of 300 r/min, the leaching rates of copper and cobalt are 88.72% and 71.75%, respectively, and the acid consumption is 5.43 t/tCu. Cathode copper is produced by leaching－extraction－electrowinning process. The production cost is $3 450.618/tCu, accounting for only 36.322% of the sales price of cathode copper, which has good economic benefits.

The selective recovery of lithium from retired ternary lithium batteries by carbothermal reduction－two-step leaching (water leaching-acid leaching) process was studied. The effects of carbothermal reduction and leaching test conditions on lithium leaching were investigated. The results show that the cathode powder and activated carbon are compounded according to the ratio of carbon mass fraction of 27%. After roasting at 600 ℃ for 90 min, the roasting product is mixed with ultrapure water according to the solid-liquid mass volume ratio of 1/18, and then 83% of the theoretical amount of sulfuric acid is added dropwise. After leaching at 65 ℃ for 20 min, the lithium leaching rate is more than 98%, while the leaching rates of nickel, cobalt and manganese are less than 5%, and the lithium recovery effect is better.