Address to the issue of high fluorine content in the flotation concentrate of a super-large Beryllianite-type uranium-beryllium co-associated ore in Xinjiang,the mineral and elemental composition,as well as the mineral dissemination characteristics were studied.Flotation process of "floating fluorite first and then beryllium" was adopted for flotation.Grinding fineness and flotation reagent system were optimized through systematic flotation condition tests. The results show that for the raw ore with a beryllium grade of 0.435% and a grinding finness of -325 mesh accounting for 93%,under the conditions of 1.5 kg/t for sodium silicate in roughing,2 kg/t for NaOH,700 g/t for oxidized paraffin soap in roughing,and 500 g/t for swept oxidized paraffin soap in scavenging,The technical indicators of flotation with beryllium grade of 3.36% and recovery rate of 81.99% are obtained. The process can achieve effective enrichment of beryllium minerals.

Preparation of alumina from fly ash by hydrochloric acid method is one of the processes with significant industrial application potential.Therefore,studying the ionic structure in acid leaching solution of fly ash is of great significance for iron removal process in the method.The ionic structure of the AlCl₃-FeCl₃-FeCl₂-HCl-H₂O system in hydrochloric acid leaching solution of fly ash at pH values of 1.0,1.5,and 2.0 was investigated by combining thermodynamic calculation,quantum chemical calculation,and Raman spectroscopy. The results show that at pH = 1.0,the main forms of Fe and Al complex ions in the system are [FeCl]⁺,[FeCl₂]⁺ and [AlCl]²⁺;when the pH rises to 1.5 and 2.0,[FeCl]⁺,[FeCl₂]⁺ and [AlCl]²⁺ transform into hydrolysis products such as [FeOH]⁺,[FeOH]²⁺ and [AlOH]²⁺ with hydroxyl ligands.The wave function analysis results indicate that at low pH,due to the high concentration of chloride ions,the chloride complexes are more stable.As the pH increases,the concentration of hydroxide ions increases,and the formation of high-coordination hydroxyl complexes becomes easier due to their higher metal-oxygen bond order and lower Gibbs free energy.

In view of the problem of secondary environmental pollution caused by acid leaching system for recycling waste lithium cathode materials,a green leaching system for obtaining cobalt carbonate with high efficiency was studied.The cobalt from waste lithium batteries was recovered by ammonia leaching—ion exchange resin method.The results show that under the conditions of ammonia concentration of 5 mol/L,NH₄Cl concentration of 0.7 mol/L,(NH₄)₂SO₃ concentration of 0.5 mol/L,temperature of 140 ℃,liquid volume to solid mass ratio of 20 g/1 L and reaction time of 50 min,the leaching rates of Co,Li and Ni are 88.0%,90.0% and 92.5%,respectively.Under the condition of temperature of 20 ℃ and flow rate of 1 mL/min,150 mL of ammonia leaching solution is selectively adsorbed by 40 g CH-90 resin,and the adsorption capacity of Co is 9.06 mg/g.FT-IR,SEM and XPS characterization results show that cobalt is exchanged with Na⁺ in the functional group of the resin and adsorbed on the resin in the form of [Co(NH₃)_i]²⁺.CoCO₃ products with purity exceeding 97.5% are obtained by using sodium carbonate precipitation to recover cobalt carbonate,after elution,precipitation and impurity removal.

The recovery of valuable components manganese and lithium from retired lithium manganese oxide batteries by carbon-thermal reduction—acid leaching combined recovery process was studied.The mixed powder of lithium manganese oxide and graphite was roasted by carbon thermal reduction,and the roasted products were characterized by XRD,XRF,SEM,TG-DTA and other technologies.The results show that the best effect is achieved by roasting at 650 ℃ for 180 min,and the lithium manganese oxide in the roasted product is completely converted into manganese monoxide and lithium carbonate.The lithium carbonate in the sample powder can be extracted by water leaching,and the leaching rate of lithium carbonate is 86.15%.Manganese ions are extracted by acid leaching of sulfuric acid.Under the acid leaching conditions of acid leaching concentration of 3.5 mol/L,acid leaching temperature of 60 ℃,acid leaching time of 3 h,and liquid volume to solid mass ratio of 8/1,the highest leaching rate of manganese ions is 88%.The method can achieve the purpose of synchronous and efficient recovery of manganese and lithium from cathode materials,and has certain application value.

Extraction and separation of valuable elements from the positive leaching solution of spent lithium-ion batteries by β-diketone/phosphate extraction system was studied.The optimal conditions for the extraction and separation of cobalt,nickel,manganese,and lithium were determined through equilibrium extraction.The results show that the β-diketone/phosphate extraction system can effectively separate cobalt,nickel,manganese,and lithium from the positive leaching solution of spent lithium-ion batteries by controlling the kinetics.Under optimized conditions,the extraction rate of cobalt,nickel,and manganese can reach 99%,and the yield of lithium can reach more than 95%.The method realizes the separation and recovery of cobalt,nickel,manganese and lithium by a single extraction system,which can provide a new process route for the recovery of waste ternary lithium batteries.

Iron and lithium were recovered from the cathode material of spent lithium iron phosphate battery using choline chloride,ascorbic acid,and ethylene glycol as a ternary deep eutectic solvent.The effects of the molar ratio of choline chloride/ascorbic acid/ethylene glycol,liquid volume to solid mass ratio,reaction temperature and time on the leaching rate of iron and lithium were investigated.The leaching mechanism was discussed through kinetic analysis and SEM characterization.The results show that under the optimal leaching conditions of choline chloride/ascorbic acid/ethylene glycol molar ratio of 1∶1∶3,liquid volume to solid mass ratio of 0.1 mL/1 mg,reaction temperature of 80 ℃ and reaction time of 1 h,the leaching rates of lithium and iron can reach 96% and 98%,respectively.The leaching process is mainly controlled by chemical reactions.The method is efficient and environmentally friendly,and can recover iron and lithium from spent lithium iron phosphate batteries.

Complex component sandstone-uranium deposits are mainly composed of conglomerate,sandstone and Slate.The uranium minerals are mainly uranite and titanium-uranium ores,including a small amount of pitchblende.The process mineralogy of complex component sandstone uranium ore was studied,and the leaching process was optimized.The results show that most of the uranium in the ore exists in the tetravalent form,and it contains a relatively large amount of calcium,magnesium,aluminium,iron and carbonate.The test results of acid leaching,enhanced leaching and column leaching show that the acid leaching process has a better effect. When 40~50 g/L H₂SO₄ is used as the leaching agent,the slag leaching rate of the acid column leaching is all greater than 90%. Comprehensively considered,it is recommended that the heap leaching process of -10 mm particle size ore be adopted in industrial production,and the mass concentration of the leaching agent H₂SO₄ is preferably 40 g/L.

The leaching of silver from failed silver-containing spent catalyst using HNO₃+H₂O₂ system was investigated.The kinetics of the silver leaching process was analyzed using the shrinkage kinematics model of liquid-solid phase reaction.The effects of leaching temperature and nitric acid concentration on the leaching rate of silver were examined.The results show that under the optimal leaching conditions of nitric acid concentration of 1.1 mol/L,leaching temperature of 50 ℃,stirring speed of 300 r/min,n(H₂O₂)∶n(Ag)=1.5∶1,leaching time of 50 min,and liquid volume to solid mass ratio of 4 mL/1 g,the leaching rate of silver can reach 94.18%.The leaching is controlled by the diffusion of the solid film,and the apparent activation energy of the leaching reaction is 15.45 kJ/mol,and the reaction order of hydrogen ion is 1.13.The method can provide reference for the research of efficient resourcing utilization of silver-containing spent catalyst.

Extraction of copper from refractory copper oxide ore using a roasting—acid leaching process was investigated.The effects of roasting and leaching conditions on the copper leaching rate were examined. The results show that under the optimal conditions of -200 mesh grinding fineness of 70%,roasting temperature of 850 ℃,roasting time of 1 h,coal addition of 8%,liquid volume to solid mass ratio of 2/1,H₂SO₄ concentration of 15%,leaching temperature of 60 ℃ and leaching time of 3 h,the copper leaching rate can reache 91.03%.The process is demonstrated to be economically efficient for extracting copper from refractory copper oxide ore and is considered to have potential for broader application.

Synthesization of mesoporous γ-AlOOH adsorbent via direct aging-ammonium salt substitution combined method in the presence of a desalting agent using metallurgical alumina hydroxide as raw material was investigated,and it was used for the adsorption of Congo red in wastewater.The physical phase and microscopic morphology of mesoporous γ-AlOOH were characterized by XRD,FT-IR,SEM,BET-BJH methods.The results show that the adsorption amount of Congo red by mesoporous γ-AlOOH adsorbent can reach 586.78 mg/g and the removal rate is 97.80% under the conditions of temperature of 25 ℃,adsorbent dosage of 100 mg,Congo red mass concentration of 300 mg/L,adsorption time of 180 min and pH=4.The adsorption process is more consistent with the pseudo-second-order kinetic model and the Langmuir isothermal adsorption model.The saturated adsorption capacity of mesoporous γ-AlOOH on Congo red is 1 965.265 mg/g at room temperature,and the adsorption process is spontaneous,heat absorption and chaotic.The main adsorption mechanism is the formation of hydrogen bonding between the adsorbent and adsorbate.

In view of the problems of high energy consumption,high equipment requirements and low flexibility in recovering rhenium from processing waste by traditional pyrometallurgical processes,the electrochemical enhanced leaching—precipitation crystallization method was studied to recover high-purity KReO₄ crystals from rhenium secondary resources.The results show that when 22%~24% HNO₃ solution is used as the electrolyte,there is no obvious passivation during the electrolysis process,and the energy consumption is stable at about 3.0 kWh/kg.During the electrolysis,Re atoms at the hexagonal lattice sites on the anode surface lose electrons,combine with hydroxyl groups and transform through low-valent oxidation states of Re to bridge oxygen connected Re(Ⅱ),and finally enter the electrolyte in the form of \mathrm{R} \mathrm{e} {\mathrm{O}}_4^{-} after reacting with the acid.When potassium salt is used as the precipitant to recover Re elements in the electrolyte,under the conditions of crystallization temperature of 25 ℃,precipitant flow rate of 6 mL/min,stirring rate of 500 r/min and crystallization time of 30 min,the precipitated KReO₄ crystals are in the shape of polyhedral spindle,with good uniformity in particle size,high recovery rate and purity of 99.95%,which can meet the requirements for hydrogen reduction to prepare metallic rhenium.The method can effectively recover rhenium processing waste and has certain promotion value.

To address fiber shrinkage embrittlement and consequent mechanical degradation during surface modification of polyamidoxime (PAO) adsorbents,a "core-shell heterostructure stress transfer" strategy was proposed.A coaxial electrospinning technique was employed to fabricate PS@PAO nanofibers with a polystyrene (PS) flexible core and rigid PAO shell.Microstructural analysis results show that PS@PAO exhibits uniform core-shell architecture (≈200 nm diameter, ≈50 nm thickness) with a specific surface area of 6.22 m²/g,representing a 38% enhancement over pristine PAO fibers. Mechanical testing results demonstrate 13.8% and 30.1% improvements in tensile strength (0.66 MPa) and Young's modulus (34.84 MPa),respectively.Dynamic contact angle measurements show that favorable hydrophilicity with water contact angle decreasing from 30° to 21° within 1 s. When PS@PAO is used to adsorb uranium from seawater with pH of 8.0 and uranium mass concentration of 16 mg/L for 48 h,the adsorption capacity is 34.14 mg/g. Adsorption kinetics analysis results indicate compliance with the pseudo-second-order model,with chelation between uranyl ions ( \mathrm{U} {\mathrm{O}}_2^{2+}) and amidoxime groups identified as the dominant mechanism.Through comprehensive investigation of material architecture,uranium extraction performance,and adsorption mechanisms,this study can provide theoretical foundations and scalable fabrication guidance for developing high-stability marine uranium extraction materials.

The removal of uranium bound to organic matter in real uranium-contaminated soil from a certain mining area was studied by using a combined oxidation washing process.The removal effects of uranium by two different new oxidation-washing systems(EDTA-H₂O₂ and SDS-H₂O₂)were compared.The effects of key parameters such as pH,liquid volume to solid mass ratio,oxidant concentration,and washing agent concentration on removal rate of uranium were investigated through single-factor experiments.The process conditions were optimized by response surface methodology,and the optimal conditions were determined.The results show that the removal rate of uranium by the EDTA-H₂O₂ system is 52.8% under the conditions of pH=4,liquid volume to solid mass ratio of 15/1,H₂O₂ concentration of 3%,and EDTA concentration of 100 mmol/L.After optimizing the process conditions by response surface methodology,the removal rate can be increased to 56.3%.The removal rate of uranium by the SDS- H₂O₂ system is 26.8% under the conditions of pH=4,liquid volume to solid mass ratio of 10/1,H₂O₂ concentration of 3%,and SDS concentration of 20 mmol/L.After optimizing the process conditions by response surface methodology,the removal rate can be increased to 29.5%.The removal effects of the two oxidation-washing systems are significantly better than those of single washing agents and single oxidants (EDTA 24.12%,SDS 0.66%,H₂O₂ 13.81%).The process can effectively break the complex of organic matter and uranium,significantly improve the remediation efficiency of real uranium-contaminated soil,and provide a feasible solution for uranium pollution control in mining areas.

Secondary aluminum dross (SAD) is a hazardous waste generated during aluminum resource recycling,containing aluminum,alumina,aluminum nitride,fluorides,chlorides,and other components. It exhibits strong chemical reactivity and leaching toxicity,making its resource utilization and environmentally friendly recovery highly significant. The denitrification of SAD by alkaline roasting,the extraction of aluminum from the roasted residue via water leaching,and the defluorination of the leachate using CaCl₂ were studied.The phase transformation,elemental distribution,and microstructure of the roasted dross were analyzed by XRD and SEM-EDS.The results show that under optimal alkaline roasting conditions of m(NaOH)∶m(SAD)=1.1,roasting temperature of 800 ℃,and roasting time of 120 min,the nitrogen removal rate can reach 98.77%.Under the best water leaching conditions of leaching temperature of 70 ℃,liquid volume to solid mass ratio of 14∶1,and leaching time of 80 min,the aluminum leaching efficiency is 91.83%,with Al,AlN,and Al₂O₃ in the roasted residue being mostly converted into soluble NaAlO₂.For fluorine removal from the leachate,CaCl₂ was employed.The results indicate that under the conditions of n(Ca²⁺)∶n(F^-) = 0.7,reaction temperature of 50 ℃,and reaction time of 90 min,the defluorination rate can reach 94.87%. The fluorine removal residue primarily consists of CaF₂ and a small amount of unreacted CaCl₂,which can be used as a flux in metal smelting.The process achieves efficient aluminum extraction from SAD while removing nitrogen and fluorine,fulfilling the objectives of harmless treatment and resource recovery.

Aiming at the problems of high sulfur content and difficult utilization of high sulfur bauxite,the alkaline leaching desulfurization process of high sulfur bauxite was studied.The effects of alkali mass concentration,liquid volume to solid mass ratio,leaching temperature and time on the desulfurization effect were investigated,and the reaction kinetics were analyzed.The results show that under the conditions of base concentration of 180 g/L,liquid volume to solid mass ratio of 8 L/1 kg,leaching temperature of 160 ℃ and leaching time of 5 h,the bauxite sulfur mass fraction after desulphurization is 0.42%.The desulfurization process is controlled by reaction-internal diffusion,and the apparent activation energy is 18.23 kJ/mol.The method can effectively reduce the sulfur content in high sulfur bauxite and is beneficial to the wide utilization of high sulfur bauxite.

In order to solve the problem of real-time and accurate parameter optimization in hydrometallurgical equipment operation,an optimization setting compensation method based on real-time data acquisition was proposed by combining the improved POPOA method and the improved JITL online learning method.The results show that compared with the traditional method,the retraining time of the modified JITL method is significantly reduced,the optimization rate is significantly increased,and the energy consumption is significantly reduced.The improved POPOA method significantly improves the performance of real-time data processing,and the processing time is about 40% shorter than that of the traditional method.The improved POPOA method reduces the load rate of the system significantly compared with the traditional method when the multi-task is running concurrently.This method can effectively improve the accuracy of operation performance evaluation,the real-time response ability of the system,and the generalization ability of the model,and reduce energy consumption and operation cost,so it has a certain application prospect.

To address the issues of simplicity and weak generalization in current fault diagnosis models,an improved CNN-Bi-LSTM model is employed for fault diagnosis in hydrometallurgical processes.Based on the diagnostic results,an enhanced random forest model is utilized to evaluate the entire hydrometallurgical process.The results indicate that the fault diagnosis accuracy can reach 90.7%,significantly surpassing accuracy of the existing rule-based diagnostic system at the factory(78.4%).Additionally,the fault detection response time is maintained within 2 seconds,ensuring real-time monitoring and rapid response during the process.

The determination of alkyl mercury in water by distillation—purge and trap/gas chromatography-cold atomic fluorescence spectrometry was studied. The detection limit,precision and accuracy of the method were determined,and the factors affecting the effect of distillation and the optimum test conditions were determined.The results show that the detection limit of methymercury is 0.003 1 ng/L and that of ethylmercury is 0.002 8 ng/L.The adding standard recovery rates of both are 99.4%~104%,and the relative standard deviations(RSD) are 1.02%~1.34%. The factors affecting distillation effect are hydrochloric acid and saturated copper sulfate adclition.For 40 mL of pure water,the optimal addition amounts of concentrated hydrochloric acid and saturated copper sulfate are 80 and 200 μL,respectively.When the addition amount of alkylmercury are 4.00,40.0 and 400 pg,the recovery rates are all in line with the quality control requirements. The recovery rate of alkyl mercury adding standard can be significantly improved after distillation treatment in actual lake water.