China Nonferrous Metallurgy

Abstract:After the high-pressure acid leaching process for extracting nickel and cobalt from laterite nickel ore, a large amount of waste residue is produced, which contains valuable elements such as iron, nickel, cobalt, manganese, chromium, and rare earths. Among them, iron is a dominant resource, with a content of 40% to 60%, primarily existing in the form of hematite. However, this residue is characterized by fine particle size, complex mineral distribution relationships, and high sulfur content, making it extremely difficult to process. Currently, the main disposal method for leaching residue of laterite nickel ore is tailings pond storage, which not only incurs high construction costs but also wastes resources and poses significant risks of dam failure. Research on treatment methods for this residue primarily focuses on producing iron concentrate, molten iron, construction materials, and high-value-added materials. The magnetization roasting-magnetic separation method for producing iron concentrate offers advantages such as low energy consumption and simple processing, but the challenge of comprehensive tailings utilization remains unresolved. The reduction smelting method for producing molten iron is highly operable and achieves high recovery rates for iron and sulfur, but it suffers from high energy consumption, and its economic viability fluctuates with iron ore prices. The preparation of construction materials from leaching residue is simple and suitable for large-scale production, but it remains at the experimental research stage due to limitations imposed by the residue’s composition and properties. The production of battery materials from leaching residue offers high resource utilization and significant product value, but it involves lengthy processes, high acid/alkali consumption, and high equipment requirements. Finally, this paper puts forward suggestions for future resource recovery and utilization of leaching residue: broaden comprehensive utilization pathways by co-processing with other industrial waste to achieve waste treatment with waste; strengthen technological coupling and innovation to improve overall utilization efficiency; and develop new technologies, reagents, and equipment to promote high-value utilization of leaching residue.

Abstract:In the process of aluminum production, the carbon anode participates in the electrolytic reaction. The theoretical carbon consumption is 333kg/t-Al, but the real consumption of the carbon anode often exceeds this theoretical amount. By analyzing the consumption mechanism of the carbon anode during aluminum electrolysis, it is found that the extra consumption of the carbon anode includes chemical and mechanical consumption. The oxidation of the carbon anode at high temperatures is the main cause of both types of consumption, either directly or indirectly. Currently, the main technologies to improve the oxidation resistance of carbon anodes include substrate modification, solution impregnation, and anti-oxidation coatings. Specific measures for substrate modification include altering the calcination temperature of petroleum coke and the baking temperature of the carbon anode, adjusting the sulfur and trace element content in the carbon anode, and using additives in the production of anode carbon blocks. Solution impregnation technology uses impregnating agents such as aluminum chloride and boron-containing compounds to enhance the oxidation resistance of the anode. Anti-oxidation coating technology involves using electrolyte coatings, carbon-based coatings, ceramic coatings, or aluminum coatings. Among these, coating protection is an important research direction for improving the oxidation resistance of carbon anodes. The cost, oxidation resistance effect, and adhesion to the carbon substrate are key research focuses for anti-oxidation coatings.

Abstract:Electroplating sludge, a hazardous waste generated during the treatment of electroplating wastewater, poses significant environmental hazards but is rich in heavy metal elements such as copper, nickel, chromium, iron, and zinc, making it a valuable secondary resource. Currently, many scholars have conducted research on the harmless disposal and resource utilization of electroplating sludge. Harmless disposal methods primarily include solidification/stabilization (S/S) and thermal treatment technologies, while resource utilization mainly encompasses hydrometallurgical leaching, pyrometallurgy-hydrometallurgy combined processes, pyrometallurgy, and materialization. Solidification/stabilization technology is significantly effective in solidifying heavy metals in electroplating sludge. However, the addition of solidifying agents increases the total mass and volume of the sludge, thereby elevating landfill burden and costs. This method is suitable for electroplating sludge containing various stable pollutants. Thermal treatment technology can improve metal fixation efficiency and achieve volume reduction of electroplating sludge, but the treatment process causes environmental pollution and increases the porosity of sintered products, limiting their reuse. Additionally, it leads to the waste of valuable metal resources. Hydrometallurgical recovery of valuable metals from electroplating sludge includes acid leaching, ammonia leaching, bioleaching, etc. Acid and ammonia leaching can recover multiple valuable metals in stages, but acid leaching faces challenges in pH regulation for separation and adaptability issues with ultrasonic-assisted processes. Ammonia leaching requires careful control of ammonia volatilization, while bioleaching is constrained by microbial strains, resulting in unstable leaching efficiency and cycles. The pyrometallurgy-hydrometallurgy combined process converts metals such as copper, nickel, and zinc in electroplating sludge into corresponding chlorides that volatilize during roasting, and transforms chromium into water-soluble phases for recovery during leaching. However, this method is limited by corrosion issues with roasting equipment, restricting its industrial application. Moreover, the high roasting temperature leads to elevated energy consumption. Pyrometallurgical processes utilize high temperatures to reduce metal oxides in sludge and recover them in the form of multi-element alloys. Different reducing agents can be selected based on the composition of the sludge. The key to this method lies in selecting appropriate additives to form a low-melting slag system. Materialization technology primarily uses hydrometallurgical or pyrometallurgical techniques to targetedly convert metal elements in electroplating sludge into functional materials such as adsorbents, catalysts, and pigments. This method features a simple process, easy operation, and is more environmentally friendly and cost-effective for resource utilization, showing broad application prospects. However, due to the complex composition of electroplating sludge, in-depth research is required to investigate the migration patterns of other heavy metals, chlorine, organic pollutants, sulfur, and other potential hazardous substances beyond target elements, as well as their influence on the performance of target materials. In practical resource utilization processes, the selection of technological methods should be based on the composition of the sludge. For single-component electroplating sludge, hydrometallurgical and materialization technologies are prioritized to reduce process energy consumption and equipment investment costs. For complex multi-metallic electroplating sludge, pyrometallurgical recovery technologies or combined pyrometallurgy-hydrometallurgy processes are preferred to achieve resource utilization while avoiding environmental risks associated with multi-metal separation in single hydrometallurgical processes.

Abstract:Cerium oxide polishing powder can form a chemical mechanical polishing (CMP) function, which meets the needs of the precision polishing field. It is characterized by a fast polishing speed, high finish, high flatness, and a long service life for polished products. In this field, the main international research directions focus on the application optimization and expansion of cerium oxide polishing powder, the influence of the size and morphology of cerium oxide particles on polishing performance, the recovery effect of the introduction of additives and cleaning methods on surface defects, and the particle size control process of polishing powder, among others. Conversely, in China, there are still shortcomings in particle size control, optimizing the production process, shortening the production cycle, and improving equipment utilization. The preparation methods of cerium oxide polishing powder are mainly solid phase, liquid phase, and gas phase. The solid phase method includes solid phase sintering and solid phase mechanical methods, which typically have higher energy consumption and longer reaction times. The liquid phase method includes hydrothermal, sol-gel, precipitation, microwave, and microemulsion methods, among others. These methods are conducted at lower temperatures and can produce polishing powder with a more uniform particle size. Nevertheless, some liquid phase methods are complex, and the treatment of solvents and by-products also needs optimization. The gas phase method mainly includes spray drying and spray pyrolysis, which can produce polishing powder with specific shapes and sizes by atomizing solutions or suspensions into fine droplets and then rapidly evaporating the solvent to form solid particles at high temperatures. With the enhancement of global environmental awareness, the development of low energy consumption and green synthesis methods has become a trend, and gas phase and other environmentally friendly synthesis methods will receive more attention. Future research will pay more attention to the particle size and morphology control of cerium oxide powder to meet the polishing needs of different fields.

Abstract:The separation and purification of vanadium from vanadium-containing solutions is the key to preparing high-purity vanadium. Conventional extraction and purification require a large amount of reagents, generate a large amount of wastewater, have high requirements for the acid resistance of equipment, and have poor ion selectivity. The pH value of Cyanex272 extraction is mainly between 2.5 and 4.0, which has a relatively low requirement for the acid resistance of the equipment. Moreover, the wastewater generated during extraction causes less harm to the environment. In this paper, the vanadium solution after calcification roasting and acid leaching of vanadium slag from a certain enterprise was used as the raw material. Cyanex272 was used as the extractant and sulfuric acid as the stripping agent to purify vanadium and prepare V₂O₅ The extraction mechanism of Cyanex272 was explored through characterization analysis, and the following main conclusions were obtained. Cyanex272 exhibits excellent selectivity for V(Ⅳ) under acidic conditions. Under the conditions of organic phase composition of Cyanex272∶TBP∶sulfonated kerosene=4∶0.5∶5.5, solution pH=3.0, extraction time 10min, extraction temperature 30℃, and O/A ratio of 1∶1, The vanadium extraction rate of single-stage extraction was 78.06%, and that of three-stage extraction reached 98.80%. Under the conditions of A sulfuric acid concentration of 3.0mol/L as the stripping agent, a stripping time of 6min, a stripping temperature of 30℃, and a stripping ratio of O/A=1∶1, the vanadium stripping rate of single-stage stripping was 80.70%, that of three-stage stripping reached 98.38%, and the purity of V₂O₅ after calcination was 99.36%. The vanadium reduced by sodium sulfite in the vanadium solution mainly exists in the form of VO²⁺. The phosphate group of Cyanex272 undergoes chelation reaction with VO²⁺ by breaking the P—O—H bond to form a stable metal complex, thereby achieving the separation and extraction of vanadium ions.

Abstract:In this study, a combined iron capture and smelting-magnetic separation process was developed for the efficient recovery of platinum group metals (PGMs) from three-way catalysts (SACs) for waste vehicles. The optimized process parameters were determined by kilogram-level intermediate frequency furnace test: calcium oxide dosage of 40%, iron powder 14%, melting temperature of 1480℃, reaction time of 45min, the content of PGMs in the slag decreased to ＜30g/t, and the capture recovery rate reached 98.55%. For the remaining Fe₃O₄/Fe₃Si encapsulated PGMs particles in the slag, wet ball milling (particle size -0.1mm≥85%) combined with 0.14T magnetic field strength magnetic separation was used to obtain a concentrate yield of 1.3%, a PGMs content of ＜0.6g/t and a magnetic separation recovery rate of 97.7%. The total recovery rate of PGMs in the whole process exceeded 99.9%, and the magnetic enrichment of Fe₃O₄ and Fe₃Si particles was the core mechanism to achieve the deep recovery of PGMs. This process significantly improves the recovery efficiency of dispersed PGMs by iron capture and smelting technology, and provides a reliable technical solution for the large-scale regeneration of SAC secondary resources.

Abstract:After the treatment of lepidolite concentrate by sulfate roasting-water leaching method, the water leaching solution contains a large amount of impurity elements such as calcium, magnesium, potassium and sodium. This study focuses on the leachate from a lithium mica smelting enterprise in Hunan Province, employing a chemical precipitation method for impurity removal. It examines two approaches: directly removing impurities with sodium carbonate and adding sodium carbonate after adjusting the pH with calcium oxide or sodium hydroxide. The effects of calcium oxide dosage, sodium carbonate dosage, reaction time, reaction temperature, and the method of adding removal agents on the impurity removal efficiency were investigated, leading to the following main conclusions. When impurities were removed directly using sodium carbonate at twice of the theoretical dosage, the calcium ion content was reduced to 21.4mg/L, the magnesium ion content was reduced to 43.3mg/L, and the lithium loss rate was 6.6%. When removing impurities by adding sodium carbonate after adjusting the pH, the pH of the leachate was primarily adjusted to 11.63, followed by continuous stirring at room temperature for 1 hour before solid-liquid separation. Then, 20% sodium carbonate solution with dosage of 1.5 times of theoretical amount was added to the pH-adjusted liquid, and after stirring at 90℃ for 1 hour and performing solid-liquid separation, the calcium ion content was reduced to 12.8mg/L, the magnesium ion concentration fell below 0.1mg/L, and the lithium loss rate was 2.1%. The stepwise filtration method proposed in this paper, involving adjusting the pH before using sodium carbonate for impurity removal from lithium mica leachate, demonstrates advantages of effective impurity removal, low lithium loss rate, and low cost, providing a reference for the optimization of lithium extraction processes in similar enterprises.

Abstract:Spent lithium iron phosphate batteries form phosphorus iron slag after selective lithium extraction. When recovering phosphorus iron from this slag, there are generally problems such as difficulty in removing aluminum and high cost of aluminum removal. This article uses a mixture of sulfuric acid and hydrated iron sulfate to pre remove aluminum from phosphorus iron slag, investigates the influence of various parameters on the aluminum removal effect, analyzes the aluminum removal mechanism, and provides a comprehensive process flow that can reduce phosphorus iron loss. The pre aluminum removal method used in the experiment is technically feasible and effective. Under the conditions of a sulfuric acid concentration of 0.3mol/L, a mass ratio of hydrated iron sulfate to phosphorus iron slag of 20%, a reaction temperature of 95℃, and a reaction time of 1 hour, the one-time aluminum removal rate can reach 91.90%; The method of using sodium hydroxide to adjust the pH value of the liquid after aluminum removal can precipitate crude iron phosphate (with high aluminum content, which is then returned to the pre aluminum removal process). This comprehensive process does not lose much iron phosphate during aluminum removal, with a one-time aluminum removal efficiency of about 84%; The mixed agent can remove impurities such as Cu and Li from the phosphorus iron slag while removing aluminum. The content of Cu and Li in the phosphorus iron slag decreased from 0.095% and 0.056% to 0.0089% and 0.0012%, respectively; After aluminum removal, some of the iron hydroxide in the phosphorus iron slag is encapsulated by crystalline iron phosphate, making it difficult to effectively adjust the iron phosphorus ratio through aging. It needs to be dissolved in acid before synthesis. The comprehensive process of aluminum removal in this experiment has low cost, good effect, simple operation, and can effectively avoid the loss of phosphorus and iron. The final impurity removal liquid (wastewater) obtained can also be reused after simple treatment, with minimal environmental impact and obvious industrial advantages.

Abstract:Aiming at the problems of particle size distribution control and anti-oxidation in the preparation of ultrafine nickel powder, this paper proposes a method for preparing ultrafine nickel powder by hydrazine hydrate-potassium borohydride reduction-hydrogen reduction system using nickel sulfate as raw material. The performance of nickel powder prepared under optimized conditions was analyzed by SEM, XRD and other characterization methods, and the following main conclusions were obtained. The optimized process conditions of hydrazine hydrate are as follows: reduction temperature 333.15K, reduction time 150min, solution pH 8.0~8.5, nickel ion molar concentration 1.25mol/L, mass ratio of hydrazine hydrate to nickel ion 1.35, mass concentration of sodium hexametaphosphate 1.5‰. Under these conditions, the nickel reduction rate can be stabilized at about 99%. The introduction of proper amount of potassium borohydride in the reduction and precipitation process of nickel powder can overcome the induction period of hydrazine hydrate reduction and improve the controllability of nickel powder particle size. Under the experimental conditions, the addition of 8.37×10^-3g/L potassium borohydride can make the particle size of nickel powder less than 1μm. The reduction of nickel powder by hydrogen can reduce the oxygen content of the product and control the length of the powder particles. Under the condition of reduction temperature of 673.15K and reduction time of 120min, the oxygen content of nickel powder is 0.56%, and the particle size is 0.51μm. The nickel powder prepared under the optimized process conditions has a purity of higher than 99.5%, a particle size of 0.5~1μm, a specific surface area of 3.6~4.24m²/g, and a loose specific gravity of 0.5~0.6g/cm³. This index has reached the level of similar foreign products due to similar domestic products. The innovation of this research lies in the introduction of potassium borohydride to control the particle size range during the preparation of nickel powder and the use of hydrogen reduction for anti-oxidation treatment. This method has short process flow, good working environment, no pollution, and has promotion value.

Abstract:In view of the problems of poor adaptability of raw materials and incomplete separation of impurities in the current preparation methods of high-purity selenium, such as oxidation volatilization method, regional melting method, vacuum distillation method, etc., this paper uses crude selenium dioxide ( purity 98%) as raw material. The 4N high-purity selenium was prepared by ion adsorption impurity removal-formic acid reduction process, and the optimization parameters of the process were investigated by single factor test. The following main conclusions were obtained. In the resin impurity removal stage, the removal rates of tellurium and arsenic were mainly investigated (other impurities had little effect on the preparation of high-purity selenium). When the solution flow rate was controlled at 2 BV, the removal rates of tellurium and arsenic were more than 99% and 96%, respectively. The optimum process parameters for the reduction of selenite by formic acid were as follows: reaction temperature 80℃, excess multiple of formic acid 1.1 times, stirring speed 200r/min and reaction time 4h. Under these conditions, the reduction rate of selenium was more than 99%. The 4N high purity selenium prepared by this method meets the requirements of standard YS/T 1354—2020 ‘selenium powder’. The results of this study can provide data and technical reference for related enterprises to prepare high-purity selenium.

Abstract:The establishment of a carbon dioxide emission accounting method in the industrial industry is of important guiding significance for dealing with climate change. Based on the standardised development of carbon emission accounting in the recycled lead industry, the carbon emission accounting method of the recycled lead industry has been sorted out and determined. With the clean recycling technology of waste lead batteries (automatic crushing and sorting-lead paste pre-desulphurisation-low-temperature melting Refining) is represented, and carbon emissions are accounted for the whole process with each recycling section as the accounting unit. We are committed to filling the gap in the carbon emission evaluation standards in the field of recycled lead, which is conducive to guiding and standardising the production and development of the recycled lead industry. The accounting results show that the processing capacity of waste lead batteries is 150000t/a of mechanical crushing and sorting-lead paste pre-desulphurisation-low-temperature smelting of recycled lead clean production process, waste lead batteries in the crushing and sorting stage, lead grid melting stage, crude lead electrolytic refining stage, lead paste pre-desulphurisation stage, desulphurisation lead paste low-temperature melting stage carbon The emissions are 0.380t CO₂ h, 0.460t CO₂ h, 0.860t CO₂ h, 1.900t CO₂ h, 5.047t CO₂ h respectively. These carbon dioxide emission indicators can also be used as the same industry and similar technology enterprises. The carbon dioxide emission accounting coefficient of the industry. Based on the results of carbon emission accounting, the characteristics of carbon emissions in the recycled lead industry are analysed. Direct emissions caused by fuel combustion are the largest source of carbon emissions in the recycled lead cleaning process, accounting for 31.57% of the total carbon emissions. The low-temperature smelting section is the largest carbon emission section of the recycled lead cleaning process, and its carbon emissions account for 58.36% of the total carbon emissions.

Abstract:Secondary aluminum ash contains a large amount of Al, Al₂O₃, as well as a certain amount of MgAl₂O₄ and AlN. Direct storage would cause environmental pollution and waste of resources. However, Al, AlN and some impurities in aluminum ash can be dissolved under alkaline leaching conditions, and the remaining residue composition is relatively close to that of magnesium-aluminum spinel material (MA). In this study, secondary aluminum ash was used as the main raw material, and the method of alkali leaching pretreatment combined with pyrometallurgical sintering was adopted. MgO was used as the modifier to prepare MA materials. The research process investigated the influence of alkali leaching process parameters on the leaching rates of Al and AlN, analyzed the alkali leaching mechanism of secondary aluminum ash, and investigated the effect of sintering process parameters on the properties of the prepared magnesium-aluminum spinel materials. The following main conclusions are obtained. The alkali leaching test shows that increasing the concentration of NaOH, raising the leaching temperature, increasing the liquid-solid ratio, and accelerating the stirring rate are all conducive to improving the leaching rates of Al and AlN. Under the optimized alkali leaching conditions of a temperature of 80℃, a NaOH concentration of 1.6mol·L^-1, a stirring rate of 300r·min^-1, and a liquid-solid ratio of 12∶1, the leaching rate of Al reached 80.35% and that of AlN reached 53.21%. The leaching reaction of Al and AlN in secondary aluminum dross is essentially a hydrolysis reaction, and the alkaline system promotes the progress of this hydrolysis reaction. Spinel materials with complete crystal form and high crystallinity can be prepared by using the alkali leaching residue as raw material and the MA material ratio of MgO under the sintering conditions of 1300-1400℃ and sintering time of 3hours. The apparent porosity of the MA material prepared at 1400℃ is 31.56%, and its compressive strength is 50.22MPa, which is higher than the national industry standard “Magnesia Bricks and Magnesia-alumina Bricks” (GB/ T2275—2007) (40MPa).