Alkaline Leaching Process Research Articles

A new iron recovery and dephosphorization approach from unroasted high‐phosphorus oolitic hematite ore via a facile chemical beneficiation process

AbstractDespite the abundance of scientific reports on the treatment of oolitic hematite, it still constitutes a global challenge. The distinctive oolitic structure within the ore is regarded as the primary factor causing the difficulty in the beneficiation of oolitic hematite ore. The objective of this study is to devise an effective approach to tackle the issue of beneficiation of oolitic hematite ore, thereby facilitating the utilization of oolitic hematite ore, which is widely abundant worldwide. Based on the Pourbaix diagram (Eh‐pH) of the Al‐Si‐H2O system, an alkaline leaching process for eliminating impurities from unroasted oolitic hematite using NaOH solution as the leaching agent was proposed. The optimal parameters were determined as a temperature of 250°C, a NaOH concentration of 16 mol/L, a liquid‐to‐solid ratio of 4 mL/g, and a reaction time of approximately 2 h. Around 54% of phosphorus (P), 21% of Al2O3, and 61% of SiO2 can be preliminarily removed from unroasted oolitic hematite ore. Through further dephosphorization with dilute sulfuric acid at room temperature, a high‐quality Fe concentrate with 63.3% Fe, 0.04% P, and a Fe recovery rate of 96% was obtained. The alkali can be readily regenerated by adding lime to the leaching solution, and the regenerated alkali has a leaching effect comparable to that of the fresh leaching solution. The recycling and utilization of alkaline leaching agents minimize production costs to the greatest extent possible. Mechanical studies have discovered that intact oolites undergo complete dissociation under alkaline leaching, enabling the leaching solution to permeate into the particles and thereby deeply eliminate impurities. The novel method has accomplished efficient processing of oolitic hematite and a reasonable control of production costs during processing, which is significant for expanding the reserves of iron ore resources.

Mineralogical examination of biomass ashes: A preliminary study for potassium enrichment and phosphoric acid (H3PO4) production

In this study, biomass fly ash was investigated and determined by XRD to be predominantly langbeinite, aphthitalite, sylvine, apatite and hydroxyapatite. An alkaline leaching study was carried out on the material in a glass reactor at l/s ratio 4 temperature 600C at pH 12.85 and after drying was analyzed by ICP. The crystal structure composition of K2SO4 72.28% and potassium content of 33.47%(w/w) was recovered. The elemental transformations in this leaching process were: K 74.16%, Mg 0.55% was found to be 0.8%. It was found that a minimum of 4.84 g H2SO4 per 100 g fly ash was required to precipitate the Ca(OH)2 migrated into the liquid solution and the overall leaching efficiency was found to be 29.85%. In the alkaline leaching process, potassium was found to be easily recovered from each biomass ash. By adding 40 g H2SO4/100 g biomass waste remaining after leaching, a phosphorus recovery of 74.9% was observed. In this preliminary study, unsatisfactory results were obtained in obtaining the industry standard H3PO4. The reason for this is that the Al, Mg and K in the fly ash are subject to selective dissolution as a result of working at high pH in the alkaline leaching process.

Innovating the recycling of silicon-based solar panels with an eco-friendly alkaline leaching process

As the accumulation of waste from solar panels and solar scraps rapidly increases, recycling these materials has become increasingly critical. Unfortunately, the complexity of disassembling solar panels often renders the recycling process unprofitable due to the low value of the recovered products, resulting in high service fees. Although numerous studies have sought to simplify the disassembly process and enhance the purity of the recovered products, these solutions often involve costly, toxic, and complex methods. In response, we introduce an innovative green recycling technique using a straightforward alkaline leaching process. This method allows for efficient recycling and refining of solar cells, yielding high-purity SiO2 (99.994 %) at a recycling rate of 92.74 % and generating green hydrogen. A preliminary cost analysis indicates that this process holds substantial commercial potential and high profitability. The proposed technique offers a new pathway towards making renewable energy sources more sustainable and economically feasible.

Microwave-assisted atmospheric alkaline leaching process and leaching kinetics of rare earth melt electrolysis slag

This study focuses on the difficulty of converting fluorinated rare earth elements into hydroxylated rare earth elements in rare earth melt electrolysis slag (RMES) and proposes the use of a microwave-assisted atmospheric alkaline leaching method for the treatment of RMES. The leaching behavior of RMES under microwave-assisted atmospheric alkaline leaching was studied, and the optimal reaction conditions were determined. Under the conditions of a reaction temperature of 150 °C, initial NaOH concentration of 60 %, NaOH-to-slag mass ratio of 4:1, microwave power of 700 W, reaction time of 120 min, and stirring speed of 300 r/min, the conversion rate of fluorinated rare earths reached 99.17 %. The apparent rate equation of the microwave-assisted atmospheric alkaline leaching process was obtained by leaching kinetic analysis, and the apparent activation energy under this process was calculated to be 54.872 kJ/mol, which was 12.458 kJ/mol lower than that achieved when conventional heating was used for leaching (67.33 kJ/mol).

Study on the treatment of pyrolysis products from the nitric acid pressure leach liquor of laterite

As the demand for nickel increases sharply, it is of great significance to extract valuable metals such as nickel and cobalt from lateritic nickel ore with huge reserves. Based on the nitric acid pressure leaching process (NAPL) of laterite nickel ore, a clean process of laterite nickel ore treatment is proposed. The process includes concentration and crystallization of pressure leaching solution of nitric acid, pyrolysis of mixed nitrate crystal, alkaline leaching of mixed oxides and acid leaching of alkaline leaching residue. Based on previous work, the high-pressure alkaline leaching process of pyrolysis products (mixed oxides) and atmospheric pressure sulfuric acid leaching process of alkaline leaching residue were studied in this paper. First, the Eh-pH diagrams of metal-water system of common elements in laterite nickel ore was summarized according to the reaction equations of different substances in aqueous solution, and the subsequent wet leaching process was guided. The following alkaline leaching experiments showed that under the conditions of 300 g/L NaOH concentration, 3.5 mL/g liquid–solid ratio, 160 °C leaching temperature and 1 h leaching temperature, 85.0 % of aluminum in mixed oxides can be leached and scandium can be enriched by 13 times. Finally, under the conditions of initial sulfuric acid concentration of 4 M, temperature of 100 °C, leaching time of 30 min and liquid–solid ratio of 4 mL/g, the leaching rates of Al, Ni, Mg, Co, Mn, Fe and Sc in alkaline leaching residue were 92.0 %, 97.5 %, 98.4 %, 92.3 %, 62.7 % and 68, respectively. A new procedure for treating lateritic nickel ore is proposed, allowing for thorough recovery and usage of important components. The technique is green and environmentally benign, resulting in closed-loop production of laterite-nickel ore treatment with numerous application possibilities.

Process simulation and life cycle assessment of hydrometallurgical recycling routes of waste printed circuit boards

Electronic waste remains a pressing global environmental concern. Addressing the proper treatment and recycling of waste printed circuit boards (WPCBs) is imperative to mitigate environmental pollution and prevent the loss of valuable resources. In this study, a comparative analysis of acidic and alkaline leaching processes for recovering base metals (Cu, Sn, Zn, and Ni) from WPCBs was conducted by process simulation (using HSC software). Life cycle inventories were compiled accordingly. Furthermore, by incorporating an alternative method for precious metal recovery, the life cycles (using openLCA software) of both processes for recovering all metals were examined. The results revealed that while the acidic leaching process showed higher recoveries for base metals and advantages in terms of lower electricity consumption and wastewater generation, it also exhibited a more pronounced environmental impact due to the chemical input in the HNO3 leaching process. Conversely, the alkaline leaching process utilized fewer chemicals, generated less solid waste and off-gases, resulting in a lower overall environmental impact. For recycling 1 t of WPCBs, the acidic leaching process achieved recoveries of 92% for tin, 99% for copper, 79% for zinc, and 94% for nickel, whereas the alkaline leaching process yielded recoveries of 96%, 93%, 78%, and 74% for the corresponding metals. Notably, the alkaline leaching process demonstrated a lower global warming potential of 11,395 kg CO2 eq, compared to the 12,459 kg CO2 eq associated with the acidic leaching process. In the evaluation of the alkaline leaching process, most environmental impacts were found to be lower than for virgin production. As a result, while the acidic leaching process is indeed more efficient in metal recycling, the alkaline leaching process emerges as a stronger contender in terms of environmental sustainability, with great potential for further technical improvements.

Innovative Recovery of Vanadium from Vanadium–Chromium-Reducing Residue by Ultrasonic-Assisted Alkaline Leaching

ABSTRACT Vanadium–chromium-reducing residue (VCRR) was generated using a sodium-roasting-leaching-precipitation process for vanadium recovery. The retained vanadium and chromium could be utilized for an important sources. Here, a novel process was developed that utilized ultrasonic to enhance the alkaline leaching process of vanadium from VCRR. The effect of various parameters was studied, and a corresponding kinetic model was established. It was found that the maximum vanadium leaching efficiency of 96.9% was obtained at optimum reaction conditions: reaction temperature of 90°C, reaction time of 60 min, dosage of NaOH at m(NaOH)/m(VCRR) = 0.5 g/g, liquid-to-solid ratio of 5 mL/g, stirring rate of 500 rpm, and ultrasonic frequency of 40 kHz. The leaching kinetics analysis showed that the leaching process was controlled by diffusion through a product layer both in normal leaching and ultrasonic leaching. The calculated Ea for vanadium leaching out was 20.56 kJ · mol−1 for ultrasonic leaching and 23.06 kJ · mol−1 for normal leaching. Ultrasonic-assisted leaching might be a potential industrial application technique for recovery of valuable metals from metal ores.

Uranium recovery from phosphate rocks/minerals – A comprehensive review

Although the use of nuclear energy is still viewed with caution due to various problems associated, such as the safe operation of 'aging' nuclear plants and the final disposal of nuclear waste, many countries in Europe are now increasingly considering nuclear energy as an alternative to reducing fossil fuel energy production. At the present stage primary sources of uranium are sufficient to meet uranium needs and secondary sources are becoming increasingly important in terms of their long-term availability and environmental friendliness. Mining/recovery of uranium from secondary sources (such as phosphate minerals) is considerably safer in terms of exposure to radiological radiation and therefore requires less stringent safety measures. Uranium is also present in waste produced from certain industrial processes such as fertilizer production industry (e.g. phoshogypsum) and in nuclear waste. This work refers to the extraction and recovery of uranium from phosphate rocks, which is directly related to the production of uranium-free phosphate fertilizers and aims to limit environmental pollution related to the use of phosphate fertilizers. Since the chemistry during the leaching step is critical to efficient uranium recovery, various investigations have been carried to study various parameters that play an important role in achieving an optimum uranium extraction from rocks. Uranium leaching is carried out under acidic or alkaline conditions and the choice of chemistry is determined by the mineralogy of the host rocks. Acid leaching yields a higher recovery than alkaline leaching because the matrix is often more resistant to alkali than to acid leaching. However, the alkaline leaching process has the advantage of leading to the precipitation of a comparatively purer product directly from the leaching liquid. In addition, the washing solutions can be regenerated for further use, making this method more environmentally friendly. Recently alternative methods have been proposed which make use of uranium complexing ligands (e.g. EDTA) to achieve more efficient and selective uranium recovery.

Influence of the structural differences between end-of-life Al-BSF and PERC modules on the Al leaching separation behavior

The booming global photovoltaic industry is bringing a large number of end-of-life photovoltaic (PV) modules. Achieving effective and green recovery of end-of-life PV modules is crucial to perform the energy and environmental sustainable development. The passivated emitter and rear cell（PERC) modules are the mainstream products in the current solar cell market, while the aluminum back surface field cell（BSF)modules were the mainstream products in the previous solar market. These two types of modules currently occupy the main share of the recycling market. There are several studies on the separation and recovery of valuable metals in the end-of-life PV modules. However, the influence of different cell processes and structures on the separation behavior of mainstream impurities in cells remains unclear. In this paper, the impacts of the structure of PERC and Al-BSF cells on the separation behavior of dominant Al impurity were studied. The structure and component difference between the PERC cells and Al-BSF cells was first clarified. The relationship between the loss rate of silicon wafer and the leaching rate of the aluminum back impurity was then studied. The results showed that the local contacts between the Al layer and Si wafers were primarily formed in PERC cells, in which the higher temperatures and longer leaching times were required to achieve the same Al leaching efficiency of Al-BSF cells. As for the Al-BSF modules, in order to better balance the Al separation and silicon loss rate during the alkaline leaching process, the reaction time should be set in the range of 7–10 min. The leaching rate of aluminum can reach 97.5% at 25 °C and 1 M NaOH for 10 min. After leaching, a layer of residual Si-O-Al alloy phase which is attached to the back surface of the two kinds of cell was observed. This study has a guiding significance for the green and efficient recovery of the silicon wafer in photovoltaic (PV) modules.

Strengthening desilication of coal fly ash by alkaline leaching with the addition of ethylene diamine tetraacetic acid

Pre-desilication, as a treatment to improve the mass ratio of alumina to silica of aluminosilicate materials, is one of the essential procedures for alumina extraction from coal fly ash (CFA). In this work, a new method for strengthening desilication efficiency of CFA was proposed by alkaline leaching with the addition of ethylene diamine tetraacetic acid (EDTA). The effects of alkaline leaching conditions on the desilication efficiency of CFA were investigated; meanwhile, the strengthening pre-desilication mechanism was further clarified by means of scanning electron microscopy-energy dispersive spectroscopy, laser particle size analysis, x-ray diffraction and nuclear magnetic resonance, etc. The results showed that the desilication efficiency of CFA was significantly increased from 38 % to 50 % with the addition of EDTA during the alkaline leaching process. The mass ratio of alumina to silica of desilicated CFA was increased from 1.41 to 1.66, accordingly. During this process, the dissolved Al(OH)4- in the desilicated solution would be complexed by the additional EDTA, which makes it not to largely consume the dissolved H2SiO42- to form an insoluble by-product of hydroxysodalite (Na8Al6Si6O24(OH)2(H2O)2). Both the reduction of cumulative amount of hydroxysodalite on the desilicated CFA and the continuous exposure of new reaction interface on the desilicated CFA contribute to the improvement of desilication efficiency.

Kinetics of alkaline leaching process of titanium (IV) from ilmenite

Kinetics of alkaline leaching process of titanium (IV) from ilmenite

Statistically optimized sequential hydrothermal route for FeTiO3 surface modification: evaluation of hazardous cationic dyes adsorptive removal.

In the present study, the performance of facile hybrid sequential chemical treatments of titanomagnetite concentrate (TC), alkaline leaching, and sodium dodecyl sulfate (SDS) modification has been evaluated for the removal of crystal violet (CV), malachite green (MG), and methylene blue (MB) cationic dyes. The physical and chemical properties of samples were characterized using scanning electron microscopy (SEM), energy dispersive X-ray (EDX), N2 adsorption-desorption, X-ray diffraction (XRD), X-ray fluorescence spectroscopy (XRF), and Fourier transform infrared spectroscopy (FTIR) analyses. Moreover, dye removal in the batch system was investigated by evaluating adsorbent dosage, contact time, initial dye concentration, pH of the solution, temperature, electrolyte concentration, adsorption isotherm, kinetic, and thermodynamic. The results showed that the maximum adsorption capacity was obtained at SDS concentration of 6mM, NaOH concentration of 9M, the temperature of 160°C, solid/liquid ratio of 4g/100mL, and the process duration of 24h. In the alkaline leaching process, forming the Na2TiO3 phase with sharp and high energy points can be improved the adsorption properties. Accordingly, the adsorption capacity and removal efficiency attained 19.84, 18.64, and 19.66mg/g and 99.21, 93.24, and 98.31% for CV, MG, and MB, respectively. Furthermore, the dye removal followed pseudo-second-order (R2 = 0.9990) and Freundlich (R2 = 0.9970) models. The evaluation of thermodynamic parameters indicated the endothermic (∆H° = 110.91J/mol) and spontaneous nature (ΔG˚ < 0) of the adsorption process. This study concluded that the modified TC had a potential ability for application in textile wastewater treatment.

Fundamental research on selective arsenic removal from high-salinity alkaline wastewater

The alkaline leaching process of arsenic-containing solid waste discharged during nonferrous metal smelting affords typical high-salinity alkaline arsenic-containing wastewater (HSAW). In this study, for the first time, Me (Ca2+ and Mg2+)–AsO43−–OH−–H2O and Me (Ca2+ and Mg2+)–AsO43−–CO32−–H2O systems are studied based on a thermodynamic equilibrium diagram and an arsenic removal experiment, proving that the removal of arsenic using single metal ions in the presence of CO32− is infeasible because of carbonate coprecipitation. Based on this observation, a new method that uses magnesium ammonium complex salts (MACSs) for HSAW treatment is proposed. Based on the thermodynamic calculations of the Mg2+–AsO43−–NH4+–CO32−–H2O system and the arsenic removal experiment, carbonate and arsenate can be selectively separated by the formation of magnesium ammonium arsenate (NH4MgAsO4·6H2O). In an arsenic solution containing 150-g/L Na2CO3, the arsenic removal rate and the arsenic grade of the precipitation product reach 90.16% and 27.13%, respectively, when the molar ratios of Mg2+/NH4+:As(V) are 1.8:1 and 2:1, respectively. The proposed method is successfully employed for treating a leaching solution of alkaline arsenic slag discharged during antimony smelting. The findings of this study will broaden the basic theory of HSAW treatment and lay a foundation for the resource treatment of arsenic-containing solid waste.

Zinc recovery from hazardous construction wastes via an integrated process of the air classification, alkaline leaching, and water washing

AbstractZinc‐ and lead‐containing wastes are often mixed with construction and demolition wastes in many factories, generating abundant of heavy metal‐enriched hazardous waste. In the present study, a novel integrated process of air classification, alkaline leaching, and water washing dechlorination was proposed for the efficient recycling of Zinc (Zn) resources. The first air classification process was realized via venturi tube, wherein the content of Zn could increase by 20 wt.%. After that, the product underwent an alkaline leaching process. Results showed that Zn recovery rate increased with fine particle sizes, and a 65% recovery rate was obtained under the following conditions of 5 mol/L NaOH, liquid/solid 10:1, and leaching time 1 h. Finally, water washing associated with microwave and ultrasonic treatments could remove over 85% of Cl and other water‐soluble salts. All the results indicated that the integrated method had an excellent recovery rate for Zn resources from construction and demolition wastes.

A New Route to Upgrading the High-Phosphorus Oolitic Hematite Ore by Sodium Magnetization Roasting-Magnetic Separation-Acid and Alkaline Leaching Process

In this paper, an innovative method is proposed to upgrade iron and remove phosphorus from high-phosphorus oolitic hematite ore by the sodium magnetization roasting–magnetic separation–sulfuric acid and sodium hydroxide leaching process. The process parameters of sodium magnetization roasting, acid leaching, and alkaline leaching were optimized. The results show that by only adopting traditional magnetization roasting–magnetic separation, an iron ore concentrate containing 57.49% Fe and 1.4% P2O5 at an iron recovery rate of 87.5% and a dephosphorization rate of 34.27% was produced, indicating that it is difficult to effectively dephosphorize and upgrade iron by the conventional magnetization roasting–magnetic separation process. The obtained rough magnetic concentrates were then subjected to acid and alkaline leaching steps, and the final product, assayed at 64.11% iron and 0.097% P2O5, was manufactured successfully. Moreover, the added NaOH could promote the mineral phase reconstruction of aluminum- and silica-bearing minerals during magnetization roasting and intensify the upgrading of iron as well as enhance the growth of iron grains.

Separation of alumina and silica from metakaolinite by reduction roasting−alkaline leaching process: Effect of CaSO4 and CaO

Limestone (CaCO3), which could promote sulfur fixation, was added to coal gangue during roasting in a circulating fluidized bed (CFB) boiler. CaO and CaSO4 were the main Ca-bearing minerals while metakaolinite was the major Al-bearing mineral in CFB slag. The effect of CaSO4 and CaO on the separation of alumina and silica from metakaolinite by reduction roasting−alkaline leaching process was studied. Results showed that metakaolinite was completely converted into hercynite and silica solid solutions (i.e., quartz and cristobalite solid solutions) by reduction roasting with hematite. More than 95% of silica in the reduced specimen was removed by alkaline leaching. The addition of CaSO4 and CaO remarkably decreased the separation efficiency of alumina and silica in metakaolinite, which could be attributed to the formation of Si-bearing minerals: (1) Fayalite and anorthite were formed during the reduction roasting process; (2) Fayalite was stable while anorthite was converted into sodalite and wollastonite during the alkaline leaching process. This study demonstrates that sulfur in coal gangue should be fixed by treating the exhaust gas instead of controlling the combustion process of CFB to achieve the comprehensive recovery of silica and alumina from the CFB slag.

Recovery of tungsten trioxide from waste diamond core drilling crowns by nitric acid leaching

A cost-effective and simple hydrometallurgy procedure for recycling the diamond core drilling crowns by nitric acid leaching, followed by the sodium hydroxide leaching, has been developed. The effect of temperature, nitric acid/sodium hydroxide concentration, leaching time and stirring rate were studied. For the acid leaching process the optimal values for parameters were 60 °C, 1.0 M HNO3, 2.5 h, and 800 rpm. It has been noted that the key factors in the alkaline leaching process were temperature and NaOH concentration. For the alkaline leaching, the optimal parameters were 100 °C, 40% concentration of NaOH solution, and leaching time of 4 h. Identification and characterisation of products in all technological processes were performed by X-Ray Diffraction (XRD), scanning electron microscope using energy-dispersive X-ray spectroscopy (SEM-EDS), and laser desorption ionisation mass spectrometry (LDI-MS). The analysis of solutions in all phases was performed by inductively coupled plasma atomic emission spectroscopy (ICP-AES). The use of LDI-MS as a new approach to identify the products after phases in the recycling process was an important part of the study, and the results were consistent with the XRD analysis. The total tungsten recovery was over 95.6%, with a high purity of the obtained tungsten trioxide.

Selective Pre-leaching of Tellurium From Telluride-Type Gold Concentrate.

With a telluride-type gold ore flotation concentrate as the research object, the Na2S + NaOH collaborative leaching process was applied to selectively separate tellurium before the cyanide leaching of gold and silver. The effects of process parameters including the type of leaching agent, the amount of leaching agent, liquid-solid ratio, leaching temperature, and leaching time on the leaching rate of tellurium were investigated. The results showed that the tellurium leaching rate could reach 78.14% under the optimum conditions of −0.038 mm (95%) grinding fineness, 80 g/L Na2S concentration, 30 g/L NaOH concentration, 4:1 liquid-solid ratio, 80°C leaching temperature and 3 h′s leaching time. The kinetic analysis showed that the leaching process of tellurium from telluride-type gold concentrate was a mixed type of chemical reaction control and diffusion control. The grain parameter in the leaching process was 0.26263 and the apparent activation energy E = 17.12 kJ/mol. Tellurium could be pre-leached from the telluride-type gold flotation concentrate through the Na2S + NaOH alkaline leaching process to achieve the effective separation of tellurium from noble metals, which, when eliminating the adverse effects of telluride on the leaching of gold and silver, provides new ideas for the extraction of rare element tellurium.

Study on the solvent extraction of thioarsenite in alkaline solutions with transformed quaternary ammonium salt

Under the advocacy of the sustainable development concept, the extraction process and reaction mechanism for arsenic in alkaline solution were studied with the aim of optimizing the alkaline leaching process of high‑arsenic copper smelting ash and promoting the recovery and harmlessness of arsenic. At 25 °C, the extractant methyltrioctylammonium underwent 6–8 conversions of Na2CO3 solution with a concentration fraction of 12%, and the extraction capacity of thioarsenite was increased to 86.91%. Furthermore, at 25 °C, at an O/A ratio of 1:1 and an phase contact time of 7 min, the thioarsenite stripping exceeded 99%, when using 4% H2O2 + saturated Na2CO3 solution as the stripping. Fourier transform infrared spectroscopy, nuclear magnetic spectroscopy, ultraviolet spectroscopy, and other analytical methods were used to determine the mechanism of the extraction process. It was evident from our experiments that the CO32−-type TOMAC and Na2CO3 solutions could be recycled and reused repeatedly. The overall process required simple materials and did not generate liquid waste. The technique involves a short process, with superior cycle performance, low cost, and high efficiency.

High-efficiency core-shell magnetic heavy-metal absorbents derived from spent-LiFePO4 Battery

Search for simple and efficient recycling methods to utilize spent lithium-ion batteries is crucial for achieving sustainable resource development and reducing the hazardous materials released from the spent batteries. Herein, we have developed a new strategy to utilize the spent LiFePO4 batteries by utilizing the cathode plate as raw material to synthesize mesoporous core-shell adsorbent Mm@SiO2 (Mm denoted as the magnetic material) through a simple alkaline leaching process. The as-converted material exhibits excellent adsorption capacity when it has been used to remove heavy metal ions in heavy metal polluted water. The adsorption capacities for Cu2+, Cd2+, and Mn2+ have been achieved up to 71.23, 80.31 and 68.73 mg g−1, respectively. The detailed adsorption mechanism has been elucidated with comprehensive characterization techniques, including TEM, XPS, NEXAS, and EXAFS, the edge shared [Cu2O8] bipyramids can be fit against the EXAFS data to represent the atomic-scale local structure after Mm@SiO2 adsorbs Cu2+. The present work demonstrates a novel routine to reutilize the spent lithium batteries, which is of great importance to achieve sustainable development based on the “waste-to-treasure” and “waste-to-control-waste” strategy for simultaneously reducing the hazardous release from industrial solid waste and heavy metal polluted water.