Recovery Of Elements Research Articles

Covalent Organic Framework Nanoarchitectonics: Recent Advances for Precious Metal Recovery.

The recovery of precious metals (PMs) from secondary resources has garnered significant attention due to environmental and economic considerations. Covalent organic frameworks (COFs) have emerged as promising adsorbents for this purpose, owing to their tunable pore size, facile functionalization, exceptional chemical stability, and large specific surface area. This review provides an overview of the latest research progress in utilizing COFs to recover PMs. Firstly, the design and synthesis strategies of chemically stable COF-based materials, including pristine COFs, functionalized COFs, and COF-based composites, are delineated. Furthermore, the application of COFs in the recovery of gold, silver, and platinum group elements is delved into, emphasizing their high adsorption capacity and selectivity as well as recycling ability. Additionally, various interaction mechanisms between COFs and PM ions are analyzed. Finally, the current challenges faced by COFs in the field of PM recovery are discussed, and potential directions for future development are proposed, including enhancing the recyclability and reusability of COF materials and realizing the high recovery of PMs from actual acidic wastewater. With the targeted development of COF-based materials, the recovery of PMs can be realized more economically and efficiently in the future.

Concentration of goethitic slimes by comprehensively exploring a ternary Collector-Frother flotation reagent and ultrasound dispersion

The recovery of elements of interest from tailings has been increasingly studied to reduce their generation and mitigate disposal risks, and in iron ore production the reality is no different. This work presents a comprehensive assessment of parameters in bench scale tests for two innovative possible solutions for recovering iron minerals contained in ultrafine tailings: a new collector reagent formulation composed of a mixture of amidoamine, etheramine and a frother, without the need of applying a depressant in reverse cationic flotation, preceded by the application of ultrasound as a tool for physical particle dispersion. The new collector presented better results than its constituents separately. The presence of a frother in its composition helps diminishing water recovery and entrainment. The main limiting factor for enhancing the quality of the concentrate is the fraction of alumina in the feed. Molecular simulations gave initial insights on the mechanisms behind the new collector and the gangue minerals interaction. The application of ultrasound at alkaline pH promotes better particle dispersion ensuring sufficient exposed surfaces for selective adsorption of the new collector. The influence of the intensity and residence time applied at the sonication step, as well as the percentage of solids are evaluated. The study shows that it is possible to increase the metallurgical recovery by 6 % for a difficult to process, goethite-rich slimes sample with 51% Fe from the Iron Quadrangle in Brazil.

Recovery of rare earth elements from mining wastewater with aminomethylphosphonic acid functionalized 3D-printed filters

Herein we report the use of nylon-12-based 3D-printed filters incorporating α-aminomethylphosphonic acid as an active additive for the recovery of Y, Nd, and Dy from the mining waste solution containing Al, K, Ca, Sc, Fe, Co, Cu, Zn, Y, Nd, Dy, and U. Nylon-12 was chosen for the polymer matrix of the filter due to its inactivity towards the studied metals. The micrometer-level structure of the filters was studied with a scanning helium ion microscope and X-ray tomography to reveal the porosity, pore size, and active additive distribution in the filters. Furthermore, FTIR spectroscopy was used to analyze the compositional changes in the 3D-printed filters after the printing and adsorption processes. Adsorption of the metals was studied at a pH range of 1–4, and the following adsorption trend Sc > Fe > U > Y, Nd, Dy > Al, Cu, Zn > K, Ca, Co was observed in each of the studied pH values. The sequential recovery process for metals was studied at pH 2, and desorption of the metals from the filters was performed with 6 M HNO3. 100 % adsorption of REEs, Fe, and U was achieved during the recovery process, and on average, over 88 % of the adsorbed Y, Nd, and Dy were desorbed from the filters. In contrast to Y, Nd, and Dy, the desorption of Sc, Fe, and U was minimal (Fe and U) or negligible (Sc) with 6 M HNO3 due to their strong coordination to the active additive. Maximum adsorption capacities for Y, Nd, Dy, and U were determined by using linear Langmuir adsorption isotherm. The best maximum adsorption capacity was determined for Sc, Qmax = 0.51 mmol/g followed by U, Nd, Dy, and Y with capacities of 0.47, 0.24, 0.23, and 0.17 mmol/g, respectively. Overall, this study achieved a complete removal of Sc, Fe, and U from the simulated mining waste solution leaving a final eluate that mainly contained Y (320 μg), Nd (350 μg), Dy (330 μg), and Al (710 μg) demonstrating the applicability of the 3D-printed filters in the recovery of Y, Nd, and Dy from the multimetal solution.

Comprehensive recovery of valuable elements from vanadium precipitation liquor using selective impurity removal-reduction precipitation process

The vanadium precipitation liquor produced by the vanadium slag extraction process contains various valuable elements, which are not utilized at high value. In this paper, chromium is recovered through aluminum salt impurity removal - chemical reduction precipitation method. Moreover, the ammonia nitrogen and sodium sulfur in the vanadium precipitation liquor are recovered by stripping and crystallization, thereby achieving comprehensive resource recovery and utilization. The aluminum salt slag removal reduction chemical precipitation method is used to separate Cr from Si, Ca and V impurities. As a result, the obtained Cr2O3 product has a purity of 98.20%, the chromium recovery efficiency is 97.58%, and the silicon removal efficiency reaches 99.66%. The ammonia nitrogen was recovered by stripping, which would be reused to the vanadium precipitation process, with an ammonia nitrogen stripping efficiency of 98%. The purity of NH4Cl is 99.13% and meets the 99-level national standard. The Na2SO4 was recovered by evaporation concentration - crystallization method, which would be reused to the vanadium slag roasting process, with a Na2SO4 recovery efficiency of 88.38%. And the purity of Na2SO4 is 98.85%, which meets the industrial standard for Class II Na2SO4 products. The final solution after comprehensive recovery of valuable elements would be reused to the leaching process, thereby achieving water circulation. This process realizes the high-value utilization of chromium, ammonia, nitrogen, sodium and sulfur, and has no wastewater discharge and is an environment-friendly route.

Technological Advancement in Solar Photovoltaic Recycling: A Review

This review examines the technological surveillance of photovoltaic panel recycling through a bibliometric study of articles and patents. The analysis considered the number of articles and patents published per year, per country, and, in the case of patents, per applicant. This analysis revealed that panel recycling is an increasingly prominent research area. However, the number of patents filed annually has varied in recent years, averaging fewer than 200 per year. The state-of-the-art review identified three main types of treatment for photovoltaic panel recycling: mechanical, chemical, and thermal. Among these, mechanical treatment serves as a preliminary stage before the recovery of valuable elements, which is achieved through chemical or thermal processes. The articles reviewed cover a range of processes, including hydrometallurgical and pyrometallurgical methods, and explore various classification processes, solvents, and oxidizing agents. In contrast, patents predominantly focus on pyrometallurgical processes. This analysis is supplemented by a survey of market-ready technologies, many of which include stages such as size reduction or delamination followed by pyrometallurgical processes. Additionally, the review highlights the collection processes implemented by some companies, noting that the volume of panels considered waste is currently insufficient to maintain a continuous and year-round operational process. This study identifies key challenges such as (i) reducing solar panel size due to the EVA polymer complicating conventional machinery use, (ii) high process costs from the need for high temperatures and costly additives, (iii) the environmental impact of thermal treatments with high energy consumption and air pollution, and (iv) the necessity for environmentally friendly solvents in hydrometallurgical treatments to reduce contamination during recycling. Future directions include developing specific machinery for panel size reduction, either creating or modifying a polymer to replace EVA for easier treatment, adopting hydrometallurgical treatments with green solvents proven effective in recycling minerals and electronic waste, and addressing the lack of detailed information on industrial processes to make more precise recommendations.

Selective leaching and recovery of rare earth from NdFeB waste through a superior selective and stable deep eutectic solvent

The recovery of rare earth elements (REEs) from NdFeB waste is crucial for advanced industrial development. The green and superior selective recovery methods can effectively avoid the problem of generating large amounts of wastewater and high pollution in the NdFeB waste recycling process. In this work, a novel deep eutectic solvent (DES) synthesized by tetraethylammonium chloride (TEAC) and levulinic acid (LA) was designed and achieved selective leaching of REEs from NdFeB waste without wastewater produced. The leaching ratio of Nd can reach up to 97.63 %, while Fe is less than 0.435 %, the separation factor can exceed 9000, and the purity of the obtained product Nd2O3 is 99.649 %. In addition, the solubility of Nd2O3 in TEAC-LA is over 75.6 g/L, the thermal stability temperature reaches 132.4 °C much higher than the working temperature, and the viscosity is 12.8 cP at the working temperature which is close to prevalent organic solvents. After three cycles, TEAC-LA still maintains its original selective leaching ability and shows excellent thermal stability. This study provides a superior selective and stable method for green and efficient recycling of NdFeB waste.

Recovery of Y(III) from wastewater by Pseudomonas psychrotolerans isolated from a mine soil

While microbial technologies, which are considered to be environmentally friendly, have great potential for the recovery of rare earth elements (REEs) from mining wastewater, their applications have been restricted due to a lack of efficient biosorbents. In this study, a strain of Pseudomonas psychrotolerans isolated from yttrium-enriched mine soil was used to recover yttrium (Y(III)) from rare-earth mining wastewater. At an initial Y(III) dose of 50 mg L−1, the amount of Y(III) adsorbed by P. psychrotolerans reached 99.9 % after 24 h. Various characterization techniques revealed that P. psychrotolerans adsorbed Y(III) mainly through complexation of oxygen-containing functional groups and electrostatic interactions. A high level of adsorption efficiency (>99.9 %) was maintained after five consecutive adsorption/desorption cycles, indicating that P. psychrotolerans was highly reusable. While the efficiency of adsorbing Y(III) by P. psychrotolerans decreased (34.4 %) in actual rare earth mining wastewater, selectivity toward other REEs (≤ 18.4 %) was still observed. Consequently, this study provides a promising green, environmentally friendly and sustainable microbial approach for the selective recovery of REEs from rare earth wastewater.

Highly selective recovery of rare earth elements from mining wastewater using phyto-synthesized biochar dispersed iron nanoparticles

Recovery of rare earth elements (REEs) from secondary sources such as mining wastewater has attracted much recent attention due to the rising global demand for REEs and the need to environmentally protect potentially contaminated mine sites. One significant issue it that most adsorbents proposed for removal and/or recovery of REEs are not generally selective for REEs. In this study, biochar dispersed iron nanoparticles synthesized using a plant extract (BC-FeNPs) were successfully used for the selective recovery of REEs from mining wastewater, where the Kd values were Nd(III) (918 mL∙g−1), Eu(III) (1253 mL∙g−1), Tb(III) (1119 mL∙g−1), Dy(III) (1032 mL∙g−1), Lu(III) (1898 mL∙g−1), compared to only 23.9 mL∙g−1 for Zn(II). The maximum adsorption efficiency of REEs for the composite was higher than the constituent parts being 88.4 % for BC-FeNPs, but only 8.72 % for biochar and 82.4 % for FeNPs, respectively. FTIR, XPS and Zeta potential analysis suggested that the observed selective adsorption of REEs, involved interactions of REEs with O and N, as well as ion-exchange with H+ from the biochar and capping layer, as well as electrostatic interactions. The desorption efficiency of bound REEs from BC-FeNPs was > 85 % when using acetic acid, and was attributed to efficient competitive ion exchange. Pearson correlation analysis further demonstrated that the adsorption mechanism included ion complexation, ion exchange and electrostatic adsorption, while the desorption mechanism was mainly via ion exchange. Overall, BC-FeNPs has significant potential to practically recover REEs from mining wastewater having demonstrated excellent reusability after five adsorption–desorption cycles.

Rare earth elements recovery and mechanisms from coal fly ash by column leaching using citric acid

Rare earth elements (REEs) play an irreplaceable role in supporting the advancement of the global economy and technology, but their limited supply drives researchers to devise effective strategies for the recovery of REEs from alternative sources, including coal fly ash (CFA). In this work, a new column leaching method using citric acid as the lixiviant was proposed to enhance the recovery of REEs from CFA at room temperature. The mechanism of column leaching was investigated based on sample properties, leaching results, and the adsorption characteristics and complexation of citric acid. REEs in the CFA were found to be hosted into the amorphous aluminosilicate matrix in the form of oxides and fluorides using SEM-EDS. Compared to the maximum recovery of total REEs (TREEs, < 30 %) in the conventional agitation leaching tests, the recovery of TREEs in the column leaching tests increased significantly to 64.8 %, while the leaching rate of impurities (Al, Ti, and Fe) was less than 13.7 %. Results of the Zeta potential measurements showed that the citric acid and rare earth cations were adsorbed on the surface of CFA particles. These adsorption phenomena may make REEs bind to the CFA surface, thereby resulting in low leaching recoveries of REEs. Rare earth-citrate complexes in the lixivium were isolated by antisolvent precipitation, demonstrating that rare earth cations can be complexed by citrate. During the column leaching process, the leached rare earth ions were complexed with citrate and then carried away with the flowing liquid phase, which improved the recovery of REEs.

Improvements in recovery of rare earth elements (REEs) from coal via fluidized-bed combustion: Thermal alteration of REE mineralogy and its impact on element extractability

The extraction of rare earth elements (REE) from coal ash has attracted significant attention, but satisfactory solutions have not yet been achieved. Indeed, the REE extractability is largely determined by the transformation behavior of REE during coal combustion. This study found that the controlled combustion conditions of circulated fluidized-bed boiler (CFB) can both utilize the heating value and enhance the recovery of REE from coal. Tests were conducted using a bituminous coal with REE concentration of approximately 149 ppm in an industrial 35 t/h CFB boiler with controlled boiler loads and coal sources. Advanced automated image analysis by scanning electron microscopy (AIA-SEM) and sequential leaching procedure were employed to clarify the speciation transformation behavior of REE in boilers and investigate the effects of these transformations on the leachability of REE. Approximately 75 % of REE partitioned into fly ash and enriched in particles < 45 μm (∼710 ppm) after combustion. Acid leaching tests under mild conditions (1 M HCl at 60 ℃ and S/L = 1:25) performed on coal, bulk ashes, and sieved ashes revealed an improvement in REE leaching ratio from coal (21.34 %) to fly ash (71.07 %). Particle size played a crucial role in acid extractability, particularly for middle and heavy REE in cyclone ash but had no influence on those in fly ash. The dominant forms of REE in coal including Ca-REE fluorocarbonates (36.7 %), aluminophosphates (12.0 %), and organic-bound REE (approximately 20 %) significantly decreased, contributing to the formation of leachable rare earth oxides (∼65 %) in ash. Furthermore, the small size of REE mineral grains in coal (over 50 % below 5 μm) and extensive mineral fragmentation occurring in boiler resulted in the enrichment and enhanced REY leachability in fine ash.

Recovery of neodymium and iron from NdFeB waste by combined electrochemical–hydrometallurgical process

The recovery of rare-earth elements and Fe from NdFeB waste has significant potential. However, conventional hydrometallurgical and pyrometallurgical processes are energy intensive, chemical consuming, and accompanied by waste emissions, resulting in low-value iron products. This study introduced a combined electrochemical–hydrometallurgical process for NdFeB waste recovery. Electrochemical analysis revealed the active dissolution of NdFeB waste under 370 mA/cm2, enabling Fe2+ reduction to Fe even in the presence of Nd3+. Electrolysis at 60 °C using Nd2(SO4)3 and FeSO4 as electrolytes, along with H3BO3 as a pH buffer, resulted in the uniform dissolution of waste, releasing Nd3+ and Fe2+ into the electrolyte, whereas Fe2+ underwent electrodeposition onto Fe at the cathode. Investigation of FeSO4 concentration revealed its influence on current efficiency, cathode-product properties, and energy consumption. When the concentration of FeSO4 was 0.5 M, a cathode product with 87.94 wt% Fe and 0.03 wt% Nd was obtained. Finally, Nd was recovered by hydrometallurgical high-temperature crystallization at 90 °C under N2 atmosphere, yielding Nd2(SO4)3·nH2O mixed crystals with 0.1 wt% Fe. The process formed a closed loop, consuming only FeSO4 and a minimal amount of H2SO4, with an energy consumption of 0.73–1.53 kW·h/kg NdFeB.

Overview of Functionalized Porous Materials for Rare-Earth Element Separation and Recovery.

The exceptional photoelectromagnetic characteristics of rare-earth elements contribute significantly to their indispensable position in the high-tech industry. The exponential expansion of the demand for high-purity rare earth and related compounds can be attributed to the swift advancement of contemporary technology. Nevertheless, rare-earth elements are finite and limited resources, and their excessive mining unavoidably results in resource depletion and environmental degradation. Hence, it is crucial to establish a highly effective approach for the extraction and reclamation of rare-earth elements. Adsorption is regarded as a promising technique for the recovery of rare-earth elements owing to its simplicity, environmentally friendly nature, and cost-effectiveness. The efficacy of adsorption is contingent upon the performance characteristics of the adsorbent material. Presently, there is a prevalent utilization of porous adsorbent materials with substantial specific surface areas and plentiful surface functional groups in the realm of selectively separating and recovering rare-earth elements. This paper presents a thorough examination of porous inorganic carbon materials, porous inorganic silicon materials, porous organic polymers, and metal-organic framework materials. The adsorption performance and processes for rare-earth elements are the focal points of discussion about these materials. Furthermore, this work investigates the potential applications of porous materials in the domain of the adsorption of rare-earth elements.

Recovery of Lithium from Industrial Li-Containing Wastewater Using Fluidized-Bed Homogeneous Granulation Technology

In this study, a novel fluidized-bed homogeneous granulation (FBHo-G) process was developed to recover lithium (Li) from industrial Li-impacted wastewater. Five important operational variables (i.e., temperatures, pH, [P]0/[Li]0 molar ratios, surface loadings, and up-flow velocities (Umf)) were selected to optimize the Li recovery (TR%) and granulation ratio (GR%) efficiencies of the process. The optimal operational conditions were determined as the following: a temperature of 75 °C, pH of 11.5, [P]0/[Li]0 of 0.5, surface loading of 2.5 kg/m2·h, and Umf of 35.7 m/h). The TR% and GR% at optimal condition could be as much as 90%. The material characterization of the recovery pellet products showed that they were highly crystallized Li3PO4 (purity ~88.2%). The pellets had a round shape and smooth surface with an average size of 0.65 mm, so could easily be stored and transported. The high purity enables them to be further directly reused as raw materials for a wide range of industrial applications (e.g., in the synthesis of cathode materials). Our calculation shows that the FBHo-G process could recover up to 0.1845 kg of lithium per cubic meter of Li-containing wastewater, at a recovery rate of ~90%. A brief technoeconomic analysis shows that FBHG process had economic viability, with an estimate production cost of USD 26/kg Li removed, while the potential gained profit for selling lithium phosphate pellets could be up to USD 48 per the same volume of wastewater and the net profit up to USD 22/m3 Li treated. In all, fluidized-bed homogeneous granulation, a seedless one-step recovery process, opens a promising pathway toward a green and sustainable recycling industry for the recovery and application of the resource-limited lithium element from nonconventional water sources.

Recovery of rare earth elements from rare earth molten salt electrolytic slag via fluorine fixation by MgCl2 roasting

Rare earth fluoride molten-salt electrolytic slag (REFES) is a precious rare earth element (REE) secondary resource, and considerable amounts of REEs exist in REFES as REF3; they are difficult to dissolve in acid or water and impede efficient REE extraction. In REFES recovery, the REF3 species in REFES are usually transformed into acid-soluble rare earth compounds by NaOH roasting or sulfating roasting and then extracted by acid leaching. Moreover, the fluorides in REFES are released as HF gas in the roasting process or enter the liquid phase during the water washing process; both of these processes cause fluorine pollution. Fixing the fluorine into the solid slag provides a way to avoid fluorine pollution. In this study, a novel method was proposed to extract REEs from REFES via MgCl2 roasting followed by HCl leaching. Thermodynamics calculations and thermogravimetry‒differential thermal analyses (TG-DTA) were conducted to investigate the reactions occurring in the roasting process. First, MgCl2 reacts with the REF3 and RE2O3 to form RECl3 and REOCl, respectively. Second, the RECl3 absorbs water and forms RE(OH)3. Third, MgCl2·6H2O is gradually dehydrated to MgCl2·2H2O and reacts with REF3 and RE(OH)3, and REOCl, MgF2 and MgO are formed. Through HCl leaching, the REOCl in the roasting products is leached by HCl acid, while fluoride remains in the solid slag as MgF2. The optimum experimental conditions are as follows: mass ratio of MgCl2 to REFES of 30%, roasting temperature of 700 °C, roasting time of 2 h, hydrochloride acid concentration of 4 mol/L, leaching time of 2 h, leaching temperature of 90 °C and leaching L/S ratio of 20:1. The efficiencies for total leaching of the REEs, La, Ce, Pr, and Nd are 99.13%, 99.20%, 98.42%, 99.38%, and 99.08%, respectively. Moreover, the concentration of fluoride in the leaching solution is 2.191 × 10−6 mol/L. This method has a short process flow with low reagent costs, and the problem of fluoride pollution from REFES recovery is solved; thus, our study has great industrial application potential.

Utilization of Eggshell Waste Calcite as a Sorbent for Rare Earth Element Recovery.

The green energy transition requires rare earth elements (REE) for the permanent magnets used in electric cars and wind turbines. REE extraction and beneficiation are chemically intensive and highly damaging to the environment. We investigated the use of eggshell waste as a sustainable alternative sorbent for the capture and separation of REE from aqueous solutions. Hen eggshell calcite was placed in multi-REE (La, Nd, Dy) solutions at 25 to 205 °C for up to 3 months. A pervasive diffusion of the REE inside the eggshell calcite was observed along pathways formed by the intracrystalline organic matrix and calcite crystal boundaries. At 90 °C, kozoite (REECO3OH, orthorhombic) spherulites precipitate on the surface of the dissolving calcite. At 165 and 205 °C, an interface-coupled dissolution-precipitation mechanism is observed, resulting in the complete dissolution of the calcite shell and its pseudomorphic replacement by polycrystalline kozoite. At 205 °C, kozoite is slowly replaced by hydroxylbastnäsite (REECO3OH, hexagonal), the stable form of the rare earth hydroxycarbonate polymorphs. Our results demonstrate two potential applications of eggshell waste for the uptake of rare earth elements in solution: at low temperatures, as a mixed organic-inorganic adsorbent and absorbent, given sufficient sorption time; and at higher temperatures, as an efficient sacrificial template for the precipitation of rare earth hydroxycarbonates.

Recovery of rare earth elements from sedimentary rare earth ore via sulfuric acid roasting and water leaching

Rare earth elements were extracted using a sulfuric acid roasting-water leaching process. The effect of acid roasting on a new type of low-grade sedimentary rare earth ore found in Guizhou Province, China was analyzed using X-ray diffraction and scanning electron microscopy. A systematic study was conducted on process parameters such as amount of acid, roasting temperature, roasting time, water leaching temperature, and leaching time. The results reveal that the total recovery of rare earth elements reaches 81.37%, which is 3.1 times higher than that achieved through direct acid leaching, under the optimal conditions. In addition, the leaching rate of heavy rare earth elements reaches 72.53%. Rare earth elements and some other valuable metals are transformed into soluble sulfate through the local decomposition of clay minerals under the action of the sulfuric acid attack. The dissolution rates of aluminum, iron, and titanium ions are 34.94%, 17.05%, and 62.77%, respectively. The precipitation rate of Ti reaches 99%, and the loss of rare earth ions in the solution is less than 1%. Meanwhile, the results of a leaching kinetics analysis indicate that the leaching process of rare ions is controlled by diffusion. Precious metal ions such as iron and aluminum in the leaching solution can reduce the adsorption of rare earth ions by kaolinite. This study efficiently recovered rare earth ions under conditions of low calcination temperature and direct water leaching, resulting in reduced energy consumption of the extraction process and acidity of the leaching solution. These findings provide a solid foundation for the further separation and extraction of rare earth ions.

Self-sulfidation roasting and flotation of electroplating sludge: An innovative technology for metal recovery

Electroplating sludge is classified as a solid hazardous waste, which presents significant environmental and human health risks, while also resulting in substantial depletion of valuable metals. The study proposed an innovative technology of self-sulfidation roasting and flotation for the efficient recovery of valuable metal elements in electroplating sludge. The distinguishing feature lied in utilizing the calcium sulfate present in the electroplating sludge as the sulfidation reagent, thereby achieving self-sulfidation roasting without the need for additional sulfidation reagents. The decomposition process of calcium sulfate and influence factors during self-sulfidation roasting were investigated by thermogravimetric analysis and roasting experiments. Under optimized conditions, copper and nickel sulfidation ratios in the electroplating sludge reached 82.80% and 97.39%, respectively, while zinc volatilization ratio reached 87.98%. XRD and SEM-EDS techniques were employed to reveal phase evolution and surface morphology of the roasted products. Finally, copper-nickel concentrate achieved recovery ratios of 64.37% for copper with a grade of Cu 2.7%, and 67.38% for nickel with a grade of Ni 12.84%. This study presented a novel and effective approach towards secondary utilization of solid wastes.

Recovery of rare earth elements from weathering crust soils using electrokinetic mining technology☆

Ion-absorption deposits (IADs) of rare earth elements (REEs) constitute the main economic resources of the essential heavy rare earth elements (HREE). Nonetheless, the existing leaching methods for extracting REEs from IADs face limitations stemming from significant environmental consequences and a poor REE recovery rate. In a recent development, electrokinetic mining (EKM) has emerged as a novel technology for extracting REEs from ion-adsorption deposits, demonstrating the potential for environmentally friendly and efficient recovery of REEs from weathering crusts. However, the transport mechanism of REE in weathering crust soil in an applied electric field remains poorly understood, and the influence of the EKM process has yet to be studied. In this study, we systematically investigated the transport characteristics of REE in weathering crust soil under the influence of an electric field. We demonstrate that the transport of REE is simultaneous with the transport of water, H+, and OH– in the applied electric field, where the EKM process is influenced by the voltage gradient, initial soil water content, electrode numbers, and power-on time. Under the optimal EKM conditions, 78.8% of REE is recovered from a 45 kg-scale weathering crust soil. Additionally, we identify the transport diversity between light REE (LREE) and HREE, which assists in the separation and pre-enrichment of LREE and HREE. Furthermore, we demonstrate that the mechanism for REE transport under an electric field involves a combination of electromigration, electroosmosis, and electrolysis working synergistically. This study provides new perspectives on the transport behavior and mechanism of REEs in weathering crust soil under an electric field. These findings pave the way for the practical implementation of EKM technology.

In Situ Modification of Activated Carbon Made from Camellia oleifera Shell with Na2EDTA for Enhanced La3+ Recovery

It is important to recover La3+ from metallurgical solutions or wastewater. However, the recovery rate of La3+ is usually less than 1% and the recovery methods are not environmentally friendly or user-friendly. Therefore, a straightforward, efficient, clean, and economically friendly method is needed. In this investigation, a modified adsorbent, COSAC-Na2EDTA-15, which was made from the Camellia oleifera shell (COS) and disodium ethylenediaminetetraacetic acid (Na2EDTA), was invented. In addition, characterization of the COSAC-Na2EDTA-15 adsorbent was conducted using SEM and XPS, and the principle of adsorption was revealed. The adsorption kinetics followed P-S-O KM, while the isotherm of COS-activated carbon (COSAC) aligned more closely with the Langmuir model. Compared to COSAC, the maximum La3+ adsorption capacity of COSAC-Na2EDTA-15 increased from 50 to 162.43 mg/g, and the content of O and N changed from 7.31% and 1.48% to 12.64% and 4.15%, respectively. The surface of the COSAC-Na2EDTA-15 exhibited abundant C, N, and O elements, and La3+ was detected on the sample surface after adsorption. The test and analysis results fully indicate that La3+ can be successfully adsorbed on the surface of COSAC-Na2EDTA-15. Because of its easy preparation, low cost, and superior performance, activated carbon made from COS finds extensive applications in the adsorption and recovery of rare earth elements.

Engineering Compounds for the Recovery of Critical Elements from Slags: Melt Characteristics of Li5AlO4, LiAlO2, and LiAl5O8.

Engineered artificial minerals (EnAMs) are the core of a new concept of designing scavenger compounds for the recovery of critical elements from slags. It requires a fundamental understanding of solidification from complex oxide melts. Ion diffusivity and viscosity play vital roles in this process. In the melt, phase separations and ion transport give rise to gradients/increments in composition and, with it, to ion diffusivity, temperature, and viscosity. Due to this complexity, solidification phenomena are yet not well understood. If the melt is understood as increments of simple composition on a microscopic level, then the properties of these are more easily accessible from models and experiments. Here, we obtain these data for three stoichiometric lithium aluminum oxides. LiAlO2 is a promising EnAM for the recovery of lithium from lithium-ion battery pyrometallurgical processing. It is obtained through the addition of aluminum to the recycling slag melt. The high temperature properties spanning from below to above the liquidus temperature of three stoichiometric Li-Al-Oxides: Li5AlO4, LiAlO2, and LiAl5O8 are determined using molecular dynamic simulations. The compounds are also synthesized via the sol-gel route. The Li+ ion exhibits the largest diffusivity. They are quite mobile already below the liquidus temperature, i.e., for LiAlO2 at T = 1700 K, the diffusion coefficient of the lithium ion equals D = 3.0 × 10-9 m2 s-1. The other ions Al3+ and O2- do not move considerably at that temperature. The diffusivity of Li+ is largest in the lithium-rich compound Li5AlO4 with D = 32 × 10-9 m2 s-1 at 2500 K. The lower the viscosity, the higher the lithium content. The Li5AlO4 exhibits a viscosity of η = 2.2 mPa s at 1328 K which matches well with the experimentally determined 2.5 mPa s at this temperature. The viscosity of LiAlO2 at 1800 K is more than two times higher. These data sets can help to describe the melts on a microscopic level and understand how the melt properties will change due to gradients in the Li/Al concentration.