Leach Liquor Research Articles

Processing of ferromanganese fumes into high-purity manganese sulphate monohydrate

ABSTRACT A novel process for recovering manganese (Mn) values from two types of ferromanganese (FeMn) fumes (wastes from smelters) was developed incorporating several stages of leaching, roasting, and precipitation. Fumes containing high K (as K2O) were first leached in water at ambient conditions to recover K, purified to remove metal impurities, and precipitated as 99.9% pure K2SO4. The residues containing Mn and remaining impurities were then leached in sulfuric acid at 80 °C using oxalic acid as reductant yielding Mn sulphate liquors. The leach liquors were then purified to remove Fe at 95°C as jarosite at pH 2.8 and Fe(III) hydroxides at pH 5.5. Base and transition metal impurities were removed as sulfides. K, Na, Ca, and Mg were separated from Mn in a novel step based on precipitation of Mn hydroxide from the Mn sulphate liquors. Subsequent redissolution of Mn hydroxide in sulfuric acid under reducing conditions using oxalic acid produced a high purity Mn sulphate liquors containing <5 ppm of these impurities. After evaporation by boiling manganese sulphate monohydrate (MSM) was crystallized, achieving purity >99.5%. Another fume containing oil contaminants having low K was subjected to sulfation roasting to produce Mn sulphate. The calcine was water leached to produce Mn sulfate liquors and treated in a similar process to make high purity MSM. Novelty aspects of the proposed process include the recovery of K as a by-product and separation of Mn2+ from other impurities by its selective precipitation to produce a high purity Mn sulphate solution for MSM recovery. Implications: Ferro-manganese smelters discharge wastes (fumes) that are collected by bag filters or electrostatic precipitators and then store them on the ground with coverage to prevent their dispersion into the surrounding atmosphere. The work presented propose a process to reprocess these wastes into high purity manganese sulphate monohydrate (MSM). Processing of wastes containing high potassium as KOH also yields high purity potassium sulfate used as fertilizers. Some smelters spray oil onto the fume dumps to prevent material being blown onto the surrounding area. The oil-contaminated fumes have to be first roasted to burn off the oil, before they can be processed. The process developed therefore converts a hazardous waste into high purity MSM and potassium sulphate used in many applications.

An innovative environmental process for the treatment of scrap Nd-Fe-B magnets

Strict environmental regulations as well as requirement of conservation of natural resources compelled the researchers to recycle the metal values from secondary resources. The scrap magnets found to be a potential alternative resource to extract rare earth metal, Nd. Present paper reports a novel process flow-sheet for the recycling of scrap Nd-Fe-B magnets to recover Nd as marketable salt and other valuable by-products. The Nd-Fe-B magnets were demagnetized, crushed and charged to atmospheric leaching resulting in ~99.99% recovery of REMs (Nd, Pr, Dy) and Fe using 1 M H2SO4 solution at room temperature for 90 min and pulp density 50 g/L. The obtained leach liquor was subjected to acid extraction procedure by mixing the liquor with 70% TEHA diluted in kerosene for 10 min, which requires five stages for complete extraction of acid from the liquor having O/A ratio 2:1. Ammonia solution was used to increase the pH of acid free leach liquor to 1.65 for the precipitation of Nd (with minor amount of Pr and Dy) and above that to precipitate Fe. Further, processing of these valuables make them industrially applicable. The leached residue was checked using Toxicity Characteristics Leachability Procedure (TCLP) test and found the remained metals within the permissible limit. This process offers a simple and efficient means to reduce the toxicological effect by recovering Nd from magnets of computer hard disks.

Extraction equilibria of cerium(IV) with Cyanex 923 followed by precipitation kinetics of cerium(III) oxalate from sulfate solution

The phenomenon studies on solvent extraction of cerium(IV) using Cyanex 923 followed by the precipitation kinetics and recovery of cerium(III) oxalate has been investigated from the sulfate media. The extraction equilibria investigated with varying concentrations of extractant into the organic phase and acid into the aqueous phase could reveal the solvation of [Ce(SO4)2·2L·HSO4-]org species. The thermodynamic properties determined at different temperatures exhibited the exothermic nature of extraction (ΔH°, −12.2 kJ·mol−1). The quantitative extraction and back-extraction of cerium(IV) in sulfate media could be achieved in three stages of counter-current contact at an organic-to-aqueous ratio of 2/3 and 3/2, respectively. Ce-enrich stripped solution was subsequently recovered as cerium(III) precipitates by treating with oxalic acid. Precipitation with the optimized dosage of Ce:H2C2O4 as 1:≥10 yielded the oxalate products of unlike physico-chemical characteristics, while precipitating at different temperatures. The precipitation kinetics was found to have good fits to the Avrami equation; where, n values < 1 indicated for instantaneous nucleation that leads to flakes-shape growth of precipitates and supported by the surface morphological study. Evaluation of the apparent activation energy (Ea, 8.6 kJ·mol−1) indicated that the precipitation reaction follows a diffusion-controlled mechanism as supported by the surface site saturation of cerium(III) oxalate through the constants of Avrami equation. The findings of the present study may supply a potential way of extracting cerium(IV) from sulfate leach liquor of rare earths ores and the precipitation recovery of cerium(III) oxalate as the suitable precursor of (nano)crystalline ceria.

Recovery of Ag, Cu, Ni and Fe from the nitrate leach liquor of waste ICs

Novel and simple flow-sheet has been reported to process integrated circuits (ICs) for effective recovery of metals. The depopulated ICs from waste PCBs were pulverised, beneficiated and leached in nitric acid. From the nitrate leach liquor, Fe was removed at pH 3.5, temperature 60 °C and air sparging. Subsequently, 99.99% Ag was precipitated as AgCl using 5% NaCl in mixing time 10 min. Further, 99.99% Cu was extracted from the filtrate using 20% LIX 984 N along with 2% isodecanol (ID) diluted in kerosene at pH 2.97, organic to aqueous phase ratio 1 in 10 min, leaving Ni in the raffinate. From the obtained pure solutions, salts of Cu and Ni could be produced using evaporation technique. The developed hydrometallurgical process flow-sheet to recover valuables from the nitrate leach liquor of waste ICs has the potential to be commercialised after pilot/ scale-up trial validations.

Iron Precipitation Strategies from Nickel Laterite Ore Sulfuric Acid Leach Liquor

ABSTRACT A two-step closed circuit, operated at 90–95ºC, was found suitable for practical iron removal from real sulfuric acid leach liquors, resulting in low losses of nickel and cobalt to the solid phase. In the first step, NH4OH was used as a reagent to treat iron-rich liquors (1.8 g.L−1 < [Fe]total <8.7 g.L−1) with an initial pH of 1.6 ± 0.1, and nickel and cobalt concentrations of 1.59 ± 0.05 g.L−1 and 108 ± 6 mg.L−1, respectively. Metal losses are greatly affected by the liquor’s pH, which increased when liquors with respective higher initial [Ni]/[Fe] and [Co]/[Fe] concentration ratios were treated. Losses were mostly associated with adsorption reactions on the surface of the precipitate, identified mainly as jarosite and goethite phases. Iron removal higher than 99% was obtained at pH 2.5–3.3, with metal losses of up to 1.8%. In the second step, milk of lime (1.5 CaO g/100 mL) was used as a reagent. A final solution containing only 2.1 mg.L−1 of iron, with nickel and cobalt losses below 1.6%, was achieved at pH 3.7–3.8; the precipitate was identified as bassanite. The precipitates presented rapid sedimentation and good filterability, which were observed visually. Filtration operation should be carried out at temperatures above 80°C. A brief review of the literature is also presented.

Recovering valuable metals from the leaching liquor of blended cathode material of spent lithium-ion battery

Blended cathode material for lithium-ion battery has been widely applied in electric vehicles, but effectively recovering valuable metals from the leaching liquor of blended cathode material is difficult due to the high content of manganese. In this study, an effective approach for recovering valuable metals from high-manganese leaching liquor of spent blended cathode materials is proposed. Firstly, 97.5% nickel and 100% cobalt in high-manganese leaching liquor are co-precipitated by sodium ethyl-xanthate under the optimum conditions of pH=5, 2 h, 50 °C and the dosage of sodium ethyl xanthate of 55 g·L-1. Then, over 95% nickel in precipitate is selectively leached into solution while almost 100% cobalt is still left in precipitate through a two-step ammonia leaching under the conditions of 2 mol·L-1 ammonia, 50 g·L-1 ammonium sulfate, L/S=8 ml·g-1, 1 h and 40 °C. Finally, the manganese and lithium left in leaching liquor are separated by solvent extraction. Over 99% manganese is effectively extracted and separated from lithium on the conditions of 40% vol D2EHPA, pH=5, 25 °C, the ratio of organic to aqueous (O:A)=1.25 and extraction time of 10 min. The whole recovering process is facile and can help effectively recycle lithium, nickel, cobalt and manganese from spent blended cathode materials.

A carrageenan/agarose composite sponge and its immunomodulatory activities toward RAW264.7.

A kind of commercial hybrid carrageenan (HC)/agarose composite sponge containing κ-, μ-, ι-, and ν-carrageenan, which could turn into hydrogel and release carrageenan at human body temperature was fabricated for immune stimulation and modulation. Release behavior demonstrated that the hybrid carrageenan contained sponge was mechanically stable and could release carrageenan constantly. RT-PCR and ELISA experiments showed that the leaching liquor of the sponge could stimulate RAW264.7 from M0 state to a polarized state by secreting more anti-inflammatory factor IL-10 than pro-inflammatory ones, such as, IL-6 and TNF-α. Transwell experiments also indicated that the leaching liquor could promote the proliferation of NIH-3T3 by stimulating RAW264.7 of M0 state after 7 days. Results of particle size and intracellular concentration analyses suggested that the released carrageenan might enter into the cellular interior of RAW264.7 in the form of microgels or protein complexes. The sponge would be a promising candidate for skin wound dressing.

Nanoporous colorant sensors and captors for simultaneous recognition and recovery of gold from E-wastes

Platinum group metals have gained significant interest due to their unique characteristics, which make them the main constituents in advanced applications. In this work, we introduce new pH-dependent optical mesocaptors for the colorimetric monitoring and separation of Au(III) from E-waste leach liquors without a preconcentration process. The mesoporous silica nanospheres are fabricated via simple, reproducible, and low-cost procedures. The optical mesocaptor is designed via indirect immobilization of thiazole yellow G (TYG) and amacid yellow M (AYM) chromogenic probes onto mesoporous nanostructured scaffolds. The silanol groups in the mesopores of silica surface robustly anchored dilauryl dimethyl ammonium bromide (DDAB) linker to induce the interactions with the TYG and AYM chelates, thereby leading to the fashioning of a stable optical mesocaptors without releasing of the chelates throughout adsorption and sensing assays. The finding provides evidence of the capability of the synthesized decorated new nanostructured sensor shows excellent sensitivity toward Au(III) with a limit of detection (LOD) as low as 1.16 µg L−1. Furthermore, the new sensors were able to selectively detect Au(III) in solution with multi ions components.

Recovery of REEs from leaching liquor of ion-adsorbed-type rare earths ores using ionic liquid based on cooking oil

Ionic liquid (IL) based on tetraheptylammonium cooking oil (TC) was used for the first time to recovery rare earth elements (REEs) from leaching liquor of ion-adsorbed-type rare earths ores. Firstly, extraction performances of tetraheptylammonium oleate (TO) and tetraheptylammonium linoleate (TL) type ILs were investigated. It was shown that these ILs had good extraction performances, such as short equilibrium time less than 1 min, good extraction capacity of 0.225 mol/L and high stripping efficiency of 99%. Additionally, REEs extraction occurred in the solutions with high acidity (pH = 0.24) because the cation (C7H15)4N+ of TO and TL could form a strong amine salt with hydrogen ion to decrease acidity of the feed solution. Also, an ion association mechanism of REEs extraction was proposed by FT-IR spectra, 1H NMR investigation and slope analysis method. Secondly, TC was synthesized in water bath by mixing tetraheptylammonium bromide with saponified peanut oil. Then, comparative study of commercial extractants showed that TC had the highest REEs recovery yield of 100% from the leaching liquor. The sustainability, biocompatibility and non-toxicity, as well as simple synthesis route and well extraction performances make these cooking oil based ILs promising for REEs recovery from the leaching liquor.

Cleaner production of rare earth elements from phosphorus-bearing sulfuric acid solution of vein deposit monazite

In the present study, the recovery of rare earth elements (REEs; viz. cerium, lanthanum, and yttrium) from sulfate leach liquor of vein deposit Korean monazite has been investigated using the solvo-chemical techniques. The dissolved phosphorus (373 mg L−1), which may significantly affect the products’ purity, was selectively removed using the benzyl-di-methyl amine (BDMA). The logarithmic distribution vs. BDMA concentration along with spectral analysis of the organic phases revealed the formation of extracted species 3[BDMA].[H3PO4]¯. The quantitative removal of phosphorus was achieved in three stages of counter-current extraction (CCE) under the organic-to-aqueous (O:A) ratio of 1.0:1.5, leaving REEs’ into the raffinate. Subsequently, Cyanex 923 was employed to selectively extract cerium(IV) from P-depleted solution under the varying parameters of extractant and sulfuric acid concentration, temperature, and O:A ratio. The quantitative extraction of cerium (∼99% from 864 mg L−1 feed concentration) showing exothermic nature was achieved at an O:A of 1:2 under three stages of CCE while using 0.15 mg L−1 extractant into organic phase and 1.12 mol L−1 H2SO4 into aqueous phase. Further, the experimental results and spectral analysis of the organic phases could reveal that cerium solvates as [Ce(SO4)2].2[Cyanex 923].[HSO4−]¯ species. Cerium was successfully stripped from loaded organic by contacting with a higher concentration of H2SO4 solution (≥4.0 mol L−1) in presence of 0.1 mol L−1 H2O2, yielding an enriched high-purity solution of 2.54 g L−1 cerium. Lanthanum and yttrium were recovered from the Ce-depleted raffinate by co-precipitation with oxalic acid, forming a mixture of C6La2O12.10H2O and C6Y2O12.4H2O. The study approaches a cleaner production of high-purity REEs for clean energy programs under Sustainable Development Goals.

The Effect of Alkali Roasting Pretreatment on Nickel Extraction from Limonite Ore by Using Dissolved SO2-Air

Extraction of limonite ore using dissolved SO2–air is an alternative hydrometallurgical method for nickel recovery. This process is carried out at atmospheric pressure and is shown to have good selectivity of nickel over iron, but with a low recovery yield. The literature refers to the application of alkali roasting as pretreatment in laterite ore leaching to increase nickel recovery. Thus, this study aims to apply the combination method of alkali roasting and leaching to extract nickel from limonite ore (1.33% Ni, 46.61% Fe) from the Southeast Sulawesi region. Three alkali compounds were included in the study (NaOH, Na2CO3 and Na2SO4). The batch-leaching process was carried out at pH 1 and 3 and temperatures of 55 and 80 °C for 180 min. The leach liquors were sampled at 15, 60, 90 and 120 min, and concentrations of the extracted metals were measured by Atomic Absorption Spectrometry (AAS). A mineralogy characterization of the raw ore and its residue after leaching was undertaken by using X-Ray Diffraction (XRD), while the thermal decomposition behavior of the ore was characterized by Thermogravimetry Analyzer (TGA)/Differential Scanning Calorimetry (DSC). The addition of Na2CO3, Na2SO4 and NaOH in the ore pretreatment increases nickel recovery from 14.80% without alkali roasting to 23.99%, 28.15% and 39.22%, respectively. The optimum extraction condition for nickel recovery is at pH 1 and a temperature of 80 °C. However, the highest Ni/Fe selectivity of 24,947 is obtained at pH 3 and a temperature of 80 °C, preceded by roasting in the absence of alkali. Compared to other hydrometallurgical processes, the process studied in this work exhibits lower recovery, but provides an alternative to extract nickel from low-grade limonite ore.

Radioactivity and mineralogical studies with an emphasis on recovery of thorium from mylonite rocks in the Abu Rusheid area, South Eastern Desert, Egypt

The Abu Rusheid area, part of the Arabian-Nubian Shield, lies about 90 km southwest of Marsa Alam City. The cataclastic rocks in this area include protomylonites, mylonites, ultramylonites, and silicified ultramylonites. Mylonite rocks make up about 65.0% by volume of the cataclastic rocks, and many of them have high concentrations of visible orange-colored thorium and yellow-colored uranium minerals. XRD and SEM analyses revealed the presence of thorite, kasolite, monazite, zircon, xenotime, hematite, and ilmenite. The mylonite rock studied here contained high concentrations of thorium: > 1% in some cases, and an average of 3962 ppm. A sample assaying 10,000 mg/L was leached at 50 °C for 1.5 h by 6 M HCl using a S/L ratio of 1/6, which dissolved about 95% of the thorium. The thorium was precipitated from the leach liquor by oxalic acid. Highly pure thorium oxide was then prepared and identified via qualitative (SEM/EDX) and quantitative (ICP-OES) analyses.

An Investigation into the Extraction Behavior of Copper from Sulfate Leach Liquor Using Acorga M5640 Extractant: Mechanism, Equilibrium, and Thermodynamics

The extraction of copper from sulfate leaching liquor was investigated using Acorga M5640 commercial extractant in kerosene diluent. The effects of important factors such as the extractant concentration, contact time, liquor pH, and temperature were examined on the separation factor Cu/Fe and the extraction efficiency of copper. The results indicated that nearly 92.32% copper was extracted from the leach liquor at 10% (v/v) Acorga M5640 concentration, 1:1 aqueous/organic phase ratio, 20 min contact time, pH 2.5, and 25 °C temperature. Also, the further increase in the values determined had no positive impact on the extraction of copper. Under these conditions, the separation factor Cu/Fe was obtained about 29.64. The extraction mechanism was studied applying the plot of log(D) – 2 pH against log [extractant] and it proved the association of two moles of Acorga M5640 for the extraction of one mole of copper. McCabe–Thiele diagram was constructed and a single-step extraction process was found to obtain the maximum extraction. Additionally, thermodynamic data were investigated and the extraction process was endothermic and non-spontaneous at 298 K (ΔH = 27.05 and ΔG = 15.52 kJ mol−1). Moreover, the loaded organic phase was stripped and 80.94% copper was recovered after 20 min using 10% sulfuric acid concentration.

Selective acid leaching of connector pins removed from waste central processing units with focus on gold recovery

This paper evaluates the recovery of gold through selective acid leaching of connector pins removed from waste central processing units (WCPU) of computers. The exclusive treatment of this electronic component appears as a promising alternative since it minimizes gold losses. The experimental approach consisted by leaching steps of different types of pins aiming at dissolving all the metals that make up them, except the precious one. Analysis of the pins (2–6 mm length and ~ 1 mm diameter) by scanning electron microscopy coupled to energy dispersive spectrometry indicated (i) substrates made by copper, iron‑nickel alloy or iron‑nickel‑cobalt alloy; (ii) thin external gold coating; and (iii) thin nickel or nickel‑cobalt sub-coatings. Chemical analysis by inductively coupled plasma optical emission spectrometry revealed that the prepared sample consists mainly by copper (43 wt%), iron (26 wt%) and nickel (20 wt%); attractive contents of cobalt (4.6 wt%) and gold (1.4 wt%) were also determined. Due to the complex geometry of the pins and distinct reactivity of the different alloys, leaching stages were conducted in H2SO4-H2O2 media (1st step), HCl-H2O2 media (2nd step) and HCl media (3rd step). Under promising leaching conditions at 25 °C, all the metals were dissolved and then fine particles and thin foils of high purity gold (~99%) were recovered. Gold losses around 2–4 wt% were observed, mainly as result of its oxidation in the HCl-H2O2 media. The approach used in this work for treating the secondary source allowed the recovery of the precious metal with minimal losses, and it has potential to recover other valuable metals from the leach liquors.

Separation of cobalt and lithium from spent lithium-ion battery leach liquors by ionic liquid extraction using Cyphos IL-101

The eventual fading and scrapping of lithium (Li)-ion batteries will result in the formation of a huge urban mine of spent batteries. Hydrometallurgy and further extraction/separation are effective ways in which to recover the valuable metal ions from such batteries. Herein, extraction and recovery of spent lithium cobalt oxide (LiCoO2) battery leach liquors using Cyphos IL-101 is proposed. Based on the optimization of extraction conditions, continuous circulating extraction, and thermodynamic analysis, Cyphos IL-101 effectively extracted cobalt (Co) ions with excellent stability and high extraction capacity (35.25 g/L). The separation factor of Co and Li (βCo/Li) was 102.11 in 0.5 M HCl at 60 °C under the optimal operation condition. Moreover, an increase in temperature increased Co extraction in the spent Li-ion battery leach liquors. The extractants before and after use were characterized using ultraviolet-visible (UV–Vis) absorption spectra, Fourier transform infrared (FTIR) spectra, and nuclear magnetic resonance (NMR) spectra. The extraction mechanism of cobalt ions by Cyphos IL-101 was suggested as that cobalt transferred to the organic phase from the aqueous phase in the form of [CoCl4]2−. 90.5% of Co and 86.2% of Li were recovered by precipitation after extraction separation in the forms of cobalt oxalate and lithium carbonate with the purity 87.4% and 74.2%. This suggested method might provide a promising strategy for recycling spent lithium cobalt oxide batteries.

Recovery of Samarium and Cobalt from Sm–Co Magnet Waste Using a Phosphonium Ionic Liquid Cyphos IL 104

Due to outstanding thermal stability and corrosion resistance, samarium–cobalt alloy is being used in rare earth magnet. So, this permanent magnet waste can be a potential source for rare earth metal samarium and critical metal cobalt. Recovery and reuse of Sm, Co from Sm–Co magnet waste can reduce the burden on primary resources and environmental issues related to primary resources. In this work, a hydrometallurgical method for recovery of samarium and cobalt from Sm–Co magnet waste has been developed using an ionic liquid Cyphos IL 104. The optimum separation condition for cobalt and samarium has been determined by varying the concentration of HCl, NaCl, and mixture of HCl and NaCl in the leach liquor. The McCabe–Thiele plot is constructed using 0.075 mol/L Cyphos IL 104. Three counter-current stages at A:O ratio of 1:1.5 are needed for complete extraction of cobalt into the ionic liquid phase leaving samarium in the raffinate. Precipitation stripping of cobalt from the loaded organic is carried out using oxalic acid and the cobalt oxalate, thus formed, is calcined at 450 °C to form Co3O4. The raffinate obtained from the counter-current extraction contains 2.02 g/L Sm and 0.02 mg/L Co. Sm2O3 from raffinate is produced by precipitation of samarium with oxalic acid followed by calcination of samarium oxalate at 800 °C. Both the oxide products are confirmed by XRD.

Solvent extraction of tungsten and rare earth with tertiary amine N235 from H2SO4–H3PO4 mixed acid leaching liquor of scheelite

Scheelite is a calcium tungstate mineral that is an important ore of tungsten. It also contains small amounts of rare-earth metals, that enter the leaching liquor along with tungsten (W) after decomposition of the ore with H2SO4–H3PO4 mixed acid. In this study, a novel process was developed to effectively extract and separate tungsten and light rare earth (LRE) from this leaching liquor. La will represent LREs including La, Ce and Nd to be extracted in this research. First, W and La were extracted from the leaching liquor using a tertiary amine (N235). The results showed that 99.8% of W and 99.5% of La were extracted in the form of La-phosphotungstic acid complex anions under the conditions of N235 10% v/v, isooctanol 4% v/v, organic/aqueous (O/A) ratio of 1:1, pH of 2.6, temperature of 30 °C, and contact time of 10 min. Then, W and La were separated using a stripping process with H2SO4 solution. Nearly 99% of the La was recovered after single-stage stripping under the optimum stripping conditions of 1.5 mol/L H2SO4 solution, 10 min stripping time and O/A 1:1 at 45 °C, while almost no tungsten was stripped. After stripping with two cross-current extraction stages, the total La stripping reached 99.5%. Finally, the tungsten loaded on the organic phase was stripped with ammonia solution. Under the same optimum conditions, W stripping exceeded 99%. Extraction and two-stage stripping enabled effective recovery and separation of W and La from the leaching liquor. Based on above results, the process flowchart for extracting tungsten and light rare earth from scheelite was proposed. The proposed method is expected to promote further exploitation of valuable tungsten and rare-earth metals from scheelite ores.

A sustainable approach for advanced removal of iron from CFA sulfuric acid leach liquor by solvent extraction with P507

The recovery of valuable metals such as lithium (Li), aluminum (Al), gallium (Ga) from coal fly ash (CFA) has attracted considerable attention in recent years. As a main component in CFA, iron (Fe) has adverse effect on the selective recovery of the metals of interest. Removal of Fe is essential for the recovery of these metals. Herein, this work reported a sustainable approach for efficient and advanced removal of Fe(III) from sulfuric acid leach liquor of CFA using 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester (P507) as extractant. The results showed that more than 97.6% of Fe(III) can be effectively removed from the synthetic solution with a Fe/Al separation factor of 2218 under the optimum conditions. The UV–Vis, 1H NMR, IR spectroscopic results indicated that the extraction mechanism of Fe(III) with P507 was proved to be ion exchange. This technique was subsequently examined to remove Fe from the sulfuric acid leach liquor of CFA. The removal of Fe remained above 95.4% after 8 cycles of extraction-stripping process with H2O2 coupled with H2SO4 as stripping agent, implying the high efficiency and long-term reusability of P507. The study developed a sustainable way for advanced Fe removal from CFA sulfuric acid leach liquor, which could facilitate the recovery of valuable metals from CFA.

Sorption behavior of uranium from Sulfate media using purolite A400 as a strong base anion Exchange resin

ABSTRACT The sorption characteristics of uranium onto a strongly basic anion exchanger, Purolite A400, were investigated. The characteristics of Purolite A400, before and after sorption of uranium, were determined using scanning electron microscopy (SEM) and energetic-dispersive X-ray (EDX). The influence of the different parameters, which may affect the process of uranium sorption process, had been tracked. Specifically, impacts of initial uranium concentration, pH, contact time and temperature, on uranium removal, were studied. In addition, the sorption kinetics and equilibrium parameters were assessed. At the optimum conditions, the adsorption capacity of uranium was found to be nearly 117.6 mg/g. The kinetics and isotherm studies could show that the uranium sorption onto Purolite A400 is following the models of pseudo-second-order and Langmuir isotherm respectively. The thermodynamic parameters had indicated that the adsorption of uranium is an exothermic and spontaneous process. The detected results of uranium sorption could reveal that Purolite A400 is a suitable material for pre-concentration of uranium (VI) species from sulfate leach liquor.

Leaching tungsten and rare earth elements from scheelite through H2SO4–H3PO4 mixed acid decomposition

Rare earth elements (REEs) possess distinctivephysicalandchemical properties and are mainly extracted from several rare earth (RE) and ion-absorbing type RE ores. Traditional RE sources are unable to meet the global demand, making it necessary to explore other resources for the extraction of rare REEs. While this is possible, the recovery of REEs from the leach residue of such resources is uneconomical due to the enormous consumption of mineralacids. H2SO4–H3PO4 mixed acid decomposition is an effective method for leaching tungsten from scheelite and is evaluated in this study for its efficient extraction of tungsten and REEs from the same mineral. The scheelite used in this study contains REEs such as lanthanum, cerium, and neodymium. The properties of the REs in the leaching system are used to evaluate our method. Additionally, the effects of the reaction conditions on the leaching of tungsten and REEs are investigated. The results show that increasing the reaction temperature and H2SO4 concentration favors the leaching process. However, high H2SO4 concentration also promotesthetransformationof gypsumfromCaSO4·2H2OtoCaSO4, which affects the leaching process negatively. Further, excessive H3PO4 concentration increases the viscosity of the leaching solution. Increasing the liquid–solid ratio is beneficial for the dissolution of scheelite. The addition of an appropriate amount of CaSO4·2H2O promotes the leaching of tungsten and REEs but the addition of PEG-400 does not. In terms of the experimental conditions used in this study, the leaching efficiency of W, La, Ce, and Nd can reach approximately 100%, 96.6%, 66.6%, and 57.1%, respectively. After the circulationleaching process, the concentrations of tungsten and the REEs in the leach liquor are increased, which proves these materials can be efficiently leached from scheelite using the H2SO4–H3PO4 mixed acid.