Alkaline Leaching Solution Research Articles

Recovery of antimony from acidic and alkaline leaching solution of low-grade antimony ore by electrowinning process

With regards to the global continuous growth in consumption of base metals such as antimony (Sb), mining companies are currently looking to improve the productivity and extraction of Sb from low grade ore in order to economically process it. With this aim, in this study, an efficient protocol was developed to recover metallic Sb from the low grade Fe3Si2O5(OH)4. Sb3O6(OH). Sb2O3 ore of the Sefid-Abe mine in Zahedan province, Iran, utilizing selective acidic or alkaline leaching followed by Sb electrowinning. A process was developed to recover antimony from an Sb2O3 ore source, using selective acidic or alkaline leaching followed by antimony electrowinning. Sb could be leached with an efficiency of 85 % using 5 M NaOH and 87 % using 5 M hydrochloric acid/5 M sulfuric acid as the digestion solution at 80 °C for 8 h. The electrowinning process plays a significant role in recovery of highly pure Sb metal. Relying on this approach, a high deposition rate and current efficiency have been generated at a deposition potential 2 V at 30 °C (80 °C for alkaline medium). This was achieved by using anode electrode made of graphite and cathode electrode made of steel (two steel electrodes were used for alkaline electrowinning). Additionally, for acidic electrowinning, exploitation of Sb from the electrolyte solution was achieved by applying a current of 0.1 A, resulting in the highest yield equal to 89 %. On the other hand, for alkaline electrowinning, the highest efficiency corresponds to the current intensity of 0.5 A, which resulted in 97 % Sb extraction from the electrolyte solution. The XRD and ICP analysis performed for samples obtained from the electrolysis confirmed that the resulting antimony ingot has an exceptionally high purity.

Selective extraction of potassium from raw nepheline materials

The global aluminium industry is ever-changing and primarily relies on bauxite as its traditional source. However, due to the finite reserves of processed bauxite, alternative sources need to be found. Nephelines have emerged as potential alumina sources. In addition to alumina, nepheline processing can yield marketable products such as soda, potash, cement, and rare metals.The objective of this study was to identify the ideal conditions for selectively extracting potassium from alkaline leaching solutions while separating it from sodium alkali, aluminium, and silicon during the comprehensive processing of nepheline syenites from the Kubasadyr deposit. This approach involved a two-stage hydrothermal leaching process with an incremental calcium addition. Under optimal conditions, the potassium extraction in the alkaline solution reached 93.91 %. The resulting leaching solution served as a feedstock for potassium sulfate production through the crystallization method. The preliminary selective separation of potassium played a crucial role in mitigating its adverse effects on the extraction of aluminium, which is the primary processed product; thus, the economic viability of production was enhanced.

Mechanism for leaching of fluoride ions from carbon dross generated in high-temperature and low-lithium aluminum electrolytic systems

Carbon dross, a hazardous solid waste generated during aluminum electrolysis, contains large amounts of soluble fluoride ions for the main components of the electrolyte (such as Na3AlF6 and NaF). Response surface methodology (RSM) was used to investigate the mechanism for fluoride ion leaching from carbon dross via water leaching, acid leaching and alkali leaching, and the kinetic and thermodynamic principles of the leaching process were revealed. The RSM predicted the optimum conditions of water leaching, alkali leaching and acid leaching, and the conditions are as follows: temperature, 50 °C; shaking speed, 213 r·min-1; particle size, 0.075 mm; shaking speed, 194 r·min-1; liquid-solid ratio, 12.6 mg·L-1; sodium hydroxide concentration, 1.53 mol·L-1; liquid-solid ratio, 25.0 mg·L-1; sulfuric acid concentration, 2.00 mol·L-1; and temperature, 60 °C，and actual results which were almost consistent with the predicted results were gained. The fluoride ions in the alkaline and acid leaching solutions were mainly the dissociation products of fluorides such as Na3AlF6, Na5Al3F14 and CaF2, as indicated by thermodynamics calculations. In particular, the fluoride compounds dissolved in alkali solution were Na3AlF6, Na5Al3F14, AlF3, ZrF4, K3AlF6, while the acid solution could dissolve only Na3AlF6 and CaF2. The leaching kinetics experiments showed that the leaching rate fit the unreacted shrinking core model [1–2/3α-(1-α)2/3 =kt] and that the leaching process was controlled by internal diffusion. This study provides theoretical guidance for the removal of soluble fluoride ions from carbon dross and will also assist in the separation of electrolytes from carbon dross. Environmental ImplicationCarbon dross, a hazardous waste generated during the aluminum electrolysis production process, contains a large amount of soluble fluoride. Improper storage will lead the fluoride ions pollution in soil, surface water or groundwater under the direct contact between carbon dross and rainfall, snow or surface runoff. The influence of wind will cause carbon dross dust to pollute further areas. With the human body long-term contact with fluoride ion contaminated soil or water, human health will be seriously harmed.

Research on synergistic disposal and high-value utilization of secondary aluminum dross and spent carbon anode

Significant hazardous by-products emerged from the aluminum electrolysis realm, including secondary aluminum dross (SAD) and spent carbon anodes (SCA), with an anticipated annual yield surpassing 800,000 tons. These solid wastes showcase distinct complexation properties enriched with aluminum and fluoride. In this paper, aluminum or aluminum compounds will be extracted from SAD by sulfuric acid first to prepare an aluminum sulfate solution, which could be used to extract cryolite in SCA by Al-F complexation. A second-step alkaline leaching is aimed at enhancing Al and F dissolution from the first-step leaching acid-leaching residue of SAD and SCA. Under optimal conditions, the ensuing data reveals the soluble components in SAD reaching a leaching zenith of 90.3%, with the Al2O3 faction standing tall at 91.08%. Simultaneously, SCA's F leaching efficiency hits an impressive 98.20%, hence the carbon residue enriched with a purity of 75.5%. Finally, the acid-leaching solution and alkaline-leaching solution could be mixed to prepare flexibly and controllably aluminum fluoride or cryolite by adjusting the pH value and Al/F molar ratio of filtrate. This avant-garde approach illuminates a sustainable, ecologically harmonious route for synergistic disposal and high-value utilization of hazardous solid wastes from the aluminum smelting process.

Efficient recycling of Au(S2O3)23− from alkaline leach solutions using diphenylamine-rich Cu-MOF via local exotherms of DMF

Recycling gold from electronic waste is essential for environmental protection and resource sustainability. However, the design of adsorbents for effective recycling of Au(S2O3)23− from alkaline leach solutions is of great challenge and has not been investigated by using MOFs. Herein, we report the design and preparation of a Cu-MOF based on the structural characteristics of Au(S2O3)23−, which is constructed by 4,4′-di carboxy diphenylamine and Cu2-paddlewheel. Cu-ADA-53 demonstrates ultra-high adsorption capacity of 206.19 mg g-1, and the gold recovery in actual alkaline leach solutions is more than 99 %. X-ray photoelectron spectroscopy, thermodynamic study, and density functional theory calculations reveal the excellent performance of Cu-ADA-53 for Au(S2O3)23− results from the abundant Lewis base diphenylamine groups and open copper sites. More importantly, different from pore diffusion-promoting adsorption, the axial coordination DMF of Cu2-paddlewheel fell off into the water phase and then released heat locally at the adsorption sites. The local exotherms increase the temperature of the adsorption sites, further promoting the adsorption of Au(S2O3)23− by Cu-ADA-53. The local exotherms at adsorption sites provide a novel insight to efficiently recover Au(S2O3)23−.

Selective recovery of lithium from high-aluminum fly ash by using alkali-dissolution-assisted HBTA–TOPO synergistic extraction

The process of recycling lithium (Li) from common coal fly ash using strong acids can release substantial quantities of hazardous low-value residues, and the complex ionic composition of the acidic leaching solution hinders the selective extraction of Li. Accordingly, a green and economical approach was developed for the selective recovery of Li and extraction of valuable Al-rich residues from high-alumina fly ash (HAFA). In HAFA, the glass phase (the main Li-bearing phase) can be selectively dissolved in the mild NaOH solution at 95 °C for 2 h, and valuable mullite phase (the main Al-bearing phase) was retained in the leaching residue (Al2O3 content 49.00 wt%). The leaching efficiencies of Li, Si, and Al were 91.60%, 43.19%, and 3.56%, respectively. Among the primary metallic elements present in HAFA, only Li, Na, K, and Al in the glass phase were readily dissolved in the alkaline solution. Meanwhile, the synergistic extraction system of HBTA and TOPO has good complexing ability with Li + compared to Al3+, K+, and Na+. Therefore, Li+ can be selectively extracted from alkaline leaching solutions using HBTA and TOPO. The Li extraction efficiency reached 99.68% after three consecutive extractions under the optimal conditions. Finally, the used alkali leaching solution and synergistic extractants can be regenerated via Ca(OH)2 precipitation and HCl reverse extraction. These findings can provide new strategies for the green and overall utilization of HAFA.

Extraction of vanadium from vanadium slag by sodium roasting-ammonium sulfate leaching and removal of impurities from weakly alkaline leach solution

In the present work, a sustainable process for extraction of vanadium (V) from vanadium slag was proposed. The vanadium slag was first roasted by Na2CO3, then leached by (NH4)2SO4. The effect of roasting and leaching conditions on V and major impurities (P, Si) leaching have been systematically investigated. The results indicated that under the optimal conditions (roasting conditions: 18% Na2CO3, 800 °C, 2.5 h; leaching conditions: 30 g/L (NH4)2SO4, 60 °C, 3 mL/g, 0.5 h), the leaching rates of V, P and Si were 91.2%, 35.3% and 2.5%, respectively. Subsequently, Al2(SO4)3·18 H2O was used to simultaneously remove P and Si from the leach solution without pH adjustment, 92.3% and 86.6% of P and Si were removed with n(Al)/n(P + Si) of 2, temperature of 70 °C and 20 min. Ammonium metavanadate in the impurity removal solution can be directly separated by cooling crystallization without adjusting pH or adding ammonium salt. Meanwhile, the wastewater can be reused for leaching, which can effectively solve the difficult treatment of wastewater containing ammonia nitrogen. Leaching residue with Fe content of 45%, which can be applied to extract iron.

Preparation of environment-friendly SiO2 aerogel based on waste boron mud and its adsorption behavior for toluene.

To mitigate the environmental hazards of boron mud waste accumulation, we prepared environmental-friendly SiO2 aerogels by extracting them through alkaline leaching treatment and optimized the experimental conditions. The optimum process parameters for alkaline leaching solution NaOH concentration, leaching temperature, solid-to-liquid ratio, and leaching time were 2 mol/L, 95 °C, 1:4, and 3 h, respectively. In this work, cheap and non-toxic hydroxy silicone oil (PDMS-OH) and hydrogen-containing silicone oil (PMHS) were used as surface modifiers instead of toxic and expensive trimethylchlorosilane (TMCS) in the SiO2 aerogel modification process. The best performance under the optimum conditions was achieved with 60% PDMS-OH-modified SiO2 aerogel. Organic liquid spills, represented by toluene, pose a great danger to the environment and water bodies. We treated free toluene on the water surface with the aerogel mentioned above and its adsorption capacity was up to 2,655 mg/g. After the adsorption of toluene, the aerogels coalesced into agglomerates for subsequent collection and handling. Furthermore, after five repeated applications, the adsorption capacity remained at 91.43% of the initial application. Overall, this research provided an inexpensive and simple solution for the treatment of organic liquids in wastewater.

Rhenium-molybdenum separation in an alkaline leaching solution of a waste superalloy by N263 extraction

The recovery of rhenium from waste superalloys has always been a research focus. This study focuses on the performance of extraction of rhenium and its separation with molybdenum in alkali leaching solution of waste superalloy by methyl trioctylammonium chloride (N263). Minitab 17 Statistical Software (Trial version, Techmax, USA) was used to create a four-factor (N263 concentration, sec-octyl alcohols concentration, extraction time and organic to aqueous ratio (O/A ratio)) two-level orthogonal experiment table. According to the orthogonal experiment results of rhenium extraction and rhenium molybdenum separation coefficient, rhenium extraction model, rhenium molybdenum separation coefficient model and the corresponding regression equation were established. The rhenium extraction performance of N263 was investigated by the rhenium equilibrium curve and NaSCN concentration. It was found that the N263 concentration was the main factor affecting rhenium extraction, and the regression equation of rhenium extraction was Re extraction/% = 0.672 + 9.880 [N263]. Moreover, N263 concentration and O/A ratio are the main factors affecting rhenium molybdenum separation coefficient. The regression equation of rhenium molybdenum separation coefficient is βRe/Mo = 37.5 + 116.88* [N263] −24.3 *(O/A ratio) −20.62 [N263]*(O/A ratio). At the same concentration of the rhenium stripping agent, the rhenium stripping efficiency is perchloric acid > nitric acid > sodium thiocyanate. However, the strong oxidability of perchloric acid and nitric acid destroys the organic phase, making the organic phase produce visible destruction. The rhenium stripping was about 75.8 % with 40 times the theoretical amount of sodium thiocyanate. The results showed that Re was easy to be extracted in the alkali solution, but it was not easy to be stripped from the loaded N263 phase.

Study on Extraction Separation of Thioarsenite Acid in Alkaline Solution by -Type Tri-n-Octylmethyl-Ammonium Chloride.

To overcome the problem of arsenic separation and enrichment from an alkaline leaching solution in arsenic-containing dust, a -type tri-n-octylmethyl-ammonium chloride (TOMAC) method for extracting thioarsenite is proposed in this paper. Considering an alkaline leaching solution as the research object, after vulcanization pretreatment, TOMAC transformation and organic phase saturated extraction capacity were measured, and the extraction mechanism was preliminarily studied. First, Cl−-type quaternary ammonium salt was effectively transformed to -type by treating organic phase with saturated NaHCO3five times. TOMAC was effectively transformed from to type by alkaline washing with 1.0 mol/l NaOH solution; this washing was repeated thrice. Thereafter, the effects of organic phase composition, phase ratio, extraction time, and temperature on the extraction and separation of arsenic were investigated. The results show that under the conditions of 30% -type TOMAC + 15% sec-octanol + 55% sulfonated kerosene, VO/VA = 1/1, and 5 min extraction at room temperature, the single-stage extraction rate of AsIII is 85.2%. The AsIII concentration in raffinate can be reduced to less than 1.33 × 10−3 mol/l by four-stage countercurrent extraction, and the extraction rate of AsIII can exceed 98.4%.

A Novel Technology to Prepare Sodium Vanadate from V–Cr-Bearing Reducing Slag

A novel process for the extraction of vanadium and preparation of sodium vanadate from V–Cr-bearing reducing slag is proposed. This process involves several steps of washing, alkaline leaching, evaporation crystallization and cooling crystallization. The washing procedure ensures that chloride and sulfate ions can be mostly washed away from V–Cr-bearing reducing slag. Subsequently, vanadium can be extracted by alkaline leaching, the leaching rate of vanadium can reach 98%. During this process, vanadium enters the leaching solution, and chromium continues to accumulate in the leaching residue. Finally, most of Na2SO4 can be removed from the alkaline leaching solution by evaporation crystallization, then the Na3VO4·12H2O can be prepared by cooling crystallization. The filtrate of cooling crystallization can be recycled as the new leaching medium with some new alkali added, which greatly reduces the discharge of alkaline wastewater and makes the whole process more environmentally friendly.

Recycling of copper telluride from copper anode slime processing: Toward efficient recovery of tellurium and copper

Copper telluride is a sort of solid waste generated from copper anode slime processing, from which the recovery of copper and particularly the high-value element of tellurium always encounter technical difficulties due to the complex physicochemical properties of tellurium. In this study, an efficient and compact process for copper telluride recycling has been developed. Efficient separation of tellurium and copper from copper telluride was achieved through a pressure oxidizing alkaline leaching process under the optimal conditions of 5 mol/L for NaOH concentration, 5:1 for liquid to solid ratio, 150 °C for temperature, 0.7 Mpa for system pressure, and 2 h for reaction time, by which over 95% of tellurium was selectively dissolved in the solution and copper was enriched in the solid phase mainly in the form of copper oxides. Tellurium was subsequently recovered as TeO2 by neutralization of the alkaline leaching solution with sulfuric acid to pH 4.5, delivering a TeO2 precipitation efficiency of over 95%. In general, the masses of tellurium and copper balanced well in both the liquid and solid phases for each step, and the overall recovery efficiencies of Te and Cu reached as high as approximately 91% and 98%, respectively.

An environmental-friendly process for recovery of tellurium and copper from copper telluride

In the present study, a novel process for effective recoveries of Te and Cu from copper telluride from anode slime processing in copper smelters was proposed. The process consists of two hydrometallurgical steps of atmospheric alkaline leaching for Cu and Te separation, and TeO2 precipitation with H2SO4 for Te recovery. The effects of NaOH concentration, liquid to solid ratio, temperature, H2O2 to Cu2Te mole ratio, and reaction time on the dissolution behavior of tellurium were investigated. A Te leaching efficiency of about 91% was obtained under the optimal experimental conditions. The results of thermodynamic and kinetic analysis indicate that a lower temperature is favorable for the dissolution of Te, and mass transfer inside the solid particle is the rate-determining step. In addition, a mechanochemical-assisted leaching was conducted, by which Te leaching efficiency was enhanced to approximately 93% with ball milling at 180 rpm for 5 h. After Te leaching, H2SO4 was utilized to adjust the pH value of the Te-containing alkaline leach solution to 4.5 for TeO2 precipitation. The crystallization of TeO2 can be completed by reacting for 1 h and the overall tellurium recovery has reached nearly 90%.

Arsenic removal from arsenic-containing copper and cobalt slag using alkaline leaching technology and MgNH4AsO4 precipitation

In this study, a novel alkaline cycle leaching and MgNH4AsO4 precipitation technology was used to remove arsenic in arsenic–containing copper and cobalt slag. The phase analysis results indicated the arsenic species mostly existed in the original slag as Cu-As alloy. By adopting NaOH-H2O2 leaching method, the arsenic leaching reached as high as 94.5% under the optimum conditions. The XRD results indicated that copper in the Cu-As alloy was oxidized to be Cu(OH)2 in high concentration NaOH solution. And the optimum conditions were established as: NaOH concentration, 2 mol/L; leaching temperature, 25℃; H2O2 concentration, 5%; and liquid-solid ratio (L/S), 9:1. Thermodynamic analysis showed that it was feasible to selectively precipitate arsenate in alkaline leaching solution in the form of MgNH4AsO4. The XRD results confirmed that the arsenic in the filter residue mainly existed in the form of MgNH4AsO4·6H2O, which has good crystallinity, contributing to a better arsenic and alkali separation. Under the optimum conditions, the content of arsenic in the purified liquid and purified residue were 39 mg/L and 24.15%, respectively. Thus, the purified alkaline solution could be circulated for the leaching process, which is beneficial to the cost reduction and environment protection.

The Removal of Silicate(IV) by Adsorption onto Hydrocalumite from the Sodium Hydroxide Leaching Solution of Black Dross

Alkaline leaching of mechanically activated black dross resulted in an aluminate(III) solution with a small amount of silicate(IV). To obtain pure aluminate(III) solution, the removal of silicate(IV) from the alkaline leaching solution was investigated by adsorption with hydrocalumite (Ca2Al(OH)6Cl·2H2O). The hydrocalumite was synthesized by the coprecipitation method. The characterization of the synthesized hydrocalumite was analyzed via X-ray diffraction (XRD), scanning electron microscopy (SEM) images and Fourier-transform infrared spectroscopy (FTIR). In our experimental conditions, silicate(IV) was selectively adsorbed onto hydrocalumite over aluminate(III). The reaction time greatly affected the removal percentage of aluminate(III) owing to mass action effect. When the reaction time was longer than 2 h, no aluminate(III) was adsorbed onto hydrocalumite and thus it was possible to selectively remove silicate(IV). When the dosage of hydrocalumite was in excess, the removal percentage of silicate(IV) was rather reduced. Complete removal of silicate(IV) from the solution was accomplished in the reaction temperature between 50 and 70 °C. By selective adsorption of silicate(IV) from the solution at the optimum condition (30 g/L hydrocalumite, 50 °C, 400 rpm, and 2 h), an aluminate(III) solution with purity higher than 99.9% was obtained. The adsorption of silicate(IV) onto hydrocalumite followed the Freundlich isotherm.

A promising approach for directly extracting lithium from α-spodumene by alkaline digestion and precipitation as phosphate

Lithium in spodumene is tightly bound by coulombic forces inside the silicon-oxide and aluminum-oxide polyhedron structure. Therefore, the key to the extraction of lithium from spodumene is destroying the aluminosilicate structure. Traditionally, the naturally occurring α-spodumene needs to be roasted at 1000 °C to be converted to β-spodumene, and is then decomposed with acid. This method greatly increases energy consumption and environmental pollution. Therefore, developing a new technology to extract lithium from α-spodumene is an urgent matter. Aluminosilicates are common gangue minerals that coexist with bauxite. Diaspore is a type of refractory bauxite, and its decomposition requires a high temperature and alkalinity during the Bayer process. However, under such conditions, aluminosilicate gangues, such as kaolinite and illite, are preferentially decomposed, which leads to a heavy burden caused by the desilication products. The main structures of spodumene and illite contain both tetrahedral [SiO4] and octahedral [AlO6] geometries, in which the Si4+ in [SiO4] is partially replaced by Al3+, which causes the cation Li+ or K+ to enter the mineral structure and maintain the overall charge neutrality. Based on the similarities between spodumene and illite, it is speculated that the Bayer method may be used to treat α-spodumene. Experiments were carried out to verify this possibility. The results indicate that α-spodumene can indeed be directly decomposed by alkali. The addition of CaO, and a higher temperature and alkalinity, had a significant influence on the decomposition reaction. Under optimized conditions, the lithium leaching efficiency was 93%. Sodium phosphate was used to precipitate lithium from an alkaline leach solution considering the low solubility of Li3PO4, and the alkali-containing mother liquor was reused. During ten consecutive cycles in the proposed flowsheet, the α-spodumene decomposition was stable, and resulted in about 90% leaching efficiency of lithium. The closed loop extraction process can significantly reduce energy consumption, and provides a new process for the sustainable development of lithium resources.

Correlations between Milling Conditions and B<sub>2</sub>O<sub>3</sub> Activity of Mechanically Activated Boron Concentrate

Facing the current inefficiency of boron extraction from boron concentrate, this paper used mechanical activation as a pretreatment method to improve the extraction of B2O3 from boron concentrate, and researched the influences of milling conditions, including the ball-sample mass ratio, the diameter of milling ball, and filling ratio of chamber, on the B2O3 activity. The changes in the properties of the milled products were investigated to analyze the mechanism of activation. These three milling conditions all affected the B2O3 activity more or less, as evidenced by the changes of B2O3 activity with these milling factors. Under optimized milling conditions, the B2O3 activity could be enhanced from 67.52% of the non-activated boron concentrate to 85.01%. The results also showed that the specific surface area increased and the crystal structure was damaged, which accelerated the reaction between the alkaline leaching solution and boron concentrate samples and enhanced the B2O3 activity of the samples.

A fluoride-free liquid-liquid extraction process for the recovery and separation of niobium and tantalum from alkaline leach solutions

The production of high purity niobium (Nb) or tantalum (Ta) compounds still relies on processes that require highly acidic conditions and concentrated HF or NH4F solutions. Recent progress in the chemistry of group V elements, along with the evolution of the environmental concerns, call for the development of more sustainable industrial processes for these strategic elements. In sharp contrast with the historical fluoride approach chosen by the Nb and Ta refining industry, there has been a recent surge in the literature concerning the chemistry of these elements in diluted alkaline or oxalate media. We here propose an innovative liquid-liquid extraction process that is amenable to separate Nb and Ta from alkaline feed solutions, can be operated at low acidity and that does not require any fluoride-based chemical. The developed strategy is as follows: first, Nb and Ta are co-extracted from their alkaline solution (pH 12) using a quaternary ammonium salt (Aliquat® 336) dissolved in an organic diluent; then, Nb is selectively back-extracted in a mixture of diluted nitric and oxalic acids; and finally, Ta is stripped in a diluted nitric media. The process also affords purifying Nb from Ti, Fe and Na. Following an optimization at laboratory scale, the full process has been validated during continuous running tests (160 h at 1.4 L h−1) using industrial Nb-Ta caustic solutions coming from the development of the so-called Maboumine process. The developed process works at room temperature, low acidity ([H+] < 1 M), exhibits high recovery yields, and allows producing Nb2O5·nH2O with purity higher than 99.5%. The process also yields a tantalum product (Ta2O5·nH2O with >20 wt% Ta) even when starting from a diluted feed (39 ppm Ta). The proposed strategy paves the way for new Nb and Ta hydrometallurgical processes operating under mild chemical environments and represents a concrete alternative to the current fluoride-intensive industrial methods.

Recovery of vanadium from secondary tailing of iron ore by salt roasting-alkaline leaching and solvent extraction processes

In this paper, the recovery of vanadium from the secondary tail deposits of iron ore by salt roasting-alkaline leaching and solvent extraction has been investigated. To extract the vanadium, after the characterizations studies (XRF, SEM, and EPMA analyzes), preliminary leaching was performed using HCl to reduce lime. Subsequently, the solid that was gained from this stage was put in the oven for palletizing and salt roasting (sodium carbonate) at the temperature of 850°C. This sample was leached in the presence of sodium bicarbonate (NaHCO3), under certain conditions. Solvent extraction experiments were carried out on the solution of carbonated leaching under different conditions. According to SEM studies, the dominant are calcium, vanadium, and phosphorus in the sample. The percentage of Vanadium (V), according to EPMA results in the 20 selected points, was determined as 0.65%. XRF analysis showed that V2O5 and CaO contents of sample were 2.04% and 51.72%, respectively. Lime was reduced to 32.59% after initial leaching with HCl. The carbonate leaching resulted in a total recovery of 93.46% at 85 °C for 60 minutes and with 40 g /l of sodium bicarbonate. In addition, under the optimum conditions of pH=6, the ratio Vaq/Vorg=1 and the organic phase concentration of 0.5 M using the anionic Aliquat 336 extractant, Vanadium can be recovered effectively (extraction efficiency=97%) from alkaline leaching solution in the form of decavanadate or and the anions in the form of or in one stage.

First-principles calculations and experiments for Ce4+ effects on structure and chemical stabilities of Zr1-xCexSiO4

Using Ce atom as the surrogate for tetravalence actinides, the structure stability, chemical bonding and solubility of Zr1-xCexSiO4 have been investigated by combing first principles calculations and static leach test (MCC-1) experimentally to understand the effect of actinides on stability of ZrSiO4. The structural analysis indicates that the high symmetry of [CeO8] polyhedron could well maintain the structural stability of Zr1-xCexSiO4. However, due to the longer and weaker CeO bond than ZrO bond, LRCe is slightly higher than LRZr. Besides, the LRZr and LRCe are highest in acid leaching solution, followed in alkaline leaching solution and deionized water system. Worth noted, the negative solution energy of −22.254 eV indicates the pre-existing Zr vacancy is energetically favorite to accommodate actinides surrogate Ce impurities or quadrivalent actinides with slight structural deformation, showing the strong structural and chemical stabilities of ZrSiO4.