Extraction of vanadium from black shale using pressure acid leaching

The extraction of vanadium from black shale was attempted using pressure acid leaching. The effects of the several parameters which included reaction time, concentration of sulfuric acid, leaching temperature, liquid to solid ratio and concentration of additive (FeSO₄) upon leaching efficiency of vanadium were investigated and a two-step counter-current leaching approach was developed. The results showed that the leaching efficiency of vanadium in the two-step process could reach above 90%. Vanadium was effectively separated and enriched by solvent extraction after leachate pretreatments, including the reduction of Fe³⁺ and adjustment of pH value. The extraction and stripping yields of vanadium were both > 98%. Ammonia was added to a stripping liquor to precipitate vanadium and then the ammonium poly-vanadate produced was calcined at 550 °C for 3 h to produce the high purity V₂O₅ powder. The overall yield of vanadium through all process stages was about 85%.     

Carrier-mediated transport of uranium from phosphoric acid medium across TOPO/n-dodecane-supported liquid membrane

Present studies deals with the application of supported liquid membrane (SLM) technique for the separation of uranium (VI) from phosphoric acid medium. Tri-n-octyl phosphine oxide (TOPO)/n-dodecane is used as a carrier and ammonium carbonate as a receiving phase for the separation of uranium (VI) from the phosphoric acid medium. Throughout the study PTFE membranes are used as a support. The studies involve the investigation of process controlling parameters like feed acidity of phosphoric acid, carrier concentration and stripping agents. The effect of nitric acid and sodium nitrate in feed is also studied. It is found that there is negligible transport of uranium (VI) from pure phosphoric acid medium but it increases to very significant amount if 2 M nitric acid is added to feed phase. More than 90% uranium (VI) is recovered in 360 min using 0.5 M TOPO/n-dodecane as carrier and 1.89 M ammonium carbonate as stripping phase from the mixture of 0.001 M H₃PO₄ and 2 M of HNO₃ as a feed. The flux and permeability coefficient are found to be 9.21 × 10^− 6 mol/m² s and 18.26 × 10^− 5 m/s, respectively. Lower concentration of phosphoric acid with 2 M HNO₃ and higher concentration of carrier is found to be the most suitable condition for maximum transport of uranium (VI) from its low-level sources like commercial phosphoric acid.     

Extraction of Fe(III) from hydrochloric acid solutions using Amberlite XAD-7 resin impregnated with trioctylphosphine oxide (Cyanex 921)

Ferric ions were efficiently removed from HCl solutions using Amberlite XAD-7 resin impregnated with trioctylphosphine oxide (Cyanex 921). Iron was removed under the form HFeCl₄ through direct binding on the resin or by extraction with Cyanex 921 involving a solvation mechanism. High concentrations of HCl and intermediary extractant loadings were required for maximum sorption efficiency and rationale use of the extractant. At intermediary extractant loading (in the range 300–450 mg Cyanex 921 g^− 1) the maximum sorption capacity increased with extractant loading. Maximum sorption capacity slightly increased with temperature, the reaction is endothermic and the enthalpy change was found close to − 30.8 kJ mol^− 1. Sorption isotherms were fitted with the Langmuir equation and maximum sorption capacity reached values as high as 20–22 mg Fe g^− 1 in 3 M HCl solutions. Despite the good fit of experimental data with the pseudo second-order rate equation, sorption kinetics was controlled by the resistance to intraparticle diffusion. The intraparticle diffusion coefficient (D_e) varying in the range 1.2 × 10^− 11–4.7 × 10^− 10 m² min^− 1 was found to increase with metal concentration and with temperature, while varying the extractant loading it reached a maximum at a loading close to 453 mg Cyanex 921 g^− 1. The desorption of Fe(III) can be achieved using 0.1 M solutions of nitric acid, sulfuric acid, sodium sulfate and even water, maintaining high efficiencies for sorption and desorption for at least 5 cycles.     

A comparative study on extraction of Fe(III) from chloride leach liquor using TBP, Cyanex 921 and Cyanex 923

A comparative study on extraction of Fe(III) from the HCl leach liquor of low grade iron ore tailings has been carried out using Tri-n-butyl phosphate (TBP), Cyanex 921 and Cyanex 923 in distilled kerosene. The percentage extraction of iron increased with increasing HCl and extractant concentrations. The extracted species in each case was found to be HFeCl₄·S. The extraction isotherms for the above extractants indicated quantitative extraction of Fe(III) in 3-stages at O:A ratio of 3:2 with TBP, and in 2-stages at O:A ratio of 1:1 with Cyanex 921 and Cyanex 923. The stripping studies of the loaded organic phases were carried out with 0.4 M HCl. The stripping isotherms indicated 2-stages at O:A ratio of 5:2 for TBP, and 3-stages at O:A ratio of 2:3 for Cyanex 921 and Cyanex 923. From the extraction studies, the extraction efficiency of the extractants for Fe(III) was in the order TBP < Cyanex 921 < Cyanex 923. Although Cyanex 923 was found to be the best extractant, the percentage stripping of Fe from the loaded Cyanex 923 was the least. The stripping of Cyanex 923 was 94.9%, but with TBP and Cyanex 921, it was 99.8% and > 99.9%, respectively.     

Selective separation of copper and nickel by solvent extraction using LIX 984N

Selective separation of copper and nickel from ammoniacal/ammonium carbonate medium was carried out by using LIX 984N diluted with deodourised kerosene. The study of the influence of equilibration time, equilibrium pH, extractant concentration and selective stripping of copper and nickel has been optimized. It was found that both the metal extractions were unaffected by the changes in pH. Nickel extraction equilibrium was reached at a longer contact time than that for copper and nickel extraction depends greatly on the extractant concentration in the organic phase. Co-extraction, ammonia scrubbing and selective stripping of copper and nickel were performed for a solution containing 3 g dm^− 3 each of copper and nickel and 60 g dm^− 3 ammonium carbonate. The extraction and the percentage of stripping for nickel and copper were almost quantitative.     

Palladium(II) complexes adsorption from the chloride solutions with macrocomponent addition using strongly basic anion exchange resins, type 1

The use of the strongly basic anion exchange resins, type 1 such as Lewatit MP-500 and Lewatit MP-500A for palladium(II) complexes adsorption has been investigated. The adsorption process was carried out from the chloride solutions with macrocomponent (sodium chloride) addition (x M HCl–1.0 M NaCl; x M HCl–2.0 M NaCl) where the concentration of hydrochloric acid was constant and equal to x = 0; 0.1; 0.5; 1.0; and 2.0 M, respectively. The breakthrough curves of Pd(II) were determined and the sorption parameters (weight and bed distribution coefficients, working anion exchange capacity) were calculated. The pseudo-second kinetic order was applied in kinetic studies as well as to calculate the kinetic parameters. The values of the working anion exchange capacities (0.029 g/cm³; 0.028 g/cm³) for Lewatit MP-500 and Lewatit MP-500A (0.028 g/cm³; 0.027 g/cm³) in the 1.0 M NaCl and 0.1 M HCl–1.0 M NaCl solutions, respectively are really close and in other solutions under discussion Lewatit MP-500 possess slightly higher values of capacities, and therefore is insignificantly more efficient in the adsorption process of palladium(II) ions than Lewatit MP-500A. The equilibrium adsorption capacities changed in the range 8.84–9.99 and 8.40–9.38 mg/g for Lewatit MP-500 as well as 8.12–9.57 and 7.26–8.85 mg/g for Lewatit MP-500A in the chloride x M HCl-1.0 M NaCl and x M HCl-2.0 M NaCl solutions, respectively. The adsorption process proceeds according to the pseudo-second kinetic order.     

Uphill permeation model of gold(III) and its separation from base metals using thiourea derivatives as ionophores across a liquid membrane

This work deals with carrier-facilitated membrane transport of Au(III) from chloride media across a polymer-immobilised liquid membrane (PILM) using as organic reagents N-(thiocarbamoyl)benzamide derivatives and N-benzoylthiourea derivatives, denoted as 2a–c and 3a–f, respectively. Both the composition of the organic membrane solvent and the type of carrier have a marked effect on gold permeation. Recovery and permeability of gold using 2a–c and 3a–f across a PILM proceed in the following order: 3e≈3d≈3c?3f>3b≈3a≈2a≈2b≈2c. In view of the performance of these carriers, 3c was selected as a metal receptor for detailed studies of Au(III) in permeation. A model is presented for the permeation of Au(III) (61 μM) in 0.5 M Cl⁻ at pH 2.5 using 3c as a membrane carrier. The mathematical equations describing the rate of permeation are derived to correlate the membrane permeability coefficient with diffusional and equilibrium parameters. The mass transfer coefficient was calculated from the described model as 1.1×10⁻⁵ m s⁻¹, and the thickness of the aqueous boundary was later calculated to be 65 μm. Several polymeric supports were tested for impregnation of the organic extractant, and Durapore (Millipore) afforded the maximum flux for Au(III), yielding a value of 1.1×10⁻¹⁴ mol m⁻² s. The relationship between flux and support characteristics is derived and a mathematical equation is presented. Of the several diluents used, cumene had the most satisfactory performance in terms of PILM stability and metal transport. Of the different reagents used, 0.5 M sodium thiocyanate in 0.5 M NaCl at pH 2.5 served most efficiently as the stripping agent. More than 80% of the Au(III) could be readily separated using 3c in the presence of various metals such as Cu(II), Fe(III) and Zn(II).     

Extraction of cadmium (II) from phosphoric acid media using the di(2-ethylhexyl)dithiophosphoric acid (D2EHDTPA): Feasibility of a continuous extraction-stripping process

The extraction of cadmium from phosphoric media has been studied. The D₂EHDTPA was used as extractant and dodecane as diluent. No third phase was observed in the investigated conditions.A continuous micro-pilot scale mixer-settler was successfully tested for both extraction and stripping. More than 99% extraction rate was obtained in steady-state conditions with a flow rate ratio Aqueous/Organic equal to 1.1. Continuous stripping was performed using HCl 4 M. More than 96% of the cadmium was stripped in one continuous mixer-settler stage for flow rate ratio equal to 0.7. Results were in good agreement with the predicted values based on the McCabe–Thiele method. Experimental mixer-settler stages behave as ideal ones (Murphree efficiency > 98%).An optimal flow sheet is proposed to purify the Wet Phosphoric Acid (WPA) and to recover a relatively concentrated cadmium solution (1 g L^? 1). Two ideal stages operating at phase ratio A/S equal to 5/1 are required for the extraction step leading to a very depleted raffinate (< 0.2 µg L^? 1). For the stripping step, six stages are required (S/A = 5/1). The recovered organic phase contains less than 2 µg L^? 1 and could be recycled in the extraction step.     

Production of high purity cobalt oxalate from spent ammonia cracker catalyst

A process was developed to produce > 99.9% pure cobalt oxalate from spent ammonia cracker catalyst pellets containing ∼ 20% Co generated at heavy water plants. A pilot plant for producing kilograms of cobalt oxalate of required purity has been set up. The process consists of leaching, oxidation, solvent extraction, ion exchange and oxalate precipitation. The major impurities present in the leach liquor after nitric acid oxidation were Fe³⁺, Al³⁺ and Ni²⁺. Fe³⁺ and Al³⁺ were separated from the leach liquor by using a solvent extraction (SX) process employing a mixed extractant system consisting of D2EHPA and TBP. Ni²⁺ was separated by ion exchange (IX) employing Dowex M4195 resin.     

Extraction of gold(III) in hydrochloric acid solution using monoamide compounds

The extraction and separation properties of Au(III) using two monoamide compounds, N,N-di-n-octylacetamide (DOAA) and N,N-di-n-octyllauramide (DOLA), which have different side chain lengths attached to the carbonyl carbon (CH₃ for DOAA and n-C₁₁H₂₃ for DOLA), were investigated. The solvent extraction of some precious and base metals (Au(III), Pd(II), Pt(IV), Rh(III), Fe(III), Cu(II), Ni(II) and Zn(II)) in HCl solutions was carried out using DOAA and DOLA diluted with n-dodecane and 2-ethylhexanol. A good selectivity for Au(III) extraction with 0.5 M extractant is obtained at lower HCl concentrations (< 3.0 M) in both systems. The extractability of Au(III) with DOAA is greater than that with DOLA. In the 0.5 M DOAA–3.0 M HCl system, a third phase is formed when the Au(III) concentration in the initial aqueous phase is over 39 g/L. In contrast, third phase formation is not found in the 0.5 M DOLA–3.0 M HCl system, and its loading capacity of Au(III) is about 79 g/L. The Au(III) extracted in the organic phase is effectively back-extracted by 1.0 M thiourea in 1.0 M HCl solution in both systems, while some thiourea is precipitated using the organic phase containing 20 g/L of Au(III). The back extraction of Au(III) using water is poor in the DOAA system, but possible in the DOLA system.     

Selective separation of nickel from cobalt in ammoniacal solutions by emulsion type liquid membranes using 8-hydroxyquinoline (8-HQ) as mobile carrier

The selective separation and concentration of nickel from ammoniacal solutions containing nickel and cobalt by an emulsion liquid membrane (ELM) technique using 8-hydroxyquinoline (8-HQ) as carrier has been examined. The emulsion liquid membrane consists of a diluent (kerosene), a surfactant (ECA 4360J), a carrier (8-HQ), and a stripping solution (0.025 M EDTA solution, buffered at pH 4.0). Cobalt (II) in the 6 M ammonia feed solution was oxidised to cobalt (III) by adding H₂O₂ and the pH adjusted to 10 with hydrochloric acid (HCl). The important variables were found to be membrane composition, ammonia concentration, diluent type, surfactant concentration, extractant concentration, EDTA concentration in the stripping solution, pH of the feed and the stripping solutions, phase ratio, and treatment ratio. It was possible to selectively extract 96.5 to 99.0% of nickel from a mixture of nickel and cobalt.     

Properties of electrolyte solutions relevant to high concentration chloride leaching. III. Solubility of pertinent solids and iron(III)/iron(II) redox potential measured in concentrated magnesium chloride solutions

Results of solubility measurements of nickel chloride, manganese chloride, iron(II) chloride, hematite and akaganeite in aqueous solutions of MgCl₂ (0.5–3.5 mol L^− 1) at temperatures of 60 and 90 °C are reported. Solubilities of metal(II) chlorides decrease almost linearly with MgCl₂ concentration due to the common ion effect. Nickel chloride and iron(II) chloride solubilities are very similar, while manganese chloride is about 30% more soluble.Hematite is more stable (i.e. less soluble) than akaganeite under all conditions investigated in this study, while ferrihydrite is considerably less stable. In other words, there is no change in the relative stabilities of these phases effected by the presence of high magnesium chloride concentrations. The solubility of all of these phases decreases with temperature and, for each temperature, the solubility constants increase linearly with the MgCl₂ concentration. The present results allow the prediction of the iron concentration as a function of the H⁺ and MgCl₂ molality at equilibrium with hematite or akaganeite.The Fe(III)/Fe(II) redox behaviour has been characterized in concentrated aqueous solutions of MgCl₂ (1.5–3.5 mol L^− 1) at a temperature of 25 °C. Standard redox potentials are ca. 100 mV lower than at infinite dilution and change linearly by only 13 mV in the range 2–4 mol L^− 1 MgCl₂.     

Separation of aluminum from rare earth by solvent extraction with4-octyloxybenzoic acid

Leaching method is usually used to extract rare earth(RE) elements from ion adsorbed RE ores.In the leaching process,some impurities such as aluminum(Al) enter the leaching solution.The separation of Al from RE by carboxylic acid extractant 4-octyloxybenzoic acid(POOA) was studied in this article.By changing the pH value,temperature,solvent,saponification degree and other parameters,the extraction and separation performance of POOA in chloride system was systematically studied.Through specific e...     

Neodymium naphthenate-loaded organic phase stripping using sodium oxalate

Neodymium naphthenate-loaded organic phase stripping using sodium oxalate solution was studied to explore the feasibility of synchronous rare earth-loaded organic phase stripping, rare earth precipitation, and blank organic phase saponification. Experimental results show that loaded organic phase stripping, rare earth precipitation, and blank organic phase saponification can be realized simultaneously. When using 20% excess of sodium oxalate over the stoichiometry with the volume ratio of organic phase to aqueous phase of 1:1 at 25 °C for 40 min, the single stage stripping rate and saponification value are about 40% and 0.29 mol/L, respectively. After 16 stages of countercurrent continuous stripping, the stripping rate of neodymium can reach 99%, the saponification value is 0.42 mol/L, the Nd³⁺ concentration in saponified organic phase is less than 0.0020 mol/L, and the main phase in precipitation is Nd₂(C₂O₄)₃·10H₂O. Afterwards, this saponified organic phase can be used in the extraction of NdCl₃ solution, and then the loaded organic phases (neodymium naphthenate) with 0.16 mol/L Nd³⁺ can be retrieved. The morphology, particle size distribution, and composition of the Nd₂(C₂O₄)₃·10H₂O products are similar to those of the current direct precipitation products. The neodymium oxide prepared by continuous calcination of neodymium oxalate meets the national standard of China (GB/T 5240?2015). These results prove the feasibility of stripping neodymium naphthenate-loaded organic phase by using sodium oxalate solution. Sodium oxalate can serve as a stripping agent, a saponifier, and a precipitator, thereby simplifying rare earth extraction and separation. This study provides theoretical and technical support for the development of a novel method for rare earth extraction and separation.     

Adsorption of rhenium(VII) on 4-amino-1,2,4-triazole resin

The adsorption properties of 4-amino-1,2,4-triazole resin (4-ATR) for Re(VII) were investigated by static and dynamic adsorption–desorption measurements with ultraviolet–visible spectroscopy. The influence of conditions such as temperature, initial solution pH and contact time on the adsorption curve was studied. It was found that the 4-amino-1,2,4-triazole resin was suitable for adsorption of Re(VII). The saturated adsorption capacity was 354 mg·g^− 1resin at pH 2.6 in HAc–NaAc medium at 298 K. The adsorption rate constant was k₂₉₈ = 8.2 × 10^− 5 s^− 1. The adsorption behavior of 4-ATR for Re(VII) obeyed the Freundlich empirical equation; whilst changes in adsorption with temperature gave an enthalpy change ΔH  = − 11.8 kJ·mol^− 1. The molar ratio of the functional group of 4-ATR to Re(VII) was about 2:1. Re(VII) adsorbed on 4-ATR was eluted by 1.0 ~ 5.0 mol·L^− 1 HCl with 100% quantitative elution in 4.0 mol·L^− 1 HCl solution. The resin can be regenerated and reused without apparent decrease in adsorption capacity.     

Solvent extraction of Ce(III) and Pr(III) with P507 using SiC foam as a static mixer

Extraction reactor is a major research area of interest within the field of rare earths extraction and separation. SiC foam offers excellent material characteristics as well as three-dimensional (3-D) reticulated structure; however, very little research has been carried out on its application in extraction reactor so far. In this work, a static mixer reactor based on SiC foam was designed and demonstrated to extract and separate Ce(III) and Pr(III) from nitric acid media by using 2-ethylhexyl phosphoric acid mono-2-ethylhexyl ester (P507) as extractant. The structure–performance relationship between SiC foam and extraction performance was studied by experiment combined with computational fluid dynamics (CFD) simulation. The experiment data are in good agreement with the simulation results. Contrast experiment by using a Kenics mixer was carried out, and SiC foam shows better extraction and mass transfer performance. Using the optimal structural SiC foam (pore size D = 2.3 mm, open porosity ε = 85%, foam length L = 80 mm), high extraction efficiency η (Pr(III): 94.6%, Ce(III): 88.5%) and separation factor β (2.27) between Ce(III) and Pr(III) is achieved at a high total throughput of 200 mL/min. Besides, overall volumetric mass transfer coefficient K_La of Pr(III) and Ce(III) are 0.519 and 0.378 s^?1 at the residence time τ of 3.6 s, respectively, which reach the high level of microchannel reactors and are better than conventional extractors and other static mixers. SiC foam is found to be applicable as a static mixer for efficient and high-throughput extraction and separation of rare earths.     

Synergistic extraction of rare earth by mixtures of 2-ethylhexyl phosphoric acid mono-2-ethylhexyl ester and di-（2-ethylhexyl） phosphoric acid from sulfuric acid medium

The extraction of Nd^3＋ and Sm^3＋, including the extraction and stripping capability as well as the separation effect of Nd^3＋ or Sm^3＋, from a sulfuric acid medium, by mixtures of di-（2-ethylhexyl） phosphoric acid （HDEHP, H2A2（0）） and 2-ethylhexyl phosphoric acid mono-2-ethylhexyl ester （HEH/EHP, H2L2（0）） were studied. The distribution ratios and synergistic coefficients of Nd^3＋ and Sm^3＋ in different acidities were also determined. A synergistic extractive effect was found when HDEHP and HEH/EHP were used as mixed extractants for Sm^3＋ or Nd^3＋. The chemical compositions of the extracted complex were determined as Nd.（HA2）2-HL2 and Sm.（HA2）2-HL2. The extraction equilibrium constants, enthalpy change, and entropy change of the extraction reaction were also determined.     

Solvent extraction of cadmium from sulfate solution with di-(2-ethylhexyl) phosphoric acid diluted in kerosene

In the present paper, solvent extraction process has been used for extraction of cadmium from sulfate solution using di-(2-ethylhexyl) phosphoric acid (D2EHPA) with 1% isodecanol in kerosene diluent expected from industrial effluents or leaching of ores/secondary materials. Different process parameters such as pH, contact time, extractant concentration, O/A ratio etc. were investigated. Results demonstrated that quantitative extraction of cadmium was feasible from 4.45 mM cadmium feed solution in single stage at equilibrium pH 4.5, time 2 min and O/A ratio 1:1 with 0.15 mM D2EHPA. The extraction mechanism of cadmium from sulfate solution by D2EHPA in kerosene could be represented at equilibrium by Cd²⁺ + 3/2 (H₂A₂)_org ⇔ CdA₂(HA)_org + 2H⁺. The loading capacity of 0.15 mM D2EHPA in sulfate solution was determined to be ∼ 8.9 mM cadmium. The loaded cadmium was effectively stripped using 180 g/L sulfuric acid. The metal or salt could be produced by electrolysis or crystallization from the stripped solution.     

The application of Primene 81R for the purification of concentrated aluminium sulphate solutions from leaching of clay minerals

A method of separating iron(III) and aluminium by solvent extraction treatment of sulphate leach liquors produced by the acid leaching of domestic non-bauxitic ores is reported. The principal impurity to be eliminated was iron(III). The amine Primene 81R was investigated for both synthetic solutions and leach liquors. The effect of aluminium, sulphuric acid, amine concentrations and other variables upon extraction was studied. The use of sulphuric acid and basic solutions as stripping agents was also studied. Extraction and stripping isotherms were used as a basis for laboratory-scale tests in single-stage and multi-stage studies. The iron concentration in the aqueous solution can be decreased from 2 kg/m³ to 0.015 kg/m³.     

The solvent extraction separation of bismuth and molybdenum from a low grade bismuth glance flotation concentrate

The main purpose of this study was to characterize and to extract bismuth and molybdenum from a low grade bismuth glance concentrate. Selective leaching of bismuth could be obtained at a temperature range 60 to 85 °C for a leaching duration of 2 h with hydrochloric acid concentration of 150 gpl, lignin calcium concentration of 0.02 gpl and using a solid–liquid ratio 1/4 g/cc. Treatment of leach liquor for the solvent extraction of bismuth with N235 showed that 8.0 × 10^− 2 M N235 in kerosene, a 3 min period of equilibration and a pH 0.2 were sufficient for the extraction of Bi(III). This bismuth-loaded organic phase was almost completely stripped using 0.5 M EDTA solution. Treatment of leached residue was dealt with by roasting in the presence of slaked lime, and followed by hydrometallurgical treatment of the roasted products. In the lime roasting process, molybdenum recoveries of around 99% were achieved when an excess of 50% lime over stoichiometric requirement was roasted at 700 °C for 2 h and the calcine was leached with 4 M HCl, at 70–80 °C for 2 h. Molybdenum then was effectively extracted from the leached residual solution with N235. An optimum pH of 0.5 was determined for molybdenum extraction. From loaded solvent, this metal was easily stripped with ammonia solutions to give a pregnant solution suitable for final recovery of metal by salt precipitation. Under the optimized conditions, the ultimate recovery rate of bismuth and molybdenum was more than 99% and 98% respectively.