Uranium recovery from strong acidic solutions by solvent extraction with Cyanex 923 and a modifier

Uranium stripping with strong acid solution is always highly desired due to its simple operation and less pollution. However, intensive acid neutralisation for uranium precipitation in the subsequent step limited its application. A new solvent extraction process has been developed to transfer uranium from strong to weak sulphuric acid solutions suitable for uranium precipitation without intensive neutralisation. An organic system consisting of 10% Cyanex 923 and 10% isodecanol as the modifier in ShellSol D70 was optimised for the process. It was found that uranium was extracted efficiently from 4 to 6 M H₂SO₄ solutions with the organic system, and it could be efficiently stripped with 0.2–0.5 M H₂SO₄ solutions. Both extraction and stripping kinetics of uranium were very fast, reaching the equilibrium within 0.5 min. Temperature between 30 and 60 °C has slight effect on uranium extraction and stripping. Four theoretical stages could effectively extract more than 98% uranium from a solution containing 17.5 g/L U and 6.0 M H₂SO₄ at an A/O ratio of 1:1.5, and it could generate a loaded organic solution containing about 12 g/L U. More than 99% U could be stripped from the loaded organic solution containing 14.6 g/L U with 0.5 M H₂SO₄ using five stages at an A/O ratio of 1:3. As a result, the loaded strip liquor containing more than 40 g/L U would be obtained which is suitable for uranium recovery by precipitation using hydrogen peroxide. A conceptual process has been proposed for uranium transfer from strong to weak sulphuric acid solutions for its recovery.     

Preparation of layered nickel aluminium double hydroxide from waste solution of nickel

The separation of nickel has been carried out from a waste solution containing 3.18 g/L Ni with other impurities such as Fe, Zn, Cu and As. Iron was removed by precipitation and Cu and Zn were removed by solvent extraction using LIX 622N and NaTOPS-99, respectively. After removal of all these impurities nickel was extracted by 1.5 M NaTOPS-99 in two counter-current stages at A:O ratio of 3:1 and the loaded organic was stripped with 30 g/L H₂SO₄ at phase ratio of unity. The strip solution of nickel was treated with Al₂(NO)₃ · 9H₂O for co-precipitation by increasing the pH of solution with 1 M NaOH up to 10. The Ni–Al layered double hydroxide was confirmed through XRD characterization.     

Recovery of boron from borax sludge of boron industry

A two-step process for boron recovery from borax sludge is proposed in the present work. The borax sludge was leached with sulphuric acid solution. Then, for the removal of alkaline species from the leachate, calcium and magnesium were precipitated by adjusting the pH of leachate using 1.5 M NaOH and 1.5 M HCl solutions. The effects of pH, temperature, concentration and time on the precipitation process were investigated. It was determined that the calcium and magnesium concentrations in the leachate were reduced from 121 mg/L to 2.6 mg/L and from 145 mg/L to 4.0 mg/L, respectively, at a pH value of 12, a temperature of 70 °C, an initial boron concentration of 500 mg/L and a precipitation time of 3 h. Under these optimum conditions, it was observed that the boron concentration in the solution decreased very slightly. In this process, the alkaline species were successfully separated from the boron.Finally, borax pentahydrate (Na₂B₄O₇·5H₂O) was produced by the evaporation of the final solution obtained after precipitation process.     

Research on the recycling of valuable metals in spent Al2O3-based catalyst

A new technology was developed to recover multiple valuable elements in the spent Al₂O₃-based catalyst by X-ray phase analysis and exploratory experiments. The experiment results showed: In the condition of roasting temperature of 750 °C and roasting time of 30 min, mol ratio of Na₂O: Al₂O₃ 1.2, the leaching rate of alumina, vanadium and molybdenum in the spent catalyst is 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by barium hydroxide and barium aluminate, the precipitation rate of vanadium and molybdenum is 94.8% and 92.6%. Al(OH)₃ is prepared from sodium aluminate solution with carbonation decomposition process, and the purity of Al₂O₃ is 99.9% after calcinations, the recovery of alumina can reach 90.6% in the whole process. The Ni–Co concentrate was leached by sulfuric acid, a nickel recovery of 98.2% and over 98.5% cobalt recovery was obtained respectively under the experimental condition of 30% (w/w) H₂SO₄, 80 °C, reaction time 4 h, liquid:solid ratio (8:1) by weight, stirring rate of 800 rpm.     

Oxidation of cyanide in effluents by Caro’s Acid

Caro’s Acid (peroxymonosulphuric acid: H₂SO₅) is a powerful liquid oxidant made from hydrogen peroxide that has been adopted for the detoxification of effluents containing cyanides in gold extraction plants in recent years.The present work reports the findings of a study on the kinetics of aqueous cyanide oxidation with Caro’s Acid. Experiments were conducted in batch mode using synthetic solutions of free cyanide. The experimental methodology employed involved a sequence of two 2³ factorial designs using three factors: initial [CN⁻]: 100–400 mg/L; H₂SO₅:CN⁻ molar ratio: 1–1.5–3–4.5; pH: 9–11; each one conducted at one level of Caro’s Acid strength which is obtained with the H₂SO₄:H₂O₂ molar ratio used in Caro’s Acid preparation of 3:1 and 1:1. The objective was the evaluation of the effect of those factors on the reaction kinetics at room temperature. Statistical analysis showed that the three investigated variables were found to be significant, with the variables which affected the most being the initial [CN⁻] and the H₂SO₅:CN⁻ molar ratio. The highest reaction rates were obtained for the following conditions: H₂SO₅:CN⁻ molar ratio = 4.5:1; pH = 9; and Caro’s Acid strength produced from the mixture of 3 mol of H₂SO₄ with 1 mol of H₂O₂. These conditions led to a reduction of [CN⁻] from an initial value of 400 mg/L to [CN⁻] = 1.0 mg/L after 10 min of batch reaction time at room temperature. An empirical kinetic model incorporating the weight of the contributions and the interrelation of the relevant process variables has been derived as: −d[CN⁻]/dt = k [CN⁻]^1.8 [H₂SO₅]^1.1 [H⁺]^0.06, with k = 3.8 (±2.7) × 10⁻⁶ L/mg min, at 25 °C.     

Kinetics of high-sulphur and high-arsenic refractory gold concentrate oxidation by dilute nitric acid under mild conditions

On one hand, high-sulphur and high-arsenic refractory gold concentrate (HGC) leads to regional ecological damage. On the other hand, it contains a lot of valuable elements. So utilization of it can bring social and environmental benefits. In this paper, the kinetics of HGC oxidation by dilute nitric acid under mild conditions was investigated. The effects of particle size (50–335 μm), reaction temperature (25–85 °C), initial acid concentration (10–30 wt.%) and stirring speed (400–800 rpm) on the iron extraction rate (C_r) were determined. It is obvious that C_r increases with the rise of initial nitric acid concentration, reaction time and stirring speed, but decreases with the increase of particle size. Oxidation kinetics indicates that the rate of reaction is diffusion controlled. The activation energies were determined to be 10.70 kJ/mol in the 10% HNO₃ and 12.25 kJ/mol in the 25% HNO₃.     

Palladium(II) removal from chloride and chloride–nitrate solutions by chelating ion-exchangers containing N-donor atoms

The chelating ion-exchangers of functional iminodiacetate (Amberlite IRC-718), amidoxime (Duolite ES-346) and aminophosphonic (Duolite C-467) groups have been applied for Pd(II) removal from the model chloride (0.1–6.0 M HCl) and chloride–nitrate (0.1–0.9 M HCl and 0.9–0.1 M HNO₃ and 0.1–1.5 M HCl and 1.9–0.5 M HNO₃) solutions. The total ion-exchange capacities as well as recovery factors of Pd(II) were determined by the batch method. The influence of acid concentrations, phase contact time and macrocomponent addition (AlCl₃, CuCl₂, NiCl₂) was studied. The results show that the ion-exchangers of functional amidoxime and iminodiacetate groups can be widely recommended for Pd(II) ion removal from anodic slimes, and used up catalysts, as well as Pd(II) trace analysis due to their high selectivity.     

Extraction of titanium (IV) from acidic media by tri-n-butyl phosphate in kerosene

The extraction of titanium (IV) from sulfate, and nitrate solutions has been studied using tri-n-butyl phosphate (TBP) in kerosene. Extraction of titanium was affected by acid concentration over the range of 0.5–4 mol L^?1. The titanium distribution coefficient reached a minimum between 1 and 2 mol L^?1 acid for both sulfate and nitrate solutions. Third phase formation was observed in the extraction of titanium from acidic media at all condition tested. At the next stage, the stripping of titanium was studied using H₂SO₄, H₂SO₄ + H₂O₂ and Na₂CO₃. The kinetics of the stripping were very slow for H₂SO₄. The use of complex forming stripping agents (H₂SO₄ + H₂O₂) and Na₂CO₃ significantly improved the kinetics of stripping. About 98% recovery was achieved by extracting titanium from an aqueous nitrate solution using TBP and stripping with sodium carbonate.     

Solvent extraction and separation of palladium(II) and platinum(IV) from hydrochloric acid medium with dibutyl sulfoxide

The solvent extraction and separation performances of Pd(II) and Pt(IV) from hydrochloric acid solutions were investigated using dibutyl sulfoxide (DBSO) diluted in kerosene. Pd(II) was strongly extracted by a lower concentration DBSO in a lower concentration hydrochloric acid solution while the reverse was obtained for Pt(IV) extraction. Based on independent extraction and separation experiments of Pd(II) and Pt(IV), the separation parameters of Pd(II) and Pt(IV), including dibutyl sulfoxide concentration, contact time of aqueous and organic phases, organic/aqueous (O/A) phase ratio and H⁺ concentration of aqueous phase, were studied in detail, and the optimal separation parameters were obtained and summarized as the following: dibutyl sulfoxide concentration 0.6–1.2 mol dm^?3, organic/aqueous (O/A) phase ratio 0.6–1.0, H⁺ concentration of aqueous phase 1.0–1.5 mol dm^?3 and contact time of two phases 5 min. The as-prepared separation parameters were corroborated by the extraction and separation from a synthetic stock solution containing Pd(II), Pt(IV) as well as several common impurities like Fe(II), Cu(II) and Ni(II). The results revealed that Pd(II) could be separated efficiently from Pt(IV) with a high separation coefficient of Pd(II) an Pt(IV) (2.7 × 10⁴) by predominantly controlling dibutyl sulfoxide and hydrochloric acid concentrations. The extraction saturation capacity of Pd(II) was determined from 1.0 mol dm^?3 HCl solution with 3 mol dm^?3 dibutyl sulfoxide and its experimental value exceeded 14 g dm^?3 under the experimental conditions.Stripping of Pd(II) from loaded organic phase was performed using a mixed aqueous solution containing NH₄Cl and ammonia solutes. Pd(II) (99.2%) was stripped using the stripping solution containing 3% (m/v) NH₄Cl and 5 mol dm^?3 ammonia, respectively.     

Density functional theory study of α-Bromolauric acid adsorption on the α-quartz (1 0 1) surface

Adsorption mechanism of collector α-Bromolauric acid (CH₃(CH₂)₉CHBrCOOH, α-BLA) on α-quartz (1 0 1) surface has been investigated by first-principles calculations based on density functional theory (DFT). The interaction energies of H₂O molecule, calcium ions (Ca²⁺), hydroxyl ions (OH⁻), calcium hydroxyl ions (Ca(OH)⁺), and α-BLA⁻ ions with α-quartz (1 0 1) surface were in the order of Ca(OH)⁺ < Ca²⁺ < OH⁻ < H₂O < α-BLA⁻. The results revealed that the collector α-BLA cannot adsorb on α-quartz (1 0 1) surface due to the hindrance of hydration shell of quartz surface, while Ca(OH)⁺ could repulse the hydration shell and consequently adsorb on quartz surface, which further leads to the adsorption of the collector α-BLA⁻ anions on Ca(OH)⁺-activated quartz surface. Mulliken populations analysis of the external oxygen atom (O2) of quartz surface, calcium atom (Ca) of Ca(OH)⁺, and oxygen atom (O1) of collector α-BLA⁻ (–OH group) shows that the electron transfer between the Ca–O1 and Ca–O2 atoms. The overlap area of electron density between Ca–O1 and Ca–O2 atoms indicates strong interactions among the three atoms of Ca, O1, and O2, suggesting that Ca(OH)⁺ ions act as a bridge between the α-quartz (1 0 1) surface and the α-BLA collector.     

The effect of [Fe]TOT on the dissolution of synthetic Pb-doped UO2 and Th-doped UO2

The dissolution of synthetic Pb-doped UO₂ and Th-doped UO₂ was systematically studied to determine the influence of leach parameters [Fe]_TOT and ORP under standard leach conditions of: T = 50 °C, [H₂SO₄] = 15 g/L (0.15 M), and UO₂ = 100 mg/L. Results demonstrated reduced uranium dissolution in both systems compared to pure UO₂. This effect was greatest for Th-doped UO₂. The decrease in uranium dissolution between the doped systems and pure UO₂ was attributed to the formation of precipitate layers at the surface of the solid, slowing down or blocking uranium release. In the case of Pb-doped UO₂, the formation of a Pb sulphate phase was directly detected but in the case of Th-doped UO₂, no layer was found. For the latter system it was postulated that passivation of the Th-doped UO₂ surface occurs due to the formation of oxidised Th-rich phases Th(OH)₄, ThO₂ and ThO₂·nH₂O at the surface of grains preventing uranium release. In tests varying the ORP, there was an approximately linear dependence of the dissolution rate on [Fe]_TOT for both systems however the rate orders indicated a step change between an ORP of 420 and 460 mV. The specific influence of Fe^II showed that both Pb–UO₂ and Th–UO₂ exhibited two distinct regions of dissolution rate dependency similar to that previously noted for pure UO₂.     

Solvent extraction of tetravalent hafnium from acidic chloride solutions using 2-ethyl hexyl phosphonic acid mono-2-ethyl hexyl ester (PC-88A)

Solvent extraction of Hf(IV) from acidic chloride solutions has been carried out with PC-88A as an extractant. Increase of acid concentration decreases the percentage extraction of metal indicating the ion exchange type mechanism. The plot of logD vs log[extractant], M is linear with slope 1.8 indicating the association of two moles of extractant with the extracted metal species. Plot of logD vs log[H⁺] gave a straight line with a negative slope of ∼2 indicating the exchange of two moles of hydrogen ions for every mole of Hf(IV). The effect of Cl⁻ ion concentration at constant concentration of [H⁺] did not show any change in D values. Addition of sodium salts enhanced the percentage extraction of metal and follows the order NaSCN > NaCl > NaNO₃ ∼ Na₂SO₄. Stripping of metal from the loaded organic (LO) with different acids indicated sulphuric acid as the best stripping agent. Regeneration and recycling capacity of PC-88A, temperature, extraction behavior of associated elements was studied.     

Leaching behaviour of natural and heat-treated brannerite-containing uranium ores in sulphate solutions with iron(III)

Uranium leaching tests were conducted on two naturally occurring, highly metamict brannerite ores from the Crockers Well and Roxby Downs deposits, South Australia. The ores were leached over a range of temperatures and Fe^(III) and H₂SO₄ concentrations. As well, samples of the ores were calcined at 1200 °C in air to investigate the effect of thermally induced recrystallisation on uranium dissolution. For the unheated samples, a maximum of ∼80% U dissolution was obtained using an Fe^(III) concentration of 12 g/L, an acid concentration of 150 g/L H₂SO₄ and a temperature of 95 °C. The heat treated samples performed poorly under identical conditions, with maximum uranium dissolution of <10% recorded. High uranium dissolution from natural brannerite can be achieved providing; (i) acid strength, oxidant strength and temperatures are maintained at elevated levels (compared to those traditionally used for uraninite leaching), and, (ii) the brannerite has not undergone any significant recrystallisation (e.g. through metamorphism).     

Self-heating activation energy and specific heat capacity of sulphide mixtures at low temperature

Energy of activation (E_a) and specific heat capacity (C_p) for mixtures of sulphide minerals that on their own do not self-heat (SH), sphalerite/pyrite, pyrite/galena, chalcopyrite/galena and sphalerite/galena, were determined using a self-heating apparatus at temperatures below 100 °C in the presence of moisture. The mixtures all gave E_a ranging from 22.0 to 27.8 kJ mol⁻¹, similar to the range reported for Ni- and Cu-concentrates. The E_a is close to that for partial oxidation of H₂S which adds to the contention that the partial oxidation of H₂S contributes to SH of sulphides at low temperature. The C_p values ranged from 0.152 to 1.071 JK⁻¹ g⁻¹ as temperature rose from 50 °C to 80 °C, similar to the reported findings on Ni- and Cu-concentrates. The role of galvanic interaction in promoting SH is tested by examining correlations with the rest potential difference of the sulphides in the mixture.     

Leaching of chalcopyrite (CuFeS2) with an imidazolium-based ionic liquid in the presence of chloride

The effects of independent variables such as, temperature, concentration of ionic liquid (1-butyl-3-methyl-imidazolium hydrogen sulphate, [bmim][HSO₄]), chloride and sulphuric acid on copper extraction from chalcopyrite (CuFeS₂) ore were studied by surface optimization methodology. The Central Composite Face approach and a quadratic model were applied to the experimental design. The optimal copper extraction conditions given by the above methodology were 20% (v/v) of [bmim][HSO₄] in water, 100 g L⁻¹ chloride, and 90 °C. The concentration of chloride and the temperature together exert a synergistic effect in enhancing chalcopyrite dissolution. Experimental data were fitted by multiple regression analysis to a quadratic equation and analyzed statistically. A model was developed for predicting copper extraction from CuFeS₂ ore with variables such as Cl⁻, [bmim][HSO₄], H₂SO₄ concentrations and temperature in the range studied. The activation energy was calculated to be 60.4 kJ/mol (temperature range 30–90 °C), indicative of chemical control of the reaction and [bmim][HSO₄] acts as an acid in the reaction.     

Process development for the recovery of europium and yttrium from computer monitor screens

This work describes the development of a process for the recovery of Eu and Y from cathode ray tubes (CRTs) of discarded computer monitors with the proposition of a flow sheet for the metals dissolution. Amongst other elements, europium and yttrium are presented in the CRTs in quantities – 0.73 w/w% of Eu and 13.4 w/w% of Y – that make their recovery worthwhile. The process developed is comprised of the sample acid digestion with concentrated sulphuric acid followed by water dynamic leaching at room temperature. In the CRTs, yttrium is present as oxysulphide (Y₂O₂S) and europium is an associated element – Y₂O₂S:Eu³⁺ (red phosphor compound). During the sulphuric acid digestion, oxysulphide is converted into a trivalent Eu and Y sulphate, in solid form, with the liberation of H₂S. In the second step, metals are leached from the solid produced in the acid digestion step by dynamic leaching with water. This study indicates that a proportion of 1250 g of acid per kg of the sample is enough to convert Eu and Y oxysulphide into sulphate. After 15 min of acid digestion and 1.0 h of water leaching, a pregnant sulphuric liquor containing 17 g L⁻¹ Y and 0.71 g L⁻¹ Eu was obtained indicating yield recovery of Eu and Y of 96% and 98%, respectively. Both steps (acid digestion and water leaching) may be performed at room temperature.     

Mechanical activation of MoS2 + Na2O2 mixtures

A study of the mechanochemical activation of molybdenum ore concentrate (MoS₂) with sodium peroxide (Na₂O₂) shows that sodium molybdate dihydrate (Na₂MoO₄ · 2H₂O) is the final crystalline product. The mechanochemical formation of sodium molybdate dihydrate is evidenced by XRD, ²³Na MAS NMR and the increasing solubility of the molybdenum in water as the oxidative reaction proceeds.     

Study of silver leaching with the thiosulfate–nitrite–copper alternative system: Effect of thiosulfate concentration and leaching temperature

Thiosulfate system is considered an interesting alternative leaching process for precious metals. Nevertheless, most of the literature published on these conventional thiosulfate leaching solutions has been focused on the use of ammonia and copper to generate the cupric tetraamine complex, which acts as a catalytic oxidant for silver. However, ammonia toxicity is also a detrimental issue in terms of the process sustainability. For that reason, thiosulfate–nitrite–copper solutions were studied as an alternative less toxic system for silver leaching.In this work, the effect of the thiosulfate concentration (0.07 M, 0.1 M and 0.15 M) and temperature (room temperature, 30, 35, 40 and 45 °C) on the metallic silver leaching kinetics is presented for the S₂O₃–NO₂–Cu system. The results show that the thiosulfate concentration plays an important role in the S₂O₃–NO₂–Cu–Ag system since it controls the silver leaching kinetics. On the other hand, an increase in temperature favors the silver recovery.Finally, the SEM–EDS analysis, the X-ray mapping and the X-ray diffractograms show that the solid silver particles are coated by a Cu, S and O layer for the 0.07 M and 0.1 M thiosulfate experiments, which is consistent with the formation of antlerite (Cu₃(SO₄)(OH)₄); while the 0.15 M thiosulfate scenario produced a layer composed only of Cu and S, revealing the formation of stromeyerite (CuAgS). The UV–Visible technique confirmed the in-situ generation of copper–ammonia complexes for the 0.07 M leaching condition; however, these complexes are not formed at the 0.15 M condition.     

Contact angle and bubble attachment studies in the flotation of trona and other soluble carbonate salts

Trona, Na₂CO₃ · NaHCO₃ · 2H₂O, is mined as the primary source for sodium carbonate production in the United States. Recent studies have shown that the flotation method can be used for pre-processing of trona ore to remove insoluble mineral contaminants for the production of soda ash (sodium carbonate). Studies with carbonate salts suggest that certain important factors can affect their flotation response, including viscosity of the brine and interfacial water structure. Flotation studies showed that contrary to the strong flotation of NaHCO₃ with both anionic and cationic collectors, Na₂CO₃ does not float at all. Based on the analysis of interfacial water structure in saturated brines, Na₂CO₃ was found to act as a strong water structure maker, whereas NaHCO₃ acts as a weak water structure maker. Bubble attachment time measurements suggest that collector adsorption at the surface of NaHCO₃ induces flotation; this is not the case for Na₂CO₃. Contact angle measurements indicated that the surface of Na₂CO₃ is hydrated to a great extent, whereas the NaHCO₃ salt surface is less hydrated. These results reveal that there is a strong relationship between the interfacial water structure and the contact angle of these salts. The less stable NaHCO₃ surface is ascribed to the interfacial water structure which allows for NaHCO₃ flotation with both anionic and cationic collectors.