Recovery of vanadium during ammonium molybdate production using ion exchange

The recovery of vanadium in ammonium molybdate production with raw sodium molybdate solution was studied. Experimental results showed that the vanadium and some of impurities P, As and Si can be adsorbed along with the molybdenum from the feed liquid by the weak base resin D314 in the pH range of 2.5-3.5 and then they can be eluted from the loaded resin with ammonia liquor. The vanadium can partially natural-precipitate from the eluted solution. The lower the pH value is and, the longer the standing time is, the less the vanadium remained in the solution will be. Standing for 24 h in pH value 6.9, the vanadium in the solution was reduced to 0.51 g/L V₂O₅. It was found that the precipitate is ammonium isopoly-vanadate and it is impure. By washing the precipitate with hydrochloric acid and ammonia solution sequentially and then roasting at 500 °C for 2 h, the product of V₂O₅ with the purity 99.12% was obtained. The impurities P, As and Si in the stood solution were removed by purifying with MgCl₂ under the pH value range of 8.0-9.0 at 60-80 °C for about 2 h, while the removal of the vanadium in the solution was performed by adsorbing with the strong base resin D296 in pH value about 7.0. The devanadiumized solution can be used to produce high-quality ammonium molybdate. The loaded vanadium resin was eluted with HCl 6 mol/L and, the eluted solution was returned to adjust the sodium molybdate solution pH value. The vanadium can be effectively separated and recovered in the process.     

Comparative study of chelating ion exchange resins for the recovery of nickel and cobalt from laterite leach tailings

Commercially available chelating resins with the iminodiacetate functional group have been evaluated for their suitability for the adsorption of nickel and other metal ions by a resin-in-pulp process from the tailings of a pressure acid leach process for nickel laterites. The Amberlite IRC 748 and TP 207 MonoPlus resins were found to be the most suitable in terms of loading capacity for nickel and kinetics of adsorption. The resin with the highest nominal capacity was observed to adsorb less nickel as a result of the adsorption of greater amounts of the impurity ions. The equilibrium loading for nickel on the preferred resin was found to be similar from the ammonium and protonated form of the resin although the kinetics of adsorption is greater when the resin is initially in the ammonium form. A study of the kinetics of the loading of nickel and cobalt from pulp has shown that the rate can be described in terms of a first-order approach to equilibrium. The optimum pH for adsorption was found to be in the range 4 to 5 as this pH is high enough to maximize the adsorption of nickel and cobalt while preventing precipitation of nickel and cobalt as hydroxides from the pulp. A method for minimizing the competition from more strongly loaded ions such as iron(III) and chromium(III) which are present in the pulp was also developed in the initial laboratory phase of the study.     

Extraction of vanadium from direct acid leach solution of converter vanadium slag

Direct acid leaching of converter vanadium slag by titanium dioxide waste is eco-friendly and efficient, but with low selectivity. This novel technique can result in a vanadium solution which contains chromium(III), aluminium(III), magnesium(II), manganese(II) and high amount of iron(II) and iron(III). Bis (2-ethylhexyl) phosphoric acid (D2EHPA) and tri-butyl-phosphate (TBP) diluted with sulphonated kerosene were applied for vanadium extraction from the multi-element leach solution. The effects of the initial pH, concentration of D2EHPA, ratio of organic to aqueous phase, and the extraction time on the extraction efficiency of vanadium were investigated in saponification and unsaponifiable systems, respectively. The results showed that the vanadium extraction percentage can be up to 97% and the iron extraction percentage can be less than 10% in a thirteen-stage counter-current simulation test and the separation coefficient between vanadium and iron can reach to 109.8. Furthermore, vanadium(IV) can also be separated from other impurities such as aluminium(III), magnesium(II), manganese(II), chromium(III) efficiently. The loaded organic phase was stripped by 184?g?L^?1 sulphuric acid solution in a three-stage counter-current stripping process and with the total vanadium stripping percentage of greater than 99.5%. In the end, the vanadium pentoxide products with a purity of 99.14% were obtained.     

特种离子交换树脂从钼酸铵溶液中深度分离钒

研究了采用特种阴离子交换树脂分离钼酸铵溶液中的钒。首先,采用静态吸附考察了料液平衡pH、料液钼浓度、氯离子浓度、吸附时间等因素对分离过程的影响;然后,进行动态交换实验,采用料液pH=8.01,接触时间60 min,处理料液为67倍树脂体积时,除钒率达到99.83%,树脂对V2O5工作交换容量为78.7 g/L,钒钼分离系数达到22522.9;用2 mol/L的NaOH溶液可以对负载树脂实现彻底解析。     

Removal of impurities from scandium chloride solution using 732-type resin

The deep removal of Al, Fe(Ⅱ/Ⅲ), Ca, Zr, Ti and Si from scandium chloride solution was carried out by using 732-type strong acid cation exchange resin. The effects of pH value, contact time and complexing agents(EDTA) on the purification process are investigated. The results indicate that the 732-type resin have a good scandium selectivity and the adsorption order is Sc Fe(Ⅲ)Al Ca Zr Ti Si in the pH range of 1-3. The separation of Sc and Zr, Si, Ti can be directly carried out because the resin have a good adsorption effect on Sc, AI and Fe(Ⅲ) but poor adsorption effect on Zr, Si and Ti under the condition of pH = 2.5 and contact time 180 min. The Fe(Ⅱ), Ca and Al are selectively adsorbed on the resin by adding reducing agent ascorbic acid and EDTA into the solution for reducing Fe(Ⅲ) to Fe(Ⅱ) and complexing Sc.By using two-step ion exchange adsorption separation method, the removal rates of Fe(Ⅲ), Ti, Al, Ca, Zr and Si are 95.5%,99.8%,100%,98.2%,98.6% and 100%,respectively.     

A two-parameter model for ion exchange process of ion-adsorption type rare earth ores

The establishment of a mathematical model for the ion exchange process is key to creating a theoretical basis for the mining of ion-adsorption type rare earth ores. Ore samples from Xinfeng, Xunwu and Anyuan were used as the test ore samples in the present study and equilibrium batch tests of ore sample leaching using various ammonium sulfate concentrations were performed. The results show that, after leaching, there is a negative exponential relationship between the ratio of the solid-phase rare earth ion concentration to the aqueous-phase rare earth ion concentration and the initial ammonium ion concentration. However, there is a linear relationship between the natural logarithm of the ratio of the solid-phase ammonium ion concentration to the aqueous-phase ammonium ion concentration and the initial ammonium ion concentration. Based on the above two functional relationships, a two-parameter model for the equilibrium ion exchange process of ion-adsorption type rare earth ores was established. Using the established model to analyze the test data the model error for the Xunwu ore sample is found to be less than 5%. The proposed model is more accurate compared with the Kerr model. The two-parameter model proposed in this study provides theoretical support for the numerical simulation of column leaching (in-situ leaching) of ion-adsorption type rare earth ores.     

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%.     

改性季铵盐从高碱度溶液中萃取钒的研究

利用改性的季铵盐萃取剂从高碱度溶液中进行萃取钒的研究.考察了萃取剂浓度、仲辛醇浓度、料液碱度、料液初始含钒浓度、相比、振荡时间及温度对钒萃取过程的影响,并绘制了钒萃取等温曲线.结果表明:在料液碱度为1.5 mol/L、料液初始V2O5浓度为3.57 g/L,相比O/A为2/1,萃取时间为10 min,萃取温度为30℃,有机相组成为(质量分数)50％CO32-型三-十八烷基甲基氯化铵+20％仲辛醇+30％磺化煤油的条件下,钒的单级萃取率可达80％,钒的饱和容量可达5.85 g/L;在此条件下,经过5级逆流萃取,萃余液中V2O5浓度可降为0.05 g/L以下,钒萃取率达到99.4％.     

Uranium sorption from phosphoric acid solutions using selective ion exchange resins: Part II. Kinetic studies

Experiments have shown that the sorption of uranium from strong phosphoric acid solutions onto four D2EHPA/TOPO based ion exchange resins and one aminophosphonic acid resin is particle diffusion controlled in the uranium concentration range 42–780 μM. Interdiffusion coefficients of about 10^?12 m²s^?1 were obtained for D2EHPA/TOPO resins and 0.14 × 10^?12 m²s^?1 with the aminophosphonic acid exchanger at 20°C in 3 M H₃PO₄. Both homogeneous particle diffusion, based on Fick's Law, and the ash layer diffusion model fitted the kinetic measurements.     

钨冶炼离子交换工艺中钨锡分离研究现状

20世纪70年代以来,由于离子交换技术的应用,我国在钨冶炼方面取得了显著的进步,产品质量名列前茅.结合国外钨冶炼工艺中除锡研究的发展情况,对我国除锡技术的发展现状进行了分析,并提出了除锡技术今后的发展方向.     

Separation and recovery of iron and scandium from acid leaching solution of red mud using D201 resin

Red mud is a byproduct of alumina refining of bauxite ores, and is a significant source for extracting scandium. However, a large amount of iron in red mud makes it difficult to recover scandium because Fe(III) and Sc(III) have similar physicochemical properties. In this study, a new method was developed for selective separation of iron and scandium in acid leachate of red mud using D201 resin. Theoretical calculations indicate that the ferric species mainly exists as FeCl₃ or FeCl₄⁻ at chloride concentration above 6.65 mol/L, while scandium still exists as ScCl₂⁺, making it possible to selectively separate iron from scandium through anion resin adsorption. The factors affecting the adsorption of iron and scandium such as chloride concentration, resin dosage, adsorption time, and temperature were evaluated in batch experiments. The Langmuir model was successfully applied to both iron and scandium adsorption, and the maximum adsorption capacities of iron and scandium are 147.06 and 0.95 mg/g, respectively, indicating a significant difference between iron and scandium. Raman analysis further demonstrates that the iron is adsorbed onto D201 resin as FeCl₄⁻ anion.     

Study on removing Mo from tungstate solution using coprecipitation adsorption method based on novel Mo sulphidation process

Abstract

A coprecipitation–adsorption method based on a novel Mo sulphidation process is proposed for deep removal of molybdenum from tungstate solutions. First, a new sulphur source, P₂S₅, is used to provide HS^? for the Mo sulphidation process. The MoO₄^2? can be thoroughly converted into MoS₄^2? while tungsten still exists as WO₄^2?. Next, in the sodium tungstate solution, 99·4% of MoS₄^2? is removed with precipitated ‘nascent’ CuS under 6 h reaction at 30°C when the molar ratio Cu/Mo is 3. Similarly, in a commercial (NH₄)₂WO₄ solution, by controlling the molar ratio Cu/Mo at 6, 98·4% of MoS₄^2? is removed after 1 h at 30°C. Subsequently, phosphorus introduced during Mo sulphidation is removed using the magnesium ammonium phosphate method allowing the ammonium paratungstate to reach the GB 10116-88 APT-0 standard.

On propose une méthode de co-précipitation et adsorption basée sur un nouveau procédé de sulfuration du Mo pour l’enlèvement en profondeur du molybdène de solutions de tungstate. Premièrement, on utilise une nouvelle source de soufre, P₂S₅, pour obtenir le HS^? pour le procédé de sulfuration du Mo. Le MoO₄^2? est entièrement converti en MoS₄^2? alors que le tungstène existe encore sous la forme de WO₄^2?. Ensuite, 99·4% du MoS₄^2? est enlevé de la solution de tungstate de sodium avec le CuS ‘naissant’ précipité, en moins de 6 heures de réaction à 30°C lorsque le rapport molaire de Cu/Mo est de 3. De façon similaire, 98·4% du MoS₄^2? est enlevé d’une solution commerciale de (NH₄)₂WO₄, après une heure à 30°C, en contrôlant à une valeur de 6 le rapport molaire de Cu/Mo. Subséquemment, le phosphore introduit lors de la sulfuration du Mo est enlevé en utilisant la méthode du phosphate ammoniaco-magnésien, permettant ainsi au paratungstate d’ammonium d’atteindre la norme GB de 10116-88 APT-0.     

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.     

Uranium sorption from phosphoric acid solutions using selective ion exchange resins: Part I. Isotherms of extraction and desorption

Experimental measurements have been made of the isotherms for the sorption and desorption of uranium from synthetic phosphoric acid solutions with two types of cation exchange resin. Work was performed with several D2EHPA/TOPO impregnated resins and with one commercially available ion exchange resin containing aminophosphonic acid groups. The effect of Fe²⁺, Fe³⁺ and Ca²⁺ on the effective capacity has been examined. Results show that D2EHPA/TOPO resins are n not affected by the presence of Fe²⁺ or of Ca²⁺. Ferrous ion, whichreduces uranium to the tetravalent oxidation state, prevents uranium sorption. The resin containing aminophosphonic acid groups possesses a higher effective capacity than D2EHPA/TOPO-based resins and was also not affected by the presence of Ca²⁺. Ferric and ferrous ions reduce the effective capacity considerably, though the resin still extracts U(IV).D2EHPA/TOPO-based resins can be eluted with 3 M H₃PO₄ at 40°C or 3 M H₃PO₄ solutions containing Fe²⁺. Ammonium carbonate was found to leach the organic extractants from the resin. The aminophosphonic acid resin can be eluted with (NH₄)₂CO₃ at 20°C producing eluate solutions highly enriched in uranium. A conceptual flowsheet for uranium recovery is proposed.     

717型阴离子交换树脂分离富集 光度法测定

研究了717型阴离子交换树脂柱分离 富集钼的条件,建立了光度法测定铜精矿和尾矿中微量钼的方法。在pH为7.5的条件下,钼以MoO４２ 形式被717型阴离子交换树脂定量吸附后,采用2 mol/L HNO3和0.5 mol/L NH4NO3混合溶液定量洗脱,消除了绝大部分共存离子的干扰。通过树脂柱分离 富集后,用硫氰酸钾分光光度法测定钼的检出限为5.6 μg/L,相对标准偏差（RSD）为2.2 % (n=5),加标回收率在96.2 %～105.7 %之间,实际样品测定结果与ICP AES方法测定值基本一致。     

Progress in green and efficient enrichment of rare earth from leaching liquor of ion adsorption type rare earth ores

Ion adsorption type rare earth ores (IATREOs) are a valuable strategic mineral resource in China, which feature a complete composition of fifteen rare earth elements and are rich in medium and heavy rare earth (RE) elements. In the leaching process for recovering rare earth elements from IATREOs, many impurities will be leached together with rare earth elements and enter the leaching liquor. An impurity removal-precipitation enrichment technique is currently applied to selectively recovery rare earth elements from the leaching liquor with the high content of impurities and low concentration of rare earth elements by using ammonium bicarbonate in the industry. However, a high loss of rare earth elements and severe ammonia nitrogen pollution are caused by this process. Therefore, more beneficial impurities removal technologies, mainly for aluminum, and green enrichment technologies with lower pollution are now urgently needed. For this purpose, this paper analyzed two aspects of research progress in recent decades: the green separation of rare earth elements and aluminum from leaching liquor and the green and efficient enrichment of rare earth elements. Finally, an approach for the high-efficiency and green enrichment of rare earth elements from leaching liquor of the IATREOs is proposed in several aspects, including impurity inhibition leaching, neutralization and impurity removal, alkaline calcium and magnesium salt precipitation enrichment, and centrifugal extraction enrichment.     

离子交换树脂在红土镍矿湿法冶炼中的应用:被吸附溶液pH值和树脂再生性能研究

离子交换树脂吸附技术在回收红土镍矿浸出液中镍、钴的应用方面还处于研究阶段,前期研究结果表明该技术可行,并给出了相应的工艺条件.本文在之前实验研究基础上,针对液相初始pH值优选范围是pH≥3的结论和树脂再生的条件和性能进行了进一步的实验,得出了更为成熟的结论.pH值对树脂吸附各金属容量影响的实验表明,将被吸附溶液pH值调...