Sulfuric Acid Leach Solution Research Articles

Selective Extraction of Cu(II) from Sulfuric Acid Leaching Solutions of Spent Lithium Ion Batteries Using Cyanex 301

Spent lithium ion batteries contain cobalt, copper, lithium, manganese and nickel. Recovery of these metals with high purity from spent lithium ion batteries is of importance. A sulfuric acid solution is generally employed for the treatment of spent lithium ion batteries, but it is difficult to separate these metal ions from the leaching solution because the divalent metal ions have similar extraction behaviors. Once Cu(II) is separated from the sulfuric acid leaching solution, the traditional solvent extraction process for the separation of Co(II) and Ni(II) can be applied to the raffinate. Therefore, selective extraction of Cu(II) was investigated in this work using Cyanex 301. The effect of the concentration of sulfuric acid and Cyanex 301 was investigated. Within the sulfuric acid concentration range (pH 5-2.0 M) employed in this work, Cu(II) was completely extracted by Cyanex 301, while the extraction percentage of other metal ions decreased to negligible when the sulfuric acid concentration increased to 2.0 M. The addition of TBP to Cyanex 301 depressed the extraction of other metal ions except Cu(II) at an initial pH of 3, facilitating the selective extraction of Cu(II). The McCabe-Thiele extraction diagram indicates that most of the Cu(II) was extracted by two count-current extraction at an A/O ratio of 2. The Cu(II) loaded in Cyanex 301 was completely stripped by using 5% aqua regia. Key words: spent lithium ion batteries, sulfuric acid, Cyanex 301, copper(II), extraction

Selective recovery and separation of Zr and Hf from sulfuric acid leach solution using anion exchange resin

Apart from zircon and baddeleyite, some other minerals also contain amounts of Zr and Hf, and H2SO4 can be the best candidate leachate for their recovery. The subsequent step in the process would be the recovery and separation of Hf and Zr from the resulting leach solution together with impurities. This reported study focused on the recovery process following the leaching of Zr and Hf from a eudialyte residue after REEs extraction. The chemistry of the solution was studied to determine the feasibility of different methods, and based on the results, an anion exchange resin was chosen to selectively recover Zr and Hf from the acidic sulfate solution. Batch and continuous ion exchange experiments using a strong anionic resin (MP600) were performed to develop the ion exchange process. The experimental results showed that Zr and Hf can be selectively recovered over other impurities. In the process of the adsorption of Zr and Hf, the bed volume reached 6 using the feed solution at a pH 0.94 at 25 °C with a contact time of 2 h. Pure Zr sulfate solution was used to scrub the load resin, after which the loaded resin was easily eluted using 10% NH4HCO3 + 1 mol/L NH4OH solution. This produced 21.23 g/L Zr, 0.03 g/L Hf, 0.01 g/L Fe, 0.01 g/L Al, 0.02 g/L Nb and < 0.001 g/L Ca in the eluate. The impurity removal from the leachate and the enrichment of Zr were achieved by the process.

Highly efficient separation and extraction of vanadium from a multi-impurity leachate of vanadium shale using tri-n-octylmethylammonium chloride

An efficient separation and extraction method for the recovery of V (V) from a sulfuric acid leaching solution of vanadium shale was developed using the ionic liquid tri-n-octylmethylammonium chloride (TOMAC). In this study, the main factors affecting the extraction of vanadium using TOMAC were investigated. Under optimum conditions, including a TOMAC concentration of 20 vol%, initial aqueous phase pH of 1.8, phase ratio (O/A) of 1:10, and extraction time of 2 min, 98.1% of vanadium was separated by a three-stage extraction, while almost no other impurities were extracted. The V (V) present in the scrubbed organic phase could be quantitatively stripped using 1.0 mol/L NaOH. Moreover, the extraction mechanism using TOMAC was analyzed by Fourier transform infrared (FT-IR) spectroscopy as well as the slope method. The obtained results supported the existence of an anion exchange mechanism between the Cl− atoms of TOMAC and H2V10O284− (HV10O285−) within the extraction system.

Recovery of Ge(IV) from synthetic leaching solution of secondary zinc oxide by solvent extraction using tertiary amine (N235) as extractant and trioctyl phosphate (TOP) as modifier

The N235–TOP–sulfonated kerosene system was applied for the recovery of Ge(IV) from a synthetic sulfuric acid leaching solution of secondary zinc oxide, containing Ge(IV), Zn(II), As(III), Fe(III), Si(IV), and Cd(II). The co-extraction of As(III) with Ge(IV) is a challenging problem in the extraction of Ge(IV) from aqueous solution by N235. However, TOP, as a modifier, suppressed the co-extraction of As(III) significantly. The effects of factors including N235 concentration, extraction time, tartaric acid to Ge(IV) molar ratio and temperature on the extractions of Ge(IV) and As(III) were studied. At an organic/aqueous (O/A) phase ratio of 1/1, aqueous phase pH of 1.2, extraction time of 10.5 min, tartaric acid-to-Ge(IV) molar ratio of 5.0, and temperature of 25 °C, the extraction efficiency of Ge(IV) reached 93.9%, whereas only 1.3% of As(III) was extracted by the 25%(v/v)N235–10%(v/v) TOP–sulfonated kerosene system in a single stage extraction. The Ge(IV) in the raffinate was less than 1.5 mg/L when employing a five-stage continuous countercurrent extraction at an O/A ratio of 1/10. In addition, aqueous NaOH solution of 0.5 mol/L was an appropriate reagent for the stripping of both Ge(IV) and As(III) at an O/A ratio of 1/1 and 25 °C for 10.0 min.

Green Recycling of Goethite and Gypsum Residues in Hydrometallurgy with α-Fe3O4 and γ-Fe2O3 Nanoparticles: Application, Characterization, and DFT Calculation

Millions of tons of hazardous iron oxide residues are produced during the iron purification process of sulfuric acid leaching solutions in the nonferrous metals hydrometallurgy industry per year. T...

Cooperative influence of D2EHPA and TBP on the reactive extraction of cobalt from sulfuric acid leach solution in a horizontal semi-industrial column

This article concerns the recovery of cobalt from a sulfuric acid leach solution in a horizontal pulsed perforated-plate column. Firstly, the influence of initial aqueous pH and the concentration of di(2-ethylhexly) phosphoric acid+tri-butyl phosphate has been investigated. It is observed that the mixture of these extractants is able to satisfactorily extract Co(II) from sulfate solution. Moreover, the influences of operational parameters including the phase flow rates along with pulsation intensity on the holdup, drop sizes and their distribution as well as extraction and stripping efficiencies and dispersion coefficients have been evaluated. Results show the highest efficiency of 95.3% and 93.2% for extraction and stripping of Co(II) at pulsation intensity of 1.30cm/s. Finally, new correlations for predicting mean drop size and holdup in the column are presented, which provide the average absolute relative error of 8.27% and 5.46% respectively.

Separation of indium from lead smelting hazardous dust via leaching and solvent extraction

A separation and extraction of indium from lead smelting hazardous dust (LSHD) was studied systematically. The novel technological keys included the pressure sulfuric acid leaching of LSHD, precipitation and enrichment of indium, atmosphere sulfuric acid leaching of indium-enriched residues and solvent extraction of indium. The optimum parameters of each procedure were obtained. More than 99% zinc and 95% indium were leached from LSHD by two-stage countercurrent pressure sulfuric acid leaching using the initial CH2SO4 of the 1st stage 30–50g/L, the initial CH2SO4 of the 2nd stage 120–150g/L, temperature 413–433K, total pressure 0.6–1.0Mpa, time 1–1.5h,L/S ratio 3–5/1 and agitation speed 400 r/min respectively. More than 97% indium was precipitated from the 1st pressure sulfuric acid leaching solution under the terminal neutralization pH 4.0, temperature 343–353K and time 1h. More than 96% indium was re-leached by two-stage countercurrent atmosphere sulfuric acid leaching from the indium-enrichment residues using L/S ratio 3–5/1, temperature 363–368K, agitation speed 500 r/min, time 0.75–1.5h, the initial CH2SO4 of the first stage and the second stage atmosphere sulfuric acid leaching with 40–50g/L and 90–110g/L, respectively. More than 98% indium was extracted by three-stage solvent extraction using 30% (v/v) di(2-ethylhexyl)phosphoric acid (D2EHPA), phase ratio 1/6, time 1min and temperature 298K.

Separation and recovery of iron impurity from a vanadium-bearing stone coal via an oxalic acid leaching-reduction precipitation process

Sulfuric acid leaching solutions of vanadium-bearing stone coal usually contain massive iron impurity, which is detrimental to vanadium extraction because iron is easy to precipitate or co-extracted during the solvent extraction and ion exchange. Meanwhile, sulfuric acid in wastewaters is environmentally hazardous. In this paper, a vanadium-bearing stone coal was leached with oxalic acid, which has the advantages of cheap biological sources and good environmental compatibility. The iron impurity was separated from the oxalic-acid leaching vanadium-bearing solution (OLVS) via reduction precipitation. During the oxalic acid leaching, 80.5% of the vanadium can be recovered, and 50.2% of the iron impurity was leached with vanadium under the conditions: a CaF2 addition of 5 wt%, an oxalic acid dosage of 60 wt%, a leaching time of 4 h, a leaching temperature of 95 °C and a water-mineral ratio of 1.0 mL/g. However, the concentrations of vanadium and iron were 1777.5 ppm and 6552.5 ppm, respectively, in the OLVS. By adopting reduction precipitation, 83.8% of the iron was precipitated as FeC2O4·2H2O with only a 2.5% vanadium loss when an iron powder dosage was 2.5 times that of the stoichiometry for 90 min at 25 °C. The oxalic-acid leaching vanadium-bearing solution after reduction precipitation (OLVS-P) contained 1733.6 ppm of vanadium and only 1059.8 ppm of iron. Analyses indicated that the as-produced FeC2O4·2H2O (P-FeC2O4·2H2O) was in a high crystalline state with a purity of 98.2%. Therefore, the iron impurity can be separated and recovered from the vanadium-bearing stone coal via the oxalic acid leaching-reduction precipitation process.

Reactive extraction of cobalt sulfate solution with D2EHPA/TBP extractants in the pilot plant Oldshue–Rushton column

In this study, a pilot plant Oldshue–Rushton extraction column has been proposed to extract and recover cobalt from sulfuric acid leach solution. The effect of D2EHPA and TBP concentration, initial aqueous pH and stripping solvent on the Co(II) extraction was investigated in the batch experiment. The results from batch experiments showed that the D2EHPA/TBP mixed extractant is a suitable choice for cobalt extraction from sulfate solution. In continuous experiments, the effects of variables such as rotor speed, dispersed and continuous phase velocities on holdup, mean drop sizes, extraction and stripping efficiencies were studied. Empirical correlations for prediction of the dispersed phase holdup and mean drop sizes in the literature were compared with the experimental results, and the modified correlations were proposed with reaction condition in the Oldshue–Rushton extraction column. The results of cobalt extraction with D2EHPA/TBP in the extraction column, showed the feasibility of the column for cobalt separation from sulfuric acid leach solution with extraction efficiency of 93% and stripping efficiency of 91.5% at high rotor speed (220rpm).

Two-Stage Separation of V(IV) and Al(III) by Crystallization and Solvent Extraction from Aluminum-Rich Sulfuric Acid Leaching Solution of Stone Coal

To improve separation of V(IV) and Al(III) from aluminum-rich sulfuric acid leaching solution of stone coal, the two-stage separation by crystallization and solvent extraction methods have been developed. A co-extraction coefficient (k) was put forward to evaluate comprehensively co-extraction extent in different solutions. In the crystallization stage, 68.2% of aluminum can be removed from the solution. In the solvent extraction stage, vanadium was selectively extracted using di-2-ethylhexyl phosphoric acid/tri-n-butyl phosphate from the crystalline mother solution, followed by H2SO4 stripped efficiently. A V2O5 product with purity of 98.39% and only 0.10% Al was obtained after oxidation, precipitation, and calcination. Compared with vanadium extraction from solution without crystallization, the counter-current extraction stage of vanadium can be decreased from 6 to 3 and co-extraction coefficient (k) decreased from 2.51 to 0.58 with two-stage separation. It is suggested that the aluminum removal by crystallization can evidently weaken the influence of aluminum co-extraction on vanadium extraction and improve the selectivity of solvent extraction for vanadium.

Direct solvent extraction of molybdenum(VI) from sulfuric acid leach solutions using PC-88A

The extractant 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester (PC-88A) was used to directly extract molybdenum(VI) from high sulfuric acid leach solutions containing 3mol/L H+. The effects of extractant concentration, solution acidity and type of stripping reagent on the recovery of molybdenum(VI) were investigated. The saturated loading capacity of molybdenum(VI) with 40% PC-88A in sulfonated kerosene was 60g/L. More than 98.7% of the molybdenum(VI) was extracted in a four-stage counter current extraction using an organic system consisting of 40% PC-88A and 60% sulfonated kerosene and a feed containing 17.3g/L Mo(VI) and 157.3g/L H2SO4 with an A/O ratio of 3/1 at 20°C for 5min. The extracted molybdenum(VI) species of Mo2O5(HA2)2 in loaded organic solutions were identified by the combination of slope and saturation methods, and further proved by FTIR spectral analysis. The stripping of molybdenum(VI) was completed in a two-stage counter current stripping using 6mol/L of NH3 with an A/O ratio of 1/1.5 at 30°C for 5min. The concentration of molybdenum(VI) in the strip liquor reached as high as 146g/L.

Sorption of rare-earth elements from phosphogypsum sulfuric acid leaching solutions

Sorption technology for extracting rare-earth elements (REEs) from sulfuric acid solutions of phosphogypsum percolation leaching has been developed. Sorption conditions that enable to reach maximum REE content in the sorbent at minimum content of thorium and calcium have been found.

Ion exchange resin selection for concentration and purification of uranium leached by sulphuric acid solution

Ion exchange is one of the most popular techniques for recovery and purification of uranium from sulfuric acid leaching solution, especially for recovery of uranium from a low uranium containing solutions. Resins commonly used are strong base or weak base anion resins with amine functional group. The anionic form of resins may be NO3-, Cl- , SO­42- or OH-. The selection of a resin depends on the uranium total exchange capacity, selectivity and the cost. The previous studies often use Amberlite IRA-420 for concentration and purification of uranium solution from Pa Lua sandstone ores. This is a good and suitable resin but high price. To diversify the resins and reduce the costs, instead of IRA-420, the authors tested two commercial resins Indion GS300 (India) and Purolite A400 (UK) in the processing of uranium solution from sandstone ores. The results showed that the uranium total exchange capacity of these resins is only about 80 - 85% over Amberlite IRA-420, but these resins should be able to be used instead of Amberlite IRA-420 due to their low cost and availability in Vietnam.

Recovery of molybdenum and vanadium with high purity from sulfuric acid leach solution of spent hydrodesulfurization catalysts by ion exchange

Valuable metals, such as Mo, V, Co and Ni together with base metals (Fe and Al) are present in the sulfuric acid leaching solutions of spent hydrodesulfurization catalysts. In order to develop a process to recover pure Mo and V from the solution, batch and continuous ion exchange experiments with strong anionic resin (AG1-x8) have been performed by employing a synthetic sulfuric acid solution. Batch experimental results indicated that the separation of molybdenum and vanadium from the solution with AG1-x8 was possible by adjusting the pH of the solution to 0.5 and 1.6, respectively. Column experimental results verified the complete separation and recovery of molybdenum and vanadium from the solution at the specified pH conditions. Molybdenum and vanadium in the loaded resin was eluted by using NaOH solution. The purity of Mo and V solution after elution was found to be higher than 99%. After elution, AG1-x8 was successfully regenerated by converting its form from hydroxide to chloride by using NaCl solution.

Recovery of water and acid from leach solutions using direct contact membrane distillation

This paper describes for the first time the use of direct contact membrane distillation (DCMD) for acid and water recovery from a real leach solution generated by a hydrometallurgical plant. The leach solutions considered contained H2SO4 or HCl. In all tests the temperature of the feed solution was kept at 60 °C. The test work showed that fluxes were within the range of 18-33 kg/m(2)/h and 15-35 kg/m(2)/h for the H2SO4 and HCl systems, respectively. In the H2SO4 leach system, the final concentration of free acid in the sample solution increased on the concentrate side of the DCMD system from 1.04 M up to 4.60 M. The sulfate separation efficiency was over 99.9% and overall water recovery exceeded 80%. In the HCl leach system, HCl vapour passed through the membrane from the feed side to the permeate. The concentration of HCl captured in the permeate was about 1.10 M leaving behind only 0.41 M in the feed from the initial concentration of 2.13 M. In all the experiments, salt rejection was >99.9%. DCMD is clearly viable for high recovery of high quality water and concentrated H2SO4 from spent sulfuric acid leach solution where solvent extraction could then be applied to recover the sulfuric acid and metals. While HCl can be recovered for reuse using only DCMD.

Solvent extraction of copper, zinc, cadmium and nickel from sulfate solution in mixer settler unit (MSU)

The disposal of large quantities of electronic scraps generated globally is causing not only environmental pollution but also loss of valuable resources. Therefore, R&D efforts have been made to develop a suitable solvent extraction (SX) process for extraction and separation of various metals viz. copper, zinc, cadmium and nickel from the solution expected from the leaching of e-scraps. Different process parameters viz. pH, aqueous to organic ratio, extraction isotherm to establish stage requirement for extraction in continuous mode, etc. have been studied to optimize the condition for the separation of metals. The studies showed the selective extraction of copper after iron precipitation from the sulfuric acid leach solution of e-wastes containing 0.92g/L Cu and minor impurities with 2-hydroxy-5-nonylacetophenoneoxime (LIX 84) above pH 2.0 in single stage. Further, extraction of zinc, nickel and cadmium was carried out using Cyanex 302. It was found that complete extraction of cadmium and zinc was taken place simultaneously using 2.5% Cyanex at eq. pH 2.1 leaving nickel in the aqueous raffinate, moreover Zn and Cd was selectively stripped out using 10% H2SO4 and 10% HCl, respectively. Based on the above studies a complete process flow sheet has been developed for separation/recovery to Cu, Zn, Cd and Ni from the leach liquor of e-waste.

Effect of colloidal potassium alum formation on vanadium recovery from acid leach solutions of stone coal

The nattier-blue colloidal particles precipitated out from the sulfuric acid leach solution of stone coal were identified to be potassium alum (KAl(SO4)2·12H2O). The particles were characterized through X-ray diffraction (XRD), scanning electron microscopy (SEM) and theoretical analysis with thermodynamics of ion crystallization. The results show that the particles belonged to potassium alum with micro-scale granular surfaces. The colloidal potassium alum formed with the temperature of the leaching solution cooling to 45°C. This observation could lead to 27.6% vanadium loss in the vanadium extraction process, which might be attributed to the chemical adsorption of vanadium in the form of the species of VO2+ on the potassium alum surfaces.

Recovery Gemanium from Pulverized Fuel Ash through the Method of H&lt;sub&gt;2&lt;/sub&gt;SO&lt;sub&gt;4&lt;/sub&gt;-NH&lt;sub&gt;4&lt;/sub&gt;F-NaClO&lt;sub&gt;3&lt;/sub&gt; - Leaching-Tannin Precipitation

This paper will conduct a systematic study on method of recovering Ge from pulverized fuel ash in Lincang prefecture, Yunnan province. Details are like this, in view of the fact that Ge contained in pulverized fuel ash is wrapped up by a lot of Silica, Cao, magnesium oxide and is difficult to react with hydrochloric acid, so, first, have the pulverized fuel ash grinded to 200 meshes; then, ammonium fluoride should be put into 85-90°Cdilute sulphuric acid solution to produce hydrogen fluoride so as to destroy silicon dioxide and calcium oxide contained in the pulverized fuel ash. Meanwhile, dilute sulphuric acid solution dissolves magnesium oxide, by doing so, Ge will be released to react with sulfuric acid and produce Sulfuric acid Ge, and Ge will go into the solution；as for low state Ge（e.g,GeO,GeS）contained in the pulverized fuel ash which is difficult to be dissolved by sulfuric acid, sodium chlorate will be used as a kind of oxidant to change bivalence Ge into tetravalence Ge, then dissolve it into the solution, after the liquid-solid separation, sulfuric acid leaching solution rich in Ge will be obtained. Then, tannic acid will be used as precipitant to recover Ge from the leaching solution, and Ge concentrate will be produced through distillation and roasting. After this, put Ge concentrate on hydrochloric acid distillation to get germanium tetrachloride, After going through strenuous evaporation, distillation and hydrolysis, germanium dioxide with high purity will be yield. This paper has conducted conditional tests and discussion on the process parameters which will influence recovery Ge from pulverized fuel ash, including the dosage of Sulfuric acid, ammonium fluoride, sodium chlorate ,tannic acid and extraction time to find out the prior controlling condition of Ge recovery，under this prior condition, Ge leaching rate is as high as 84%，and tannic sinking Ge rate is as high as 98%. The pulverized fuel ash, after recovering Ge, can be used to backfill the mine pit, besides, most of the lixivium after recovering Ge can be recycled, and the rest of it will be released after neutralized up to the standard by lime. This process enjoys the following quality: high Ge recovery rate, less time used during the process ,less consumption of auxiliary material, easier to control the process, avoiding the potential pollution on the environment and lower cost involved in it, therefore, which is a kind of efficient and economic process of recovering Ge from pulverized fuel ash.

Comparison of ion exchange and solvent extraction in recovering vanadium from sulfuric acid leach solutions of stone coal

Leaching of roasted stone coal with sulfuric acid results in a mixed vanadium solution containing Fe(III), Fe(II), Al(III) and Mg(II). Separation of V from synthetic sulfate solution containing these impurities has been investigated by ion exchange and solvent extraction. In ion exchange experiments, anion exchange resin ZGA414 was tested as its optimum adsorption capacity compared with D202, D453, D301FC and ZGA351 resins, and D2EHPA and TBP diluted with kerosene were employed in solvent extraction. Ion exchange tests indicated that only V(V) was loaded from the synthetic solution at pH>1.5, while it was difficult to separate V(V) from Fe(III), which also made resin toxic. Solvent extraction experiments revealed that V(IV) had a better extraction ratio than that of V(V), while Fe(III) had a serious effect on the extraction of V(IV). The co-extraction ratio of Al(III) and Mg(II) can be decreased by controlling their concentrations lower than 10g/L. Counter-current experiments with D2EHPA presented that 99% of V(IV) was extracted from the real leach solution after reduction process, leaving most of Fe(II), Al(III) and Mg(II) in the raffinate.

Study on Recovering of Value Metals from Acid Leaching Solution of Roasted Electroplating Sludge

Chelating electroplating sludge contains high grade of toxic metals or heavy metals, so this sludge should be treated and recovered of valuable metals from it. This paper mainly studied on the recovery process of Copper, Nickel and Zinc process from sulfuric acid leaching solution of roasted chelating electroplated sludge. It provided an effective technique for recovery of valuable metals from acid leaching solution. The efficiency of the process had been measured with the metal recovery rate. Experimental results showed that recovery rate of Copper and Nickel can respectively achieve at 99.30% and 97.00% from acid leaching solution after removal iron with zinc powder used reducing agent. The process has simple operation and high recovery rate. It can solve the environmental pollution problem of electroplating sludge and valuable metals can be recovered from which. It could bring better economic and environmental benefits to the enterprise.