Selective recovery of vanadium as AMV from calcium vanadate sludge by direct AS leaching process: An industrial approach.

Generation of calcium vanadate waste sludge their management and treatment.is one of the major problem of metal processing industry. In this paper, we have proposed a simple process for the selective recovery of vanadium as ammonium metavanadate (AMV) from the calcium vanadate sludge using ammonium sulphate (AS) as a leaching agent. Under the optimum leaching condition (pH-7.5, temperature-80 °C, time-1 h, AS reagent-0.5 M) it is possible to leach out 82% of V values from the calcium vanadate sludge. The overall recovery of V is 81% with 98.5% AMV product purity. The AMV product quality from AS leach process has been compared with conventional H2SO4 leach process. The proposed process has major advantages such as, better economic benefits, less chemical consumption, minimum effluent recycling and less waste generation.

Separation of platinum and rhodium from chloride solutions containing aluminum, magnesium and iron using solvent extraction and precipitation methods.

The solvent extraction and precipitation methods have been used to develop a process to separate platinum and rhodium from a synthetic chloride solutions containing other associated metals such as (mg/L): Pt-364, Rh-62, Al-13880, Mg-6980, Fe-1308 at <1M HCl acidity. At pH 3.4, the quantitative precipitation of Al and Fe was achieved using 10 wt% Na(3)PO(4)·12H(2)O, with ~4% loss of Pt and Rh due to adsorption phenomenon. The selective separation of platinum was carried out with 0.01 M Aliquat 336 (a quaternary ammonium salt) at an aqueous to organic ratio (A/O) of 3.3 in two stages. Stripping of Pt from loaded organic (LO) at O/A ratio 6 with 0.5 M thiourea (tu) and HCl indicated that ~99.9% stripping efficiency. In stripping studies, needle like crystals of Pt were found and identified as tetrakis (thiourea) platinum (II) chloride ([Pt(tu)(4)]Cl(2)). The selective precipitation of rhodium was performed with (NH(4))(2)S from platinum free raffinate with a recovery of >99%.

Process for the separation and recovery of palladium and platinum from spent automobile catalyst leach liquor using LIX 84I and Alamine 336.

Spent catalysts from automobile industry contain environmentally critical and economically valuable metals such as Pt, Pd, Fe, Ni, Mn, and Cr. In this paper, we present a process for the selective separation and complete recovery of palladium (Pd) and platinum (Pt) from hydrochloric acid leach liquors of spent automobile catalyst employing solvent extraction method. Typical composition of leach liquor used for the present study contains (mg/L): Pd-150, Pt-550, Mn-500, Ni-1000, Fe-1500, Cr-100 and 3 M HCl. Selective separation of Pd from the leach liquor is achieved with 0.5 vol.% LIX 84I (2-hydroxy-5-nonylacetophenone oxime in a mixture with a high flash point hydrocarbon diluent) in kerosene at an aqueous to organic (A/O) ratio of 3 in 2 stages, with an enrichment factor of three. Quantitative stripping of Pd from loaded organic is achieved with 0.5 M thiourea and 1 M HCl. Co-extraction of Fe and Pt with 5 vol.% Alamine 336 (tertiary amine of mixed tri-octyl/decyl amine) in kerosene followed by selective scrubbing of Fe with dilute HCl and complete stripping of Pt from loaded organic was proposed with 0.5 M thiourea and 0.1 M HCl. Purity of Pd and Pt strip solutions are 99.7%. Finally, the present process can solve environmental related issues and at the same time recover valuable metals in pure form.

Nickel recovery from spent Raneynickel catalyst through dilute sulfuric acid leaching and soda ash precipitation.

Pharmaceutical industry makes extensive use of Raneynickel catalyst for various organic drug intermediates/end products. Spent catalysts contain environmentally critical and economically valuable metals. In the present study, a simple hydrometallurgical process using dilute sulfuric acid leaching was described for the recovery of nickel from spent Raneynickel catalyst. Recovery of nickel varied with acid concentration and time, whereas temperature had negligible effect. Increase of S/L ratio to 30% (w/v) showed marginal effect on nickel (90%) recovery, whereas Al recovery decreased drastically to approximately 20%. Under the optimum conditions of leaching viz: 12 vol.% H(2)SO(4), 30 degrees C, 20% solid to liquid (S/L) ratio and 120 min reaction time, it was possible to recover 98.6% Ni along with 39.2% Al. Leach liquor [pH 0.7] containing 85.0 g/L Ni and 3.25 g/L Al was adjusted to pH 5.4 with 30 wt.% alkali for quantitative aluminum removal. Nickel loss was about 2% during this Al removal step. Nickel from the purified leach liquor was recovered as nickel carbonate by adding required amount of Na(2)CO(3). The purity of NiCO(3) product was found to be 100% with a Ni content of 48.6%. Na(2)SO(4) was recovered as a by-product with a purity of 99%. Complete process is presented.

Selective recovery of molybdenum from spent HDS catalyst using oxidative soda ash leach/carbon adsorption method.

The petroleum refining industry makes extensive use of hydroprocessing catalysts. These catalysts contain environmentally critical and economically valuable metals such as Mo, V, Ni and Co. In the present study, a simple hydrometallurgical processing of spent hydrodesulphurization (HDS) catalyst for the recovery of molybdenum using sodium carbonate and hydrogen peroxide mixture was investigated. Recovery of molybdenum was largely dependent on the concentrations of Na2CO3 and H2O2 in the reaction medium, which in turn controls the pH of leach liquor and the presence of Al and Ni as impurities. Under the optimum leaching conditions (40 g L(-1) Na2CO3, 6 vol.% H2O2, room temperature, 1h) about 85% recovery of Mo was achieved. The leach liquor was processed by the carbon adsorption method, which selectively adsorbs Mo at pH around 0.75. Desorption of Mo was selective at 15 vol.% NH4OH. With a single stage contact, it was found possible to achieve >99%, adsorption and desorption efficiency. Using this method, recovery of molybdenum as MoO3 product of 99.4% purity was achieved.

Liquid-liquid extraction of tetravalent zirconium from acidic chloride solutions using Cyanex 272.

The liquid-liquid extraction of zirconium(IV) from acidic chloride solutions was carried out with Cyanex 272 as an extractant diluted in kerosene. An increase of the acid concentration decreased the percentage extraction of metal, which indicates that the extraction follows ion exchange-type mechanism: MO2+(aq) + 2(HA)2(org) <--> MO (HA2)2(org) + 2H+(aq), where, M = Zr(IV); HA = Cyanex 272. The extraction of Zr(IV) increases with an increase of the extractant concentration. In a plot of log D vs. log[extractant], M is linear with a slope of approximately 2, indicating the association of two moles of extractant with the extracted metal species. On the other hand, the extraction decreases with an increase of the H+ ion concentration. A plot of log D vs. log[H+] gave a straight line with a negative slope of 1.7, indicating the exchange of two moles of hydrogen ions for every mole of Zr(IV). The effect of the Cl- ion concentration at a constant concentration of [H+] did not show any change in the D values. The addition of sodium salts enhanced the percentage extraction of metal, and followed the order of NaSCN > NaNO3 > Na2SO4 > NaCl. The stripping of metal from the loaded organic (L.O) with different acids indicated sulfuric acid to be the best stripping agent. An increase of the temperature during the extraction and stripping stages increases the metal transfer, showing that the process is exothermic. The synergism, regeneration and recycling capacity of Cyanex 272; the extraction behavior of associated elements, such as Hf(IV), Ti(IV), Al(III), Fe(III); and IR spectra of the extracted Zr-Cyanex 272 complex were studied.

Extractive spectrophotometric determination of Cobalt(II) in synthetic and pharmaceutical samples using Cyanex 923.

Cyanex 923 has been proposed as a sensitive analytical reagent for the direct extractive spectrophotometric determination of cobalt(II). Cobalt(II) forms a blue-colored complex with Cyanex 923 in the organic phase. The maximum absorbance of the complex is measured at 635 nm. Beer's law was obeyed in the range 58.9 - 589.0 microg of cobalt. The molar absorptivitiy and Sandell's sensitivity of the complex was calculated to be 6.79 x 10920 l mol(-1) cm(-1) and 0.088 microg cm(-2), respectively. The nature of the extracted species was found to be Co(SCN)2 x 2S. An excellent linearity with a correlation coefficient value of 0.999 was obtained for the Co(II)-Cyanex 923 complex. Stability and regeneration of the reagent (Cyanex 923) for reuse is the main advantage of the present method. The method was successfully applied to the determination of cobalt in synthetic mixtures and pharmaceutical samples was found to give values close to the actual ones. Standard alloy samples, such as high-speed tool BCS 484 and 485, have been tested for the determination of cobalt for the purpose of validating the present method. The results of the proposed method are comparable with atomic absorption spectrometry and were found to be in good agreement.

Solvent extraction of Ni(II) from sulfate solutions with LIX 84I: flow-sheet for the separation of Cu(II), Ni(II) and Zn(II).

This paper reports on solvent-extraction studies of Ni(II) from sulfate solutions with LIX 84I (2-hydroxy-5-nonylacetophenoneoxime) as the extractant. The extraction of metal depends on the equilibrium pH of the aqueous phase and the extractant concentration. The transfer of metal follows a cation exchange-type mechanism: Ni2+ + 2HA --> NiA2 + 2H+. Extraction varies with the nature of the diluents. Temperature has no effect on the extraction of metal. The extraction behavior of associated metals clearly demonstrates the application of LIX 84I as the extractant for the separation of Cu(II), Ni(II) and Zn(II). Based on the results, a flow sheet of the process was developed.