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Abstract

Leach solutions and wastes of Bayer process are important resources for metals such as aluminum and vanadium. Despite the fact that vanadium cake is precipitated and removed in the Seydisehir Eti Aluminum Facility (Turkey), it cannot be used due to low metal content and impurities it contains. Within the scope of this study, research and development of environmentally acceptable, technically sound and low-cost chemical leaching and recovery methods were conducted for the recovery of vanadium from the by-product cake of the Bayer process. In the conducted studies, a sample of vanadium cake was used after its detailed characterization. Roasting tests were performed in order to remove the arsenic in the vanadium cake; however, it was found that roasting was not effective in removing the arsenic from the cake. The performance of different reagents were examined in chemical leaching tests (H2O and H2SO4 leaching, H2SO4 leaching with the addition of NaSO3, and NH4F); in the H2SO4 leaching tests performed with the addition of Na2SO3, the concentration of the reagents and the effect of temperature on the efficiency of vanadium recovery (max. 93.09%) were determined with the full factorial experimental design method, the outcomes were evaluated with ANOVA (variance analysis) method, and empirical models were formed. In lab and semi-pilot scale leaching tests, vanadium recoveries were 96.34% and 94.76% respectively. Vanadium was precipitated with NaOH and FeSO4 and almost all vanadium (95.8%) was obtained as Fe3(VO4)2. Cost analysis and economic evaluation have shown the economic feasibility of the leaching and recovery processes proposed.

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... In medical studies, vanadium compounds have been proposed for the treatment of diabetes, cancer, and other diseases [8], and these diverse chemical applications consume 5% of vanadium [2]. The latest developments in the use of vanadium are the storage of energy in newly developed batteries (vanadium redox battery (VRB)) [2,4], the production of glasses, the filtering of windows against UV rays, and the production of vitamin A tablets [9]. International Journal of Mining and Geo-Engineering IJMGE Table 1. ...
... three methods of purification are chemical precipitation, solvent extraction, and ion exchange. When the leach liquor impurities are high and vanadium concentration is low, solvent extraction and ion exchange method are used to treat [4,9,26]. Different studies have focused on vanadium recovery using solvent extraction with different organic extractants [67][68][69][70]. ...
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Vanadium is a strategic metal and its compounds are widely used in industry. Vanadium pentoxide (V2O5) is one of the important compounds of vanadium, which is mainly extracted from titanomagnetite, phosphate rocks, uranium-vanadium deposits, oil residues, and spent catalysts. The main steps of vanadium extraction from its sources include salt roasting, leaching, purification, and precipitation of vanadium compounds. In the hydrometallurgical method, first, the vanadium is converted to a water-soluble salt by roasting, and then the hot water is used to leach out the salt-roasted product and the leach liquor is purified by chemical precipitation, solvent extraction, or ion exchange processes to remove impurities. Then, a red cake precipitates from an aqueous solution by adjusting the conditions. To provide high pure vanadium pentoxide, it is necessary to treat the filtered red cake in ammonia solution. So, ammonium metavanadate (AMV) is precipitated, calcined, and flaked to vanadium pentoxide. In the pyrometallurgical method, vanadium-containing concentrate is smelted, and by forming titanium-containing slag and molten pig iron, oxygen is blown into pig iron in a converter or shaking ladles, and vanadium is oxidized to produce vanadium-rich slag. In the next step, the slag is roasted and treated by the hydrometallurgical process. In this paper, the industrial processes and novel developed methods are reviewed for the extraction of vanadium pentoxide.
... The removal of V from process liquors is a side benefit of process lime addition. V precipitates as calcium vanadate, as an impurity in tri-calcium aluminate (Ca 3 Al 2 (OH) 12 ), or as Na 7 (VO 4 ) 2 F·19H 2 O, as identified by Okudan et al. [25,45,46]. Our study did not detect any V in the aluminium hydroxide product (< 10 mg/kg). ...
Conference Paper
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Bauxites contain trace elements which have not been leached from their parent rock, but have instead remained in the composition of bauxite. During the refining of alumina from bauxites, these trace elements will also be introduced to the Bayer process along with the major bauxite constituents. This paper describes a study on the distribution of the trace elements gallium (Ga), vanadium (V), cerium (Ce), yttrium (Y) and thorium (Th) through the Bayer process. The first four elements can potentially be extracted as Bayer process by-products, whereas Th should be analysed due to its potentially adverse impact. Case-by-case examination showed that most of the trace elements end up entirely in bauxite residue. It was found that Ga accumulation was in an average range compared to previous reports and there is potential for the economic extraction of this metal. V was also found to accumulate in Bayer liquor in a similar amount to that reported previously, but was absent from the aluminium hydroxide product, with the majority ending up in bauxite residue. Practically all of the Ce, Y and Th were found in bauxite residue after bauxite processing. For the trace elements entirely ending up in bauxite residue, a method is proposed for predicting their content in residue based on their concentration determined in bauxite feed.
... Annually, approximately 120 million tonnes of red mud is generated with an estimated global inventory of over 3 billion tonnes . For possible metal recovery techniques from bauxite or red mud are leached where Al, V, and Ga also dissolve in acidic leach solution (Gladyshev et al., 2013a(Gladyshev et al., , 2015Liu and Li, 2015;Okudan et al., 2015). The rate of generation of red mud tends to increase, concomitant with the rapid depletion of high grade bauxites and increasing demand for aluminium (Abdulvaliyev et al., 2013). ...
... The S/L ratio, leaching time, temperature and concentration of the leaching reagents were studied under the experimental conditions (reagents and their concentration ranges have been stated according to the results of preleaching tests) shown in Table 2. H 2 O 2 is widely known to be able to play a role as an oxidizing and reducing agent (Vegliò et al., 2006), and the purpose of its use in the experiment was to see the behavior of V during acidic leaching. At the end of the leaching procedure, the solution was filtered , leach liquor was separated from the solid and the concentration of V in the leach liquors was determined by titration with 0.2 M KMnO 4 solution (Okudan et al., 2015 ). V recovery percentages were calculated according to Eqs. (4) and (5) by taking the average of the three analysis. ...
Article
Catalysts are used extensively in industry to purify and upgrade various feeds and to improve process efficiency. These catalysts lose their activity with time. Spent catalysts from a sulfuric acid plant (main elemental composition: 5.71% V2O5, 1.89% Al2O3, 1.17% Fe2O3 and 61.04% SiO2; and the rest constituting several other oxides in traces/minute quantities) were used as a secondary source for vanadium recovery. Experimental studies were conducted by using three different leaching systems (citric acid with hydrogen peroxide, oxalic acid with hydrogen peroxide and sulfuric acid with hydrogen peroxide). The effects of leaching time, temperature, concentration of reagents and solid/liquid (S/L) ratio were investigated. Under optimum conditions (1:25 S/L ratio, 0.1M citric acid, 0.1M hydrogen peroxide, 50°C and 120min), 95% V was recovered in the presence of hydrogen peroxide in citric acid leaching.
... The total reserves of V 2 O 5 in vanadium titano-magnetite reach 25.96 million [2][3]. The recovery of vanadium has been extensively studied from all kinds of vanadium resources [4][5][6]. ...
... Recovery techniques of base metals, including alumina, soda, ferric oxide and titanium oxide, from red mud are of much interest (Balomenos et al., 2011;Lindsay, 2011;Samouhos et al., 2013;Gladyshev et al., 2015;Liu and Li, 2015;Okudan et al., 2015;Qu et al., 2015). However, few systematic investigations have been conducted regarding the Ga extraction from the residue. ...
Article
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... The removal of V from process liquors is a side benefit of process lime addition. V precipitates as calcium vanadate, as an impurity in tri-calcium aluminate (Ca 3 Al 2 (OH) 12 ), or as Na 7 (VO 4 ) 2 F·19H 2 O [71,73,74]. Our study as well as the regular monitoring in the plant materials did not detect any V in the aluminium hydroxide product (<10 mg/kg). ...
Article
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The aim of this work was to achieve an understanding of the distribution of selected bauxite trace elements (gallium (Ga), vanadium (V), arsenic (As), chromium (Cr), rare earth elements (REEs), scandium (Sc)) in the Bayer process. The assessment was designed as a case study in an alumina plant in operation to provide an overview of the trace elements behaviour in an actual industrial setup. A combination of analytical techniques was used, mainly inductively coupled plasma mass spectrometry and optical emission spectroscopy as well as instrumental neutron activation analysis. It was found that Ga, V and As as well as, to a minor extent, Cr are principally accumulated in Bayer process liquors. In addition, Ga is also fractionated to alumina at the end of the Bayer processing cycle. The rest of these elements pass to bauxite residue. REEs and Sc have the tendency to remain practically unaffected in the solid phases of the Bayer process and, therefore, at least 98% of their mass is transferred to bauxite residue. The interest in such a study originates from the fact that many of these trace constituents of bauxite ore could potentially become valuable by-products of the Bayer process; therefore, the understanding of their behaviour needs to be expanded. In fact, Ga and V are already by-products of the Bayer process, but their distribution patterns have not been provided in the existing open literature.
... Vanadium is usually of +4 or +5 valence in solutions. The recovery of V(IV) and V(V) in acidic sulfate media and alkaline media have been studied separately [10][11][12][13][14][15][16][17]. The extraction of Fe(III) using di(2-ethylhexyl) phosphate (D2EHPA) in chloride solutions has also been widely studied [13,[18][19][20][21]. Studies on the solvent extraction of vanadium from chloride solution with trace impurities or the separation of vanadium and iron in acidic sulfate media have also been reported [22][23][24]. ...
Article
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In this study, scandium and lithium extractions were investigated using the atmospheric pressure agitation leaching method at acidic medium. The leaching tests were carried out at two stages. To remove the ionic impurities such as Na and Al first stage leaching was performed at relatively higher pH. Following solid-liquid separation of leach cake of the first stage leaching, it was subjected to the second stage leaching. The second stage leaching resulted in 95.1% Sc and 94.7% Li extractions. The overall Sc and Li recoveries were determined as 82.4% and 86.5%, respectively. Regarding the kinetic studies, it was understood that scandium and lithium leaching processes were controlled by a combination of chemical reaction and ash diffusion models. In this case, the activation energies were determined as 29.52 and 30.22 kJmol-1, respectively for scandium and lithium. As a result, while direct H2SO4 leaching of red mud is a challenge due to physical and chemical problems, an alternative solution was suggested using H2SO4
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The paper provides an overview of the methods used for processing of red mud to extract rare earth elements (REEs). Red mud is a toxic and highly alkaline waste. Several methods have been adopted and being practiced all over the world for the processing of red mud. Complex processing of red mud is cost-effective since red mud contains iron, aluminum, titanium, calcium, rare earth metals etc. It has been observed that the acid leaching of red mud can almost completely recover the rare earth elements in the solution with various individual techniques and also a combination of them. Therefore, the choice of extraction method depends on the form in which the element occurs in the solution. However, relatively low concentrations of rare earth in the solution and significant amount of impurities increase the cost of getting the final commercial products. To ensure the cost-effectiveness of the process involving rare earth’s extraction from red mud, it is necessary to increase their content by several times. This article presents the various studies that have been carried out in these aspects and the possibility of making this resource a sustainable one for REE extraction with a special focus on scandium replenishment.
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The recovery of vanadium from stone coal acid leaching solution by coprecipitation, alkaline roasting and water leaching was studied. A method used to recover vanadium from stone coal acid leaching solution was developed, and it primarily included vanadium coprecipitation with iron in the solution, vanadium extraction by alkaline leaching from the precipitate, and vanadium pentoxide preparation with the alkaline leaching solution. Experiments found that the vanadium in stone coal acid leaching solution can be effectively enriched in the precipitate obtained by adding 3.64 g NaClO3 per liter solution with initial pH 1.73 under stirring for 0.5 h at 95 °C. By roasting the mixture of 25 g of the precipitate with 22.5 g NaOH at 170 °C for 1.0 h, and then water leaching the roasted mixture at 98 °C for 1.0 h under stirring with L/S ratio of 3.3:1 mL/g, 97.0% of vanadium was extracted from the precipitate. After purifying with MgCl2, the vanadium pentoxide with purity 99.3% was obtained by adding NH4Cl to precipitate ammonium vanadate from the solution at pH about 2.0, and then roasting the ammonium vanadate at 520 °C for 2.0 h. The essential components of the coprecipitate are KFe3(SO4)2(OH)6 and HNaV6O16·4H2O.
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A D2EHPA and TBP mixed solvent system diluted with kerosene were used for the selective extraction of vanadium(IV) from iron(II) from an acidic stone coal leach solution that was reduced using sodium sulfite. Extraction studies were carried out under different pH and solvent concentrations, and optimized conditions were determined. The loaded organic was stripped with sulfuric acid. The number of stages required for the extraction and stripping of vanadium were determined from a McCabe-Thiele plot and confirmed by counter-current simulation studies. Results demonstrate that the extraction of vanadium increases as the initial pH and the D2EHPA concentration increases in the organic phase. A six-stage counter-current extraction simulation test was conducted over a period of 40h at a initial pH of 2.48 with 10% (v/v) D2EHPA and 5% (v/v) TBP mixed extractant and resulted in a vanadium extraction of 97% for the feed solution containing 5.78g/L V2O5 and 10.86g/L total Fe. The loaded organic phase that contained 5.34g/L V2O5 and 1.0g/L Fe can be completely stripped by three-stage counter-current stripping with 1.5mol/L H2SO4 at a phase flow ratio of O/A=5:1 to give a strip solution containing 26.3g/L V2O5 and 0.72g/L Fe.
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An effectively new technology of extracting vanadium from stone coals by high concentration sulfuric acid was researched. The effect of the concentration of sulfuric acid, leaching temperature, leaching time and helper leaching agent on the extraction of vanadium was explored. The results show that the optimal conditions of extraction are as follows: the concentration of sulfuric acid is 6 mol/L, the ratio of liquid to solid is 3:1; the temperature is 90 °C; the leaching time is 3–5 h, the diameter of the ore particle is less than 180 μm, and the concentration of helper leaching agent R is 6%. Under these conditions, the extraction of vanadium can reach 95.86%.
Article
During the recovery of vanadium pentoxide from vanadium containing sludge of alumina plants, huge amount of effluents containing NH3, P, V and NaCl, is generated. Due to its toxic nature, it can neither be discharged to land or water nor can be stored as such, hence a process has been developed to recover ammonia as salts and vanadium pentoxide to be recycled. Thus on using sulphuric acid, about 75% of ammonia was recovered as ammonium sulphate and while using HCl about 89% of ammonia was obtained as ammoniun chloride. On refluxing calcium vanadate with H2SO4 about 94% of V2O5 is obtained which may be recycled to the main stream of V2O5 plant.
Article
The loading of V on weak base resin D314 from sulphuric acid leach solutions of stone coal containing 2.06 g/L V (V2O5) was found to be 260 mg/mL with contact time of 60 min at pH=4, giving a recovery of 99%. Stripping of loaded resin was excellent using 3 mol/L NaOH, with a maximum V concentration of 150 g/L in the eluate. The resin features stable reusability when regenerated using 1 mol/L H2SO4 or 2 mol/L HCl. The loading of V on the resin in the industrial trial was found to be as high as 280 mg/mL with a maximum V concentration of 200 g/L in the eluate. The final product of V2O5 with high purity (>99%) meeting the standard specification was produced. Application of this process in industry is expected to make an important impact on the extraction of vanadium from the leach solution of stone coal and to deliver good economical benefits.
Article
This investigation resulted in the development of a flowsheet for the recovery of tungsten and vanadium from a leach solution of tungsten alloy scrap. The leach solution contained 95 kg/m3 WO3, 0.175 kg/m3 V and some other impurities. In the purification step, MgCl2 combined with Al2(SO4)3 was employed to remove the majority of Si, Al, P and Fe from the solution. This step was followed by converting tungsten and vanadium into the calcium salt with the addition of Ca(OH)2. When the mixed calcium salt was treated with formic acid solution, vanadium was selectively dissolved and separated from tungsten, while tungsten was rejected in the residue. The vanadium solution was then purified with H2SO4 and the ammonium meta-vanadate was precipitated with NH4OH to give 86.87% vanadium recovery. The leach residue of formic acid solution was dissolved with hot HCl solution in the presence of oxygen, resulting in the conversion of CaWO4 to H2WO4. This step was followed by NH4OH addition to convert H2WO4 to (NH4)2WO4 and removal of the remaining impurities. Finally, ammonium para-tungstate (APT) was crystallized from the solution of (NH4)2WO4 by evaporation. The purity of both tungsten and vanadium products met the standard specifications.
Article
Extraction of vanadium from black shale was attempted in pressure acid leaching. The chemical components of the sample obtained from Guizhou Province of China show that it contains 3.258% V2O5, 52.880% SiO2 and 16.140% Al2O3. Phase analyses of vanadium indicates vanadium mainly exists in the free oxide and mica. Vanadium contents in the two phases are 18% and 53%, respectively. The contents of V3+, V4+ and V5+ are almost equal. Under the optimum parameters of one-step leaching (reaction time of 3 h, sulfuric addition of 25%, temperature of 150 °C, liquid to solid ratio of 1.2 mL/g, catalyst (FeSO4) addition of 5% and size of 85% particle 0.074 mm), about 77% of vanadium is recovered. After two-step countercurrent leaching, the leach recovery of vanadium can reach above 90%. Air replacing oxygen is completely feasible. The impurity metals can dissolve into solution in different degrees.
Article
Due to numerous co-properties in chemistry of molybdenum and vanadium, it is relatively more difficult to remove vanadium from molybdate solution. To produce high-quality ammonium molybdate with the molybdenum materials containing vanadium, the separation of vanadium from the molybdate solution under batch and column mode using ion exchange technique with the strong base resin D296 was studied. In the solution, the pH value was about 7.2, the molybdenum concentration was in the range of 60–80 g/L, vanadium concentration was near 0.6 g/L, chloride ions 20 g/L. The concentration of vanadium (V2O5) was not more than 0.01 g/l in the effluent until effluent/resin volume ratio over 20 at 25±0.5 °C for contact time 60 min and, the average concentration of vanadium (V2O5) was less 0.008 g/L in the effluent collected till the breakthrough point 0.02 g/L V2O5. It was found that the separation can be only performed in the pH range of 6.5–8.5 and, chloride ions have an important influence on the separation as well. When Cl− concentration is increased near 70 g/L, it is impossible to remove vanadium from the solution with the resin. The loaded resin was stripped and regenerated using HCl 6 mol/L, the desorption ratio of vanadium was over 98.5%.
Article
A novel technology characterized by higher recovery of vanadium and which was environmentally-friendly was developed to recover vanadium from stone coal. Vanadium in stone coal could be leached by NaOH solution after roasting stone coal at 850°C for 3h. H2SO4, Mg(NO3)2 and ammonia were employed, respectively, in two steps to remove the impurities of Si and Al from the leach liquor. After extracting vanadium from the leach liquor with 10vol% N235, 20vol% secondary octyl alcohol and 70vol% sulfonated kerosene, 1.5mol/L NaOH was used as a stripping agent to strip vanadium from extracting solution. Adding 80g/L NH4NO3 to the stripping solution at 30–40°C and pH 7.5, vanadium could be crystallized as ammonium metavanadate. Roasting ammonium metavanadate at 540°C for 1h, the purity of V2O5 met the standard specification. The total recovery of vanadium reached 67.39%, which was higher than the classical technology.
Article
This paper reports on laboratory work for the development of a process flowsheet for the recovery of pure vanadium from sodium vanadate-containing sludge, typically analysing about 20 wt% V2O5. The sludge is generated as a by-product during the production of alumina from some bauxite ores by the Bayer process. The process developed consists essentially of (1) leaching of the sludge in hot water to solubilize the vanadium values, (2) adsorption of vanadium on activated charcoal, (3) desorption of vanadium with a suitable eluent, (4) precipitation of vanadium bearing cake from the strip solution, and (5) calcination of the cake to yield pure V2O5. The influence of various operational parameters on the process of adsorption, desorption and precipitation has been studied in detail. It has been possible to achieve high recovery of a 99.9% purity V2O5 product by allowing the adsorption to take place at a pH of around 2.5, desorption with 10% ammonia solution, and precipitation by acidification, all at temperatures around 85°C. Based on the experimental work, a schematic flow chart for the recovery of vanadium oxide from the vanadium bearing Bayer sludge has been presented.
Article
Extraction of vanadium from electrostatic precipitator (EP) ashes containing a large amount of ammonium sulfate has been studied under weakly acidic and reducing conditions. Vanadium was extracted 95% or more with 0.1 - 0.2 M H/sub 2/SO/sub 3/ at 323-363 K, with an ash/reagent ratio of 1/4, a H/sub 2/SO/sub 3//V/sub 2/O/sub 5/ molar ratio of 2, and an extraction time of 45 min. Dithionate ion was not detected in the extraction solution by the precipitation-atomic absorption method.
Article
A commercially available extractant, LIX®63 was used to investigate the extraction of molybdenum(VI) and vanadium(V) from a synthetic sulphuric acid leach solution of spent hydrodesulphurisation catalysts. The molybdenum(VI) and vanadium(V) were extracted and separated completely from other metals at pH 1.5. The loaded organic solution was easily stripped by NaOH solution with excellent phase separation. After the precipitation of most of the vanadium(V) as ammonium vanadate from the loaded strip liquor, a pure molybdate solution could be obtained by further removing the small amount of vanadium(V) remaining in the filtrate using Aliquat 336 at pH 8.5. Therefore, both pure molybdenum(VI) and vanadium(V) products can be obtained. The separation of nickel(II) and cobalt(II) from iron(III) and aluminium(III) in the raffinate after the recovery of molybdenum(VI) and vanadium(V) could be achieved by ion exchange with Dowex M4195 resin. A process flowsheet has been developed to recover the valuable metals from leach solutions of spent hydrodesulphurisation catalysts.
Article
Vanadium is an important by-product that is used almost exclusively in ferrous and non-ferrous alloys due to its physical properties such as high tensile strength, hardness, and fatique resistance. Vanadium consumption in the iron and steel industry represents about 85% of the vanadium-bearing products produced worldwide. The ubiquitous vanadium is employed in a wide range of alloys in combination with iron, titanium, nickel, aluminum, chromium, and other metals for a diverse range of commercial applications extending from train rails, tool steels, catalysts, to aerospace. The global supply of vanadium originates from primary sources such as ore feedstock, concentrates, metallurgical slags, and petroleum residues. Vanadium-bearing host minerals consist of carnotite, mottramite, patronite, roscoelite, and vanadinite. Deposits of titaniferous magnetite, uraniferous sandstone, bauxite, phosphate rock, crude oils, oil shale and tar sands host vanadium. Apart from titanomagnetite and ilmenite ore deposits containing vanadium, slags from the ferrous industry are a major source of supply. At present, known world reserves are expected to supply the next century’s needs. Vanadium-bearing materials are treated by means of several processes such as calcium reduction, roast/leach, solvent extraction and ion exchange to recover vanadium either as metal, ferrovanadium, vanadium pentoxide, or in the form of various chemicals. The recovery of aluminum and magnesium metal from smelters and refineries generates vanadium and associated compounds. Countries such as China, South Africa, and Russia are the largest world producers of ferrovanadium and its toxic oxides while about 40 other countries contribute smaller quantities in different forms for global consumption. Australia is poised to become a major player for this essential substance during the next decade. The supply and demand of vanadium products during the past 20 years has been relatively stable and subject to a gradual decline in delivered price. The paper describes established industrial processes for recovery of vanadium from sources such as raw ore and process reverts. The comprehensive condensation of pertinent facts is intended to provide a single reference source rather than the reader perusing many articles.
Article
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%.
Article
Molybdenum-containing catalysts are mostly used in petroleum-refining industries for mild hydrogenation processes. These catalytic processes generate huge quantity of spent catalyst. With the increasing demand of metal values and environmental awareness, catalysts can serve as a secondary source for metal recovery.Spent hydro-refining catalyst mainly consists of 10–20% MoO3, 6–8% CaO, 0–12% V2O5, 0.5–6% NiO, 10% S, 10–12% carbon and the balance is Al2O3. The spent catalyst was roasted with soda ash to convert molybdenum into a water-soluble compound. Various parameters like temperature, time and soda ash addition have been studied thoroughly and conditions for the maximum recovery of molybdenum have been established. It has been found that at 600 °C with 30 min of retention time and using 12 wt.% soda ash, it is possible to extract 92% of molybdenum from the spent catalyst. The molybdenum is extracted as sodium molybdate. This sodium molybdate is further purified using carbon adsorption method to obtain a pure grade of molybdenum trioxide.
Article
The hybrid ion exchanger consisted of PONF-g-GMA anion fibrous exchanger and IRA-96 bead-type anion exchanger was developed by combining different types of layers with hot-melt adhesive. Its ion exchange capacity and the pressure drop with flow rate of water were measured and the adsorption of vanadium (V) ions on the hybrid ion exchanger was evaluated with various process parameters such as pH, initial concentration, and temperature. It was observed that the adsorption kinetics of vanadium (V) ions on the hybrid ion exchanger could be analyzed with pseudo-second-order model.
Article
In order to reduce the environmental impact due to land disposal of oil fly ash from power plants and to valorize this waste material, the removal of vanadium was investigated using leaching processes (acidic and alkaline treatments), followed by a second step of metal recovery from leachates involving either solvent extraction or selective precipitation. Despite a lower leaching efficiency (compared to sulfuric acid), sodium hydroxide was selected for vanadium leaching since it is more selective for vanadium (versus other transition metals). Precipitation was preferred to solvent extraction for the second step in the treatment since: (a) it is more selective; enabling complete recovery of vanadate from the leachate in the form of pure ammonium vanadate; and (b) stripping of the loaded organic phase (in the solvent extraction process) was not efficient. Precipitation was performed in a two-step procedure: (a) aluminum was first precipitated at pH 8; (b) then ammonium chloride was added at pH 5 to bring about vanadium precipitation.
Article
The Puertollano Integrated Coal Gasification Combined Cycle (IGCC) Power Plant (Spain) fly ash is characterized by a relatively high content of Ga and V, which occurs mainly as Ga2O3 and as Ga3+ and V3+ substituting for Al3+ in the Al-Si fly ash glass matrix. Investigations focused on evaluating the potential recovery of Ga and V from these fly ashes. Several NaOH based extraction tests were performed on the IGCC fly ash, at different temperatures, NaOH/fly ash (NaOH/FA) ratios, NaOH concentrations and extraction times. The optimal Ga extraction conditions was determined as 25 degrees C, NaOH 0.7-1 M, NaOH/FA ratio of 5 L/kg and 6 h, attaining Ga extraction yields of 60-86%, equivalent to 197-275 mg of Ga/kg of fly ash. Re-circulation of leachates increased initial Ga concentrations (25-38 mg/L) to 188-215 mg/L, while reducing both content of impurities and NaOH consumption. Carbonation of concentrated Ga leachate demonstrated that 99% of the bulk Ga content in the leachate precipitates at pH 7.4. At pH 10.5 significant proportions of impurities, mainly Al (91%), co-precipitate while >98% of the bulk Ga remains in solution. A second carbonation of the remaining solution (at pH 7.5) recovers the 98.8% of the bulk Ga. Re-dissolution (at pH 0) of the precipitate increases Ga purity from 7 to 30%, this being a suitable Ga end product for further purification by electrolysis. This method produces higher recovery efficiency than currently applied for Ga on an industrial scale. In contrast, low V extraction yields (<64%) were obtained even when using extreme alkaline extraction conditions, which given the current marked price of this element, limits considerably the feasibility of V recovery from IGCC fly ash.
Qualitative Analytical Chemistry (Spanish)
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Natural Resources Canada, Minerals & Resources Sector, Canada Minerals Yearbookpp
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Perron, L., 2001. Vanadium. Natural Resources Canada, Minerals & Resources Sector, Canada Minerals Yearbookpp. 59.1-59.7.
Vanadium market in the world present status, price trends and future prospects
  • B V R Raja
Raja, B.V.R., 2007. Vanadium market in the world present status, price trends and future prospects. Steelworld 19-22 (February).