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Metal extraction from spent sulfuric acid catalyst through alkaline and acidic leaching
Abstract
Spent catalyst from manufacture of sulfuric acid production (main elemental composition: 3.5% V, 0.63% Ni, 7.9% Fe and 9.64% Si) can be used as a secondary source of vanadium and nickel. Extraction of these metals was studied using two different leaching systems (alkaline and acidic). Statistical design of the experiments and ANOVA (analysis of variance) were performed in order to determine the main effects and interactions of the factors under research, which were roasting, leaching temperature, concentration of the leaching reagent (H2SO4 and NaOH), liquid/solid (L/S) ratio (at 100 mL of liquid reagent), and presence of hydrogen peroxide.The results obtained after acidic treatment show that nickel extraction yield of 96% is achieved after roasting at 600 °C, followed by leaching with 5 mL/g 1 M sulfuric acid at 80 °C for a 30 min reaction time. The highest vanadium extraction yield was 59% after roasting at 400 °C and leaching at 80 °C by 0.3 M sulfuric acid for 6 h and 10 mL/g L/S ratio.A full factorial experiment was also performed by application of sodium hydroxide for vanadium extraction in the second leaching system. The highest vanadium extraction yield after alkaline treatment was 78%, obtained through roasting at 400 °C, leaching at 80 °C by 4 M NaOH for 2 h and 10 mL/g L/S ratio. Because nickel is not dissolved by sodium hydroxide, a sequential acidic leaching was conducted using the alkaline leaching residue, obtaining a nickel extraction yield of 88%.
... Hydrometallurgical processes are more preferred than pyrometallurgical methods due to their low energy consumption and low gas emissions and high recovery of metals [16]. In the hydrometallurgical processes, pretreatments such as oxidation and roasting are used to ensure that the precious metals pass into the solution in high efficiency [17,18]. However, leaching of the spent catalyst can be carried out directly in the presence of acidic or alkali solvents [16]. ...
... The dissolution rate of molybdenum was found to be 81.92% at 600°C roasting temperature and 120 min. However, since there was no significant increase in metal leaching efficiencies after 120 min, subsequent experiments were carried out at this roasting temperature and time [15,17]. ...
... Fig. 7. Effect of liquid/solid ratio on molybdenum leaching. This can be attributed to a reduction in the boundary layer thickness around the particle and an increase in the amount of metal passing into the solution due to an increase in the stirring speed [17]. The highest leaching values of molybdenum metal reached 94.75% at 600 rpm stirring speed. ...
... Hydrometallurgical processes are more preferred than pyrometallurgical methods due to their low energy consumption and low gas emissions and high recovery of metals [16]. In the hydrometallurgical processes, pretreatments such as oxidation and roasting are used to ensure that the precious metals pass into the solution in high efficiency [17,18]. However, leaching of the spent catalyst can be carried out directly in the presence of acidic or alkali solvents [16]. ...
... The dissolution rate of molybdenum was found to be 81.92% at 600°C roasting temperature and 120 min. However, since there was no significant increase in metal leaching efficiencies after 120 min, subsequent experiments were carried out at this roasting temperature and time [15,17]. ...
... Fig. 7. Effect of liquid/solid ratio on molybdenum leaching. This can be attributed to a reduction in the boundary layer thickness around the particle and an increase in the amount of metal passing into the solution due to an increase in the stirring speed [17]. The highest leaching values of molybdenum metal reached 94.75% at 600 rpm stirring speed. ...
... Normally potassium sulfate is used as promoter but in recent years also cesium sulfate has been used. The carrier material usually is silica in different forms [2]. ...
... These catalyst need to be appropriately stored and in way protected from weather conditions. The presence of sulfate, free sulphite and water in spent catalysts may result in an acidic leachate (pH about [1][2] and in environmental contamination with harmful compounds contained in those catalysts [1]. Vanadium and nickel are considered major sources of contamination. ...
... The toxicity of vanadium and its effect on health is highly dependent on its form. Vanadium pentoxide is one of the most toxic forms [2]. Moreover, it is slightly soluble in water [5]. ...
Vanadium has many industrial applications and its contribution to environmental contamination is growing every day. Spent catalyst used in the sulfuric acid production is considered a hazardous residue due to the presence of vanadium pentoxide which is the most toxic form of the vanadium. Because of this, currently, these wastes are incinerated. Thus, the removal of vanadium from catalysts is justified from environmental and economic viewpoints. The aim of this work is to evaluate the effect caused by the application of the electrokinetic remediation technique in the removal of vanadium of spent catalyst used in sulfuric acid industry when submitted to different conditions of electrical potential. The experiments were performed under 6.7 and 20.0 V m⁻¹ constant potential gradient. The results showed that the maintenance of a potential gradient of 20.0 V m⁻¹ resulted in an increase of 16% in vanadium removal but an increase in energy consumption per kg of remediated catalyst of about 80 times compared to the application of 6.7 V m⁻¹. It was also observed that the intensity of the applied electric potential modify the pH of the medium.
... A number of hydro-and pyro-metallurgical processes have been proposed for metal recovery from secondary raw materials. The metals are recovered as mixed solutions and then separated through conventional separation techniques (solvent extraction, selective precipitation and ion-exchange) (Ognyanova et al. 2009). ...
... It is well known that vanadium can be extracted from solution by hydrolysis as V 2 O 5 (Muzgin and Khamzina, 1981). Some authors (Vegliծ et al. 2006;Ognyanova et al. 2009;Erust et al. 2016) reported about using hydrogen peroxide (H 2 O 2 ) as an oxidizing agent for vanadium leaching, leading to an intensification of the V 2 O 5 hydrolytic extraction. Our previous studies showed the possibility to use H 2 O 2 for oxidation of SVC leaching solutions and further extraction of V 2 O 5 from solutions by thermohydrolysis with boiling time 5 min (Orehova et al. 2013). ...
Spent vanadium catalysts of sulfuric acid production (main elemental composition in wt%: 7.5 V, 9.1 K, 10.2 S, 23.2 Si and 1.4 Fe) can be used as a secondary source of vanadium. Extraction of vanadium was studied using two-step leaching (acidic and reductive) of spent vanadium catalysts with further oxidizing of leaching solutions. The factors leaching and hydrolysis temperature, concentration of leaching (H2SO4, Na2SO3) and oxidizing ((NH4)2S2O8) reagents, solid/liquid ratio, mixing parameters, and time of leaching and thermohydrolysis were systematically investigated. The solubility of V2O5 was investigated as a function of temperature, pH of sulfuric acid solutions, and concentration of Na2SO3. The kinetics of V2O5 solubility and reduction were also studied. The vanadium leaching yield after a two-step recovery was 98 wt% after acidic (H2SO4, pH 1.2–1.3) leaching with ultrasonic treatment for 5 min at ambient temperature, followed by reductive leaching in 0.01 Mol/L Na2SO3 solution for 15 min at ambient temperature. The highest vanadium extraction yield from leaching solutions was 98 wt% obtained through oxidizing of leaching solutions by 30 wt%. (NH4)2S2O8 with a molar ratio n(V2O5)/n((NH4)2S2O8) of 5/1 for a reaction time of 5 min at 80–90 °C. the extracted vanadium product was V2O5 with a purity of 85–87 wt%. The technological scheme has been developed to recycle all obtained products and sub-products
... 44 Keywords: Vanadium-bearing spent catalyst; Vanadium leaching; leaching efficiency; 45 potassium ferric sulfate 46 47 48 49 50 51 52 J o u r n a l P r e -p r o o f 65 vanadium from spent catalysts and other similar vanadium-bearing resources [4][5][6][7] . The 66 proposed methods can be classified into three groups: alkali leaching [2, 8] , acid 67 leaching [9][10][11][12][13][14] , and bioleaching [15][16][17][18][19][20][21] . Alkali leaching includes two steps: roasting, 68 followed by leaching with alkali solution or ammonia water [22, 23] ; however, spent 69 catalyst from the sulfuric acid industry has a high sulfur content, which will produce 70 SO 2 gas in the roasting process, resulting in secondary pollution; Besides, high 71 content of silica in the spent catalyst causes high alkali consumption. ...
... In the process of sulfuric acid production, 55 vanadium-bearing catalyst is used for the transformation between SO 2 and SO 3 [1] . 56 After a period of use, some components are deactivated resulting from sintering, 57 thermal aging, or poisoning; thus, the catalyst loses its function and is replaced, 58 thereby creating spent catalyst [2] . Storage of spent catalyst would occupy vast 59 expanses of land; furthermore, the vanadium in spent catalyst dissolves easily in water, 60 which would result in heavy metal pollution [3] . ...
The sulfuric acid industry produces large volumes of spent catalyst containing vanadium. To develop an efficient method for vanadium recovery, hydrogen peroxide was used to oxidize sulfur in the spent catalyst, forming sulfuric acid, which, in turn, leached the metals from the catalyst. Experimental results showed that the leaching efficiency of vanadium gradually increased with hydrogen peroxide concentration; other parameters had minor effects. The hydrogen peroxide concentration, stirring rate, and pH had negative effects on leaching percent of potassium. Leaching efficiency of iron decreased with an increase in leaching time, stirring rate, and pH; with the increase in hydrogen peroxide concentration and liquid:solid ratio, leaching efficiency of iron first increased and then decreased. The optimum leaching conditions for vanadium were hydrogen peroxide concentration of 148 g/L, leaching time of 5 min, leaching temperature of 25 °C, stirring rate of 300 r/m, and liquid:solid ratio of 3:1, and leaching efficiencies of vanadium, potassium, and iron were 92.2%, 48.2%, and 46.3%, respectively. For potassium and iron, the leaching efficiencies of 70.9% and 70.5% were achieved using pure water leaching under the conditions of leaching time of 30 minutes, liquid to solid ratio of 4:1, and the temperature of 25 °C.
... Therefore, (bio)hydrometallurgical approaches are believed to be more environment-friendly due to low-energy demand. The reagents used in hydrometallurgical studies were both inorganic and organic, acidic, and alkaline [23][24][25][26][27]. In particular, organic acids such as citric acid might be used in metal extraction due to their environmental safety, natural origin, and high degradability [25,26]. ...
... The vanadium ions presumably did not form any complexes with citric acid, but are just dispersed in the solution, in which they are stable due to low pH [45,46] and bonded with one another by oxygen bridging [47]. Therefore, it may be easy to precipitate vanadium ions using, e.g., ammonium salts or simply by increasing the pH of the solution, as proved by other authors [24,25]. The V 2 O 5 concentration after CA leaching (Table 3) was determined at 1940 mg kg −1 , which is equal to 1087 mg kg −1 of V. When one compares this value to the initial V concentration (30,836 mg kg −1 ), the leaching yield is at the level of 96.5%, which is even higher than determined by ICP-MS (93%). ...
Spent catalysts being considered hazardous wastes exhibit a high metal content in mobile forms. In addition, growing demand for circular economy policy applications requires proper utilization of these wastes. This study aimed at the assessment of vanadium leaching from spent desulfurization catalyst derived from sulfuric acid plant dump located nearby a copper smelter. Chemical and phase composition of the catalyst has been characterized. The extraction has been performed using chemical (0.1-M and 1-M citric acid) and biological (biotic solution with Acidithiobacillus thiooxidans) methods, using different experimental parameters (pulp density, particle size, leaching time) to observe V leaching behavior and kinetics. The results revealed that both citric acid and bacteria carried out the extraction process well. The optimal parameters for acid leaching were < 0.2-mm particle size and 2% pulp density, which allowed to leach out 95% of V from spent catalyst within 48 h. The bacterially mediated extraction resulted in 93% V leached out within 21 days with 2% pulp density. The experiments showed that V present in the catalyst is susceptible to bioleaching and organic acid leaching with the latter being a quicker process.
... [19] Therefore, pyrometallurgical pre-treatments such as oxidation and roasting are applied to the spent HDS catalyst to ensure that precious metals are passed into the solution with high efficiency. [5,6,[30][31][32][33] There are extensive studies on combination of hydro-and pyro-metallurgical processes for recovery metals from the spent HDS catalysts in the literature; however, the number of studies carried out in the presence of alkaline reactants is limited. The aim of the present paper is to examine the leaching behaviors and affecting parameters on the leaching process in the presence of KOH as an alkali leaching agent. ...
... Therefore, the solubility of Mo and Al is high while the solubility of Co and Ni is quite low for these leaching conditions as above-mentioned in the reactions. [30,38] For this reason, the amounts of Mo and Al passing into the solution were investigated in this study. ...
The aim of present study is to examine the dissolution and kinetics of valuable metals from spent hydrodesulfurization catalyst using potassium hydroxide. Due to low solubility of metal oxides of Ni and Co in alkali medium, only the recovery of Mo and Al was investigated. Optimum roasting temperature and time were determined 600°C and 60 min., respectively. The maximum dissolution yields of Mo (83.54%) and Al (36.89%) were reached under optimum leaching conditions. The leaching behavior of both metals is represented by the shrinking core model and liquid film diffusion is controlled the process. Activation energies found to be 14.71 kJ/mol and 17.84 kJ/mol for Mo and Al, respectively.
... Supported vanadium oxide catalysts have been shown to be active for the oxidation of SO 2 in sulfuric acid industry [1,2]. Typically, the catalyst is composed of vanadium pentoxide (4-10 wt%) and promotive oxides or pyrosulfates/sulfates of alkaline metals such as potassium (15-20 wt%), sodium (2-5 wt%) and/or cesium (5-15 wt%) supported on SiO 2 [3,4]. ...
... So far, most processes for treating spent vanadium oxide containing catalysts have been performed toward recovery of vanadium pentoxide by using a 3-step approach, involving acid leaching, oxidation and precipitation [2,[23][24][25]. Drawbacks of this pathway to the situation of spent sulfuric acid catalysts lay in the low content of vanadium to be recovered from the spent catalyst, operation requirement at elevated temperatures with large amount of high concentration acid solution as well as recovery efficiency versus vanadium oxide purity trade-off. ...
In this work, the efficiency of the regeneration process of spent V2O5 catalyst from sulfuric acid plant under different atmospheres (5%O2/N2 or air) was evaluated. Temperature-programmed results showed that the observed reduction profiles of the samples are attributed to the reduction of amorphous V⁺⁵ and low-valence V+5−x species at low temperatures followed by the reduction of their crystalline structures at high temperatures. Significantly low values of SO2 conversion of the spent samples can be explained by the significant drop in quantity of all vanadium species, coupled with their structural change to more thermally stable forms. It was found that the exposure of the spent catalyst to 5%O2/N2 stream at 550 °C for 1 h allowed at first the re-oxidation of amorphous low-valence V species and second the dissolution of crystalline low-valence V species, thus resulted in recovery of their catalytic activity for SO2 oxidation. However, the regeneration in air was less effective than in 5%O2/N2 stream. This is supposedly due to the differential behaviors of the spent sample in different oxidative streams toward re-oxidizing low-valence V species and re-dissolving V precipitates.
... The general mechanism for direct leaching of metal oxides with strong inorganic acids is based on the fact that the oxides shell and core become rapidly hydroxylated, followed by the successive adsorptions of hydrogen ions, anions of the acid and again hydrogen ions at hydroxylated sites. [28] In our case, when leaching a solid (mineral sludge) with liquid (strong inorganic acid), the desired solid (metals) is extracted to the liquid phase while unleached solid (mineral sludge with unleached metal minerals) remains. The removal of the solid as the liquid penetrates into the particle leads to shrinking of a diameter of unleached core with time. ...
... [31] This depends on the proportion of the soluble constituent present, its distribution throughout the original solid, the nature of the solid and the original particle size. [28] Due to the strong corrosive potential of inorganic acids the solid phase may be easily dissolved, therefore solubilizing the target metals. If the soluble material is surrounded by a matrix of insoluble matter, the solvent must diffuse inside to contact and dissolve the soluble material and then diffuse out. ...
This study investigated the leaching yields of Mo, Ni and Co from a mineral sludge of a metal recycling plant generated by rainfalls. The investigated mineral sludge had a complex heterogeneous composition, consisting of particles of settled soil combined with the metal bearing particles (produced by catalysts, metallic oxides and battery recycling). The leaching potential of different leaching reagents (stand-alone strong acids (HNO3 (68%), H2SO4 (98%) and HCl (36%)) and acid mixtures (aqua regia (nitric + hydrochloric (1:3)), nitric + sulfuric (1:1) and nitric + sulfuric + hydrochloric (2:1:1))) was investigated at changing operational parameters (solid - liquid (S/L) ratio, leaching time and temperature), in order to select the leaching reagent which achieves the highest metal leaching yields. Sulfuric acid (98% H2SO4) was found to be the leachant with the highest metal leaching potential. The optimal leaching conditions were a three stage successive leaching at 80 °C with a leaching time of 2 h and S/L ratio 0.25 g L-1. Under these conditions, the achieved mineral sludge sample leaching yields were 85.5, 40.5 and 93.8% for Mo, Ni and Co, respectively. The higher metal leaching potential of H2SO4 in comparison with the other strong acids/acid mixtures is attributed to the fact that H2SO4 is a diacidic compound, thus it has more H+ ions, resulting in its stronger oxidising power and corrosiveness.
... Various attempts have been made to remove these metals by conventional hydrometallurgical processes that have employed strong acid (8 M H 2 SO 4 ), strong alkali (4 M NaOH), salt roasting with Na 2 CO 3 or NaCl followed by water or Na 2 CO 3 (30 g/L) leaching (Kar et al., 2005;Ognyanova et al., 2009;Park et al., 2007). Different pyrometallurgical techniques such as smelting, calcination, and anhydrous chlorination have also been used to recover metals from spent catalyst (Kar et al., 2005). ...
... A lower Ni yield during alkali leaching of spent sulfuric acid catalyst has been reported (Ognyanova et al., 2009). In that study, the residue was further subjected to acid leaching due to the lower yields of Ni under alkaline conditions. ...
... Various hydrometallurgical processes have been used to remove metal from spent catalyst. The hydrometallurgical processes have used high concentration of acid (8 M H 2 SO 4 ) and alkali (4 M NaOH) for removing metals from spent catalyst (Ognyanova et al., 2009;Park et al., 2007). Although, hydrometallurgical processes have shown reasonable metals extraction efficiencies, the use of high strength acids and alkali, secondary pollution and expensive downstream processing has restricted their usage on a larger scale. ...
... Ni was not leached out, whereas low Al yield was obtained using alkali leaching in comparison with bioleaching from both the AS and RS. The low Ni yield during alkali leaching has also been reported in another study conducted in alkali solution (NaOH) using spent sulfuric acid catalyst (Ognyanova et al., 2009;Villarreal et al., 1999). In contrast, leaching of Mo increased substantially in both the AS and RS during second stage alkali leaching. ...
The effect of sequential leaching such as bioleaching followed by bioleaching and bioleaching followed by chemical leaching is aimed at enhancing metal (Mo, Ni, V and Al) dissolution from a differently pretreated (acetone washed/decoked) spent catalyst. The X-ray photoelectron spectroscopy characterization of spent catalyst samples suggested the presence of metals in their oxide and sulfide forms. Bioleaching followed by bioleaching with either Acidithiobacillus thiooxidans (Ni-100%, Al-55%, Mo-81% and V-100%) or Acidithiobacillus ferrooxidans (Ni-94%, Al-55%, Mo-77% and V-99%) significantly enhanced removal of Al, Ni, and V from acetone washed (AS) spent catalyst compared to decoked spent catalyst (RS). In contrast, bioleaching using either A. thiooxidans or A. ferrooxidans followed by alkali leaching remarkably enhanced removal of Mo from both AS and RS, although higher yields were achieved using AS. Bioleaching using A. thiooxidans followed by alkaline leaching is an optimum strategy yielding a maximum of 96% Mo in 125 h from AS. A field emission scanning electron microscopic study revealed only minor stretches of Mo in the treated AS.
... In recent years, different physicochemical processes such as pyrometallurgy and hydrometallurgy have been employed to remove the metals from spent catalyst. [2,3] The use of hydrometallurgy has been found to be effective, but also poses major drawbacks. Pyrometallurgical processes are energy intensive and also result in the emission of harmful gases. ...
... [9] The statistical significance of the model equations (Eqs. [2][3][4][5] and the model terms were assessed by the F-test for analysis of variance (ANOVA) is presented in Table 4. A P-value < 0.05 indicates that the model was statistically significant at the 5% level. ...
A central composite design (CCD) combined with response surface methodology (RSM) was employed for maximizing bioleaching yields of metals (Al, Mo, Ni, and V) from as-received spent refinery catalyst using Acidithiobacillus thiooxidans. Three independent variables, namely initial pH, sulfur concentration, and pulp density were investigated. The pH was found to be the most influential parameter with leaching yields of metals varying inversely with pH. Analysis of variance (ANOVA) of the quadratic model indicated that the predicted values were in good agreement with experimental data. Under optimized conditions of 1.0% pulp density, 1.5% sulfur and pH 1.5, about 93% Ni, 44% Al, 34% Mo, and 94% V was leached from the spent refinery catalyst. Among all the metals, V had the highest maximum rate of leaching (Vmax) according to the Michaelis-Menten equation. The results of the study suggested that two-step bioleaching is efficient in leaching of metals from spent refinery catalyst. Moreover, the process can be conducted with as received spent refinery catalyst, thus making the process cost effective for large-scale applications.
... Table 2 lists the best metal leaching agent and the best experimental conditions for leaching discovered in the literature. [59][60][61][62][63] In the reaction container, the leaching solution (225, 375, and 525 ml) was charged with the specific amount of spent catalyst (15 g) and HNO 3 (150, 250, and 350 ml) + H 2 SO 4 (75, 125, and 175 ml). A syringe filter (Whatman Puradisc 4, 4 mm diameter, 0.45 m, PTFE) was used to separate the solution once the required time had passed, and the weight % of Ni and V in the residue was determined by X-ray fluorescence (XRF). ...
The spent desulfurization catalysts are considered as a major secondary source of valuable metals. The content of nickel and vanadium present in these catalysts, accompanied by environmental rules, have attracted the scientists to explore diverse options for their effectual processing. The electrometallurgy recovery of Ni and V from the spent desulfurization Ni‐Mo‐V/Al2O3 catalyst is described in this study. Using flat plate graphite electrodes, the electrochemical deposition of Ni and V from spent catalyst in acid solution (HNO3/H2SO4) was investigated. By the central composite design of the response surface methodology, the effect of the operating factors was examined and optimized. At the ideal conditions of reaction temperature of 84.0 and 42.0 °C, electrolysis time of 5.6 and 4.4 h, liquid/solid ratio of 22.7 and 15.4 ml/g, and current density of 229.0 and 255.6 A/m2, respectively, the recovery efficiency of Ni and V was 81.96 and 93.07%. The statistical analysis revealed that the expected data (R2=0.9984, and R2=0.9883) were in good agreement with the observed data (R2=0.9984), with an average variation from experimental data of 0.78 and 0.65% for the optimum conditions of Ni and V recovery, respectively. It shows that the Ni and V nanoparticles deposited have a spherical form with purity of 84.39 and 90.76%, respectively. Because of its great efficiency and purity, the current study can provide a dependable procedure for extracting Ni and V from solid waste.
... Due to the cheap operational and energy costs, hydrometallurgical processing is frequently chosen as the preferred method for recovering nickel and zinc from fly ash [25]. Direct acidic leaching is used to dissolve almost all valuable metals into the solution from various types of solid wastes and then extract these metals separately by appropriate methods [41][42][43][44][45][46][47][48][49][50][51][52][53][54][55][56]. ...
... Due to the cheap operational and energy costs, hydrometallurgical processing is frequently chosen as the preferred method for recovering nickel and zinc from fly ash [25]. Direct acidic leaching is used to dissolve almost all valuable metals into the solution from various types of solid wastes and then extract these metals separately by appropriate methods [41][42][43][44][45][46][47][48][49][50][51][52][53][54][55][56]. ...
The water leaching solid residues (WLSR) obtained from salt-roasting Egyptian boiler ash are considered an essential secondary resource for (13%) nickel and (5.6%) zinc extraction. Hence, the current study aims for the cost-effective and high purity Ni, Zn, Fe and Mg metal ion extraction from (WLSR) using a sulfuric acid leaching process. The factors affecting the percentage recovery of Ni, Zn, Fe and Mg from WLSR, including leaching temperature, time, acid concentration and solid/liquid ratio, have been investigated. The obtained leaching solutions were analyzed chemically using ICP, and the different precipitates were analyzed mineralogically using XRD and EDX analysis and chemically using XRF. The maximum percentage recovery of Ni, Zn, Fe and Mg was 95.02%, 90.13%, 66.29% and 75.73%, which was obtained under the optimum leaching conditions of 8% H2SO4 concentration and 1/15 solid/liquid ratio at 85 °C for 240 min. The effect of pH, Fe2O3 dosage as nucleating agent and the precipitation duration on iron removal and Ni and Zn loss have been thoroughly studied. It has been found that >95% of the contained iron impurity can be removed , while nickel and zinc losses are around 4.2% and 3.8%, respectively. Additionally, a pH of 6 and 0.45 mol/L concentration of H2C2O4 was utilized to precipitate Mg as MgC2O4.2H2O, demonstrating that the precipitation efficiency of Mg reaches 96.9%. Nickel and zinc precipitation efficiency was 92.25% and 85.51%, respectively, by raising the solution pH to approximately 9. The kinetic of Ni and Zn dissolution has been investigated to explain the mechanism prevalent and the factors influencing the leaching process. It has been found that the nickel leaching kinetic is controlled by both diffusion through an inert porous layer and by chemical reaction with an activation energy of 20.25 kJ.mol −1. Meanwhile, the kinetic of zinc leaching is controlled by solid product layer diffusion with an activation energy of 11.67 kJ mol −1 .
... Hydrometallurgical processes are more preferred than pyrometallurgical method because of low energy consumption and low gas emission and recovery of metals by high efficiency [15]. In hydrometallurgical processes, pre-treatments such as oxidation and roasting are applied to the spent catalyst to ensure that precious metals are passed to the solution in high yields [16,17]. However, leaching process of the spent catalyst may be directly carried out in the presence of acid or alkali leaching agents [15]. ...
The spent hydrodesulphurization (HDS) catalyst is an important secondary source for Ni, Mo, Co and Al metals. The high yield recovery of these metals is quite difficult due to the carbon accumulated on the catalyst surface and the stability of the metal oxides. Therefore, the leaching process in the presence of sodium persulfate (Na2S2O8) was carried out after the roasting pre-treatment to remove the carbon from the spent HDS catalyst structure and convert the metal oxides to the soluble form in this study. The optimum experimental conditions were determined as roasting temperature, 500 °C; roasting time, 120 min.; particle size, +75–30 μm; liquid/solid ratio, 12.5 ml/g; Na2S2O8 concentration, 0.4 M; leaching temperature, 50 °C; leaching time, 90 min and stirring speed, 400 r/min. Recovery of Mo (89.8%), Co (86.5%) and Ni (81.2%) from leach solution were achieved by precipitation method. The liquid film diffusion control mechanism best represents the proposed leaching process. On the other hand, the magnitude of Ea values (<20 kJ/mol) for Mo, Co, Ni and Al metals indicates that the leaching process is controlled by liquid film diffusion mechanism.
... 15% H 2 SO 4 0.5 60 10 70% V [7] Spent catalyst (V = 3.5, Ni = 0.63, Fe = 7.9, Si = 9.64, Al = 0.07, P = .08, S = 3.62) 1.5% H 2 SO 4 6 80 10 60% V [8] 18% vanadium, with low intensities of iron and aluminum. ...
This work aims to develop a green method for the recovery of vanadium from spent contact process catalyst (SVC) with 6-7% V 2 O 5. The efficiency of leaching was evaluated using oxalic acid by varying pulp density, time, particle size, and temperature. It is found that with 0.25 M of oxalic acid, almost 98% V, 42% Fe and 24% Al was recovered in 2 h at 323 K and pulp density of 20% using <50 µm particle size. The leaching kinetics and mechanism was further established by characterizing the feed and residues by XRD and SEM. Separation of vanadium was performed using ion-exchange and precipitation. ARTICLE HISTORY
... 15% H 2 SO 4 0.5 60 10 70% V [7] Spent catalyst (V = 3.5, Ni = 0.63, Fe = 7.9, Si = 9.64, Al = 0.07, P = .08, S = 3.62) 1.5% H 2 SO 4 6 80 10 60% V [8] 18% vanadium, with low intensities of iron and aluminum. ...
This work aims to develop a green method for the recovery of vanadium from spent contact process catalyst (SVC) with 6–7% V2O5. The efficiency of leaching was evaluated using oxalic acid by varying pulp density, time, particle size, and temperature. It is found that with 0.25 M of oxalic acid, almost 98% V, 42% Fe and 24% Al was recovered in 2 h at 323 K and pulp density of 20% using <50 µm particle size. The leaching kinetics and mechanism was further established by characterizing the feed and residues by XRD and SEM. Separation of vanadium was performed using ion-exchange and precipitation.
... When using three leaching stages with 15% of sulphate acid (100°C and solid/liquid ratio of 0.2), oxidative precipitation, and extraction with 60-70% of HNO2 (120°C) (Khorfan & Reda, 2001), a recovery efficiency in vanadium around 70% is obtained. Another research for separating nickel and vanadium from the spent catalyst for making sulphate acid shows that acidic leaching (0.3-1M H2SO4) provides recovery efficiency in vanadium of 59% when combined with alkaliacidic leaching (4 M NaOH, continued with 0.5 M H2SO4) to reach 78% (Ognyanova et al., 2009). The use of oxalate acid 0.50 M and 0.66 M H2O2 to process the spent catalyst Ni-Mo/Al2O3 is known to provide vanadium yield of 80% at a temperature of 70°C (Szymczycha-Madeja, 2011). ...
Utilization of spent catalysts serves to meet the needs of vanadium and to overcome the environmental problem since vanadium is categorized as a hazardous, toxic material. Vanadium from the spent catalysts can be recovered in the form of V2O5 or NH4VO3. However, vanadate compounds (NH4VO3) are considered to be more valuable due to their higher price and easier production process, compared with V2O5. This study aims to find adequate operating conditions to obtain high yields and high purity of NH4VO3 crystals. The results showed that the presence of iron compounds in the extract made the crystals contaminated by brownish colour, so it decreased its purity. Therefore, iron compounds need to be separated first with precipitation. Crystals of NH4VO3 with yield of 60% on spent catalysts and purity of 75% were obtained by extraction using solvent Na2CO3 1.887 M for 60 minutes at room temperature with weight ratio of V2O5 in spent catalyst toward solvent volume (Rvp), 0.006 gram V2O5/mL Na2CO3. It was then continued by precipitation of iron compounds at pH of 12 for 2 hours and crystallization of NH4VO3 using NH4Cl 11.215 M for 4 - 5 hours at 60oC.
... In the above-mentioned hydrogenation process, alumina-supported catalysts, containing 5-14% Mo and 1-4% Co, are widely generally used. These catalysts have a very high resistance to degradation (poisoning) and can be easily regenerated and maintain their catalytic activity for a long time (Ognyanova et al. 2008). ...
In present study, the leaching kinetics of the spent Mo–Co–Ni/Al2O3 catalyst was investigated in the presence of formic acid as an organic leaching agent. Firstly, the spent catalyst was roasted in different roasting temperature (200–700 °C) and time (15–240 min), the maximum metal extraction was achieved that at 500 °C with 90 min. Then, the leaching experiments were carried out to determine the influences of process parameters following; particle size, liquid/solid ratio, formic acid concentration, leaching temperature, leaching time and stirring speed. According to the experimental results, the highest dissolution rates of molybdenum (Mo, 75.82%), cobalt (Co, 96.81%), nickel (Ni, 93.44%) and aluminum (Al, 19.46%) were reached under optimum experimental conditions; particle size +75 − 30 µm; liquid/solid ratio 10 ml/g; formic acid concentration 0.6 M; leaching temperature 80 °C; leaching time 90 min and stirring speed 300 r/min. Moreover, the leaching kinetics clearly reveal that the leaching reaction is controlled by liquid film diffusion and that the activation energy values (Ea) of Co, Ni, Mo and Al were to be 24.49, 25.98, 32.36 and 33.47 kJ/mol, respectively. In conclusion, the leaching process can be conducted in the presence of formic acid for the various industrial wastes in similar structure and composition to Mo–Co–Ni/Al2O3 spent catalyst.
... The hydrometallurgical processes have shown high metal extraction efficiencies; however, these processes use a high concentration of sulfuric acid as 8 M and alkali as 4 M NaOH, which represents a secondary pollution with a consequent expensive treatment. [5] Pyrometallurgical techniques such as smelting, calcination, and anhydrous chlorination have also been employed to recover metals from spent catalysts with high efficiencies; however, these processes have enormous consumption of energy and emit toxic gases to the environment. [6] Bioleaching has demonstrated high efficiency of metal recovery from rocks with a low energy requirements and low generation of wastes. ...
The aim of the present study was to isolate microorganisms able to tolerate Ni²⁺ and V⁵⁺ from different sites located close to a mineral mine in Guanajuato, Mexico, and then to evaluate their ability to remove metals contained in a spent catalyst. Seventeen isolates were obtained; among them seven presented a minimum inhibitory concentration (MIC) higher than 200 mg/L of Ni²⁺ and V⁵⁺ each. Nickel and Vanadium removal was evaluated in 9 K liquid medium added with spent catalyst at 16% (s/v) pulp density and incubated at 30 °C, 150 rpm for 7 days. Only three isolates which were coded as PRGSd-MS-2, MNSH2-AH-3, and MNSS-AH-4 showed a significant removal at the end of treatment corresponding in mg kg⁻¹ (or percentage metal removal) of 138 (32%), 123 (29%), and 101 (24%) for Ni, respectively; and 557 (26%), 737 (34%), and 456 (21%) mg kg⁻¹ for V, respectively. The same isolates were capable to remove also Al, Fe, As, and Mg at different extent. Cell morphology changes were observed, in comparison to the control system at the end of biological treatment as a higher quantity of spores for MNSH2-AH-3, 2 μm cells in pairs for MNSS-AH-4, also long chain-vegetative cells having inclusions into the cell surface were observed for PRGSd-MS-2. The three isolated microorganisms were identified by sequencing of the 16S gene as Bacillus thuringiensis, Bacillus megaterium, and Bacillus sp, respectively, suggesting its potential use in the treatment of this solid industrial waste.
... Leaching at atmospheric pressure has been recently receiving more attention despite its less efficiency, longer reaction time and higher acid consumption. In this framework, conventionally, sulfuric acid is used as leaching reagent for the recovery of nickel from catalysts [12][13][14][15] because of its lower corrosion property and lower cost. In fact, Ivascanu and Roman [16] recovered 99% of Ni from a spent catalyst in the form of nickel sulfate by leaching with sulfuric acid under the following conditions (sulfuric acid concentration of 80%, 50 min of reaction and 70°C). ...
In this study the leaching of NiO from NiO/α-Al2O3 catalyst in acidic (HCl, HNO3 and H2SO4) and ammoniacal ((NH4)2CO3, CH3COONH4) media was investigated. The effects of leachant concentration, liquid/solid ratio, stirring speed and temperature were studied. It was found that 100% of nickel was dissolved after 30min of reaction with HCl at 2M, 80°C and liquid/solid ratio of 50mL/g, while HNO3 and H2SO4 dissolved 77.15% and 46.12% respectively under the same operation conditions. Mixing two strong acids led to a synergetic effect on nickel leaching at the beginning of the reactions followed by a rapid stabilization in dissolution. Ammoniacal leaching was less efficient than acidic one registering 41.43% with ammonium carbonate and 29.16% with ammonium acetate after 180min. However, the addition of chlorides to ammonium carbonate led to totally dissolve NiO.
... Spent NH 3 -SCR catalyst generally contains titanium oxide and metals such as W and V in appreciable concentrations, and thus could be used as a secondary source for these valuable metals. In general, the recovery of valuable metals from spent catalysts has been carried out by conventional methods [9][10][11][12][13], such as acid leaching, caustic leaching, and salt roasting followed by leaching with water. Among these, caustic leaching seems to be the best option for the recovery of SCR catalyst, since WO 3 is insoluble in acid. ...
The recovery of tungsten (W) from a honeycomb-type spent selective catalytic reduction (SCR) catalyst using an alkaline leaching–ion exchange method was investigated. Spent SCR catalyst mainly consists of TiO2 and other oxides (6.37% W, 1.57% vanadium (V), and 2.81% silicon (Si), etc.). The ground catalyst was leached at the optimal conditions, as follows: NaOH concentration of 0.3 kg/kg of catalyst, pulp density of 3%, leaching temperature of 70 °C, particle size of −74 µm, and leaching time of 30 min. In this study, the leaching rate values of V andWunder the above conditions were 87 wt %, and 91 wt %, respectively. The pregnant solution was then passed through a strong base anion exchange resin (Amberlite IRA900). At high pH conditions, the use of strong base anion exchange resin led to selective loading of divalent WO4²⁻ from the solution, because the fraction of two adjacent positively-charged sites on the IRA900 resin was higher and separate from the coexisting VO4³⁻. The adsorbed W could then be eluted with 1 M NaCl + 0.5 M NaOH. The final concentrated W solution had 8.4 g/L of W with 98% purity. The application of this process in industry is expected to have an important impact on the recovery of W from secondary sources of these metals.
... Some studies indicate chemical leaching using inorganic or organic acids (Ognyanova et al., 2009; Mirsha et al., 2010; Srichandan et al., 2014a) and microbial leaching (Jadhav & Hocheng, 2012; Noori et al., 2014) as promising alternatives for the recovery of metals from spent catalysts. However, these works report chemical or biotechnological processes from oil refining spent catalysts pretreated with operations such as washing with solvents, heat treatment, and milling or sieving (Srichandan et al., 2014b; Thosar et al., 2014; Bharadwaj & Ting, 2013). ...
The aim of this work is to evaluate the recovery of metals from a spent catalyst of diesel hydrodesulfurization (HDS), without any chemical, thermal, or physical pre-treatment, using the Acidithiobacillus thiooxidans FG-01 strain and chemical leaching. The spent catalyst is a hazardous waste due to the presence of organochlorine compounds, petroleum hydrocarbons, and heavy metals. Chemical leaching tests were performed with solutions of citric acid (20 g L-1), oxalic acid (20 g L-1), citric and oxalic acid mixture (1:1, 20 g L-1), and sulfuric acid (10 g L-1). The best results of metals recovery through chemical reaction were obtained when using the inorganic acid solution. The two-step bioleaching assays were conducted in 500 mL Erlenmeyer flasks, with a pulp density of 27.5 g L-1, and a sulfuric acid concentration of 20 g L-1. There were no differences between chemical and microbial leaching results. The A. thiooxidans FG-01 is a promising strain due to the recovery of Al (~65%), V (~25%), and Mo (~25%) from the raw solid residue.
... SO 2 is converted to sulphur trioxide (SO 3 ) through a reversible reaction where vanadium is used as a catalyst (~4-9 wt% V 2 O 5 ). Extremely high temperatures are required for this reaction to take place without a catalyst (Ognyanova et al., 2009a;Ognyanova et al., 2009b). The third stage is the conversion of SO 3 to sulphuric acid (Hiji et al., 2014). ...
The pyrite ashes formed as waste material during the calcination of concentrated pyrite ore used for producing sulphuric acid not only has a high iron content but also contains economically valuable metals. These wastes, which are currently landfilled or dumped into the sea, cause serious land and environmental pollution problems owing to the release of acids and toxic substances. In this study, physical (sulphation roasting) and hydrometallurgical methods were evaluated for their efficacy to recover non-iron metals with a high content in the pyrite ashes and to prevent pollution thereby. The preliminary enrichment tests performed via sulphation roasting were conducted at different roasting temperatures and with different acid amounts. The leaching tests investigated the impact of the variables, including different solvents, acid concentrations and leach temperatures on the copper and cobalt leaching efficiency. The experimental studies indicated that the pre-enrichment via sulphation roasting method has an effect on the leaching efficiencies of copper and cobalt, and that approximate recoveries of 80% copper and 70% cobalt were achieved in the H2O2-added H2SO4 leaching tests.
... Investigations have also been carried out to recover nickel and cobalt from spent catalysts using different methods such as adsorption, ion exchange, precipitation and solvent extraction. The spent catalysts were first leached by HCl (Parhi et al., 2013), H 2 SO 4 (Barik et al., 2012), oxalic acid (Anna, 2011) or NaOH (Ognyanova et al., 2009), followed by a serious of separation processes to ultimately produce the regenerated products such as NiSO 4 (Juferi et al., 2010), Co 3 O 4 (Padhan and Sarangi, 2014), CoC 2 O 4 (Nguyen and Lee, 2015). ...
... (2)). The oxidation of SO 2 to SO 3 is slow and needs a very high temperature to have a realistic rate without the use of a catalyst (Ognyanova et al., 2009). ...
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.
... Hydrometallurgical processes mainly involve high concentrations of acid and base for dissolving the metals from spent catalyst. 2,3) However, the use of high concentration of chemicals and their downstream processing is still a bottleneck for applying the process on a larger scale. Similarly, pyrometallurgical processes have also been tested to recover the metals from the spent catalyst, but high energy consumption and emission of toxic gases into the atmosphere hindered their wide applicability. ...
This study investigates the effectiveness of bioleaching in recovery of metals (Al, Ni, V and Mo) from raw petroleum refinery spent catalyst using Acidithiobacillus ferrooxidans (At. ferrooxidans) and Acidithiobacillus thiooxidans (At. thiooxidans). It was found that bioleaching with At. ferrooxidans or At. thiooxidans resulted in higher leaching yields of Ni (55.858.6%) and V (33.033.4%) as compared to Al (9.310%) and Mo (3.95.8%). After 168 h of bioleaching with either At. ferrooxidans or At. thiooxidans, the remaining metals in the bioleached spent catalyst samples were present in stable forms (oxidizable and residual fractions). Bioleaching also led to increase in the reduced partition index of all the metals in the bioleached residues (Ni: 0.630.65, Al: 0.98, V: 0.900.91, Mo: 0.800.83) as compared to feed spent catalyst (Ni: 0.14, Mo: 0.63, V: 0.70, Al: 0.94). The low 'risk assessment code' (RAC) values of the bioleached residues as compared to feed spent catalyst indicated that bioleached residues posed low or no risk to the environment. The results of the present study suggested that the bioleaching with either At. ferrooxidans or At. thiooxidans is effective in leaching of Ni and V, whereas leaching of Mo and Al requires further treatment.
... In some cases, alkaline solutions can also be used to selectively leach metals (Ognyanova et al., 2009). The researchers showed that vanadium can be extracted from a spent sulphuric acid catalyst by alkaline leaching. ...
Nickel is an important metal, heavily utilized in industry mainly due to its anticorrosion properties. As a consequence, nickel containing wastes such as spent batteries and catalysts, wastewater and bleed-off electrolytes are generated in various processes. These wastes could have a negative impact on the environment and human health if they contaminate soil, water and air. The present review addresses the environmental and economical aspects of nickel recovery/removal from various types of wastes. The main physico-chemical technologies for processing various effluents and wastewaters containing nickel are reviewed and discussed. Nickel recovery from spent batteries, catalysts, electronic waste and other sources is described. Hydrometallurgical approaches are emphasized. Recovery of nickel from wastes is important not only for economical aspects, but also for environmental protection.
... However the V recovery could be enhanced to 78% by alkali leaching of the roasted spent catalyst using NaoH at 80 • C in 2 h. 194 Bioleaching of spent catalysts rich in Ni, V and Mo using iron/sulphur oxidizing bacteria by addition of both S 0 and Fe(II), with studies of possible toxicity, revealed that addition of Fe(II) ion resulted in higher recovery of 83%, 90% and ∼35% for Ni, V and Mo, respectively. The reason for low recovery of Mo could be possibly due to less liberation of Mo from the organic entrapment of the spent catalyst. ...
Waste generated by the electrical and electronic devices is huge concern worldwide. With decreasing life cycle of most electronic devices and unavailability of the suitable recycling technologies it is expected to have huge electronic and electrical wastes to be generated in the coming years. The environmental threats caused by the disposal and incineration of electronic waste starting from the atmosphere to the aquatic and terrestrial living system have raised high alerts and concerns on the gases produced (dioxins, furans, polybrominated organic pollutants, and polycyclic aromatic hydrocarbons) by thermal treatments and can cause serious health problems if the flue gas cleaning systems are not developed and implemented. Apart from that there can be also dissolution of heavy metals released to the ground water from the landfill sites. As all these electronic and electrical waste do posses richness in the metal values it would be worth recovering the metal content and protect the environmental from the pollution. Cyanide leaching has been a successful technology worldwide for the recovery of precious metals (especially Au and Ag) from ores/concentrates/waste materials. Nevertheless, cyanide is always preferred over others because of its potential to deliver high recovery with a cheaper cost. Cyanidation process also increases the additional work of effluent treatment prior to disposal. Several non-cyanide leaching processes have been developed considering toxic nature and handling problems of cyanide with non-toxic lixiviants such as thiourea, thiosulphate, aqua regia and iodine. Therefore, several recycling technologies have been developed using cyanide or non-cyanide leaching methods to recover precious and valuable metals.
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Due to the increasing demand for the scale-up of metal–organic frameworks (MOFs) and the implementation of their potential in more industrial applications, many challenges need to be overcome related to the cost and availability of some precursors and the MOF synthesis conditions. In order to achieve a level allowing the industrialization of the production of MOFs, it is highly important to develop alternative methods for more sustainable and green synthesis of MOFs and their derived materials. In the context of organic and inorganic precursors shortage, the use of waste derived resources as a reliable source for both organic and inorganic elements is now being targeted for the production of these high-value materials, while considering the associated cost, as well as the environmental impacts of the precursors and the synthesis procedures. Combining two or more waste materials as precursors can be highly beneficial value and cost-wise for materials. This review is focused on the use of sustainable approaches and low-cost strategies for the synthesis of MOFs based mainly on the extraction of organic and inorganic precursors from recyclable waste, more specifically i) PET waste bottles as a valuable source for organic ligands and ii) electronic inorganic solid materials and other solid wastes containing metallic precursors recovered from industrial wastewater and electroplating sludge (EPS), petroleum refinery waste, etc. as a source for metallic ions, in addition to the inorganic insoluble precursors as non-conventional resources to yield pristine MOF, MOF composites and their derived materials such as carbons and oxide composites for advanced applications. Some advantages of these MOFs resourced from wastes will be given and compared to their homologue prepared from commercial precursors. This review will give an overview of how alternative routes from organic and inorganic wastes are important for the new trend of MOFs.
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.
Sülfürik asit üretiminde katalitik indirgemeyi sağlayan vanadyum katalizörlerinin yaygın kullanımı, tehlikeli atık olarak kabul edilen kullanılmış katalizörlerin zamanla daha da artmasına neden olmaktadır. Kullanılmış vanadyum katalizörler (KVK), yüksek oranda SiO₂ ve ağır metal içeriğinin yanında kritik metal listesinde yer alan vanadyumu da içermektedir. Döngüsel ekonomi politikası uygulamalarına yönelik artan talep, bu atıklardan vanadyumun kazanımı için tekno-ekonomik açıdan uygun bir yol geliştirmeyi gerektirmektedir. Bu çalışmada, kimyasal liç (1 M sülfürik asit ve %1 h/h hidrojen peroksit) ve biyoliç (Acidithiobacillus ferrooksidans, Acidithiobacillus thiooxidans ve Leptospirillum ferrooxidans içeren karışık bakteri kültürü) yöntemleri kullanılmış ve KVK’lardan vanadyum kazanımı değerlendirilmiştir. Katalizörlerde bulunan vanadyum, hidrometalurjik ve biyohidrometalurjik yöntemlerle yüksek verimle (%96,8 ve %97,1) kazanılmıştır. Geliştirilen modelleme de biyohidrometalurjik yöntemin yatırım maliyetinin 3,8 yılda geri karşılanacağı ve geri ödeme yüzdesi %89,32 olarak öngörülmüştür. Hidrometalurjik yöntemde ise, yatırım maliyetinin 1,2 yılda karşılanacağı ve geri ödeme yüzdesinin %80,3 olduğu belirlenmiştir. Bu sonuçlar hidrometalurjik yaklaşımın daha hızlı, biyohidrometalurjik yaklaşımın ise daha ekonomik bir yöntem olduğunu göstermiştir.
The present study deals with the potential application of the two adsorbents Dowex M4195 (M4195) and Lewatit MonoPlus TP220 (TP220) for the removal of vanadium(V) from the model and real aqueous solutions. The kinetic and equilibrium studies as well as the effects of pH, ion exchanger dose, contact time and initial concentration were studied using the batch method. Both the Langmuir and Freundlich adsorption isotherms and the pseudo first order, pseudo second order, intraparticle diffusion models were used to describe the adsorption process. The maximum adsorption capacity obtained from the Langmuir isotherm was found to be 248 mg/g and 209 mg/g on TP220 and M4195, respectively. The kinetics followed the pseudo second order reaction. Desorption, reusability of ion exchangers and vanadium(V) removal in the continuous system (column studies) were determined.
The solvent extraction of vanadium (V) from Na2CO3–NaHCO3 solution was attempted using Aliquat-336 (methyltrioctylammonium chloride) extractant dissolved in sulfonated kerosene as diluent with 2-octanol as phase modifier. The influence factors, such as Na2CO3 concentration, NaHCO3 concentration, temperature, extractant concentration and impurity P, on vanadium extraction were investigated. The results revealed that the single stage extraction efficiency of vanadium reached above 89% from 20 g/L Na2CO3 – 20 g/L NaHCO3 solution under the following conditions: extractant concentration of 15%, room temperature of 20–30 °C, 2-octanol of 6%, phase ratio (O/A) of 1:1, and oscillation time of 1 min. The addition of NaHCO3 into Na2CO3 solution could lower the pH value of the Na2CO3 solution, improve extraction efficiency of vanadium, and not introduce other impurities. The phosphate anion did not interact with the vanadate anion in the Na2CO3–NaHCO3 solution. The P in the Na2CO3–NaHCO3 solution had little influence on vanadium extraction.
This paper addresses the sustainability of vanadium, taking into account the current state-of-the-art related to primary and secondary sources, substitution, production, and market developments. Vanadium plays a critical role in several strategic industrial applications including steel production and probable widespread utilization in next-generation batteries. Confirming the importance of vanadium, the European Commission identified and formally registered this metal on the 2017 list of Critical Raw Materials for the European Union. The United States and Canada have also addressed the importance of this metal. Like the European economy, the American and Canadian economies rely on vanadium and are not globally independent. This recognized importance of vanadium is driving many efforts in academia and industry to develop technologies for the utilization of secondary vanadium resources using hydrometallurgical and pyrometallurgical techniques. In this paper, current efforts and their outcomes are summarized along with the most recent patents for vanadium recovery.
An environmentally friendly approach was reported for recovering vanadium (V) from aqueous media with aqueous two-phase system (ATPS) formed by poly (ethylene glycol) 2000 (PEG2000) and sodium sulfate with octadecyl amine ethoxylate ethers as extractants. The effect of aqueous pH, the type of extractants and phase-forming salts, vanadium concentration and extracting temperature on the extraction performance of vanadium was investigated. The researched results indicated that the extraction rate of vanadium has remarkable relationship with vanadium species in aqueous phase, and the selected extractants can extract efficiently vanadium from salt-rich to PEG200-rich phase. The initial and equilibrium pH in aqueous phase has no relationship with vanadium concentration, and there is no new covalent bond generation between polyvanadate anion and extractant in PEG 2000-rich phase by FTIR, which can be deduced that polyvanadate anion is extracted into PEG 2000-rich phase due to the electrostatic attraction interaction between protonated extractant and polyvanadate anion. Adding chloride and nitrate in aqueous two-phase system could remarkably decrease the extraction rate of vanadium, and then sulfate is in favor of extracting vanadium. The extraction rate of vanadium decreases with increasing temperature from 40.0 °C to 60.0 °C, which showed the phase transfer of vanadium from salt-rich phase to PEG2000-rich phase is exothermal, resulting from the difference of interaction energy among all compositions in aqueous two-phase system. The vanadium can be easily separated out from the solution containing Fe(III), Co(II), Ni(II), Cu(II), Zn (II), Al(III), Cr(III) and Mn(II) under the suitable conditions. The stripping rate and precipitation rate of vanadium can be easily achieved to 90% and 80%, respectively, in single stage stripping with ammonium carbonate solution as stripping agent and phase-forming salt.
Although there have been numerous studies on separation and purification of metallic minerals by hydrometallurgy techniques, applications of the chemical techniques in separation and purification of non-metallic minerals are rarely reported. This paper reviews disparate areas of study into processing and purification of quartz (typical non-metallic ore) in an attempt to summarize current work, as well as to suggest potential for future consolidation in the field. The review encompasses chemical techniques of the quartz processing including situations, progresses, leaching mechanism, scopes of application, advantages and drawbacks of micro-bioleaching, high temperature leaching, high temperature pressure leaching and catalyzed high temperature pressure leaching. Traditional leaching techniques including micro-bioleaching and high temperature leaching are unequal to demand of modern glass industry for quality of quartz concentrate because the quartz products has to be further processed. High temperature pressure leaching and catalyzed high temperature pressure leaching provide new ways to produce high-grade quartz sand with only one process and lower acid consumption. Furthermore, the catalyzed high temperature pressure leaching realizes effective purification of quartz with extremely low acid consumption (no using HF or any fluoride). It is proposed that, by integrating the different chemical processes of quartz processing and expounding leaching mechanisms and scopes of application, the research field as a monopolized industry would benefit.
Aim to the separation problem of high concentration V–Fe in stronger acidic leaching solution of vanadium slag, a novel chelating extraction system is proposed to separate V and Fe. Influence factors including initial pH value, the extractant concentration, the phase ratio, temperature and time are investigated. The experiment results show that single extraction rate of V and Fe reach 85.57 and 0.39% respectively, when pH value is −0.4, organic phase composition is 30% Mextral 973H + 70% sulfonated kerosene oil, O:A = 1:1, stirring time is 10 min, extraction temperature is 30 °C. The industrial experiment results of three levels countercurrent extraction show that the total extraction rate of vanadium and iron can reach 97.44 and 0.23%, separation factor is 6410. The effective separation of V and Fe in high acid and high iron solution can be accomplished.
In the current study, a novel chelating extractant Mextral 973H was employed to selectively extract V(V) from a low pH sulfuric acid solution containing high concentration of Fe (Ш). The extraction of V(V) is strongly dependent on solution pH, and the separation coefficient (β(V/Fe)) between V(V) and Fe(Ш) can be greater than 720 when the solution pH < 0.5. A cation exchange between -H and VO2⁺ was proposed as the main extraction mechanism at pH < 1.5, while at pH >1.5, a solvating reaction between the extractant and metal-oxo cluster anions took place. The effects of solution pH, acid medium types, contacting time and temperatures on the extraction and separation of V have been discussed. FT-IR and ¹HNMR were applied to confirm the extraction mechanism and also the degradation behavior of Mextral 973H in some extreme conditions. The loaded vanadium was effectively stripped using a diluted 0.2M Na2CO3 solution and a total stripping extent of 97% was achieved. The extraction process by Mextral 973H was finally simulated using Aspen Plus simulator comparing to traditional process, the results obtained were in good recovery efficiency and with the advantage of acid and alkali saving.
Two-stage hydrolytic purification of acidic solution of reductive leaching of spent vanadium catalysts for sulfuric acid production from arsenic and iron impurities has been investigated. After oxidation of Fe(II) and As(III) with hydrogen peroxide and neutralization of vanadium (IV) solution with lime to рН 2.8 at 25 °С, the concentration of iron and arsenic has been reduced from 4.5 to 1.6 g·L− 1 and from 1.0 to 0.6 g·L− 1, correspondingly. After neutralizing excessive acidity to рН 8.5 at 60 °С and oxidation of vanadium, the concentration of iron and arsenic in obtained solutions of sulfates and vanadates of alkali metals does not exceed 0.1 mg·L− 1. Lime's calcium ions are bound with excessive sulfate ions to form gypsum, which allows to preserve the composition of promoter salts and to return them into production of catalysts after vanadium separation. Washing of the second stage precipitation with initial leaching solutions allows to avoid vanadium losses, guarantees return of impurities to the first stage and their extraction in a convenient form for arsenic burial. The kinetic equation for an oxidation of vanadium (IV) solutions with atmospheric oxygen during their neutralization with lime in presence of ferric compounds has been acquired.
Reductive leaching of vanadium from spent vanadium catalysts for sulfuric acid production by sulfur dioxide, with the use of weakly acidic aqueous solutions of sulfurous acid, is studied. It is shown that the rate and completeness of vanadium leaching is related to the presence of admixtures of polyvalent metals. For catalyst with high iron content (1.5 %) the effect of different parameters on the kinetic of vanadium dissolution was determined. The acquired data was analyzed for correspondence to the mathematical models for heterogeneous processes. The best description of the leaching process is given by the model which assumes the process limitation by diffusion of two reactants trough the insoluble layer of forming product. It was found that these reactants are hydrogen and vanadyl ions reaction orders for which are 0.48 and -0.7, respectively. Temperature increase has a significant effect on the leaching rate (activation energy is 49.24 kJ ∙ mol-1). This value of activation energy can be explained by the change in permeability of the film of insoluble product with temperature increase, which improves diffusion rate of reactants.
In recent years the recycling of refractory metals has become more important. Specific reasons for this trend are local scarcities of raw materials because of remote deposits and the economic incentive of secondary metallurgy. At present there is generally no lack of supply of refractory metals like vanadium, nickel and molybdenum in the sector of ferro-alloys. Nevertheless, basic research projects that deal with an improvement of existing recycling routes have already been started. This paper describes known and new residues containing refractory metals which are feasible in secondary metallurgy. A comparison of various solid wastes like trash of refining crude oil or cogeneration plant residues will be shown. Basically, recycling can be done in two procedural ways with combinations of different technologies. On the one hand, a common process is leaching in numerous hydrometallurgical steps. On the other hand, pyrometallurgical techniques are available, which represents the focus of the current study. For the smelting of secondary raw materials and the refining to a marketable product, their melting behavior and characteristics of the produced slag should be well known. Hence, practical analyses have been done with a hot stage microscope and samples of three different initial conditions. At first, synthetic mixtures of lime, silica and alumina provided a basis which is visualized in a ternary system. Furthermore, a multi component, aluminum oxide rich waste material from the industry was molten with various contents of lime. Thirdly, synthetic mixtures with minor elements as impurities were heated to 1700 °C. Each heating process in the hot stage microscope offers values of the area and the shape of especially prepared samples, so a comparison of synthetic mixtures and industrial waste can be done. Based on this, the effect of minor elements on the ternary system is discussed.
Study on base metals, such as tin, nickel, iron, copper, from waste CPU has been performed using a hydrometallurgical route which consisted in oxidative leaching and chemical precipitation. The chemical leaching has been applied on concentrate obtained after milling and magnetic separation of these scraps. Four replicate leaching tests, using 1.89M sulfuric acid as leaching reagent and 10 ml of hydrogen peroxide (30%wt) as oxidant at 70°C for 24 hours were performed. The obtained results after this chemical treatment showed that average of extraction yield for copper was 1.37%, 70.82% for nickel, 99.86% for iron and 98.74% for tin,. All four pregnant solutions were mixed and then shared in two in order to study the precipitation of Sn, Fe and Ni by adding solution of NaOH with a concentration of 10% and 30%, respectively, for pH level augmentation until 7. The total recovery of tin from solution was achieved, while nickel and iron were precipitate in a very low concentration probably due their high solubility and to the presence of other elements.
Large amounts of three different kinds of zinc plants residues are annually generated in Iran. Hot filtercake (known as HFC) is one of them which could be used as a promising secondary resource of zinc, cobalt and manganese. Unfortunately, despite its valuable elements, HFC isn't treated and the most of its hazardous constituents are easily exposed to the environment. For the first time, in this paper, zinc was selectively leached from HFC employing alkaline leaching. Secondly, it was tried to optimize leaching through L16 taguchi experiments design to obtain the maximum recovery of zinc. In addition, the effect of each affecting parameter was studied.
With the increase in environmental awareness, the disposal of any form of hazardous waste has become a great concern for the industrial sector. Spent catalysts contribute to a significant amount of the solid waste generated by the petrochemical and petroleum refining industry. Hydro-cracking and hydrodesulfurization (HDS) catalysts are extensively used in the petroleum refining and petrochemical industries. The catalysts used in the refining processes lose their effectiveness over time. When the activity of catalysts decline below the acceptable level, they are usually regenerated and reused but regeneration is not possible every time. Recycling of some industrial waste containing base metals (such as V, Ni, Co, Mo) is estimated as an economical opportunity in the exploitation of these wastes. Alkali roasted catalysts can be leached in water to get the Mo and V in solution (in which temperature plays an important role during leaching). Several techniques are possible to separate the different metals, among those selective precipitation and solvent extraction are the most used. Pyrometallurgical treatment and bio-hydrometallurgical leaching were also proposed in the scientific literature but up to now they did not have any industrial application. An overview on patented and commercial processes was also presented.
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Annually, a great amount of zinc plants residue is produced in Iran. One of them is hot filter cake (known as HFC) which can be used as a secondary resource of zinc, cobalt and manganese. Unfortunately, despite its heavy metal content, the HFC is not treated. For the first time, zinc was selectively leached from HFC employing alkaline leaching. Secondly, leaching was optimized to achieve maximum recovery using this method. Effects of factors like NaOH concentration (C = 3, 5, 7 and 9 M), temperature (T = 50, 70, 90 and 105 °C), solid/liquid ratio (weight/volume, S/L = 1/10 and 1/5 W/V) and stirring speed (R = 500 and 800 rpm) were studied on HFC leaching. L16 orthogonal array (OA, two factors in four levels and two factors in two levels) was applied to determine the optimum condition and the most significant factor affecting the overall zinc extraction. As a result, maximum zinc extraction was 83.4 %. Afterwards, a rough test was conducted for zinc electrowinning from alkaline solution according to the common condition available in literature by which pure zinc powder (99.96 %) was successfully obtained.
In the studies on the recovery of vanadium from vanadium catalyst extracts, three types of polymer strongly acidic ion exchangers were used. The ion exchange resins differed in terms of granularity and their ion exchange capacity. As a result, breakthrough curves were made for three main components of the test extract, i.e.: ions of vanadium, iron and potassium. On this basis the optimum conditions for the removal of iron ions from the solution were defined and the technological concept of the process in the semi-technical scale was proposed.
Annually, thousands tons of Co residues called Hot Filter Cake (HFC) are produced from Co neutralization step of Iranian zinc plants. With respect to the composition of HFC )i.e., 15-25% Zn, 0.5–1.5% Co, 3–8% Mn(, it can be used as a secondary source of zinc, cobalt and manganese. In the present study, for the first time, treatment of HFC for separation and recovery of zinc has been studied. The residue was treated by employing selective alkaline leaching, in order to recover the maximum amount of zinc, followed by zinc electrowinning process. Therefore, a solution was obtained from alkaline leaching under the optimum condition of 75 °C, sodium hydroxide of 8 M, S/L of 1:10, and stirring speed of 600 rpm, having zinc recovery of 88.5 %. In the following step, the electrowinning process under the optimum working conditions being as current density= 350 A/m2 and time= 10 hours, was carried out to produce the zinc powder with high purity of 99 percent. Finally, a simple and effective conceptual flow diagram was proposed for the process.
The fly ash generated from oil burning is collected with a dust collector installed on a boiler. Recently, The necessity of developing a new process for recycling fly ash has recently received considerable attention due to the rather short life of landfills used for the disposal of a fly ash. Many tests have therefore recently been carried in order to develop a recycling process for useful components such as V and Ni following the concept of zero emission. The results can be summarized as follows: (1) V ions in the fly ash can be leached out by mixing with an equal quantity of water and most of the components that remain in the residue can then be leached out by adding an: aqueous ammonia solution. Using this two-stage leaching method, all components of interest can be leached out without heating or using significant amounts of chemicals. (2) It was possible to recover of Ni and V ions by a solvent extraction method which resulted in a high purity of the recovered compounds. (3) The sulfate ions, Mg ions, and ammonium ions that remained in the solution after the solvent extraction process to recover Ni and V ions were recoverable through a crystallization and distillation process as calcium sulfate, magnesium hydroxides and aqueous ammonia solution, respectively. (4) Based on results of this work, a new recycling system which does not produce secondary waste was developed for fly ash generated from oil burning.
Vanadium has many industrial uses and its contribution to environmental contamination is increasing all the time. Recovery of vanadium pentoxide from spent sulphuric acid catalysts was performed using a three-step process involving acid leaching, oxidation and precipitation. Several different acids were used in the leaching process. Finally, sulphuric acid was used in various concentrations, solid to liquid ratios, stirring times and temperatures. A high solid/liquid ratio in the leaching stage was used to obtain high concentration of vanadium pentoxide and low acid consumption that allowed direct precipitation without the use of extraction by rather expensive organic solvents. Sodium carbonate solution of one mole/liter concentration was used in the precipitation stage. An industrial application including material balance and operating conditions with an overall vanadium pentoxide recovery efficiency of (70%) was proposed.
Abstrakt V tomto článku sa študovali rôzne luhovacie systémy na získavanie zinku a horčíka z použitých alkalických a zinok-uhlíkových batérií. Experiment na získanie zinku a horčíka, prítomných vo forme prášku, sa realizoval dvomi kyslo-redukčnými luhovaniami: (1) kyselina sírová/kyselina oxálová a (2) kyselina sírová/peroxid vodíka. Pre každý systém sa vykonala štatistická analýza (ANOVA) na vyhodnotenie výsledkov extrakcie horčíku a zinku. Navrhol sa hydrometalurgický proces na recykláciu alkalických a zinok-uhlíkových bateriek. Abstract In this paper different leaching systems for the recovery of zinc and manganese from spent alkaline and zinc-carbon batteries have been studied. The experimental tests for the recovery of zinc and manganese present in the powder have been carried out by two acidic-reductive leachings: (1) sulphuric acid/oxalic acid and (2) sulphuric acid/hydrogen peroxide. For each system the analysis of variance (ANOVA) has been performed to evaluate the behaviour of the manganese and zinc extraction yields. Afterwards a hydrometallurgical process has been proposed for the recycling of alkaline and zinc-carbon batteries.
The chemical, physical and mechanical properties of Va-base alloys and unalloyed Va, its interaction with molten metals and gases and its applications are described. The sources and processing methods are outlined. The properties of vanadium compounds (oxides, hydroxides, peroxy compounds, vanadates, vanadium bronzes, sulphates and nitrates, halides and oxyhalides, and interstitial compounds) are presented. The reduction of vanadium compounds, refining (pyrovacuum treatment, electrorefining, iodide refining, zone melting, electrotransport, gettering) and the toxicology of vanadium are discribed. (A.P.)
The various process options available to recover vanadium from spent dehydrosulphurisation catalysts, sulphuric acid catalysts, and alumina sludge residues from the Bayer Process are reviewed, and the fate of other metal impurities such as Mo, Ni, Co, Al and Fe are considered. Most processes give an impure V2O5 product, but selective solvent extraction of vanadium from the impure leach solution allows high purity product to be obtained. A comparison is made between the D2EHPA and Amine extractants with regard to the vanadium species and other metals extracted. The results of studies with tertiary and quaternary amines in acidic media are reported, and the relative performance of quaternary amines in extracting a range of vanadium(V) anionic species between pH 6–13 is presented. It is shown that quaternary amines offer the greatest flexibility for treating acidic, neutral, or alkaline liquors depending on the process of choice.
This paper presents a brief review of extractive metallurgy of tantalum starting from processing of its ore to two pure intermediates K2TaF7 and Ta2O5 and their conversion to pure tantalum metal by various technically feasible processes. Though tantalum metal can be produced by several means only two processes – sodium reduction of K2TaF7 and fused salt electrolysis of K2TaF7 in the presence of oxide, have been successful on industrial scale. Besides providing salient features of these two processes, the paper presents brief accounts of studies carried out on the reduction of oxide by metallic reductants – calcium and aluminium as well as nonmetallic reductants – carbon and carbon–nitrogen. The crude metal obtained by various reduction techniques outlined are purified either by solid state pyrovacuum treatment or by melt refining in an electron beam furnace. Mechanism of refining processes taking place during these post reduction treatments are also included in the review.
Recovery of vanadium from a spent catlaysts leaching process has been studied using primary amine PRIMENE 81R, resulting in an industrial multistage process for the treatment of these effluents. An extraction mechanism for V(V) is proposed for this amine in acidic media, verifying the influence of pH on the process. Adequate ranges for the variables: O/A ratio, organic phase composition, pH, stirring speed and phase separation speed were fixed and simulated in industrial conditions. Vanadium is finally recovered by means of precipitation as ammonium metavanadate and later calcination to obtain vanadium pentoxide of commercial grade.
Flyash is a powdery residue generated by the power stations that use heavy oil as the source of fuel. The ash poses a threat to the environment due to the presence of certain heavy elements such as chromium. At the same time, it contains valuable metals, including vanadium and nickel, that are extractable if an economical and environmentally acceptable process can be developed. To this end, Mintek, in co-operation with Oxbow Carbon and Minerals, undertook a major research program over a two-year period in an attempt to find the most suitable technology to deal with this ash. After considering various alternatives, a process flowsheet was developed and used as a basis to identify the testwork required for the major processing units. The flowsheet consisted of a drying stage at moderate temperatures of about 150 °C, a de-carburization and de-sulphurization (pre-treatment) stage, and finally a smelting step in a DC arc furnace.Testwork carried out at the 40 kW scale, using pre-treated flyash, indicated the feasibility of producing a ferrovanadium alloy containing more than 15% vanadium and 6% nickel, with vanadium recovery of more than 89%, and a disposable slag based on USA EPA safe disposal criteria. Several parameters were studied in order to optimise the smelting stage including operating temperature, aluminium addition, the use of ferrosilicon as the reducing agent and lime (flux) addition. The results of the investigation are presented in this paper.
The recovery of vanadium from heavy oil fly ash having a high carbon content was performed using a four-step process consisting of a preliminary burning in order to reduce the carbonaceous fraction, followed by an acid leaching and an oxidative precipitation of vanadium pentoxide. The preliminary burning was conducted in the temperature range 650 to 1150 °C, below the initial deformation temperature (IDT) of the fly ash. The temperature of the preliminary burning step was revealed to be a significant parameter. Above 950 °C various phenomena (fusion, volatilisation of V, formation of V–Ni refractory compounds) occurred that adversely affected the recovery of vanadium. The burning temperature of 850 °C was found to be the best as a result of the trade-off between the overall vanadium recovery yield (83%) and the V2O5 weight percentage in the precipitate (84.8%).
This work concerns a three-step process for the recovery of vanadium from heavy oil and Orimulsion combustion fly ashes. This consisted of acid leaching, oxidation and precipitation of the vanadium pentoxide, followed by washing of the precipitate. Preliminary tests were conducted to investigate the effect of some operating parameters for the various steps of the process. After these preliminary tests, the recovery of vanadium from the fly ash samples was performed on a laboratory scale and the overall yield of the process was determined. By washing the precipitate, it was possible to reduce the concentration of the impurities and to allow its use for the production of ferrovanadium alloys.
Annual production of oil-fired fly ash in Taiwan is approximately 43 000 tons, of this approximately 13 000 tons is electrostatically precipitated, the rest is cyclonically collected. Structurewise, both consist of porous unburned carbon, vanadium and nickel oxide, and water-soluble sulfate. Electrostatically precipitated fly ash contains large amounts of ammonium sulfate. If these ashes are not properly disposed of, they become environmental problems, such as dusting, leakage of acid liquids, and pollution with heavy metals. This paper discusses the experimental extraction of vanadium and nickel from oil-fired fly ash. The results indicated that leaching of oil-fired fly ash in 0.5 N of sulfuric acid led to an extraction of 65% vanadium, 60% nickel, and 42% iron, along with an increase in the concentration of sulfuric acid. When leached in 2 N sodium hydroxide solution, the extraction of vanadium was 80%, and the extraction of nickel was negligible. If leached in an ammonia water, the extraction of nickel increased, along with an increase in the concentration of ammonia in water. When leached with 4 N ammonia water, the extraction of nickel was 60%, the extraction of vanadium was less than that obtainable from leaching in sulfuric acid solution or in sodium hydroxide solution. If electrostatically precipitated fly ash was leached in a solution of 0.25 N ammonia water and 2 N ammonium sulfate, it would yield an extraction of 60% nickel and 8% vanadium—leading to a selective extraction of nickel. This study has established an extraction flowsheet in which fly ash was first leached in an ammoniacal solution containing ammonium sulfate to recover nickel. The leached residues were then leached in an alkaline solution to recover vanadium.
Fly ash samples were collected from the flue lines of two different oil-fired power plants and analyzed by a variety of analytical procedures designed to determine the V cations extractable from the samples. Both VOâ/sup +/ and VO/sup 2 +/ were shown to be present in the samples. The V(V) cation, VOâ/sup +/, was the principal species extracted from these samples.
Fly ashes resulting from the combustion of fuel containing high concentrations of vanadium that can be slightly removed by water and more efficiently by alkaline or acid solutions. This uncontrolled release can contaminate water sources and requires appropriate storage of fly ashes. This study investigated the possibility of cleaning the ashes by leaching the material and recovering vanadium by solvent extraction (for metal concentration solutions higher than 200 mg V L-1) using several amine extractants (Primene JM-T, Amberlite LA-2, Alamine 336, and Alamine 304), a quaternary ammonium salt (Aliquat 336), and by a sorption process (for low-metal concentration solutions) using chitosan. Extraction and stripping were investigated with liquid extractants and showed that Aliquat 336 was the best of these extractants. However, since Aliquat 336 exhibits a greater difficulty at stripping, secondary or tertiary amine extractants appear more suited for the extraction process. Vanadium sorption occurs on chitosan through anion exchange with a maximum sorption capacity of 400 to 450 mg V g(-1) at pH 3. The treatment of acid leachates with chitosan does not appear possible, since it requires a pH control to pH 3, which precipitates ferric ions and coprecipitates vanadium. Alternative routes could be the alkaline leaching of fly ashes and a further pH control.
Chitosan is very efficient at removing vanadium from dilute solutions: sorption capacity can reach 400−450 mg V g-1 under optimum experimental conditions, which correspond to pH 3. In acidic solutions, the chitosan's amine groups are protonated, and vanadate anions can be exchanged with counterions bound to −NH3+ sites. A correlation is observed between the speciation diagram and the sorption isotherms, and it appears that decavanadate species are more favorable to sorption than other anionic vanadate species. Selecting experimental conditions under which decavanadate ions are the predominant form of vanadium results in enhanced sorption capacities and improved sorption kinetics. The results obtained in the study of vanadate desorption confirm the high affinity of chitosan for these polyoxoanions.
Recovery of vanadium and nickel from an oil fly ash is studied in a two-step leaching process, carried out under ambient pressure without calcination. In the leaching, nickel is dissolved with NH4Cl in the first step, followed by vanadium leaching with Na2CO3 in the second step. Both leachants depress the dissolution of iron and aluminium from the ash and hence the selective leachings are feasible. Nickel was recovered in 87% yield from the leach liquor by precipitation with Na2S, the purity of NiS being over 99% on the basis of the target metals (V, Ni, Fe, Al and Mg). For the recovery of vanadium, which has a relatively low concentration in the leach liquor, solvent extraction with TOA is preferable to concentration, prior to crystallization with NH4Cl. The experimental data showed that the vanadium concentration of the stripping solution is increased by about 15 times that of the leach liquor, which results in 78% yield as NH4VO3 crystals.
Extraction of molybdenum and vanadium from ammonia leaching residue (main chemical composition: 2.05% Mo, 0.42% V, 65.6% Al2O3 and 10.7% SiO2) of spent catalyst was investigated by roasting the residue with soda carbonate, followed by hydrometallurgical treatment of the roasted products. In the roasting process, over 91.3% of molybdenum and 90.1% of vanadium could be extracted when a charge containing a sodium carbonate to spent catalyst ratio of 0.15 was roasted at 750 °C for 45 min and the roasted mass was leached with water (liquid to solid ratio of 2) at 80–90 °C for 15 min. After the purification of leach liquor, an extraction solvent consisting of 20 vol.% trialkylamine (N235, commercialized in China) and 10 vol.% secondary octyl alcohol (phase modifier) dissolved in sulfonated kerosene was used to extract molybdenum and vanadium in leach liquor. 10 wt.% ammonia water was used as stripping agent. Adding 30 g/l NH4NO3 to the stripping solution and adjusting the pH to 7–8.5, over 99% of vanadium can be crystallized as ammonium metavanadate. Over 98% of molybdenum can be crystallized as ammonium polymolybdate when pH is between 1.5 and 2.5 (pH is adjusted by HNO3). Ammonium metavanadate and ammonium polymolybdate were calcinated at 500–550 °C, the purity of MoO3 and V2O5 was 99.08% and 98.06% respectively. In the whole process, 88.2% of molybdenum and 87.1% of vanadium could be achieved. The proposed roasting, leaching and separation steps give a feasible alternative for the processing of ammonia leaching residue of spent catalyst and can be applied in the comprehensive utilization of low grade molybdenum ores.
The results of a leaching kinetics study of spent nickel oxide catalyst with sulfuric acid are presented. The effects of spent catalyst particle size, sulfuric acid concentration, and reaction temperature on Ni extraction rate were determined. The results obtained show that extraction of about 94% is achieved using −200+270 mesh spent catalyst particle size at a reaction temperature of 85 °C for 150 min reaction time with 50% sulfuric acid concentration. The solid/liquid ratio was maintained constant at 1:20 g/ml. The leaching kinetics indicate that chemical reaction at the surface of the particles is the rate-controlling process during the reaction. The activation energy was determined as about 9.8 kcal/mol, which is characteristic for a surface-controlled process.
The necessity of recycling spent batteries in order to recover metals and to protect the environment is emphasized. Some experiments have been made to develop a hydrometallurgical process for metal recovery in alkaline manganese batteries. The components are easily separated: zinc can be electroplated and manganese(II) can be oxidized to dioxide; Mn(IV) is solubilized in acidic hydrogen peroxide and reprecipitated by bases.
A hydrochloric acid leaching process for the recovery of nickel, as nickel oxide, from low-grade spent catalyst analysing 17.7% Ni was studied. The effects of acid concentration, temperature, etc., on the extraction of nickel in chloride solution are first reported. Two different methods were examined for the separation of impurities from nickel chloride solution. In the first method, nickel was precipitated as solid nickel chloride by saturating the solution with hydrogen chloride gas. The effects of repeated leach/precipitation cycles were also investigated and it was found that the purity of the precipitated nickel chloride decreased as the number of cycles increased. Further purification was achieved by washing and anion exchange treatment. In the second method, nickel chloride solution was purified by removing impurities such as copper by cementation and iron and aluminium by oxidation/pH adjustment process. The preparation of nickel oxide from both methods was carried out by first precipitating nickel hydroxide, followed by calcination, to give an oxide with a purity suitable for the smelting process.
Power plant process simulation software is well-suited for the modelling of energy systems and more importantly, tools for analysing the energy efficiency are often built into the software. This work presents the development of a simulation model for a sulphuric acid plant using a commercial software package for power plant process simulation. This will be of value to for instance small consultant and engineering companies involved with audits and analysis of energy systems. For small sized companies the cost of acquiring and maintaining many different specialised software packages will be noticeable. However, companies involved with audits and analysis of energy systems will in most cases have access to at least one software package for power plant process calculations. The use of this kind of software for also modelling chemical plants would be valuable to these companies. The results of this work shows that it is possible to use an inexpensive but powerful power plant process simulation software for modelling a common chemical process as a part of a large energy system.
Three hydrometallurgical processes for industrial wastes treatment are presented. The main separation techniques are: solvent extraction, leaching-precipitation, electro-oxidation, and ion exchange. Recovery of gold from solid wastes generated in the electronic and jewellery industries consists of thermal degradation, two-stage leaching with nitric acid solution to remove silver and other metals and then with aqua regia to dissolve gold, selective solvent extraction of gold with diethyl malonate, and reduction of gold from the organic phase.Vanadium recovery from residue ashes after burning heavy oil fractions consists of alkaline leaching of vanadium, filtration, neutralization of sodium vanadate solution, precipitation of ammonium metavanadate, drying of the precipitate, and adsorption of the remaining vanadium from the filtrate on an anionite. From the remaining ashes nickel is recovered using acidic leaching, filtration, precipitation of ammonium-nickelous sulphate, filtration, and drying.The third process concerns processing of electroplating sludges and waste waters containing chromium and copper. The waste waters are electro-oxidized to transform Cr(III) into chromate. Then metal cations are separated on a cationite. The purified electroplating baths are recycled directly to electroplating; other solutions are first concentrated using anionite, followed by sodium chromate eluate conversion into concentrated chromic acid solution. The sludges accumulated from waste water processing by hydroxide precipitation are re-dissolved in chromic acid solution generated progressively by circulation between the dissolving and electro-oxidation steps. The concentrated chromic acid solution obtained is purified on the cationite and recycled.
Fly ash sample were collected from the flue lines of two different oil-fired power plants and analyzed by a variety of analytical procedures designed to determine the V cations extractable from the samples. Both VO2+ and VO2+ were shown to be present in the samples. The V(V) cation, VO2+ was the principal species extracted from these samples.
Debris flows originating in colluvial deposits within hillside swales cause significant property damage and loss of life in
steep, soil-mantled environments and will continue to do so for as long as mitigation measures are not adopted for both new
and existing development. Colluvium-filled swales constitute mappable debris flow source areas and should be identified routinely
as potential hazards in geotechnical feasibility and design-level investigations. Toward this end, we propose a swale classification
system that incorporates the state of activity, slope, and depth of colluvial fill within a swale. Recognition and siting
considerations allow for debris flow hazard mitigation by avoiding locations downslope of colluvium-filled swales, along low-order
channelways, or on debris fans. Alternative mitigations, particularly appropriate for areas of existing development, include
removal of colluvial deposits, subdrain installation, or downslope intervention to reduce the impact of a debris flow.
This study investigates the possibility of recovering nickel from the spent catalyst (NiO/Al2O3) resulting from the steam reforming process to produce water gas (H2/H2O) in many industries. In the extraction process, nickel is recovered as sulfate using sulfuric acid as a solvent. The considered parameters affecting nickel recovery were acid concentration, temperature and time of digestion solid:liquid ratio, particle size and stirring rate. Nickel was to be directly recovered as a sulfate salt by direct crystallization method. The conversion was 99% at 50% sulfuric acid concentration, solid: liquid ratio (1:12) by weight, particle size less than 500 micron for more than 5 h and 800 rpm at 100 degrees C.
Proposed technique in this investigation is given for vanadium and nickel enrichment in the Egyptian boiler ash. Among the various concepts for recovery of vanadium and nickel from boiler ash, the pyro-metallurgical approach is technically feasible, but is not cost-effective from an operational economy standpoint. Another technically viable process which, however, needs further development and presented in this investigation, is the hydrometallurgical processing that involves acid leaching under oxygen pressure of ground ash, followed by electrolytic separation of nickel from sulphate solution and vanadium is then neutralized and precipitated by adjustment the pH value and calcined to produce V2O5.
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.
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Possible flow sheet for selective extraction of V and Ni from spent sulfuric acid catalyst by means of a two-step leaching
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Fig. 6. Possible flow sheet for selective extraction of V and Ni from spent sulfuric acid catalyst by means of a two-step leaching.
Nickel recovery from spent catalysts: I. Solvation process
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