• Resources Recycling
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• v.31 no.2
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• pp.49-55
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• 2022

In the present study, the thermodynamic evaluation of the sulfate-roasting process was conducted to extract vanadium from the Korean vanadium titano-magnetite ore. The leaching efficiency of vanadium and other impurities was analyzed for varying roasting temperatures and addition of Na₂SO₄. In the case of sulfate roasting, the roasting temperature was 200 ℃ higher than that previously observed Na₂CO₃ roasting. However, the higher leaching efficiency of vanadium and lower leaching efficiency of other impurities, such as aluminum and silicon, were observed. The high selectivity for the extraction of vanadium in sulfate roasting would result from the reaction mechanism between SO₂ gas and vanadium concentrate.

• Resources Recycling
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• v.31 no.2
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• pp.33-39
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• 2022

In this study, analytical thermochemical modeling factors that contribute to maintaining a specific temperature range during vanadium roasting as a pretreatment using a rotary kiln are investigated. The model-related mechanisms include thermochemical reaction rates, heat balance, and heat transfer, through which the resultant temperature can be estimated intuitively. Ultimately, by optimizing these parameters, the ideal roasting temperature in the kiln is ≈1000 ℃ (or ≈1273 K) for long-term operation. Therefore, the heat generated from hydrocarbon (natural gas) fuel combustion and ore oxidation reactions, as well as the radiant heat transferred to ores, are assessed. In addition, thermochemical methods for relieving the temperature gradient in order to maintain the optimum temperature range of the rotary kiln are suggested.

• Resources Recycling
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• v.30 no.4
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• pp.27-34
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• 2021

In this study, the leaching behavior of vanadium (V) was investigated through CaO roasting and sulfuric acid leaching from domestic V-bearing titanomagnetite (VTM). Changes in the phase according to the amount of CaO added and roasting temperature were analyzed. Regardless of the roasting conditions, perovskite (CaTiO₃) was preferred to form. When the CaO content was increased, the calcium ferrite (CaFeO_x) phase was formed; otherwise, ferrite (Fe₂O₃) was preferred. After CaO was roasted, leaching was performed for 6 h with 1M sulfuric acid at 50℃ and a 10% solid-liquid ratio. Results of leaching revealed that when the roasted product was sintered, V was not sufficiently oxidized, and the leaching efficiency decreased. In addition, when the roasting temperature was low, the leaching efficiency of V decreased due to the influence of unreacted excess CaO. To lower the leaching efficiency of iron and titanium in VTM concentrates, suppressing the formation of CaTiO₃ and CaFeO_x was necessary by minimizing the amount of CaO added. Consequently, a leaching efficiency of 86% V, 4.3% Fe, and 6.5% Ti was obtained when the roasted product of 1150℃ and 10 wt% CaO was leached.

• Resources Recycling
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• v.31 no.2
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• pp.56-62
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• 2022

This study investigated the water leaching behavior of vanadium in Na₂CO₃-roasted vanadium-bearing titaniferous magnetite (VTM) concentrate. The magnetic concentrate and Na₂CO₃, mixed in a mass ratio of 4:1, were roasted at 1050 ℃, kept for 3 h, and ground to a size of D₅₀ = 48.79 ㎛ using a rod mill. The effects of leaching temperature and pulp density on water leaching were then investigated. The results show that the vanadium leaching efficiency decreased to 90.4%, 88.2%, and 83.8% as the temperature increased to 25, 55, and 85 ℃, respectively, whereas it remained almost constant 90.4%, 87.0%, and 87.0% as the pulp density increased to 10, 50, and 100 w/v%, respectively. Based on the preliminary leaching results, multi-stage leaching was conducted with the experimental conditions of 25 ℃, 100 w/v%, 300 rpm, and 1 h. The vanadium concentration in the final leaching solution was determined as 16.20 g/L after four stages of leaching. Thus, a high-concentration sodium vanadate solution was prepared by multi-stage leaching.

• Resources Recycling
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• v.32 no.1
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• pp.42-49
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• 2023

The glass ceramic secondary resource containing Li-Al-Si is used in inductor, fireproof glass, and transparent cookware and accounts for 14% of the total consumption of Li, which is the second most widely used after Li-ion batteries. Therefore, new Li resources should be explored when the demand for Li is exploding, and extensive research on Li recovery is needed. Herein, we recovered Li from fireproof Li-Al-Si glass ceramic, which is a new secondary resource containing Li. The fireproof glass among all Li-Al-Si glass ceramics was used as raw material that contained 1.5% Li, 9.4% Al, and 28.9% Si. The process for recovering Li from the fireproof glass was divided into two parts: (1) calcium salt roasting and (2) water leaching. In calcium salt roasting, a sample of fireproof glass was crushed and ground below 325 mesh. The leaching efficiency was compared based on the presence or absence of heat treatment of the fireproof glass. Moreover, the leaching rates based on the input ratios of calcium salt, Li-Al-Si glass, and ceramics and the leaching process based on calcium salt roasting temperatures were compared. In water leaching, the leaching and recovery rates of Li based on different temperatures, times, solid-liquid ratios, and number of continuous leaching stages were compared. The results revealed that fireproof glass ceramics containing Li-Al-Si should be heat treated to change phase to beta-type spodumene. CaCO₃ salt should be added at a ratio of 6:1 with glass ceramics containing Li-Al-Si, and then leached 4 times or more to achieve a recovery efficiency of Li over 98% from a solution containing 200 mg/L of Li.

• Resources Recycling
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• v.4 no.3
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• pp.2-9
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• 1995

The rccoverg and separation pracess of nikcl, vanadium and molybdenum from spent dcsulfilrizing catalyst ofpetrochemical rndustries was studied. Tnis process was canied out wet process which is consist of roasting, ammonialeaching and solve111 exDaction techniqcs. The metal ions of NI, V and Mo as vduable compollents were treated byroasting them a1 low lernperatuc, 400$^{\circ}$C in first dep, and then dlssah'ed nu1 at 80$^{\circ}$C wlth ammonium cabonate mlulion.Aftcr cooling them a1 room tempertaure, vanadium wa rccavered from mathcr iiquur in thc f n m of precipitate, sodiumvanadales The Secand slep, roasting the catalyst which is added sodium carbonate ul IOOO"C, was employed. Leachingwith distilled ~ a l e rga ve a iwo phase resultant, solutio~c~a ntaning Ni, V and Mo and solid residue containing sibca,alurmniu~n and iron. A solvcnt exlclction technique uslng vvriuus extractanls, MSP-8, TOIUC, LIX64Pi was eflecnve farthc extraclion and scparation ol thrcc mcfals from thc ammonical 11qou1 thc ammonical 11qou1.

• Resources Recycling
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• v.5 no.1
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• pp.14-20
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• 1996

For recycl~ng alumhx~rn dross, the cbaraclcrislics of dross and ils reutll~zalin~af~te r plocsssing should he considered. Nurn~llumd ross was classified according to iB sire in this shldy. Tbe dross larger Illan 3001~1w as directly re-meltcd to recover aluminum, and the s~nallerd ~ossw as leachcd and riln~tedl o scparalc 111s eri~lerls alt and to oxidize the rnelals contained in the dross. As a rcwlt, amount of the dross uscd lo be dircarded ahcr processing could hc reduced Also, lhc chem~cal culnposit~ons of a domzstic alumlnum dross and the changes in con~posilii~ndsu ring processings were investigated. and Cound that most mctaU'ic aluminum in the dross was changed inlu alu~n~numox ide lhruugli the roasting. Tile processed dross would he ulillzed for malerials such as alumina, alumma cemenl or tilcs.

• Resources Recycling
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• v.31 no.2
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• pp.40-48
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• 2022

In this study, the effects of solution components were investigated in the recovery of vanadium as ammonium metavanadate from vanadium-ore-salt roasting-water leaching solution. The vanadium-containing solution is strongly alkaline (pH 13), so the pH must be lowered to 9 or less to increase the ammonium metavanadate precipitation efficiency. However, in the process of adjusting the solution pH using sulfuric acid, aluminum ions are co-precipitated, which must be removed first. In this study, aluminum was precipitated in the form of an aluminum-silicate compound using sodium silicate, and the conditions for minimizing vanadium loss in this process were investigated. After aluminum removal, the silicate was precipitated and removed by adjusting the solution pH to 9 or less using sulfuric acid. In this process, the concentration and addition rate of sulfuric acid have a significant influence on the loss of vanadium, and vanadium loss was minimized as much as possible by slowly adding dilute sulfuric acid. Ammonium metavanadate was precipitated using three equivalents of ammonium chloride at room temperature from the aluminum-free, aqueous solution of vanadium following the pH adjustment process. The recovery yield of vanadium in the form of ammonium metavanadate exceeded 81%. After washing the product, vanadium pentoxide with 98.6% purity was obtained following heat treatment at 550 ℃ for 2 hours.

• Resources Recycling
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• v.3 no.3
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• pp.21-26
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• 1994

In present study an attempt has been made to extract the high purity $Al_2O_3$from domestic aluminous shale. The processes of the salt roasting with ammonium sulfate, extraction with sulfuric acid and calcination were adopted. In the extraction of alumina, the effects of the sulfuric acid concentration, the reaction time and the temperature has been investigated. The reaction products were analyzed by X-ray diffraction, DTA-TG, chemical analysis and SEM. The results are summerized as follows: 1)The pretreatment conditions were 0.6M-$(NH_4)_2SO_4$and $650^{\circ}C$ in sintering temperature. 2) The optimum extraction conditions were $10%-H_2SO_4$ and 240 minutes in acid treating time. 3)Physical properties of sintering materials were confirmed as $\alpha-Al_2O_3$ by X-ray diffraction method and the purity of $\alpha-Al_2O_3$ was 99.23%.

• Resources Recycling
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• v.33 no.2
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• pp.24-36
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• 2024

The lithium-ion battery recycling process has been classified into direct recycling, hydrometallurgical process, and pyrometallurgical process. The commercial process based on the hydrometallurgical process produces black mass through pretreatment processes consisting of dismantling, crushing and grinding, heat treatment, and beneficiation, and then each metal is recovered by hydrometallurgical processes. Since all lithium-ion battery recycling processes under development conducts hydrometallurgical processes such as leaching, after the pretreatment process, to produce precursor raw materials, this article suggests a classification method according to the pretreatment method of the recycling process. The processes contain sulfation roasting, carbothermic reduction roasting, and alloy manufacturing, and the economic feasibility of the lithium-ion battery recycling process can be enhanced using unused by-products in the pretreatment process.