• Journal of the Mineralogical Society of Korea
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• v.32 no.3
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• pp.185-200
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• 2019

This study aimed to liberate gold from invisible gold concentrate (Au = 1,840.00 g/t) through microwave nitric acid leaching experiments. For the purpose, this study conducted microwave-nitric acid leaching experiments and examined nitric acid concentration effect, microwave leaching time effect and sample addition effect. The results of the experiments were as follows: Au (gold) contents were not detected in all of the microwave leaching conditions. In the insoluble-residue, weight loss rate tended to decrease as the nitric acid concentration, microwave leaching time and sample addition increased. In an XRD analysis with solid-residue, it was suggested that gypsum and anglesite were formed due to dissolution of calcite and galena by nitric acid solution. When a fire assay was carried out with insoluble-residue, it was discovered that gold contents of the solid-residue were 1.3 (Au = 2,464.70 g/t) and 28.8 (52,952.80 g/t) times more than those of concentrate. But in the gold contents recovered, a severe gold nugget effect appeared. It is expected that the gold nugget effect will decrease if a sampling method of concentrate is improved in the microwave-nitric acid leaching experiments and filtering paper with smaller pore size is used for leaching solution and burned filter paper is used for sampling in lead-fire assay.

• Applied Chemistry for Engineering
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• v.9 no.3
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• pp.407-412
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• 1998

This study was carried out to investigate the optimum leaching conditions for the sulfation and water leaching, and separation of cerium from rare earth elements in leached solution by acid-adjusting method. The optimum conditions for the sulfation and water leaching from bastnasite concentrates are that the equivalent ration of sulfuric acid to concentrates is 2.5, calcination temperature and time are $600^{\circ}C$ and 2 hrs respectively, and the pulp density in the water leaching is 9.1%. The yield of rare earth oxide is about 93% at the above condition. The process of recovery of cerium hydroxide from leached solution by acid-adjusting method was carried out as following steps. The first step is the oxidation of the solution at pH 5 by using twice the equivalent of $H_2O_2$ solution as an oxidant. The second step is the precipitation to obtain cerium complex salt and cerium hydroxide after lowering the solution to pH 2. The last step is the oxidation-precipitation by using equivalent of $H_2O_2$ solution. From these results, it was possible to prepare cerium hydroxide with the yield of 60% and the quality of 80%.

• Clean Technology
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• v.30 no.1
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• pp.28-36
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• 2024

As demand for electric vehicles increases, the market for lithium-ion batteries is also rapidly increasing. The battery life of lithium-ion batteries is limited, so waste lithium-ion batteries are inevitably generated. Accordingly, lithium was selectively preleached from waste lithium iron phosphate (LiFePO₄, hereafter referred to as the LFP) cathode material powder among lithium ion batteries, and iron phosphate (FePO₄) powder was recovered. The recovered iron phosphate powder was mixed with alkaline sodium carbonate (Na₂CO₃) powder and heat treated to confirm its crystalline phase. The heat treatment temperature was set as a variable, and then the leaching rate and powder characteristics of each ingredient were compared after water leaching using Di-water. In this study, lithium showed a leaching rate of approximately 100%, and in the case of powder heat-treated at 800 ℃, phosphorus was leached by approximately 99%, and the leaching residue was confirmed to be a single crystal phase of Fe₂O₃. Therefore, in this study, lithium, phosphorus, and iron components were individually separated and recovered from waste LFP powder.

• Resources Recycling
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• v.14 no.6 s.68
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• pp.28-36
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• 2005

A hydrometallurgical process to leach cobalt from the waste cathodic active material, $LiCoO_{2}$, and subsequently to separate it by solvent extraction was developed. The optimum leaching conditions for high recovery of colbalt and lithium were obtained: 2.0 M sulfuric acid, 5 $vol.\%$ hydrogen peroxide, $75^{\circ}C$ leaching temperature, 30 minutes leaching time and an initial pulp density of 100 g/L. The respective leaching efficiencies for Co and Li were $93\%$ and $94.5\%$. About $85\%$ Co was extracted from the sulfuric acid leach liquor containing 44.72 g/L Co and 5.43 g/L Li, using 1.5 M Cyanex272 as an extractant at the initial pH 5.0 and in organic to aqueous phase ratio of 1.6:1 under the single stage extraction conditions. The Co in the raraffinate was completely extracted by 0.5 M Na-Cyanex272 at the inital pH 5.0, and an organic to aqueous phase ratio of 1;1. The cobalt sulfate solution of higher than $99.99\%$ purity could be recovered from waste $LiCoO_{2}$, using a series of hydrometallurgical processes: sulfuric acid leaching of waste $LiCoO_{2}$- solvent extraction of Co by Na-Cyanex 271 - scrubbing of Li by sodium carbonate solution - stripping of Co by sulfuric acid solution.

• Resources Recycling
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• v.30 no.6
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• pp.43-52
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• 2021

In this study, the optimal nitration process for selective lithium leaching from powder of a spent battery cell (LiNi_xCo_yMn_zO₂, LiCoO₂) was studied using Taguchi method. The nitration process is a method of selective lithium leaching that involves converting non-lithium nitric compounds into oxides via nitric acid leaching and roasting. The influence of pretreatment temperature, nitric acid concentration, amount of nitric acid, and roasting temperature were evaluated. The signal-to-noise ratio and analysis of variance of the results were determined using L₁₆(4⁴) orthogonal arrays. The findings indicated that the roasting temperature followed by the nitric acid concentration, pretreatment temperature, and amount of nitric acid used had the greatest impact on the lithium leaching ratio. Following detailed experiments, the optimal conditions were found to be 10 h of pretreatment at 700℃ with 2 ml/g of 10 M nitric acid leaching followed by 10 h of roasting at 275℃. Under these conditions, the overall recovery of lithium exceeded 80%. X-ray diffraction (XRD) analysis of the leaching residue in deionized water after roasting of lithium nitrate and other nitrate compounds was performed. This was done to determine the cause of rapid decrease in lithium leaching rate above a roasting temperature of 400℃. The results confirmed that lithium manganese oxide was formed from lithium nitrate and manganese nitrate at these temperatures, and that it did not leach in deionized water. XRD analysis was also used to confirm the recovery of pure LiNO₃ from the solution that was leached during the nitration process. This was carried out by evaporating and concentrating the leached solution through solid-liquid separation.

• Korean Chemical Engineering Research
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• v.53 no.1
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• pp.46-52
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• 2015

A leaching kinetics was conducted for the purpose of recovery of praseodymium in sulfuric acid ($H_2SO_4$) from REE slag concentrated by the smelting reduction process in an arc furnace as a reactant. The concentration of $H_2SO_4$ was fixed at an excess ratio under the condition of slurry density of 1.500 g slag/L, 0.3 mol $H_2SO_4$, and the effect of temperatures was investigated under the condition of 30 to $80^{\circ}C$. As a result, praseodymium oxide ($Pr_6O_{11}$) existing in the slag was completely converted into praseodymium sulfate ($Pr_2(SO_4)_3{\cdot}8H_2O$) after the leaching of 5 h. On the basis of the shrinking core model with a shape of sphere, the first leaching reaction was determined by chemical reaction mechanism. Generally, the solubility of pure REEs decreases with the increase of leaching temperatures in sulfuric acid, but REE slag was oppositely increased with increasing temperatures. It occurs because the ash layer included in the slag is affected as a resistance against the leaching. By using the Arrhenius expression, the apparent activation energy of the first chemical reaction was determined to be $9.195kJmol^{-1}$. In the second stage, the leaching rate is determined by the ash layer diffusion mechanism. The apparent activation energy of the second ash layer diffusion was determined to be $19.106kJmol^{-1}$. These relative low activation energy values were obtained by the existence of unreacted ash layer in the REE slag.

• Economic and Environmental Geology
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• v.53 no.2
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• pp.159-166
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• 2020

In this research, we investigate the effect of microwave pretreatment on the recovery of gold from the gold concentrates by thiourea leaching. The changes in mineral phases by decomposition of pyrites in the gold concentrates using microwave were observed, and the result of microwave irradiation showed that the temperature of the irradiated sample increases with increasing irradiation time. With the reaction of temperature increases, Sulfur (S) in pyrites was converted to sulfur dioxide (SO₂), and then the content of S in the sample was reduced. The analytical results of XRD and SEM-EDS showed that pyrites are converted to magnetite and hematite, and its surfaces are changed to a porous shape where micro-cracks are developed. The Au leaching efficiency from the irradiated gold concentrates using thiourea-mixed solvent increased with the increases of irradiation time and solvent concentration. The experimental results considering leaching parameters indicate that the mechanism of microwave irradiation increases the maximum leaching efficiency and leaching rate of the gold concentrates, and the solvent does a role for the increasing of leaching rate constant.

• Proceedings of the IEEK Conference
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• 2001.10a
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• pp.448-452
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• 2001

An Experimental study was carried out to develop a simple method of processing copper waste sludge which is produced by PBC manufacturing. The procedure is based on leaching of wet sludge in 2N H$_2$SO$_4$, and the solid / liquid ratio is controlled approximately at 1/10. The recovery of copper is 85.4%, and pH of the leachate is 3.20. Adding ammonia solution into leachate forms ammine, and hydroxide compounds derived from other impurities in leachate at pH 10. The hydroxide compound can be treated by ferrite process, and the product is a stable oxide compound. Then the ammine solution is heated to evaporate ammonia, and the copper hydroxide is formed. Heating at 8$0^{\circ}C$by aeration, copper hydroxide is transformed into copper oxide with a purity of 98.4%. This process can recover most copper from sludge and the residue can be stabilized by the formation of a stable oxide compound which is not hazardous to environment.

• Resources Recycling
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• v.21 no.6
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• pp.65-73
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• 2012

Selective catalytic reduction(SCR) catalysts are obtained from de-NOx system of thermoelectric power plant. A process was developed for valuable metals such as vanadium and tungsten recovery from spent SCR catalyst by using soda roasting followed by water leaching. Spent SCR catalyst having $V_2O_5$(1.23 mass %) and $WO_3$(7.73 mass %). For getting soluble metal forms of the targeted metals like vanadium and tungsten soda roasting process was implemented. In soda roasting process, sodium carbonate added 5 equivalent ratio at roasted temperature $850^{\circ}C$ with 120 min roasted time for $544{\mu}m$ particle size of spent SCR catalyst. After soda roasting process moved to water leaching for roasted spent catalyst. Before leaching process the roasted spent catalyst was grinded up to $-45{\mu}m$ size. The leaching time is 30 min at $40^{\circ}C$ temperature, 10 % pulp density. The final leaching efficiency obtained 46 % of vanadium and 92 % of tungsten from present process.

• Journal of the Korean Ceramic Society
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• v.48 no.2
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• pp.189-194
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• 2011

Metallurgical grade silicon was recovered from slurry waste for ingot sawing process by acid leaching and thermal treatment. SiC abrasive was removed by gravity concentration and centrifugation. Metal impurities were removed by the acid leaching using HF/HCl. The remaining SiC was separated by the thermal treatment at $1600^{\circ}C$ in an inert atmosphere by the difference in melting points. The purity of the obtained silicon was found to be around 99.7%.