• KEPCO Journal on Electric Power and Energy
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• v.7 no.1
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• pp.153-159
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• 2021

A two-step process for increasing the leaching efficiency of yttrium and neodymium from coal fly ash were investigated at solid loadings of 5.0 g ash ~1,000 g ash/l of 1.0 N~10.0 N H₂SO₄, temperature ranging from 30℃ to 90℃, ultrasonic leaching time of 1~10 hours, and ultrasonic power of 25~200 W. The yttrium and neodymium from coal fly ash were effectively leached into ion phases by step change of the first conventional dissolution at room temperature and then the second heating process with the aid of ultrasonic wave, and maximum leaching efficiency of yttrium and neodymium obtained were 66 % and 63 %, respectively. The activation energies for the leaching reaction of yttrium and neodymium at second heating process dependent on leaching time and temperature were derived to be 41.540 kJmol^-1 and 507.92 kJmol^-1, respectively. The optimum conditions for the maximum leaching of yttrium and neodymium were found to be the solid loading of 250 g ash/l of H₂SO₄, solvent concentration of 2.0 N H₂SO₄, and second step process of temperatures of 30℃ for 3 hours and then 90℃ for 4 hours with ultrasonic intensity of 100 W.

• KEPCO Journal on Electric Power and Energy
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• v.3 no.1
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• pp.29-34
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• 2017

Leaching kinetics for extracting yttrium from the coal fly ash was investigated in the presence of sulfuric acid during extraction. The leaching kinetics of yttrium were conducted at reactant densities of 5~1,000 g coal fly ash per L of $1.0{\sim}10.0N\;H_2SO_4$, agitation speed of 250 rpm and temperature ranging from 30 to $90^{\circ}C$. As a result, the leaching kinetic model was determined in a two-step model based on the shrinking core model with spherical particles. The first step was proceeded by chemical reaction at ash surface, and the second step was proceeded by ash layer diffusion because the leaching conversion of yttrium by the first chemical reaction increases with increased the time irrelevant to the temperature whereas it increases with increased the leaching temperature. The activation energy of the first chemical leaching step was determined to be $1.163kJmol^{-1}$. After the first chemical reaction, the activation energy of ash layer diffusion leaching was derived to be $41.540kJmol^{-1}$. The optimum conditions for leaching the yttrium metal of 60 % were found to be the slurry density of 250 g fly ash per L of $H_2SO_4$, solvent concentration of $2.0N\;H_2SO_4$, second step leaching of temperatures of $30^{\circ}C$ for 3 hours and then $90^{\circ}C$ for 3 hours at agitation rate of 250 rpm.

• Clean Technology
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• v.29 no.2
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• pp.87-94
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• 2023

The recovery of valuable metals from waste lithium-based secondary batteries is very important in terms of efficiently utilizing earth's limited number of resources. Currently, the cathode material of a LiFePO₄ battery, a type of battery which is widely used in automobiles, contains approximately 5% lithium. After use, the lithium in these batteries can be used again as a raw material for new batteries through lithium recycling. In this study, low-concentration sulfuric acid, a commonly used type of inorganic acid, was used to selectively leach the lithium contained in a waste LiFePO₄ cathode material powder. In addition, in order to compare and analyze the leaching efficiency and separation efficiency of each component, the optimalleaching conditions were derived by applying a two-step leaching process with pulp density being used as a variable during leaching. When leaching with pulp density as a variable, it was confirmed that at a pulp density of 200 g/L, the separation efficiency was approximately 200 times higher than at other pulp densities because the iron and phosphorus components were hardly leached at this pulp density. Accordingly, the pulp density of 200 g/L was used tooptimize the leaching conditions for the selective leaching and recovery of lithium.

• Resources Recycling
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• v.20 no.3
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• pp.48-54
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• 2011

Sulphuric acid leaching was conducted to extract the metal values from spent refinery catalyst. More than 95% of Ni and V and 30% of Mo could be leached out in 1 M sulphuric acid and 1 hr of leaching time. The decrease in Mo leaching was due to typical characteristic of Mo matrix. The activation energies of the leaching reactions showed the dissolution process follows a diffusion control mechanism. In order to leach out all Mo, further the leaching experiments were conducted with sulfur free spent refinery catalyst. For sulfur free spent refinery catalyst, a two step process of leaching with 1 M sulphuric acid followed by sodium carbonate washing showed better leaching than a two step leaching process with sodium carbonate followed by sulphuric acid washing, with almost 99% leaching of Ni, Mo and V. Solvent extraction using LIX 841 were conducted for a leach liquor containing Ni, 2 g/L; V, 9 g/L, Mo, 0.6 g/L. More than 98% of Mo was extracted from the leach liquor at A:O ratio of 5:2 in a 2 stage process. Similarly V was extracted at A:O ratio of 5:3 in a 2 stage process with 82% of total V extraction.

• Journal of Powder Materials
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• v.16 no.5
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• pp.351-357
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• 2009

A two-step recovery method was developed to produce copper powders from copper chloride waste solution as byproducts of MoO$_3$ leaching process. The first step consisted of replacing noble copper ions with external Fe$^{3+}$ ions which were formed by dissolving iron scraps in the copper chloride waste solution. The replaced copper ions were subsequently precipitated as copper powders. The second step was cementation of entire solution mixture to separate (pure) copper powders from aqueous solution of iron chloride. Cementation process variables of temperature, time, and added amount of iron scraps were optimized by using design of experiment method and individual effects on yield and efficiency of copper powder recovery were investigated. Copper powders thus obtained from cementation process were further characterized using various analytical tools such as XRD, SEM-EDS and laser diffraction and scattering methods.Cementation process necessitated further purification of recovered copper powders and centrifugal separation method was employed, which successfully yielded copper powders of more than 99% purity and average 1$\sim$2$\mu$m in size.

• Resources Recycling
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• v.11 no.3
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• pp.10-16
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• 2002

In the leaching of $LiCoO_2$with a strong acid such as sulfuric and nitric acid, an additional step was needed to recover cobalt and lithium separately from spent lithium ion batteries (LIBs). The leaching of $LiCoO_2$in an oxalic acid solution was investigated to recover cobalt selectively using a low solubility of cobalt oxalate at low pH. Leaching efficiency of 95% of lithium and less than 1% of cobalt were obtained when pure $LiCoO_2$powder was leached in 3M oxalic acid at $80^{\circ}C$ and 50 g/L pulpdensity. Under the above leaching conditions, complete dissolution of lithium was accomplished with mere 0.25% of cobalt in the solution when the cathodic active material collected from spent LIBs was employed. The lithium in the leaching solution can be recovered as a form of carbonate or hydroxide depending on the addition of $Na_2$$CO_3$or LiOH.

• Resources Recycling
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• v.29 no.1
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• pp.17-24
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• 2020

In this study, rare earth elements (REE) leaching from a refractory REE ore containing goethite as a major gangue mineral was conducted, introducing a two-stage method of chemical decomposition-acid leaching. At the chemical decomposition step, using one of alkaline agent, NaOH, the ore was decomposed, changing NaOH concentration from 20 to 50 wt% at 10% (w/w) of pulp density and the maximum temperature achieved without boiling at each NaOH concentration. With increasing NaOH concentration, light REE (Ce, La and Nd) and iron were concentrated in the solid phase which is the decomposed product, while aluminum (Al) and phosphorus (P) were removed to the liquid phase, and their concentrations in the solid phase were down to 0.96 and 0.17%, respectively. In addition, through XRD analysis, it was found that the crystallinity of goethite was considerably decreased. At the acid leaching step, the product decomposed by 50 wt% NaOH was leached at 3.0 M HCl and 80 ℃ for 3 hr, then the REE leaching efficiency was above 94% (Ce 80%), and the leaching efficiencies of Al and P were decreased to 12 and 0%, respectively. Therefore, in terms of both REE leaching efficiency and impurity removal, those decomposition and leaching conditions were chosen as optimum processing methods of the investigated material. In terms of REE leaching mechanism, because REE and iron leaching efficiencies showed the positive correlation each other, so it can be concluded that decreasing crystallinity of goethite affect the improvement of REE leaching.

• 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%.

• Applied Chemistry for Engineering
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• v.19 no.6
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• pp.668-673
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• 2008

The objective of this study is to optimize the leaching conditions of sequential leaching and extracting processes for selective Ni recovery from spent Ni-Mo-based catalyst. The selective Ni recovery process consists of two processes of leaching and extracting process. In this 2-step process, Ni component is dissolved from solid spent Ni-Mo-based catalyst into leaching agent in leaching process and sequentially extracted to Ni complex with an extracting agent in the extracting process. The solutions of nitric acid ($HNO_3$), ammonium carbonate ($(NH_4)_2CO_3$) and sodium carbonate ($Na_2CO_3$) were used as a leaching agent in leaching process and oxalic acid was used as an extracting agent in extracting process. $HNO_3$ solution is the most efficient leaching agent among the various leaching agent. Also, the optimized leaching conditions for the efficient and selective Ni recovery were the leaching temperature of $90^{\circ}C,\;HNO_3$ concentration of 6.25 vol% and elapsed time of 3 h. As a result, Nickel oxalate having the highest yield of 88.7% and purity of 100% was obtained after sequentially leaching and extracting processes under the optimized leaching conditions.

• Analytical Science and Technology
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• v.37 no.1
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• pp.63-78
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• 2024

The continuous increasing of the global demand of copper and nickel metals raises the interest in developing alternative technologies to produce them from copper sulfide ore. Also, in line with Egypt's vision 2030 for achieving the sustainable socioeconomic development which aims at developing alternative and eco-friendly technologies for processing the Egyptian ores to produce these strategic products instead of its importing. These metals enhance the advanced electrical and electronic industries. The current work aims at investigating the recovery of copper and nickel from Abu Swayeil copper ore using pug leaching technique by sulfuric acid. The factors affecting the pug leaching process including the sulfuric acid concentration, leaching time and temperature have been investigated. The copper ore sample was characterized chemically using X-ray fluorescence (XRF) and scanning electron microscope (SEM-EDX). A response surface methodology develops a quadratic model that expects the nickel and copper leaching effectiveness as a function of three controlling factors involved in the procedure of leaching was also investigated. The obtained results showed that the maximum dissolution efficiency of Ni and Cu are 99.06 % and 95.30%, respectively which was obtained at the following conditions: 15 % H₂SO₄ acid concentration for 6 hr. at 250 ℃. The dissolution kinetics of nickel and copper that were examined according to heterogeneous model, indicated that the dissolution rates were controlled by surface chemical process during the pug leaching. The activation energy of copper and nickel dissolution were 26.79 kJ.mol^-1 and 38.078 kJ.mol^-1 respectively; and the surface chemical was proposed as the leaching rate-controlling step.