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1

Kinetics study and reaction mechanism for titanium dissolution from rutile ores and concentrates using sulfuric acid solutions

Ismael Mohamed H., Mohammed Hesham S., El Hussaini Omneya M., El-Shahat Mohamed F.

Physicochemical Problems of Mineral Processing

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2022

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Vol. 58, iss. 1

138--148

EN

Recent developments of acid leaching of titanium concentrates and ores have produced renewed industrial and commercial interest. However, the leaching kinetics and mechanism of these concentrates and ores had received little attention. This work, therefore, addresses the leaching kinetics and mechanism of Ti from a rutile concentrate in sulfuric acid solution. The leaching reaction was controlled by diverse parameters like temperature, particle size, acid concentration, liquid/solid (L/S) ratio, and stirring speed. The leaching kinetics was investigated using the Shrinking Core Model in order to determine the optimum criteria which control the reaction. The kinetics analysis showed that the rate of dissolution of Ti increased by increasing reaction temperature, L/S ratio, and stirring speed, while it decreased upon increasing particle size. The kinetics analysis revealed that the dissolution reaction is controlled by the chemical reaction at the rutile particle surface. Applying the Arrhenius relation, the apparent energy of activation Ea for the leaching reaction was calculated to be 23.4kJ/mol. A semi-empirical overall rate equation was introduced to describe the combined effects of the process variables upon the rate of the dissolution reaction: [formula]

2

Experimental and kinetic study of zinc leaching from metallurgical slag by 5-sulfosalicylic acid

Wang Long, Gao Hanyu, Song Shimin, Xue Na, Zhang Jinxia, Yang Siyuan, Liu Cheng

Physicochemical Problems of Mineral Processing

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2021

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Vol. 57, iss. 3

8--20

EN

As an organic acid with the characters of low corrosivity and extensive source, 5-sulfosalicylic acid (5-SSA) was firstly utilized as a potential leaching reagent for the recovery of zinc from metallurgical slag. Effects of stirring speed, leaching temperature, 5-SSA concentrations and size fraction on the leaching zinc leaching rate were investigated. A zinc leaching efficiency of 94.2% was achieved under the appropriate operating conditions (450 rpm of stirring speed, 50 ℃ of leaching temperature, 0.3 mol/L of 5-SSA concentration and d90=65 µm of size fraction), indicating that 5-SSA was an excellent leaching reagent of zinc oxide. SEM-EDS and specific surface aperture analyzer further reveal the well-developed micropores and cracks from zinc metallurgical slag, which could be assigned to the removal of zinc oxide encapsulated in the sample. In addition, the leaching kinetics of zinc metallurgical slag in the 5-SSA was studied. It was found that the surface chemical reaction model satisfactorily predicted the zinc leaching rate. A reaction kinetic equation was finally established for the zinc leaching rate.

3

Effect of externally adding pyrite and electrical current on galvanic leaching of chalcopyrite concentrate

Asgari Kaveh, Hassanzadeh Ahmad, Nazari Sabereh, Vakylabad Ali Behrad, Hosseinzadeh Mostafa

Physicochemical Problems of Mineral Processing

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2021

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Vol. 57, iss. 2

105--119

EN

Although the operating properties of GalvanoxTM leaching have been widely studied in the literature, several factors concerning chalcopyrite passivation during the process remain unknown so far. The present work hence aims at investigating the significant effect of externally added pyrite features with a particular focus on its particle size (d80 of 0.52, 20, 45 and 2000 µm) through a series of experiments performed in a 2-L stirred-tank electro-reactor. To this end, the role of pyrite: chalcopyrite ratio (0.49:1, 2:1 and 4:1) and presence of electrical current were examined while the rest of the parameters kept constant (80 °C temperature, 400–500 mV (Ag/AgCl) redox potential, pulp density of 10% (w/v), and stirring rate of 1200 rpm). Plus, kinetic models of the leaching tests were studied based on the diffusion and chemical controlling concepts. It was found that the coarser the pyrite particles, the more favorable the copper extraction from the concentrate due to acceleration of reactions in the cathodic electrode and high mass transfers. However, this was in contradiction with the existing reports in the literature. Moreover, galvanic interactions became intensive in the presence of pyrite meaning extensive chalcopyrite dissolution with significantly reduced passivation. Ultimate copper extraction values of 24.17±1.25%, 55.79±0.91% and 57.26±1.59% were resulted at Py:Cp ratios of 0.49:1 (natural), 2:1 and 4:1, respectively. The results showed that maximum copper recovery of 67.32±2.34% was obtained at an optimum condition of pyrite grain size=2000 µm, Py:Cp=4:1, current application=500 mA, 8 h and 80 °C. Finally, detailed kinetic modeling indicated that the chemical control mechanism was dominant in the early reaction stages (t<3.5 h) concerning the availability of fresh surface for chemical agents; however, the second half of the process (8.0 h>t>3.5 h) was controlled by the diffusion control.

4

Removal of calcium from magnesite flotation concentrate by selective leaching and kinetics analysis

Hu Xiaoxing, Zhu Yangge

Physicochemical Problems of Mineral Processing

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2020

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Vol. 56, iss. 5

939--948

EN

Dolomite is the main impurity mineral in magnesite ore, affecting the quality of magnesite products. This study proposed a selective leaching process to reduce the calcium content from the magnesite flotation concentrate using hydrochloride acid (HCl). Laboratory scale tests were conducted to explore the influence of operation factors including HCl concentration, leaching temperature and time, stirring speed on the leaching recovery. The results showed that leaching recovery of CaO and MgO increased with increasing HCl concentration, temperature and time. The CaO leaching recovery is always significantly higher than that of MgO, indicating a good selectivity of HCl to leach dolomite from magnesite. The leaching kinetics of both dolomite and magnesite follow the Avrami model and are diffusion controlled, with an activation energy of 43±}1 kJ×mol-1 and 25±}4 kJ×mol-1, respectively.

5

Leaching kinetics of iron from coal-series kaolin by oxalic acid solutions

Shao Hui, Ji Mengjiao, Yu Mingming

Physicochemical Problems of Mineral Processing

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2020

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Vol. 56, iss. 1

134--146

EN

Removal of iron from coal-series kaolinite and its kinetics by chemical leaching were investigated in this paper. The optimal removal rate of iron reaches 29.09% (the content of Fe2O3 decreased from 0.55 to 0.39%) under the conditions of 100 °C, 0.4 M C2H2O4, and liquid-to-solid ratio of 4 cm3/g for 120 min. The leaching process was successfully described by the Avrami model and mainly controlled by internal diffusion control of shrinking core model, the apparent activation energy is 18.46 kJ/mol. The experimental results were consistent with the internal diffusion control of shrinking core model, improving acid concentration and leaching temperature could promotes the removal of iron from coal-series kaolin.

6

Recovery of Eu and Y from Waste Fluorescent Lamps

Vu H. N., Pham T. D., Formanek J., Dvorak P.

Inżynieria Mineralna

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2017

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R. 18, nr 1

23--28

EN

Eu and Y were extracted from waste fluorescent lamps containing ~0.3% Eu and 7% Y by acidic leaching, hydrolytic precipitation and/or solvent extraction. The leaching tests showed that about 90% Eu and 95% Y were leached in 3M HCl or 3M HNO3 at 80°C, liquid to solid ratio 10:1 in 30 min. Leaching in H2SO4 provided lower Eu and Y extraction efficiency. Only around 85% Eu and 80% Y were extracted at temperatures higher than 70°C. Eu and Y started precipitating simultaneously at pH 1.5 and the precipitation completed at ~ pH 2. Solvent extraction of Eu and Y from sulfate solutions using D2EHPA is strongly influenced by solution pH. At pH 1.4 about 80% Eu and 100%Y were extracted while at pH 1.16 only 25% Eu was extracted together with 90% Y.

PL

Eu i Y były odzyskiwane z odpadów lamp fluorescencyjnych zawierających ok. 0.3% Eu i 7% Y za pomocą metody ługowania kwaśnego, wytrącania hydrolitycznego i/lub ekstrakcji rozpuszczalnikowej. Próby ługowania wykazały, że około 90% Eu i 95% Y można wyługować w 3M roztworze HCl lub 3M roztworze HNO3 w temperaturze 80°C, stosunku fazy ciekłej do stałej 10:1 w czasie 30 min. Ługowanie w H2SO4 daje niższy współczynnik odzysku Eu i Y. Jedynie około 85% Eu i 80% Y odzyskano w temperaturze wyższej niż 70°C. Eu i Y zaczęły wytrącać się jednocześnie w pH 1.5, a wytrącanie zakończyło się przy pH około 2. Ekstrakcja rozpuszczalnikowa Eu i Y z roztworów siarczanów z wykorzystaniem D2EHPA zależy silnie od pH roztworu. Przy pH = 1.4 około 80% Eu i 100%Y zostało wydzielone, podczas gdy przy pH = 1.16 wyekstrahowano tylko 25% Eu i 90% Y.

7

Investigation of kinetics and mechanism of priceite leaching in sulphuric acid solutions

Gur A.

Physicochemical Problems of Mineral Processing

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2015

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Vol. 51, iss. 2

719--730

EN

The leaching kinetics and mechanism of priceite having the formula of 4CaO5B2O37H2O was investigated in sulphuric acid solutions. For the dissolution process, the effects of reaction temperature, sulphuric acid solution concentration, solid/liquid ratio, particle size, and stirring speed were investigated as effective parameters for the experiments. The experimental data indicated that the dissolution rate increased with the increasing reaction temperature, the decreasing particle size, and the solid/liquid ratio. The conversion rate increased up to 1.0 mol/dm3 with the increasing acid concentration, and then decreased with the increasing acid concentration over concentrations of 1 mol/dm3. It was also determined that the stirring speed had no significant effect on the dissolution rate. The dissolution process of priceite in sulphuric acid solution was tested with regard to heterogeneous and homogeneous reaction models, and it was found that the dissolution rate was controlled by first order pseudo-homogeneous reaction model. The activation energy of the dissolution process was determined as 26.07 kJmol-1. The results were evaluated graphically and statistically. The experimental data were found to fit well with the mathematical model.