Mineral Processing
Mohammadreza Shahbazi; Hadi Abdollahi; Sied Ziaeddin Shafaei; Ziaeddin Pourkarimi; Sajjad Jannesar Malakooti; Ehsan Ebrahimi
Abstract
Tabas coal possesses favorable plastometric properties that make it suitable for use in metallurgical industries as coking coal. However, its high sulfur content, which stands at approximately 2%, poses a significant environmental pollution risk. Additionally, reducing ash content to below 10% is a critical ...
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Tabas coal possesses favorable plastometric properties that make it suitable for use in metallurgical industries as coking coal. However, its high sulfur content, which stands at approximately 2%, poses a significant environmental pollution risk. Additionally, reducing ash content to below 10% is a critical objective of this study to prevent a decline in coal's thermal efficiency in the metallurgical industries. This research work investigates the removal of sulfur and ash from Tabas coal samples using the biological methods including bioflotation and bioleaching. Initially, a combination of mesophilic bacteria including Acidithiobacillus ferrooxidans, Acidithiobacillus thiooxidans, and Leptosprillium ferrooxidans were employed in the bioflotation method to detain pyrite sulfur in the Tabas coal samples. The highest reduction percentages of pyrite sulfur and ash were equal to 62% and 54.18%, respectively. In the next stage, bioleaching experiments were conducted, the effect of the test time, percentage of bacteria by volume, percentage of coal solids, and absence of bacteria on the amount of sulfur and ash removal was investigated. The test time emerged as the most critical factor. The best sulfur removal was achieved using bioleaching, with a maximum removal of 72.43%, observed for the PE coal sample. Bioflotation also achieved significant sulfur removal, with a maximum removal of 61% observed for the same sample. On the other hand, the best ash removal was achieved using bioflotation, with a maximum removal of 68.98% observed for the PE coal sample, and a maximum removal of 69.34% observed for the B4B2 coal sample using bioleaching. Finally, this research work conducted a comparison of biological methods to determine the amount of sulfur and ash reduction achieved. The results showed that both bioleaching and bioflotation were effective for coal desulfurization and ash removal, with bioleaching performing slightly better for sulfur removal and bioflotation performing slightly better for ash removal.
Laleh Sohbatzadeh; Sied Ziaedin Shafaei Tonkaboni; Mohammad Noaparast
Abstract
In this research work, with a simple, safe, and environmentally friendly approach to hydrometallurgy, a method for the recovery of lithium (Li), cobalt (Co), and nickel (Ni) from LIBs is suggested. The cathode materials are leached by malonic acid, as the leaching agent, and ascorbic acid, as the reducing ...
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In this research work, with a simple, safe, and environmentally friendly approach to hydrometallurgy, a method for the recovery of lithium (Li), cobalt (Co), and nickel (Ni) from LIBs is suggested. The cathode materials are leached by malonic acid, as the leaching agent, and ascorbic acid, as the reducing agent in the first process, and by L-glutamic acid, as the leaching agent, and ascorbic acid, as the reducing agent in the second process. In order to optimize the leaching parameters including temperature, organic acid concentration, ascorbic acid concentration, type of organic acid, pulp density, and time, response surface methodology (RSM) of the experimental design process is used. According to the results, compared to L-glutamic acid in the second process, the leaching recovery increase considerably with malonic acid in the first process. This normally occurs due to the higher solubility of malonic acid in water, which results in a better complexation and a higher chelation rate. By contrast, as solubility of L-glutamic acid in water is low, metal-acid surface reaction and poor complexation are unavoidable. According to the statistical analysis of the results and validation testing, optimal experimental leaching occurs at the reaction temperature of 88 °C, organic acid concentration of 0.25 M, ascorbic acid concentration of 0.03 M, pulp density of 10 g/L, and leaching time of 2 h, via which metal recovery of 100% Li, 81% Co, and 99% Ni is achieved. Before and after acidic leaching, the sample active materials are qualitatively and quantitatively analyzed using X-ray diffraction, X-ray fluorescence, particle size analyzer, scanning electron microscope, energy dispersive spectroscopy, and atomic absorption spectroscopy.
Mineral Processing
V. Radmehr; Seyed Z. Shafaei; M. Noaparast; H. Abdollahi
Abstract
This paper presents a new approach for flotation circuit design. Initially, it was proven numerically and analytically that in order to achieve the highest recovery in different circuit configurations, the best equipment must be placed at the beginning stage of the flotation circuits. The size of the ...
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This paper presents a new approach for flotation circuit design. Initially, it was proven numerically and analytically that in order to achieve the highest recovery in different circuit configurations, the best equipment must be placed at the beginning stage of the flotation circuits. The size of the entering particles and the types of streams including pulp and froth were considered as the basis for specialization of the flotation processes. In the new approach, the flotation process plays as the two functions of primary and secondary concentrations. The proposed approach was applied to a lead flotation circuit of a lead-zinc flotation plant. The results obtained showed that in most traditional-oriented circuits, a large part of the streams containing valuable metals were returned to the rougher stage, which, in turn, reduced the efficiency and caused perturbation. In the new approach, providing more control over unit operations in the circuit could provide a higher performance. In addition, in cases where zinc minerals are liberated from their gangue in coarse size, the new approach, by generating coarse-grained tailing, can prevent excessive grinding of zinc minerals in the feed into the zinc flotation circuit.
Mineral Processing
M. Salehfard; M. Noaparast; Seyed Z. Shafaei; H. Abdollahi
Abstract
A lead-zinc carbonate ore sample containing 2.5% Pb and 9.39% Zn was used in this research work. The sample was prepared from the Darreh-Zanjir mine located in the Yazd province (Iran). Influences of the influential factors on flotation of smithsonite and cerussite were investigated. Among the different ...
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A lead-zinc carbonate ore sample containing 2.5% Pb and 9.39% Zn was used in this research work. The sample was prepared from the Darreh-Zanjir mine located in the Yazd province (Iran). Influences of the influential factors on flotation of smithsonite and cerussite were investigated. Among the different parameters involved, dosages of the dispersant, depressants, sulfidizing agent, and collectors de-sliming prior to lead or zinc flotation were essential for the effective recovery and grade of the Zn and Pb flotation concentrates. In addition, the anionic, cationic, and mixed (cationic/anionic) collectors were employed for flotation of smithsonite. The results of reverse and cumulative flotation of both Zn and Pb were relatively low in comparison with the direct process without depressant. Flotation of smithsonite using mixed collectors (Armac C+KAX) showed promising results. Also dosages of chemicals in the cleaning stage for the Zn and Pb concentrates were optimized, and finally, after the cleaner stage for both lead and zinc, a cerussite concentrate with Pb grade and recovery of 49.82% and 60.06%, respectively, and smithsonite concentrate with Zn grade and recovery of 35.47% and 68.56%, respectively, were obtained under the optimal conditions. Furthermore, kinetics of Zn and Pb oxide mineral flotations in the rougher and cleaner stages were studied, which showed that some kinetics models, especially the classical first-order model, could predict the flotation behaviour of the Zn and Pb oxide minerals.
Mineral Processing
S. Nazari; Seyed Ziaedin Shafaei; M. Gharabaghi; R. Ahmadi; B. Shahbazi
Abstract
In this work, the effects of the types of frother (MIBC, pine oil, and A65) and operational parameters (impeller speed and air flow rate) on the flotation of quartz coarse particles was investigated using nano bubbles (NBs). Quartz particles of the size of -425+106 mm and three types of frother were ...
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In this work, the effects of the types of frother (MIBC, pine oil, and A65) and operational parameters (impeller speed and air flow rate) on the flotation of quartz coarse particles was investigated using nano bubbles (NBs). Quartz particles of the size of -425+106 mm and three types of frother were used for the flotation experiments. Also the impeller speed was 600 to 1300 rpm, and the air flow rates were 30 and 60 L/h. In the absence of NBs, the maximum recovery was achieved with the pine oil frother, an impeller speed of 1000 rpm, and an air flow rate of 60 L/h. In the presence of NBs, the maximum recovery was achieved using pine oil at an impeller speed of 900 rpm and an air flow rate of 30 L/h. However, increasing the recovery in the presence of NBs, compared to the absence of NBs for MIBC, was more than the other two frothers, and the recovery using this frother to increase up to 25% but using pine oil, the recovery increased up to 23%. The lowest recovery in the presence of NBs was obtained using A65. Also the use of NBs increased recovery in all the three fractions compared to the absence of NBs but the presence of NBs increased the recovery of particles with size of -212+106 mm more than the particle size in the ranges of -300+212 and -425+300 mm.
Mineral Processing
M. Naderi; Seyed Z. Shafaie; M. Karamoozian; Sh. Gharanjik
Abstract
In this work, the parameters affecting the recovery of copper from the low-grade sulfide minerals of Sarcheshmeh Copper Mine were studied. A low-grade sulfide ore was used with a copper grade of 0.25%, which was about 28% of the mineral oxide, and the sulfide minerals made up the rest. Much more sulfide ...
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In this work, the parameters affecting the recovery of copper from the low-grade sulfide minerals of Sarcheshmeh Copper Mine were studied. A low-grade sulfide ore was used with a copper grade of 0.25%, which was about 28% of the mineral oxide, and the sulfide minerals made up the rest. Much more sulfide minerals were found to be pyrite and most of the gangue minerals were quartz, anorthite, biotite, and muscovite. In order to investigate, simultaneously, the solids (10 to 20%) and acidity (1.5 to 2.5) and shaking (110 to 150 rpm), the separation of bacteria from Sarcheshmeh Copper Mine was carried out. After adjustment of the sample, bio-leaching tests were performed in accordance with the pattern defined by the software DX7 in shaking flasks, and the Cu recovery was modeled and optimized using the response surface methodology. The influential parameters were comprehensively studied. The central composite design methodology was used as the design matrix to predict the optimal level of these parameters. Then the model equation was optimized. The results obtained showed that increasing solids (from 10 to 20%) was bad for bacteria. The highest copper recovery was equivalent to 69.91%, obtained after 21 days at 35 degrees using the Acidi Thiobacillus Ferrooxidans bacteria and a K9 medium with a pulp density of 10% and pH 1.5.
Seyyed M. Seyyed Alizadeh Ganji; S. Z. Shafaie; N. Goudarzi
Abstract
This work was aimed to evaluate and compare the performances of the solvents D2EHPA (Di-(2-ethylhexyl) phosphoric acid), Cyanex 272 (bis (2,4,4-trimethylpentyl) phosphinic acid), and a mixture system of D2EHPA and Cyanex272 in the separation of some rare earth elements (REEs) including lanthanum, gadolinium, ...
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This work was aimed to evaluate and compare the performances of the solvents D2EHPA (Di-(2-ethylhexyl) phosphoric acid), Cyanex 272 (bis (2,4,4-trimethylpentyl) phosphinic acid), and a mixture system of D2EHPA and Cyanex272 in the separation of some rare earth elements (REEs) including lanthanum, gadolinium, neodymium, and dyspersym from a nitric acid solution. The results obtained showed that Cyane272 had the lowest separation factor in the separation of Dy, La, Nd, and Gd from each other. Also it was found that a mixture system of D2EHPA and Cyanex272 had the best performance in the separation of the investigated REEs, owing to the higher separation factors for Dy/Nd and Dy/Gd, as well as the lower extraction efficiencies for Gd (64.54%), La (30.07%), and Nd (26.47) from Dy (99.92). It was also determined that the separation factors forDy/Nd and Dy/Gd were 720.05 and 3640.27, respectively, using their mixture system.
M. Shamsi; M. Noparast; Seyyed Z. Shafaie; M. Gharabaghi; S. Aslani
Abstract
Copper smelting slags are hard materials. Therefore,to recover their copper by flotation method, grinding should be carried out to obtain optimal particle size. Copper smelting slags in the Bardeskan district, with work index of 16.24 kwh/st, were grinded for 65 minutes to reach an acceptable degree ...
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Copper smelting slags are hard materials. Therefore,to recover their copper by flotation method, grinding should be carried out to obtain optimal particle size. Copper smelting slags in the Bardeskan district, with work index of 16.24 kwh/st, were grinded for 65 minutes to reach an acceptable degree of freedom for the flotation tests, with particle size of 80%, smaller than 70 μm. With this grinding time, degree of freedom for copper-bearing minerals was achieved 85-90%. The floatation method performed and the procedure used for the optimization of the effective parameters were described in this paper. The results obtained for the flotation tests, carried out at the optimal conditions after grinding the slags (with a grinding time of65 minutes), showed 62.23% of copper recovery, while, by flotation of copper slags at optimal conditions after increasing the grinding time to 85 minutes (d80 = 48µ), the Cu recovery was increased to 79.89%.
Bahman Ghobadi; Mohammad Noaparast; Seid Ziaedin Shafaei; Majid Unesi
Abstract
This study investigates the optimization of gold dissolution from Aghdarre ore. Therefore, a laboratory investigation was initiated, to improve the leaching conditions with the objective of maximizing mill capacity with no reduction on gold recovery. It was observed that the time reduction from 25 to ...
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This study investigates the optimization of gold dissolution from Aghdarre ore. Therefore, a laboratory investigation was initiated, to improve the leaching conditions with the objective of maximizing mill capacity with no reduction on gold recovery. It was observed that the time reduction from 25 to 15 hours did not change the gold recovery. In the other words, it indicated that a capacity of 140t/h can be sustained without detrimental effect on gold recovery. The optimum parameters were 700g/t NaCN, 46% solid in pulp, pH=10, and d80=45 microns using the Taguchi method. So, the gold recovery was obtained 90.71%. Also, it was concluded that the NaCN concentration was the most effective parameter and the solid percent plus retention time had the lowest effects on this process.
