Pub Date : 2024-09-24DOI: 10.1016/j.mineng.2024.109006
Copper oxide ore is the important component of copper resources and cannot be effectively recovered by conventional beneficiation method. In this work, the low-content and complex copper oxide ore was efficiently recovered by magnetic-leaching process. Under the optimized magnetic separation condition (feed ore size: −2 mm, rod matrix size: 3 mm, magnetic induction intensity: 1.5 T, pulsating frequency: 100 r/min), copper grade of the concentrate was enriched from 1.50 % to 8.60 % with copper recovery of 76.63 %. Compared with direct leaching, magnetic-leaching process decreased acid consumption for 1 t Cu from 13.63 t to 2.92 t. Magnetic susceptibility measurement and DFT calculation results indicated that the Cu atom of played a critical role in determining magnetism for copper-containing minerals.
{"title":"Recover of Cu from copper oxide ore using magnetic separation-leaching process and magnetism study of copper-containing minerals","authors":"","doi":"10.1016/j.mineng.2024.109006","DOIUrl":"10.1016/j.mineng.2024.109006","url":null,"abstract":"<div><div>Copper oxide ore is the important component of copper resources and cannot be effectively recovered by conventional beneficiation method. In this work, the low-content and complex copper oxide ore was efficiently recovered by magnetic-leaching process. Under the optimized magnetic separation condition (feed ore size: −2 mm, rod matrix size: 3 mm, magnetic induction intensity: 1.5 T, pulsating frequency: 100 r/min), copper grade of the concentrate was enriched from 1.50 % to 8.60 % with copper recovery of 76.63 %. Compared with direct leaching, magnetic-leaching process decreased acid consumption for 1 t Cu from 13.63 t to 2.92 t. Magnetic susceptibility measurement and DFT calculation results indicated that the Cu atom of played a critical role in determining magnetism for copper-containing minerals.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-24","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314257","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-24DOI: 10.1016/j.mineng.2024.108911
The useful mineral apatite in phosphate ore has a fine embedded particle size, requiring fine grinding for monomer dissociation, which will lead to a large number of fine apatite generation, resulting in low flotation efficiency and other problems. In this study, the effects of pH, agent concentration and stirring rate upon agglomeration behavior and rheological properties of fine apatite were investigated. The reasonable agglomeration mechanisms were proposed by means of Fourier Transform Infrared Spectroscopy (FTIR), X-ray Photoelectron Spectroscopy (XPS), zeta potential analysis. Moreover, the interaction between particles was evaluated by using the rheological parameter apparent viscosity, and the internal connection between fine apatite agglomeration and pulp rheology was revealed. The results indicate that at an oleic acid (OA) concentration of 400 mg/L, stirring rate of 500 rpm, and pH of 9, the average particle size (dmean) of apatite increases from the original 7.37 μm to 104.7 μm, and the apparent viscosity rises from 2.18 mPa·s to 17.40 mPa·s. In the process of fine apatite agglomeration, OA is adsorbed onto the surface of apatite particles through chemical complexation, which makes the apatite surface hydrophobic, and thus causing agglomeration between apatite particles due to hydrophobic attraction. In addition, the flotation models of apatite were proposed before and after agglomeration, and there is a correlation among particle agglomeration, apparent viscosity and particle interactions, forming a triangular relationship similar to an “iron triangle”. This study provides novel insights into the regulation of hydrophobic agglomeration of fine particles.
{"title":"Study of oleic acid-induced hydrophobic agglomeration of apatite fines through rheology","authors":"","doi":"10.1016/j.mineng.2024.108911","DOIUrl":"10.1016/j.mineng.2024.108911","url":null,"abstract":"<div><div>The useful mineral apatite in phosphate ore has a fine embedded particle size, requiring fine grinding for monomer dissociation, which will lead to a large number of fine apatite generation, resulting in low flotation efficiency and other problems. In this study, the effects of pH, agent concentration and stirring rate upon agglomeration behavior and rheological properties of fine apatite were investigated. The reasonable agglomeration mechanisms were proposed by means of Fourier Transform Infrared Spectroscopy (FTIR), X-ray Photoelectron Spectroscopy (XPS), zeta potential analysis. Moreover, the interaction between particles was evaluated by using the rheological parameter apparent viscosity, and the internal connection between fine apatite agglomeration and pulp rheology was revealed. The results indicate that at an oleic acid (OA) concentration of 400 mg/L, stirring rate of 500 rpm, and pH of 9, the average particle size (<em>d<sub>mean</sub></em>) of apatite increases from the original 7.37 μm to 104.7 μm, and the apparent viscosity rises from 2.18 mPa·s to 17.40 mPa·s. In the process of fine apatite agglomeration, OA is adsorbed onto the surface of apatite particles through chemical complexation, which makes the apatite surface hydrophobic, and thus causing agglomeration between apatite particles due to hydrophobic attraction. In addition, the flotation models of apatite were proposed before and after agglomeration, and there is a correlation among particle agglomeration, apparent viscosity and particle interactions, forming a triangular relationship similar to an “iron triangle”. This study provides novel insights into the regulation of hydrophobic agglomeration of fine particles.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-24","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314258","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-24DOI: 10.1016/j.mineng.2024.109005
Copper, nickel, iron, zinc, lead, and molybdenum are among the important metals commonly found as sulfide minerals in nature. Some of these metals are also found as sulfides in intermediate products like matte and mixed sulfide precipitates (MSP). The extraction processes of these metals from their sulfide forms through either pyrometallurgy or hydrometallurgy encounter various challenges. This study aimed to explore the viability of solvometallurgical processes (solvoleaching) as potential alternatives for the extraction of these metals from their sulfides (CuS, Ni3S2, FeS, ZnS, PbS and MoS2). The solvometallurgical processes employed a range of solvoleaching systems, such as D2EHPA+MnO2+H2O, ammoniacal solvoleaching comprising of LIX 84-I+NH4OH+H2O2, HCl-equilibrated TBP (TBP-HCl) and EG (ethylene glycol)-based solvents containing each of HCl, FeCl3, ChCl and NH4Cl. The solvoleaching mechanism of the metal sulfides in each solvoleaching system was discussed. The results were corroborated with thermodynamic tools like Eh-pH diagrams, Gibbs free energy change, and solvation properties of the metal ions, as well as the hard and soft acid-base (HSAB) theory. Analyses of the resulting leach solutions, utilizing suitable analytical techniques such as UV–Vis spectroscopy and FTIR, provided additional support for the results. A comparison made between the solvoleaching and classical aqueous leaching, showing comparable results, indicated the viability of solvoleaching processes as alternatives for the sustainable extraction of these metals from their sulfides.
{"title":"Investigation on solvometallurgical processes for extraction of metals from sulfides","authors":"","doi":"10.1016/j.mineng.2024.109005","DOIUrl":"10.1016/j.mineng.2024.109005","url":null,"abstract":"<div><div>Copper, nickel, iron, zinc, lead, and molybdenum are among the important metals commonly found as sulfide minerals in nature. Some of these metals are also found as sulfides in intermediate products like matte and mixed sulfide precipitates (MSP). The extraction processes of these metals from their sulfide forms through either pyrometallurgy or hydrometallurgy encounter various challenges. This study aimed to explore the viability of solvometallurgical processes (solvoleaching) as potential alternatives for the extraction of these metals from their sulfides (CuS, Ni<sub>3</sub>S<sub>2</sub>, FeS, ZnS, PbS and MoS<sub>2</sub>). The solvometallurgical processes employed a range of solvoleaching systems, such as D2EHPA+MnO<sub>2</sub>+H<sub>2</sub>O, ammoniacal solvoleaching comprising of LIX 84-I+NH<sub>4</sub>OH+H<sub>2</sub>O<sub>2</sub>, HCl-equilibrated TBP (TBP-HCl) and EG (ethylene glycol)-based solvents containing each of HCl, FeCl<sub>3</sub>, ChCl and NH<sub>4</sub>Cl. The solvoleaching mechanism of the metal sulfides in each solvoleaching system was discussed. The results were corroborated with thermodynamic tools like Eh-pH diagrams, Gibbs free energy change, and solvation properties of the metal ions, as well as the hard and soft acid-base (HSAB) theory. Analyses of the resulting leach solutions, utilizing suitable analytical techniques such as UV–Vis spectroscopy and FTIR, provided additional support for the results. A comparison made between the solvoleaching and classical aqueous leaching, showing comparable results, indicated the viability of solvoleaching processes as alternatives for the sustainable extraction of these metals from their sulfides.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-24","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"https://www.sciencedirect.com/science/article/pii/S0892687524004345/pdfft?md5=c18c9d0db11fc5f3bdfb22b9aa3eb366&pid=1-s2.0-S0892687524004345-main.pdf","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314255","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"OA","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-23DOI: 10.1016/j.mineng.2024.109012
Accelerated production of lithium-ion batteries (LIBs) implies an increase in the raw materials demand, especially for metals like lithium, cobalt, and nickel. Spent LIBs recycling guarantees the regeneration and reincorporation of valuable materials into the manufacturing industry; therefore, recycling methods and techniques must be optimized. In this investigation, alkaline and reductive acid leaching processes were evaluated and compared in order to determine the effect of parameters such as pH, temperature, and reagents concentrations to achieve selective leaching processes. This study demonstrated that strongly alkaline solutions (NaOH) do not ensure selective lithium and aluminum dissolution. Also, a solid compound, can be formed at pH ∼14, negatively affecting the lithium extraction. On the other hand, reductive acid leaching, with acid sulfuric and hydrazine sulfate (H2SO4 + N2H6SO4) solutions resulted in an efficient system, extracting ≥90 % of Ni, Co, and Mn at 40 °C. Hydrazine is essential as a reductant, although it must be added in excess (40 % excess with respect to the Co, Ni, and Mn content) to suppress copper dissolution. Furthermore, this work demonstrated the possibility of processing the entire spent LIBs sample once the discharge and crushing stages were concluded, avoiding physical separation, without affecting the leaching efficiency and contributing to process economy.
{"title":"A comparative study of discharging and leaching of spent lithium-ion battery recycling","authors":"","doi":"10.1016/j.mineng.2024.109012","DOIUrl":"10.1016/j.mineng.2024.109012","url":null,"abstract":"<div><div>Accelerated production of lithium-ion batteries (LIBs) implies an increase in the raw materials demand, especially for metals like lithium, cobalt, and nickel. Spent LIBs recycling guarantees the regeneration and reincorporation of valuable materials into the manufacturing industry; therefore, recycling methods and techniques must be optimized. In this investigation, alkaline and reductive acid leaching processes were evaluated and compared in order to determine the effect of parameters such as pH, temperature, and reagents concentrations to achieve selective leaching processes. This study demonstrated that strongly alkaline solutions (NaOH) do not ensure selective lithium and aluminum dissolution. Also, a solid compound, <span><math><mrow><mi>L</mi><mi>i</mi><msub><mrow><mi>A</mi><mi>l</mi></mrow><mn>2</mn></msub><msub><mrow><mo>(</mo><mi>O</mi><mi>H</mi><mo>)</mo></mrow><mn>6</mn></msub><msub><mrow><mi>O</mi><mi>H</mi></mrow><mrow><mo>(</mo><mi>s</mi><mo>)</mo></mrow></msub></mrow></math></span> can be formed at pH ∼14, negatively affecting the lithium extraction. On the other hand, reductive acid leaching, with acid sulfuric and hydrazine sulfate (H<sub>2</sub>SO<sub>4</sub> + N<sub>2</sub>H<sub>6</sub>SO<sub>4</sub>) solutions resulted in an efficient system, extracting ≥90 % of Ni, Co, and Mn at 40 °C. Hydrazine is essential as a reductant, although it must be added in excess (40 % excess with respect to the Co, Ni, and Mn content) to suppress copper dissolution. Furthermore, this work demonstrated the possibility of processing the entire spent LIBs sample once the discharge and crushing stages were concluded, avoiding physical separation, without affecting the leaching efficiency and contributing to process economy.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-23","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314252","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-23DOI: 10.1016/j.mineng.2024.109007
Crompton et al. (2023) developed a new algorithm for describing the performance of coarse particle flotation. They used the flotation rate constant, k, normalised by the maximum rate constant, kmax, for the pure mineral, as a proxy for the fractional surface liberation. The algorithm was used to produce the partition curve for a separation performed by a novel device, the CoarseAIR™. Part I of this new study re-visits the former work, particularly the batch mechanical flotation responses of the steady state samples from the CoarseAIR™. The flotation responses were deconvolved to the corresponding distributions of rate constants for the three streams, and in turn used to produce the partition curve for the coarse particle flotation. The algorithm used to produce the distribution of rate constants was driven towards a simple functional form by minimising its overall curvature. The steady state samples from any coarse particle flotation system can be assessed in this way. Part II of this study focuses on the reproducibility of the approach, and hence the uncertainty, using a batch mechanical cell to simulate the coarse particle flotation, and in turn the steady state feed, product and reject samples.
{"title":"A new method for assessing coarse particle flotation performance Part I − On the deconvolution of the flotation response","authors":"","doi":"10.1016/j.mineng.2024.109007","DOIUrl":"10.1016/j.mineng.2024.109007","url":null,"abstract":"<div><div><span><span>Crompton et al. (2023)</span></span> developed a new algorithm for describing the performance of coarse particle flotation. They used the flotation rate constant, k, normalised by the maximum rate constant, k<sub>max</sub>, for the pure mineral, as a proxy for the fractional surface liberation. The algorithm was used to produce the partition curve for a separation performed by a novel device, the CoarseAIR™. Part I of this new study re-visits the former work, particularly the batch mechanical flotation responses of the steady state samples from the CoarseAIR™. The flotation responses were deconvolved to the corresponding distributions of rate constants for the three streams, and in turn used to produce the partition curve for the coarse particle flotation. The algorithm used to produce the distribution of rate constants was driven towards a simple functional form by minimising its overall curvature. The steady state samples from any coarse particle flotation system can be assessed in this way. Part II of this study focuses on the reproducibility of the approach, and hence the uncertainty, using a batch mechanical cell to simulate the coarse particle flotation, and in turn the steady state feed, product and reject samples.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-23","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"https://www.sciencedirect.com/science/article/pii/S0892687524004369/pdfft?md5=35c4f31c91a262131e56ed5d52c08e9f&pid=1-s2.0-S0892687524004369-main.pdf","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314248","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"OA","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-23DOI: 10.1016/j.mineng.2024.108979
Separating complicated linked fluorite and calcium-containing gangue minerals efficiently is challenging with a standard oleate (NaOL) collector due to its limited selectivity and flexibility. In order to improve the efficiency of separating fluorite ore through flotation, a new collector called (E)-N-(2-hydroxyethyl)octadec-9-enamide (NOE) was created and compared to the standard collector sodium oleate (NaOL). The experimental results indicate that the optimal pH for flotation is 9, while the optimal dosages of NOE and NaOL are 150 g/t and 300 g/t, respectively. During bench-scale flotation studies, the use of a novel collector (NOE) at half the dose of the usual collector NaOL (175 g/t compared to 350 g/t) resulted in a 9.15 % improvement in the recovery of fluorite.
{"title":"Froth flotation separation of fluorite ore using(E)-N-(2-hydroxyethyl)octadec-9-enamide as the flotation collector","authors":"","doi":"10.1016/j.mineng.2024.108979","DOIUrl":"10.1016/j.mineng.2024.108979","url":null,"abstract":"<div><div>Separating complicated linked fluorite and calcium-containing gangue minerals efficiently is challenging with a standard oleate (NaOL) collector due to its limited selectivity and flexibility. In order to improve the efficiency of separating fluorite ore through flotation, a new collector called (E)-N-(2-hydroxyethyl)octadec-9-enamide (NOE) was created and compared to the standard collector sodium oleate (NaOL). The experimental results indicate that the optimal pH for flotation is 9, while the optimal dosages of NOE and NaOL are 150 g/t and 300 g/t, respectively. During bench-scale flotation studies, the use of a novel collector (NOE) at half the dose of the usual collector NaOL (175 g/t compared to 350 g/t) resulted in a 9.15 % improvement in the recovery of fluorite.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-23","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314242","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-23DOI: 10.1016/j.mineng.2024.109004
In this study, lithium extraction from a Li-rich kaolin is performed by roasting with Na2SO4 and water leaching. The thermal analysis of the Li-rich kaolin is characterized by TGA and DSC analysis. The comparison tests are conducted through roasting alone and H2SO4 leaching, in the aspects of ions concentration, Li recovery and solid wastes. Under Na2SO4 roasting and water leaching, 84 % of Li is leached out, which is a little lower than the H2SO4 method. Meanwhile, the leached Al and Fe are very low, which is highly beneficial to the purification process. After precipitation, a Li2CO3 product of 95 % in purity is obtained. However, for the H2SO4 method, Li2CO3 is hardly obtained because the Li is mainly lost in the purification process. And the process pH is much lower, leading to a high amount of solid wastes. This study might give a clue for the lithium recovery from Li-rich kaolin resources.
{"title":"Lithium extraction from a Li-rich kaolin resource through Na2SO4 roasting and water leaching","authors":"","doi":"10.1016/j.mineng.2024.109004","DOIUrl":"10.1016/j.mineng.2024.109004","url":null,"abstract":"<div><div>In this study, lithium extraction from a Li-rich kaolin is performed by roasting with Na<sub>2</sub>SO<sub>4</sub> and water leaching. The thermal analysis of the Li-rich kaolin is characterized by TGA and DSC analysis. The comparison tests are conducted through roasting alone and H<sub>2</sub>SO<sub>4</sub> leaching, in the aspects of ions concentration, Li recovery and solid wastes. Under Na<sub>2</sub>SO<sub>4</sub> roasting and water leaching, 84 % of Li is leached out, which is a little lower than the H<sub>2</sub>SO<sub>4</sub> method. Meanwhile, the leached Al and Fe are very low, which is highly beneficial to the purification process. After precipitation, a Li<sub>2</sub>CO<sub>3</sub> product of 95 % in purity is obtained. However, for the H<sub>2</sub>SO<sub>4</sub> method, Li<sub>2</sub>CO<sub>3</sub> is hardly obtained because the Li is mainly lost in the purification process. And the process pH is much lower, leading to a high amount of solid wastes. This study might give a clue for the lithium recovery from Li-rich kaolin resources.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-23","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314247","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-23DOI: 10.1016/j.mineng.2024.109001
Magnetic separation is a primary method for processing iron ore and plays a crucial role in both current beneficiation practices and other fields. Extensive research has been conducted on the motion behavior of magnetic particles within magnetic separation equipment. However, force analysis, particularly the calculation of magnetic forces, remains imprecise when dealing with irregularly shaped particles. Accurate prediction of magnetic particle behavior requires precise magnetic force calculations. This study introduces micromagnetic simulations to accurately compute the magnetic forces on irregular magnetic particles. Micromagnetic simulations can determine the precise magnetic moments and magnetic induction intensities within each microelement of the particle. The results of these simulations will be validated using magnetic force microscopy (MFM). The findings indicate that traditional magnetic force calculations deviate from the precise calculations presented in this study. For irregular particles, the computational errors in repulsive and attractive forces are 770% and 576% higher, respectively, compared to spherical particles. This underscores the necessity of considering particle shape in realistic magnetic force calculations. Additionally, both the MFM measurement images and the simulated magnetic force maps exhibit bright and dark regions correlated with particle shape, demonstrating that micromagnetic simulation results can be verified through MFM measurements. This paper proposes an experimentally verifiable method for accurately calculating the magnetic forces on magnetic particles using micromagnetic simulations, which holds significant implications for designing more efficient and precise magnetic separation equipment.
{"title":"Accurate calculation of magnetic forces on magnetic mineral particles using micromagnetic simulations","authors":"","doi":"10.1016/j.mineng.2024.109001","DOIUrl":"10.1016/j.mineng.2024.109001","url":null,"abstract":"<div><div>Magnetic separation is a primary method for processing iron ore and plays a crucial role in both current beneficiation practices and other fields. Extensive research has been conducted on the motion behavior of magnetic particles within magnetic separation equipment. However, force analysis, particularly the calculation of magnetic forces, remains imprecise when dealing with irregularly shaped particles. Accurate prediction of magnetic particle behavior requires precise magnetic force calculations. This study introduces micromagnetic simulations to accurately compute the magnetic forces on irregular magnetic particles. Micromagnetic simulations can determine the precise magnetic moments and magnetic induction intensities within each microelement of the particle. The results of these simulations will be validated using magnetic force microscopy (MFM). The findings indicate that traditional magnetic force calculations deviate from the precise calculations presented in this study. For irregular particles, the computational errors in repulsive and attractive forces are 770% and 576% higher, respectively, compared to spherical particles. This underscores the necessity of considering particle shape in realistic magnetic force calculations. Additionally, both the MFM measurement images and the simulated magnetic force maps exhibit bright and dark regions correlated with particle shape, demonstrating that micromagnetic simulation results can be verified through MFM measurements. This paper proposes an experimentally verifiable method for accurately calculating the magnetic forces on magnetic particles using micromagnetic simulations, which holds significant implications for designing more efficient and precise magnetic separation equipment.</div></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-23","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314241","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-22DOI: 10.1016/j.mineng.2024.108993
<div><div>This paper presents and analyzes the geometallurgy of ores at the Tenke-Fungurume district, a major global source of Cu and Co. Ores at depth are mainly Cu, Cu-Fe, and Cu-Co sulfides (chalcocite group, chalcopyrite, bornite, carrollite, pyrite) hosted in variably silicified dolomite and shale layers. Within roughly 200 m of the surface the sulfides are overprinted by malachite, heterogenite, brochantite, pseudomalachite, and chrysocolla. A mixed zone at intermediate depth contains metals as cobaltoan dolomite, sphaerocobaltite, cuprite, and native copper in addition to sulfides and oxides (<em>sensu lato</em>).</div><div>Geometallurgical properties are highly variable and anisotropic due to the wide unit-to-unit variation in lithology and alteration. Average crusher work index varies from 6.3 to 8.6 kWh/t across units. Average grinding Bond work index ranges from 10.8 kWh/t in the basal argillaceous conglomerate to 15.4 kWh/t in the overlying, variably silicified dolomitic shales. In these dolomitic shales, both work indices correlate strongly with rock uniaxial compressive strength. The same correlation is present in an unlaminated silicified algal dolomite, in which abrasion index also correlates with quartz content. There are no other observed correlations among comminution characteristics, geomechanical properties, and mineralogy.</div><div>Flotation tests show average recoveries of 87% Cu and 61% Co in the sulfide stage, 61% Cu and 40% Co in the oxide stage. Highest recoveries came from the variably silicified dolomitic shales, the lowest from the extremely phyllosilicate-rich basal unit. The main losses in flotation are of oxide minerals and some chalcocite. Apart from the rejection of cobaltoan dolomite, most flotation problems are due to muscovite, biotite, and chlorite, which report to concentrate and may cause sliming and other sources of loss.</div><div>Leach test recoveries for Cu and Co show strong inverse correlations with the fraction of metal hosted in sulfides. Oxide-dominated ores are generally 80–90 % leachable, with Cu recoveries < 20 % for primary sulfides other than chalcocite. In leach testing, Co recoveries exceed Cu recoveries in the same samples, though ultimate Co recovery in practice is lower due to more locking and minor Co hosted in insoluble chlorites. Acid consumption is a function of dolomite concentration and varies from a low of 226 kg/t in a slightly dolomitic shale to a high of 435 kg/t in a laminated dolomite. Acid consumption by other gangue minerals is undetectable over the timescale of testing.</div><div>Ores are divided into one of six types at the mine: sterile; leached; oxide; sulfide; oxide-dominant mixed; and sulfide-dominant mixed ores. The first two of these are below cutoff grade and contain few or no Cu or Co minerals. Higher-grade material is subdivided into the last four categories based on results from quick-leach testing and 1-hour acid-soluble Cu analysis, along with logged miner
{"title":"Geometallurgy of the Tenke-Fungurume sediment-hosted copper-cobalt district, D.R. Congo","authors":"","doi":"10.1016/j.mineng.2024.108993","DOIUrl":"10.1016/j.mineng.2024.108993","url":null,"abstract":"<div><div>This paper presents and analyzes the geometallurgy of ores at the Tenke-Fungurume district, a major global source of Cu and Co. Ores at depth are mainly Cu, Cu-Fe, and Cu-Co sulfides (chalcocite group, chalcopyrite, bornite, carrollite, pyrite) hosted in variably silicified dolomite and shale layers. Within roughly 200 m of the surface the sulfides are overprinted by malachite, heterogenite, brochantite, pseudomalachite, and chrysocolla. A mixed zone at intermediate depth contains metals as cobaltoan dolomite, sphaerocobaltite, cuprite, and native copper in addition to sulfides and oxides (<em>sensu lato</em>).</div><div>Geometallurgical properties are highly variable and anisotropic due to the wide unit-to-unit variation in lithology and alteration. Average crusher work index varies from 6.3 to 8.6 kWh/t across units. Average grinding Bond work index ranges from 10.8 kWh/t in the basal argillaceous conglomerate to 15.4 kWh/t in the overlying, variably silicified dolomitic shales. In these dolomitic shales, both work indices correlate strongly with rock uniaxial compressive strength. The same correlation is present in an unlaminated silicified algal dolomite, in which abrasion index also correlates with quartz content. There are no other observed correlations among comminution characteristics, geomechanical properties, and mineralogy.</div><div>Flotation tests show average recoveries of 87% Cu and 61% Co in the sulfide stage, 61% Cu and 40% Co in the oxide stage. Highest recoveries came from the variably silicified dolomitic shales, the lowest from the extremely phyllosilicate-rich basal unit. The main losses in flotation are of oxide minerals and some chalcocite. Apart from the rejection of cobaltoan dolomite, most flotation problems are due to muscovite, biotite, and chlorite, which report to concentrate and may cause sliming and other sources of loss.</div><div>Leach test recoveries for Cu and Co show strong inverse correlations with the fraction of metal hosted in sulfides. Oxide-dominated ores are generally 80–90 % leachable, with Cu recoveries < 20 % for primary sulfides other than chalcocite. In leach testing, Co recoveries exceed Cu recoveries in the same samples, though ultimate Co recovery in practice is lower due to more locking and minor Co hosted in insoluble chlorites. Acid consumption is a function of dolomite concentration and varies from a low of 226 kg/t in a slightly dolomitic shale to a high of 435 kg/t in a laminated dolomite. Acid consumption by other gangue minerals is undetectable over the timescale of testing.</div><div>Ores are divided into one of six types at the mine: sterile; leached; oxide; sulfide; oxide-dominant mixed; and sulfide-dominant mixed ores. The first two of these are below cutoff grade and contain few or no Cu or Co minerals. Higher-grade material is subdivided into the last four categories based on results from quick-leach testing and 1-hour acid-soluble Cu analysis, along with logged miner","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-22","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142314239","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
Pub Date : 2024-09-21DOI: 10.1016/j.mineng.2024.108997
In order to improve the motion characteristics of particles in vertical roller mills (VRMs), the assumption that different structures of helical guide blades affect the internal flow field of the VRMs was put forward. The distributions of fluid velocity and vorticity in VRMs were analyzed, and the mechanism affecting the motion of particles and the separation performance was studied. The study took the MMLM2550 (supplied by Jiangsu Dahuan Group, China; grinding table diameter of 2550 mm) VRM as the research object, whose structure was improved by adding helical guide blades. The results showed that, with the increase in the width of blade, the air flow trajectory and the distribution of vorticity improved, which was conducive to the transport of particles. However, changing the thickness of the blade had little effect on the internal physical field of the VRM and particle motion characteristics. Therefore, the influence of blade’s thickness on these factors was ignored. With the increase in the number of helical guide blade turns, the direction of the helical guide blade remained consistent with the trajectory of the movement of particles, making it more conducive to the discharge of particles from the VRM. The height of the helical guide blade had a great influence on the flow field and the particle motion characteristics. The smaller the blade height, the greater the reflux in the primary separation zone, and stronger the irregular circulatory motion of particles, thus increasing the movement time and the distance of particles. With the increase in the number of blades, the airflow speed increased in the flow channel, so that the particles moved with the airflow at high speed, while the movement time was effectively shortened. The study provides valuable guidance for the improvement of VRM structures, and serves as a reference for characterizing the motion of particles and enhancing the internal flow field in VRMs.
{"title":"Numerical Study of Flow Field and Particle Motion Characteristics on Raw Coal Vertical Roller Mill Circuits","authors":"","doi":"10.1016/j.mineng.2024.108997","DOIUrl":"10.1016/j.mineng.2024.108997","url":null,"abstract":"<div><p>In order to improve the motion characteristics of particles in vertical roller mills (VRMs), the assumption that different structures of helical guide blades affect the internal flow field of the VRMs was put forward. The distributions of fluid velocity and vorticity in VRMs were analyzed, and the mechanism affecting the motion of particles and the separation performance was studied. The study took the <strong>MMLM2550</strong> (supplied by Jiangsu Dahuan Group, China; grinding table diameter of 2550 mm) VRM as the research object, whose structure was improved by adding helical guide blades. The results showed that, with the increase in the width of blade, the air flow trajectory and the distribution of vorticity improved, which was conducive to the transport of particles. However, changing the thickness of the blade had little effect on the internal physical field of the VRM and particle motion characteristics. Therefore, the influence of blade’s thickness on these factors was ignored. With the increase in the number of helical guide blade turns, the direction of the helical guide blade remained consistent with the trajectory of the movement of particles, making it more conducive to the discharge of particles from the VRM. The height of the helical guide blade had a great influence on the flow field and the particle motion characteristics. The smaller the blade height, the greater the reflux in the primary separation zone, and stronger the irregular circulatory motion of particles, thus increasing the movement time and the distance of particles. With the increase in the number of blades, the airflow speed increased in the flow channel, so that the particles moved with the airflow at high speed, while the movement time was effectively shortened. The study provides valuable guidance for the improvement of VRM structures, and serves as a reference for characterizing the motion of particles and enhancing the internal flow field in VRMs.</p></div>","PeriodicalId":18594,"journal":{"name":"Minerals Engineering","volume":null,"pages":null},"PeriodicalIF":4.9,"publicationDate":"2024-09-21","publicationTypes":"Journal Article","fieldsOfStudy":null,"isOpenAccess":false,"openAccessPdf":"","citationCount":null,"resultStr":null,"platform":"Semanticscholar","paperid":"142270697","PeriodicalName":null,"FirstCategoryId":null,"ListUrlMain":null,"RegionNum":2,"RegionCategory":"工程技术","ArticlePicture":[],"TitleCN":null,"AbstractTextCN":null,"PMCID":"","EPubDate":null,"PubModel":null,"JCR":null,"JCRName":null,"Score":null,"Total":0}
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