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This is an article in the field of metallurgical engineering. At present, the cumbersome product routing of FeV alloy causes high energy consumption and hazardous leaching liquor waste. Therefore, a two-step direct reduction to produce FeV alloy from vanadium slags using photovoltaic silicon wafer cutting waste and aluminum ash as reductant, was investigated to solve the above problems. In the present work, the evolution of viscosity and the melt structure during the two-step reduction process of vanadium slags was investigated. The results showed that at 1 873 K, FeO-V2O3-Cr2O3-MnO-SiO2-TiO2-CaO slag system corresponding to the prereduction process, with the decrease of FeO content and the increase of SiO2 content in the slags, the size of FeCr2O4 spinel decreased gradually. The transition of the main influencing factor of the melt structure from FeO to SiO2 resulted in the increase of structure complexity, the structural complexity of the slag system increased from 0.176 to 2.517, the viscosity of the slag system increased from 0.092 Pa·s to 0.476 Pa·s. FeO-Cr2O3-MnO-V2O3-Al2O3-SiO2-TiO2-CaO slag system corresponding to the final reduction process, with the decrease of FeO, Cr2O3, MnO, V2O3 content and the increase of Al2O3 content in the slag, the size of MnCr2O4 spinel decreased gradually and the slag system comprised from a solid-liquid mixture to liquid, the transition of the main influencing factor from SiO2 to Al2O3 resulted in the transformation of the main structure of the slag system from silicate to aluminosilicate, the structural complexity of the slag system increased from 0.252 to 2.248, and the viscosity of the slag system increased from 0.091 Pa·s to 0.372 Pa·s.

This is an article in the field of metallurgical engineering. Mangane-zinc spinel ferrite has been widely used in various fields due to its advantages of high permeability， high frequency and low loss. In order to make full use of Fe, Zn, Mn and other elements in electric arc furnace dust (EAFD)， Mn-Zn spinel ferrite was prepared by solid phase sintering method with alkali leaching zinc-containing EAFD as raw materials and MnSO4·H2O added. The effects of temperature， mass ratio (RZ/M) and pretreatment on the synthesis and magnetic properties of Mn-Zn ferrite were investigated by XRD, SEM-EDS and VSM. The results show that with the increase of reaction temperature，the speed of inter-grain merging in the Mn-Zn ferrite increases， the magnetic saturation induction and crystal grain size increase， and the coercivity decreases. When RZ/M changes to 1∶1.0，the magnetic saturation induction， coercivity and grain size increase. After pretreatment，the content of SiO2 in the EAFD decreases，and the magnetic saturation strength of the synthesized manganese-zinc ferrite increases. Therefore, after pretreatment with 2 mol/L alkali solution， the performance of manganzn-spinel ferrite synthesized at the condition of mass ratio 1∶1.0 and calcination temperature 1 100 ℃ is the best. Its saturation magnetic induction intensity (Ms) is 17.902 emu/g, and coercivity (Hc) is 3.21 kA/m.

This is an article in the field of metallurgical engineering. Aiming at the limitation of low reactivity of barite and few chemical processing methods, a new soda ash conversion method is proposed. Firstly, the effect of barite reaction particle size, temperature, and sodium carbonate concentration on the production rate of barium carbonate by the hydrochloric acid hydrolysis method is explored. The results show that the sodium carbonate concentration has the greatest impact on the reaction conversion rate. At limited experimental conditions, when sodium carbonate is used when the concentration reaches 28.57%, the highest formation rate of barium carbonate reaches 90.05%. Secondly, in the subsequent reaction, ammonium chloride is selected to re-convert the barite conversion product to prepare ammonium carbonate. The barium carbonate produced by this method has a higher yield and purity. The new process breaks through the traditional misconception that barium sulfate does not react with soda ash. The process is simple and feasible, and ammonia and carbon dioxide can be recycled and reused, making the entire process a large cycle process, in line with the concept of green environmental protection, and recrystallization. The preparation of barium carbonate from stone provides a new production process.

This is an article in the field of metallurgical engineering, which proposed a new idea of mixing molten blast furnace slag with steel slag to achieve self-pulverization of steel slag, and systematically studied the influence mechanism of special components of CaF2 and CeO2 in blast furnace slag of Baotou Iron and Steel Company on the self-pulverization of steel slags. Using analytically pure CaO, SiO2, CaF2 and CeO2 as starting materials, pure C2S specimens were prepared with n ( CaO ) : n ( SiO2 ) = 2 : 1, then different mass fractions of CaF2 and CeO2 were added, respectively. The effect of CaF2 and CeO2 on the self-pulverization behavior of C2S was investigated by calcining the specimen at 1 450 °C for 1 h and cooling with the furnace. The results showed when CaF2 was incorporated into C2S and the content of CaF2 was less than 2.5%, it had little effect on the self-pulverization of C2S. However, when the content of CaF2 was more than 2.5%, the specimen had different degrees of agglomeration, which affected the self-pulverization of C2S. Considering that excessive CaF2 would affect the environment, the content of CaF2 should be controlled below 2.5%. When CeO2 incorporated into C2S and the content was 0.25%, C2S could still be completely self-pulverized, but when the CeO2 content was more than 0.5%, C2S no longer self-pulverized.Therefore, the CeO2 content must be less than 0.25 %, and the self-pulverization effect of C2S was relatively good.

This is an article in the field of metallurgical engineering. Torecover copper and iron from a copper smelting slag flotation tailingwith iron grade of 37.42% and copper grade of 1.86%, according to the characteristics of its ore properties, the bituminous coal was used as a reductant, the process of direct reduction roasting - magnetic separation was applied.The experimental results show that copper smelting slag, bitumite and quicklime (as a co-reducing agent) are mixed in the mass ratio of 100∶25∶20, and at the conditions of roasting temperature of 1 200 ℃, time of 80 min, the roasted ores gringding fineness of -0.045 mm 80%, and magnetic strength of 0.139 T, the copper-bearing iron ore concentrate can be obtained with an iron grade of 90.98%, iron recovery of 90.13%, copper grade of 5.82% and copper recovery of 87.38%, thus realizing the comprehensive recovery of copper and iron from copper smelting slags.

This is an article in the field of metallurgical engineering. A beneficiation and smelting plant processes two kinds of gold ores: primary ore and oxidized ore. The primary ore is treated by conventional flotation process. The gold recovery rate of flotation concentrate is about 85%, which is sold to smelting enterprises outside the province for cyanide leaching. The oxidized ore is treated by stirring cyanidation leaching activated carbon adsorption gold loaded carbon analytical electrodeposition gold slime smelting ingot casting process. The output gold ingot is sold outside, and the total recovery rate of gold dressing and metallurgy is about 90%. At present, with the deepening of mining depth, the ore properties change, the amount of oxidized ore gradually decreases, and the amount of primary ore gradually increases, resulting in excessive ore treatment in flotation system and insufficient ore treatment in leaching system, affecting the economic benefits of enterprises. Therefore, the enterprise considers adding the primary ore that cannot be treated by the flotation system to the existing oxidized ore all slime cyanidation leaching system to increase the gold ingot output and economic benefits of the enterprise. In order to enable the existing full slime cyanidation leaching system to leach the primary ore, based on the field investigation, the grinding fineness, reagent dosage and dissolved oxygen condition tests were carried out for the primary ore, and the alkali leaching pretreatment test and mixed oxidation ore leaching test were carried out in combination with the actual production. Based on these test and research results, specific rectification requirements and measures are put forward for the production site.

This is an article in the field of metallurgical engineering. For the vanadium precipitation waste liquid of a waste catalyst in a zinc smelter in Shaanxi, the vanadium in the vanadium precipitation waste liquid was recovered by solvent extraction method. The effects of N235 volume fraction, extraction ratio, extraction time and extraction stage on the vanadium extraction rate were investigated. The results showed that the extraction system with 30%N235+5%TBP+65% sulfonated kerosene was used for four-stage countercurrent extraction, the extraction ratio of VO/VA= 2∶1, the extraction time was t=5 min, and the extraction rate of vanadium was 95.68%. The volume fraction of Na2CO3 was 5%, the extraction ratio of V'o/V'A =4∶1, the extraction time was t=4 min, and the extraction rate of vanadium was 98.25% after three-stage counter-current extraction.

This is an article in the field of metallurgical engineering. In order to explore the leaching mechanism of zinc oxide from zinc-containing dust by choline chloride and oxalic acid, the interaction of choline chloride and oxalic acid with 1∶1 and 1∶2 eutectic solvent on zinc oxide surface was simulated by means of quantum mechanics, and the simulation results were verified by experiments. The results show that the intermolecular hydrogen bond formed by choline chloride and oxalic acid plays an important role in the stability of eutectic solvent. Ch+ gains electrons, OA and Cl- lose electrons. ZnO(001) surface is the complete cleavage surface of zinc oxide, oxygen and zinc atoms form tetrahedral structure, oxygen atom is easy to gain electrons and zinc atom is easy to lose electrons during the reaction with the outside world. The zinc oxide loses electrons during the interaction between cholinergic chloride-oxalic acid and zinc oxide, and cholinergic chloride-oxalic acid gains electrons. The interaction energy △E=-819.6896Ha between ChCl-2OA and ZnO is small, indicating that zinc oxide is more likely to react with ChCl-2OA. The radial distribution function shows that the chemical adsorption is the main factor in the interaction between choline chloride and oxalic acid and zinc oxide, and the physical adsorption is the secondary factor. The contribution of Cl-Zn formed by chemical adsorption is greater than that of O-Zn. Through the pure mineral test, it can be verified that ChCl-2OA has a better leaching effect on zinc oxide, which verifies the accuracy of molecular simulation of the interaction mechanism between choline chloride and oxalic acid and zinc oxide, and provides a theoretical basis for the leaching of zinc-containing dust with eutectic solvent.

This is an article in the field of metallurgical engineering. Titanium gypsum is an industrial waste residue mainly composed of calcium sulfate dehydrate when the titanium dioxide is produced by the sulfuric acid method. At present, the utilization rate of titanium gypsum is relatively low due to its high impurity content , difficulty in dehydration and poor mechanical properties of the subsequent products. A large amount of titanium gypsum stored in the open air has caused great environmental and economic stress on the titanium dioxide enterprises, which is a serious waste of land resources. In this article, the current situation and research progress of global titanium gypsum in building materials, chemical industry, agriculture and climate change were summarized. The suggestions and prospects for the effective reduction and high-quality resource utilization of titanium gypsum from the aspects of structure adjustment, technological innovation, green development, standard formulation and policy support were provided.

This is an article in the field of metallurgical engineering. The preparation of black copper electrolyte by removing copper from secondary refining slag of high copper and high tin anode furnace by hydrogen peroxide oxidation and acid leaching was studied. The process conditions related to the removal efficiency of the copper and arsenic were researched including the hydrogen peroxide dosage, reaction temperature, sulfuric acid concentration and reaction time. Results show that, under the following optimal conditions, including refining slag of 20.0 g, 2.5 mol/L sulfuric acid of 200 mL, solid-liquid ratio of 1/10, temperature of 65 ℃, dosage of hydrogen peroxide of 10.0 mL, stirring speed of 400 r/min and leaching time of 60 min, the leaching ratios of Cu and As can reach 88.98% and 87.33%, respectively, and the leaching ratios of Ni and Sn were only 7.72% and 1.34%, respectively. The concentration of copper ion in the leaching solution was 48.48 g/L, which can be used as the supplement of black copper electrolyte.

This is an article in the field of metallurgical engineering. In order to strengthen the effective separation and recovery of zinc and iron resources in zinc containing electric arc furnace dust(EAFD), and reduce the consumption of carbon reducing agent, the idea and method of calcified carbothermal reduction roasting with high basicity burden prepared from EAFD was proposed. Thermodynamic calculation and test study were combined, the main phase transformation during the calcified carbothermal reduction roasting process of EAFD was analyzed to explore calcified carbothermal reduction reaction behavior and path. It is shown that the reduction of ZnFe2O4 to Fe0.85-xZnxO inhibits the reduction and volatilization of zinc when the molar ratio of carbon to oxygen is less than 0.6 and the temperature is lower than 1000 ℃. While the calcification of ZnFe2O4 and Fe0.85-xZnxO to Ca2Fe2O5 by adding CaO can promote the reduction of zinc and Ca2Fe2O5 will be further reduced. When the temperature is lower than 1100 ℃ and nc/no< 1.0, the reaction path of calcified carbothermal reduction roasting is ZnFe2O4 + CaO → Ca2Fe2O5 + ZnO → Ca2Fe2O5 + Zn (g) and Fe0.85-xZnxO + CaO → Ca2Fe2O5 + ZnO + FeO → Ca2Fe2O5 + Fe + Zn (g). Both of these reactions can promote the release of zinc. The results show that CaO can promote zinc volatilization in the range of nc/no 0.4～1.2 and calcination temperature 1000～1100 ℃. The dezincification rate of calcified carbon thermal reduction roasting at nc/no = 1.0 is close to that at nc/no = 1.2, both of which are about 90%, so the consumption of carbon reducing agent can be reduced and energy consumption can be saved.

This is an article in the field of metallurgical engineering. In order to better realize the leaching of zinc oxide from zinc-containing dust, Materials Studio software based on density functional theory was used to simulate and optimize the crystal structure of zinc oxide and the structure of three kinds of choline deep eutectic solvents, and the mutual adsorption model of the two was calculated. The calculated results show that the ZnO(001) plane is a complete cleavage plane, and the highest occupied state near the Fermi level shifts to the left, and the peak of the highest occupies state increases, and the p orbital of O and the d orbital of Zn are more active, which are the active sites of the ZnO(001) plane. Optimization of three choline hyper eutectic solvent shows that choline chloride forms multiple intermolecular hydrogen bonds with three different hydrogen bond donors centered on chlorine atom. The adsorption model of deep eutectic solvents and zinc oxide was calculated using Forcite module. The results shows that the interaction strength of zinc oxide with three choline deep eutectic solvent is Chcl-MA >Chcl-Urea>Chcl-Eg. The radial distribution function shows that malonic acid makes it easier for Cl in choline chloride to chemisorb with Zn. In the three kinds of hydrogen bond donors, the activity of oxygen atomic functional groups forming chemical bonds with Zn is C=O, N-O and C-O, indicating that C=O in malonic acid makes the combination of Chcl-MA and ZnO more stable. Results show that the leaching temperature of Chcl-MA is 70 ℃ and the liquid-solid ratio is 10∶1. Under the condition of leaching time of 1 h for 60 min, single mineral zinc oxide can be almost completely leached, and the leaching effect is far greater than that of Chcl-Urea and Chcl-Eg, thus proving the accuracy of molecular simulation and providing theoretical guidance for the leaching of zinc-containing dust in choline deep eutectic solvent.

This is an article in the field of metallurgical engineering. The vanadium ore with vanadium containing sericite and illite in Shaanxi province was taken as the research object, low-temperature sulfuric acid ripening-water leaching process was used to extract vanadium. The effects of ripening temperature, ripening time, sulfuric acid dosage, sodium chloride dosage, leaching temperature, leaching time and liquid-solid ratio on vanadium leaching rate were investigated. The results show when 25% sulfuric acid and 0.8% sodium chloride are added to the stone coal vanadium ore, the sample is aged at 130 ℃ for 8 h, and the sample is leached at 55 ℃ for 2 h under the conditions of liquid-solid ratio of 1.5∶1, the vanadium leaching rate can reach more than 89%. This indicates that it is feasible to adopt low temperature sulfuric acid ripening and water leaching process to extract stone coal type vanadium ore.

This is an article in the field of metallurgical engineering. High purity hafnium diboride (HfB2) powders were synthesized from hafnium oxide (HfO2) and boron carbide (B4C) powders via sodium chloride as the molten salt through boron/carbon thermal reduction. Investigations were carried out to determine the effect of synthesis parameters (reaction temperature, holding time) and boron content on the phase composition and microstructure of the synthesized hafnium diboride powders. The results demonstrated when sodium chloride was used as the molten salt medium, the conversion of HfO2 to HfB2 started at 1300 ℃, which was significantly lower than the temperature required for conventional hafnium diboride. The purity of the HfB2 powder was the highest when the boron excess was 20%, with the reaction temperature and holding time being 1400 ℃ and 2 h, respectively. HfB2 crystalline peaks were prominent in X-ray diffraction (XRD) and scanning electron microscopy (SEM) confirmed HfB2 in a densely agglomerated structure.

This is an article in the field of metallurgical engineering. China is short of bauxite resources, which makes the high alumina fly ash a kind of potential bauxite resources. The extraction of alumina oxide from fly ash will be used as a reserve technology, which will make the in-depth development and research constantly pushed forward. The pre-desilication soda lime sintering method of alumina extraction from fly ash has been introduced as the only large-scale and industrialized manufacture technology. It can provide reference information for technology development and the industrialization, and make the national bauxite strategic reserve system safe.

This is an article in the field of metallurgical engineering. Large production and storage capacity of phosphogypsum has become an important factor restricting the green and healthy development of phosphorus chemical industry. Removing impurity and whitening of phosphogypsum is a necessary way for its high consumption. Analysis results showed that there was still a small amount of black organic matter in the phosphogypsum after removing impurity and desilication by flotation, which affected its whiteness. In the paper, calcination and acid leaching were adopted to treat the phosphogysum. The whiteness of whitened phosphogysum significantly increased from 51.5% to 92.7% at the conditions of calcination temperature 600 ℃, calcination time 70 min, acid leaching time 2 h, sulfuric acid concentration 1.5 mol/L, acid leaching liquid to solid ratio 5∶1 and acid leaching temperature 90 ℃. The whitened phosphogypsum was mainly CaSO4, and contained a small amount of CaSO4·2H2O and CaSO4·0.5H2O. The particles of the whitened phosphogypsum were irregular in shape of 15～20 μm, and the surface was smooth. The study provided a guidance for the comprehensive utilization of phosphogypsum with high consumption.

This is an article in the field of metallurgical engineering. To provide a theoretical basis for FeO composition control and quality optimization of sinter, the metallurgical properties of sinter ore used in 5800 m3 blast furnace of Shagang, such as drum strength, low-temperature reduction pulverization, medium-temperature reducibility were investigated. The effect of FeO content on the performance of sinter and the factors affecting FeO content were explored through statistical data analysis. The results show that based on Shagang's current raw material quality, ore blending structure and process operating conditions，the increase of FeO content can improve the drum index and RDI+3.15 of sinter. When RI is more than 85%, the FeO content becomes the limiting factor of RI, and the appropriate value of FeO content is about 9.5%. The increase of R, SiO2 content, fine powder and fuel ratio will cause the increase of FeO content, MgO content and negative pressure will cause the decrease of FeO content. Data analysis shows if R is 1.9, the contents of SiO2 and MgO are about 5.2% and 1.8%, respectively. The ratio of fine powder is less than 11%, the ratio of fuel is between 4.5 and 5%, and the negative pressure is below 16 KPa, the average FeO content of sinter is about 9.5%, the corresponding drum index is about 80%. The average values of RDI+3.15 and RI are above 75% and 80%, respectively.

This is an article in the field of metallurgical engineering. With the increasing attention of iron and steel enterprises to pollutant emission, non blast furnace ironmaking process has gradually become a hot issue. At present, the industrialized non blast furnace ironmaking technology in China is mainly Corex process of Baogang and HIsmelt process of Shandong Molong. It is also a hot process of non blast furnace ironmaking technology in China. This article expounds the processes of Corex process and HIsmelt process, compares the technical indexes, advantages and disadvantages of the two processes, discusses the research status of the two processes in China, and looks forward to their development direction in combination with the characteristics of each process.

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Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

This is an article in the field of metallurgical engineering. In this article, TiH2 powder was selected as the experimental raw material to study the sintering behavior of TiH2 powder under different pressing pressure and holding time in cold isostatic pressing, to reveal the effects of pressing pressure and holding time on the shrinkage as well as the relative density of TiH2 sintered specimens, in order to improve the density of the billet and the performance of the products. The results show that the shrinkage of the sintered sample decreases with the increase of the pressing pressure, and when the pressing pressure reaches 400 MPa, the pressing pressure has little effect on the relative density of the sintered sample. For the consideration of the sample preparation efficiency and the pressing pressure limit of the cold isostatic press, the pressing pressure should not exceed 400 Mpa. The shrinkage of sintered samples decreases with the increase of holding time, and when the holding time reaches 20 min, the relative density of sintered samples changes little with the extension of holding time. For the consideration of sample preparation efficiency, the suitable holding time is 15～20 min.

This is an article in the field of metallurgical engineering. There are high contents of gallium and scandium in bauxite, which has great recovery value. The study on the leaching behavior of gallium and scandium in bauxite leaching process is of great significance for the extraction and utilization of gallium and scandium. The leaching behavior of gallium and scandium from gibbsite bauxite was studied. The results showed that gallium was easy to be leached into the solution. Increasing the temperature, prolonging the reaction time and increasing the mass concentration of sodium hydroxide were conducive to the leaching of gallium. The leaching rate of gallium could reach over 73%, and adding calcium oxide reduced the leaching rate of gallium. The gallium content in the solution was accumulated during the cyclic digestion process. After 3 cycles, the gallium concentration in the decomposition mother liquor reached 24.2 mg/L. Both red mud and decomposition mother liquor had high gallium content and had recycling value. Scandium was not easy to be leached. After digestion, 97.4% of scandium was enriched in red mud. The object of scandium recovery and extraction was mainly red mud.

This is an article in the field of metallurgical engineering. For recycling theferrophosphorus slag which a by-products of phosphorus chemical industry,the thermodynamic calculations and experimentally investigations of the removal of SiO2 by added CaO were carried out. The results of thermodynamic calculations of the reaction between SiO2 and CaO show that: the Gibbs free energy values of generating CaO·SiO2, 2CaO·SiO2, 3CaO·SiO2 and 3CaO·2SiO2 are less than zero, when the temperature in 900～1700 K range and molar ratio of(SiO2∶CaO)= 1∶1～3. It is feasible to generate calcium silicate by adding CaO, which can be used as a method of desilicification for ferrophosphorus slag. The experimental results show that:there are two layers of the products,i.e. Ca2SiO4, Ca3SiO5in the upper layer and Fe2P in the lower layer when the pressure between 10～40 Pa and the temperature 1573 K and molar ratio of SiO2∶CaO = 1∶1. The final products were testified to be ω（Si）

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Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

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Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

This is an article in the field of metallurgical engineering. In view of the problems of low compressive strength and high reduction temperature in the gas-base reduction of a high-phosphorus iron ore, a new process of oxidation roasting, gas-based reduction followed by magnetic separation was proposed. The effect of oxidation temperature and the types and dosages of dephosphorization on the compressive strength of the oxidized pellets were investigated, and the conditions of oxidation roasting that met the strength requirements of the shaft furnace were found. On this basis, the effects of reducing temperature, total reducing gas flow rate, reducing gas composition and reducing time on iron recovery and dephosphorization were studied. The results showed that under the conditions of 10% Na2CO3 dosage, oxidation temperature 1200 ℃, oxidation time 60 min, reduction temperature 950 ℃, flow rates of H2 and CO were 3.75 L/min and 1.25 L/min, respectively, and reduction time 180 min, Iron grade, iron recovery and phosphorus content were 91.15%, 93.07% and 0.14%, respectively. The SEM results show that the phosphorus in the powdered reduced iron exists in the form of mechanical inclusions, and the phosphorus is removed by grinding-magnetic separation.

This is an article in the field of metallurgical engineering. A copper cobalt ore Congo is an oxide ore with Cu 3.43% and Co 0.42% respectively. In this paper, the five stage cycle leaching process of leaching solution was used for the copper and cobalt leaching. The optimum conditions were as follows: the amount of sulfuric acid was 7.4% of the ore mass, the amount of sodium sulfite was 1.68 times of the theoretical amount, the grinding particle size of -74 μm accounts for 75%, leaching temperature was 45 ℃, the single-stage leaching time was 4 hours, and the liquid-solid volume mass ratio was between 2/1 and 3/1, the results showed that the leaching rate of copper and cobalt were 96.85% and 95.67% respectively. According to this leaching process, the amount of sulfuric acid was reduced by 6 kg/t compared with the primary leaching, and the total solution amount in the leaching process was reduced by about 1/4, which made the amount of subsequent cobalt precipitation and copper extraction solution was reduced.

This is an essay in the field of metallurgical engineering. Calcium chloride was used as chlorination agent to conduct the heavy metals removal experiments for coal fly ash, investigating the effects of calcination temperature, calcination time and amount of CaCl2 on the removal rate of heavy metals. The results of calcination experiments and thermodynamic calculation demonstrate the mechanism of heavy metals removal in fly ash by using chlorinating volatilization method; further, the kinetic model of chlorinating process was discussed. Moreover, the coal fly ash, as raw material, was used to produce soil conditioner which was studied to discover its fertilizing effect by pot experiments. Studies found that the remaining contents of Pb and Cd in mixture were 10.22 g/t and 0.54 g/t, respectively, which meant 89.26% of Pb and 76.88% of Cd were decreased after addition of CaCl2 at high calcination temperature for 40 min. The mineral of mullite, quartz, alumina and disodium oxide in fly ash contributed to the decomposition of CaCl2. PbO and CdO mainly reacted with HCl formed by the decomposition of CaCl2 to form volatile chloride, meaning the oxides do not directly react with CaCl2. thermodynamic models fitting implied that the chlorination of Pb and Cd were controlled by the chemical reaction, and the apparent activation energy of reactions was 84.54 and 44.96 kJ/mol, respectively. Further, the pot experiments showed that the synthesized soil conditioner can improve the condition of the soil, promote nutrient absorption of chives, and improve the quality of chives. The effective calcium content of soil conditional product was increased after the reduction of heavy metals by adding CaCl2. This research provides a method for comprehensive utilization of fly ash.

This is an essay in the field of metallurgical engineering. Taking a lithium mica ore in Jiangxi as raw material, the roasting-leaching, mixed acid ripening, direct acid leaching and alkali pressure cooking process were explored. Finally, the sulfate roasting-water leaching method was used to extract lithium from lithium mica ore. At the same time, the effects of roasting temperature, roasting time, additive type, additive amount, liquid-solid ratio and leaching temperature on lithium leaching rate were studied. The results show that roasting temperature has a great influence on lithium leaching rate, and the leaching effect of lithium is better in the appropriate roasting temperature range. 40% potassium sulfate, 20% sodium sulfate and 20% calcium oxide were added to the lithium mica ore, calcinated at 900 ℃ for 1 h, and the calcine was leached at room temperature for 1 h according to the liquid-solid ratio of 1∶1. The lithium leaching rate reached 94.87%. This shows that the effect of using sulfate as an additive to roast and extract lithium is better. Through the study of the roasting mechanism, it can be seen that the mineral structure is reconstructed after the addition of sulfate after high temperature roasting. The sodium and potassium ions in the ore are replaced with lithium ions in the lithium mica, so that they are separated from the insoluble aluminum silicate minerals to generate soluble lithium sulfate, which is then immersed in water and then into the solution.

This is an essay in the field of metallurgical engineering. Bayer red mud and sintered red mud were added to Ningxia coal in a certain proportion to study the influence of the two different red mud on the ash melting temperature of Ningxia coal. Factsage software was used to study the ash melting temperature change mechanism. The experimental results show that Ningxia coal ash contains a large amount of mullite, which causes the ash melting point of coal ash to be higher, and adding red mud to it can effectively reduce the melting point of coal ash. With the increase in the amount of Bayer red mud added, the composition of mullite in coal ash gradually decreases, while anorthite and albite gradually increase, reducing the melting temperature of coal ash. With the increase of sintered red mud addition, the content of mullite in coal ash gradually decreases, and the content of feldspar gradually increases. The main reason that red mud reduces the melting temperature of coal ash in Ningxia is that the alkaline oxides in the red mud react with the acidic oxides in the coal ash to form low-melting minerals and the interaction between these minerals generates low temperature. Communion.

Copyright of Multipurpose Utilization of Mineral Resources / Kuangchan Zonghe Liyong is the property of Multipurpose Utilization of Mineral Resources Editorial Office and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

This is an essay in the field of metallurgical engineering. In the hydro-zinc smelting system, the large amount of chloride ions seriously restricts the normal production. In this essay, a new type of X agent is used for the dechlorination test. Neutral leaching solution was used as the research object to explore the influence of X agent addition, neutral leaching solution temperature, pH value, reaction time, and impurities F− and Mg2+ on the chlorine removal rate. The test results show that when the dosage of X agent is 8 times, the reaction time is 60 min, the reaction temperature is 60 ℃, and the pH value is less than 5.5, X agent has the relatively highest rate of chlorine removal, about 85%; among them, F− no obvious effect on the chlorine removal rate; when the Mg2+ concentration exceeds 16 g/L, the chlorine removal rate of the X agent drops slightly.

This is an essay in the field of metallurgical engineering. Aiming at the problem that the rare earth enrichment after magnetic separation of iron ore tailings in Yen Bai Province, Vietnam cannot be directly utilized, the sulfuric acid roasting process and extraction and enrichment transformation technology are used to carry out experimental development. The process parameters of sulfuric acid roasting, water leaching, impurity removal, transformation and other processes were systematically studied and observed how work on the effect of the experiment and the most optimum technological conditions were determined. The experiment showed the acid and ore ratio (volume: mass) is 0.7, the roasting temperature is 250 degrees Celsius, the roasting time is 2 hours, the leaching water and ore ratio is 4, and the roasting and leaching process was carried out under the above conditions. Impurity removal was carried out with a concentration of 0.1 g/mL and a pH of 4.2 as the endpoint. The grade 5 counter-current extraction was carried out under the O/A of 1.5~2, the 5 counter-current back extraction was carried out under the O/A rate of 10:1, the overall recovery rate reached 90.1%-92.88%, the rare earth concentrations is 1.5 mol/L, the content of iron and aluminum is lower than 100 μg/mL, and met the requirement of the industrial applications, and realized the utilization and development of the accompanied rare earth resources of the iron ore tailings in Yen Bai Province, Vietnam.

This is an essay in the field of metallurgical engineering. This work aims to investigate and optimize the effects of the leaching parameters on the selective leaching of cadmium from bottom blowing melting furnace dust. The response surface method was applied for the optimization of selective leaching parameters of cadmium from the dust. The leaching recovery of cadmium (Y1) was taken as the response variable, where the leaching time (A, h), concentration of sulphuric acid (B, mol/L), and leaching temperature (C, ℃) were considered as the independent variables (factors). The research shows that the parameters of the response surface quadratic regression model are significant. Among the model parameters, A, B, C, A2 and C2 are important model items. The order of influence of temperature, time and sulfuric acid concentration on Cd leaching rate is C>A>B. The F value of AB was 2.16, the F value of AC was 0.051, and the F value of BC was 0.2. The order of influence on Cd leaching rate during the interaction process was AB>BC>AC. When the temperature is 75 ℃, the time is 1.62 h, and the sulfuric acid concentration is 0.26 mol/L, the predicted cadmium leaching rate is 99.2%. Through three experiments, the average value of the cadmium leaching rate is 98.18%, and the deviation from the predicted value is 1.02, Higher is the optimum process condition for leaching of cadmium.

This is an essay in the field of metallurgical engineering. The effects of crushing particle size and leaching conditions on the leaching of lithium cobalt from spend lithium-ion batteries in the citric acid environment were studied, which provided some reference for the recovery of different types of mixed spend lithium-ion batteries. The results showed that Mn， Al， Ni, Co and Li accounted for a large proportion of the contents in mixed different types of spend lithium-ion batteries. Due to the larger recovery value of Co and Li and simplify the recovery process, only Co and Li be recovered. The crushing particle size has little influence on the leaching rate of lithium cobalt in the range of -5 mm, while the leaching rate of lithium cobalt in the larger crushing particle size is not high, which may be lithium and cobalt wrapped in the middle of particles and cannot contact with the leaching solution, thus reducing the leaching effect. For mixed different types of spend lithium-ion batteries，under the conditions of -5 mm particle size, citric acid concentration 1.0 mol/L, leaching temperature 65 ℃, solid-liquid ratio 1 g/100 mL, H2O2 concentration 3%, leaching time 55 min, stirring speed 30 r/min, the lithium leaching rate of 97.86% and cobalt leaching rate of 98.01% are obtained.