Reaction behaviors of Pb and Zn sulfates during reduction roasting of Zn leaching residue and flotation of artificial sulfide minerals

To evaluate the feasibility of recovering Pb and Zn sulfides and Ag-containing minerals from Zn leaching residue by the process of reduction roasting followed by flotation, the reaction behaviors of Pb and Zn sulfates during this process were investigated. Chemical analysis showed that the transformation ratios of PbSO_4 and ZnSO_4 could reach 65.51% and 52.12%, respectively, after reduction roasting, and the introduction of a sulfidation agent could improve the transformation ratios of these sulfates. scanning electron microscopy–energy dispersive spectroscopy(SEM–EDS) revealed that temperature obviously affects the particle size, crystal growth, and morphology of the artificial Pb and Zn sulfide minerals. Particle size analysis demonstrated that the particle size of the materials increases after roasting. Flotation tests revealed that a flotation concentrate composed of 12.01 wt% Pb, 27.78 wt% Zn, and 6.975 × 10~(-2) wt% Ag with recoveries of 60.54%, 29.24%, and57.64%, respectively, could be obtained after roasting.     

利用自动解离分析方法研究硫化铜尾矿的矿物学特征:谦比希铜矿尾矿的实例研究

As ore grades constantly decline, more copper tailings, which still contain a considerable amount of unrecovered copper, are expected to be produced as a byproduct of froth flotation. This research reveals the occurrence mechanism of copper minerals in typical copper sulfide tailings using quantitative mineral liberation analysis (MLA) integrated with scanning electron microscopy–energy dispersive spectroscopy (SEM–EDS). A comprehensive mineralogical characterization was carried out, and the results showed that almost all copper minerals were highly disseminated within coarse gangue particles, except for 9.2wt% chalcopyrite that occurred in the 160–180 μm size fraction. The predominant copper-bearing mineral was chalcopyrite, which was closely intergrown with orthoclase and muscovite rather than quartz. The flotation tailings sample still contained 3.28wt% liberated chalcopyrite and 3.13wt% liberated bornite because of their extremely fine granularity. The SEM–EDS analysis further demonstrated that copper minerals mainly occurred as fine dispersed and fully enclosed structures in gangue minerals. The information obtained from this research could offer useful references for recovering residual copper from flotation tailings.     

Process mineralogy as a key factor affecting the flotation kinetics of copper sulfide minerals

The aim of this study is to apply process mineralogy as a practical tool for further understanding and predicting the flotation kinetics of the copper sulfide minerals. The minerals' composition and association, grain distribution, and liberation within the ore samples were analyzed in the feed, concentrate, and the tailings of the flotation processes with two pulp densities of 25 wt% and 30 wt%. The major copper-bearing minerals identified by microscopic analysis of the concentrate samples included chalcopyrite(56.2 wt%), chalcocite(29.1 wt%),covellite(6.4 wt%), and bornite(4.7 wt%). Pyrite was the main sulfide gangue mineral(3.6 wt%) in the concentrates. A 95% degree of liberation with d_(80) 80 μm was obtained for chalcopyrite as the main copper mineral in the ore sample. The recovery rate and the grade in the concentrates were enhanced with increasing chalcopyrite particle size. Chalcopyrite particles with a d_(80) of approximately 100 μm were recovered at the early stages of the flotation process. The kinetic studies showed that the kinetic second-order rectangular distribution model perfectly fit the flotation test data. Characterization of the kinetic parameters indicated that the optimum granulation distribution range for achieving a maximum flotation rate for chalcopyrite particles was between the sizes 50 and 55 μm.     

Mineralogical characterization of copper sulfide tailings using automated mineral liberation analysis: A case study of the Chambishi Copper Mine tailings

As ore grades constantly decline, more copper tailings, which still contain a considerable amount of unrecovered copper, are expected to be produced as a byproduct of froth flotation. This research reveals the occurrence mechanism of copper minerals in typical copper sulfide tailings using quantitative mineral liberation analysis(MLA) integrated with scanning electron microscopy–energy dispersive spectroscopy(SEM–EDS). A comprehensive mineralogical characterization was carried out, and the results showed that almost all copper minerals were highly disseminated within coarse gangue particles, except for 9.2 wt% chalcopyrite that occurred in the 160–180 μm size fraction. The predominant copper-bearing mineral was chalcopyrite, which was closely intergrown with orthoclase and muscovite rather than quartz. The flotation tailings sample still contained 3.28 wt% liberated chalcopyrite and 3.13 wt% liberated bornite because of their extremely fine granularity.The SEM–EDS analysis further demonstrated that copper minerals mainly occurred as fine dispersed and fully enclosed structures in gangue minerals. The information obtained from this research could offer useful references for recovering residual copper from flotation tailings.     

A novel process for Fe recovery and Zn,Pb removal from a low-grade pyrite cinder with high Zn and Pb contents

Comprehensive utilization of pyrite cinders is increasingly important because of their huge annual outputs and potential valuable metals recovery to cope with the gradual depletion of high-grade mineral resources. In this work, a new process, i.e., a high-temperature chlorination–magnetizing roasting–magnetic separation process, was proposed for recovering Fe and removing Zn, Pb from a low-grade pyrite cinder containing 49.90 wt% Fe, 1.23 wt% Zn, and 0.29 wt% Pb. Various parameters, including the chlorinating conditions(dosage of Ca Cl2, temperature, and time) and the magnetization roasting conditions(amount of coal, temperature, and time) were investigated. The results indicate that the proposed process is effective for Fe recovery and Zn, Pb removal from the pyrite cinder. Through this process, 97.06% Zn, 96.82% Pb, and approximately 90% S can be removed, and 89.74% Fe is recovered as magnetite into the final product under optimal conditions. A purified magnetite concentrate containing 63.07 wt% Fe, 0.16 wt% P, 0.26 wt% S, and trace amounts of nonferrous metals(0.005 wt% Cu, 0.013 wt% Pb, and 0.051 wt% Zn) was obtained. The concentrate can be potentially used as a high-quality feed material for producing oxidized pellets by blending with other high-grade iron ore concentrates.     

Use of centrifugal-gravity concentration for rejection of talc and recovery improvement in base-metal flotation

The possibility of using a centrifugal-gravity concentrator to reject Mg-bearing minerals and minimize metal losses in the flotation of base metals was evaluated. Sample characterization, batch scoping tests, pilot-scale tests, and regrind-flotation tests were conducted on a Ni flotation tailings stream. Batch tests revealed that the Mg grade decreased dramatically in the concentrate products. Pilot-scale testing of a continuous centrifugal concentrator (Knelson CVD6) on the flotation tailings revealed that a concentrate with a low mass yield, low Mg content, and high Ni upgrade ratio could be achieved. Under optimum conditions, a concentrate at 6.7% mass yield was obtained with 0.85% Ni grade at 12.9% Ni recovery and with a low Mg distribution (1.7%). Size partition curves demonstrated that the CVD also operated as a size classifier, enhancing the rejection of talc fines. Overall, the CVD was capable of rejecting Mg-bearing minerals. Moreover, an opportunity exists for the novel use of centrifugal-gravity concentration for scavenging flotation tailings and/or after comminution to minimize amount of Mg-bearing minerals reporting to flotation.     

Preparation of ferronickel from nickel laterite via coal-based reduction followed by magnetic separation

The sticking phenomenon between molten slag and refractory is one of the crucial problems when preparing ferronickel from laterite ore using rotary hearth furnace or rotary kiln processes. This study aims to ameliorate sticking problems by using silicon dioxide (SiO₂) to adjust the melting degree of the briquette during reduction roasting. Thermodynamic analysis indicates that the melting temperature of the slag gradually increases with an increase in the SiO₂ proportion (SiO₂/(SiO₂ + Al₂O₃ + MgO) mass ratio). Experimental validations also prove that the briquette retains its original shape when the SiO₂ proportion is greater than 75wt%, and sticking problems are avoided during reduction. A ferronickel product with 8.33wt% Ni and 84.71wt% Fe was prepared via reductive roasting at 1500℃ for 90 min with a SiO₂ proportion of 75wt% and a C/O molar ratio of 1.0 followed by dry magnetic separation; the corresponding recoveries of Ni and Fe reached 75.70% and 77.97%, respectively. The microstructure and phase transformation of reduced briquette reveals that the aggregation and growth of ferronickel particles were not significantly affected after adding SiO₂ to the reduction process.     

Formation of calcium titanate in the carbothermic reduction of vanadium titanomagnetite concentrate by adding CaCO_3

The formation of calcium titanate in the carbothermic reduction of vanadium titanomagnetite concentrate(VTC) by adding CaCO_3 was investigated. Thermodynamic analysis was employed to show the feasibility of calcium titanate formation by the reaction of ilmenite and Ca CO_3 in a reductive atmosphere, where ilmenite is more easily reduced by CO or carbon in the presence of CaCO_3. The effects of CaCO_3 dosage and reduction temperature on the phase transformation and metallization degree were also investigated in an actual roasting test. Appropriate increase of CaCO_3 dosages and reduction temperatures were found to be conducive to the formation of calcium titanate, and the optimum conditions were a CaCO_3 dosage of 18 wt% and a reduction temperature of 1400°C. Additionally, scanning electron microscopy–energy dispersive spectrometry(SEM–EDS) analysis shows that calcium titanate produced via the carbothermic reduction of VTC by CaCO_3 addition was of higher purity with particle size approximately 50 μm. Hence, the separation of calcium titanate and metallic iron will be the focus in the future study.     

添加CaCO3后碳热还原钒钛磁铁矿精矿过程中钛酸钙的形成

The formation of calcium titanate in the carbothermic reduction of vanadium titanomagnetite concentrate (VTC) by adding CaCO₃ was investigated. Thermodynamic analysis was employed to show the feasibility of calcium titanate formation by the reaction of ilmenite and CaCO₃ in a reductive atmosphere, where ilmenite is more easily reduced by CO or carbon in the presence of CaCO₃. The effects of CaCO₃ dosage and reduction temperature on the phase transformation and metallization degree were also investigated in an actual roasting test. Appropriate increase of CaCO₃ dosages and reduction temperatures were found to be conducive to the formation of calcium titanate, and the optimum conditions were a CaCO₃ dosage of 18wt% and a reduction temperature of 1400°C. Additionally, scanning electron microscopy–energy dispersive spectrometry (SEM–EDS) analysis shows that calcium titanate produced via the carbothermic reduction of VTC by CaCO₃ addition was of higher purity with particle size approximately 50 μm. Hence, the separation of calcium titanate and metallic iron will be the focus in the future study.     

Improved hemimorphite flotation using xanthate as a collector with S(Ⅱ) and Pb(Ⅱ) activation

The flotation of hemimorphite using the S(Ⅱ)–Pb(Ⅱ)–xanthate process,which includes sulfidization with sodium sulfide,activation by lead cations,and subsequent flotation with xanthate,was investigated.The flotation results indicated that hemimorphite floats when the S(Ⅱ)–Pb(Ⅱ)–xanthate process is used; a maximum recovery of approximately 90% was obtained.Zeta-potential,contact-angle,scanning electron microscopy–energy-dispersive spectrometry(SEM–EDS),and diffuse-reflectance infrared Fourier transform spectroscopy(DRIFTS) measurements were used to characterize the activation products on the hemimorphite surface and their subsequent interaction with sodium butyl xanthate(SBX).The results showed that a Zn S coating formed on the hemimorphite surface after the sample was conditioned in an Na2 S solution.However,the formation of a Zn S coating on the hemimorphite surface did not improve hemimorphite flotation.With the subsequent addition of lead cations,Pb S species formed on the mineral surface.The formation of the Pb S species on the surface of hemimorphite significantly increased the adsorption capacity of SBX,forming lead xanthate(referred to as chemical adsorption) and leading to a substantial improvement in hemimorphite flotation.Our results indicate that the addition of lead cations is a critical step in the successful flotation of hemimorphite using the sulfidization–lead ion activation–xanthate process.     

Two-stage reduction for the preparation of ferronickel alloy from nickel laterite ore with low Co and high MgO contents

The preparation of ferronickel alloy from the nickel laterite ore with low Co and high MgO contents was studied by using a pre-reduction-smelting method. The effects of reduction time, calcination temperature, quantity of reductant and calcium oxide (CaO), and pellet diameter on the reduction ratio of Fe and on the pellet strength were investigated. The results show that, for a roasting temperature >800℃, a roasting time >30 min, 1.5wt% added anthracite coal, 5wt% added CaO, and a pellet size of~10 mm, the reduction ratio of Fe exceeds 70% and the compressive strength of the pellets exceeds 10 kg per pellet. Reduction smelting experiments were performed by varying the smelting time, temperature, quantity of reductant and CaO, and reduction ratio of Fe in the pellets. Optimal conditions for the reduction smelting process are as follows:smelting time, 30-45 min; smelting temperature, 1550℃; quantity of reductant, 4wt%-5wt%; and quantity of CaO, 5wt%; leading to an Fe reduction ratio of 75% in the pellets. In addition, the mineral composition of the raw ore and that during the reduction process were investigated by process mineralogy.     

Effects and mechanisms of fluorite on the co-reduction of blast furnace dust and seaside titanomagnetite

The co-reduction roasting and grinding magne -tic separation of seaside titanomagnetite and blast furnace dust was investigated with and without fluorite addition at a reduction roasting temperature of 1250℃ for 60 min, a grinding fineness of -43 μm accounting for 69.02wt% of the total, and a low-intensity magnetic field strength of 151 kA/m. The mineral composition, microstructure, and state of the roasted products were analyzed, and the concentrations of CO and CO₂ were analyzed in the co-reduction roasting. Better results were achieved with a small fluorite dosage (≤ 4wt%) in the process of co-reduction. In addition, F- was found to reduce the melting point and viscosity of the slag phase because of the high content of aluminate and silicate minerals in the blast furnace dust. The low moisture content of the blast furnace dust and calcic minerals inhibited the hydrolysis of CaF₂ and the loss of F-. Compared with the blast furnace dust from Chengdeng, the blast furnace dusts from Jiugang and Jinxin inhibited the diffusion of F- when used as reducing agents, leading to weaker effects of fluorite.     

A novel process for Fe recovery and Zn,Pb removal from a low-grade pyrite cinder with high Zn and Pb contents

Comprehensive utilization of pyrite cinders is increasingly important because of their huge annual outputs and potential valuable metals recovery to cope with the gradual depletion of high-grade mineral resources. In this work, a new process, i.e., a high-temperature chlorination-magnetizing roasting-magnetic separation process, was proposed for recovering Fe and removing Zn, Pb from a low-grade pyrite cinder containing 49.90wt% Fe, 1.23wt% Zn, and 0.29wt% Pb. Various parameters, including the chlorinating conditions (dosage of CaCl₂, temperature, and time) and the magnetization roasting conditions (amount of coal, temperature, and time) were investigated. The results indicate that the proposed process is effective for Fe recovery and Zn, Pb removal from the pyrite cinder. Through this process, 97.06% Zn, 96.82% Pb, and approximately 90% S can be removed, and 89.74% Fe is recovered as magnetite into the final product under optimal conditions. A purified magnetite concentrate containing 63.07wt% Fe, 0.16wt% P, 0.26wt% S, and trace amounts of nonferrous metals (0.005wt% Cu, 0.013wt% Pb, and 0.051wt% Zn) was obtained. The concentrate can be potentially used as a high-quality feed material for producing oxidized pellets by blending with other high-grade iron ore concentrates.     

Evolution of ferronickel particles during the reduction of low-grade saprolitic laterite nickel ore by coal in the temperature range of 900–1250℃with the addition of CaO–CaF2–H3BO3

The method of producing ferronickel at low temperature(1250–1400℃)has been applied since the 1950s at Nippon Yakin Kogyo,Oheyama Works,Japan.Limestone was used as an additive to adjust the slag composition for lowering the slag melting point.The ferronickel product was recovered by means of a magnetic separator from semi-molten slag and metal after water quenching.To increase the efficiency of magnetic separation,a large particle size of ferronickel is desired.Therefore,in this study,the influences of CaO,CaF₂,and H₃BO₃ additives on the evolution of ferronickel particle at≤1250℃were investigated.The experiments were conducted at 900–1250℃with the addition of CaO,CaF₂,and H₃BO₃.The reduction processes were carried out in a horizontal tube furnace for 2 h under argon atmosphere.At 1250℃,with the CaO addition of 10 wt%of the ore weight,ferronickel particles with size of 20μm were obtained.The ferronickel particle size increased to 165μm by adding 10 wt%CaO and 10 wt%CaF₂.The addition of boric acid further increased the ferronickel particle size to 376μm,as shown by the experiments with the addition of 10 wt%CaO,10 wt%CaF₂,and 10 wt%H₃BO₃.     

菱铁矿在煤基直接还原条件下的转化过程

为研究菱铁矿在强还原气氛下加热过程中铁矿物的转化过程和规律,采用热重分析、X射线衍射和扫描电镜等手段研究了嘉峪关某菱铁矿石在煤基直接还原过程中菱铁矿的热行为和不同条件下焙烧产物中铁矿物的存在形式等.结果表明,菱铁矿在煤基直接还原条件下转化为金属铁的历程为FeCO3→Fe3O4→FeO→Fe.转化过程分为菱铁矿分解和铁氧化物还原两个阶段;热分解阶段在556.6℃时基本结束,最终产物为Fe3O4;铁氧化物的还原阶段在556.6℃以后、1200℃时完全结束,最终产物为金属铁.     

A novel process for the recovery of iron,titanium, and vanadium from vanadium-bearing titanomagnetite:sodium modification-direct reduction coupled process

A sodium modification-direct reduction coupled process was proposed for the simultaneous extraction of V and Fe from vanadium-bearing titanomagnetite. The sodium oxidation of vanadium oxides to water-soluble sodium vanadate and the transformation of iron oxides to metallic iron were accomplished in a single-step high-temperature process. The increase in roasting temperature favors the reduction of iron oxides but disfavors the oxidation of vanadium oxides. The recoveries of vanadium, iron, and titanium reached 84.52%, 89.37%, and 95.59%, respectively. Moreover, the acid decomposition efficiency of titanium slag reached 96.45%. Compared with traditional processes, the novel process provides several advantages, including a shorter flow, a lower energy consumption, and a higher utilization efficiency of vanadium-bearing titanomagnetite resources.     

鲕状赤铁矿悬浮焙烧的磁性研究

针对鄂西某鲕状赤铁矿进行悬浮焙烧研究,并采用振动样品磁强计、X射线衍射分析仪、穆斯堡尔谱仪分析还原温度、还原时间、氧化温度、颗粒粒度对焙烧物料磁性和物相组成的影响规律.结果表明:铁矿石经悬浮焙烧后磁性明显增强,且焙烧物料磁性与强磁性铁矿物的含量呈正比.当还原温度为550~650℃时,还原物料的磁化强度和比磁化率随还原温度的升高而升高,超过700℃后则随之降低.延长还原时间可提高还原物料的磁化强度和比磁化率.焙烧物料中γ-Fe2O3含量随氧化温度升高而增加,在氧化温度为350℃时物料中γ-Fe2O3的含量达到最大值.当焙烧物料颗粒粒度小于15μm时,颗粒的磁化强度和比磁化率随之降低,而剩磁和矫顽力则随之增加.     

复杂银铜精矿的沸腾焙烧综合回收工艺

采用沸腾焙烧综合回收工艺对呷村复杂银铜精矿难浸出问题进行实验,以沸腾焙烧脱硫、脱砷→硫酸浸出铜、锌→氯盐浸出锑、银→NaClO3氧化浸金→碳铵转化铅的工艺进行处理,实现铜、锌、银、锑、铅、金、砷和硫等有价元素的综合回收.在硫酸化沸腾焙烧过程中控制1·1倍空气过剩系数、0·25~0·35m·s-1的工况炉膛线速度以及600~630℃的焙烧温度,可以避免高铅复杂银铜精矿的烧结,脱硫率为49·68%,烟气中SO体积分数为5·5%可满足制酸要求.     

生活垃圾焚烧飞灰中Zn、Pb、Cu的浸出特性

测定了飞灰中Zn、Pb、Cu的浸出毒性及浸出率与pH值的关系,并利用MINTEQA2模型对它们的浸出率以及不同pH值下的主要存在形态进行了模拟。结果表明:Pb、Cu、Zn的模拟浸出值与实测浸出值基本相符,模拟浸出值可反映实际浸出情况;pH值在8~12之间时Zn、Cu、Pb的浸出率接近0;pH<7.5时,Zn的浸出率高达70%~80%,pH值>12时,随着pH值的增加,Pb与Zn的浸出能力逐渐增强;pH<2时,Zn、Cu、Pb的浸出率较大,表明三种重金属在强酸条件下容易浸出。了解重金属在不同pH值下的浸出情况可以帮助预测各种环境中重金属的潜在危害性,并可通过改变环境pH值控制重金属的浸出。