Deselenization and detellurization of precious-metal ore concentrates by swelling oxidizing roasting and successive alkaline leaching

A new technique of swelling oxidizing roasting and alkaline leaching was proposed for deselenization and detellurization of precious-metal ore concentrates. Alkali-metal and alkaline-earth-metal chlorides and carbonates were preliminarily selected as swelling agents. The roasting removal rate and alkaline leaching rate of selenium and tellurium were investigated, and NaCl was selected as an appropriate swelling agent. Furthermore, the effects of various factors on the selenium gasification rate and leaching rate of selenium and tellurium were investigated. The results show that the selenium gasification rate reaches 88.41% after swelling oxidizing roasting for 2 h at 510℃ using an NaCl dosage coefficient of 100 and a sulfuric acid dosage coefficient of 1.3; the amorphous elemental tellurium is completely transformed into TeO₂. The roasted product is subjected to alkaline leaching using a 100 g/L NaOH solution, which results in a selenium leaching rate of 10.51%, a total selenium removal rate of 98.92%, and a tellurium leaching rate of 97.64%. In the alkaline leaching residue, the contents of selenium, tellurium, gold, platinum, and palladium are 0.7825%, 5.492%, 8.333%, 0.2587%, and 1.113%, respectively; the precious metals are enriched approximately sixfold.     

Recovery of valuable metals from a low-grade nickel ore using an ammonium sulfate roasting-leaching process

Metal leaching from a low-grade nickel ore was investigated using an ammonium sulfate roasting-water leaching process. The nickel ore was mixed with ammonium sulfate, followed by roasting and finally leaching with water. During the process the effects of the amount of ammonium sulfate, roasting temperature, and roasting time on the leaching recovery of metal elements were analyzed. The optimum technological parameters were determined as follows:ammonium sulfate/ore ratio, 0.8 g/g; roasting temperature, 400℃; and roasting time, 2 h. Under the optimum condition the leaching recoveries of Ni, Cu, Fe, and Mg were 83.48%, 76.24%, 56.43%, and 62.15%, respectively. Furthermore, the dissolution kinetics of Ni and Mg from the nickel ore was studied. The apparent activation energies for the leaching reaction of Ni and Mg were 18.782 and 10.038 kJ·mol-1, which were consistent with the values of diffusion control reactions. Therefore, the results demonstrated that the leaching recoveries of Ni and Mg were controlled by diffusion.     

Extraction of vanadium from high calcium vanadium slag using direct roasting and soda leaching

The extraction of vanadium from high calcium vanadium slag was attempted by direct roasting and soda leaching. The oxidation process of the vanadium slag at different temperatures was investigated by X-ray diffraction (XRD), scanning electron microscopy (SEM), and energy dispersive spectroscopy (EDS). The effects of roasting temperature, roasting time, Na₂CO₃ concentration, leaching temperature, leaching time, and liquid to solid ratio on the extraction of vanadium were studied. The results showed that olivine phases and spinel phases in the vanadium slag were completely decomposed at 500 and 800℃, respectively. Vanadium-rich phases were formed at above 850℃. The leaching rate of vanadium reached above 90% under the optimum conditions:roasting temperature of 850℃, roasting time of 60 min, Na₂CO₃ concentration of 160 g/L, leaching temperature of 95℃, leaching time of 150 min, and liquid to solid ratio of 10:1 mL/g. The main impurities were Si and P in the leach liquor.     

低品位钼精矿的钼提取研究

采用焙烧—氨浸—渣碱浸工艺对某低品位钼精矿进行钼提取的研究.结果表明,碳酸钠的加入能有效分解钼酸钙,提高钼的提取率.焙烧—氨浸阶段最优工艺条件为焙烧温度600℃,焙烧时间2 h,氨浸温度80℃,氨水过量系数1.4,碳酸钠用量467 kg/t,液固比4.浸出渣中的钼分别采用酸法(HCl)和碱法(Na2CO3+NaOH)进行提取.结果表明:酸法仅可回收渣中34.92%的钼,而且操作过程不易控制,不适合实际应用;碱法(Na2CO3+NaOH)处理工艺中碳酸钠用量533 kg/t,氢氧化钠用量433 kg/t.放大实验结果显示整个流程钼的回收率达到96.8%.     

高镍铜阳极泥预处理富集金银的研究

由于高镍铜阳极泥是典型的难处理铜阳极泥,故以高镍铜阳极泥为原料,考察了温度、时间、液固比等因素对贱金属硒、铜和镍脱除效果的影响.研究结果表明,经过两次焙烧和浸出,可脱除995%的硒、997%的铜、9335%的镍和浸出9876%的银,且金从193g·t-1富集到1820g·t-1,增加了8~9倍.第一段焙烧和浸出条件:温度650℃、焙烧时间1h、酸泥质量比12、浸出温度55℃、浸出时间1h、液固质量比6.第二段焙烧和浸出条件:焙烧温度500℃、焙烧时间3h、酸泥质量比12、浸出温度55℃、液固质量比6、浸出时间1h.经过预处理之后,阳极泥的量减少为原来的1126%,大大提高了后续回收工序中的设备处理能力.     

银金精矿的焙烧条件对焙砂中金银提取的影响

在热分析的基础上对吉林某浮选银精矿的预处理过程进行了研究。研究了焙焙烧温度、焙烧时间奶对金精矿中硫，碳脱除率的影响，着重研究了有添加剂存在时焙烧工艺条件对银精矿焙砂的金银提取性能的影响，并了添剂在硫酸化焙烧过程中的作用。试验结果表明：有添加剂存在时的焙少经稀硫酸预浸，其预浸渣于用硫脲法或氰化法浸金时，其金的浸率将大于９５％，争遥总浸出率也分别达９５％     

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.     

A method for recovery of iron,titanium, and vanadium from vanadium-bearing titanomagnetite

An innovative method for recovering valuable elements from vanadium-bearing titanomagnetite is proposed. This method involves two procedures:low-temperature roasting of vanadium-bearing titanomagnetite and water leaching of roasting slag. During the roasting process, the reduction of iron oxides to metallic iron, the sodium oxidation of vanadium oxides to water-soluble sodium vanadate, and the smelting separation of metallic iron and slag were accomplished simultaneously. Optimal roasting conditions for iron/slag separation were achieved with a mixture thickness of 42.5 mm, a roasting temperature of 1200℃, a residence time of 2 h, a molar ratio of C/O of 1.7, and a sodium carbonate addition of 70wt%, as well as with the use of anthracite as a reductant. Under the optimal conditions, 93.67% iron from the raw ore was recovered in the form of iron nugget with 95.44% iron grade. After a water leaching process, 85.61% of the vanadium from the roasting slag was leached, confirming the sodium oxidation of most of the vanadium oxides to water-soluble sodium vanadate during the roasting process. The total recoveries of iron, vanadium, and titanium were 93.67%, 72.68%, and 99.72%, respectively.     

钼精矿焙烧处理过程中典型杂质元素的热力学行为及分析

通过模拟实验结合X射线荧光光谱仪(X-ray fluorescence spectrometer,XRF)、电感耦合等离子光谱发生仪(inductively coupled plasma spectrometer,ICP)及矿物解离分析仪(mineral liberation analyser,MLA)等研究了钼精矿焙烧处理流程中多种杂质元素间的相互作用,在MLA对物相定量分析的基础上,采用Factsage7.0软件分析了钼精矿焙烧过程中不同杂质元素反应进行的热力学条件。结果表明：杂质元素在钼精矿及后续处理流程中的分布存在明显的粒度偏析,主要表现为K、Si等杂质更多以大分子量的矿物形式赋存在较粗粒度的钼精矿中,而Fe、Ca、Cu等杂质则更多以FeS₂、CaSO₄、CuFeS₂等小分子量的化合物形式赋存在较细粒度的钼精矿中。FeS₂、CaSO₄和SiO₂等杂质会在高温焙烧过程中与MoO₃形成致密度较高的混合物,降低Mo的浸出率。云母在钼...     

氯化铵焙烧生石灰提取可溶性钙

为实现生石灰的高附加值利用,采用水浸氯化铵与生石灰的焙烧熟料提取氧化钙,考察了焙烧温度、焙烧时间及铵矿物质量的比对生石灰中氧化钙提取率的影响.通过正交试验,确定了最佳提取氧化钙焙烧条件.研究结果表明:最佳的焙烧工艺为焙烧温度250℃,焙烧时间90min,铵矿物质的量比2.2∶1,此时,氧化钙的提取率可达97.26%.利用差热法进行了机理分析,结果表明氯化铵焙烧生石灰过程总体分三个化学反应步骤.     

高铝硅氰化渣中铁回收工艺

研究一种处理磁选前高铝硅氰化渣的新工艺。采用复合添加剂焙烧-水浸-磁选工艺对一种铁品位为27.69%(质量分数),SiO2含量为23.9%,Al2O3含量为6.35%的高铝硅氰化渣进行杂质与铁分离的研究。研究结果表明:在最佳焙烧条件下,当水浸温度为60℃,液固比为15:1,水浸时间为5 min,转速为20 r/min,在激磁电流为2 A时,可获得铁品位57.11%,铁的回收率为72.58%的铁精矿。铁的品位和回收率都比单纯的复合添加剂还原焙烧-磁选法所获得的铁精矿的指标高,铁的品位提高了10%左右,回收率提高了30%左右。X线荧光(XRF),X线衍射(XRD)及能谱(EDS)分析研究结果表明:经水浸后,复合添加剂焙烧过程中所产生的可溶性复杂杂质化合物被洗除,不溶性物质经磁选后随之进入非磁性物,实现铁与杂质矿物之间的有效分离。     

Investigation on microwave heating for direct leaching of chalcopyrite ores and concentrates

The use of microwave energy in materials processing is a relatively new development presenting numerous advantages because of the rapid heating feature. Microwave technology has great potential to improve the extraction efficiency of metals in terms of both a reduction in required leaching time and an increase in the recovery of valuable metals. This method is especially pertinent in view of the increased demand for environment-friendly processes. In the present study, the influence of microwave heating on the direct leaching of chalcopyrite ores and concentrates were investigated. The results of microwave leaching experiments were compared with those obtained under conventional conditions. During these processes, parameters such as leaching media, temperature, and time have been worked to determine the optimum conditions for proper copper dissolution. Experimental results show that microwave leaching is more efficient than conventional leaching. The optimum leaching conditions for microwave leaching are the solid-to-liquid ratio of 1:100 g/mL, the temperature of 140℃, the solution of 0.5 M H₂SO₄ + 0.05 M Fe₂(SO₄)₃, and the time of 1 h.     

含钼铜精矿氧化焙烧-浸出分离钼研究

四川攀西地区的含钼铜精矿中,由于钼、铜矿物组成复杂,共生关系紧密,提出了氧化焙烧-浸出工艺进一步分离钼。将试样置入焙烧炉中进行氧化焙烧,使硫化物转化为CuO、MoO3、Fe2O3等氧化物后;添加NaOH、H2O与MoO3反应生成可溶性Na2MoO4,浸出渣作为铜精矿产品。研究结果表明:铜、钼等以氧化物形式赋存于焙烧矿中,氧化焙烧矿中的硫含量较低为0.53%,硫以SO2气相形式挥发;在焙烧温度650℃、焙烧时间120 min、氢氧化钠用量为30%、浸出温度60℃、浸出时间120min、浸出液固比2∶1的综合条件下,钼的浸出率为94.24%,铜精矿(浸出渣)中铜的质量分数为24.27%,钼分离效果明显。     

载金硫化物焙烧--自浸出过程研究

针对传统氧化焙烧-氰化浸金工艺环境污染严重的现状,采用焙烧-自浸出工艺提取载金硫化物中的金.研究焙烧温度、焙烧时间和试样量对单质硫转化率和金浸出率的影响,通过X射线衍射分析、扫描电镜观察、能谱分析等手段分析焙烧过程中载金硫化物中硫的物相转变规律.载金硫化物中黄铁矿发生热分解反应生成单质硫和磁黄铁矿,随焙烧温度的升高和焙烧时间的延长,黄铁矿的特征衍射峰强度逐渐减小直到消失,磁黄铁矿的特征衍射峰逐渐生成并增强,原本致密状的黄铁矿颗粒变得疏松多孔.50 g试样在氮气流量1 L·min-1、焙烧温度800℃、焙烧时间60 min的条件下,单质硫的转化率达到42.53%,金浸出率达到88.70%,实现载金硫化物的高效非氰浸出.     

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.     

NaOH-Ca(OH)2分解独居石精矿的研究

为降低独居石分解工艺的加碱量,并提高独居石分解率,本研究在NaOH-Ca(OH)₂体系中对独居石精矿进行焙烧分解,采用XRD对焙烧产物进行物相分析,并结合焙烧矿中稀土元素在盐酸中的浸出率判断独居石分解效果.实验分别研究了NaOH添加量、Ca(OH)₂添加量、焙烧温度以及焙烧时间对独居石精矿分解的影响.结果表明,在NaOH-Ca(OH)₂体系中,独居石精矿分解的最佳工艺条件:NaOH添加量为25%,Ca(OH)₂添加量为20%,焙烧温度为800℃,焙烧时间为1.5h.该焙烧条件下独居石全部分解为稀土氧化物,浓盐酸对稀土浸出率可达到98%左右.与现有工业生产工艺相比,本研究工艺中碱添加量降低55%左右,独居石分解率提高2%左右.     

采用焙烧-水浸法从报废锂离子动力电池中回收金属

锂离子动力电池在新能源汽车中已获得广泛应用,其报废后Li、Ni、Co、Mn等金属清洁高效回收对促进有色金属循环利用具有重要意义.从LiNi_0.5Co_0.2Mn_0.3O₂为正极材料的锂离子动力电池中回收Li、Ni、Co、Mn,并采用TG-DSC、XRD、ICP-OES、XPS、热力学分析等研究了回收过程物相演变规律及影响金属回收率的主要因素.结果表明:由LiNi_0.5Co_0.2Mn_0.3O₂与NaHSO₄·H₂O组成的混合物,经过焙烧后Li、Ni、Co、Mn元素的赋存状态发生改变,从不溶于水的复杂金属氧化物形式,转化为可溶于水的金属硫酸盐形式.焙烧产物在一定条件下用水浸出后,Li、Ni、Co、Mn元素以金属离子的形式转移到水溶液中获得回收.混合物的组成、焙烧温度对Li、Ni、Co、Mn元素在焙烧产物中的赋存形式呈现制约关系,也是影响Li、Ni、Co、Mn金属回收率的主要因素.     

Thermodynamics and phase transformations in the recovery of zinc from willemite

Willemite is a common component of zinc and lead metallurgical slags that, in the absence of effective utilization methods, cause serious environmental problems. To solve this problem and increase zinc recovery, we proposed a novel extraction method of zinc from willemite by calcified roasting followed by leaching in NH₄Cl-NH₃·H₂O solution. The thermodynamics and phase conversion of Zn₂SiO₄ to zinc oxide (ZnO) during calcified roasting with CaO were investigated. The mechanism of mineralogical phase conversion and the effects of the CaO-to-Zn₂SiO₄ mole ratio (n(CaO)/n(Zn₂SiO₄)), roasting temperature, and the roasting time on zinc-bearing phase conversion were experimentally investigated. The results show that Zn₂SiO₄ was first converted to Ca₂ZnSi₂O₇ and then to ZnO. The critical step in extracting zinc from willemite is the conversion of Zn₂SiO₄ to ZnO. The zinc percent leached in the ammonia leaching system rapidly increased because of the gradual complete phase conversion from willemite to ZnO via the calcified roasting process.     

粉煤灰中稀有金属镓-铌-稀土的联合提取

 采用碱法烧结-分步浸出法, 对重庆安稳电厂循环流化床粉煤灰中Ga、Nb、REE 等稀有金属进行了联合提取实验。结果表明, 粉煤灰加无水碳酸钠在860℃下烧结30 min, 采用水浸法提取Ga, 采用酸浸法提取REE, Ga、REE 的提取率分别达到84.70%和80.07%;Nb 在两步浸出实验中的浸出率均低于1%, 但在酸浸滤渣中得到富集。采用D201 离子交换树脂和NH₄Cl(0.5 mol/L)溶液在40℃下对富Ga 水浸滤液中的Ga 进行吸附与解吸附, Ga 的吸附率(27.99%)、解吸附率(37.33%)偏低, 可能与解吸附液用量不足和水浸滤液中Al 离子竞争吸附有关, 后续将通过改进实验条件提升Ga 的分离提取效果;酸浸液中REE 及酸浸渣中Nb 的分离提取工艺尚需进一步研究。通过上述稀有金属Ga、Nb、REE 的联合提取及后续工作, 可实现安稳电厂粉煤灰的高附加值利用, 有效缓解粉煤灰造成的环境污染。     

多金属矿选择性浸出的物理化学

对难以分选的铜、铝、锌复合矿,可经混合浮选,得到多金属混合精矿后,再用冶炼方法分离提取各种金属,以实现矿产资源的综合利用。本文论述了选择性硫酸化焙烧及选择性浸出的物理化学,为选、冶结合处理多金属复合矿提供了理论依据。