Selective reduction process of zinc ferrite and its application in treatment of zinc leaching residues

The traditional zinc hydro-metallurgy generates a large amount of zinc ferrite residue rich in valuable metals. The separation of iron is crucial for resource recycling of valuable metals in zinc ferrite residue. A novel selective reduction roasting？leaching process was proposed to separate zinc and iron from zinc leaching residue which contains zinc ferrite. The thermodynamic analysis was employed to determine the predominant range of Fe3O4 and ZnO during reduction roasting process of zinc ferrite. Based on the result of thermodynamic calculation, we found thatV（CO）/V（CO＋CO2） ratio is a key factor determining the phase composition in the reduction roasting product of zinc ferrite. In the range ofV（CO）/V（CO＋CO2） ratio between 2.68% and 36.18%, zinc ferrite is preferentially decomposed into Fe3O4 and ZnO. Based on thermogravimetric （TG） analysis, the optimal conditions for reduction roasting of zinc ferrite are determined as follows： temperature 700？750 °C, volume fraction of CO 6% and V（CO）/V（CO＋CO2） ratio 30%. Based on the above results, zinc leaching residue rich in zinc ferrite was roasted and the roasted product was leached by acid solution. It is found that zinc extraction rate in zinc leaching residue reaches up to 70% and iron extraction rate is only 18.4%. The result indicates that zinc and iron can be effectively separated from zinc leaching residue.

Sulfidation roasting of zinc leaching residue with pyrite for recovery of zinc and iron

Zinc leaching residue(ZLR) contains high content of valuable metals such as zinc and iron. However, zinc and iron mainly exist in the form of zinc ferrite, which are difficult to separate and recover. This study proposed a new process involving sulfidation roasting, magnetic separation and flotation to recover zinc and iron in ZLR. Through sulfidation roasting of ZLR with pyrite, zinc and iron were converted into ZnS and Fe3 O4. The effects of pyrite dosage, roasting temperature and roasting time on the sulfidation of zinc in ZLR were investigated. The results showed that the sulfidation percentage of zinc reached 91.8% under the optimum condition. Besides, it was found that ball-milling was favorable for the separation and recovery of zinc and iron in sulfidation products. After ball-milling pretreatment, iron and zinc were enriched from sulfidation products by magnetic separation and flotation. The grade of iron in magnetic concentrates was 52.3% and the grade of zinc in flotation concentrates was 31.7%, which realized the recovery of resources.

Recovery of valuable metals from zinc leaching residue by sulfate roasting and water leaching

Zinc leaching residue(ZLR),produced from traditional zinc hydrometallurgy process,is not only a hazardous waste but also a potential valuable solid.The combination of sulfate roasting and water leaching was employed to recover the valuable metals from ZLR.The ZLR was initially roasted with ferric sulfate at640°C for1h with ferric sulfate/zinc ferrite mole ratio of1.2.In this process,the valuable metals were efficiently transformed into water soluble sulfate,while iron remains as ferric oxide.Thereafter,water leaching was conducted to extract the valuable metals sulfate for recovery.The recovery rates of zinc,manganese,copper,cadmium and iron were92.4%,93.3%,99.3%,91.4%and1.1%,respectively.A leaching toxicity test for ZLR was performed after water leaching.The results indicated that the final residue was effectively detoxified and all of the heavy metal leaching concentrations were under the allowable limit.

Zinc extraction from zinc oxidized ore using(NH4)2SO4 roasting−leaching process

An improved method of(NH4)2SO4 roasting followed by water leaching to utilize zinc oxidized ores was studied.The operating parameters were obtained by investigating the effects of the molar ratio of(NH4)2SO4 to zinc,roasting temperature,and holding time on zinc extraction.The roasting process followed the chemical reaction control mechanism with the apparent activation energy value of 41.74 kJ·mol^−1.The transformation of mineral phases in roasting was identified by X-ray diffraction analysis combined with thermogravimetry–differential thermal analysis curves.The water leaching conditions,including the leaching temperature,leaching time,stirring velocity,and liquid-to-solid ratio,were discussed,and the leaching kinetics was studied.The reaction rate was obtained under outer diffusion without product layer control;the values of the apparent activation energy for two stages were 4.12 and 8.19 kJ·mol^−1.The maximum zinc extraction ratio reached 96%while the efficiency of iron extraction was approximately 32%under appropriate conditions.This work offers an effective method for the comprehensive use of zinc oxidized ores.

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 PbSO4 and ZnSO4 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 12.01wt%Pb,27.78wt%Zn,and 6.975×10^(−2)wt%Ag with recoveries of 60.54%,29.24%,and 57.64%,respectively,could be obtained after roasting.

Laboratory Study on the Removal of Fluorine from High Fluorine Bearing Zinc Oxide Ore

An ever increasing demand for zinc has resulted in worldwide efforts to exploit complex and lean grade reserves of zinc oxide ore. In this study experiments were done on zinc recovery from high-fluorine bearing zinc oxide ore. First the effect of different variables on fluorine removal was investigated. Optimum experimental conditions occurred when the temperature was 1173 K, roasting time was 90 min and air flow was at a velocity of 5 m/min, the fluorine removal from the samples reached over 93%. The results obtained indicate that fluorine removal is not enhanced by adding diluted sulfuric acid or water as a binder. Second the roasted materials were leached with a diluted sulfuric acid, neutralized, flocculated and filtrated. The which satisfies the requirement of fluorine residual fluorine content in the leached content in the leached solutions for the solutions was less than 38 mg/L production of electrolytic zinc.

锌浸渣浮选银精矿还原焙烧-低酸浸出脱锌

以某湿法炼锌厂产出的锌浸渣浮选银精矿为原料,采用回转炉还原焙烧—低酸浸出工艺进行脱锌研究。结果表明:在碳粉加入量10%、焙烧温度600℃、时间120 min、原料粒度-0.1 mm、回转炉转速5 r/min的条件下,铁酸锌还原焙烧分解效果最优,焙烧产物中可溶锌率达到85.83%。在硫酸终点pH为2、液固比5 mL/g、浸出温度70℃、浸出时间90 min优化条件下,对焙烧产物进行选择性浸出,锌和铁的浸出率分别为84.23%和34.71%。浸出渣中未溶解铁锌氧化物主要为还原焙烧过程中团聚成块的大颗粒及被硫酸铅熔化包裹形成的颗粒。

锌粉置换渣中镓、锗的强化浸出试验研究

针对锌粉置换渣提取Ga、Ge过程浸出效果不佳的问题,本研究首先开展了常压硫酸浸出试验,考察硫酸浓度、反应温度、液固比、反应时间对Ga、Ge浸出率的影响,并对浸出过程进行了动力学分析;而后分别尝试了还原浸出(先后以Na_(2)SO_(3)、SO_(2)为还原剂)、硫酸化焙烧-水浸两种工艺,对锌粉置换渣开展了Ga、Ge强化浸出试验。结果表明:常压硫酸浸出工艺在硫酸浓度150 g/L、反应温度90℃、液固体积质量比5 mL/g、反应时间4 h、搅拌转速400 r/min的适宜条件下,Ga、Ge浸出率分别为68.4%、61.7%;酸浸渣主要物相为PbSO_(4)和ZnAl_(2)O_(4),增加酸度可促使残留的黄钾铁矾、铅铝矾进一步溶出;Ga和Ge常压硫酸浸出过程主要受内扩散控制,反应表观活化能分别为18.06 kJ/mol和16.02kJ/mol。对一段常压硫酸浸出渣进行二段还原强化浸出,以Na_(2)SO_(3)为还原剂,在50 g/L H_(2)SO_(4)+3 g/L Na_(2)SO_(3)条件下,Ga、Ge浸出率达89.4%和70.1%;以SO_(2)为还原剂,在SO_(2)分压0.3 MPa、硫酸浓度50 g/L、反应温度110℃、、反应时间3 h条件下,Ga、Ge浸出率达88.9%和86.7%。硫酸化焙烧-水浸工艺对锌粉置换渣中Ga的浸出效果改善明显,但对Ge浸出不利,酸渣体积质量比宜为1/2.5 mL/g,并在焙烧温度300℃、焙烧时间3 h、水浸温度80℃、液固体积质量比5 mL/g、浸出时间2 h的条件下,Ga浸出率可提高到89.1%。

锌浸渣还原焙烧-磁选回收铁

在查明锌浸渣工艺矿物学的基础上,采用还原焙烧将铁酸锌分解为氧化锌和磁性氧化铁,再通过磁选的方法回收铁,达到锌、铁分离的目的。实验考查了焙烧温度、焙烧时间、还原剂用量对铁酸锌分解率、铁回收率和铁品位的影响。结果表明:在焙烧温度为950℃、焙烧时间为1 h及还原剂添加量为10%和5%的条件下,铁酸锌分解率达到72.05%,铁回收率可达到91.79%,精矿中铁的品位为50%左右。焙烧及磁选过程中颗粒的团聚包裹是铁精矿品位不高的主要原因。

锌浸渣硫酸铵焙烧-选择性浸出回收锌

针对锌浸渣中锌难于选择性浸出回收的难题,提出硫酸铵焙烧-选择性浸出回收锌的新工艺。该工艺通过硫酸铵焙烧改变锌浸渣中锌铁物相,在浸出过程对锌进行选择性浸出回收。研究硫酸铵加入量、焙烧温度、焙烧时间等工艺参数对铁酸锌分解和锌铁浸出的影响,并获得最佳的工艺参数,即硫酸铵和铁酸锌质量比为4、一段焙烧温度和时间分别为450℃和90 min,二段焙烧温度和时间分别为650℃和60 min。在该条件下,锌浸出率可以达到92.63%,而铁的浸出率仅为2.04%,实现了锌浸渣中锌的选择性浸出。

湿法炼锌沉铁渣和浸锌渣的焙烧预处理

采用电阻炉对湿法炼锌沉铁渣和浸锌渣进行焙烧预处理。考察氧气浓度、焙烧温度、焙烧时间和气体流量对沉铁渣脱硫率以及焙烧矿硫含量的影响;利用XRD和SEM-EDS技术对沉铁渣焙烧矿和浸锌渣焙烧矿的物相及微观形貌进行表征。结果表明:焙烧脱硫处理湿法炼锌沉铁渣和浸锌渣是可行的;沉铁渣焙烧预处理的合适工艺条件为氧气浓度70%,气体流量0.2 L/min,经1300℃焙烧反应20 min,沉铁渣脱硫率为98.2%,焙烧矿硫含量从8.15%降低到0.253%(质量分数),达到炼铁入炉原料硫含量的要求;沉铁渣焙烧矿的物相组成主要是Zn Fe2O4,浸锌渣焙烧矿的物相组成主要是Zn Fe2O4和Pb2SO5,Pb2SO5的生成表明浸锌渣中Pb SO4分解脱硫不彻底,这是浸锌渣焙烧矿硫含量偏高的主要原因。

从高酸浸出锌渣中回收银研究

针对某锌冶炼厂高酸浸出渣含银高的特点,采用硫酸化焙烧—水浸脱锌铁—氯化浸银—冷却结晶PbCl2—铅片置换沉银工艺,对高酸浸出锌渣进行了回收银研究。结果表明,锌和铁的浸出率分别达到92.66%和94.67%,浸出液返回炼锌主流程生产电锌;银和铅的浸出率分别达到94.17%和97.89%;用铅片置换得到粗银粉,银置换率达到99%以上。

浸锌渣综合利用新工艺及镓的富集行为

研究了浸锌渣还原焙烧分选综合回收有价元素新工艺 ,并采用电子显微镜、能谱仪和扫描电镜等分析了还原焙烧渣中金属的性质 .研究结果表明 :当还原温度为 110 0℃、还原时间为 15 0min时 ,还原焙烧渣中铁的金属化率、镓的回收率、锌的挥发率分别为 95 .10 % ,89.10 % ,98.4 2 % .还原焙烧渣经破碎、磨矿、磁选分离获得的磁性产物中含Fe 90 .16 % ,Ga的质量浓度为 2 16 4g t;Fe ,Ga的回收率分别为 87.78% ,92 .4 2 % ;还原焙烧渣中金属铁是镓的主要载体矿物相 ,镓具有明显的亲铁特性 ;

从锌浸出渣中湿法回收锌

研究了采用硫酸焙烧—水浸出工艺从锌浸出渣中回收锌。考察了硫酸用量、焙烧时间、焙烧温度、液固体积质量比及浸出时间对锌浸出率的影响。试验结果表明:锌浸出渣与70%浓硫酸混合,在250℃下焙烧2.5h,然后以水为浸出剂,按液固体积质量比4∶1、常温下浸出1h,锌浸出率达85%以上。

从锌冶炼渣中回收银的试验研究

某锌冶炼渣含高品位银360 g/t,银在浸出渣中的形态比较复杂,且绝大部分银被黄钾铁矾包裹,回收困难。根据浸出渣性质进行了试验研究,最终采用焙烧浸出除铁、焙烧浸出渣浮选富集银,得到的银精矿品位为3 899 g/t,银回收率为88.09%,实现了资源的综合回收。

浸锌渣中有价元素的综合利用

采用还原焙烧分选法 ,从湿法炼锌的浸锌渣中富集回收有价元素 Ga、Ge、Ag等。结果表明 :当还原焙烧温度 1 1 0 0℃、时间 3h,焙烧产品磁选所得精矿 TFe、Ga、Ge含量分别为 90 .67%、1 997g/ t、1 4 1 0 g/ t, 回收率分别为 88.1 2 %、88.1 0 %和 98.33% ; 尾矿含 Ag1 4 0 3g/ t,回收率为85.85%。

焙烧制度对煤泥-浸锌渣冷固结球团还原行为的影响

基于直接还原法探讨了焙烧制度对煤泥-浸锌渣冷固结球团中锌、铅挥发率和铁金属化率的影响,分析了焙烧制度对球团中含锌、铅、铁化合物相变的影响,试验确定了焙砂磨矿—弱磁选回收其中铁的工艺和效果。结果表明:在1 250℃焙烧90 min,可使球团中锌、铅的挥发率分别达到98.87%、95.39%,铁的金属化率达到98.66%;焙砂中未见锌、铅单质及其化合物,只存在大量的金属铁,且金属铁颗粒多数大于30μm;焙砂采用2段磨矿、2段弱磁选流程处理,可同时获得含铁91.20%、回收率为30.32%的金属铁粉和铁品位为61.58%、回收率为50.01%的铁精矿,铁总回收率达80.33%。

硫精矿焙砂提金及铜锌综合回收试验研究

针对云南某矿山硫精矿焙砂中铜、锌、硫、砷含量高及铁氧化物包裹金等问题,开展了常规氰化浸出和酸浸—氰化浸出两种不同方案回收金的试验研究。结果表明,常规氰化浸出金的浸出率为84.52%,酸浸—氰化浸出的酸浸过程铜、锌浸出率分别为40.25%、38.79%,后续金的浸出率为85.82%。综合比较,酸浸—氰化浸出工艺更优,比常规氰化浸出的浸出率高1.54%,氰化钠的消耗量降低1 kg/t以上,同时还可以综合回收铜、锌。

锌湿法冶炼渣的自热焙烧无害化处理

针对湿法炼锌常规浸出渣和氧压浸出渣(均为危险废物)组成的高硫混合渣,系统考察了自热焙烧反应的各因素(温度、硫含量等)对渣中重金属固化效果的影响,并采用BCR连续提取法研究了焙烧前后渣中重金属的化学形态分布。结果表明,在焙烧温度1000℃、混合物料硫质量分数20%、空气流量1.6 m^(3)/h、焙烧时间7 h的条件下,渣中Zn、Pb、As、Cd等毒性浸出结果均可满足国家标准,渣率可降至约70%,焙烧热处理后渣中的Zn、Pb、As及Cd均以相对稳定的残渣态为主要化学形态,为锌湿法冶炼渣的低成本无害化与减量化处理提供了理论基础。

黄铵铁矾渣的焙烧酸浸行为研究

对黄铵铁矾渣的组成结构及其焙烧酸浸行为进行研究,以达到使渣中的铟、锌和铁得到高效分离和利用的目的.由MLA分析推测黄铵铁矾渣的结构式为NH4Fe3（SO4）2（OH）6,渣中的铁主要包含在黄铵铁矾结构中,锌主要包含在水锌矾Zn[SO4]·H2O结构中.在680～720℃焙烧1.5h后酸浸,铟的浸出率大于82％,锌的浸出率大于95％,铁的浸出率小于10％,实现了黄铵铁矾渣中的锌、铟和铁的有效分离.