云南某含锡多金属硫化矿选矿实验研究

对云南某含锡多金属硫化矿进行了工艺矿物学和选矿实验研究. 结果表明,矿石中铅锌品位低,铅、锌矿物相互交代、包裹,嵌布粒度不均匀,采用优先浮铅、再选锌的原则流程,利用铅矿物与锌、硫矿物间可浮性差异较大的特点,采用石灰、亚硫酸钠和硫酸锌抑制锌、硫,以乙基黄药为铅捕收剂优先浮选铅矿物,选铅尾矿用硫酸铜作活化剂活化闪锌矿选锌;锡矿物与黄铁矿、磁黄铁矿等矿物共生关系复杂,且嵌布粒度较细,选锌尾矿经脱硫浮选后采用重磁联合流程回收锡矿物. 通过闭路实验,得到含铅40.92%、银1610.53 g/t、铅回收率81.25%、银回收率77.03%的铅精矿,锌精矿含锌43.23%、回收率为85.92%,硫精矿含硫42.57%,作业回收率为87.65%,锡精矿含锡42.38%,作业回收率为59.29%.     

从云南某高硫铅锌矿尾矿中综合回收锌的试验研究

云南某高硫铅锌矿尾矿平均含Pb 0.84％、Zn 3.67％,综合回收价值较高,其中含铅矿物主要为方铅矿和白铅矿,含锌矿物主要为闪锌矿和菱锌矿.实验室选矿试验研究表明,采用"硫化矿优先浮选-浮硫尾矿脱泥-氧化锌浮选"工艺可较好地实现尾矿中有价矿物的高效回收,进一步研究发现在氧化锌浮选过程中,粗选和扫选分别采用不同碳链长...     

山东某铅锌银多金属矿选矿试验研究

针对山东某铅锌银多金属矿嵌布关系复杂多变的特点,采用银铅锌混合浮选后铅锌分离和铅优先浮选两种技术路线进行试验.采用铅优先浮选技术路线使铅锌分离,得到了较好的选别指标,铅精矿中铅品位为70.04％,铅回收率97.75％,银品位1541 g/t,银回收率79.56％;锌精矿中锌品位43.73％,锌回收率82.55％.     

云南某富银锌精矿锌银浮选分离实验研究

云南某富银锌精矿中银主要以类质同象形式存在于白铅矿中,本工作以该矿样为研究对象,根据其性质,采用抑铅浮锌的工艺流程进行浮选,考察了磨矿细度、抑制剂、活化剂及捕收剂等因素对浮选分离指标的影响。结果表明,–19+10 μm粒级中银含量最高。在磨矿细度–74 μm占90%,硅酸钠用量为2000 g/t,硫酸铜用量为200 g/t,丁基黄药用量为300 g/t,松醇油用量为30 g/t的条件下,1次粗选、1次精选、1次扫选,中矿顺序返回的全流程闭路实验,可获得含锌61.08%,回收率95.89%的锌精矿和含银1548.32 g/t,回收率为71.17%的银精矿,实现了锌银的浮选分离及伴生银的高效富集。     

CaO-SiO2-FeO体系中选择性分离锌铅的实验研究

为了探讨CaO-SiO2-FeO体系中选择性分离回收Zn, Pb的可行性,在系统分析该体系中Zn, Pb挥发行为的基础上,研究了温度、基体成分及Cl含量等因素如对Zn, Pb挥发率的影响. 结果表明,低温、低碱度、低FeO含量、低Cl含量条件下,Zn与Pb的挥发率差异达到最大,原因在于该种条件下Zn倾向于以ZnFe2O4等复杂化合物形式滞留在渣基体中,而Pb以氯化物形式挥发;在此基础上,对实验后的渣基体进行升温、并添加一定量的C粉等还原剂后,渣中Zn会以单质形式挥发分离,从而实现粉尘及飞灰熔融处理过程中Zn、Pb的选择性分离回收.     

The Effect of Copper Bearing Particles Liberation on Copper Recovery from Smelter Slag by Flotation

The impact of the liberation of copper bearing particles on copper recovery in the flotation has been the subject of research in this paper. Tests have shown that grinding of material highly impacts the recovery rate in the flotation process. Results of flotation of smelter slag samples with different contents of grain size fraction ?74 µm: 50%, 60%, 70%, 80%, and 90%, have shown that by an increasing the content of the ?74 µm fraction in the feed, the recovery of copper was increasing as well. The highest recovery rate of copper was obtained with 90% of grain size fraction ?74 µm in the feed. The microscopic analysis of concentrates have shown that increased content of grain size fraction ?74 µm, was followed by increased contribution of liberated particles in the 0 to 70 µm fraction, while the contribution of middlings had been decreasing.     

云南某高硫铅锌矿流程优化试验研究

针对云南某铅锌选矿厂产出的硫精矿中铅锌品位高、铅精矿中铅品位低的问题,考查了目的矿物的解离情况,通过铅硫混合粗精矿再磨提高了目的矿物的单体解离度,优化了工艺流程。试验结果表明:在磨矿细度为-325目质量分数占70%时,采用铅硫混合粗精矿再磨、脱锌扫选尾矿返至锌作业的工艺流程,可得到铅品位为61.23%、铅回收率为85.68%的铅精矿,锌品位为49.65%、锌回收率为93.38%的锌精矿;与不再磨流程相比,铅精矿中铅的品位提高了2.05个百分点,回收率提高了1.75个百分点,锌精矿中锌的回收率提高了2.28个百分点,同时铅精矿中锌品位及硫精矿中铅、锌品位均有所降低。     

Importance of mineralogy and reaction kinetics for selecting leaching methods of copper from copper smelter slag

Ammoniacal leaching and sulfuric acid leaching were explored separately to recover copper from smelter slag. The overall 75% copper recovery was achieved from the smelter slag under optimum condition using ammoniacal leaching. The low copper recovery in this process is directly related to the complex slag mineralogy and dissolution kinetics. To overcome such problems, sulfuric acid leaching was applied separately. This method provided an overall 89% copper recovery under atmospheric condition. The use of sulfuric acid as the leaching agent was found to be more advantageous than ammonia due to the complex interlocking nature of copper bearing particles and different reaction kinetics of the slag.     

碳酸钠对镍渣碳热还原的催化作用

以全铁含量39.40wt%和SiO2含量32.50wt%的镍渣为原料,针对其中的铁以铁橄榄石形式存在难以直接还原磁选提铁的问题,在镍渣中添加不同质量比的碳酸钠促进镍渣碳热还原,进行了热力学计算和实验验证。结果表明,碳酸钠添加量由0增加至6wt%时,还原产物中铁的金属化率和回收率不断增大,继续增大碳酸钠添加量至8wt%时,铁的金属化率和回收率略有减小。不添加碳酸钠的还原产物中铁粒径很小,平均粒径为6 ?m,难还原的铁橄榄石大量存在,而加入6wt%碳酸钠的还原产物中铁粒径粗大,平均粒径增大至17 ?m,铁橄榄石含量明显降低,金属铁的XRD衍射峰强度明显增加。     

湖北某胶磷矿浮选工艺研究

以PF-QWS为正浮选捕收剂,HY-2B为反浮选捕收剂,碳酸钠为调整剂,水玻璃、六偏磷酸钠、磷酸、硫酸为抑制剂,对湖北某中低品位(17.09％)硅钙质磷矿石进行正反浮选工艺研究,在常温( 30℃)条件下,获得了P2O528.31％、MgO 1.91％、回收率82.86％的精矿指标.     

A Study on the Utilization of Ultrasonic Pretreatment in Zinc Flotation

This paper deals with the investigation of the effect of ultrasonic pretreatment on the mechanical flotation of zinc from lead-zinc-copper ore. After the chosen flotation parameters such as KAX dosage, frother amount, flotation time, stirring rate and amount of Na₂SiO₃, were optimized by flotation experiments without ultrasonic pretreatment, flotation experiments with ultrasonic pretreatment were carried out under these optimized conditions. The results obtained by both methods were compared. A zinc concentrate containing 8.52% Zn was obtained from feed containing 1.60%Zn with 24.64% recovery using mechanical flotation method without ultrasonic pretreatment. On the other hand, a concentrate containing 18.73%Zn was obtained with 33.18% recovery using flotation with ultrasonic pretreatment. It has been concluded that, ultrasonic pretreatment improved the grade and recoveries in zinc flotation due to the energy released by cavitation bubbles generated in the ultrasonic cleaning bath. XRD analysis were also supported the positive effect of ultrasonic pretreatment on flotation.     

氧化铅锌矿制活性氧化锌湿法工艺及其反应条件研究

对氧化铅锌矿制活性氧化锌湿法工艺进行了研究 ,包括酸浸工艺条件的优化、锌液净化条件、碱锌合成沉淀剂的选择及其反应条件。该工艺浸出率达 97%以上 ,锌总回收率达 92 %以上 ,并可实现铅锌分离     

碳酸钠在云南胶磷矿正浮选中的作用效果探索

碳酸钠作为一种无机调整荆,在胶磷矿选矿中被广泛用作pH调整剂、脉石细泥分散剂、活化剂等.本文对云南胶磷矿进行了加与不加碳酸钠对比试验,结果表明:在相同的刮泡时间内,加碳酸钠浮选流程的精矿产事为81.04%、累积回收率为90.97%,而不加碳酸钠的两次纯水玻璃浮选工艺漉程试硷的精矿产率分别为70.49%和74.28%、累积回收率分别为80.07%和74.85%,说明碳酸钠在胶磷矿正浮选中的活化作用非常明显.     

腾冲高泥氧化锌矿选矿实验研究

针对腾冲某氧化锌矿嵌布粒度细、含泥量高的技术难题进行了选矿实验研究. 结果表明,该矿中锌品位为6.65%、铁品位为12.31%,锌氧化率达92.83%,锌主要以异极矿形式存在,铁主要以磁铁矿形式存在. 经多个实验流程对比分析,采用预先洗矿(洗去小于37 mm粒级的颗粒)、沉砂磨矿后先磁选后浮选的流程,确定最佳磨矿细度为小于75 mm的颗粒达91.74%,弱磁选电流为4 A,粗选药剂用量为碳酸钠2 kg/t、六偏磷酸钠2 kg/t、硫化钠13 kg/t、KPR 2 kg/t. 采用闭路选矿流程,得到锌品位为27.89%、回收率75.26%的氧化锌精矿和铁品位为61.25%、回收率51.65%的铁精矿. 通过Zeta电位分析捕收剂KPR与异极矿的作用机理.     

磷化渣废液除锌制备高纯度磷酸三钠

为了实现磷化渣碱溶后所得废液的综合利用,采用中和沉淀法和硫化钠沉淀法去除磷化渣废液中的杂质锌,对中和沉淀法和硫化钠沉淀法的工艺条件进行优化,并由此设计磷化渣废液除锌制备高纯度磷酸三钠的生产工艺。对于中和沉淀法,在反应体系pH=8.85,反应时间为10 min时,Zn2+去除率可达97.96%;对于硫化钠沉淀法,在反应体系pH=6.85,硫化钠投加量与Zn2+的摩尔比为1:1,反应时间为10 min时,Zn2+去除率可达99.80%。相比于中和沉淀法除锌制备磷酸三钠工艺,硫化钠沉淀法除锌制备磷酸三钠工艺具有更高的磷酸三钠纯度,为98.85%,固相产品中Zn2+含量更低,仅为0.0004%,符合化工行业分析纯标准。该工艺大幅度地回收了磷化渣废液中的PO43-离子和Na+离子,为实现磷化渣综合利用的工业化提供新思路。     

资源化消纳高浸铅银渣底吹+侧吹炼铅工艺优化

铅锌资源互补是企业实现多金属综合回收的主流方向。针对汉中锌业底吹+侧吹炼铅系统混配高浸铅银渣、冶炼烟灰等含铅物料后,面临的氧化熔炼热量不足、炉渣过程控制困难、还原渣稳定性差等难题,在对原有工艺综合评价基础上,结合热力学分析和炉渣性能测定结果,提出了以渣型调控为主线的工艺优化思路。研究表明,底吹炉补热方式由硫磺为主转向以烟煤为主、硫铁矿为辅,可提升氧化熔炼温度至1050℃,还原熔炼降温至1250℃;控制还原熔炼渣FeO/SiO2质量比在1.2~1.6,CaO/SiO2质量比在0.4~0.6之间,ZnO≤20wt%,可确保混配含铅废料的直接炼铅过程顺行。工艺优化后,氧料比由120 Nm3/t降至110 Nm3/t,混配高浸铅银渣比例升至24wt%;氧化熔炼与还原熔炼烟尘率同比降低13.47%与15.82%,粗铅成本同比降低10.63%。     

云南高泥尾矿铅锌分离实验研究

针对云南高泥尾矿铅锌嵌布粒度细、泥化严重、性质相对复杂的技术难点,进行了铅锌分离实验研究,采用泥砂分级别浮选-重选联合工艺流程,有效分离了铅锌,实现了该复杂尾矿资源的综合回收利用。结果表明,该矿中铅含量为4.29wt%,锌含量为4.99wt%,铅主要以白铅矿和铅铁矾的形式存在,铅和铁相互交代形成不同的包裹形式,分离难度极大;锌主要以氧化锌的形式存在,氧化程度较深,锌氧化率达99%,且主要为难选的异极矿。最终通过闭路选矿流程,获得铅品位33.87%,回收率62.53%的铅精矿;铅精矿中银品位142.50 g/t,银回收率30.92%;获得锌品位15.21%、回收率47.82%的锌精矿。     

浮选法富集低品位柱硼镁石矿的工艺研究

以油酸钠和十二烷胺醋酸盐(DAA)为复合捕收剂,以偏硅酸钠为调整剂,利用反浮选工艺富集青海大柴旦低品位柱硼镁石(MgO·B2O3·3H2O)矿。实验结果表明,采用反浮选法可将柱硼镁石与石膏、碱式碳酸镁、食盐等杂质分离。通过实验确定了适宜的工艺条件,即:捕收剂十二烷胺醋酸盐用量1.5 kg/t,油酸钠用量0.15 kg/t,调整剂偏硅酸钠0.92 kg/t;液固比6∶1;浮选温度30～35℃;浮选时间10 min;矿浆pH=9.0～9.2。在此条件下,一次浮选所得1#硼精矿中B2O3品位由原矿的7.82%提高至12.3%,产率达60%以上。     

Effect of separation mechanism on the kinetics of Zn(II) flotation

ABSTRACT

In this study, the kinetics study and separation mechanism of zinc ion removal from synthetic wastewater were investigated using the foam separation process. With increasing the pH, Zn(II) recovery increased, in which foam fractionation occurred at pH 1.5, 3 and 5.5 and ion flotation at pH 8. The kinetics study was also used to determine separation mechanism. The order of Zn(II) foam fractionation was n = 1, while the order of Zn(II) ion flotation was n = 2 at the first stage and n = 0 at the second stage. Water recovery kinetics also increased with increasing ionic strength.     

Evaluation of diffusion mechanism and structural characterizations of ZnS compound during chemical reaction of Zn and S in liquid phase

The present study investigates the diffusion mechanism, morphology and structural characteristics of crystalline zinc sulfide (ZnS) during the reactive diffusion process. The samples with nominal composition of Zn₅₀S₅₀ were prepared via capsulation of high purity of zinc and sulfur followed by the annealing process at various reaction temperatures. The prepared samples were characterized using X-ray diffraction, differential scanning calorimetry and scanning electron microscopy. The structural measurements confirm the formation of zinc sulfide with wurtzite structure during the annealing process at 550°C. The wurtzite allotrope of ZnS is not stable at a high annealing temperature and is transformed to cubic zinc-blende structure. During the annealing process, a continuous layer of ZnS compound forms at the Zn/S interface. Both Zn and S diffuse into the formed ZnS layer and the growth of that occurs mainly toward the Zn side. Under this condition, Kirkendall voids form and accumulate near the ZnS/S interface.