低品位硫矿石利用扩建项目增量内部收益率问题及分析

某矿山低品位硫矿石利用扩建项目经济评价,根据"有项目""无项目"评价体系,采用增量内部收益率和增量项目净现值指标进行评价分析,通过计算2个指标的评价结论截然相反。分析发现矿山作为资源项目具有不可再生性和耗竭性,可采资源是恒定的,导致"有项目""无项目"时计算期不同,出现多个内部收益率,指标无参考意义。通过对项目效益和费用进行统计分析,对矿山原有的剥离、建筑构筑物、总图等工程根据一定的比例按利旧处理,与新增投资一起计入投资费用,按新建项目体系进行经济评价,认为是一种比较好的处理方式,满足评价体系要求。

复杂微细粒低品位硫回收试验研究

针对凡口铅锌矿选硫尾矿中复杂微细粒低品位硫的特点,采用浮选法回收硫,考察了试样的自然可浮性以及捕收剂用量、起泡剂用量、浮选浓度对浮选指标的影响。试验结果表明,经一粗一精一扫和粗硫精矿脱铅锌后,可得到硫精矿品位43.02%,硫回收率77.85%的指标。

高硫低品位硫酸渣“热液碱熔分”试验研究

对云南某高硫低品位硫酸渣进行了“热液碱熔分”试验研究,采用HSC、SEM-EDS、XRD等分析检测结果阐述了提铁除杂的机理,结果表明:原料中Fe品位为51.34%,S质量分数为1.95%,杂质矿物主要由SiO_(2)和Al_(2)O_(3)组成;在NaOH质量分数为30%、熔分温度为200℃、熔分时间为90 min、液固比为2∶1的条件下,获得了Fe品位为60.03%、S质量分数为0.05%的合格铁精矿。基础理论分析及样品特性研究结果表明,高硫低品位硫酸渣经过热液碱熔分后,Fe、O元素明显富集,含S、Si、Al矿物由难溶性向可溶性转变。研究成果可以为类似硫酸渣的二次利用提供理论基础和技术支撑。

云南某低品位硫-氧混合型铜矿石浮选试验

云南某铜矿石属典型的低品位、高氧化率硫-氧混合型铜矿石,含铜0.33%,其中硫化铜占有率为49.09%,氧化铜占有率为50.91%。为确定该矿石的开发利用工艺,进行了选矿试验研究。结果表明,在磨矿细度为-0.074 mm占86.40%的情况下,采用1粗3精1扫流程浮选硫化铜矿物、1粗3精1扫流程浮选氧化铜矿物,可获得铜品位18.58%、回收率77.55%、金品位4.23 g/t的铜精矿。试验指标良好,实现了低品位硫-氧混合型铜矿石中铜、金的高效综合回收,可作为该矿石开发利用工艺设计的依据。

高钙低品位硫氧混合锌矿氧化焙烧矿相的转变规律

针对低品位硫氧混合锌矿含钙高的特点,在低温下对其进行氧化焙烧,通过控制气相中的硫势和氧势,使矿石的矿相发生转变,促进锌在氨性体系中溶解;同时利用矿石中钙的固硫作用,使硫转化成CaSO_4,有效避免SO_2对环境的污染.结果表明,锌品位为5.45%的硫氧混合锌矿在300℃下焙烧3h,矿石中的异极矿和闪锌矿转化成了易溶出的ZnO,方铅矿和黄铁矿分别转化成了PbSO_4和Fe_2O_3,方解石转化成了CaSO_4,所得焙砂在NH_3-(NH_4)_2SO_4-H_2O体系中45℃下浸出,锌浸出率可达73.27%;焙烧温度大于300℃时,焙砂中会产生大量铁酸锌,使锌浸出率下降.

低品位高硫铝土矿反浮选同步脱硫硅试验

重庆某低品位高硫铝土矿石Al2O3含量为59.79%,Si O2含量为13.06%,S含量达1.82%,铝硅比为4.58;矿石中的含铝矿物主要为一水硬铝石,一水软铝石少量;硅酸盐矿物种类较复杂,主要有绿泥石、高岭石、伊利石和石英等;硫主要以黄铁矿的形态存在。矿石中高岭石、伊利石、绿泥石等共生关系密切,且与一水硬铝石的嵌布关系较复杂,黄铁矿晶形较规则、嵌布粒度较粗。为确定该矿石的开发利用工艺,进行了选矿试验研究。结果表明,矿石在磨矿细度为-0.075 mm占77%、矿浆p H=8.5的情况下,以碳酸钠为p H调整剂,无水硫酸铜为活化剂,无机高分子聚合硅酸盐为抑制剂,丁基钠黄药为脱硫捕收剂,松醇油为起泡剂,自主研发的多胺类组合药剂为脱硅捕收剂,采用1粗2扫1精、中矿顺序返回流程处理,可获得Al2O3含量为65.35%、S含量为0.19%、Si O2含量为9.85%、铝硅比为6.63、Al2O3回收率为83.47%的铝精矿,试验指标较理想。

广东某含硫铁低品位铜矿石选矿工艺研究

广东某含硫铁低品位铜矿石主要有用元素铜、硫、铁品位分别为0.51%、27.68%、34.07%。铜赋存状态复杂,以次生硫化铜形式存在的铜占总铜的54.91%,水溶性铜占总铜的26.39%,采用常规浮选方法选别铜回收率低。为探索该矿石中铜、硫、铁的高效分选工艺,对其进行了选冶工艺研究。结果表明:原矿磨细至-0.074 mm占72%时,采用p H=3的硫酸溶液为浸出剂,在液固比为4 m L/g、搅拌转速为1 400 r/min、浸出时间为24 h条件下浸铜,可以获得铜浸出率为93.33%的指标;铜浸渣经自来水搅拌洗涤至p H=6以后,以丁黄药为捕收剂、2号油为起泡剂,经1粗1扫硫浮选,可获得硫品位为48.44%、对铜浸渣回收率为95.57%的高品质硫精矿;浮硫尾矿在磁介质为2mm棒介质、脉动冲程为16 mm、冲次为280次/min、背景磁感应强度为0.6 T条件下,经1次高梯度强磁选选铁,可获得铁品位为51.42%、对铜浸渣回收率为17.02%的铁精矿。以上试验结果说明,采用铜浸出—硫浮选—铁磁选的工艺流程可以实现矿石中铜硫铁的有效分离。

低品位含金硫精矿碱硫氧压浸金试验

对某低品位含金硫精矿开展氧压酸浸产元素硫、碱硫氧压浸取金银试验。酸浸试验最佳条件为:浸出温度130℃、浸出时间300min、浸出压力1.6 MPa、液固比4:1、始酸浓度20g/L、木质素2‰;碱硫氧压浸金试验最佳条件为:浸金剂(石灰+硫磺)为总干矿的40%、硫碱质量比0.35、催化剂Cu(NH_3)_4^(2+)0.06mol/L、活化剂木质素占干矿的0.2%、氧压0.4 MPa、温度105℃、浸金时间5h,在此条件下金、银的浸出率分别为88.23%和61.35%。

低品位含金硫精矿综合回收金选冶试验

介绍了某公司铜硫矿石浮选硫精矿在金品位为1.6 g/t,硫品位为49.68%的条件下,通过焙烧—浸出—碳浆吸附工艺综合回收金,最后获得的金总回收率达到80.57%,同时产生的选金后的尾砂含铁品位为66.45%,含硫量为0.66%,可做为铁精矿直接出售,经济效益显著。

锌浸出渣-低品位氧硫铅锌矿富氧侧吹熔池熔炼渣型理论研究

锌浸出渣-低品位氧硫铅锌矿富氧侧吹熔池熔炼工艺可实现低品位氧硫铅锌矿的高效利用,但目前存在能耗高、生产率低的问题,主要是由于熔渣熔化温度高、流动性差。采用Factsage热力学软件,通过计算研究不同熔渣成分对熔渣熔化温度的影响。结果表明,熔渣的熔化温度随着渣矿质量比的减小而逐渐增大;在同一渣矿质量比下,二元碱度对熔渣熔化温度影响巨大,其随着碱度减小呈现先降低后升高的趋势。当渣矿质量比为3.07、w(CaO)/w(SiO_(2))=0.5时,熔渣熔化温度达到最低值1327℃。

西南某地区低品位高硫铝土矿选矿试验研究

以西南某地区含一水软铝石低品位高硫铝土矿为研究对象,该矿石平均硫含量1.98%,平均A/S 4.32,微细粒级产品A/S略低,粗粒级产品A/S略高;主要有用矿物为一水硬铝石及少量一水软铝石;主要脉石矿物为绿泥石、高岭石及伊利石,在细粒级中绿泥石含量较高。本研究进行了磨矿细度试验、多种预先分级浮选工艺研究、脱硅捕收剂优化研究及选矿闭路试验。开路试验采用原矿预选分级-浮选脱硫脱硅方法,铝精矿产率为62.15%时,精矿A/S达到6.83,尾矿A/S为1.56。闭路试验数据表明,当铝精矿产率为66.92%时,精矿A/S达到6.65;铝精矿和原矿筛分出(-0.023 mm)的综合产率为78.81%,A/S为5.76;尾矿A/S为1.73。

广西某高硫低品位碳酸锰矿选矿试验研究

广西某低品位锰矿石含锰7.34%、含硫6.47%,锰主要以碳酸锰形式存在。针对矿石性质,对锰的回收进行了浮选、跳汰重选、摇床重选、湿式强磁选探索试验,试验结果表明,湿式强磁选是处理该矿合理的工艺流程。通过一次粗选、两次精选、一次扫选强磁选、中矿返回再选的试验流程,获得锰品位17.24%、硫品位4.81%、锰回收率81.11%的磁精矿,将磁精矿磨矿至-74μm占90%,进行一次粗选一次扫选脱硫,最终获得锰品位19.72%、硫品位0.78%、锰回收率80.25%的锰精矿。

某低品位铜硫矿综合回收浮选工艺技术试验研究

针对该低品位铜硫矿的矿物特征,采用抑硫浮铜的优先浮选方案,配合石灰、硫化钠组合抑制剂抑制黄铁矿,成功实现了铜硫的有效分离。该浮选工艺流程结构简单,最终取得了铜精矿品位20.02%,回收率94.33%;硫精矿品位47.33%,回收率79.24%较为理想的试验指标。

某含硫铝土矿石高效、低耗分选工艺研究

为了高效、低耗开发利用广西某含硫低品位铝土矿石,采用阶段磨矿与分级浮选相结合的工艺进行了矿石选矿试验。结果表明,矿石在磨矿细度为-0.074 mm占65%的情况下采用1粗2扫3精、中矿顺序返回流程脱硫,脱硫尾矿中的+0.074 mm粒级1次浮选粗粒铝土矿,粗粒铝土矿浮选尾矿再磨至-0.074 mm占96%的情况下与脱硫浮选尾矿中的-0.074 mm粒级合并1粗2扫3精浮选细粒铝土矿,最终获得S品位为40.54%、Al2O3含量为25.12%、Si O2含量为8.54%、S回收率为81.32%的硫精矿,以及Al2O3含量为65.17%、Si O2含量为8.13%、S含量为0.28%、铝硅比为8.01、Al2O3回收率为79.56%的铝土矿精矿。

从河口铜矿石中回收铜铁硫的选矿试验

云南河口铜矿石含Cu 0.59%、S 4.57%、Fe 26.98%,属伴生硫铁的低品位硫化铜矿石,铜、硫、铁在矿石中分别主要以黄铜矿、黄铁矿、磁铁矿形式存在,但有少部分黄铜矿与黄铁矿形成固熔体。采用铜硫混合浮选—铜硫分离浮选—浮选尾矿弱磁选工艺对该矿石进行综合回收铜、硫、铁的选矿试验,得到了铜品位为18.03%、铜回收率为93.07%的铜精矿,硫品位为52.02%、硫回收率为56.34%的硫精矿和铁品位为61.90%、铁回收率为27.38%的铁精矿,从而为该矿石的合理开发利用提供了技术依据。

某含金银硫精矿焙烧—氰化浸出试验研究

针对云南某低品位含金银硫精矿硫含量高,直接氰化浸出金、银回收率低等问题,进行了焙烧—氰化浸出试验研究。结果表明:硫精矿经过焙烧脱硫后氰化浸出,金、银浸出率得到大幅提高,在焙烧温度600℃,焙烧时间3 h,石灰调p H值至11,搅拌速度1 380 r/min,磨矿细度-0.043 mm占92.48%,氰化钠用量3 kg/t,矿浆液固比3∶1,浸出时间36 h的最佳条件下,金浸出率可达81.41%,银浸出率46.77%,较直接氰化浸出分别提高42.22百分点和26.77百分点。

Co-extraction of valuable metals and kinetics analysis in chlorination process of low-grade nickel-copper sulfide ore

To efficiently co-extract Ni and Cu from low-grade nickel-copper sulfide ore,chlorination roasting with NH;Cl followed by a water leaching process was investigated.The results show that 98.4%Ni and 98.5%Cu can be synchronously extracted when the ore particle size is 75-80μm,the roasting time is 2 h,the mass ratio of NH;Cl to ore is 1.6:1 and the roasting temperature is 550°C.The evolution behavior of various minerals was elucidated using X-ray diffraction(XRD)coupled with scanning electron microscopy(SEM).The kinetics of the chlorination process based on the differential thermal and thermogravimetric analysis(DTA-TG)data was analyzed by Kissinger method and Flynn-Wall-Ozawa(FWO)method.The chlorination process of low-grade nickel-copper sulfide ore mainly contains two stages:the decomposition of NH;Cl and the chlorination of ore.The maximum apparent activation energies(Ea)at two stages are determined to be 114.8 and 144.6 kJ/mol,respectively.The condensed product of exhaust gas is determined to be ammonium chloride,which can be recycled as the reactant again,making the process economic and clean.

Study on hydrometallurgical process and kinetics of manganese extraction from low-grade manganese carbonate ores

In order to utilize low-grade manganese ore resources effectively, a hydrometallurgical process was developed for manganese extraction in dilute sulfuric acid medium, and the kinetics of leaching manga- nese was also investigated. At room temperature, manganese from low-grade manganese carbonate ores was extracted by sulfuric acid leaching without reductants. During the extracting process, single-factor analysis method was used to evaluate the effects of grinding fineness, sulfuric acid concentration, liquid-to-solid ratio, agitation rate and leaching time on the leaching efficiencies of Mn and Fe. The optimal leaching conditions are determined as coarse particles of below 2 mm size （without ball-milling）, sulfuric acid concentration of 0.86 mol/L, liquid-to-solid ratio of 5：1, agitation rate of 150 r/rain and leaching for 180 min at room temperature. Under the optimal conditions, the leaching efficiencies of Mn and Fe are 96.21g and 13.35%, respectively. In addition, through the experiments at different temper- atures, it is found that the leaching process follows the shrinking core model under the conditions of changing acid concentration and intermittent reaction device. Moreover, the apparent activations of effective diffusion and chemical reaction in the kinetic model are calculated to be 18.83 and 27.15 kJ/mol, respectively.

Leaching kinetics of low-grade copper ore with high-alkality gangues in ammonia-ammonium sulphate solution

The leaching kinetics of low-grade copper ore with high-alkality gangues was studied in ammonia-ammonium sulphate solution.The main parameters,such as ammonia and ammonium sulphate concentrations,particle size,solid-to-liquid ratio and reaction temperature,were chosen in the experiments.The results show that the increase of temperature,concentrations of ammonia and ammonium sulphate is propitious to the leaching rate of copper ore.The leaching rate increases with the decrease of particle size and solid-to-liquid ratio.The leaching rate is controlled by the diffusion through the ash layer and the activation energy is determined to be 25.54 kJ/mol.A semi-empirical equation was proposed to describe the leaching kinetics.

Recovery of Zn, Pb, Fe and Si from a low-grade mining ore by sulfidation roasting-beneficiation-leaching processes

To recover Zn, Pb, Fe and Si from a low-grade mining ore in the Lanping basin, Yunnan Province, China, a novel technology using the roasting with pyrite and carbon followed by beneficiation and hydrochloric acid leaching was proposed. Firstly, several factors such as pyrite dosage, roasting temperature, carbon powder dosage, holding time and particle size affecting on the flotation performance of Zn(Pb) and magnetic separation performance of Fe were simultaneously examined and the optimum process parameters were determined. A flotation concentrate, containing 17.46% Zn and 3.93% Pb, was obtained, and the Zn and Pb recoveries were 86.04% and 69.08%, respectively. The obtained flotation tailing was concentrated by a low-intensity magnetic separator. The grade of iron increased from 5.45% to 43.45% and the recovery of iron reached 64.87%. Hydrochloric acid leaching was then carried out for the magnetic separation tailing and a raw quartz concentrate containing 81.05% SiO2 was obtained. To further interpret the sulfidation mechanism of smithsonite, surface morphology and component of the sample before and after reactions were characterized by XRD and EPMA-EDS. The aim was to achieve the comprehensive utilization of the low-grade mining ore.