应用电感耦合等离子体发射光谱法测定岩矿中的金

用电感耦合等离子体发射光谱法测定锑金精矿中金量解决了样品在烧样除炭过程中因锑的熔解而结块、在溶解实验过程中因大量锑的水解包裹使金的溶解变得较为困难的问题。采用电感耦合等离子体发射光谱法测定金，克服了原子吸收法测定的线性范围小，而稀释倍率大，造成分析误差偏大的不足。方法的检出限为0．1μg／g。金加标回收率96％～104％，精密度（RSD，n=6）〈3．0％。适宜于锑金精矿中金的测定。     

辽宁某高硫钼矿选矿试验研究

针对辽宁某高硫钼矿进行了选矿试验研究。在磨矿细度为-200N占67％时,经两粗两扫获得钼粗精矿及浮钼尾矿,对钼粗精矿进行再磨再选后,获得钼品位49．43％,钼回收率91．11％的钼精矿。同时对浮钼尾矿中的硫进行浮选回收,采用一粗、一扫、两精的浮选流程,获得了品位为52％,回收率为72．04％的硫精矿。     

氢化物-原子荧光光谱法依次测定锌精矿中砷、锑、铋、锡

用氢化物～原子荧光光谱法，一次性分解样品，在L2半胱氨酸存在下，实现了锌精矿中As、sb、Bi、Sn的连续测定。其回收率分别为93．8％～104．0％、95．0％～108．3％、93．0％～110．6％和95．7％～104．7％；检出限分别为0．35、0．31、0．19和0．35μg／L。用该方法分析锌精矿样品，结果满意。     

西藏甲玛铜钼矿石选矿试验研究

针对西藏甲玛铜钼矿石进行选矿工艺试验研究。该矿石中主要目的矿物为黄铜矿、铜的次生矿物及辉钼矿。通过铜钼混合浮选-铜钼混合精矿再磨后铜钼分离的选别工艺流程及适宜的药剂制度,取得了较好的选矿指标。小型闭路试验获得的指标为：钼精矿品位47.71%,钼回收率73.20%；铜精矿品位32.11%,铜回收率94.22%；铜精矿含金16.65 g/t,含银351.7 g/t,金回收率47.88%,银回收率67.30%。     

黑龙江省双鸭山市某石墨矿选矿工艺流程试验研究

原矿固定碳含量6.06%,为晶质鳞片石墨,+100目鳞片石墨固定碳含量占总固定碳含量的48.22%。通过对原矿进行化学分析、矿物组成及特征研究,采用一次粗磨快速浮选、粗精矿经1次精选、再经3次再磨4次精选后筛出大鳞片石墨,筛下精矿产品继续精选,得到大鳞片含量高,精矿品位高的浮选精矿产品。最终精矿石墨总回收率为94.09%,其中+100目精矿固定碳回收率34.18%;大鳞片石墨保护率达到70.89%;-100目精矿固定碳回收率59.91%。     

无捕收剂浮选提高金矿回收率试验研究

为建设环保矿山并提高金矿回峰率,进行了无捕收剂浮选的试验研究。对磨矿细度、调整剂用量、硫化钠用量、起泡剂用量等条件进行了研究。结果表明,采用无捕收剂浮选工艺,在原矿金品位16．75g／t,磨矿细度细度-0．074mm占60％、硫化钠用量600克／吨,2#油用量234．8g／t、浮选时间24min的情况下,可获得金精矿平均品位45．96g／t、回收率96．33％的技术指标。比选矿厂原生产指标（92％左右）提高了4％以上的回收率。     

碘量法测定锌精矿中的砷

试样以硝酸、氯酸钾分解、硫酸冒烟,再以碘量法测定锌精矿中的砷含量。方法准确、简便,加标回收率为98．20％-101．6％,RSD〈2．0％,方法测定范围0．20～3．00％。     

碱性模式氢化物发生-原子荧光光谱法测定多金属矿中的微量铋

建立了碱性模式氢化物发生-原子荧光光谱法测定多金属矿石中微量铋的分析方法，对仪器和实验条件进行了最佳化选择，讨论了共存离子的影响。方法经国家标准物质分析，测定值与标准值吻合，方法精密度（RSD，n=12）为2．97％，回收率为95％～104％。     

吉林省靖宇县二道阳岔铁矿矿石可选性分析及探究

总体来看,原矿矿石成分比较单一,选矿试验证明矿石为易选矿石,采用磁选法进行选别,单一磁选工艺流程,经过一次磨矿、一次初选,粗精矿再磨后再次精选,获得的精矿品位TFe67.20%,回收率72.66%。除磁铁矿及少量的磁黄铁矿外,其他非磁性矿物均进入尾矿,其他含铁矿物分布局限,质量分数低,且不易利用,不需另行选矿。     

中关铁矿钴回收试验研究

通过试验研究,确定对磁选尾矿中Co金属作为副产物回收的可能性,并确定合适的浮选工艺,在不影响原有磁选工艺的前提下,加入适当的调节剂CaO、起泡剂2#油、捕收剂丁基黄药,浮选钴精矿品味即达到0.341%,回收率71.06%,经济效果显著。     

Sulphidic refractory gold ore pre-treatment by selective and bulk flotation methods

In sulphidic refractory gold ores, fine gold particles may be highly disseminated and locked up in sulphide minerals such as pyrite and arsenopyrite. Hence, pre-treating processes are necessary to extract gold for further processing methods like cyanidation. In this paper, the selective and bulk flotation methods are considered as primary sulphidic refractory gold ore pre-treating options. The results show that it is impractical to remove arsenic (As) and antimony (Sb) from downstream processes by using selective flotation. It was observed that 22% of total Sb and 31.1% of As remained in the final tailings. In this case, 24% of gold was recovered in the initial stage of flotation. On the contrary, in the bulk method of flotation, which was aimed at maximizing the gold (Au) recovery, the maximum Au recovery of 90.6% obtained after 60 min of flotation at the grind size with d₈₀ of 146 μm. The bulk flotation method as a pre-treating method resulted in the concentrate with less toxic chemical elements such as As and Sb.     

Optimal Dearsenification Parameters of Gold Sulfide Concentrate with a High As Content

The gold sulfide concentrate with a high As content in Liangshan District, Sichuan Province, China, is a potentially important resource. This paper describes experiments of dearsenification of gold concentrate in a weakly reduced atmosphere in a rotary pipe furnace. The results showed that the optimal parameters were a temperature range of 650-700℃, 15%-16% CO2 of gas and a reaction time of 30-40 min. The removal rate of arsenic and sulfur was over 95% and 25%-28%, respectively. With further oxidization and roasting, residue sulfur in the roasted materials was dropped to below 4%, and the cyanide leaching recovery of gold was over 92%.     

The mechanism of suspension reduction on Fe enrichment with low-grade carbonate-containing iron ore

Carbonate-containing iron ore is a refractory iron resource with a complex symbiotic relationship between iron minerals and gangue. An efficient suspension magnetizing roasting technology was proposed to extract the iron mineral from the carbonate-containing iron ore. The results demonstrated that a concentrate containing 67.20% total Fe and 83.35% iron recovery was obtained under the optimal conditions: the roasting temperature at 550 °C for 15 min with a CO concentration of 15 at.%, a total gas flow of 400 mL/min, and the grinding size of the roasted product was ?0.045 mm accounting for 85.39 wt%. The phase transformation and microstructure evolution of the samples before and after roasting were characterized. The investigation revealed that the iron mineral phases were selectively transformed from weakly magnetic hematite and siderite to strongly magnetic magnetite, disrupting the particle structure. Subsequently, the iron mineral was effectively enriched into concentrate via magnetic separation.     

Improvement of the leach efficiency of carbonaceous gold concentrates using reduction roasting pretreatment technology

Carbonaceous gold concentrates were pretreated using a reduction roasting method. The roasting process was investigated by thermogravimetry-mass spectrometry and X-ray diffraction. The characteristics of the thermal reaction process and gas release, and the removal mechanism for pyrite and organic carbon were analyzed. Results showed that the thermal decomposition and reduction reactions caused the highest weight loss. The main gold-locked mineral of pyrite was eliminated by decomposition and reduction reaction to produce FeS and S₂. The “preg-robbing” ability of organic carbon was eliminated by thermal decomposition to converted into graphite. After reduction roasting, exposed gold content increased from 34.65% to 83.25%, organic carbon removal rate was 98.38%, and the gold leaching rate increased significantly from 2.43% to 73.7%. The technology solved the gold locking and “preg-robbing” problems of organic carbon, prevented the production of SO₂, which realized a highly efficient and clean pretreatment for carbonaceous gold concentrates.     

Gold nanolenses generated by laser ablation-efficient enhancing structure for surface enhanced Raman scattering analytics and sensing

Nanoaggregates formed by metal spheres of different radii and interparticle distances represent finite, deterministic, self-similar systems that efficiently concentrate optical fields and act as "nanolenses". Here we verify experimentally the theoretical concept of nanolenses and explore their potential as enhancing nanostructures in surface enhanced Raman scattering (SERS). Self-similar structures formed by gold nanospheres of different sizes are generated by laser ablation from solid gold into water. These nanolenses exhibit SERS enhancement factors on the order of 10(9). The "chemically clean" preparation process provides several advantages over chemically prepared nanoaggregates and makes the stable and biocompatible gold nanolenses potent enhancing structures for various analytical and sensing applications.     

Metal induced crystallization: Gold versus aluminium

In this work metal induced crystallization was studied using aluminium and gold deposited over 150 nm amorphous silicon films grown by LPCVD. Aluminium and gold layers with thickness between 1 and 5 nm were deposited on the silicon films and after that, the samples were annealed at 500^∘C from 5 up to 30 h. When the crystallization is induced through a gold layer, the Si crystalline fraction is higher than when using aluminium. For samples crystallized for 30 h at 500^∘C with 2 nm of metal a crystalline fraction of 57.5% was achieved using gold and only 38.7% when using aluminium.     

Application of three xanthates collectors on the recovery of nickel and pentlandite in a low-grade nickel sulfide ore using optimum flotation parameters

One of the challenges encountered when processing nickel ore is the presence of other minerals like chalcopyrite, pyrrhotite, pyrite and amphiboles in mineral dressing. Hence, there is need to investigate the selectivity of some collectors at optimum flotation parameters (particle size, milling time, pH and flotation cell) to maximize nickel and pendlandite recovery. Three xanthate collectors, namely potassium amyl xanthate (PAX), sodium isobutyl xanthate (SIBX) and sodium ethyl xanthate (SEX), were investigated in this study. In order to achieve this aim, the physical treatment of nickel sulfide ore and percentage chemical composition of nickel sulfide ore were investigated. The analyses of the concentrates were done using x-ray diffraction (XRD) and x-ray fluorescence spectroscopy (XRF). The optimum nickel and pentlandite recovery was achieved at particle size of 64.73 µm milled for 1.5 h, pH of 9.65 and using a new flotation cell. Also, SEX collector recovered more nickel and pentlandite contents compared to PAX and SIBX collectors when optimum parameters of particle size, milling time, pH and flotation cell were considered. In conclusion, the results showed that the selectivity of SEX is the strongest and will recover more pentlandite in a complex and low-grade nickel sulfide ore.     

Stochastic production phase design for an open pit mining complex with multiple processing streams

In a mining complex, the mine is a source of supply of valuable material (ore) to a number of processes that convert the raw ore to a saleable product or a metal concentrate for production of the refined metal. In this context, expected variation in metal content throughout the extent of the orebody defines the inherent uncertainty in the supply of ore, which impacts the subsequent ore and metal production targets. Traditional optimization methods for designing production phases and ultimate pit limit of an open pit mine not only ignore the uncertainty in metal content, but, in addition, commonly assume that the mine delivers ore to a single processing facility. A stochastic network flow approach is proposed that jointly integrates uncertainty in supply of ore and multiple ore destinations into the development of production phase design and ultimate pit limit. An application at a copper mine demonstrates the intricacies of the new approach. The case study shows a 14% higher discounted cash flow when compared to the traditional approach.     

Effects of type and dosages of organic depressants on pyrite floatability in microflotation system

Pyrite is considered as a disturbing sulfide mineral in the beneficiation of sulfide minerals and coal and in order to prevent floating of pyrite in the flotation, different inorganic, organic and biological depressants could be used. The most important organic pyrite depressants are starch, dextrin, guargum, carboxymethyl cellulose (CMC), ethylene diamine tetra acetic acid (EDTA), Diethylenetriamine (DETA). In this paper, flotation experiments were carried out by modified Halimond tube. At first, flotation experiments were performed to obtain optimal pyrite flotation condition (suitable collector and frother for maximum recovery). In optimal condition (collector: KAX?+?KIPX , 16?ppm; collector condition time: 4?min; frother: Poly Propylene Glycol (A65), 20?ppm; frother condition time: 30?s), pyrite recovery increased to 95.63%. Flotation kinetics experiments were performed and various kinetic models were evaluated on pyrite flotation. The results showed that flotation kinetics of pyrite is second-order. After obtaining optimal condition for flotation and kinetic studies, the effects of type and dosage of different organic depressants including starch, dextrin, guargum, carboxymethyl cellulose, EDTA and DETA at three concentrations (50, 100, 150?ppm) at neutral pH were comprehensively investigated. The results of the experiments showed that DETA can significantly reduce the recovery of pyrite at 150?ppm by 44%.