某高硫含砷碳低品位难处理金矿选矿试验研究

某低品位金矿石原矿含金1.68 g/t,砷0.43%、碳0.40%、硫3.20%,金以显微或次显微形式浸染于毒砂、黄铁矿、褐铁矿中,具有载金矿物粒度细、砷和碳含量高等特点,是典型的低品位含砷碳极难处理 金矿石,严重影响金的浮选指标。为回收利用矿石中的金,分别进行了直接全泥氰化浸出、重选、浮选三种方案对比试验研究。结果表明,直接全泥氰化浸出率仅5%,重选金精矿回收率不足10%,浮选可获得金品位 15.04 g/t、回收率77.13%的金精矿。由于浮选金精矿含砷、碳、硫有害元素均较高,浮选尾矿含金0.42 g/t,损失较高,因此试验采用焙烧预处理以脱除金精矿和尾矿中的有害元素,然后焙砂氰化浸出回收金。最终 试验采用浮选—金精矿焙烧氰化浸出—尾矿焙烧氰化浸出联合工艺,得到金总回收率70.66%的较好指标,有效地回收了矿石中的金。     

Treatment of pyritic matrix gold–silver refractory ores by ozonization–cyanidation

Most of the gold and silver produced worldwide are extracted by the cyanidation process. The recovery of the precious metals involves two distinct operations: the oxidative dissolution of gold and silver by an alkaline cyanide solution, and the reductive precipitation of metals from the solution. From the cyanidation point of view, gold and silver ores can be classified as free milling, and refractory ores. The term “refractory ore” defines those materials that when submitted to a conventional cyanidation process, show low recoveries (<80%) or high consumption of reactants [Weir, D., Berezowsky, M., 1984. Gold Extraction from Refractory Concentrates, Sherrit Research Centre, Alberta, Canada, pp. 1–26; Haque, K.E., 1992. The Role of Oxygen in Cyanide Leaching of Gold Ore, CIM Bulletin 85963, pp. 31–38]. These refractory ores are usually pretreated by some oxidizing process after which gold and silver can be recovered by standard cyanidation process. Since ozone gas (O₃) is a strong oxidizing, it may be regarded as a promising alternative in the treatment of refractory ores.The present work summarizes the results obtained when two pyritic refractory ores from Mexican sites (samples A and B), were pretreated with an oxygen/ozone stream in acid media before cyanidation. Two contacting methods were studied: the indirect method (contacting the ore three times with ozone saturated water), and the direct method (direct addition of ozone to the mineral slurry). Sample A did not show any difference in recoveries with indirect pretreatment, while the direct pretreatment reduced the cyanidation time for maximal gold and silver recovery from 40 to 24 h. Sample B, only tested with indirect contact method, increased the gold recovery from 53% to 88% and the silver recovery from 26% to 78%.     

某氧化金矿氰化浸出工艺研究-增刊

某金矿为氧化性金矿,金品位3g/t,含砷0.86%,采用全泥氰化工艺处理得到金回收率为72.53%;采用加入常规氧化剂预处理-氰化浸出工艺得到金回收率最高为74.89%;采用碱浸-氰化浸出工艺最佳条件下,金浸出为81.97%。碱浸-氰化浸出工艺可作为该氧化金矿提金处理工艺。     

某难浸浮选金精矿碱式预处理-氰化提金工艺

某浮选金品位65．20g／t，含砷15．40％、硫25．64％。84％以上金被黄铁矿、毒砂和脉石包裹，为难浸金精矿。在常温常压下进行碱式预处理。再接氰化和炭吸附提金。结果表明。在NaCN用量6．5kg／t和炭浆浓度17．5g／L。炭浸24h条件下，金浸出率达93．42％。金吸附回收率达99．67％。     

贵州某难浸金矿原矿焙烧-氰化提金工艺试验研究

针对贵州某难浸金矿的矿石性质,确定采用原矿焙烧—氰化提金工艺。结果表明,金浸出率可由直接氰化的25%提高到90%,为该类难浸金矿提金提供了一条有效的途径。     

新疆难处理金精矿焙烧预氧化—氰化提金工艺试验研究

对新疆某难处理金精矿进行工艺矿物学研究和焙烧预氧化—氰化提金工艺研究。试验结果表明,通过对该金精矿进行焙烧预氧化处理后,氰化金的回收率达91.42%,比常规氰化回收率提高了50.60%。这对于该类型金矿资源的充分利用有着重要意义。     

福建某低品位金铜混合矿石综合回收利用工艺研究

福建某低品位金铜混合矿石含Au 0.36 g/t、Cu 0.29%、Ag 7.4 g/t、S 4.02%,若直接氰化,铜进入金氰化浸出系统,不但得不到回收,还会恶化选金指标,增加生产成本。针对该低品位金铜混合矿,采用浮选+氰化联合工艺进行选别。浮选作业考察了磨矿细度、石灰用量、捕收剂种类、分散剂种类对浮选指标的影响,结果表明,在磨矿细度为-0.074 mm 60%、石灰用量为1500 g/t、Z-200作捕收剂、水玻璃作分散剂时,浮选效果最佳,闭路实验获得铜精矿含Au 16.74 g/t、Cu 20.21%,金、铜回收率分别为61.90%和87.09%。将浮选尾矿进行氰化浸出,考察了氰化钠浓度和氰化时间对金浸出率的影响,结果显示,在氰化钠初始浓度300 mg/L浸出24 h,金浸出率为71.26%。全流程Au回收率达到89.05%,Cu回收率达到87.09%,最终达到综合高效回收矿石中金铜的目的,为此类资源的开发提供了技术支撑。      

富氧氰化法从砷钴硫化矿浸钴渣中提金的研究

从含砷的多金属硫化矿中提金,一向棘手。本文提供了采用富氧氰化法从砷钴硫化矿加压氧化浸钴渣中提金的试验结果,其目的是要开辟一条湿法提金新工艺,解决这一难题。扩大试验结果完全验证了小型试验结果的可靠性。文中详细地叙述了富氧氰化原理,并对酸浸渣(细菌浸钴渣)做了富氧氰化与常规空气氰化提金的对比扩大试验。扩试数据表明,富氧氰化比常规空气氰化提金效果好,指标高,是一种很有前途的提金方法。     

河南某难处理金矿石选冶工艺对比研究

针对河南某难处理金矿石品位低、黄铁矿含量高、部分载金硫化物氧化严重,以及金嵌布粒度极细的特点,开展了详尽的浮选及全泥氰化浸出试验。试验结果表明：采用浮选工艺,所得精矿的金品位和金回收率仅为18.72 g/t和72.55%;而采用全泥氰化浸出工艺,在磨矿细度为-0.074 mm占90%,矿浆液固比为2∶1,加石灰调浆5 h使矿浆pH值稳定在11.5左右,氰化钠用量为1 kg/t,氰化浸出时间为72 h的条件下,金的浸出率可达81.11%。因此,推荐采用全泥氰化浸出工艺处理该矿石。     

微细粒浸染包裹含砷金矿石金的回收

提供了一种微细粒浸染包裹含砷金矿石的选冶联合工艺,包括浮选、碱性常温常压强化碱浸预氧化和氰化。先对含砷金矿石进行浮选,获得含金63.80 g/t、产率5.51%、金回收率92.08%的浮选金精矿,然后对金精矿进行超细磨和碱性常温常压强化碱浸预氧化,氧化渣金的浸出率88.56%,金的选冶总回收率81.55%。     

Leaching and CIL processes in gold recovery from refractory ore with thiourea solutions

In this study, the applicability of leaching and CIL processes in gold recovery with thiourea method, alternative to the cyanidation from the refractory Gümüşhane-Kaletaş/Eastern Black Sea Region (Turkey) ore was investigated.The experiments were conducted at laboratory conditions using ore samples of which approximately 80% were ground to −0.038 mm. The grade of the ore samples was 6.8 g Au/ton. At the first part of the experimental studies, assuming that the gold could be recovered with CIC and CIP processes, the effects of pH, thiourea, oxidizing agent consumption, and leaching time on leaching were investigated. Then, on the basis of the optimum pH and reagent consumption values obtained in the first part (pH = 1.5, 15.2 kg thiourea/ton ore, 140.9 kg iron(III) sulfate/ton ore and 46.2 kg sulfuric acid consumption/ton ore) and adding 50 kg activated carbon/ton ore at the beginning of experiments, the gold leaching extents were obtained for the same leaching times. At this part, the applicability of CIL process in gold recovery with thiourea was investigated for the first time. As a result of the experiments, although higher gold leaching extents were obtained in CIL process, the increase in extent was about maximum 8% and the highest gold leaching extent was obtained as 75% at the end of the 5th hour.     

山东某金矿氰化浸出金的研究

针对矿石风化严重、具有多孔状结构的细粒自然金的特点,采用氰化浸出工艺回收金.用石灰作为保护碱,氰化钠作为浸出剂,通过优化工艺条件,在原矿金品位为4.45g/t时,可获得金浸出率为97.30％的指标.     

老挝某金矿重选-重选尾矿氰化浸金实验

为预先回收老挝某金矿石中的中粗粒金,开展了重选-重选尾矿氰化浸金实验,结果表明,在磨矿细度-0.074 mm粒级占75%、重力值为60G、重选流态化水流量3.6 L/min、给料速度500 g/min条件下,尼尔森重选获得的金精矿品位为15 812.50 g/t,回收率达到21.94%;在磨矿细度-0.074 mm粒级占90%、矿浆浓度40%、CaO用量3 000 g/t、预处理2 h、NaCN用量800 g/t、浸出时间32 h条件下对重选尾矿进行氰化浸金,金浸出率达到74.24%。两种工艺联合最终获得金总回收率96.18%。     

The characterisation and behaviour of carbonaceous material in a refractory gold bearing ore

Sulfide ore and a flotation concentrate from Fosterville contained 0.76% carbonate carbon, 0.18% native carbon and 0.20% organic carbon of which 4.5 ppm were n-alkane hydrocarbons. The concentrate yielded 0.99% native carbon, 0.28% carbonate carbon and 0.11 % organic carbon of which 19.7 ppm was n-alkane hydrocarbons. IR spectroscopy of the fulvic acid fraction of the flotation concentrate showed it to be similar to humic acid.The addition of representative levels of n-alkane hydrocarbons, carbonate carbon and humic acid to the oxidised ore indicated that these components had negligible effect on gold recovery. The addition of 0.2% native carbon decreased gold recoveries from 84.4% to 68.8% while 0.2% activated carbon reduced recovery to 3.1%. Native carbon when acidified, mimicking the environment of bacterial oxidation, resulted in recovery dropping from 84.4% to 61.0%.Gold recovery for a standard sulfide float was 92.5%. Sodium naphthalene sulfonate (SNS) was the only depressant to lower native carbon levels in the concentrates without reducing gold recovery. 275 g/t SNS depressed 45% of the native carbon and gold recovery was unchanged. Nitric acid oxidation and subsequent cyanidation of the SNS concentrates lead to overall gold recovery improving from 88.3% for a standard sulfide float to 92.0%.     

The development of an environmentally friendly leaching process of a high C,As and Sb bearing sulfide gold concentrate

An environmentally friendly leaching process, consisting of the pretreatment of alkaline pressure oxidation and thiosulfate leaching, has been developed to efficiently extract gold from a high carbon, arsenic and antimony bearing sulfide gold concentrate. The Au extraction from the concentrate by direct cyanidation was very low mainly due to the encapsulation of gold by associated minerals and the preg-robbing effect of graphite and organic carbon. The pretreatments of permanganate oxidation and oxidative roasting both could effectively liberate encapsulated gold and eliminate the preg-robbing effect on cyanidation. However, the reagent dosage of permanganate oxidation was high and the final oxidation solution contained substantial quantities of toxic ions. The flue gas of oxidative roasting also contained a lot of poisonous oxides, and the extremely drastic reaction environment of roasting led to the secondary encapsulation of gold by newly generated oxides. The pretreatment of alkaline pressure oxidation effectively liberated encapsulated gold with less than stoichiometric reagent dosage and simultaneously relieved the secondary encapsulation of gold, but could not completely remove graphite and organic carbon. Because carbonaceous matter had a weak affinity for gold thiosulfate complex, the Au extraction by thiosulfate leaching after the pressure oxidation achieved 86.1% whilst the thiosulfate consumption was 35.3 kg/t-concentrate. This process of alkaline pressure oxidation–thiosulfate leaching neither used toxic reagent nor released poisonous gas, and furthermore its effluents contained few toxic ions.     

A review of the selective leaching of gold from oxidised copper-gold ores with ammonia-cyanide and new insights for plant control and operation

The ammonia-cyanide leach system was first patented over 100 years ago and stands out as a unique method of selectively leaching up to 90% Au and <1% Cu from oxidised copper-gold ores using <10% NaCN used in conventional cyanidation processes. However, the system has proved to be difficult to control and predict performance with different ores due to a lack of understanding of the chemistry and mechanism and proper process control. Several laboratory studies have been carried out using various empirical concentrations of ammonia and cyanide with mixed success and only a few commercial operations have been successful.This paper reviews some of the recent applied and fundamental studies on the leaching of copper-gold ores with the ammonia-cyanide system and provides insights into the mechanism to give a better appreciation of the key parameters required for the optimum leaching of gold with minimum copper dissolution. Recommended leach compositions, Eh and pH are provided to enable process control measures to be adopted for a variety of ores. The selective recovery of gold from the leach solutions by cementation or adsorption onto activated carbon or ion-exchange resins is also discussed.     

提高硫酸化焙砂中金和铜浸出率的研究与实践

为了提高硫酸化焙砂中金和铜的浸出率,降低尾渣金品位,减少铜对氰化浸出过程的影响,考察了焙砂粒度、硫酸浓度、温度对硫酸脱铜率和脱铜渣氰化浸金率的影响。结果表明,焙砂(矿粉粒度-0.045 mm粒级占90.16%)在酸度25 g/L、液固比1.5∶1、80 ℃下浸出2 h,硫酸脱铜率达93.62%。脱铜渣在NH₄HCO₃用量10 kg/t、液固比1.5∶1、NaCN浓度0.10%条件下浸出60 h,金浸出率高达98.04%。根据研究结果,通过提高硫酸脱铜温度、硫酸浓度和氰化浸出过程增加旋流器和浸出槽数,采用两段浸出-两段洗涤措施,对现有生产流程进行了优化,铜和金回收率得到了明显提高,获得较好的经济效益。     

某难处理金矿石选冶工艺研究

为了给某难处理金矿石的开发提供技术依据,对其进行了详尽的选冶工艺试验研究。结果表明：采用单一浮选工艺处理该矿石,在-200目占80%的磨矿细度下,可以获得金品位为57.32 g/t、金回收率为84.00%的金精矿;采用浮选-尾矿氰化浸出工艺处理该矿石,可以先在-200目占70%的磨矿细度下获得金品位为60.09 g/t、金回收率为82.26%金精矿,然后在-200目占90%的再磨细度下获得金浸出率为10.70%的浸出液,金的总回收率达92.96%。根据试验结果,推荐采用浮选-尾矿氰化浸出工艺。     

难处理矿石提金工艺研究

介绍了处理秦皇岛某金矿矿石的研究结果。主要进行了混汞及预处理条件试验，提出了处理该矿的最佳工艺为混汞－预处理－氰化。在最佳工艺条件下，混汞可以回收45．1％的金，预处理后进行氰化金的浸出率达94．13％。     

某难浸金矿堆浸尾矿的利用

介绍了西北某难浸金矿堆浸尾矿利用的试验结果,先采用浮选工艺,获得产率４．０９％ 、金品位６４．７２ｇ／ｔ、回收率７０．７７％ 的浮选精矿;浮选精矿经焙烧后氰化浸出,金浸出率达９５．７３％ ;浮选尾矿直接氰化浸出,金的浸出率可再增加１４．１０％ ,从而获得金总收率８１．８５％ 的优异指标。初步技术经济分析结果表明,采用本文介绍的方法利用该类尾矿资源,经济效益较好。