国外某低品位氧化矿浮选-氰化法回收金的试验研究

为提高国外某低品位氧化矿中金的回收率,本文采用浮选-氰化工艺流程：在原矿磨矿细度-74 μm占74.65%,pH=9.0,调整剂氧化钙用量1000g/t、活化剂硫酸铜用量200g/t、捕收剂丁铵黑药60g/t、丁基黄药用量120g/t、起泡剂松醇油用量60g/t时,采用1粗2精2扫闭路试验流程,获得金品位24.30g/t,金回收率72.17%。进一步对浮选尾矿氰化浸出,金浸出率可达92.31%。     

提高青海某金矿浮选回收率试验

为了解决青海某金矿浮选回收率偏低的问题,针对矿石特性,开展了提高金回收率的试验研究,在原药剂制度基础上,采用单因素试验方法和组合用药的技术措施,成功引进新型组合调整剂M11＋NaCO3,取得了较理想的试验指标。同时在选矿厂应用新的试验成果,生产实践表明,金精矿品位大于30 g/t,浮选回收率由72.61%提高到81.55%,企业经济效益显著。     

某含金矿石提高金回收率浮选试验研究

某含金矿石中可供回收的有价元素金含量2.02 g/t,为提高该矿石中金的回收率并获得较佳的技术指标,同时解决金精矿含砷、碳等问题,在一系列矿石性质研究的基础上,对该矿石进行了工艺矿物学研究及浮选工艺试验研究。试验研究结果表明:在磨矿细度-0.074 mm65%、pH值调整剂氧化钙用量t、活化剂硫化钠+硫酸铜用量(1 500+400)g/t、捕收剂Y-89用量120 g/t、起泡剂2~#500 g/油用量45 g/t的条件下,采2粗2精3扫浮选工艺流程,最终得到了金精矿品位33.70 g/t、金回收率78.01%的试验指标;同时为综合开发利用该类金矿石提供了可靠的技术依据。     

福建川石金矿浮选试验研究及生产实践

福建川石金矿选矿厂前期采用常规浮选工艺生产金精矿,浮选采用一次粗选、三次精选、四次扫选工艺流程,丁基铵黑药和丁基黄药作为捕收剂、松醇油为起泡剂、硫酸铜为活化剂、石灰用作p H值调整剂,选矿厂尾矿中金品位为0.82g/t,金选矿回收率为61.87%。经过工艺流程的优化,调整为一次粗选、两次精选、三次扫选,捕收剂丁基铵黑药30 g/t,丁基黄药100 g/t,起泡剂松醇油30 g/t、SC 15 g/t。最终选矿厂尾矿中金品位降至0.28 g/t,金的回收率为87.32%,年增加经济效益可达1 259.42万元。     

某难选低品位金矿的选矿试验研究

对某难选低品位金矿进行的研究结果表明,载金黄铁矿的嵌布粒度微细和易浮脉石矿泥是影响金精矿金品位和回收率的两个重要因素,在强化载金黄铁矿捕收的同时,选择合适的调整剂可以显著提高金的选矿指标。原矿金品位为1.69 g/t时,实验室闭路试验可获得金回收率69.85%、金品位20.47 g/t的金精矿。     

从某氰化尾渣中回收金的研究

针对辽宁某金矿氰化尾渣泥化严重且其中载金矿物被强烈抑制的问题, 进行了脱泥试验和浮选药剂条件试验, 最终提出磨矿-矿泥分散和抑制-硫化矿活化浮选工艺流程, 采用碳酸钠和水玻璃为组合调整剂, 戊黄药和丁铵黑药为组合捕收剂, 硫酸铜为活化剂, 金回收效果明显, 金精矿中金品位从4.87 g/t提升至26.03 g/t, 回收率为65.12%。     

新疆某黄铁矿型低品位金矿石浮选试验

新疆某低品位黄铁矿型金矿石中,脉石矿物蚀变严重,矿泥含量较高,对载金矿物的上浮影响较大。通过强化对矿泥的分散和脉石矿物的抑制,以及对载金矿物的活化和捕收研究,确定了碳酸钠+水玻璃作调整剂、硫酸铜作活化剂、异戊基黄药+丁铵黑药作组合捕收剂、RB-3作起泡剂的药剂制度,采用现场2粗2精3扫流程处理金、银品位分别为1.49 g/t和6.11 g/t的试样,可获得金品位为20.07 g/t、金回收率为67.49%、银品位为42.45 g/t、银回收率为34.81%的金精矿。     

提高新疆某地金矿浮选回收率试验研究

通过大量的试验研究,发现浮选中采用合理的调整剂及分段加药方式可使某金矿金的回收率提高近3%,且金精矿的金品位提高10g/t以上。此工艺为现厂下一步技改提供了可靠的技术依据。     

低品位含金硫化矿选矿试验研究

根据矿石性质,采用混合浮选工艺流程处理某低品位含金硫化矿,在原矿金品位1.03 g/t的条件下,通过添加新型抑制剂DPS强化易浮细粒脉石的上浮和加强含金矿物的活化,大幅度提高了选矿试验指标,金品位由15.25 g/t提高到17.17 g/t,金回收率由91.96%提高到94.55%。     

甘肃省礼县某金矿选矿试验研究

甘肃省礼县某金矿石含金5.67 g/t、银3.1 g/t、含硫2.67%,矿石性质研究结果表明,矿石中金矿物嵌布粒度细,裸露半裸露金可浮性较差是影响金回收的主要原因。试验通过对金矿物活化剂和捕收起泡剂研究,确定了碳酸钠作调整剂、T31作活化剂,丁基黄药作捕收剂,A25作捕收起泡剂的药剂制度,小型试验和工业试验均取得了良好的技术指标。     

Evaluation of leaching parameters for a refractory gold ore containing aurostibite and antimony minerals: Part I – Central zone

A cyanidation study was conducted on a mild refractory gold ore sample from the Central zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. Gold, at a grade of 8.00 g/t, is present as native gold, electrum and aurostibite. The ore also contains 2.8% pyrrhotite, together with several antimony minerals (0.8% berthierite and gudmundite, 0.18% native antimony and stibnite). It also exhibits weak preg-robbing properties with 0.16% organic carbon. Aurostibite, a gold antimony compound, is particularly known to be insoluble in cyanide solution. The antimony dissolves in cyanide solution to form antimonates, which retards gold dissolution. Industrial practice of extracting gold from aurostibite generally consists of producing a flotation concentrate, which is leached in a pipe reactor at low alkalinity and high oxygen pressure with about 20 g/L cyanide.The proposed new approach is efficient and allows the extraction of gold directly from an ore at atmospheric pressure and a low cyanide concentration at pH 10.5. The effects of grinding, pre-treatment, lead nitrate, kerosene and cyanide concentrations have been investigated. The maximum gold extraction obtained on the ore was 87.9% using 800 ppm NaCN, 500 g/t lead nitrate, 30 g/t kerosene, DO (dissolved oxygen) 10 ppm and pH 10.5 in 168 h. The associated cyanide consumption was 1.3 kg/t. The additions of lead nitrate and kerosene increased gold extraction. In comparison to a P₈₀ of 74 μm, a P₈₀ of 30 μm significantly increased gold extraction. Gold in solid solution in gudmundite and arsenopyrite was believed to be responsible for the un-leached fraction until mineralogical analysis of hydroseparation concentrates of leach residues showed that most of the un-leached gold occurs as aurostibite, either as locked grains in sulphides/sulpharsenides or as grains with passivation rims of an Au–Sb–O phase. Coarse gold was also found. Gold extraction was not sensitive to cyanide concentration from 250 to 1200 ppm NaCN and high pH was detrimental. Decreasing the cyanide concentration reduced the cyanide consumption from 1.39 to 0.85 kg/t. The removal of coarse gold using a Knelson concentrator and a Mosley table prior to leaching increased the gold extraction to 90.4% (leach residue at 0.77 g/t).     

江西金尾矿资源的性质与矿物学特征研究

黄金矿山选矿中95%以上为尾矿,实现金尾矿综合利用的前提是充分查明其性质与特征。针对江西某金尾矿,采用化学分析、ICP和γ能谱仪,查明了金尾矿的化学成分和放射性核素限量,利用岩矿鉴定、EMPA和EDS分析等手段,研究了该物料的资源性质与矿物学特征,分析了其综合利用前景。结果表明：该金尾矿中含金0.495 g/t,含银1.70 g/t,且金以铁等氧化矿物中包裹金存在,占比58.18%,显微镜下未发现残余的金颗粒。金尾矿中SiO₂、Al₂O₃分别为62.21%和13.54%,残余的矿物主要为石英、绢云母和方解石,其中绢云母具有较好的回收前景。放射性核素限量合格,若制备建材制品产销与使用不受限制。该试样筛析-0.074 mm占41.58%,残余含金矿物的单体解离度不高,含金矿物多赋存在比重较高的重矿物或金属矿物中,仍有部分提金潜力。绢云母多为细粒级产出,片径为0.01～0.2 mm,提取时需加强分级和细粒易泥化物的分散,防止影响其品质。     

Enhanced leachability of gold and silver in cyanide media: Effect of alkaline pre-treatment of jarosite minerals

Composite samples of tailings containing gold (1.35 g/t) and significant amounts of silver (155 g/t) were subjected to batchwise cyanide leaching to assess the feasibility of extracting gold and silver. The tailings are waste solids arising from flotation and leaching operations whereby the flotation product (sphalerite concentrate) is calcined and then solubilised into dilute sulphuric acid solution and eventually sequestered from the electrolyte by electrowinning. Silver and gold are part of the zinc refinery residue, flotation tailings and to a limited extent the calcine leach tailings. Mineralogical results showed that composite tailings are refractory in nature (44% quartz, 17% silico aluminates and 12% jarosites).The concept of enhancing gold and silver recovery from the tailings focused on firstly decomposing the jarosite minerals by alkaline pre-treatment and then secondly leaching with cyanide solution. These two steps ensured that free gold and silver found in the zinc refinery residue and in the jarosite minerals could be leached simultaneously. The composite tailings were treated with Ca(OH)₂ solutions and then heated to 90 °C for 2 h to decompose the silver-bearing mineral (Ag,PbFe₃(SO₄)₂(OH)₆). The alkaline pre-treated tailings were then subjected to cyanide leach tests at different NaCN dosages (2.5–10 kg/t) and particle size (96–200 μm). Without an alkaline pre-treatment stage, leach efficiencies achieved were 41% and 25% for gold and silver, respectively at 40 °C and 8 h mixing time. But, better leach efficiencies (55% for Au, 81% for Ag) were achieved after the feed was pre-treated with Ca(OH)₂. The leaching mechanism of gold was explained by the shrinking sphere model denoted by surface chemical reaction.     

河北承德某金矿选矿试验研究

河北承德某含黄铁矿石英脉型金矿石,金品位为10.4 g/t,黄铁矿是主要的载金矿物,金的嵌布粒度大小不等.经过探索试验确定该矿石适宜采用单一浮选工艺流程,选矿试验结果表明:在磨矿细度-0.074 mm含量占59.44％条件下,以碳酸钠为pH值调整剂、丁基黄药为捕收剂,经一次粗选、一次精选、两次扫选,可获得精矿金品位17...     

老挝某金尾矿浮选—氧化浸出试验研究

老挝某金矿浮选尾矿中金品位为2.99 g/t,含金量较高,为高效回收该尾矿中的金,开展了详细的选矿试验研究。尾矿中金主要赋存于硫化物中,其次是连生体金和单体金,决定采取浮选—氧化浸出联合选别流程。采用石灰作为pH调整剂,以六偏磷酸钠和水玻璃为脉石矿物抑制剂,以CuSO₄和Pb(NO₃)₂为活化剂,以丁铵黑药和丁基黄药作为捕收剂,在条件试验的基础上对金矿浮选尾矿通过“一粗三精三扫”的闭路流程试验,获得了金品位为28.57 g/t、回收率为67.36%的金精矿;同时,所得尾矿金品位为1.05 g/t。再以高锰酸钾作为氧化剂,以氰化钠作为浸出剂对该尾矿进行氧化浸出,最佳条件下金的浸出率为43.57%,相对于原矿浸出率为14.22%,且浸渣中金品位仅为0.6 g/t,金总回收率为81.58%,取得了较为满意的试验指标。      

The influence of pyrite pre-oxidation on gold recovery by cyanidation

The influence of pyrite pre-oxidation in alkaline solutions on gold recovery by cyanidation from Twin Creek refractory gold ore in which pyrite was identified as the major sulfide mineral has been investigated with the aid of electrochemical measurements, leaching experiments, and direct analysis of reaction products for selected residues. It was found that gold recovery by cyanidation in bottle roll experiments mainly depended on the extent of pyrite pre-oxidation. The rate of pyrite oxidation in alkaline solutions measured by electrochemical measurements, including chronoamperometry and linear sweep voltammetry, increased with an increase in pH, potential, and temperature. All alkaline reagents used for the electrochemical measurements, NaOH, NH₄OH, Na₂CO₃ and Ca(OH)₂, showed a similar effect on pyrite oxidation kinetics. However, the results of alkaline pre-oxidation for pyrite of the Twin Creek refractory gold ore suggested that NaOH and Na₂CO₃/Ca(OH)₂ were superior to Ca(OH)₂. Without pre-oxidation, cyanide leachable gold was found to be only 20% which could be increased to 70% under appropriate pre-oxidation conditions. At the same time, cyanide consumption decreased from 2.5 kg/t ore to 1.5 kg/t ore.Selected residues after pre-oxidation and cyanidation were examined by X-ray diffraction. Backscattered electron images of pyrite particles in these residues were taken. The reaction products at the surface of pyrite particles were found to be iron-, silicon-, and calcium-bearing compounds with variable amounts of sulfur as determined by X-ray energy dispersion analysis. Additionally, some mineral fines, such as aluminum and/or potassium-bearing minerals, were found to be present at the partially oxidized pyrite surface.     

云南某低品位氧化铅锌矿选冶联合新工艺研究

针对云南某低品位氧化锌矿的性质特点,提出了一种选冶联合工艺,即脱泥—浮选预富集脱钙—湿法炼锌的流程。在磨矿细度-0.074 mm含量为85%,Na_2S用量为8 000 g/t、组合抑制剂用量六偏磷酸钠(100+100)g/t+水玻璃(500+500)g/t,HHA为捕收剂的条件下,经脱泥作业后两次粗选,最终获得了氧化锌粗精矿品位和回收率分别为18.09%和77.57%的良好指标,且粗精矿中的氧化钙含量由原矿中的21.71%降到了13.25%,高达63.50%的氧化钙杂质得到脱除。通过该工艺流程,实现了保锌脱钙的目标。全流程工艺总锌回收率为90.27%,每吨锌综合耗酸量仅为3.60 t,大幅度降低了湿法炼锌的耗酸量。     

Bulk flotation of auriferous pyrite and arsenopyrite by using tertiary dodecyl mercaptan as collector in weak alkaline pulp

Laboratory and industrial scale experiments were conducted to investigate the effect of tertiary dodecyl mercaptan (TDM) as a collector on the flotation of auriferous pyrite and arsenopyrite. The optimum recovery of gold associated with auriferous sulphides was obtained by adding a mixture of TDM and sodium butyl xanthate, together with only a little CuSO₄ as an activator in a weak alkaline pulp adjusted by NaOH. A two-month industrial trial at the Liumei plant in Guangxi, China showed that an average gold recovery of 90.8% into a concentrate assaying 81.1 g/t Au from a feed assaying 2.9 g/t Au could be achieved at pH 8–8.5 using TDM as a collector.