Apparent activation energy for spontaneous combustion of sulfide concentrates in storage yard

In order to evaluate the spontaneous combustion hazard of sulfide concentrates in storage, three different kinds of sulfide concentrates （sulfur-rich sulfide concentrate, iron sulfide concentrate and copper sulfide concentrate） were obtained from a storage yard in Dongguashan Copper Mine, China. The reaction processes at different heating rates of 5, 10, 15, 20, and 25 ℃/min in air flow from ambient temperature to 1 000 ℃ were studied by TG-DTG-DSC analysis. By the peak temperatures of DTG curves, the whole reaction process for each sample was divided into different stages, and the corresponding apparent activation energies were calculated by the Ozawa-Flynn-Wall method. It is found that the reaction process of each sample is considerably complex; the apparent activation energy values change from 36 to 160 kJ/mol in different temperature ranges; sulfur-rich sulfide and iron sulfide concentrates have lower apparent activation energy than copper sulfide concentrate below 150 ℃; so they are more inclined to cause spontaneous combustion at ambient temperature.

Ultrasonic-assisted Oxidation Leaching of Nickel Sulfide Concentrate

The feasibility of oxidation leaching process of nickel from nickel sulfide ore and the form of different components in the lixivium has been studied at first. The method of leaching nickel sulfide concentration directly by oxidants with existence of ultrasonic has been advanced. The process of leaching nickel from nickel sulfide concentration by using the system of persulfate and silver has been determined. The influence of different factors on the leaching rates of nickel,such as with and without ultrasonic,the concentration of Na2S2O8,liquid-solid ratio and the concentration of AgNO3 have been explored. The results show that:(1) in the oxidative leaching system,nickel can be leached completely as Ni2+ or NiSO4(aq.) from nickel sulfide concentration in theory;(2) the nicopyrite can not exist steadily in the persulfate acid leaching system,but Ni2+can. Meanwhile,sulfur may be existed steadily in the leaching process;(3) nicopyrite with much lower electrostatic potential may be leached earlier than chalcopyrite and sideropyrite;and(4) the overall reaction rate of the leaching process can be enhanced with ultrasonic radiation,but it does not change the mechanism of leaching. The more oxidant concentration or higher liquid-solid ratio is,the higher leaching percentage of nickel. The leaching percentage of nickel can be increased significantly by adding a small amount of AgNO3 during the leaching processes. Under the same conditions,the higher concentration of AgNO3 is,the higher leaching yields of nickel will be obtained.

Pressure acid leaching of zinc sulfide concentrate

Effects of particle size of the zinc sulfide concentrate,leaching temperature,solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied.The results indicate that the additive can improve the reaction kinetics and the conversion rate.And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur.The reasonable experiment parameters are obtained as follows:the leaching temperature 150℃,oxygen partial pressure 1 MPa,additive amount 1%,solid-to-liquid ratio 1:4,leaching time 2 h,initial sulfuric acid concentration 15%,and particle size less than 44μm.Under the optimum conditions,the leaching rate of the zinc can reach 95%and the reduction rate of the sulfur can reach 90%.

APPLICATIN OF ATTRITION GRINDING IN ACID LEACHING OF NICKEL SULFIDE CONCENTRATE

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Sodium sulfide leaching of low-grade jamesonite concentrate in production of sodium pyroantimoniate

Sodium sulfide leaching of a low-grade jamesonite concentrate in the production of sodium pyroantimoniate through the air oxidation process and the influencing factors on the leaching rate of antimony were investigated. In order to decrease the consumption of sodium sulfide and increase the concentration of antimony in the leaching solution, two-stage leaching of jamesonite concentrate and combination leaching of high-grade stibnite concentrate and jamesonite concentrate were used. The experimental results show that the consumptions of sodium sulfide for the two-stage leaching process and the combination leaching process are decreased by 20% and 60% compared to those of one-stage leaching process respectively. The final concentrations of antimony in the leaching solutions of both processes are above 100g/L.

Pressure leaching of bismuth sulfide concentrate containing molybdenum and tungsten in alkaline solution

The leaching results of bismuth sulfide concentrate containing molybdenum and tungsten in air-H2O2-NaOH system, pressure-O2-Na2CO3 system and pressure-O2-NaOH system were investigated. The results show that the extraction of molybdenum, tungsten and sulfur goes up with the increase of NaOH concentration, oxygen partial pressure and reaction time. The extraction of molybdenum and tungsten also rises up with temperature, but the leaching ratio of sulfur increases initially to a peak of 98% at 150℃ and then decreases with the increase of temperature. Under the optimal conditions, the extraction of molybdenum, tungsten and sulfur is more than 95.6%, 93.8% and 96.0%, respectively, and the main phases of residue are Bi2O3 and Fe2O3. Therefore, the method of pressure leaching in alkaline solution is provided as an effective separation of molybdenum, tungsten and sulfur from bismuth and a beneficial pretreatment for consequent process.

Electrogenerative leaching of nickel sulfide concentrate with ferric chloride

In order to utilize the chemical energy in hydrometallurgical process of sulfide minerals reasonably and to simplify the purifying process, the electrogenerative process was applied and a dual cell system was introduced to investigate FeCl3 leaching of nickel sulfide concentrate. Some factors influencing the electrogenerative leaching, such as electrode structure, temperature and solution concentration were studied. The results show that a certain quantity of electrical energy accompanied with the leached products can be acquired in the electrogenerative leaching process. The output current and power increase with the addition of acetylene black to the electrode. Varying the components of electrode just affects the polarization degree of anode. Increasing FeCl3 concentration results in a sharp increase in the output of the leaching cell when c（FeCl3） is less than 0.1 mol/L. The optimum value of NaCl concentration for electrogenerative leaching nickel sulfide concentrate with FeCl3 is 3.0 mol/L. Temperature influences electrogenerative leaching by affecting anodic and cathodic polarization simultaneously. The apparent activation energy is determined to be 34.63 kJ/mol in the range of 298 K to 322 K. The leaching rate of Ni2+ is 29.3% after FeCl3 electrogenerative leaching of nickel sulfide concentrate for 620 min with a filter bag electrode.

Recovery of Lead from Sulfide Concentrate after Mechanochemical Activation Using Nitric Acid

Decomposition of lead sulfide concentrates in nitric acidic solutions after ferric(Ⅲ)nitrate addition and mechanical preparation has been investigated.It was found out that the decomposition can be achieved in HNO_(3) solution(>1.5 mol·L^(-1))at the temperature above 85℃.The leaching rate of lead can be increased by means of mechanochemical activation.The use of[Fe(NO_(3))_(3)-HNO_(3)-H_(2)O]aqueous salt system allows the decomposition of lead sulfide concentrate at room temperature and in this case the concentration of HNO_(3) does not exceed 0.05-0.1 mol.L^(-1),The degree of lead recovery into solution from PbS concentrate reaches 99.3%-99.6c70 with the mechanochemical activation of this concentrate.The insoluble residues after the leaching contain SiO_(2),S^(0),FeCO_(3) and PbSO_(4).

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%.

Recovery of Copper from Copper Sulfide Concentrate by Sulfation Roasting

Sulfation roasting was studied to extract copper from the copper sulfide concentrate. Sodium sulfite was added as sulfation agent to the copper sulfide concentrate during roasting in this study. Sulfur removal rate at different roasting temperatures was investigated, and the effects of roasting temperature, roasting time, and the amount of sodium sulfite on copper extraction were studied. Copper extraction higher than 96% was achieved at optimum roasting conditions.

Separation of sulfide lead-zinc-silver ore under low alkalinity condition

A complex lead-zinc-silver sulfide ore containing 2.98% Pb, 6.49% Zn and 116.32×10^-4 % Ag （mass fraction） from Yunnan Province, China, was subjected to this work. Research on mineral processing was conducted according to the properties of the lead-zinc-silver ore. Under low alkalinity condition, the lead minerals are successfully separated from the zinc minerals with new reagent YZN as zinc depressant, new reagent BPB as lead collector, CuSO4 as zinc activator and ethyl xanthate as zinc collector. The associated silver is mostly concentrated to the lead concentrate. With the process utilized in this work, a lead concentrate of 51.90% Pb with a recovery of 82.34% and a zinc concentrate of 56.96% Zn with a recovery of 81.98% are produced. The silver recovery in the lead concentrate is 80.61%. Interactions of flotation reagents with minerals were investigated, of which the results indicate that the presence of proper amount of Na2S can precipitate Pb^2＋ and has a sulfidation on oxidized lead minerals. The results also show that NazCO3 and YZN used together as combined depressants for sphalerite can signally improve the depressing effect of new reagent YZN on sphalerite.

Separation and enrichment of elemental sulfur and mercury from hydrometallurgical zinc residue using sodium sulfide

The separation and enrichment of mercury and the recovery of elemental sulfur from flotation sulfur concentrate in zinc pressure leaching process were carried out by sodium sulfide leaching and carbon dioxide precipitating. The results show that the leaching rate of elemental sulfur is more than 98%, and 98.13% of mercury is enriched in the residue, under the optimized conditions of sodium sulfide concentration 1.5 mol/L, liquid/solid ratio 6：1 and leaching time 30 min at room temperature. In addition, the content of mercury is enriched 5.23 times that in the leaching residue. The elemental sulfur is precipitated from leaching solution under conditions of carbon dioxide flow rate 200 mL/min and blowing time 150 min, while solution is stirred adequately. The recovery efficiency of elemental sulfur reaches 97.67%, and the purity of elemental sulfur is 99.75%, meeting the requirements of industrial first-rate product standard according to the national standard of GB/T 2449-2006 （PRC）.

Platinum-group Elements Geochemistry of the Yangliuping Magmatic Ni-Cu-PGE Sulfide Deposit:Implications of Its Genetic Link with the Extrusive Basalts

Primitive mantle normalized Platinum group elements (PGE) concentration patterns for the Zhengziyanwo intrusion and Dashibao Formation basalts are of positive slope, similar to most of the world class magmatic Ni Cu PGE sulfide deposits. Characters of this intrusion and its related ores and Dashibao Formation basalts are their negative Pt anomaly and high concentration of Rh relative to Pt and Pd, facts being interpreted to be the results of crystallization and fractionation of Pt alloys and spinel phase free crystallization history for the magma, respectively. PGE parameters of the Dashibao Formation basalts are consistent with the general trend of those found for the Zhengziyanwo intrusion, and this might infer a genetic link between them.

Metallogeny of the Baiyangping Lead-Zinc Polymetallic Ore Concentration Area, Northern Lanping Basin of Yunnan Province, China

The Lanping Basin in the Nujiang-Lancangjiang-Jinshajiang （the Sanjiang） area of northeastern margin of the Tibetan Plateau is an important part of eastern Tethyan metallogenic domain. This basin hosts a number of large unique sediment-hosted Pb-Zn polymetallic deposits or ore districts, such as the Baiyangping ore concentration area which is one of the representative ore district. The Baiyangping ore concentration area can be divided into the east and west ore belts, which were formed in a folded tectogene of the India-Asia continental coUisional setting and was controlled by a large reverse fault. Field observations reveal that the Mesozoic and Cenozoic sedimentary strata were outcropped in the mining area, and that the orebodies are obviously controlled by faults and hosted in sandstone and carbonate rocks. However, the oreforming elements in the east ore belt are mainly Pb-Zn -Sr-Ag, while Pb-Zn-Ag-Cu-Co elements are dominant in the west ore belt. Comparative analysis of the C-O-Sr-S-Pb isotopic compositions suggest that both ore belts had a homogeneous carbon source, and the carbon in hydrothermal calcite is derived from the dissolution of carbonate rock strata; the ore- forming fluids were originated from formation water and precipitate water, which belonged to basin brine fluid system; sulfur was from organic thermal chemical sulfate reduction and biological sulfate reduction; the metal mineralization material was from sedimentary strata and basement, but the difference of the material source of the basement and the strata and the superimposed mineralization of the west ore belt resulted in the difference of metallogenic elements between the eastern and western metallogenic belts. The Pb-Zn mineralization age of both ore belts was contemporary and formed in the same metaliogenetic event. Both thrust formed at the same time and occurred at the Early Oligocene, which is consistent with the age constrained by field geological relationship.

Coupling of anionic wetting agents to dust of sulfide ores by dropping liquid method

By using the experimental approach of dropping liquid, the coupling of three anionic wetting agents with ten dust samples of sulfide ores was studied, and particularly the wetting effects of the wetting agents on the sulfide dust influenced by factors of agent concentration and sulfate additive in the wetting agent solutions were investigated. The results show that when the solution temperature is about 20 ℃, all the selected wetting agents are effective to most dust samples, but the effect is different. Wetting agents are more effective to the dust which is difficult to be wetted. Wetting agent solution with sodium sulfate can improve the wetting ability of sulfide dust. For sodium dodecyl benzene sulfonate, the suitable concentration of sodium sulfate is 12 mmol/L. The cost of wetting agents can be reduced because the sodium sulfate is much cheaper than many surfactants. Since the dust of sulfide ores is composed of various minerals and elements, the whole effect of depressing dust should be considered while innovating a wetting agent.

硫化铜精矿氧压浸出工艺研究

针对硫化铜精矿火法冶炼工艺副产硫酸,在硫酸滞销地区难以大规模应用的难题,研究了硫化铜精矿氧压浸出工艺。对某硫化铜精矿原料进行了成分、物相及有价元素赋存状态进行了分析,从理论上分析了硫化铜精矿氧压浸出的可行性,研究了一段氧压浸出、两段氧压浸出工艺的浸出效果。结果表明,采用两段逆流氧压浸出,一段控制温度135~140℃、总压1.3~1.4 MPa、时间2~2.5 h、分散剂加入量4‰~5‰、沉矾剂加入量2.5%~3%,二段控制温度165~170℃、总压1.4~1.5 MPa、时间2.5~3 h、分散剂加入量6‰~7‰、沉矾剂加入量3%~3.5%,铜浸出率为90.49%~92.25%。一段浸出液含铜90.78~92.34 g/L、铁10.81~13.84 g/L、硫酸28~30 g/L,可送后续除铁净化—电积生产电铜。

硫化锌精矿氧压浸出过程中硫回收率的影响机制

锌氧压浸出特点是硫化锌精矿直接湿法浸出,硫不进入烟气,而是直接转化成单质硫回收。该工艺生产流程更简单,环境友好,综合能耗低,硫磺产品相对更容易储存、运输及销售。因此该技术的推广应用对锌冶炼行业节能减排乃至碳达峰碳中和目标的实现意义重大。从精矿矿物特征及金属的浸出率、浸出过程中硫元素的酸化率、硫磺结晶过程控制、硫浮选效率以及硫精矿热过滤温度等方面着手,研究其对硫磺回收率的影响机制。

硫化钠体系中异极矿表面的硫化特性及硫化产物解析

通过多种分析检测方法揭示硫化钠体系中异极矿表面的硫化特性和机理,并对异极矿表面硫化产物进行分析测试。研究结果表明,硫化钠浓度对异极矿表面的硫化效果具有显著影响,即提高硫化钠浓度能促进矿浆溶液中硫组分与矿物表面反应,从而提高异极矿表面硫化产物的含量及矿物表面的疏水性。XPS和ToF-SIMS分析表明,异极矿表面的硫化产物并非单一的硫化物,而是由硫化物(S^(2−))和多硫化物(S_(n)^(2−))组成,且矿物表面多硫化物含量随硫化钠浓度的增加而增加,从而增强异极矿表面硫化产物的活性。

锌氧压酸浸硫精矿和硫尾矿的矿物学特征与环境活性

采用电感耦合(ICP-OES)、X射线衍射(XRD)、扫面电镜-能谱(SEM-EDS)、红外光谱(FTIR)及差热差重(TG-DTA)等多种检测方法对锌氧压酸浸硫精矿和硫尾矿的化学组成、粒度分布、物相结构等进行研究,并采用毒性浸出程序(TCLP)法考察其短期环境活性。研究结果表明:硫精矿1和硫精矿2主要以球状的单质硫存在,其中,少量硫精矿1以小颗粒的闪锌矿、黄铁矿等附着于块状的铁矾颗粒上,平均粒径为22.1μm;少量硫精矿2以块状的闪锌矿、黄铁矿和小颗粒存在,平均粒径为59.2μm。2种硫精矿在185~340℃时会剧烈燃烧,热稳定性较差。2种硫尾矿主要以块状的铅黄铁矾和黄钠铁矾存在,其中,硫尾矿1平均粒径为21.1μm;硫尾矿2含一定的赤铁矿(16.86%),平均粒径为5.3μm。硫精矿和硫尾矿中锌和镉元素在弱酸条件下不稳定,硫尾矿1中锌和镉的浸出量分别超标1.93倍和2.50倍;硫精矿2的分别超标2.03倍和6.56倍,硫尾矿1的分别超标3.46倍和5.94倍,硫尾矿2的分别超标11.10倍和8.28倍。

从低浓度硫钴精矿浸出液中萃取分离铜钴试验研究

研究了采用P204+磺化煤油体系从除杂后的低浓度硫钴精矿浸出液中萃取分离铜、钴,考察了萃原液pH、萃取剂浓度、萃取相比、萃取时间、反萃取剂浓度、反萃取相比等对铜钴分离效果的影响。结果表明:在萃原液pH=4、P204体积分数25%、萃取相比V_(O)/V_(A)=1/1、萃取时间3min、反萃取剂硫酸浓度0.10mol/L、反萃取相比V_(O)/V_(A)=2/1条件下,对硫钴精矿除杂后液进行两级逆流萃取、四级逆流反萃取试验,萃余液中钴质量浓度为338.80mg/L,铜质量浓度为2.96mg/L,铜萃取率为98.32%,90.03%的钴留在萃余液中,经反萃取铜钴可得到进一步分离,有利于下一步制备高纯度硫酸钴。