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.

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。

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）.

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。

云南某贫细硫化铜矿石选矿工艺优化研究

云南某低品位细粒嵌布的铜矿石铜、硫品位分别为0.161%和9.704%,有价铜矿物嵌布粒度细,铜硫分离富集难度较大。为实现该硫化铜矿石资源的高效开发利用,在选厂原选矿流程及原矿工艺矿物学研究基础上,进行了系统的磨浮试验。结果表明:矿石采用磨矿—1粗2扫铜硫混合浮选—混浮粗精矿再磨—1粗3精2扫铜硫分离浮选闭路流程处理,可获得铜品位14.69%、铜回收率63.15%的铜精矿,硫品位29.16%、硫回收率75.10%的硫精矿,试验铜精矿品位较现场提高近3个百分点,铜回收率提高15个百分点以上,显著优于现场生产指标,达到了理想的综合回收铜硫效果。研究结果可作为工艺流程优化的依据。

盐度胁迫下亚硝酸盐型同步脱氮除硫性能急性响应及功能菌群分析

针对生物处理高盐含氮废水过程中亚硝酸盐积累问题,因亚硝酸盐同步脱氮除硫工艺具有高基质耐受性和高基质去除性能,为探究该工艺对盐度(氯化钠的质量分数)胁迫的急性耐受性能,分别考察了盐度梯度和基质质量浓度对亚硝酸盐型同步脱氮除硫工艺性能的单一急性影响。升高基质质量浓度显著提高了基质去除速率(显著性值p<0.05),硫化物和亚硝酸盐中S和N的去除速率最高可达(88.7±0.57) mg·L^(-1)·h^(-1)和(5.51±0.21) mg·L^(-1)·h^(-1);而升高盐度则明显抑制了基质去除速率(p<0.01),硫化物和亚硝酸盐中S和N的去除速率降至(23.33±1.10) mg·L^(-1)·h^(-1)和(1.98±0.09) mg·L^(-1)·h^(-1)。并探究了不同盐度梯度和基质质量浓度联合作用下该工艺性能应激响应。当NaCl质量分数<2.4%时,适当升高基质质量浓度有助于提高硫化物去除速率,NaCl对硫化物去除速率的促进率分别为0.28(180 mg·L^(-1))和0.27(240 mg·L^(-1));而不论基质质量浓度如何,升高盐度皆使亚硝酸盐去除速率显著下降(p<0.01),且抑制率随基质质量浓度升高而增大。经高通量测序技术分析,该工艺优势功能菌属为sulfurovum、rhodobacteraceae、sulfurimonas和microbacterium,其相对丰度分别在28.55%~11.25%。

云南某低品位斑岩型铜矿浮选回收试验研究

云南某大型斑岩型铜矿铜品位0.32%,伴生有钼,具有综合回收价值。采用一次粗选、两次精选、两次扫选工艺流程综合回收矿石中的铜和钼等金属。系统研究工艺参数条件如原矿磨矿细度、粗精矿再磨细度、捕收剂、抑制剂、起泡剂种类及用量等对铜回收的影响,并考察伴生金属钼、金和银的回收情况。结果表明,捕收剂PL307和起泡剂HCLL对铜浮选回收效果好,在原矿磨矿细度-0.075mm占比65%、粗精矿再磨细度-0.045mm占比90%、捕收剂PL-307用量120g/t、抑制剂石灰用量300g/t、起泡剂HCLL用量20g/t的最佳参数条件下,经过闭路试验可获得产率1.23%、铜品位22.84%、铜回收率87.79%的选矿指标,铜精矿产品含钼0.398%、金4.33g/t、银62.50g/,其中钼回收率为71.99%、金回收率为66.57%、银回收率为14.48%,伴生金属钼和金也得到了较好的回收,选别指标好。

Detecting H2S Gas Concentration by 1,8-Naphthalimides Fluorescent Probe

针对传统硫化氢检测方法灵敏度低的问题,以1,8-萘酰亚胺为荧光基团,基于H_2S的还原性,通过在荧光分子结构上引入具有氧化性的硝基,合成一种可与硫化氢气体发生氧化还原反应生成有荧光响应的小分子荧光探针。该探针本身荧光十分微弱,且荧光峰值在λ=467nm和λ=522nm处。与H_2S反应之后,522nm处的荧光效应消失,467nm处的荧光效应显著增强。测定小分子荧光探针在通入H_2S气体前后的荧光光谱,分析467nm处的荧光强度与气体浓度关系。结果表明:荧光光谱法检测出的H_2S气体浓度与荧光强度之间存在很强的线性关系,相关系数为0.979 3,最低可检测极限可达0.88×10^(-6 )mol·L^(-1)量级。表明基于1,8-萘酰亚胺衍生物的荧光光谱检测法可为油气田H_2S气体浓度的的快速测定提供参考。