磷砷渣返回压煮处理的工艺研究

分析了在黑钨精矿碱压煮条件下 ,磷砷渣中钨和污染因子砷的行为。通过试验 ,摸索了磷砷渣返回压煮处理的工艺条件。在过碱系数 1.42 ,加入氧化剂等添加剂的情况下 ,适量的磷砷渣与精矿一起压煮 ,精矿分解率达 98.8% ,溶液质量满足后续工序要求 ,同时实现了有害磷砷渣向无害钨渣的转化。     

银精矿碱法熔炼工艺的扩大试验

用正交试验法研究了熔炼温度、配煤量、配碱量及银精矿与铅精矿的质量比等因素对银精矿碱法熔炼中银、铅直收率的影响,建立了相应的回归模型;确定了苏打钾碱熔炼工艺的最佳工艺条件为：熔炼温度１０００～１０５０℃,停留时间１～２ｈ,配煤率４％～６％,钾碱率５％～７．５％及混矿比１∶１。此时,可获得含铅≥９８％,含银约１％的粗铅,铅、银的直收率分别大于９６％和９２％。     

MACA体系中循环浸出低品位氧化锌矿制备电解锌

以云南兰坪低品位氧化锌矿及其循环浸出渣的浮选精矿为原料,常温常压下在MACA(金属氨络合物)体系中进行循环浸出。浸出液先净化除砷和锑、再两段锌粉逆流置换深度净化,两次净化后液电积制取电解锌。考察工艺中循环浸出率、净化率、物质平衡以及电解锌质量和电耗等技术经济指标。结果表明:先用MACA法处理原矿粉,再浮选硫化锌的选冶结合流程是合适的兰坪低品位氧化锌矿的处理方案,原矿锌的平均浸出率为70.48%,其氨可溶锌浸出率达到89.14%,浮选精矿锌的浸出率为79.75%,杂质元素的净化率达到95%,电解锌纯度达到99.98%,电流效率可达97.02%。     

煤系硫铁矿浮选尾矿热化学活化脱硅制备铝精矿

提出热活化脱硅技术处理某煤系硫铁矿浮选尾矿制备铝精矿,对制备氧化铝精矿的工艺制度及脱硅机理进行研究。结果表明:该尾矿适宜的热化学活化脱硅制度为活化焙烧温度1 150℃、焙烧时间15～20 min、碱浸溶硅温度125～140℃、溶出时间30 min、NaOH浓度140 g/L。在此条件下,对Al2O3和SiO2含量分别为46.22%和28.33%(质量分数)的硫铁矿浮选尾矿,焙砂SiO2溶出率达到71.91%,所得铝精矿中Al2O3含量达69.29%,铝硅比5.59。XRD结果表明:硫铁矿尾矿中伊利蒙脱石、高岭石和叶腊石等铝硅酸矿物在焙烧过程中活化分解生成无定形SiO2和少量莫来石,与此同时,一水硬铝石转变成α-Al2O3。在焙砂的碱浸过程中,无定形SiO2溶解于NaOH溶液被脱除,而α-Al2O3和莫来石不能溶解,同时生成的水合铝硅酸钠(Na8Al6Si6O24(OH)2(H2O)2)将导致SiO2溶出率降低。焙烧过程中尾矿中的黄铁矿转化为赤铁矿、锐钛矿部分转化成金红石,在碱浸过程中它们均不会溶解而进入铝精矿中。     

白钨精矿的机械活化碱分解

测定了氢氧化钠与白钨反应的平衡常数，初步查明了有关动力学规律性，从理论到实践完善了氢氧化钠分解白钨精矿的新工艺。在１６０℃下，碱用量为理论量的２．２０￣２．４０倍时，保温１ｈ左右，白钨精矿分解率达９８．３３％￣９９．３９％。     

白云鄂博混合型稀土精矿在惰性气氛下的热分解行为（英文）

为了开发和应用白云鄂博混合型稀土精矿的先进冶炼技术，采用Kissinger公式、TGA-DSC和XRD等分析方法，研究在氮气氛下白云鄂博混合型稀土精矿的热分解行为，包括热分解动力学、物相变化规律、铈氧化效率以及物相变化对稀土浸出率的影响。结果表明：在500～550℃焙烧时，焙烧质量损失率约10%、热分解活化能(E_a)为148 k J/mol。550℃焙烧2 h，白云鄂博混合型稀土精矿中氟碳铈矿完全分解，并转化为稀土氧化物和氟氧化物，铈氧化率最大值为0.58%。600℃焙烧2 h，稀土最大浸出率达49.1%。     

微波焙烧预处理难浸含金硫精矿

对难浸含金硫精矿进行微波焙烧,考察微波功率、矿量(即样品质量)和焙烧时间对样品质量损失率和浸出率的影响。结果表明:当微波功率为16 k W、焙烧时间为50 min、矿量为900 g时,样品质量损失率可达28.12%,浸出率可达71.56%,较原矿直接碘化浸出率(9.82%)有了大幅提高。利用XRD和SEM技术分析焙烧前后样品的成分和表面形貌,微波焙烧后的样品较原矿更为松散、多孔,更利于浸出。     

H2SO4分解铁-FeCl3浸出焙烧金精矿中金

采用硫酸分解焙烧金精矿,金从黄铁矿中解离的同时金得到了富集,可采用氯化铁溶液非氰浸出金。研究了硫酸浓度及过量系数、分解温度对铁分解率的影响,优化工艺条件为,焙烧温度180 ℃,反应时间90 min,硫酸过量系数1.2,在此条件下,铁分解率为92.14%,金含量从原来的51.7 g/t提高到106.1 g/t;研究了反应温度、液固比对氯化铁溶液浸出硫酸浸出渣中金的影响,优化浸出条件为,液固比1.5,80 ℃浸出90 min,在此条件下,金浸出率96.8%。     

难浸砷金精矿的碱性常温常压预氧化

本文提供了1种难浸金精矿的湿法预氧化新工艺,它包括细磨、强化碱浸预氧化、氰化和炭吸附。在螺旋搅拌式塔式磨浸机中,先将目的难浸金精矿细磨至98%<37μm,然后在40%的矿浆质量浓度、11℃的环境温度和0.1MPa的环境压力下强化碱浸24h,NaOH的消耗量为88kg/t矿,仅为相同氧化率条件下将砷硫氧化成砷酸盐和硫酸盐所需理论碱耗量的30%。预氧化完成后经36h的氰化浸出和炭吸附,金的浸出率从预氧化前的24.6%提高到95.4%,金的吸附率99.2%,NaCN的消耗4kg/t矿。整个提金工艺的成本约300元/t矿。     

高砷难处理金精矿细菌氧化-氰化提金

通过在高砷金精矿中配入不同比例的低砷碳酸盐型金精矿,使其所含硫、砷及铁等主要矿物成分含量发生变化,研究给矿中铁砷摩尔比对难处理高砷金精矿细菌氧化一氰化浸出效果的影响.结果表明:含砷金精矿中铁砷摩尔比直接影响细菌预氧化的效果,同时也影响细菌的活性和溶液中铁砷摩尔比的变化,给矿中铁砷摩尔比越高,溶液中的铁砷摩尔比也越高,且随着给矿中铁砷摩尔比的增加,溶液中铁砷摩尔比的变化幅度加大,给矿中铁砷摩尔比介于4.6~2之间,有利于细菌预氧化和氰化浸出,铁、砷氧化率分别由6.14%和7.38%提高到89.90%和93.60%,金、银浸出率分别由64.18%和35.93%提高到97.78%和88.83%,较好地改善细菌氧化效果,稳定和优化细菌预氧化过程.     

A decomposing technique for scheelite concentrate and low-grade scheelite concentrate

The effect of different decomposition conditions on tungsten recovery for scheelite concentrate has been examined. The results show that tungsten recovery can be more than 98% under decomposing conditions as follows: the amount of caustic soda is 2.2 and 3.2 times of theoretical respectively, ratio of water and ore is 0.7-0.8, temperature is 160℃, and preservation time is 2.0 h for scheelite concentrate (63.21% WO3) and low grade scheelite concentrate (55.17% WO3).     

A new collector for wolframite slime flotation

With aniline and salicylaldehyde as main materials, a new collector for wolframite slime was synthesized. In a pulp of natural pH value, this collector can collect wolframite effectively. Its selectivity is similar to that of benzyl arsenic acid and better than that of sodium oleate. With this collector, a wolframite rough concentrate with grade 30.12 % WO3 and recovery 91.50 %, and a concentrate with grade 58.66 % WO3 and recovery 85.00 % were obtained respectively from a wolframite ore containing 4.08 % WO3.     

Leaching of complex sulphide concentrate in acidic cupric chloride solutions

The chemical analysis of a complex sulphide concentrate by emission spectrometry and X-ray diffraction shows that it contains essentially copper, lead, zinc and iron in the form of chalcopyrite, sphalerite and galena. A small amount of pyrite is also present in the ore but does not be detected with X-ray diffraction. The cupric chloride leaching of the sulphide concentrate at various durations and solid/liquid ratios at 100 ℃ shows that the rate of dissolution of the ore is the fastest in the first several hours, and after 12 h it does not evolve significantly. If oxygen is excluded from the aqueous cupric chloride solution during the leaching experiment at 100 ℃, the pyrite in the ore will not be leached. The determination of principal dissolved metals in the leaching liquor by flame atomic absorption spectrometry, and the chemical analysis of solid residues by emission spectrometry and X-ray diffraction allow to conclude that the rate of dissolution of the minerals contained in the complex sulphide concentrate are in the order of galena 〉 sphalerite 〉 chalcopyrite.     

高砷硫金矿的预处理

黄铁矿和砷黄铁矿是高砷硫金矿和精矿的2种主要成分，它们将Au包裹在其中，用一般已知的方法处理该矿，存在Au收率低、过程复杂，易造成环境污染等问题。目前对高砷硫金矿的预处理方法主要有氧化焙烧法、加压氧化法、细菌氧化法、硝酸分解法、真空脱砷法等。其中真空脱砷法处理高砷硫金精矿能保护环境免受污染，同时使As作为对环境无污染产物析出，是一种理想的预处理方法。     

A Novel Self-Heated Roasting Technology for Molybdenum Concentrate

对钼精矿焙烧过程进行了热力学分析。结果表明,焙烧反应属于放热反应,反应一旦开始,在工业规模生产条件下完全可以自热进行。热分析实验表明,氧气浓度越高,越有利于钼精矿的转化率提高;物料粒度越小,越有利于焙烧反应完全程度的提高。对传统钼精矿焙烧回转窑的热平衡测试显示,煤作为外热源经燃烧后提供热量为4.42 GJ/t的烟气供入回转窑内,烟气中的氧浓度为10%。由于主反应区温度高,造成物料结块,阻碍了钼精矿的焙烧。本研究开发了钼精矿的自热焙烧新工艺,该工艺通过设置换热器回收主反应区放出的化学热,并将预热后的空气用于脱硫区的补热,提高了焙烧气的氧浓度。取消了传统回转窑的燃煤过程,整个焙烧过程仅靠6.45 GJ/t的化学反应热即可维持,没有含碳燃料的输入和CO2的排放,其节能环保效益显著。     

Preparation of High－purity Sodium Tungstate from Low－grade Tung－sten－concentrate with a High Content of Calcium and Other Impurities

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多金属硫化精矿中铜、锌和铁的生物浸出(英文)

利用嗜温混合菌Acidithiobacillus ferrooxidans,Acidithiobacillus thiooxidans和Leptospirillum ferrooxidans对低品位复杂Cu-Zn-Pb-Fe-Ag-Au硫化精矿在曝气生物浸出反应器中进行生物浸出。该菌种为从塞尔维亚Bor地下铜矿的酸性溶液中筛选出一种嗜热嗜酸菌。营养液为p H 1.6的9K营养液。87%的矿物粒度大于10μm,矿浆密度为8%(w/v)。在测试条件下,锌、铜和铁的浸出率分别达到89%、83%和68%。动力学分析表明,浸出过程与Spencer-Topley模型相符,受局部反应扩散控制。     

Wolframite Conversion in Treating a Mixed Wolframite–Scheelite Concentrate by Sulfuric Acid

Complete wolframite conversion in sulfuric acid is significant for expanding the applicability of the sulfuric acid method for producing ammonium paratungstate. In this paper, the conversion of wolframite in treating a mixed wolframite–scheelite concentrate by sulfuric acid was studied systematically. The results show that the conversion of wolframite in sulfuric acid is more difficult than that of scheelite, requiring rigorous reaction conditions. A solid H₂WO₄ layer forms on the surfaces of the wolframite particles and becomes denser with increasing H₂SO₄ concentration, thus hindering the conversion. Furthermore, the difficulty in wolframite conversion can be mainly attributed to the accumulation of Fe²⁺ (and/or Mn²⁺) in the H₂SO₄ solution, which can be solved by reducing Fe²⁺ (and/or Mn²⁺) concentration through oxidization and/or a two-stage process. Additionally, the solid converted product of the mixed wolframite–scheelite concentrate has an excellent leachability of tungsten in an aqueous ammonium carbonate solution at ambient temperature, with approximately 99% WO₃ recovery. This work presents a route for manufacturing ammonium paratungstate by treating the mixed concentrate in sulfuric acid followed by leaching in ammonium carbonate solution.     

Decomposition reaction kinetics of Baotou RE concentrate with concentrated sulfuric acid at low temperature

Baotou RE concentrate was decomposed with concentrated sulfuric acid by controlling the roasting temperature below 500°C.Thermogravimetry-differential thermal analysis(TG-DTA) and chemical analytical methods were used to study the thermal decomposition process and the thermal decomposition effect.The Freeman-Carroll method was applied to analyze the TG-DTA curves.The activation energy, reaction order, and reaction frequency factor at different stages were calculated.The Satava method was used to deduce the reaction mechanism and the relative reaction rate during the thermal decomposition process.