高硅锌精矿焙烧过程中硅的控制及热平衡计算

针对高硅锌精矿焙烧过程中焙砂可溶硅高、沸腾炉易结块、浸出固液分离困难等问题,论文以现场生产焙烧工艺参数为基础,研究了低温和高温焙烧对焙砂中可溶硅含量的影响,并基于MatCal软件对沸腾炉焙烧工艺进行热平衡计算。结果表明:在焙烧条件基本相同的情况下,随着硫化锌精矿焙烧温度的增加,焙砂中的可溶硅也增加。当焙砂中可溶硅高于3.18%会出现浓密机上清液跑混、低浸浓密底流矿浆过滤困难、净液中除杂钴偏高等问题。经MatCal模拟计算后,理论消耗空气50361.328Nm3/h,低温焙烧的平均风量47102.8m3/h,高温焙烧平均风量48005.7m3/h,实际的焙烧中平均风量偏低,需要增加沸腾炉的风料比。     

The effect of temperature and culture history on the attachment of Metallosphaera hakonensis to mineral sulfides with application to heap bioleaching

Temperatures in excess of 60 °C are required for efficient bioleaching of chalcopyrite. Within heaps, colonisation of the mineral with thermophilic archaea is important in reaching and maintaining these high temperatures. The effect of temperature and culture history on the attachment of Metallosphaera hakonensis, an extreme thermophilic acidophile identified as a key player in heap bioleaching, to sulfide concentrates and low-grade ore was investigated in shake flasks and packed beds. Attachment studies were conducted at 25 °C, 45 °C and 65 °C. The results show a clear relationship between increasing temperature and attachment efficiency for both suspended and packed bed systems. Attachment at 25 °C was low. Increasing the temperature to 45 °C improved attachment efficiency by between 50% and 100% while a further increase to 65 °C improved attachment by an additional 20-50%. Cells cultured on elemental sulfur as energy source prior to contacting showed 1.3 times greater affinity for the mineral concentrate than those cultured on sulphide mineral concentrates or ferrous sulphate. In contrast to previous studies using mesophilic organisms the selective attachment of Metallosphaera to sulfide minerals, relative to gangue, was less pronounced. Attachment efficiency was lower in the packed bed system which more closely mimicked flow through a heap. The cell surface properties surface charge and hydrophobicity as well as metabolic activity were investigated to provide insight into the observed phenomena. The data suggest that retention of thermophiles within the heap could be enhanced by a secondary inoculation following elevation of the temperature above 40 °C by the mesophilic pioneer species.     

Dispersion effect on a lead–zinc sulphide ore flotation

This investigation was performed with samples from a lead–zinc sulphide deposit aiming at studying the influence of the dispersion degree of the particles in the pulp on lead and zinc flotation. Samples of ore and also of the minerals sphalerite, galena, pyrite, and dolomite were selected for the experiments. Nine types of dispersing agents and six blends among them were employed.A set of three dispersing agents was selected for the lead flotation and another set of three was chosen for zinc flotation. The criteria for the reagents selection were: high dispersion degree for galena and low for the other species, high dispersion degree for sphalerite and low for the other species, low dispersion degree for pyrite and high for the other species, and high dispersion degree for all species.Lead flotation experiments were performed under three conditions aiming at verifying the influence of the dispersing agent, of the pH, and of sodium carbonate. The zinc flotation tests were carried out at pH 10.5, modulated with lime.The use of dispersing agents in lead flotation did not improve the overall efficiency of the circuit for, despite improving the lead metallurgical recovery, they increase significantly the zinc losses in the lead concentrate.Sodium carbonate presented a low dispersion degree and did not affect the lead flotation results when compared with those achieved at natural pH and at pH 9.8 modulated with lime.Two dispersing agents were particularly effective in zinc flotation: dispersant 3223, a sodium polyacrylate, and sodium hexametaphosphate. Both reagents significantly enhanced zinc recovery without impairing the concentrate quality.     

Enhanced separation of mineral sands using the Reflux Classifier

The Reflux Classifier consists of a conventional fluidized bed, with a set of parallel inclined plates immediately above. The fluidized suspension passes into the inclined channels where relatively fast settling particles segregate, slide down the plates and return to the fluidized bed. The slower settling particles pass up through the inclined channels and into the overflow. By increasing the aspect ratio of the inclined channels, the effective sedimentation area of the vessel increases, in turn increasing the hydraulic capacity of the device. Laskovski et al. [Laskovski, D., Duncan, P., Stevenson, P., Zhou, J., Galvin, K.P., 2006. Segregation of hydraulically suspended particles in inclined channels. Chem. Eng. Sci., doi:10.1016/j.ces.2006.08.024], however, have shown that there exists an asymptotic limit to this hydraulic capacity, due to an increased tendency for particle re-suspension within the inclined channels as the aspect ratio increases. Laskovski et al. (2006) showed that the re-suspension facilitates the separation of the particles on the basis of density, reducing the dependence on the particle size. A conventional fluidized bed does not offer the benefits of the mechanism identified by Laskovski et al. (2006). In this study, the Reflux Classifier was used to recover and concentrate the heavy minerals presented in a low grade feed of mineral sands. Enhanced separation was achieved using inclined channels having a large aspect ratio of about 200, thus promoting the particle segregation of the denser particles while also promoting the re-suspension and hence hydraulic conveying of the lower density silica sand. With the aspect ratio of approximately 200, recoveries of heavy minerals in excess of 90% were achieved at solids throughputs of up to 40 t/m² h. A heavy mineral recovery of 97% was achieved at a solids throughput of 21 t/m² h. The separations resulted in heavy mineral grades approaching 100% in the size range of 90–180 μm.     

微细粒人造硫化锌矿的选择性絮凝试验研究

人造硫化锌矿为冶炼废水处理过程产生的中和渣经水热硫化合成后生成的一类硫化物。由于其粒度细,表面自由能高,各组分互相团聚,很难用浮选的方法进行分离。试验采用选择性絮凝的方法,在酸性条件下,用六偏磷酸钠作为分散剂,用糯米淀粉作为絮凝剂,能使硫化锌矿锌的品位由14%提高至28%左右,回收率达到90%。通过絮凝前后硫化矿的XRD图谱对比分析发现絮凝后的硫化矿中,硫化锌的特征峰较絮凝前高且尖锐,可知絮凝后的硫化锌更加接近天然闪锌矿,为后续提取作业创造了条件。     

复杂高铅低铁硫化锌精矿氧压浸出

以选矿尾矿经二次选矿获得的铅锌混合精矿为主要研究对象,针对其含铅高、铁较低的特点。对比研究了不同硫化锌精矿的氧压浸出效果,分析了高铅低铁硫化锌精矿的氧压浸出行为。结果表明,低铁锌精矿需在二段补加5 g/L的铁传递氧,锌浸出率达99%以上,铜浸出率约90%。铁大多以二价铁的形式随锌进入到浸出液,少部分入渣,以黄铁矿的形式存在,并有少量的铁氧化物;铅、银、硅沉淀入渣并在渣中富集,浸出渣可实现铅、银等有价金属的回收,精矿中的硫主要以单质硫的形式入渣。在两段氧压逆流浸出中,二段浸出液中铜会沉淀进入一段渣,在系统里循环累积,直至平衡,终渣含铜0.16%,一段浸出液含铜1.00 g/L,具有较高回收价值。     

某低品位铅锌硫化矿浮选试验研究

某硫化铅锌矿含铅锌原矿品位低、嵌布粒度细、伴生关系复杂。通过多种方案的比较,采用优先浮选抑锌浮铅的选别流程,试验采用乙硫氮作为优先选铅的捕收剂,石灰作为调整剂以及黄铁矿的抑制剂,硫酸锌和亚硫酸钠作为闪锌矿的抑制剂,之后利用硫酸铜作为闪锌矿的活化剂,用丁基黄药作为捕收剂来实现铅与锌的有效分离。试验获得铅精矿含铅51.00%、铅回收率86.63%、含银518 g/t、银回收率47.41%,锌精矿含锌51.20%、锌回收率85.27%、含银234 g/t、银回收率38.38%。     

豫西某难选铅锌矿选矿试验研究

豫西某铅锌矿有用矿物共生关系密切、嵌布粒度较细, 采用铅锌等可浮、铅锌分离-硫化锌浮选-氧化铅浮选工艺, 成功实现了该矿的铅锌回收与分离, 并有效回收了氧化铅矿物, 最终获得了铅品位、回收率分别为58.95%、68.67%的铅精矿和锌品位、回收率分别为48.67%、66.06%的锌精矿。     

某高硫铅锌矿石选矿试验

某高硫铅锌矿石中磁黄铁矿和黄铁矿含量大、铅锌嵌布关系复杂、嵌布粒度细等,以新药剂BK-509和BK-512抑制硫化铁矿物,采用磁选-铅锌依次优先浮选工艺进行了铅、锌、硫分离试验。结果表明,矿石在磨矿细度为-0.074 mm占90%的情况下,经1粗1精弱磁选、2粗2扫浮选选铅、铅粗精矿再磨至-0.043 mm占85%情况下4次精选、铅扫选尾矿1粗2扫选锌、锌粗精矿再磨至-0.043 mm占90%情况下4次精选,获得了铅品位为56.71%、回收率为76.85%的铅精矿,锌品位为45.98%、回收率为75.57%的锌精矿。试验的铅、锌精矿指标理想,可作为铅锌回收工艺流程设计的依据。     

山西某铅锌矿的分选研究

以山西某铅锌硫化矿为研究对象,在工艺矿物学研究基础上采用铅锌优先浮选流程,确定了最佳药剂制度,最终获得了铅品位为53.06%、回收率为90.48%的铅精矿和锌品位为54.02%、回收率为87.12%的锌精矿。     

某铅锌硫化矿选矿试验研究

以山西某铅锌硫化矿为研究对象,试验采用抑锌浮铅的优先浮选流程,通过条件试验,确定了最佳的药剂制度,获得了铅品位为38.92%、回收率为90.65%的铅精矿和锌品位为50.64%、回收率为88.58%的锌精矿,为选矿厂的改扩建提供了依据。     

青海某高硫高铁铅锌硫化矿选矿试验

青海某铅锌硫化矿中矿物间共生包裹关系复杂,主要硫化物磁黄铁矿占金属矿物总量的51%,锌以铁闪锌矿形式存在,锌硫分离难度较大。试验采用铅锌硫依次浮选流程,小型闭路试验获得铅品位51.58%,回收率89.98%的铅精矿和锌品位42.74%,回收率81.81%的锌精矿以及硫品位35.70%,回收率72.72%的硫精矿,试验指标较为理想。     

云南沧源某氧化铅锌矿浮选工艺研究

以云南省沧源县某深度氧化,且锌主要以异极矿形式存在的难选铅锌矿石为研究对象,进行了丁黄药直接浮选方铅矿、硫化-丁黄药浮选白铅矿、硫化-苯硫酚浮选异极矿的工艺技术条件研究。采用试验确定的先浮铅后浮锌、先选硫化矿后选氧化矿的闭路浮选工艺处理该矿石,获得了铅品位为53.93%、含锌13.13%、银品位为1 751.90 g/t的铅精矿,锌品位为31.82%、含铅2.75%的锌精矿,以及铅品位为33.38%、锌品位为19.10%、银品位为694.85 g/t的铅锌混合精矿,铅锌银的综合回收率均超过98%。     

A mechanistic model to predict matte temperatures during the smelting of UG2-rich blends of platinum group metal concentrates

High matte temperatures can be related to numerous catastrophic furnace failures in the platinum group metal (PGM) industry where chromite-rich upper group 2 (UG2) concentrates are smelted. Chromite rich concentrates require high slag temperatures as well as sufficient mixing to suspend the chromite spinel particles in the slag and prevent settling in a so-called “mushy” layer consisting of a three phase emulsion of slag, matte and chromite particles. To achieve sufficient bath mixing and to melt and suspend chromite spinel build-up, high hearth power densities are utilised. However, high hearth power densities in conjunction with a heat-isolating concentrate layer, leads to high side wall heat fluxes which motivated the use of intensive cooling in the furnace side wall so that a slag freeze lining can be formed. If matte temperatures are above the slag liquidus temperature, any matte that comes into contact with the freeze lining can destroy the freeze lining. Moreover, if the matte temperature exceeds ca. 1500 °C, chemical thermodynamics indicate that matte has the ability to sulfidise MgO-Fe_xO-Cr₂O₃ refractories, leading to rapid wear of refractories exposed to high temperature flowing matte. Models are derived for the concentrate-to-matte and slag-to-matte droplet heat transfer. Calculations using the derived models, physical properties and furnace operating conditions give realistic matte temperatures and show that matte temperatures rapidly increase as the concentrate bed becomes matte drainage rate limiting. It is shown that for each concentrate blend mean particle size and mineralogy, there is a maximum smelting rate above which the concentrate bed becomes rate limiting with regards matte drainage, thereby significantly contributing to matte preheating, prior to further heat absorption from the slag layer.     

某硫酸渣中铅、锌的氯化焙烧动力学研究

硫酸渣是一种大宗固体工业废弃物,铁含量较高,含量偏高的铅、锌往往是制约其作为铁资源利用的重要因素。氯化焙烧-磁化焙烧-磁选工艺则可成功脱除铅、锌,获得高铁低铅锌铁精矿。为揭示硫酸渣氯化焙烧过程中各主要相态的铅、锌发生氯化反应的限制环节,以及氯化反应的速率和氯化焙烧机理,以CaCl₂为氯化剂,对某硫酸渣进行了氯化焙烧动力学研究。结果表明：①铁、铅、锌含量分别为49.90%、0.29%和1.23%,锌绝大部分为氧化态,铅主要为氧化态,其次是硫酸铅和其他形态铅,在CaCl₂与硫酸渣的质量比为6%的情况下,延长氯化焙烧时间或提高焙烧温度,锌、铅的氯化挥发脱除率均上升,1 000 ℃时焙烧5 min,锌、铅的脱除率分别达86.99%和83.14%,为后续磁化焙烧-磁选制备高铁低杂铁精矿创造了良好的条件。②相比较而言,氯化焙烧脱锌比脱铅更容易。③900～1 050 ℃时锌氯化挥发的表观活化能为42.07×10³ J/mol,受化学反应控制;900～950 ℃时铅氯化挥发的表观活化能为43.88×10³ J/mol,受化学反应控制;1 000～1 050 ℃时铅氯化挥发的表观活化能为20.34×10³ J/mol,受扩散控制。④强化铅、锌的氯化挥发脱除,除了提高温度,还可通过增加固体氯化剂用量或提高硫酸渣固体颗粒的孔隙率和比表面积来实现。     

三相流化床中硝酸氧化难选冶金精矿动力学研究

以难选冶金精矿为原料, 三相流化床中硝酸氧化金精矿反应过程中铁的转化率为实验指标, 考察了气速、硝酸浓度、温度和粒径等因素对铁转化率的影响, 研究了三相流化床中硝酸氧化金精矿的反应动力学。结果表明: 在三相流化床中, 铁的转化率受气速、硝酸浓度、温度和粒径等因素的影响。随着气速、硝酸浓度及反应温度的增加, 转化率均有所提高, 而随着金精矿粒径的增加, 转化率降低。三相流化床中硝酸氧化难选冶金精矿符合颗粒缩小的缩核模型, 反应活化能为43.2 kJ/mol, 属于化学控制。     

综合回收某硫铁矿石中伴生铜锌的研究

从某硫铁矿石中综合回收铜锌等伴生金属的关键是铜锌分离技术,根据该矿矿石含硫高、铜锌含量低且致密共生等特点,制定了优光浮选工艺流程,在中性矿浆中采用以硫化钠为主的组合抑制剂,成功地实现了铜铸浮选分离,获得了铜精矿、锌精矿及硫精矿三种合格产品,较好地实现了矿石资源的综合利用。     

复杂铜铅锌多金属硫化矿选矿试验研究

针对某地含银铜铅锌多金属硫化矿易浮难分、嵌布粒度极不均匀的特点,采用优先浮选工艺流程,以硫化钠消除次生铜离子的影响,组合药剂浮选铜铅锌,铜铅粗精矿再磨显著提高铜铅锌分选效果,获得了较佳的分选指标,铜精矿含铜23.44%、回收率88.83%,铅精矿含铅54.43%、回收率84.28%,锌精矿含锌55.72%、回收率83.72%。     

A new method for testing the self-heating character of sulphide concentrates

A rapid, inexpensive, and quantitative method has been developed to quantify the self-heating potential of sulphide concentrates. The method consists of measuring the sulphur dioxide evolved when a concentrate is heated and oxidized in air at 300 °C. Sulphur dioxide is the reaction product of the oxidation of elemental sulphur present in the concentrate. We found a strong correlation between the output of the sulphur method (i.e. the elemental sulphur grade) and the output of the United Nations self-heating protocol (i.e. the temperature rise). Both methods provide evidence of the extent of prior oxidation of the sulphide concentrate, but do not inform about the rate at which the concentrate was oxidized.     

Effect of grinding media on zinc depression in a lead cleaner circuit

Laboratory testwork, investigating the effect of high chrome grinding media in a lead regrind application has on subsequent metallurgical performance, was conducted at a large silver–lead–zinc operation in Australia. The initial data showed that the use of a more inert grinding media could have increased zinc losses to the lead cleaner concentrate if careful attention was not paid to alloy selection.Diagnostic tests showed that iron hydroxide surface coatings generated by grinding media corrosion reactions are an effective depressant for sphalerite in this ore body, even though it is known that an excess of these coatings could depress both galena and sphalerite flotation. These tests demonstrated that a 1% chrome alloy produced the desired pulp chemical conditions to yield an increase in lead concentrate grade through the rejection of sphalerite from the lead circuit.A plant trial was conducted in one of the two parallel grinding/flotation trains, and data collected for statistical analysis. During the plant trial, pulp chemical surveys of the regrind circuit were also taken to compare the effect of grinding media on the cleaner one feed slurry pulp potentials, dissolved oxygen, pH, temperature and EDTA extractable iron.The statistical analysis showed clearly that the change to 1% chrome grinding media had a significant positive impact on improving galena/sphalerite selectivity during lead cleaner flotation and improved the lead concentrate grade. The improved metallurgical performance is explained in terms of modified pulp chemistry.