Method of burning the molybdenite concentrates

A new method of burning the molybdenite concentrate is developed. An installation is used in which material is burned in the vortical flow of the oxidizer with obtaining the cinder. It is established that, at a temperature of 600°C and air consumption of 100–120 m 3 /h, 98.7% of MoS₂ transforms into MoO₃ and 85% of Re sublimes in 0.5 h. The sulfur content in the cinder is <1.8%. Upon treating the cinder with a potassium carbonate solution, up to 96.5% Mo is extracted. This method provides a rather high productivity, it is rather simple, it does not require large expenses for equipment and exploitation, it is promising, and it can be used for burning other materials containing sulfides of metals.     

Appraisal study on processing low-grade molybdenite concentrates

This paper investigates how the low-temperature roasting of low-grade molybdenite concentrates ([Mo] < 20%) produced at different deposits and the posterior leaching of stubs with soda and alkali affect Mo recovery in a solution. Based on the results obtained, it has been revealed that there do seem to be sufficient reasons for using soda as a selective chemical for leaching stubs; the roasting and leaching modes have been optimized and a processing flowchart of the above concentrates has been proposed. Such a flowchart ensures molybdenum recovery in a solution at 87–97%, depending on the concentrate’s phase composition.     

Autoclave leaching of molybdenite concentrates with catalytic additives of nitric acid

The scientific theoretical prerequisites for the use of small additives of nitric acid for the intensification of the autoclave oxidative leaching (AOL) of sulfide concentrates are checked in order to increase the degree of extraction of the base metal with a simultaneous decrease in the cost of the process. The technological parameters of the AOL conditions applied to molybdenite are also refined, and recommendations for a practical application of the apparatus of the process and for a decrease in its cost are made.     

Variants of processing molybdenite concentrates involving the use of preliminary mechanical activation

A review of possible ways to process molybdenite concentrates of various grades by both pyrometallurgical and hydrometallurgical methods is given and the most rational are noted. Prospects for processing low-grade molybdenum raw materials by nitric acid autoclave-less leaching with the complete transfer of molybdenum into the solution in the composition of sulfate complexes is shown. The variant that intensifies the nitric acid leaching of low-grade (6–17% Mo) concentrates obtained by the flotation of ores of the Bugdainskoe deposit via their preliminary dry mechanical activation is suggested.     

金铜硫化矿火法冶炼技术的发展

针对黄金冶炼企业的工艺技术现状,提出了造锍捕金的技术发展方向。对涉及造锍捕金的铜火法冶炼中的熔炼和吹炼工艺特点进行述评,详细介绍了底吹连续炼铜技术的进展,指出连续炼铜技术将成为新建或搬迁、升级改造金、铜冶炼项目的必然选择。     

Agglomeration of particles during roasting of zinc sulfide concentrates

This study was undertaken in an attempt to explain the agglomeration of zinc concentrate particles that occurs during fluid bed roasting. In order to simulate extensive particle contact, pure ZnS, FeS, and (Zn, Fe)S particles were roasted in a fixed bed at temperatures between 1223 K and 1323 K using an argon-oxygen gas mixture withP _O2 = 0.21 atm. The results were compared to the reaction products and morphology of an agglomerated commercial concentrate. The reactants were contained in MgO crucibles suspended in a thermobalance and the weight change was recorded as the reaction proceeds. The resulting data were used to identify the specific reaction that takes place for each of the reactants. Products were identified by X-ray diffraction analysis. The morphology of the products was analyzed using SEM and confirmed that the formation of a liquid phase composed of a solution of Fe-S-O causes particles to stick to each other, thus causing agglomeration. This phenomenon was observed for both the FeS and (Zn, Fe)S samples but not for the ZnS. The morphology and the X-ray diffraction pattern of the reaction products of the synthetically prepared (Zn, Fe)S showed remarkable similarity to that of the roasted industrial concentrate.     

Lime-enhanced reduction of sulfide concentrates: A thermodynamic discussion

The need to control or eliminate sulfur dioxide emissions from sulfide smelters has increased the drive to develop new processes for the extraction of metal from sulfide minerals. One such process currently gaining interest is the lime-enhanced reduction of metal sulfides. This paper discusses the thermodynamic aspects relevant to chalcopyrite, chalcocite, and pyrrhotite reduction with hydrogen, carbon monoxide, or carbon in the presence of lime. The effects of temperature and gas composition on sulfide reduction are also discussed. ΔG°vs T diagrams for lime-enhanced sulfide reductions with hydrogen, carbon monoxide, and carbon are constructed and discussed together with some experimental results produced by the authors.     

The bioleaching of different sulfide concentrates using thermophilic bacteria

The bioleaching of different mineral sulfide concentrates with thermophilic bacteria (genusSulfolobus @#@) was studied. Since the use of this type of bacteria in leaching systems involves stirring and the control of temperature, the influence of the type of stirring and the pulp density on dissolution rates was studied in order to ascertain the optimum conditions for metal recovery. At low pulp densities, the dissolution kinetic was favored by pneumatic stirring, but for higher pulp densities, orbital stirring produced the best results. A comparative study of three differential concentrates, one mixed concentrate, and one global concentrate was made. Copper and iron extraction is directly influenced by bacterial activity, while zinc dissolution is basically due to an indirect mechanism that is activated in the presence of copper ions. Galvanic interactions between the different sulfides favors the selective bioleaching of some phases (sphalerite and chalcopyrite) and leads to high metal recovery rates. However, the formation of galvanic couples depends on the type of concentrate.     

The extraction of gold from sulfide concentrates into molten lead

The results of investigation of the extraction of gold from mineral sulfide ore into molten lead at low temperatures (350–550°C) are given. The questions of solubility of Au in Pb in the mentioned temperature range and the kinetics of dissolution of metal are analyzed. The condition that determines extraction of gold is the presence of a processing medium, which provides the decomposition of carrier minerals of gold and prevents the anodic oxidation of the lead alloy. The alkali melt is suggested to be used as such a medium. The procedures of preparation of sulfide Au-containing concentrates to the extraction associated with obtaining the cakes and their charging into the extraction apparatus are considered. Extraction proceeds in the transient region. In this case, high characteristics of extraction are achieved. The reasons for collector dispersion, which conditions the losses of gold, are established.     

Improvement of modes and circuits of bacterial leaching of sulfide concentrates

These studies are devoted to the effect of the feed circuit of a series of chemostats on the characteristics of bacterial leaching of sulfide concentrates. The parameters and modes of the process are investigated at a laboratory installation consisting of five operating reactors, each with an active volume of 1.5 l and a conditioning tank. The initial material was a gold-bearing arsenical pyrite flotation concentrate containing Au, Ag, and As, as well as a specially adapted bacterial strain, namely, Acidithiobacillus ferrooxidans. It is established that double-flow feeding of a series of chemostats makes it possible to increase the leaching yield up to a 70% excess of the critical rate of dilution under the single-flow feeding mode at the same degree of oxidation of output sulfide minerals.     

Characteristics of electrochemical reactions accompanying chlorinating annealing of copper sulfide concentrates

The mechanism for reactions accompanying the low-temperature roasting of a copper sulfide concentrate with potassium and sodium chlorides without the formation of sulfurous gases and their evolution into the atmosphere is suggested. The obtained cake is leached with water under pH correction in the vicinity of 6. Water-soluble sulfates transfer into the solution, while copper and iron hydroxides remain in the cake. The further leaching of the cake by an aqueous sulfuric acid at the final pH 2.5–3.0 makes it possible to completely transfer copper into the solution, while iron completely remains in the solid phase. The recovery of copper into the solution under the optimal annealing conditions (t = 450°C, τ = 1 h) and the above conditions of double leaching is no lower than 95%. Chlorinating annealing at temperatures below and above 450°C results in a decrease in the recovery of copper into the final product of copper sulfate. The found maximum is explained.     

Selective recovery of copper from copper-nickel sulfide concentrates by applying segregation technology

Copper can be selectively segregated from roasted copper-nickel sulfide concentrates. Conditions are reported for desulfurizing-roasting followed by segregation treatment to selectively reduce and segregate copper without nickel contamination. A process is described which is applicable for treating low grade sulfide concentrates such as are produced in Botswana (4.6 pct Cu, 3.1 pct Ni) and higher grade concentrates that can potentially be produced from sulfide deposits in Minnesota (13 pct Cu, 2.5 pct Ni). Steps include roasting to remove at least 97 pct of the sulfur, segregation treatment of calcine and flotation to recover a copper product, a middlings fraction which is recycled to the desulfurizing roast and a tailings fraction which can be subsequently treated for nickel recovery. LAMAR D. COFFIN, formerly Senior Research Metallurgist at AMAX Base Metals R&D, Inc., has retired.     

某难处理硫化金精矿加压氧化一氰化浸金试验研究

对云南某难处理硫化金精矿进行加压氧化一氰化浸金试验研究,考察了加压氧化各因素对氰化浸金的影响。加压氧化最优条件为：固液比1：4,木质素磺酸钠5g／t,硫酸初始质量浓度10g/L,温度190℃,压力2．0MPa,反应时间4h,搅拌转速450r／min。金精矿经加压氧化一氰化浸出获得了97．55％的较高金浸出率。     

An investigation of certain regularities of oxygen-flash smelting of copper sulfide concentrates using computer simulation

Dependences of temperature of the products of oxygen-flash smelting of copper sulfide charge on the furnace capacity, copper content in the charge, oxygen content in the blast, and charge humidity are investigated. A mathematical model is suggested that was obtained by implementing the planned simulation experiment. Mass and heat balances are calculated using developed software. Independent variables are ranged by the force of their effect on the slag temperature. The strongest effects are exerted, first of all, by the copper content in the matte, followed by the amount of copper and sulfur in the charge.     

硫精矿除杂提纯浮选工艺回收利用硫酸烧渣中的铁

硫酸烧渣是硫铁矿制酸氧化焙烧产物;从硫酸烧渣中选铁的工艺技术指标一直不高,其主要原因是硫铁矿氧化焙烧过程中生成的氧化铁矿物颗粒微细,高温时新生成的氧化铁矿物颗粒会与杂质和脉石矿物颗粒相互包裹、相互黏结、相互污染。该文将硫酸烧渣选铁改为硫精矿再浮选提纯硫化铁,即通过提纯硫酸原料中硫化铁的质量分数,从而去除原料中的脉石和杂质,使硫酸原料中硫品位达到50%~52%(黄铁矿型原料)以上,硫、铁回收率均达到90%~92%;采用该高纯硫精矿制造硫酸,硫酸烧渣中铁品位达到63%~67%,使硫酸烧渣全部直接成为铁精矿,无需再选矿,达到了有效利用硫酸烧渣中铁的目的。该工艺能够获得较高技术指标的原因是硫精矿除杂提纯浮选是硫化矿选硫,采用高纯硫精矿制酸,避免了非目的矿物污染硫化铁的氧化焙烧过程以及硫酸烧渣选铁时杂质含量高、铁品位低、选矿技术指标低等问题。     

Lime-enhanced hydrogen reduction of molybdenite

Kinetics of the direct hydrogen reduction of a high-grade (59 pct Mo) molybdenite (MoS₂) concentrate was investigated in the presence of lime as a function of the quantity of lime in the charge, hydrogen flow rate, temperature, and time of reduction. Lime was found to enhance tremendously the reduction rate of MoS₂ and drastically reduce H₂S emission into the off gas to negligible levels. Successful application of the lime-hydrogen reduction technique was found to depend on the employment of low hydrogen flow rate and moderate temperatures of reduction. In these laboratory studies, best results were obtained with a lime addition ≥ three times the theoretical requirement and at 1173 K in 3.6 ks employing a hydrogen flow rate of 3.33 cm³s^-1. The results were tested for the treatment of a low-grade (41 pct Mo) molybdenite concentrate. In this latter case, the procedure consisted of upgrading the concentrate by acid leaching (with dil HC1+HF) followed by lime-hydrogen reduction. The influence of quantity of acids, temperature, and time of leaching were investigated to optimize the conditions required for upgrading the MoS2 concentrate. The molybdenum powders obtained from the highgrade as well as upgraded molybdenite concentrates had 96 to 97 pct purity and could be further refined to 99.9 pct by electron-beam melting. Based on this lime-enhanced hydrogen reduction concept, a new ‘Leach-Reduction-Melting’ approach has been suggested as an alternative to the traditional methods of molybdenum extraction.     

Acid bake-leach-flotation treatment of offgrade molybdenite

A process has been developed and piloted to reject Cu, Pb, Fe, P, Bi, Sb, elemental sulfur, and insol from molybdenite concentrate without decomposing the contained molybdenite and rhenium. The impure concentrate is first heated with concentrated sulfuric acid to attain a 260° to 290°C final bed temperature. This bake not only oxidizes the contaminating sulfides, but also activated the concentrate for later rejection of insol by flotation. Two stages of leaching follow the sulfation step. The first is an acidified water leach which extracts the oxidized Cu, Fe, and P; the second leach rejects the Pb, Bi, and Sb into a warm concentrated solution of sodium chloride. Final upgrading is achieved by conventional froth flotation. The chemistry of the process as well as results obtained from its implementation in a 5000 lb per day pilot plant are presented.