A tellurium recovery process from refractory tellurium-containing tailings was studied. The feasi-bility was discussed of tellurium recovery by the integrated process flowsheet of flotation, gravity separa- tion and magnetic separation. The results show that two types of concentrate from flotation and gravityseparation are obtained by one-stage grinding and flotation of raw tailings, gravity separation of flotation tailings, and magnetic separation of gravity-separation concentrate. The average tellurium grade of themixed concentrate is 1.65% ,and its total recovery is up to 45.97%. And the concentrate product can be used as metallurgical raw materials for further tellurium extraction.

Technological conditions were studied of cobalt enrichment from magnetite tailings by segrega- tion roasting-magnetic separation. The results from orthogonal experiments show that carbon powder a- mount is the strongest influencing factor on cobalt content of the concentrate product, followed by calcina- tion temperature, and sodium chloride amount is the last one. Higher carbon powder amount and roasting temperature are conducive to cobalt enrichment. The cobalt grade of the concentrate reaches 0. 044 5% with a cobalt recovery of 75.35% ,under the conditions of 1 200 ℃ roasting temperature,20% carbon pow- der amount,4% sodium chloride amount and 2 h roasting time.

Thermal balance was tested of traditional roasting rotary furnaces of molybdenum concentrate, and its roasting process was analyzed. It is pointed out that MoS2 roasting can take place based on self-heat- ing under the industrial-scale production conditions, and proper cooling measures should be adopted to pre- vent overheating. Therefore, carbon-free roasting technology of molybdenum concentrate was developed, and thermal balance tests and technical analysis of the corresponding rotary furnace were carried out. The results show that outer heat source used in traditional roasting no longer appears in carbon-free roasting, because roasting heat is enough to ensure spontaneous occurrence of the reaction. In fact,heating is needed just at the beginning of the reaction to make molybdenum concentrate on fire and after violent reaction to remove residual sulphur, without additional heating during operation. The above technology has obvious advantages in saving energy and emission reduction, improving yield and quality of molybdenum calcine, re- ducing heat loss, and increasing SO2 concentration of flue gas.

Abstract. Experimental study was conducted on molybdenum leaching with NaOH from molybdenum con- centrate. The effects were investigated of leaching temperature and time, particle size of molybdenum con- centrate,stirring speed and the coefficient β （i. e. the mass ratio of NaOH to molybdenum concentrate）, and liquid to solid ratio on molybdenum leaching rate. The optimal technical parameters of molybdenum leaching were also determined. The average molybdenum leaching rate is 99.76% from the molybdenum concentrate with particle size of less than 0. 1 ram,under the conditions of leaching for 90 rain at 85 ℃, stirring speed of 450 r/min,β value of 1.2,and liquid to solid ratio of 4：1. Study on the above leaching process provides reliable technical parameters for processing of refractory molybdenum concentrate.

A new process of molybdenum-tungsten extraction from high pressure leaching solution of crude molybdenum-tungsten oxide concentrate was studied,including the steps of solution pre-treatment, molyb- denum-tungsten precipitation, solid-liquid separation, product drying and so on. When the high pressure leaching solution contains 27.26 g/L Mo and 5.70 g/L WOa ,Mo and WOa precipitation rates are 99.94% and 85.08 %, respectively, under the conditions of pH value of 7-8, precipitation time of 1-1.5 h, reaction temperature of 80 ℃, and CaC12 dosage of 2. 15 times the theoretical value, and the calcium molybdate- tungstate product contains 36.08% Mo and 7.11% WO3 ,with impurity of 0. 012% P and 0. 004 5% As. The new process has simple flowsheet,good operability and high molybdenum-tungsten recovery.

With molybdenum concentrate leaching solution as the object of study, effective separation ofrhenium from molybdenum was achieved by ion exchange experiments. The results show that the adsorp- tion rate of rhenium is 99.97% and its saturated adsorption capacity is 43.66 g/kg,with 201# resin as the ion adsorbent and under the optimal conditions of leaching solution pH of 8. 5 and flow velocity of 20 mL/min. The desorption rate of rhenium is 99.99% with 9% NH4SCN as the strippant. 201~ resinhasgood regeneration performance, suitable for recycling.

P204-sulphonated kerosene system was used to extract A1,Fe and Mn impurities from lepidolite leaching solution to separate the main metal Li. Influences were investigated of phase ratio, extractant con- centration,initial water phase pH value and extraction time on A1,Fe and Mn extraction rates. The optimal extraction technical conditions were determined as follows：P204 concentration of 1.5 tool/L, phase ratio of 1/1 ,initial water phase pH value of 2.5 ,and extraction time of 18 min. Under the above conditions and by single-stage extraction, the extraction rates of AI,Fe and Mn are 99.1 %,99.3% and 99.0%, respectively, only with lithium loss of 6. 9%. The optimum separation are realized of aluminum, iron and manganese from lithium.

MoO3 powders were reduced by two stages with graphite powders as the carbon source. The re- duction products at different stages were characterized by XRD,FESFM and EDS,and then a preliminary discussion was made on the possible reaction mechanism during carbon reduction. The results show that Mo4O11 may be the transitional phase during the carbon reduction of MoO3 to MoO2. MoO3 can be com- pletely reduced with carbon by reaction for 1 h both at 650 ℃ and 1 250 ℃ to obtain metallic molybdenum of high crystallinity and uniform particle size,containing a small amount of a-Mo2 C byproduct.

Using waste tantalum as the raw material and high-carbon iron as the carbon carrier, the fine grained TaC was produced by solid-liquid diffusion reaction during smelting combined with acid leaching. The products were characterized by XRD and SEM. The results show that carbon atoms react with tanta lure by diffusion during smelting and heat preservation, to form TaC grains in a-Fe matrix. After acid leac hing,shattering and ball milling of the TaC/Fe complex,the fine-grained TaC can be obtained with a parti cle size of about μm.

The ultrafine silver powders were prepared from AgNO3 by chemical liquid-phase reduction,with ascorbic acid as the reducing agent and carboxymethyl cellulose as the dispersant. The effects of feeding way,dispersant dosage,pH value and reaction temperature on silver powder properties were emphatically researched. The results show that feeding way and dispersant dosage directly affect the dispersity of silver powders. As pH value increases,the reaction particle size of silver powders becomes smaller while the dis- persity decreases. The reaction temperature shows a significant influence on the properties of silver pow- ders. Spherical,sphere-like and dendritic silver powders are respectively obtained at 30 ℃, 35~45 ℃ and 50 ℃. Meanwhile,the dispersity and tap density of silver powders decrease when the temperature increa- ses. The optimized preparation conditions are：the feeding way of adding AgNOa solution into reducing so- lution,m（CMC）/m（C6 He O6） value of 0. 015, pH value of 2 and reaction temperature of 30℃. Under the conditions,the highly dispersed ultrafine silver powders,with a tap density of 4.6 g/mL and a mean parti- cle size of 3.95 μm,can be prepared to meet the requirements of silver paste on solar cell front.