Using W,C,Co, VC and Cr₃C₂as raw materials, and compounding them according to WC-10 Co-0.9 VC-0.3 Cr₃C₂,WC-10 Co cemented carbide was prepared by plasma ball milling and one-step carbonization-sintering method.The test results show that the average grain size of WC-10 Co cemented carbide is 0.25 μm, the density is 99.2% and the hardness is 92.3 HRA.The friction coefficient of WC-10 Co cemented carbide decreases with the increase of load, and the average friction coefficient is within 0.48~0.57.Compared with commercial YG8 cemented carbide, WC-10 Co cemented carbide prepared by one-step carbonization-sintering method has shorter running time, lower friction coefficient and higher wear resistance.The simulation results show that the maximum Hamilton contact stress caused by grinding ball effect on the subsurface of the WC-10 Co cemented carbide prepared by one-step carbonization-sintering method is 1 563.9 MPa, which is far lower than 1 847.1 MPa of the commercial YG8 cemented carbide.

Using cubic boron nitride(CBN),tungsten carbide and cobalt powder as experimental raw mate-rials, after high temperature vacuum treatment, PCBN composites with different WC-Co contents were synthesized by high temperature and ultra-high pressure sintering on a six-sided top press.The phase, compactness, hardness, abrasion ratio, impact toughness and micro morphology of the PCBN composites were analyzed.The results show that in the experimental range, when the WC-Co content is 11%,the bonding force between CBN and the binder is weak, the internal structure of the PCBN composite is relatively loose, the porosity of the sample is the highest, and the wear ratio and microhardness reach the maximum, 7 200 and 35.6 GPa, respectively.When the WC-Co content is 20%,the porosity on the surface of the PCBN composite is significantly reduced, the internal structure is compact, the porosity is the lowest, the density is significantly improved, and the flexural strength and impact times reach the maximum, 750 MPa and 80 times, which are 22% and 45% higher than those of the PCBN composite with 11% WC-Co content, respectively.

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The method of sulfation roasting-selective decomposition-water leaching was used to recover rare earth from CeFeB waste,and meanwhile the reaction mechanism of CeO₂ and Ce valence change in the re-covery process were investigated.The results show that the optimal recovery process conditions are sulfuric acid concentration of 14.5mol/L,sulfuric acid dosage of 2times the theoretical amount,sulfation roasting temperature and time of 300℃and 60min,and selective decomposition temperature and time of 700℃and 120min.Under the above optimized conditions,the rare earth leaching rate reaches 95.46%,the impurity Fe enters into the filter residue in the form of Fe₂O₃,and the rare earth oxide Nd_0.20Ce_0.80O_1.90is ob-tained after the leaching solution is precipitated with oxalic acid and the precipitate is roasted.The products of sulfation roasting process are Ce(SO₄)₂,RE₂(SO₄)₃and Fe₂(SO₄)₃.Ce₂(SO₄)₃,Fe₂O₃and indecompo-sable other RE₂(SO₄)₃are obtained in the selective decomposition process,and the decomposition product SO₃further reacts with unreacted CeO₂to form Ce₂(SO₄)₃.During the water leaching process,the water-soluble Ce₂(SO₄)₃and RE₂(SO₄)₃dissolve into the leachate,and the water-insoluble Fe₂O₃enters into the filter residue,thereby realizing the effective separation and recovery of rare earth from CeFeB waste.

Titanium dioxide waste water is a kind of acidic waste water containing a lot of iron ions, which has important recovery value.Using titanium dioxide waste water as raw material, ferrous oxalate was prepared by adding sodium oxalate.The effects of sodium oxalate dosage, reaction temperature, stirring strength and reaction time on the precipitation rate of iron were analyzed.The phase and microstructure of the precipitated products were analyzed by XRD and SEM-EDS.The results show that the precipitation rate of iron is more than 82% under the optimized conditions of titanium dioxide waste water volume of 50 mL,temperature of 20 ℃,sodium oxalate dosage of 12 g, stirring strength of 300 r/min and reaction time of 30 min, and the precipitate is the granular crystal of ferrous oxalate dihydrate with a purity of more than 98%.This study provides a method for efficient recovery of iron resources in acidic solution system.

The purification process and mechanism of hafnium metal by iodination-electron beam smelting were studied by thermodynamic analysis.The results show that when hafnium is purified by iodization, Zr and Ti cannot be separated because the binding ability of Zr, Ti and Hf to iodine is very close.When hafnium is purified by electron beam smelting, the vapor pressure of W,Nb, Zr and Ti is lower than or close to that of Hf, which makes them impossible to separate.The decomposition pressure of HfN near the melting point of Hf is very low, which is not easy to decompose and cannot be separated.In general, after metal hafnium iodination-electron beam smelting treatment, except for Zr and Ti which cannot be separated, other elements with relatively high content can be effectively separated.The actual production data further proves the correctness of the theoretical analysis of hafnium iodination-electron beam smelting purification mechanism, and Fe, Cr, Al, N,O and Si can be effectively separated from Hf.

A method for recovering rubidium and cesium was proposed from rubidium-cesium alum, a by-product of lithium extraction from lepidolite.The rubidium-cesium alum was leached with sodium hydro-xide, and the effects of sodium hydroxide concentration, liquid-solid ratio and leaching time on the leaching rate of rubidium and cesium were investigated.The results show that the leaching rate of rubidium-cesium is above 99% under the optimized leaching conditions of leaching temperature 25 ℃,NaOH solution concentration 1.5 mol/L,liquid to solid ratio 8∶1,and leaching time 40 min.Rubidium-cesium was extracted from alkaline leaching solution with t-BAMBP,and the four-stage countercurrent extraction was carried out under the conditions of temperature 25 ℃,pH=13.75,C_t-BAMBP=0.6 mol/L,O/A=1.5∶1 and extraction time 5 min.The extraction rates of rubidium and cesium are respectively over 92% and 99.9%,only 1.53% K impurity is extracted, and the elements of Li, Na, Al are not extracted.The loaded organic phase is washed and selectively stripped for rubidium and cesium, and the obtained rubidium chloride and cesium chloride solutions have purity of more than 92% and 98%,respectively.The rubidium in the washing solution is recovered and enriched, and the rubidium chloride solution with a concentration of 21 g/L and a purity of more than 98% is obtained.

Rhenium is mainly associated with copper-molybdenum ore.Rhenium enters the contaminated acid with flue gas during copper smelting.Due to the complex composition of the contaminated acid and its low rhenium content, there are few production examples of extracting and separating rhenium from copper smelting contaminated acid at present.The research progress is reviewed of extraction and separation technology of rhenium from copper smelting contaminated acid at home and abroad in recent years, and the development trend of rhenium extraction technology is prospected.