A, calcined soda, and baking soda object to flooding and flooding is in addition to tungsten. To this end, a soda amount of 1.25 times the theoretical amount was added to the tin slag, and a small amount of charcoal was added for grinding. It is then calcined in a rotary kiln, controlled at a temperature of 850 to 900 ° C, and calcined for 45 minutes. The calcined slag is wet- milled in a ball mill, then immersed in water, and the solid-liquid ratio is 1:2.5-3, the temperature is 90 ° C, and the time is one hour. After clarification, the mixture was filtered, and the synthetic white tungsten was recovered from the filtrate. Residue containing sodium and sodium tantalum niobium send the next processing step. Soda roasting, the de-tungstening efficiency of water immersion is 56%, and some silicon, arsenic , phosphorus , aluminum , tin, etc. can be removed.
Second, dilute acid and de-desilication tin residues were present in most of the sodium form of silicon, silicon is off with dilute hydrochloric acid (7 to 95%) leaching the sodium silicate into a silicic acid into solution. The solid-liquid ratio of the leaching is 1:6, and the mixture is filtered immediately after stirring for 1 to 3 minutes. The desiliconization string can reach 60 to 70%, and the load sawing grade in the filter residue can be 2.5 to 3 times. The de-siliconized filter residue is further leached with jade 2 to 14% hydrochloric acid, and stirred at a solid-liquid ratio of 1:6 at a temperature of 80 to 90 ° C for 2 hours to remove tin, so that tin enters the solution in the form of SnCl 4 and removes part. Impurities such as iron and calcium. After filtration, the filter cake that is enriched tantalum and niobium thereof. If the enriched material is too low and contains high levels of impurities, it must be further treated with alkali and acid leaching. The enrichment components of the above raw materials are listed in Table 2. The recovery rate of the above-mentioned enrichment stage is 71.5 to 85.5%; the impurity efficiency is 42 to 70% of tungsten and 50 to 70% of tin.
Third, the use of hydrofluoric acid decomposition is decomposition of tantalum, niobium oxide can be dissolved in hydrofluoric acid generated and the properties of H 2 TaF 7 H 2 NbF 7, according to the solid-liquid ratio of 1: 2.5 was added 6-7 equivalents of hydrofluoric acid to In the reaction pot, then slowly add the concentrate, stir and leaching for 2 hours, add sulfuric acid (5-6 equivalents) to adjust the acidity, the decomposed solution contains lanthanum, cerium oxide 130-150 g / liter, and a small amount of tungsten, tin and other impurities, rare earth elements and the residue was precipitated as alkali metal and other insoluble fluoride. The decomposition efficiency of lanthanum and cerium is relatively complete, and the recovery rate can reach 98 to 99%.
Fourth, tantalum, niobium separating the extraction-containing tantalum and niobium was treated with tributyl phosphate (TBP) as extractant, separating tantalum and niobium extraction. For this purpose, in a box extractor, after 10 stages of extraction, 12 stages of pickling; then, after stripping, the stripping enthalpy is level 7, and the stripping enthalpy is level 14, and finally the sputum and sputum are obtained in volume. The organic extractant is returned for use and the raffinate is discarded. The separated liquid tantalum and niobium were added liquid ammonia (pH≥9) and borax precipitate was then immersed in addition to tungsten, tin, iron and other impurities by alkaline and acid leaching. The produced barium hydroxide and barium hydroxide are filtered, washed, dried (or calcined) to produce barium hydroxide and barium hydroxide (or barium oxide and alumina) products, the ingredients of which are listed in Table 3.
V. and smelting reduction of potassium tantalum fluoride is potassium or a fluoro niobium at 850 to 900 deg.] C under argon, sodium metal reduction using a reducing furnace, i.e., to obtain tantalum powder and niobium powder: K 2 TaF 7 + 5Na = Ta + 5NaF + 2KF It is also possible to reduce the cerium oxide by carbon or lanthanum lanthanum in a vacuum reduction furnace at 1750 to 1800 ° C to obtain a ruthenium. The ruthenium is hydrogenated, pulverized and dehydrogenated to obtain a ruthenium powder: The tantalum powder and niobium powder are vacuum-smelted in an electron beam melting furnace, a sintering furnace or an electric arc furnace to remove gaseous impurities and volatile non-metallic impurities for further purification, and then smelt into ingots. Six, tungsten and tin recovery flooding outputs sodium tungstate (containing WO 3 20 g / l) for the production of synthetic scheelite, sodium tungstate solution which often contain silicon, arsenic, phosphorus and other impurities, these impurities removed Is to use " magnesium ammonium purification method", that is, first add hydrochloric acid and ammonium chloride to desiliconize the solution, then add MgCl 2 to remove arsenic and phosphorus, so as to produce H 2 SiO 3 , Mg( NH 4 )PO 4 with low solubility. After precipitation with Mg(NH 4 )AsO 4 , after clear separation, the supernatant was heated to 80-90 ° C, and CaWO 4 was precipitated with a saturated CaCl 2 solution, which was filtered, washed, and dried to obtain artificial white tungsten.
Whether the solution of desiliconization or de-tinization of dilute acid contains different amounts of tin (6-12 g/L), it is necessary to add iron scraps in the reservoir to reduce tin, make SnCl 4 become SnCl 2 , and then use electricity. The cathode method can obtain cathode tin (75-85% Sn) as an intermediate product. The above process for recovering ruthenium, osmium and tungsten from the slag, the total recovery string of ruthenium and osmium is 65-85%, and the tungsten is 66-70%. ,
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