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Antonio Baker pyrochlore beneficiation plant
Antonio Baker (Niobec) Company owned St. Huo Nuolei niobium mine in Quebec City, northeast Xikutimi 11 kilometers St. Huo Nuolei village. A monocarbonate antimony deposit. The ore contains niobium (Nb 2 O 5 ) from 0.58% to 0.66%. Niobium is mainly mineral pyrochlore, niobium iron ore, grain particles less than 0.2 mm. Gangue minerals are mainly calcite, white mica. The mine is one of the world's major production bases for tantalum raw materials.
The size of the ore dressing plant is 2085 tons/day. Two-stage flotation process of grinding (FIG. 1), the whole process is divided desliming flotation carbonate, then desliming, magnetic separation, flotation pyrochlore, pyrite flotation, pyrochlore concentrate leaching and dephosphorization Leaching slag floats eight parts of sulfur. The ore is first crushed to -20 mm by Alice Chalmers (AC) cone crusher , and the Harding 2.4×3.75 m rod mill is ground to 0.4 mm, and the first DSM arc is screened (screen 0.4 Millimeter), the material under the sieve is fed into the second DSM curved screen (0.2 mm), and the material of the first and second sieves is fed into a 3.2×3.8 m ball mill by a spiral classifier with a grinding particle size of 95%- 0.2 mm, with DSM curved screen constitutes a closed circuit. The second arcuate sieve material under DSM (-0.2 mm) into a set of φ254 mm Krebs cyclone, the ore concentration of 18%, a pressure of 1.2 to 1.3 kg / cm 2. The cyclone overflows into another set of φ102 mm Krebs cyclones (pressure 1.8 kg/cm 2 ) and removes -10 μm fine mud. The two-stage cyclone maintains the stability of the ore by means of a return overflow circulation. The underflow of the cyclone (concentration is about 55%), and the +40 micron material is sent to the coarse grade carbonate flotation circuit. The -40 to +10 micron material is sent to a fine grade carbonate flotation circuit.
Figure 1 Niobek Yellow Chlorite Concentrator Process
Carbonate flotation
Emulsified fatty acid is used as collector , sodium silicate is used as inhibitor and softener of pyrochlorite, and 25%-30% of carbonate is floated under the condition of natural pH of about 8, wherein the lost yellow chlorite accounts for 2% of total 5%. The tailings of the two carbonate flotation sections were removed from the -10 micron fine mud by two sets of cyclones φ254 mm and φ100 mm, and the hard water was replaced by high-quality drinking water, so that the total salt content was greatly reduced. The cyclone grit is selected from two series of drum-shaped Eriez magnetic separators to select magnetite. Non-magnetic materials are sent to the glaucon flotation section for treatment.
Chlorophyll flotation
The non-magnetic material is flocculated with oxalic acid, fluorosilicic acid and emulsified fatty diamine acetate under conditions of pH 6.8 to 7.5. After five times of selection, the limestone coarse concentrate was added with oxalic acid and fluorosilicic acid to adjust the pH value of the slurry. The pH of the selected one to five times was 5.5, 4.5, 3.5, 3.0, and 2.7. The average enrichment ratio is 1.9 times each time. The concentrate contains Nb 2 O 5 45% to 50%, and also contains some pyrite, carbonate, apatite and other minerals. The recovery rate is generally 80%. Desulfurization and dephosphorization are required to obtain the product. Chlorite concentrate.
Pyrite flotation
The foam product of the Chlorophyll flotation circuit is adjusted with NaOH to adjust the pH of the slurry to 11, the cassava starch is added to inhibit the yellow chlorite, and the potassium pentyl sulphate is floated to the pyrite. After a rough selection and secondary selection, 95% of the yellow iron can be obtained. The mine floated out. The product in the tank (the pyrite concentrate contains less than 0.1% sulfur) is sent to the leaching operation for further processing.
Leaching
The product in the flotation tank of the pyrite is fed into the 3.657-meter concentrator for concentration. The concentrator sand (concentration: 55%) is sent to the 1524×2438 mm leaching tank, and HCl (1.816 kg/ton) is added for leaching to make part of the ash. The stone dissolves. The leaching residue is sent to another concentrator to remove the solution and washed with a small amount of fresh water. The grit is given to the leaching tank of the second leaching operation, and 227 g/ton of HCl is added, and the pH is 0.5. After two consecutive leaching, the phosphorus content can be reduced to less than 0.1%. The leaching slag is sent to the second section of the pyrite flotation process.
Leaching slag
The leaching slag is fed into the filter, the filter cake is sent to the agitation tank for slurry adjustment, CuSO 4 is added to activate a small amount of remaining sulfide, and the pH is adjusted to 10.5 by adding NaOH, and the sulfide is floated by the addition of xanthate (PAX). The product in the tank (vitite final concentrate) contains Nb 2 O 5 60% to 62%, S0.1%, P 2 O 5 0.07%, and SiO 2 2.02%. This process is very effective in treating carbonate glauconite ore and has obtained a US patent.
Araxa yellow-green water barium strontium ore concentrator
The Araksa Mine is located in Minas Gerais, south of Araxa, Brazil. A carbonate rock composite deposit. Ore mainly composed of goethite, limonite, magnetite, barite, monazite, titanium, iron, gorceixite, quartz and barium strontium pyrochlore water (i.e., barium pyrochlore containing Nb 2 O 5 63.42%, Mineral composition such as Ta 2 O 5 0.15%, BaO 16.51%). The ore contains Nb 2 O 5 2.5% to 3%, Re 2 O 3 4.44%, ThO 2 0.13%, TiO 2 3.6%, and ZrO 2 0.2%. The leeches of the chlorite are finer and generally do not exceed 1 mm. The reserves of the mine (Nb 2 O 5 ) account for more than 70% of the world's resources, and are the world's largest resource base. The current cut-off grade (Nb 2 O 5 ) is controlled at around 2%. What is being mined is an enriched ore body located in the central part of the structure, which is formed by deep weathering of carbonate bedrock and strong residual enrichment. The ore dressing grade (Nb 2 O 5 ) is 3%, and the combined process of smelting and smelting is used to produce yttrium oxide and lanthanum iron. The whole process consists of three parts: mineral processing, leaching and smelting.
Dressing
The size of the ore dressing plant is 3,500 tons/day, and the annual output of the glauconite concentrate is 42,000 tons. The process of mineral processing equipment is shown in Figure 2. The ore is first passed through a 50 mm 2.1 x 4.2 m Faco (AC) vibrating screen, and the material on the screen (greater than 50 mm, the yield is less than 10% of the ore) is given to the Hazemag impact crusher . Broken to -50 mm, combined with vibrating screen material (less than 50 mm) into the ball mill, grinding to 95%-150 mesh (0.104 mm), ball mill and φ508 mm cyclone constitute closed circuit. The cyclone overflow is selected by a double cylinder weak magnetic separator. Non-magnetic products are sent to the de-sludge section. Because the slime has a great influence on the flotation of the chlorite, and most of the strontium is enriched in the 37-5 micron size, effectively removing the -5 micron fine mud is very important for the flotation operation. Therefore, the three-stage cyclone is used for de-sludge, and each de-sludge section has a de-sludge operation. The first stage of desilting uses a set of φ381 mm cyclones, the underflow of the cyclone is fed into the scrubber and scrubbed, and another set of φ102 mm cyclones is used for the second stage of de-sludge, the second stage of de-sludge and the first stage Scoured and de-sludged as well as mud. The φ381 mm de-drip cyclone underflow and the φ102 mm cyclone underflow are the ore feeding of the main flotation circuit rough selection tank. The φ102 mm cyclone overflows into the third-stage cyclone (φ25 mm) for de-sludge and de-sludge, but no need to scrub again. The φ25 mm cyclone underflow is the fine-grain flotation circuit rough selection tank. Give the mine.
Figure 2 Equipment flow of Araksa é’¡ Chlorite concentrator
Chlorophyll float is selected from amine cation collector, a wetting agent and sodium fluorosilicate for 15 minutes. Sodium fluorosilicate is used as actin activator. Hydrochloric acid is added to control pH 2.5-3.5. The main flotation circuit is thick. The tank is selected for rough selection and four times for selection. The coarse-grain flotation coarse selection tank is subjected to rough selection and one selection. Obtained flotation total concentrate Nb 2 O 5 55%～60%, P0.3%～0.8%, S0.02%～0.2%, Pb0.1%～0.2%, Fe 2 O 3 2%～8%, SiO 2 0.1% to 0.5%, ThO 2 2% to 3%, U 3 O 8 0.05% to 0.1%, BaO 15% to 18%, CaO0% to 0.2%, and the ore recovery rate is about 70%. Due to the high content of phosphorus, sulfur and lead , it needs to be sent to the factory for further processing.
Leaching
The leaching process uses high-temperature calcination of calcium chloride to volatilize lead chloride. Calcium replaces the ruthenium in the crystal lattice of the leeches, forming ruthenium chloride, and then washing with hydrochloric acid to dissolve phosphorus and sulfur. The leaching process is shown in Figure 3. The flotation concentrate is first filtered, and the filter cake is mixed with 25% calcium chloride and 5% lime, and calcined at a temperature of 800 to 900 ° C in a rotary kiln. After the calcined material was cooled, it was leached and filtered with 5% hydrochloric acid at a solid concentration of 50%. This step was carried out in two stages. The leaching residue is washed with water at a concentration of 50%. After filtration and drying, the obtained leached concentrate contains Nb 2 O 5 59%-65%, P0.05%-0.1%, S0.01%-0.05%, Pb0.01 %～0.05%, Fe 2 O 3 2%～8%, SiO 2 0.1%～0.5%, ThO 2 2%～3%, U 3 O 8 0.05%～0.1%, BaO1%～3%, CaO15%～ 20%.
Figure 3 Alasha mink yellow earth leaching process
The PbCl 2 and HCl discharged in the calcination are recovered through a cooling tower, a venturi, and a recovery tower. All the collected gas dust is sent to a 13.72m thickener, the concentrator is sanded into the filter, and the filter cake is returned to the starting point of the process; and the overflow containing the hydrochloric acid and lead chloride and all the waste water from the factory are passed through the milk. After processing, it is sent to the tailings pond.
Smelting
Smelting and smelting is divided into two parts: ferroniobium production and bismuth oxide production.
Niobium iron production: intermittent aluminum thermal reduction process. According to the pyrochlore leaching concentrate, the charge ratio accounted for 61.54%, iron oxide accounted for 13.68%, aluminum powder 20.51%, fluorite 2.56%, and stone 1.71%. Batch charge first mixed in a rotary mixer for 90 minutes and then a steel cylinder lined reactor was charged with magnesium sand, the reactor diameter of 3.7 meters, 1.8 meters high, is placed above the circular pits, with the pits Lime and fluorite are lined and prepared in advance in a silica sand bed. The bottom of the outer layer of the reactor is sealed with sand embankment. After the reaction material is installed, potassium chlorate and aluminum powder are placed on the top, and the mixture is ignited by fire. The reaction is automatically carried out for 15 minutes, the reaction temperature is up to 2400 ° C, and the reaction is cooled for 16 hours. The slag was removed, and the bottom ferroniobium was cooled for a few hours, taken out, placed in the air for 12 hours, and then crushed to less than 100 mm with a jaw crusher and stored in a 1000 kg steel vessel. The slag composition after reduction was: Al 2 O 3 48%, CaO 25%, TiO 2 4%, BaO 2 %, Re 2 O 3 4%, Nb 2 O 5 trace, ThO 2 2%, and U 3 O 8 0.05%.
Production of cerium oxide: In 1979, a factory for the production of cerium oxide was built. In 1981, the company researched a method (the method was not reported) to build a high-purity cerium oxide plant for optical and electronic applications. A pilot plant for the construction of ultrapure crystalline cerium oxide, while producing high-purity ferroniobium and nickel bismuth using industrial cerium oxide.