PROCESS FOR THE PRODUCTION TITANIUM CONCENTRATE HAVING A CHEMICAL COMPOSITION SIMILAR TO ILMENITE FROM HIGHLY IMPURE ANATASE ORES

Abstract

The present invention is related to a process for concentrating highly impure titanium ore, such as anatase, with the main purpose of obtaining a final concentrate having a chemical composition similar to that of ilmenite. The proposed process is based on the utilization as best as possible of the iron contents of the raw ore, in such a way that the final concentrate obtained can be used as an intermediate raw material for the production of titanium slag.

Description

DESCRIPTION OF THE INVENTION

[0001] The present invention discloses an innovative process for producing a titanium concentrate having contents of titanium and iron close to those usually found in ilmenite concentrates. Said material, herein called “artificial ilmenite”, is obtained by applying pyro and hydromerallurgical processes to the treatment of highly impure titanium ore, a class into which can be found in Brazilian ores, the predominant titanium source of which is anatase.

[0002] The processes for the production of titanium and its compounds are schematically shown in FIG. 1. Presently, there are only two ore sources commercially exploited around the world: rutile (TiO2) and ilmenite (FeTiO3). Several alternatives involving physical and/or chemical beneficiation processes for concentrating such ores have been proposed (HENRY, J. L. et all., “Bureau of Mines Development of Titanium Production Technology”, Bulletin 690, U.S. Bureau of Mines, 1987). The products obtained this way have essentially two destinations: direct production of metallic titanium and production of titanium pigment. The latter is presently carried out through the so-called sulphate, and chloride processes, wherein the latter process is the main one and that with the highest growing potential.

[0003] Among the items shown in FIG. 1, the production of titanium slag (85% TiO2 equivalent) is of utmost importance due to the fact that this material can be directly used in the two presently available routes for the production of TlO2 pigment, by far the main application of titanium based raw materials. The following advantages can be mentioned for the production of slag: possibility of using titanium ore with a lower impurity content (ilmenite), utilization of the iron content of said ore through the production of pig iron, in addition to the already mentioned advantage of producing a material which can employed in the existing processes for the production of pigment. Estimates for the titanium slag consumption by the pigment industry show a present increase of up to 600,000 metric tons until year 2000 (ANON., “Titanium Minerals 1994—A Complete Cost Analysis”, AME Mineral Economics, 1994).

[0004] Brazil possesses approximately 20% of the known titanium ore reserves and the main concentration thereof can be found in alkaline pipes in the western region of Minas Gerais State and south of Goias State. The ore, that covers phosphate beds, is of complex nature, being the result of alterations in different types of rocks. The main difficulty in developing a process for recovering titanium amounts from this type of ore is the selection of single operations that show high efficiency and selectivity simultaneously. As a consequence, when trying so produce a concentrate rich in titanium the overall mass recovery falls dramatically, rendering the process uneconomical. Treatments of such kind have been tried in the past (PIXAO, J. M. J. and MENDONCA, P. A. F., “Process for Concentrating Titanium Ores”, Brazilian patent PI 7507645; PAIXAO, J. M. J. and MAGALHAES, G., “Process for Beneficiating Titanium Ores”, Brazilian patent PI 7604532), resulting in alternatives of limited practical use, since mass recoveries have been comprised within the 5 to 10% range only.

[0005] The process here presented has been developed with the main purpose of overcoming the severe drawbacks found in the processes avaliable to date for the treatment of Brazilian anatase ores. As shown later; the solution for such a problem is obtained through by the utilization of iron amounts usually associated to titanium in the several anatase deposits found in Brazil.

[0006] The objective of the present process is the recovery of a titanium concentrate having a chemical composition similar to that of ilmenite; this concentrate being the intermediate raw material for the production of titanium slag. On the other hand, the process for producing slag is based on the carbothermic reduction of a part of the titanium contained in the ore—leading to a material with a stoichiometry close to Ti3O5—and all the iron from the concentrate to FeO—contained in the slag phase—and metallic Fe. Other elements such as aluminum, silicon, alkaline and alkaline-earth metals are not reduced and thus constitute deleterious impurities that must be removed from the titanium raw material. Phosphorus is also an undesirable impurity due to the contamination caused in the resulting pig iron. As can be seen hereinbelow, the process developed makes it possible to obtain a high degree of removal of such impurities, thus providing concomitantly a large recovery of titanium and iron contained in the raw ore.

[0007] The process described in the present invention is based on the use of the following sequence of operations to the processing of titanium ores: disintegration, screening through 6 mm, crushing, classification, low intensity magnetic separation for removing coarse magnetite, attrition and slimes removal for discarding the minus 74 &mgr;m (200 mesh) fraction, calcination in the presence of sodium carbonate (Na2CO3) and/or potassium carbonate (K2CO3), comminution of the calcination product to a particle size below 1 mm, leaching in alkaline medium (pH in excess of 10) and later acid leaching. Each of such steps is detailed below.

[0008] First of all, the raw ore is subjected to a disintegration step in a drum washer, followed by screening in a vibrating screen having a 6 mm (¼″) opening, and crushing in a conical crusher. The screen operates in closed loop with the crusher, that is, only the >6 mm fraction is crushed. The screened material is classified in a spiral classifier which is adjusted for providing a product with a particle size above 48 mesh (300 &mgr;m). The overflow of the classifier with a particle size of −48 mesh is discarded.

[0009] The classLfied ore is then subjected to a low intensity magnetic separation (approximately 800 Gauss), preferably in a wet drum separator. The magnetic fraction is discarded. The non magnetic fraction undergoes a slimes removal operation in an attrition cell with the purpose of removing clay-minerals particles that remain adhered to the surface of the anatase grains. The ore from the attrition step is classified in a spiral classifier adjusted in such a way that a 200 mesh (74 &mgr;m) cut can be obtained. The overflow of the classifier, with a particle size of −200 mesh, is discarded. For the purposes of the present invention, the material subjected to the above mentioned sequence of operations that is, disintegration, screening, crushing, classification, magnetic separation, attrition and slimes removal, is called mechanical concentrate.

[0010] In the calcination step, the mechanical concentrate is mixed with sodium carbonate (Na2CO3) and/or potassium carbonate (K2CO 3) and heated to a temperature within the 900 to 1300° C. range, preferably 1100° C., being kept at the desired temperature for a time period between 20 to 60 minutes, preferably 30 minutes, in an oxidizing environment. The amount of carbonate to be added, either as powder or in aqueous solution, depends on the content of Al2O3 and SiO2, ranging from 60 to 150 kg per metric ton of ore to be treated. The main purpose of this calcination is to produce soluble sodium compounds (Na2O.Al2O3, Na3O.P2O5, Na2O.SiO2 among others) at elevated temperatures, which are removed in the leaching steps that follow. No agglomeration or sintering was observed in the ore during calcination. Therefore, a simple disintegration of the calcined product is reqgired in order to match its grain size with the next leaching step. In laboratory scale, this is carried out by compressing the calcined product in a rubber mantle, whereas in industrial scale it can be carried out through the use of conventional comminution equipment (hammer or jaw crusher) and care should be taken when regulating the operation of said equipment in order to avoid the generation of excessive fines. The calcined disintegrated product with a particle size below 1 mm feeds the leaching operations described hereinbelow.

[0011] The first leaching step is performed in alkaline medium (pH >10) by using as solvent any alkaline substance such as NaOH, KOH, Na2CO3, or K2CO3. In this step, the temperature is in the 60 to 90° C. range, preferably 70° C., and the time is 1 hour. After leaching, the material is filtered, and the liquor is reserved for the production of Al2O3, P2O3 or SiO2 and the recovery of the alkaline compound used. The solid residue of this step feeds the next step, the acid leaching.

[0012] The product of the alkaline leaching is then subjected to leaching with HCl or H2SO4, preferably HCl, at concentrations ranging from 30 to 50 g of HCl per liter of solution. The temperature is in the 60 to 90° C. range, preferably 70° C., for a period of 20 to 240 minutes, preferably 60 minutes. After leaching, the pulp is filtered and the liquor is used in the recovery of rare earth compounds contained in the initial ore, and the solid, after drying, constitutes the final product, which is called artificial ilmenite. For the purposes of the present invention, the sequence of operations including calcination, disintegration, alkaline leaching and acid leaching shall be referred to as thermo-chemical treatment.

[0013] The spirit and scope of the present invention may be fully understood based on the examples given below. It should be noticed that said examples are merely illustrative and shall not limit the process developed. The main steps of the process are shown in the schematic representation of FIG. 2.

EXAMPLE 1

[0014] An anatase ore sample was subjected to the mechanical processing steps with the exception of the attrition and slime removal operations. That is, its initial processing included disintegration, screening, crushing, classification and magnet to separation. Said mechanical concentrate was then calcined to 1100° C. for 1 hour in the presence of soda ash (Na2CO3) by using an amount of 95 kg of carbonate per metric ton of ore, corresponding to the sodium required for reacting switch all aluminum, silicon and phosphorus contained in the mechanical concentrate. After disintegration, the calcined product underwent leaching with caustic soda (25 g/l NaOH) at 70° C. and 35% solids for 1 hour, followed by leaching with hydrochloric acid (30 g/l HCl), also at 70° C., 35% solids and 1 hour The material from the leachings, after separation of the liquor through filtration and drying, is the final product. The chemical compositions of the raw ore, mechanical concentrate and final product are given in Table 1, where are also shown the weights of each of said materials along the processing. It can be seen that the proposed process leads to high overall recovery of titanium and iron, 100% and 94% (compared to the mechanically treated product), respectively, with a high extraction of the main contaminants. However, it is possible to further increase the removal of impurities through the inclusion of attrition and slimes removal operations at the end of the mechanical treatment, thus bringing forth a higher efficiency in the thermo-chemical treatment, as shown in Example 2. 1 TABLE 1 Example 1 - contents (% by weight) of main constituents in the ore in different steps of the concentration process weight, kg (1) (2) (3) 100 30.7 26.9 TiO2 26.6 40.5 46.5 Fe2O3 43.2 37.7 40.5 Al2O3 8.72 6.39 5.08 P2O5 4.56 4.42 1.21 SiO2 3.14 2.16 0.83 CaO 0.44 0.84 0.29 BaO 0.77 0.89 0.19 MgO 0.85 0.73 0.73 MnO N.A. 0.68 0.38 Others 11.72 5.69 4.29 (1) - raw ore (2) - concentrate after magnetic separation (3) - final concentrate (after thermo-chemical treatment) N.A. - not available

EXAMPLE 2

[0015] The same anatase ore sample was subjected to an identical sequence of mechanical processing operations described in Example 1, including attrition and slimes removal steps. Next, it was calcined at 1000° C. for 1 hour in the presence of soda ash (Na2CO6), by using an amount of 80 kg of carbonate per metric ton of ore. This amount of soda ash is lower than that of the previous example, taking unto account that the attrition and slime removal operations provide a good removal of impurities, such as Al, P and Si. By disintegration, the particle size of the calcination product was adjusted for the next step of acid leaching carried out under the following conditions: HCl (50 g/l), temperature=70° C. and time=60 minutes. The material from leaching, after separation of the liquor through filtration and drying, is the final product. As shown in Table 2, a concentrate rich in titanium and iron is obtained, with Ti recovery of 92% and Fe recovery of 84%. The beneficial effect of the use of attrition and slimes removal steps at the end of the mechanical treatment is made evident by the low content of the main contaminants given in Table 2. Based on these values, the following extraction degrees of said contaminants are calculated: 91% for barium, 84% for phosphorus, 83% for calcium, 75% for silicon and 51% for aluminum. The concentrate obtained in the present invention, with more than 50% TiO2, presents a chemical composition comparable to that of some ilmenites available in the international market, as shown in Table 3. 2 TABLE 2 Example 2 - contents (% by weight) of the main constituents of the ore in different concentration process steps weight, kg (1) (2) (3) (4) 100 30.7 27.3 21.7 TiO2 26.6 40.5 43.5 50.4 Fe2O3 43.2 37.7 38.9 41.1 Al2O3 8.72 6.39 5.11 3.96 P2O3 4.56 4.42 3.29 0.88 SiO2 3.14 2.16 1.38 0.68 CaO 0.44 0.84 0.75 0.18 BaO 0.77 0.89 0.96 0.10 MgO 0.85 0.73 0.72 0.71 MnO N.A. 0.68 0.76 0.51 Others 11.72 5.69 4.67 1.49 (1) - raw ore (2) - concentrate after the magnetic separation (3) - concentrate after the attrition and mud removal (4) - final concentrate (after the thermo-chemical treatment) N.A. - not available

[0016] 3 TABLE 3 Chemical compositions of some titanium concentrates “Orissa” “R. Bay” Present Ilmenite Ilmenite Compound Invention (India) (South Africa) TiO2 50.4 49.7 49.2 Fe2O3 41.1 49.4 51.0 SiO2 0.68 0.95 0.58 CaO 0.18 0.07 0.02 MgO 0.71 0.76 0.52 MnO 0.51 0.61 1.0

[0017] The use of the concentrate provided by the present invention in the process of titanium slag manufacture can be assessed by means of a calculation based on a computer program. In such calculation, it is assumed that all titanium in the concentrate is incorporated to the slag and that all oxygen removed in the reduction reactions combines with carbon, producing gaseous carbon monoxide (CO). The results of the calculation are given in Table 4, wherein the composition of a slag produced in industrial scale can also be found. It can be seen that both materials show similar compositions, indicating that the concentrate obtained through the present invention is a suitable raw material for the production of titanium slag. 4 TABLE 4 Composition of the slag obtained by smelting the concentrate of the present invention (simulation) and that of a commercially available slag % BY WEIGHT Compound Present Invention “Richards Bay” Slag TiO2 (1) 82.5 85.0 Ti2O3 (2) 51.1 32.7 FeO 10.0 10.0 P2O5 0.14 0.01 SiO2 1.1 1.6 CaO 0.29 0.20 MgO 1.16 1.10 MnO 0.83 1.7 (1) - total titanium expressed as TiO2 (2) - trivalent titanium expressed as TiO2

Claims

1. A process for the production of titanium concentrate with a chemical composition similar to that of ilmenite from highly impure anatase ores, based on the use of the following sequence of operations for the treatment of titanium ores wherein the predominant titanium source is anatase: disintegration, screening through 6 mm, crushing, classification, low intensity magnetic separation for removing coarse magnetite, attrition and slimes removal for discarding the minus 74 &mgr;m (200 mesh) fraction, calcination in the presence of sodium carbonate (Na2CO3) and/or potassium carbonate (K2CO3), disintegration of the calcination product to a particle size below 1 mm, leaching in alkaline medium (pH in excess of 10) and acid leaching.

2. The process for the production of titanium concentrate with a chemical composition similar to that of ilmenite from highly impure anatase ores according to

claim 1, wherein the calcination in the presence of sodium carbonate and/or potassium carbonate of the disintegrated ore, screened through 6 mm, subjected to magnetic separation, attrition and slimes is carried out in a temperature range of 900 to 1300° C., preferably 1100° C., in an oxidizing environment, for a period of 20 to 60 minutes. preferably 30 minutes, with an amount of 60 to 150 kg of carbonate per metric ton of dry ore.

3. The process for the production of titanium concentrate with a chemical composition similar to that of ilmenite from highly impure anatase ores according to claims 1 and 2, wherein the alkaline leaching of the disintegrated ore, screened through 6 mm, crushed, subjected to magnetic separation, attrition and slimes removal, calcined in the presence of Na2CO3 an/or K2CO3 and having a particle size below 1 mm is carried out at a concentration of alkaline solvent (NaOh, KOH, Na2CO3 or K2CO3) lower than 50 g/l, for a time of 60 minutes and at a temperature in the 60 to 90° C. range, preferably 70° C.

4. The process for the production of titanium concentrate with a chemical composition similar to that of ilmenite from highly impure anatase ores according to

claim 1,

2 and 3, wherein the acid leaching of the disintegrated ore, screened through 6 mm, crushed, subjected to magnetic separation, attrition and shines removal, calcined in the presence of Na2CO3 and/or K2CO3 and having a particle size below 1 mm and subjected to alkaline leaching is carried out with hydrochloric acid (HCl) or sulphuric acid (H2SO4) at a concentration lower than 50 g/l for a period of 20 to 240 minutes, preferably 60 minutes, and at a temperature in the 60 to 90° C. range, preferably 70° C.