RU2786064C2 - Titanium dioxide production - Google Patents

Abstract

FIELD: chemical industry; environmental protection.

SUBSTANCE: invention relates to the chemical industry and environmental protection; it can be used for the production of pigment titanium dioxide. A titanium-containing residue from hydrochloric acid leaching of titanium-magnetite ore is subjected to autothermal boiling with concentrated sulfuric acid, at a ratio of 1:1.27 g/g, respectively, at a temperature of 175-190°C, for 3-4 h, then, it is subjected to leaching in diluted sulfuric acid, at a temperature of about 60°C, for 15 h. The resulting suspension is filtered for extraction of black liquor, which is then hydrolyzed and calcined to obtain pigment titanium dioxide. The extraction of titanium into black liquor is at least 98%.

EFFECT: invention allows for disposal of a side product of ore recycling – a titanium-containing residue from leaching to obtain a useful and demanded product – a pigment titanium dioxide, as well as for reduction in environmental pollution.

21 cl, 16 tbl, 3 dwg, 1 ex

Description

The technical field to which the invention belongs

[0001] The present invention relates to a method for producing titanium dioxide. More specifically, the titanium dioxide produced by the process of the present invention is for use in the production of titanium dioxide pigment.

[0002] Even more specifically, the proposed method uses a residue from the leaching of a titanomagnetite type ore as a feedstock.

State of the art

[0003] In the international patent application PCT/AU 2011/000519 (WO 2011/143689), filed by the present applicant, a new hydrometallurgical process for the extraction of vanadium from titanomagnetite ores is described. The process described in application PCT/AU 2011/000519 uses a combination of acid leaching, solvent extraction and separation to selectively recover valuable metals. PCT/AU 2011/000519 also describes a leaching feedstock containing some iron, said iron being extracted together with vanadium. The iron is extracted in the acid leaching step along with the vanadium, since the vanadium is contained in a titanomagnetite matrix. After that, iron, together with vanadium, is sent to the stages of solvent extraction and separation for subsequent removal.

[0004] In order to increase the overall extraction and recovery of vanadium, it is advantageous to minimize the amount of iron or any other low value material in the leaching feedstock. In addition, improving the quality of the leach feedstock minimizes operating costs and capital costs, as additional process steps to process significant amounts of iron after the leach step are substantially eliminated.

[0005] International Patent Application PCT/AU 2018/050310 (WO 2018/184067), also filed by the present Applicant, describes a process for preparing a leaching feedstock in which an ore or concentrate containing vanadium and iron is first fed into a reduction step for obtaining a reduced ore or concentrate, and then passing the reduced ore or concentrate to a ferric leaching step to obtain a ferric leaching product containing iron and a ferric leaching residue containing vanadium, wherein the ferric leaching residue is suitable for use as feedstock for leaching for extraction and recovery of vanadium.

[0006] International Patent Application PCT/AU 2018/050310 also describes feeding ferric leach residue to an acid leach step that produces both an acid leach containing vanadium and an acid leach containing titanium. The acid leaching step is carried out using hydrochloric acid (HCl) at a concentration of from about 15% to 32% by weight and preferably from about 15% to 20% by weight. The acid leaching step is carried out at atmospheric pressure and at a temperature in the range of approximately 25°C to 100°C. Preferably, the temperature is in the range from about 60°C to 80°C.

[0007] The processes described in both International Patent Applications PCT/AU 2011/000519 and PCT/AU 2018/050310 form a pressure leach residue that has a high titanium content. Applicants have found that such a leach residue can serve as a potential feedstock for a titanium pigment manufacturing plant. However, this leach residue is much finer than conventional feedstock and contains more silica, which can be expected to adversely affect the efficiency of a standard or known titanium pigment plant chlorinator.

[0008] One of the objectives of the method according to the present invention is, in essence, the solution of the aforementioned problems of the prior art, or at least the creation of a useful alternative.

[0009] Throughout the description, unless the context otherwise requires, the term "comprise" or variants thereof such as "comprises" or "comprising" should be understood to mean the inclusion of a specified integer or group of integers, but without the exclusion of any other integer numbers or groups of integers.

[00010] Throughout the description, unless the context otherwise requires, the term "comprise" or variations such as "comprises" or "comprising" should be understood to mean the inclusion of a specified integer or group of integers, but without excluding any other integer or group of integers.

[00011] Each document, reference, patent application or patent cited in this text is expressly incorporated herein in its entirety by reference, which means that it should be read and considered as part of this document. The fact that a document, reference, patent application or patent cited in this text is not repeated in it is for reasons of brevity only.

[00012] Reference to the material cited or the information contained therein should not be taken as an admission that the material or information was part of the common knowledge or was known in Australia or any other country at the time of filing this application, or any application, under which priority can be claimed.

SUMMARY OF THE INVENTION

[00013] In accordance with the present invention, a method for producing titanium dioxide is provided, which consists in the fact that:

(i) subjecting the titanium-containing leach residue to a pulping step with concentrated sulfuric acid, and

(ii) then subject this residue to leaching in dilute sulfuric acid to obtain,

black liquor, from which titanium dioxide is then obtained, and the residue from the leaching is the residue from the hydrochloric acid leaching of the titanomagnetite type ore.

[00014] Preferably, titanium dioxide is used to produce a titanium dioxide pigment.

[00015] Preferably, the concentration of acid HCl is:

a. from about 15% to 32 wt. %, or

b. from about 15% to 20 wt. %.

[00016] Preferably, the feedstock for leaching the titanomagnetite type ore is a ferric iron leachate. The ferric leaching step is preferably carried out with ferric chloride. Preferably, the ferric chloride concentration is:

a. from about 10 to 40 wt. %,

b. from about 25 to 35 wt. %, or

c. about 28% of the mass.

[00017] Preferably, the titanium-containing leach residue has a P ₈₀ :

a.≤125 µm,

b.≤45 µm, or

c. < 40 µm.

[00018] In one form of implementation of the present invention, the extraction of titanium in black liquor is at least 98%.

[00019] Preferably, the cooking step is carried out at a temperature of:

a. above 175°C or

b. around 190°C.

[00020] Preferably, the cooking step is carried out for a period of:

a. approximately 3 to 4 hours, or

b. about 3 hours.

[00021] Also preferably, the mixture of leach residue and concentrated sulfuric acid in the cooking step has a ratio of approximately 1:1.27 (g/g).

[00022] Preferably, cooking step (i) provides at least 1.9 grams of concentrated H ₂ SO ₄ for every gram of TiO ₂ in the titanium-containing leach residue.

[00023] Preferably, the dilute sulfuric acid in the leaching step (ii) is about 6% sulfuric acid.

[00024] Also preferably, the leaching step (ii) is carried out at a temperature of about 60°C.

[00025] Preferably, the leaching step (ii) is carried out for about 15 hours.

[00026] Even more preferably, the leaching step (ii) is carried out at about 20% solids.

[00027] Preferably, the black liquor is recovered from the slurry obtained in the leaching step (ii) by filtration and the solids are washed to recover the titanium.

[00028] In one embodiment of the present invention, the cooking step proceeds autothermally.

[00029] In another form of the present invention, the cooking step further includes an initial dilution of the acid. Preferably, the acid is diluted with water to about 88-92%.

Brief description of the drawings

[00030] The present invention will be described hereinafter, by way of example only, with reference to an embodiment thereof and the accompanying drawings, in which:

Fig. 1 is a graph of the start temperature of a test carried out in a large oven showing the heating rate of the sample versus the heating rate of the furnace reference thermocouple;

Figure 2 is a graph of the start temperature of a test run in a smaller forced-ventilation oven with a fast temperature response, also showing the rate of sample heating versus the rate of heating of the furnace reference thermocouple;

3 is a differential scanning calorimetry plot of ilmenite and pressure leach residue showing exothermic reactions occurring therein when mixed with acid.

Preferred embodiment of the invention

[00031] International patent application PCT/AU 2018/050310 (WO 2018/184067), the entire content of which is incorporated herein by reference, describes a process for producing a leaching feedstock, in which:

submitting an ore or concentrate containing vanadium and iron to a reduction step to obtain a reduced ore or concentrate; and

the reduced ore or concentrate is fed to the ferric iron leaching stage to obtain an alkaline solution of ferric iron containing iron and a residue from ferric leaching containing vanadium,

wherein the ferric leaching residue is suitable for use as a leaching feedstock for the extraction and recovery of vanadium. In one form of implementation of this invention, the ore or concentrate, in addition to vanadium and iron, also contains titanium.

[00032] The reduction step is preferably carried out using a carbon reducing agent. Preferably, coke is used as the carbon reducing agent. More preferably, the concentration of coke, expressed as a ratio to the stoichiometric ratio of carbon required to reduce iron, is from about 0.8 to 6.5. Also preferably, the concentration of coke is from 2.5 to 5.

[00033] Without wishing to be bound by a particular theory, the carbon/sample ratio, referred to as the stoichiometric carbon coefficient, is calculated using the average composition of titanomagnetite, which can be, for example, 4FeO.3Fe ₂ O ₃ .2TiO ₂ , along with the following reactions:

4FeO _(s) + 4C _(s) → 4Fe _(s) + 4CO _(g) , and

3Fe ₂ O _3(s) + 9C _(s) → 6Fe _(s) + 9CO _(g) .

[00034] Based on these reactions and the composition of the titanomagnetite, the stoichiometric coefficient of carbon is 0.153 (wt carbon/wt concentrate).

[00035] According to the description, the recovery step is carried out in the temperature range from approximately 900°C to 1200°C. More preferably, the recovery step is carried out in the temperature range from about 1000°C to 1100°C. The residence time in the recovery step is preferably 1 to 3 hours. More preferably, the residence time in the recovery step is about 2 hours.

[00036] In one embodiment, the recovery step is carried out using reformed natural gas.

[00037] Preferably, the percentage of metallized iron in the reduced ore or concentrate is from about 50 to 100%.

[00038] The ferric leaching step is preferably carried out with ferric chloride. Preferably, the concentration of ferric chloride is in the range from about 10 to 40 wt. %. More preferably, the concentration of ferric chloride is in the range from about 25 to 35 wt. %. Also preferably, the concentration of ferric chloride is about 28 wt. %.

[00039] Even more preferably, the step of leaching the ferric iron is carried out at a temperature in the range from about 60°C to 110°C at atmospheric pressure. More preferably, the step of leaching ferric iron is carried out at a temperature in the range from about 60°C to 80°C at atmospheric pressure. The residence time in the ferric leaching step is preferably 1 to 5 hours. More preferably, this time is from 1 to 3 hours.

[00040] The solids content in the ferric iron leaching step is preferably in the range of about 3 to 20 wt. %. More preferably, the solids content is in the range from about 3 to 14 wt.% or also preferably from 4 to 5 wt. %.

[00041] Those skilled in the art will appreciate that the solids content during the ferric leaching step will depend on the amount of reduced iron in the reduced ore or concentrate and the solubility of any ferrous chloride that is formed during the ferric leaching process.

[00042] In one embodiment of the invention described in international patent application PCT/AU 2018/050310 (WO 2018/184067), the method further includes:

feeding the ferric iron leaching residue to the acid leaching stage to obtain an acid leaching product containing vanadium and an acid leaching residue.

[00043] Preferably, the acid leach residue contains titanium.

[00044] The acid leaching step is preferably carried out using hydrochloric acid (HCl). More preferably, the concentration of acid HCl is in the range from about 15% to 32 wt. %. Also preferably, the concentration of acid HCl is in the range from about 15% to 20 wt. %.

[00045] The acid leaching step can be carried out at atmospheric pressure or under pressure. The atmospheric acid leaching step is preferably carried out at a temperature in the range of about 25°C to 100°C. Even more preferably, the atmospheric pressure acid leach step is preferably carried out at a temperature in the range of about 60°C to 80°C.

[00046] In one embodiment, the percentage of metallised iron in the reduced ore or concentrate is preferably about 50 to 70% for an atmospheric pressure acid leach step, or about 70 to 100% for a pressure acid leach. .

[00047] The pressure acid leaching step is preferably carried out at a temperature in the range of about 120 to 180°C, more preferably in the range of about 140 to 160°C, and even more preferably at a temperature of about 150°C.

[00048] The residence time in the atmospheric acid leaching step is preferably about 0.5 to 10 hours. More preferably, the residence time in the atmospheric pressure acid leach step is in the range of about 6 to 8 hours.

[00049] Preferably, the pressure acid leach step has a residence time in the range of about 0.5 to 4 hours. More preferably, the pressure acid leaching step has a residence time in the range of about 3 to 3.5 hours.

[00050] The solids content in the acid leaching step is preferably in the range of about 10 to 30 wt. %. More preferably, the solids content in the acid leaching step is in the range of about 10 to 15 wt. %. Even more preferably, the solids content in the acid leach step is about 11 wt. %.

[00051] Those skilled in the art will appreciate that the conditions of the acid leach step, eg, HCl acid concentration, residence time, and solids content, are controlled to minimize free acid at the end of the acid leach step. Preferably, the free acid concentration at the end of the acid leaching step is in the range of about 10 to 40 g/l.

[00052] In accordance with the present invention, a method for producing titanium dioxide is provided. The method includes pulping a titanium-containing material, such as a leach residue, with concentrated sulfuric acid. The leach residue can be, for example, the leachate of titanomagnetite ore in hydrochloric acid as described above. The proposed method further includes leaching the sulfuric acid pulp with a weak solution of sulfuric acid. This produces a "black liquor" containing, for example, about 80 g/l Ti, 8 g/l Fe, 0.5 g/l V and about 440 g/l free acid. The recovery of titanium to black liquor exceeds 98%, with about 79% of iron and 90% of vanadium also being recovered to black liquor from the leach residue.

[00053] The preferred conditions for the extraction of titanium by the proposed method were achieved in the first boil at 190°C for three hours using a mixture of leach residue and concentrated sulfuric acid in a ratio of 1: 1.27 (g/g). For a given leach residue containing 67.3% TiO ₂ , the cooking acid requirement is calculated to be 1.9 grams of concentrated H ₂ SO ₄ for every gram of TiO ₂ contained in the sample.

[00054] The boil residue is then further leached with dilute, eg, 6% H ₂ SO ₄ acid, at about 60° C. for about 15 hours (20% solids in a shaking incubator) to produce black liquor. Separation of solid and liquid phases can be achieved by simple filtration.

[00055] Some dilution of the acid at the beginning of the boil is carried out to generate enough heat to initiate a potentially autothermal process. Comparative thermograms of acidic slurries of ilmenite (which is known to flow autothermally via the sulfate pathway) and the leach residue prepared as described above show similar heat generation in the initial mixing step and suggest that an autothermal pulping reaction is also possible for the titanium-containing leach residue. obtained as described above.

[00056] More extensive tests were also carried out to establish whether titanium could be recovered from black liquor and to provide guidance values for its grade and recovery. Titanium was extracted from black liquor by hydrolysis, and a fine (p80 ~ 10-12 μm) white powder was obtained with a TiO ₂ content of 74.2%, the extraction of titanium was 80%. Calcination (1000°C) of the hydrolyzed precipitate gave a weight loss of 22%, indicating a final TiO2 content of 95%.

[00057] The thus obtained crude titanium dioxide can be subjected to surface treatment to obtain a product with the desired characteristics for titanium pigment.

[00058] Next, a method for producing a titanium dioxide pigment according to the present invention will be described with reference to the following non-limiting example.

Example

[00059] The method was carried out using a concentrate from the Mount Peak ore body (Mt Peake) of the applicant, which has a relatively coarse particle size distribution (~p80 150 μm) and a chemical composition close to that expected for the proposed commercial plant.

[00060] As described above in relation to the decision of International Patent Application PCT/AU 2018/050310 (WO 2018/184067), the concentrate was subjected to reduction roasting and ferric chloride leaching to first remove the bulk of the iron from the sample. The ferric chloride leach residue was pressure leached using hydrochloric acid (20% HCl, 20% solids) at 150° C. for three hours. The solid leach residue was separated from the leach liquor, washed and dried to give a pressure leach residue.

[00061] The leach residues obtained from this concentrate served as the main feedstock used to study the potential of the method of the present invention in recovering titanium from the Applicant's concentrate.

[00062] The recovery of titanium by the method of the present invention was carried out using concentrated sulfuric acid pulping followed by leaching with dilute sulfuric acid (6%) at low temperature (60°C) to obtain a rich or "black" liquor.

[00063] The high dissolved titanium black liquor was then hydrolyzed to precipitate hydrated titanium oxide, which was then calcined to recover the TiO ₂ product.

[00064] The following describes the specific technological steps for implementing the method of the present invention:

1. A sample was prepared from the concentrate using roasting, ferric chloride leaching and pressure leaching in HCl as described above;

2. Approximately 50 g of the leach residue was accurately weighed into a tared 400 ml flash tank;

3. The sample was placed in a fume hood and concentrated sulfuric acid (88-98%) was slowly added to it with constant stirring with a glass/plastic rod;

4. This homogeneous slurry/vessel was then transferred to a Birlec™ oven preheated to the desired temperature. The furnace was ventilated so that flue smoke could be removed;

5. The mixture was boiled at various temperatures for various times before being cooled to 60°C;

6. Dilute sulfuric acid (6%) was slowly added to the boil residue with gentle agitation (glass or plastic rod) until a dilute slurry was obtained;

7. The tank was then placed in a laboratory oven (shaking incubator) at 60°C and stirred for the first 2 hours. After that, the sample was left overnight at a temperature;

8. The liquor from the leach was recovered by centrifugation (in early trials) or filtration (in recent trials) and the weight and density of the liquor was recorded (Anton Paar DMA™35).

9. The solid residue was washed with deionized water (repulped) and filtered again;

10. The washes were repeated and for each the mass and specific gravity of the liquor were recorded; and

11. The finally washed solids were transferred to a glass beaker and dried overnight at 105°C. The final dry weight was recorded.

[00065] The free acid of the leaching and washing solutions was determined by titration with alkali (using EDTA) and the solutions were analyzed for metals using inductively coupled plasma (ICP). Solid residue analyzes were performed using X-ray fluorescence (XRF) as well as alkali melt/ICP. Metal recovery and mass balances were determined from liquor and residue analyzes.

[00066] Titanium was recovered from sulfuric acid leach liquors (black liquor) by hydrolysis as described by Grzmil B.U. and Bogumil K. “Hydrolysis of titanium sulfate compounds.” Chemical papers - Slovak Academy of Sciences February 2008.

[00067] A portion of black liquor (100-150 g) was taken, accurately weighed into a small beaker and placed on a hot plate (~80°C).

[00068] Based on the free acid titration data, the volume of water needed to dilute the black liquor sample to achieve a final free acid concentration of 150 g/L or less (~300-400 ml) was calculated. This volume of deionized (DI) water was added to an 800 ml tared beaker and heated on a hot plate with stirrer to >80°C.

[00069] At a temperature (80-90°C), black liquor was added dropwise to the stirred hot water. The dilution of black liquor is exothermic and the hydrolysis reaction occurs almost at the boil. Black liquor was added over 30-40 minutes and the stirred mixture was maintained at 90° C. for 3 hours. After cooling, the solution was allowed to settle and the final weight was recorded.

[00070] Separation of the solid and liquid phases was carried out using a centrifuge, the precipitate was washed twice with deionized water, and the mass and specific gravity of the liquor from the leach and all washes were recorded. The washed solids were dried overnight at 105°C.

[00071] The above-described concentrate, obtained from Applicant's Mount Peak orebody, was treated by reduction roasting, ferric chloride, and pressure leaching, also described above, to produce a leach residue on which the rest of the titanium recovery tests were conducted. The following table 1 lists the main elements of the chemical composition of the concentrate from the applicant's Mount Peak orebody.

[00072] Table 1

[00073] Calcination of the concentrate was carried out in three campaigns, a total of 17 batches of 300 g each, in a rotary pot kiln at a temperature of about 1050-1100°C for 2 hours. The first campaign was run with a carbon stoichiometry of 0.8 which gave an iron removal rate of only 51%. This was less than desired and two subsequent campaigns were run with a carbon stoichiometry of 1.0. This gave a more satisfactory iron recovery of about 80%.

[00074] The ferric chloride leach residues became the feedstock for pressure leaching. Two batches were leached with 20% HCl at 16% solids, 150° C. and 400 kPa for 3 hours. The slurry was filtered, the leachate and washed solids were analyzed for metal recovery and mass balance data. The following table 2 summarizes the chemical composition of the material at each step (where FCLR is the ferric chloride leach residue and pTLR is the pressure leach residue).

[00075] Table 2

[00076] Initial sulfuric acid digestions for titanium recovery were carried out on ferric chloride and leach residues (TLR), and the bulk of the leach testing work was performed on pressure leach residues (PTLR) obtained from concentrate produced from the Mount ore body Peak of the applicant. In order to obtain comparative data, several sulfuric acid digests were also carried out on ferric chloride leach residues (FCLR, feedstock for pressure leaching) to obtain comparative data.

[00077] Analysis of major elements in various feedstocks is summarized in Table 3 below. The TLR residue was fine (P ₈₀ <45 µm) and PTLR obtained from the larger concentrate from Applicant's Mount Peak orebody showed a P _{80 of} about 125 µm.

[00078] Table 3

[00079] An initial digestion test was conducted to evaluate the methodology using ferric chloride leach residue (FCLRpp) and pressure leach residue (TLRpp) feedstocks from Applicant's pilot plant, with the above fine particle size P ₈₀ <40 µm.

[00080] Duplicate samples of FCLRpp (50 g) were mixed with concentrated sulfuric acid (117 g) and heated to 300°C for 4 hours. The resulting sintered mass was cooled and leached with water (800 ml deionized) at 80° C. for 2 hours. The results of these tests (S19-S1, S2) are shown in the following table 4.

[00081] Table 4 - Initial test results

[00082] Titanium leaching under these experimental conditions gave poor results with a recovery of 32-38%. A similar boil/leach attempt was made on the leach residue in the TLRpp pilot plant using 100 g of sample and 200 g of acid boiled at 280°C for 4 hours. This leach was modified to use less acid (200 ml 50% sulfuric acid for 2 hours) and the filter residues were then washed twice with 50% acid. The aim of this attempt was to keep the leached titanium in solution (TiOSO ₄ , which readily precipitates TiO ₂ when diluted) and to recover any soluble titanium entrained in the solids. This method allowed to increase the extraction of titanium by 51% (S19 S3, table 4). An additional 3 hour boil at 250° C. and leaching with 6% acid for 2 hours without washing the residue (S19 S4, Table 4) returned to a titanium recovery of 36%.

[00083] It was decided to adopt moderate cooking temperatures with longer leaching times (using 6% acid) and multiple washes of residues. This modified pulping/leaching process was used to recover titanium from a pressure leach residue obtained from Applicant's Mount Peak concentrate (PTLR-1).

[00084] Duplicate samples (50 g) were mixed with concentrated sulfuric acid (~100 g) and boiled at 250°C for 3 hours. The resulting precipitate was cooled and leached with dilute sulfuric acid (200 g 6%) for 15 hours at 60° C. (stirred incubator). The liquor from the leach was separated from the solids and analyzed to determine the amount of titanium recovered into the solution. The solid residue was washed (4 times with 6% H ₂ SO ₄ ), dried and analyzed to determine the mass balance. The results of this test are presented in Table 5 below. Estimated recovery data are for XRF and ICP assays and include recovery of metals after wash steps.

[00085] Table 5

(A) XPF analysis

(B) Heat/ICP analysis

[00086] This procedure used a longer leaching time with dilute acid to recover and stabilize the titanium in solution. Along with washing the residues, this allowed over 98% of the titanium to be recovered from PTLR-1. This high recovery of titanium was re-confirmed by analysis of both liquor and solids, with very good agreement on the mass balances.

[00087] Although most of the iron and vanadium were leached from the sample in the previous processing steps described above, this procedure also achieved high recovery rates for these residual metals (~80 and 90% for Fe and V, respectively) from PTLR-1.

[00088] The liquors from the sulfate leach, referred to herein as "black liquor", from PTLR-1 were essentially dark green (industrial liquors are black). This is thought to be due to the lower iron content (4.5%) of the feedstock used. The following table 6 provides information on the composition of the leach liquor from the results of both leach tests (pure leach liquor).

[00089] Table 6

[00090] Only 76 g (46 ml) of concentrated leach liquor (out of 200 ml of dilute acid added) was recovered from sample S19 S5 due to the high evaporation rate during leaching, and therefore the metal content of the leach liquor was high. The titanium recovered from the leach liquor was about 46% titanium in PTLR-1, while two washes (170 ml total) of the sulfate leach residue recovered most of the remaining 54% titanium from the captured solution.

[00091] In the second leach (S19 S6), evaporation losses were minimized and only 232 g (163 ml) of green liquor was recovered from the leach with a titanium content of about 80 g/l. The recovered leach liquor contained 70% titanium in PTLR-1, which increased to >97% after two washes (2 x 120 ml) with deionized water.

[00092] The recovery of titanium in these initial tests was high, and two additional sulphate pulps were carried out to confirm. Titanium recovery tests (S19-9, 10) showed a titanium recovery above 99% as seen in the following Table 7 and confirmed the high recovery data obtained in earlier tests.

[00093] A series of tests were then performed in an attempt to optimize some experimental conditions and reduce reagent consumption, which is also shown in Table 7.

[00094] Cooking temperatures (11-13) were reduced in steps of 25°C from 200°C to 150°C to determine the minimum required cooking temperature. Cooking at temperatures above 175°C kept the titanium recovery above 97%, but reduced it to 89% when the cooking temperature dropped to 150°C. It is assumed that the target cooking temperature is about 190°C. Therefore, most of the subsequent tests to optimize other variables were carried out at 190°C or higher.

[00095] Table 7

[00096] The effect of total acid addition and acid concentration on titanium recovery data was investigated in Tests 14-22 of Table 7. Acid additions for these cooking tests were based on the titanium content of the feedstock. However, XRF analyzes for this batch of feedstock (PTLR-2) were not available and acid values were calculated from a single ICP analysis from pressure leach data (61% TiO ₂ ).

[00097] The test results clearly showed a marked reduction in titanium recovery for all of these tests and convincingly demonstrated that the acid addition rates were insufficient for the titanium content of this feedstock. The XRF data reverted to 67.3% TiO ₂ , confirming insufficient acid levels for these tests, and were therefore repeated (23-27) with higher acid levels recalculated. This resulted in titanium recovery values above 90% for all acid tests except for test 27 (81.9%). This test used more acid than the others but boiled at a lower temperature of 175°C.

[00098] Acid tests show that for every gram of TiO ₂ content in a raw material sample, a minimum of 1.9 g H ₂ SO ₄ (100%) is required. Therefore, a 50 g sample with 67.3% TiO ₂ requires a minimum acid addition (100%) of 64 g (or 65 g of 98% H ₂ SO ₄ ) to achieve high (>90%) titanium recovery. The minimum acid addition for tests 23-27 was 65 g commercial grade 98% sulfuric acid (tests 24 and 26). The maximum acid addition for tests 14-22 was only 55 g H ₂ SO ₄ (or 1.6 g H ₂ SO ₄ (100%) per gram TiO ₂ ).

[00099] As noted above, the sulphate pulping and leaching of titanium from ilmenite proceed autothermally, and no additional heat input is required to carry out the pulping. However, the pressure leach residue feedstock to the process of the present invention contains more TiO ₂ than ilmenite, but significantly less iron. Therefore, it was not known whether the original leach residue used as feedstock for the proposed process could produce enough heat in the initial exothermic reaction to allow the brew to proceed without additional heating.

[000100] Samples of PTLR-3 (50 g) were mixed with sulfuric acid (65 g) and water (10 g) in a small ceramic dish connected to a glass-encased thermocouple. An initial rise in temperature after stirring was recorded and the sample was then slowly heated in an attempt to trace any onset of an exothermic reaction.

[000101] The first test was conducted in a large furnace used for pulping and leaching. This furnace has a large thermal mass, and the controlled heating rate of the sample lagged significantly behind the heating rate of the control furnace thermocouple, as seen in Figure 1. The result was a sample temperature plot with little detail and a curve that broadly followed the profile of the furnace.

[000102] The second test shown in FIG. 2 was a repeat of the first, using a much smaller forced-ventilation oven with fast temperature response. This sample showed a similar initial temperature rise after mixing with acid and water (20-76°C) and then placed in the oven. The temperature of the sample rose to an oven temperature of 100° C. after 40 minutes, but no significant temperature deviation was observed in the sample other than an increase in the oven temperature set point.

[000103] To increase the sensitivity of the test, and also in the hope of detecting any exothermic reactions, a small sample of PTLR-3 was thermally analyzed as shown in FIG. 3. A mixture of acid and water was prepared separately and cooled with ice to remove the initial heat from the dilution. The acid was then mixed with the pressure leach residue and placed in a small ceramic sample holder in a thermogravimetric differential scanning calorimeter (TGA/DSC) and heated at a rate of 2.5°C/min to 220°C. To obtain control data for comparison, a similar analysis of the ilmenite sample was also performed.

[000104] Even without the initial heat of dilution, PTLR-1 and ilmenite exhibit slight exothermic reactions when mixed with acid (31-81°C). The exothermic peaks begin to return to baseline but stop at about 80°C in both samples. This indicates that an additional exothermic reaction occurs near this temperature. Both samples return to baseline around 135°C (little heat absorption) and then exhibit large exothermic reactions (cooking) between 160-210°C.

[000105] The DSC curves for ilmenite and PTLR-3 are very similar, and have the largest discrepancy in the amount of heat released at the peak of exothermic decomposition between 160-210°C. The energy released by ilmenite (based on the peak areas of incomplete reactions) is approximately twice that of the PTLR-3 sample.

[000106] Since the cooking reaction starts at about 160° C., the initial exothermic reactions will require the mixtures to be heated to this temperature to initiate the autothermal cooking process. The low temperature portions of the plots have very similar profiles, and since the ilmenite reaction is known to be autothermal, the PTLR reaction is expected to be the same.

[000107] The recovery of titanium from leach liquors was determined by hydrolyzing a small portion of the leach liquors to produce hydrated titanium oxide (TiO(OH) ₂ ). An initial leach liquor test (S19H1 using S19 S5 leach liquor) was performed by simply boiling the liquor to initiate hydrolysis. This resulted in small amounts of sediment, but eventually resulted in gels that did not promote hydrolysis or prevent solid-liquid separation. To address these issues, a procedure was adopted to dilute the liquors from hot water leaching to the desired final free acid content (Grzmil et al., 2008).

[000108] In the first (failed) hydrolysis test, a mixture of leach liquor (40 g) and first wash solution (180 g) was used to obtain a solution of approximate composition suitable for hydrolysis, which is presented in table 8. Initial boiling of this solution led to the formation gel from which no precipitate was obtained. An additional liquid sample during the boiling phase (~45 g) was an attempt to monitor the progress of hydrolysis (through changes in the liquor assays), but it stopped after gel formation.

[000109] Table 8

[000110] This gel sample was diluted back by adding water (50 ml). Approximately 114 g of liquor remained (with 446 g/l free acid) which was then treated according to the method described by Grzmil et al. hydrolysis. The volume of hot water was calculated to achieve a final free acid value below 150 g/l.

[000111] Table 9 shows the analysis of the two streams of the original liquor and 1 and 2 hour samples. The mass balance has been corrected for this group of samples. In this case, the total contribution of titanium to the hydrolysis process was (4.45+7.54-1.411-2.027-0.036-0.108) = 8.41 g (Table 12).

[000112] Table 9

[000113] The spent solution from hydrolysis, including a single wash (Table 10 below), retained about 1.68 g of titanium, which was calculated to be an 80% recovery of titanium (Table 12). Similar calculations for iron and vanadium give recovery of these metals to precipitated solids of 18.8% and 4.5%, respectively.

[000114] Analysis of the dried solid precipitate (table 11) showed a titanium content of 44.5%, which gives a recovery of 70.3% of titanium by hydrolysis. The calculated extraction of iron and vanadium was 20.5 and 4.4%, respectively.

[000115] Table 10

[000116] Table 11

[000117] Table 12

* Feed+LL1+W1 - (liquor samples + solids samples). Raw material adjusted for 1 hour and 2 hour sample groups)

[000118] The hydrolyzed product, a fine white powder, was calcined at 1000° C. to remove water of hydration. The weight loss during the calcination step was 22%. Adjusted for this weight loss, the TiO ₂ content of the calcined product is about 95% (Table 13 below). The main pollutant is iron (~1.5%).

[000119] Table 13

[000120] Ignition of the product gives a weight loss of 22%. The expected loss of 1 mole of water is approximately 18% and suggests either the presence of additional water or decomposition of other metal hydroxides.

[000121] The recovery of titanium from the second leach liquor (S19 S6) was a simpler process in which only primary leach liquor and water were used to recover titanium using the Grzmil et al., 2008 process. A sample of leach liquor (150 g, 105.2 ml) was heated and added dropwise to hot deionized water (397 ml) over 40 minutes. The volume of water required was determined from the initial free acid content of the leach liquor (446 g/l) to give a final free acid below 100 g/l. Table 14 shows the volumes of solution after hydrolysis and analyzes.

[000122] Table 14

[000123] Analysis of the leach liquor showed 80 g/L Ti with a free acid concentration of 446 g/L. The target concentration of free acid (<100 g/l) was lowered with the intention of simply obtaining a product with sufficient recovery and purity for analysis.

[000124] The spent solution from hydrolysis after cooling was separated from the precipitated solids and analyzed. The precipitated solids were washed twice with deionized water and both the solids and the washings were analyzed. Table 15 shows the data of waste solutions from hydrolysis. Sludge analyzes are shown in Table 16.

[000125] Table 15

[000126] Table 16

[000127] The following table 17 shows data on the recovery of titanium from solution and solid samples. The mass balance data improved from the previous test and the overall titanium recovery was slightly higher. Iron contamination was found to be less, but the titanium content of the hydrolysis product (uncalcined) was found to be very similar to the previous test (~44.5% Ti). The calcination of this product gave a very similar weight loss (22.6%) and the final purity of TiO ₂ is the same as the original (S19 H1) - about 95.9%.

[000128] As can be seen from the above description, the sulphate digestions of the pressure leach residue in accordance with the present invention produced a black liquor containing about 80 g/l Ti, 8 g/l Fe, 0.5 g/l V and about 440 g/l free acid. In two washing steps, the recovery of titanium in the leach liquors exceeded 98%, while about 79% iron and 90% vanadium were also recovered from the pressure leach residue.

[000129] Testing shows that the optimal conditions for kraft pulping were 190° C. for three hours using a 1:1.27 (g/g) mixture of pressure leach residue and concentrated sulfuric acid. The pressure leach residue showed 67.3% TiO ₂ , so the calculations show that for every gram of TiO ₂ content in the sample, 1.9 g of concentrated H ₂ SO ₄ (100%) is required.

[000130] Dilution of the acid with water to approximately 88-92% H ₂ SO ₄ is recommended to initiate an exothermic reaction when preparing the slurry for cooking. After pulping, the black residue is repulped/leached in 6% H ₂ SO ₄ acid at 60° C. for 15 hours. The black liquor is recovered by filtration and the solids are washed (2 x weight of feedstock with 6% H ₂ SO ₄ ) to achieve high titanium recovery.

[000131] Tests have shown that cooking at temperatures below 180°C showed a rapid decrease in titanium recovery, as well as insufficient addition of acid (<1.9 g H ₂ SO ₄ per gram TiO ₂ ). Feed samples with higher iron content (pressure leach residue=1.9% Fe) may also require additional acid if they differ significantly from the current pressure leach residue samples.

[000132] Comparative thermograms of acidic pressure leach residue slurries and ilmenite show that they both generate the same amount of heat during the initial mixing stage. This suggests that the autothermal pulping reaction observed for ilmenite may occur with the pressure leach residue, and thus it is contemplated that this may reduce or eliminate the need for external heating during the pulping step.

[000133] Some initial recovery of titanium from black liquors from leaching was achieved through hydrolysis. The titanium in the liquors from the sulfate leach was hydrolyzed by dilution in hot water to give a fine white powder containing 74.2% TiO ₂ . The extraction of titanium by hydrolysis was about 80%.

[000134] Calcination (1000°C) of the product showed a weight loss of 22%, which was slightly higher than the loss due to one mole of water (18%), and provided a calculated content of 95% TiO ₂ in the calcined product.

[000135] The crude titanium dioxide thus obtained can be subjected to a surface treatment to obtain a product with the characteristics required for a titanium pigment product.

[000136] Modifications and variations that will be obvious to a qualified specialist, should be considered falling within the scope of the present invention.

Claims (36)

1. A method for producing titanium dioxide, which consists in the fact that: (i) subjecting the titanium-containing leach residue to a pulping step with concentrated sulfuric acid, and (ii) then subject this residue to leaching in dilute sulfuric acid to obtain black liquor, from which titanium dioxide is then obtained, and the residue from the leaching is the residue from the hydrochloric acid leaching of the titanomagnetite type ore. 2. The method of claim 1 wherein the titanium dioxide is used to produce the titanium dioxide pigment. 3. The method according to claim 1 or 2, in which the concentration of hydrochloric acid is: (i) from about 15 to 32 wt. % or (ii) from about 15 to 20 wt. %. 4. The method according to any one of the preceding claims, wherein the titanomagnetite type ore leaching feedstock is a ferric iron leachate. 5. The method of claim 4, wherein the ferric leaching step is carried out with ferric chloride. 6. The method according to claim 5, in which the concentration of ferric chloride is: (i) from about 10 to 40 wt. %, (ii) from about 25 to 35 wt. % or (iii) about 28 wt. %. 7. A method according to any one of the preceding claims, wherein the titanium-containing leach residue has a P ₈₀ equal to: (i) ≤125 µm, (ii) ≤45 µm or (iii) <40 µm. 8. A process according to any one of the preceding claims, wherein the recovery of titanium into black liquor is at least 98%. 9. A method according to any one of the preceding claims, wherein the cooking step is carried out at a temperature of: (i) above 175°C or (ii) about 190°C. 10. A method according to any one of the preceding claims, wherein the cooking step is carried out for a period of: (i) approximately 3 to 4 hours, or (ii) about 3 hours. 11. A process according to any one of the preceding claims, wherein the mixture of leach residue and concentrated sulfuric acid in the cooking step has a ratio of approximately 1:1.27 (g/g). 12. A process according to any one of the preceding claims, wherein at least 1.9 grams of concentrated sulfuric acid is provided in the cooking step (i) for every gram of TiO ₂ in the titanium-containing leach residue. 13. A process according to any one of the preceding claims, wherein the dilute sulfuric acid in the leaching step (ii) is about 6% sulfuric acid. 14. A process according to any one of the preceding claims, wherein the leaching step (ii) is carried out at a temperature of about 60°C. 15. A method according to any one of the preceding claims, wherein the leaching step (ii) is carried out for about 15 hours. 16. A process according to any one of the preceding claims, wherein the leaching step (ii) is carried out at about 20% solids. 17. A process according to any one of the preceding claims, wherein the black liquor is recovered from the slurry obtained in the leaching step (ii) by filtration and the solids are washed to recover the titanium. 18. A method according to any one of the preceding claims, wherein the cooking step (i) proceeds autothermally. 19. A method according to any one of the preceding claims, wherein the cooking step (i) further comprises an initial dilution of the acid. 20. The method of claim 19 wherein the acid is diluted with water to approximately 88-92%. 21. A process according to any one of the preceding claims, wherein the black liquor is subjected to hydrolysis and calcination to recover the titanium dioxide.

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