AU2015328791B2 - Method for processing alumina-containing raw material and method for breaking down alumina-containing raw material during processing - Google Patents

Abstract

The invention relates to chemistry and metallurgy and is intended for the processing and breakdown of alumina-containing raw material. A processing method is carried out as a cyclic process that includes: a breakdown stage, in which raw material is decomposed using a reagent solution containing ammonium bisulphate and sulphuric acid and the resultant slurry is separated, producing undecomposed solid residues, which are rinsed with water, and an alum mother solution, wherein the alum mother solution and the rinse waters are collected separately; a purification stage, in which the rinse waters are purified of iron then combined with the alum mother solution and cooled, and crystals of ammonium alum are isolated, wherein the sulphuric acid used for the reagent solution in the breakdown stage is separated off; a precipitation stage, in which aluminium hydroxide is precipitated from the purified alum solution with ammonia; a separation stage for separating off aluminium hydroxide; a stage for obtaining solid ammonium sulphate and a stage for the thermal decomposition thereof into ammonium bisulphate and ammonia, which can be used in the breakdown stage, during the preparation of the reagent solution, and in the precipitation stage, respectively. The invention makes it possible to process any alumina-containing raw material at low temperatures, to reduce reagent loss and to reduce the volume of reagents which needs to be replenished during the process.

Description

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Method for alumina-bearing raw material processing and a method for digestion alumina-bearing raw material as part of processing thereof

Technical Field

This invention relates to the chemical industry and metals industry, and, more particularly, to a method for alumina-bearing raw material proc15 essing and its sub-process for liberating such raw material as part of processing thereof to produce smelter-grade alumina and by-products.

Related Art

Alumina-bearing (aluminous) raw material processing methods, to which the first invention of the proposed group relates, are commonly known.

Thus, known are the alkali processes (Bayer process and modifications thereof) used to process alumina-bearing raw materials (Lainer, A.I., Yeremin N.I., Lainer Yu.A., Pevzner I.Z. Alumina Refining. M., Metallurgy, 1978, 394 pages. [1]; Troitskiy I.A., Zheleznov V.A. Metallurgy of

Aluminum. - M., Metallurgy, 1977, p. 42-116 [2]; RF Patent

No.RU2360865, pubd. 10.07.2009 [3]; RF Patent No.RU 2193525, pubd.

27.11.2002 [4]). The above methods require the use of high-grade (i.e., low-silica) bauxitic raw materials that are becoming less and less available, since mineable reserves of such bauxites are finite everywhere. Processing of low-grade bauxites, as well as aluminum silicate raw materials, such as nepheline, according to the Bayer process and similar processes is technologically and economically unsound. The reason is that in such technological process the silicon oxide chemically interacts with the alkali, and a large amount of both alkali and aluminum is lost to capture the silicon oxide due to the formation of a mixed compound, i.e., hydrated sodium alu10 mino silicate.

Close to the alkali processes are those based on raw material lime, or soda-lime, or alkali-lime sintering with subsequent washing of the sintered material with water or aqueous soda solution: Matveev, V.A. Physical and Chemical and Technological Bases for Improving Efficiency of

Nepheline-Bearing Raw Material Integrated Processing by Acid Methods. A Thesis for PhD in Engineering, Apatyty, 2009, 299 pages. [5]. Now about 40% of alumina refined in Russia is produced from the Kola Peninsula nepheline and nepheline syenites from Siberian deposits using this method (ref. to [5], and: Isakov, E.A., Pikalevo Industrial Group “Gli20 nozyom” in New Realities. Non-Ferrous Metals, 1997, No.4, p. 8 [6]). However, the above methods are substantially constrained by their high energy intensity and necessary production and processing of large amount of sodium-calcium-silicate byproducts. In future, in view of the increasing energy prices, cost efficiency of the sintering process may decrease dramati25 cally.

An alternative to the alkali and sintering methods for low-grade bauxite and non-bauxite (aluminosilicate) aluminum raw material processing are acid processes providing for silica separation as early as at the initial stages of the process without adding any reagents to bind the silica [1, 2, 5]. Furthermore, the acid processes are much less energy intensive than the sintering processes.

Known are nitric acid methods for aluminous raw material processing (ref. to [1, 5], and: Vaitner, V.V., A Research into Nitric Acid Processing of Aluminosilicates to Produce Aluminum Oxide. A Thesis for Candidate in Engineering Sciences, Yekaterinburg, 2004, 146 pages [7], and RF Patents: No.RU 2202516, pubd. 20.04.2003 [8]; No.RU 2215690, pubd.

10.11.2003 [9]; No.RU 2372290, pubd. 10.11.2009 [10]; No.RU 2460691, pubd. 10.09.2012 [11]). These processes (in various modifications) include treatment of the raw material with hot concentrated nitric acid, filtering the resulting slurry and washing the precipitate, further processing of the filtrate including segregation of intermediate aluminum compounds and sepa15 ration thereof from iron compounds, and, preferably, thermal hydrolysis of the intermediate aluminum compounds to recover nitric acid and produce hydrated or dry aluminum oxide, i.e., alumina. All nitric acid processes feature a number of common disadvantages, including: poor filterability of the slurry produced by decomposing easy to liberate aluminosilicate raw mate20 rials, such as nepheline ore and concentrates (even where flocculants are used), making the process more costly and labor intensive; low digestion degree achieved by direct decomposition of bauxite and kaolinite-boehmite raw materials, requiring pre-calcination thereof, which also makes the process more costly due to increasing energy intensity; poor rate of nitric acid recovery and, therefore, the need to deliver and consume large quantities of fresh acid; environmental hazard from the process due to nitrogen oxide formation and significant economic losses incurred in converting thereof into nitric acid.

Known are hydrochloric acid processes (ref. to [7], and: Schwartzman, B.H., Acid-Based Methods for Alumina-Bearing Raw Material Proc5 essing. M.: Tsvetmetinformatsia, 1964. 82 pages [12]; Pustilnik, G.L., Pevzner, I.Z., Acid-Based Methods for Processing Low-Grade AluminumBearing Raw Materials. M.: Tsvetmetinformatsia, 1978. 52 pages. [13]). These methods comprise pre-calcining the raw material, digestion, siliceous slug separation by filtering, de-ironing by various methods, including evaporation of the filtrate solution to produce crystalline “yellow salt” (A1C1₃ 6H₂O admixed with iron), the latter being cooled and washed with hydrochloric acid to produce “white salt” (pure aluminum chloride crystals), which, in turn, is calcined at temperatures above 1000°C to produce alumina and to recover hydrogen chloride in the form of a mixture with wa15 ter vapors to be returned, after absorption, to the head of the process as 30% hydrochloric acid. In recent years, there has been a revive of interest in these methods owing to the development of efficient chloride calcination and hydrochloric acid recovery plants (Herbert Weissenbaeck, Benedikt Nowak, Dieter Vogl, Horst Krenn. Development of Chloride Based Metal

Extraction Techniques - Advancements and Setbacks, Proceedings of Nickel-Cobalt-Copper Conference of ALTA-2013, 29 May - 1 June, 2013 Perth, WA., Melbourne, Australia, p. 360 [14])· However, the alumina produced through such refining technique requires further de-ironing [13], In recent years, Orbite Aluminae Inc., of Canada, presented a method based on the above technique, comprising further operations to remove iron by extraction (RF Patent No. RU 2471010, pubd. 27.12.2012 [15]). The major disadvantages of the above methods [7, 12 - 15] include stricter corrosion resistance requirements to the equipment, and significant power usage for the hydrochloric acid recovery in the environment where energy costs are growing faster than those of the smelter-grade alumina [7].

Commonly known are the sulfuric acid methods for alumina5 bearing raw material processing (ref. to [1, 7,12], and: Lainer, Yu.A., Integrated Processing of Aluminous Raw Materials by Acid-Based Methods.

M. : Nauka, 1982, p. 208 [16]; Zapolskiy, A.K., Sazhin, V.S., Zakharova,

N. N., Crystallization of Basic Aluminum Sulfate Salts. Alumina Chemistry and Technology. Novosibirsk, Nauka, 1971, p. 430 - 438 [17]; Zapolskiy,

A.K., Sulfuric Acid Processing of High Silica Aluminous Raw Material.

Kyiv, Naukova Dumka, 1981, p. 198 -200 [18]; Paweena Numluk and Aphiruk Chaisena. Sulfuric Acid and Ammonium Sulfate Leaching of Alumina from Lampang Clayll E-Joumal of Chemistry. 2012. V 9, No.3. p. 1364 - 1372, http://www.ejchem.net [19]). According to these methods, pre-calcined or crude ore is treated with sulfuric acid. Aluminum sulfate salts are then extracted from the de-ironed sulfuric solutions: aluminum sulfate, alum or basic salts to produce alumina by direct calcination or aluminum hydroxide calcination, previously extracted with ammonia [12], One major advantage of using sulfuric acid to process aluminous raw mate20 rials is that it allows employing lower cost equipment owing to the vast track record of anti-corrosion protection in the sulfuric acid production and other industrial sectors. Where the sulfuric acid-based decomposition method is used with subsequent aluminum separation in the form of ammonia alum, the slurry resulting from the separation is filtered, and the re25 suiting filtrate is added with ammonium sulfate; herein, to prevent formation of ferriammonium sulfates co-crystallizing with the intermediate product, residual iron is reduced to ferrous iron using ammonium bisulfite ([12] and RF Patent No.RU 2337877, pubd. 10.11.2008 [20]), aluminum chips (Sandler, E.M., Lainer, Yu.A., Lainer, A.I., Chizhikov, D.M. Product DeIroning in Processing Nephelines by Sulfuric Acid-Based Method. NonFerrous Metals Industry. News of Higher Education Institutions, 1962,

No.2,p.30-33 [21])or other reducing agents, such as sulfur dioxide gas (Funaki K. Sulfuric acid process for obtaining pure aluminum oxide from its oresll Bull, of the Tokyo Inst, of Technology, 1980, No.l [22]). Similar to other acid-based methods, after digestion by sulfuric acid-based methods it is difficult to separate the solid phase from the liquid phase. Furthermore, power consumed in high-temperature drying (500°C) and direct calcination (1300°C) of aluminum sulfates and their derivatives is at least equal to that of the hydrochloric acid-based methods, while the mixture of SO₂ and SO₃ gases, resulting from calcination, requires extra capital and operational expenditures to recover (synthesize) sulfuric acid and to return it to the head of the process for aluminous raw material digestion.

An alkali and acid method is known from RF Patent No. RU 2440296 (pubd. 20.01.2012 [23]). The method comprises acid treatment, preferably with the nitric acid, of the red mud resulting from the raw material conventional alkali decomposition, instead of the original raw material acid treatment. The resulting solution is boiled out, and the precipitate is thermally decomposed at a temperature up to 600°C to produce sodium aluminate (a mixture of sodium and aluminum oxides). The method is described by the inventors thereof as versatile and addressing the Bayer alkali process’ disadvantages, since it allows not losing aluminum as part of the hydrated aluminosilicates remaining in the mud. The rationale provided in the patent [23] consists in that, as such, hydrated aluminosilicates may be decomposed with a weak acid at a conventional temperature. However, since the method comprises the alkali treatment step, it actually retains the Bayer process’ disadvantages as identified above and becoming apparent when processing a high-silica raw material, while further extraction of aluminum compounds from the mud does not warrant for the asso5 ciated costs. This is due to that the alkalinous (after pre-treatment of the original raw material) mud has first to be neutralized with a large quantity of acid (this is being omitted in the method description [23]) before any additional quantity of weak acid may act on the hydrated aluminosilicates as expected. The large quantity of acid is not recoverable, while the resulting weak solution boiling out and high-temperature thermal decomposition of the precipitate involve high additional energy consumption. Furthermore, high wetting of the mud during the treatment thereof with weak acid necessitates re-drying of the mud.

Known are the alumina-bearing material processing methods based on salts or salt brines. For example, one know method described in US Patent No. US 1,426,891 (pubd. 22.08.1922 [24]) for processing aluminosilicates, in particular, clay matters, includes boiling in an ammonium bifluoride (NH₄F ‘ HF) solution to obtain precipitates containing aluminum hydroxide and aluminum fluoride that are separated, dried, and decom20 posed by open water steam at temperatures of 300 to 400°C to obtain aluminum hydroxide and to recirculate hydrogen fluoride, and to obtain a solution containing mixed soluble aluminum and ammonia bifluorides, which are also decomposed by long thermal hydrolysis to produce aluminum hydroxide and to recirculate ammonia and a portion of ammonia difluoride in the form of vapors. Certain issues are associated with the method, including non-recoverable consumption of a portion of the initial ammonia difluoride, and strict requirements to the equipment in terms of resistance to corrosive fluorhydric acid and ammonia difluoride vapors.

Further, known are the methods comprising alumina-bearing material, primarily aluminosilicates in the form of clay matter, calcining to5 gether with ammonium sulfate (USSR Inventor’s Certificate No. SU42067, pubd. 31.03.1935 [25]; Ullmann B. Encyklopadie der technischen Chemie, Auflage, Urban & Schwarzenberg, Miinchen & Berlin. 1954, Bd. 3, 401 420 [26]; Grim R.E. Applied Clay Mineralogy, McGraw-Hill, New York, 1962, p. 335 - 345 [27]; G. Bayer, G. Kahr, and M. Mueller-Vonmoos. Re10 actions of ammonium sulphates with kaolinite and other silicate and oxide minerals, Clay Minerals, 1982, V.17, p. 271 - 283 [28]). The calcining is done at temperatures of around 300 to 500°C to obtain a sinter containing mixed aluminum and ammonium sulfates and aluminum sulfate to be leached out from the sinter by water to produce ammonium alum solution.

The latter is cleared from iron impurities and converted to aluminum hydroxide through ammonia precipitation and consecutive precipitation by fluorides and ammonia. Herein, the process cycle includes ammonia produced at the head of the process at the calcining step. The issues associated with the method are energy losses due to calcining of the whole amount of the alumina-bearing material with ammonium sulfate, and non-recoverable losses of ammonium sulfate due primarily to ammonium sulfate decomposition during calcining wherein sulfur dioxide is released. Further, hard to filter complex compounds, as well as hard to de-iron compounds, are produced during calcining.

The closest to the proposed method is a method for processing alumina-bearing raw materials, provided by M. Buchner as far back as in the early twenties of the last century (Patents [29]: British Patent

No.GB195,998, pubd. 12.04.1922; US Patent No.US 1,493,320, pubd. 06.05.1924; USSR naTernr No. SU 11489, pubd. 30.09.1929). The method includes a circular (closed cycle) process, comprising the steps of prior thermal decomposition of ammonium sulfate into ammonia and ammonium bisulfate; dissolving the latter, and autoclave treatment of the aluminabearing raw material with hot ammonia hydrosulfate (bisulfate) solution admixed or not admixed with ammonia sulfate; ammonium alum solution filtering and aluminum hydroxide precipitation therefrom with the previously segregated ammonia; ammonium sulfate recovery from the mother liquor and return to the head of the process. In contrast to the methods [25 28], this method does not include direct involvement of the raw material in the calcination process, thus allowing pure ammonia sulfate to be decomposed at temperatures below 300°C and power consumption to be reduced. Furthermore, alumina-bearing raw material treatment with hot ammonia hydrosulfate solution does not result in any loss of reagents in the form of sulfur dioxide. Finally, this method provides for producing intermediate compounds that are easy to filter and clear from iron. The method provided by M. Buchner is one of the so-called “Named Processes in Chemical Technology” and is also known as Aloton (Encyclopedic Dictionary of

Named Processes in Chemical Technology /Ed. Alan E. Cornyns Boca Raton: CRC Press LLC, 2000, 2-nd Ed., p.19 (Aloton) [30]). The method was implemented at pilot facilities in Germany in the late twenties of the last century and in the USA (State of Oregon) in 1944. However, it has not found industrial application. This was primarily due to the fact that readily available low-silica bauxites make the Bayer alkali process more competitive than the hydrosulfate process, provided by M. Buchner and allowing alumina production from aluminosilicates, thus explaining why the attempts to employ the method, provided by Buchner, never resulted in in10 dustrial-scale application (Bretsznajder St. Otrzymyxvanie estow kwasu ortokrzemowego w fazie gazovey I St. retsznajder, W. Kawecki // Rocz. Chem,- 1955,- 29. - s. 287- 299 [31]; Bretsznajder St. Nova metoda otrzymywania hutniczego tlenku glinovego i innych zwiazkow glinu z glin //

Przem. Chem.- 1963,- V.42, No. 12. - s. 677-683 [32]).

In the current context with low-silica bauxites being a rather scanty resource, it becomes reasonable to revive M. Buchner’s method; however, the issues associated with the described hydrosulfate method [29] still restrict its wide commercialization.

The main disadvantage associated with M. Buchner’s method [29] is its inadequate versatility. The method is applicable only to easy to liberate aluminosilicate minerals (O’Connor, D. J. Alumina Extraction from Non-bauxitic Materials, Aluminium-Verlag, Dusseldorf, 1988, 159 p. [33]), while not allowing a high grade, commeasurable with the corre15 sponding Bayer process’ parameters, of aluminum hydroxide (or alumina after hydroxide calcining) decomposition and recovery from bauxitic raw materials, including high-silica bauxitic raw materials, processing of which is now the priority task. This is due to: a) high aluminum content and low content of alkali and alkali-earth metals in them; b) better availability. An20 other disadvantage follows from the fact that the inventor of the method [29] recommends autoclave decomposition of even aluminosilicates at relatively high temperatures, i.e. at least 200°C (since the necessary completeness of the raw material decomposition will not be achieved at lower temperatures).

Further disadvantage of the method is that excessive energy is consumed to boil out the ammonium sulfate residual solution during process11 ing thereof to obtain solid ammonium sulfate, which is due to limited solubility of ammonium alum in the mother liquor after decomposition where high ammonium hydrosulfate concentrations in the initial reagent solution were used. This necessitates using a dilute reagent solution, and, conse5 quently, to the need to circulate large amounts of water in the circular process. Another disadvantage is due to the non-recoverable usage of the hydrosulfate reagent associated with the relatively high content of alkali and alkali-earth components in the aluminosilicate raw material. This is also the reason for gradual loss of the reagent solution efficiency from cy10 cle to cycle, where recycled ammonium hydrosulfate is used in the circular process.

Finally, the patents [29] does not provide any details of the deironing or silicon removal methods for further producing smelter-grade pure alumina, despite of the fact that the patents do contain information about the possibility to achieve such grade, which requires rather fine purification.

The above disadvantages are not addressed by an embodiment of the M. Buchner’s method [29], specifically, that, where ammonium hydrosulfate admixed with ammonium sulfate is used as the reagent solution.

The methods for alumina-bearing material digestion as part of processing thereof, to which the second proposed invention relates, are included as a step into the above methods [1 - 13,15 - 29], while RF Patent No. RU 2153466 (pubd. 27.07.2000 [39]) covers only this step.

According to the method [39], alumina-bearing raw material diges25 tion includes thermal treatment at a temperature of 400 to 550°C by magnesium chloride aqueous solutions as a reagent and digestion of the sinter with subsequent aluminum extraction from the resulting solution. This method is close to the known sintering methods [5, 6] (with respect to the raw material digestion), and, therefore, retains their disadvantage of high energy intensity, though somewhat reduced in this method.

The present invention, related to a method for alumina-bearing raw material digestion as part of processing thereof, is closest to the digestion method executed as part of M. Buchner’s method [29],

Alumina-bearing raw material digestion as part of processing thereof according to M. Buchner’s method includes preparing and heating of a reagent solution containing ammonium hydrosulfate; raw material treatment with the resulting solution and the resulting slurry separation into an undecomposed solid residue and a mother liquor of ammonium alum and other alums for subsequent extraction of aluminum compounds therefrom.

In contrast to the sintering methods [5, 6] and the methods [2528], where the alumina-bearing raw material is calcined together with ammonium sulfate, the raw material digestion by M. Buchner’s method neither includes direct involvement of the raw material in the calcining process, nor requires subsequent treatment at temperatures of least 400°C, in contrast to the method [39], thus reducing energy consumption. However, this method allows producing intermediate compounds that are easy to filter and clear from iron.

However, digestion as part of M. Buchner’s method is only applicable to easy to decompose aluminosilicate minerals (ref. to [33]), but not to the bauxitic raw materials, including high-silica bauxitic raw materials. Furthermore, though no high-temperature treatment (e.g., as common to the methods [25 - 28, 39]) is done, according to M. Buchner’s method it is still preferable to decompose the raw material in an autoclave at a relatively high temperature (at least 200°C), since the necessary completeness of the raw material decomposition will not be achieved at lower temperatures.

Summary of the Invention

A first of the provided inventions, related to a method for alumina-bearing raw material processing, is aimed at overcoming the above disadvantages of the closest known method and at accomplishing the technical result of providing for processing of any alumina-bearing raw mate10 rial, while reducing energy consumption in both the initial processing of the whole amount of the initial raw material through lower temperatures during such processing and at subsequent steps through enabling reduction in the amount of water circulated in the circular process; the invention is further aimed at reducing reagent losses and the amount of replenishment thereof during the circular process. Further types of the technical result to be accomplished may be pointed out below in this Summary of the Invention and the examples illustrating the invention.

The present method for alumina-bearing material processing, as well as the closest known method [29], is a circular process, comprising:

- a digestion step, including:

preparing and heating a reagent solution containing ammonium hydrosulfate;

decomposing the alumina-bearing raw material with the hot reagent solution to produce a slurry in the form of a solution of ammonium alum and other alums together with solid decomposition residues;

separation of the slurry into a solid and a liquid phases to obtain undecomposed solid residues and alum mother liquor;

- purification step where iron-free alum solution is obtained;

- precipitation step where aluminum hydroxide, precipitated from the de5 ironed alum solution, is obtained by treating the solution with ammonia;

- precipitated aluminum hydroxide separation step where a semi-product in the form of such hydroxide is produced, while obtaining a residual solution of ammonium sulfate produced at the precipitation step;

- solid ammonium sulfate production step;

- thermal decomposition step, where ammonium hydrosulfate and ammonia are produced for using at the digestion step to prepare the reagent solution and the precipitation step, respectively.

In contrast to the closest known method, to accomplish the above technical result, the method according to the present invention includes:

- reagent solution preparation at the digestion step, adding sulfuric acid thereto, and the slurry separation into the solid and the liquid phases at this step includes water washing of the solid residue, the alum mother liquor and washwater being collected separately;

- at the purification step, at least the latter of the alum mother liquor and the washwater is de-ironed by precipitation, after which the heated washwater and the alum mother liquor are combined to obtain a pre-purified mother liquor, the pre-purified mother liquor is then processed via a sequence of opera25 tions, including:

reducing the iron content of the liquor to ferrous iron, cooling the solution to crystallize ammonium alum, separating the alum crystals from the mother liquor and dissolving them in pure water to obtain a de-ironed alum liquor, the de-ironed alum liquor being delivered to the precipitation step;

- the method further comprises recovering sulfuric acid from the pre5 purified mother liquor, obtained at the purification step, from which ammonium alum crystals have been separated as part of the processing, to which end the liquor is flown through a column containing a strongbase anion exchange resin in a sulfate form, the sulfuric acid retarded at the anion exchange resin at the column is rinsed off with pure water, and the column continues in use, while the recovered sulfuric acid is returned to the head of the process to prepare the reagent solution at the digestion step, the mother liquor after processing, pre-purifying, ammonium alum crystal separating, and passing through the column, is combined with the ammo15 nium sulfate residual solution obtained at the separation step, the combined solution is then used as feedstock at the solid ammonium sulfate production step.

The alumina-bearing raw material processing according to the present invention, providing, instead of the treatment with ammonium hydro20 sulfate or mixed ammonium hydrosulfate and ammonium sulfate solution, for the raw material treatment with an ammonium hydrosulfate solution admixed with sulfuric acid, brings a totally new quality to the process, making it versatile and allowing processing, alongside with aluminosilicates (nephelines, kaolines, sillimanites, argillite, ashes, etc.), of hard to de25 compose alumina-bearing raw materials, such as bauxites and even red muds, i.e. bauxite processing waste.

It should also be noted that, being more efficient, the reagent solution according to the present method enables decomposing easy to decompose raw materials, for example, such aluminosilicates as nepheline, at temperatures lower than those of the M. Buchner’s methods (down to

75°C).

Such effect could not have been hypothesized, since the quantity of acid added to the reagent solution according to the present method is incommensurably less than required by the balance for an acid-based method. This effect appears even at an acid concentration of about 0.05% or less, including as low as the determination error. It may be suggested that, when added to the reagent solution, sulfuric acid acts as a catalyst. Initially, as its concentration is growing by few percent, the raw material decomposition rate in increasing, but hereafter the rate growth decelerates, thus making unfeasible the use of a concentration above 25%, since it does not provide further acceleration or increase in the decomposition grade, while making the process more complex and requiring more costly equipment.

It is sufficient to add sulfuric acid in an amount of few percent by weight, since this allows significantly increasing the solubility of potassium alum in the hot mother liquor after decomposition, and to bring the ammo20 nium hydrosulfate concentration in the reagent solution to 8.2 mol/1 (65 % by weight), thus providing for a substantial reduction in the amount of water, from which vapor is produced in the circular process, and, therefore, for a reduction in energy consumption.

The issue of silicon, which ingresses at excessive concentrations into the aluminum-bearing filtrate after the raw material decomposition as part of the process according to the M. Buchner’s method, is addressed by the present method as early as at the decomposition step. Where a reagent solution containing sulfuric acid is used, substantially all silicon remains in the solid phase, i.e. the undecomposed residue.

The present method for the process, where a concentrated reagent solution with a sulfuric acid content of several percent is used, substantially mitigates the de-ironing issue and allows using a low-cost sound scheme. This owes to the fact, that, if the present method is used, even with substantially equal (weight) content of aluminum and iron in the bauxites being decomposed, the aluminum content of the alum mother liquor resulting from the slurry separation is higher by more than an order of magnitude than the iron content. It further appears that commensurable concentrations of aluminum and iron are present in the washwater containing substantially lesser portion of dissolved aluminum compound, and, therefore, the washwater may be processed separately for de-ironing, and it is easier to de-iron than concentrated alum mother liquor. Furthermore, a substantial portion of iron is retained in the solid residue resulting from decomposition.

Based on the initial alumina-bearing raw material content of alkali and alkali-earth metals and magnesium forming non-hydrolysable sulfate salts, a certain quantity of a main reagent is lost in any circular process, in20 eluding that according to the M. Buchner’s method. This is illustrated by the chemical equation of decomposition of a bauxite with a typical molar composition:

0.57Al₂0₃‘0.23Fe₂0₃‘ (0,12Si0₂-0.05Ti0₂-0.01MgO ‘ 0.01K₂O ‘ ‘0.005Na₂0 · 0.005CaO) (1)

According to the Buchner’s method:

0.57Al₂0₃‘0.23Fe₂0₃‘(0.12Si0₂‘0.05Ti0₂‘0.01MgO ‘0.01K₂O ‘ ‘0.005Na₂0 0.005CaO) + 4.86 NH₄HSO₄ + 17.13H₂O = = 1.14A1NH₄(SO₄)₂T2H₂O + 0.46 FeNH₄(SO₄)₂· 12H₂O + + 1.63 (NH₄)₂SO₄+ 0.12H₂SiO₃ + 0.05Ti0₂-2H₂0 + + 0.01MgS0₄ ’ 7H₂O + 0.01K₂S0₄ ‘ H₂O + 0.005Na₂S0₄’10H₂O + + 0.005CaS0₄‘2H₂0 (2) and according to the present method:

0.57Al₂0₃’0.23Fe₂0₃'(0.12Si0₂’0.05Ti0₂'0.01MgO ’0.01K₂O ’ • 0.005Na₂0 · 0.005CaO) + 4.8 NH4HSO4 + 17.13H₂O + 0.03H₂S0₄ = = 1.14A1NH₄(SO₄)₂12H₂O + 0.46 FeNH₄(SO₄)₂T2H₂O + 1.6 (NH₄)₂SO₄ + 10 + 0.12H₂SiO₃ + 0.05Ti0₂‘2H₂0 + 0.01MgS0₄‘7H₂0 + + 0.01K₂S0₄ · H₂O + 0,005Na₂S0₄'10H₂O + 0.005CaS0₄'2H₂O (3)

The circular process is then closed by the processes formally described by the below chemical equations. For the two methods:

According to the Buchner’s method:

3.2 (NH₄)₂SO₄ + 1.63 (NH₄)₂SO₄ = 4.83 (NH₄)₂SO₄ (4)

4.83 (NH₄)₂SO₄ - 4.83 NH₃ + 4.83 NH4HSO4 (5)

According to the present method:

3.2 (NH₄)₂SO₄ + 1.6 (NH₄)₂SO₄ = 4.8 (NH₄)₂SO₄ (6)

4.8 (NH₄)₂SO₄ =4.8 (NH₄)₂SO₄ = 4.8 NH₃ + 4.8 NH4HSO4 (7)

In the former case, the main reagent, i.e. ammonium hydrosulfate, consumption is slightly greater than regeneration thereof in each cycle of the circular process; furthermore, the unrecoverable losses associated with sodium, potassium, calcium, and magnesium sulfate formation involve a reagent, i.e. ammonium hydrosulfate, that is more expensive than an acid. In the latter case, according to the present method, the hydrosulfate cycle is completely closed.

Further, according to the present method, the ammonia cycle is also completely closed (ref. to Equations (3) and (7)). The circular process according to M. Buchner’s method includes production of more ammonia than required, as is evident from comparison of Equations (3) and (5). Where a closed ammonia cycle is used, and only 4.8 mole of ammonia per one mole of the initial alumina-bearing source material is produced, the re10 agent becomes less “acid” (i.e., less efficient) at each subsequent cycle due to the ammonium hydrosulfate enriching with ammonium sulfate.

The provided method, through including an operation where sulfuric acid is recovered by passing the acid mother liquor, obtained by processing at the purification step and separated from ammonium alum crystals, through a column containing a strong-base anion exchange resin in a sulfate form, further allows returning sulfuric acid, which may be abundant in the liquors resulting from the alumina-bearing material decomposition, to the head of the circular process.

The described effects inherent to the provided method explain why

M. Buchner method’s [29] disadvantages may be overcome, and the technical result being the object of this invention may be accomplished, by the provided method.

Detailed description of the present invention will be elaborated below with reference to the preferred attributes of carrying out its operations, including in various embodiments thereof and subject to specific conditions in which the method is implemented.

It is inexpedient to prepare the reagent solution at the digestion step with ammonium hydrosulfate concentration below 5%, since such concentration does not allow digestion of even easily decomposable raw materials. Neither it is expedient to prepare the reagent solution with ammonium hydrosulfate mass concentration therein exceeding 65%, since it will be difficult to maintain solution of potassium alum after decomposition even at a temperature close to 100°C. Herein, the alum precipitates in the solid form and becomes part of the undecomposed residue, thus resulting in losses of aluminum or requiring the use of large amounts of washwater;

furthermore, substantial portion of the leached aluminum is contaminated with iron contained in the washwater in concentrations commensurable with those of aluminum. Overall, it reduces the process’ cost efficiency.

However, it is expedient to add sulfuric acid to the reagent solution until reaching a concentration of 1 to 5%. As explained above, within this range the rate of raw material decomposition is higher. Further increase in the acid content does not provide any substantial increase in the aluminum recovery degree, while resulting in additional costs associated with adding larger amounts of acid into the circular process and processing of larger masses of the reagent solution.

Temperature is maintained within the range of 75 to 180°C at the digestion step and further at the purification step up to and inclusive of the iron reduction to ferrous iron. Ammonium alum precipitates at temperatures below 75°C, thus not allowing the process to be carried out as appropriate. Decomposition rate increases with the temperature increase, how25 ever, increasing the temperature above 180°C results in extra costs that are not offset by the decrease in decomposition duration. Equipment employed in the method is based on selected temperature conditions, e.g., autoclave or microwave equipment is used where the temperature is above 100°C.

Under the above conditions, the decomposition duration may be 2 to 5 hours.

It is expedient to cool the pre-purified mother liquor, after deironing, ammonium alum crystallization, separation thereof from the liquor, and dissolving in pure water to obtain an iron-free alum solution, at a temperature of or below 20°C. At higher temperatures, a substantial portion of aluminum remains dissolved and does not become part of the final product, thus making the processing process uneconomical.

However, before dissolving ammonium alum crystals, crystallized and separated from the mother liquor, which has been cooled after the recovery operation, in pure water, it is preferable to wash the crystals with a concentrated ammonium sulfate solution cooled to at least 20°C. This fa15 cilitates finer clearing from residual iron.

Desalted water, obtained at the solid ammonium sulfate production step, may be used as the pure water to dissolve the above crystals.

It is reasonable to use a weight ratio of the reagent solution to the raw material under processing of at least 3:1. This is due to the fact that to provide for efficient decomposition it is necessary that the weight of the reagent solution components is greater than the aggregate weight of the raw material components, with which the former are interacting during the decomposition. Even if such raw material is supposed to contain alumina substantially free from iron impurities, the above ratio should be applied to a raw material with an alumina content of up to 15%. Any lower ratios would correspond to off-grade raw materials with lower alumina content. For the same reason, there is no point in applying a ratio greater than 10:1. Even if the aggregate content of aluminum oxides and iron is close to 100%, even such ratio provides an excess of the reagent solution components in equivalent calculus. Further increasing of the above ratio will not result in any substantial increase in the aluminum recovery degree, while resulting in energy losses associated with processing (evaporation) of a large amount of water in the circular process.

Slurry separation into the liquid and the solid phases (alum mother liquor and undecomposed solid residue) at the digestion step may be done by such known methods as filtration, centrifuging, decantation, being equivalent in terms of their effect on the capability to accomplish the technical result being the object of the present method.

Washwater (or washwater and alum mother liquor) de-ironing by precipitation at the purification step after the slurry separation at the digests tion step, including solid residue washing with water and separate collection of the alum mother liquor and washwater, may be done by ammoniating the washwater, i.e. by adding ammonia thereto, and separating the resulting iron hydroxide.

Herein, there is an increase in the liquor pH. It is expedient to pre20 vent the pH value rising above 4. This is due to the finding that in actual practice of the process according to the present method, where pH > 4, some of the recovered aluminum is lost as a result of its hydroxide coprecipitation with iron hydroxide.

Simultaneous boiling down and pH increasing improves the iron 25 hydroxide ex-solution.

The de-ironing is preferably performed by a polyacrylamide-based flocculent, thus facilitating aggregation of particles and improving separation thereof.

The above de-ironing of the washwater (or alum mother liquor and 5 washwater) is carried out until a ratio of at least 10:1 of aluminum weight concentration to that of iron in the pre-purified mother liquor is achieved. At lower ratios, the reducing agent amount that has to be added to the prepurified mother liquor during further processing is too high with reference to the resulting final product weight, thus making the process uneconom10 ical.

The purification step of the pre-purified mother liquor processing includes an operation for the mother liquor iron content reduction to ferrous iron, the reducing agent being one of ammonium sulfite, sulfur dioxide gas, or metallic aluminum powder. These reducing agents are preferable, since using them does not result in undesirable formation of components not present in the above liquor.

The reduction is done to prevent formation of ferriammonium sulfate co-crystallizing with ammonium alum.

Where the mass concentration ratio of aluminum to iron in the iron20 free alum solution, obtained by processing of the pre-purified mother liquor, is less than 1500:1, the solution is further de-ironed to achieve or exceed the above ratio by one of the molecular sorption, ion-exchange, liquidphase or solid-phase extraction methods using sorbents, ion-exchange resins or extracting agents with iron-selective functional groups.

This condition is due to the fact that it is expedient to precipitate aluminum hydroxide from a finely purified solution. The above ratio is exactly that, which provides for an iron oxide content in the alumina, to be further obtained from the aluminum hydroxide, of up to 0.05%, which meets the smelter-grade alumina standards.

The most advantageous way of sulfuric acid recovery from the prepurified mother liquor, obtained at the purification step of this liquor processing after ammonium alum crystals were separated therefrom, is NewChem Treatment, where a NewChem Column (as described below) is used as the column containing a strong-base anion exchange resin in a sulfate form, wherein the space not filled with the ion-exchange resin is filled with an organic liquid immiscible either with water or aqueous solutions.

Solid ammonium sulfate may be obtained, for example, by evaporating the solution resulting from combining the processed pre-purified alum mother liquor, flown through the NewChem Column, with the ammonium sulfate residual solution resulting from precipitated aluminum hydroxide separation.

Herein, desalted condensed water is further obtained, which may subsequently be used as the pure water to rinse off the acid retarded at the

NewChem Column anion exchange resin, and to dissolve the ammonium alum crystals isolated and separated from the pre-purified mother liquor cooled after the recovery operation.

The present method may further comprise a step, where alumina is obtained by dehydrating and calcining the semi-product, i.e. aluminum hy25 droxide, obtained at the precipitated aluminum hydroxide separation step.

The aforementioned NewChem Treatment, named so by the inventors herein (Khamizov, R.Kh., Krachak, A.N., Khamizov, S.Kh., Separation of ionic mixtures in columns with two liquid phases, Sorption and Chromatographic Processes, V.14, No.l, (2014), p. 14-23 [34]), is similar to that disclosed by RF Patent No. RU 2434679 (pubd. 27.11.2011, [35]), and is a variation of an acid retardation method: Hatch MJ and Dillon JA, Acid retardation. A simple physical method of separation of strong acids from their salts. I&EC Process Design and Development. 2/4: 253 - 263 (1963) [36],

Where used with sulfate media, its essence is that, when concentrated solutions of sulfate and sulfuric acid mixtures are flown through nanoporous sorbents, including gel strong-base anion exchange resins in a sulfate form, the acid’s low-hydration molecules and ion pairs, due to their smaller size, are retarded in the pores, while salt ion pairs of high hydration pass through the sorbent bed. The process is carried out in a cyclic mode, each cycle comprising a sorption stage and a desorption stage. After the acid “breakthrough” to a certain predetermined level at the sorption stage, the pure acid retarded at the column is desorbed plainly with water. This acid retardation method is limited by that, for an efficient acid from salt separation, the sorption column must have substantially no free space between the sorbent granules. To this end, the conventional commercial embodiment of the method employs tightly clamped highly-pressurized beds containing “compressed” sorbent granules (US Patent No. US 4,673,507, pubd. 16.06.1987 [37]; Sheedy M, Recoflo ion exchange technology. Pro25 ceedings of the TMS Annual Meeting held in 1998 in San Antonio Texas (1998) [38]). However, this approach is unfit for mixed colloidal systems or suspensions produced in iron-bearing solution processing (when neutral26 izing the passing through solution to retard the acid). According to the NewChem Technology, columns are employed, wherein the sorption bed is flooded with an organic liquid (e.g., decanol, pelargonic acid), being retained within the column during processing of mixed solutions containing acids and their salts. Operations with alumina-bearing material solutions become easier, since: a) supersaturated solutions and colloidal systems containing iron compounds are stabilized at NewChem Columns, and precipitation is occurring outside them; b) the solutions under processing are passed (in contrast to the conventional acid retardation method) from the top downwards, thus making easier the colloidal particle removal from the column.

Mild hydrolysis processes with iron hydroxide isolation, as well as with the acid recovery and return, according to the present method occur in accordance with the below chemical equations. They are provided for all processes occurring as part of processing an acid mother liquor containing the following components: (NH₄)₂SO₄, NH₄HSO₄, FeSO₄ (after recovery), NH₄Fe(SO₄)₂ (residual iron(III) after recovery), H₂SO₄, and H₂O (there is no reference to residual aluminum in the form of sulfate complexes, since ammonium alum is not subjected to mild hydrolysis with an anion20 exchange resin).

Retarding of a free acid remaining in the acid mother liquor delivered to the NewChem Treatment (R - anion-exchange resin functional group):

[R-so₄·· h₂o] + h₂so₄ = [R-SO₄··-h₂so₄] + H₂O (8)

Acid retardation as part of hydrolysis of residual iron (III) sulfate complex in the initial mother liquor with iron hydroxide precipitation in the filtrate downstream of the NewChem Column:

[R-SO₄· · H₂O] + NH₄Fe(SO₄)₂-12H₂O = Fe(OHy + NH₄HSO₄ + +[R-S0₄---H₂S0₄] + 10H₂0 (9)

Acid retardation as part of the reduced iron (II) sulfate hydrolysis to obtain iron hydroxide in the filtrate with subsequent oxidation thereof by aerial oxygen, including iron (II) hydroxide precipitation within this filtrate:

2[R-SO₄·· H₂O] +2FeSO₄ + 2H₂O = 2Fe(OH)₂+ 2[R-SO₄’·ySOJ (10)

2Fe(OH)₂+ H₂O+ V₂O₂ = 2Fe(OHy (11)

Rinsing (desorption) of dissolved sulfuric acid, retarded according to reactions (8) -(10), for returning thereof to the head of the circular process:

4[R-SO₄---H₂SO₄] +4H₂O = 4[R-SO₄---H₂O] + 4H₂SO₄ (12)

Sulfuric acid interaction with ammonium sulfate in the residual solution delivered to the ammonium sulfate recovery and processing steps:

4H SO, + 4(NH ) SO, = 8NH HSC)

4' 4⁷2 4 4 4 (13)

Lumped reaction in view of the processes of (8)-(13):

H₂SO₄ + NH₄Fe(SO₄)₂‘12H₂O + 2FeSO₄ + 4(NH₄)₂SO₄ + >/₂O₂ = 3Fe(OHy + 9NH₄ HSO₄ + 4H₂O (14)

Another important effect following from the above chemical equations and used in the present method is that the acid, returned to the circular process, enables ammonium hydrosulfate regeneration (under mild conditions, without thermal decomposition) from a substantial portion of residual ammonium sulfate. Herein, the ammonium hydrosulfate amount is equivalent to that of iron compounds involved in the mild hydrolysis process.

Finally, the method according to the first proposed invention employs another advantage of interfacing with the NewChem acid retardation technology, specifically, ferric iron reduction to ferrous iron in the combined solution before precipitation and separation of ammonium alum. Such reduction allows carrying out the processes described by Equations (10) and (11), which, where performed, substantially reduce the likelihood of iron (III) hydroxide precipitation in the sorption column.

A second proposed invention, related to a method for alumina15 bearing raw material digestion as part of processing thereof, is aimed at accomplishing the technical result of providing for processing of any alumina-bearing raw material without high-temperature (200°C and above) treating thereof, while increasing the aluminum compound decomposition and recovery degrees.

The present method for alumina-bearing material digestion, as well as the closest known digestion method disclosed by M. Buchner’s Patents [29], comprises: preparing and heating a reagent solution containing ammonium hydrosulfate, decomposing the alumina-bearing raw material with the reagent solution to produce a slurry in the form of a solution of ammo25 nium alum and other alums together with solid decomposition residues, and separation of the slurry into a solid and a liquid phases to obtain undecomposed solid residues and alum mother liquor

In contrast to the closest known method, to accomplish the above technical result, the digestion method according to the present invention comprises: reagent solution preparation includes adding sulfuric acid thereto; and the slurry separation into the solid and the liquid phases includes water washing of the undecomposed solid residue and separately collecting the alum mother liquor and washwater for further use as solutions for subsequent extraction of aluminum compounds therefrom.

The method according to the present invention, providing for alumina-bearing raw material decomposition by an ammonium hydrosulfate solution admixed with sulfuric acid, makes the process versatile and allowing digestion, alongside with aluminosilicates, of hard to decompose alumina-bearing raw materials, such as bauxites and even red muds, i.e. baux15 ite processing waste. Being more efficient, the reagent solution according to the present method enables lower-temperature decomposition of easy to decompose raw materials, for example, such aluminosilicates as nepheline,

i.e., the method becomes versatile, while employment of the described techniques in separating the resulting slurry ultimately increases the degree of aluminum compound extraction.

The effect appears even at an acid concentration of about 0.05% or less, including as low as the determination error. Initially, as its concentration is growing by few percent, the raw material decomposition rate in increasing, but hereafter the rate growth decelerates, thus making unfeasible the use of a concentration above 25%, since it does not provide further ac30 celeration or increase in the decomposition grade, while making the process more complex and requiring more costly equipment.

The present method, where a concentrated reagent solution with a sulfuric acid content of several percent is used, even with substantially equal (weight) content of aluminum and iron in the bauxites being decomposed, the aluminum content of the alum mother liquor resulting from the slurry separation is higher by more than an order of magnitude than the iron content. It further appears that commensurable concentrations of aluminum and iron are present in the washwater, which, if collected separately, may subsequently be processed separately.

It is inexpedient to prepare the reagent solution with ammonium hydrosulfate concentration below 5%, since such concentration does not allow decomposing of even easily decomposable raw materials, such as nepheline. Neither it is expedient to prepare the reagent solution with am15 monium hydrosulfate mass concentration therein exceeding 65%, since it will be difficult to maintain solution of potassium alum after decomposition even at a temperature close to 100°C. Herein, the alum precipitates in the solid form and becomes part of the undecomposed residue, thus resulting in losses of aluminum.

It is expedient to add sulfuric acid to the reagent solution until reaching a concentration of 1 to 5%. As explained above, within this range, with an increase in concentration, the rate of raw material decomposition is higher. Further increase in the acid content has little effect on the decomposition rate, and, therefore, is inexpedient, since it results in increased acid consumption and the need to process a larger mass of the reagent solution.

When implementing the present method, temperature is maintained within the range of 75 to 180°C. Ammonium alum precipitates at temperatures below 75°C, thus not allowing the process to be carried out as appropriate. Decomposition rate increases with the temperature increase, how5 ever, increasing the temperature above 180°C results in extra costs that are not offset by the decrease in decomposition duration. Equipment employed in the method is based on selected temperature conditions, e.g., autoclave or microwave equipment is used where the temperature is above 100°C.

Under the above conditions, the decomposition duration may be 2 to 5 hours.

It is reasonable to use a weight ratio of the reagent solution to the raw material to be liberated of at least 3:1. This is due to the fact that to provide for efficient decomposition it is necessary that the weight of the reagent solution components is greater than the aggregate weight of the raw material components, with which the former are interacting during the decomposition. Even if such raw material is supposed to contain alumina substantially free from iron impurities, the above ratio should be applied to a raw material with an alumina content of up to 15%. Any lower ratios would correspond to off-grade raw materials with lower alumina content. For the same reason, there is no point in applying a ratio greater than 10:1. Even if the aggregate content of aluminum oxides and iron is close to 100%, even such ratio provides an excess of the reagent solution components in equivalent calculus.

Slurry separation into the liquid and the solid phases (alum mother liquor and undecomposed solid residue) may be done by such known methods as filtration, centrifuging, decantation, being equivalent in terms of their effect on the capability to accomplish the technical result being the object of the present alumina-bearing raw material digestion method.

Brief Description of the Drawings

The present inventions are illustrated in the diagrams of Figs. 1-6 and the examples provided below.

The diagram, presented in multiple parts in Figs. 1-4, shows a sequence and an interrelation of the operations according to the first of the present inventions, related to a method of processing of alumina-bearing materials, without the features characterizing the method in particular em10 bodiments thereof.

The diagram, presented in two parts in Fig. 5 and Fig. 6, is a simplified illustration of an industrial process of alumina-bearing raw-material processing according to the above method.

The examples also provide a general illustration of the alumina15 bearing material processing method, including a number of features preferable in various particular embodiments of the method and various possible combinations of such features.

The method according to the second of the proposed inventions is illustrated by the pertinent parts of the above diagrams and the pertinent examples, since it is a sub-process of the method according to the first invention. Thus, pertinent to the second invention are: Fig. 1, Blocks 1-3, Fig. 5, paragraphs A, B, and Example 1 paragraph C sub-paragraph 1, as well as Examples 4, 5, 36 so far as they pertain to the above paragraphs of Example 1, and the table to Examples 6-35.

Detailed Description of the Embodiments

The diagram shown in Figs. 1 - 4 is a graphic representation of the alumina-processing method features in wordings, sometimes abbreviated, but still close to those used above in the summary of this invention. Herein, the rounded blocks pertain to the features of the invention. The diagram parts in Fig. 1, Fig. 2, Fig. 3, Fig. 4 are marked A, B, C, D, E, F, G, H to show the interconnection of the parts.

The below notes to the diagram, shown in Fig. 5 and Fig. 6, and illustrating an industrial process, use the terms specific to such purpose of the diagram and intended for those skilled in process engineering, and, therefore, may, in some instances, be somewhat different from those used in the summary of this invention. This reason also determines the sequence of description.

The diagram parts in Fig. 5, Fig. 6 are marked K, L, Μ, N, P, to show the interconnection of the parts. The diagram depicts a closed circular process of alumina-bearing material processing, wherein a reagent solution is prepared using a portion of desalted water obtained at the final steps of the process, specifically at the solid ammonium sulfate production step (Block 8). Further, washwater resulting from aluminum hydroxide extrac20 tion from the de-ironed alum solution during ammoniation thereof (Block 7; the washwater is shown in the diagram as “washwater 2”).

Further, the reagent solution is produced using the solid ammonium hydrosulfate obtained as part of the circular process after the solid ammonium hydrosulfate separation and decomposition operation (Block 9). Fur25 ther still, the reagent solution is prepared using partially the acid returned from the NewChem Treatment step (Block 11), and partially a fresh batch of technical sulfuric acid.

The resulting reagent solution is heated (Block 1), and the hot solution is mixed with ground alumina-bearing raw material. Decomposition (Block 2) includes holding the mixture hot in a non-sealed (open) reactor or autoclave (subject to the type of raw material under processing) for 2 to 5 hours. The slurry resulting from decomposition is separated (Block 3), for example, by filter presses or vacuum filters to obtain a primary hot filtrate (alum mother liquor separated from solid decomposition residue). The fil10 ters containing the above residue are then washed with hot water (e.g., with process softened water), thus obtaining the washwater.

The washwater is boiled down and partially ammoniated (Block 4) to convert the acid solution to a weakly acid solution enabling iron hydroxide precipitation and separation, for example, by filtering.

The washwater remaining after filtration (hot) is combined with the hot primary filtrate to obtain a combined filtrate (i.e., pre-purified mother liquor), where a reducing agent, e.g., ammonium sulfite, is added to (Block 5). After reduction, the resulting solution is cooled for the ammonium alum to crystallize, the crystals are separated from the acid pre-purified mother liquor via filtration (and, optionally, washed with cool concentrated ammonium sulfate solution as is shown below in Example 1, though not shown in this diagram), the crystals are then dissolved in desalted water to obtain a de-ironed alum solution, briefly termed “secondary solution” in the diagram (Block 6).

The secondary solution is ammoniated (Block 7) with ammonia obtained at the final steps of the process, i.e., thermal decomposition of the previously obtained solid ammonium sulfate (Block 9). Herein, pure aluminum hydroxide is isolated, filtered from the ammonium sulfate residual solution, and washed with desalted water at a filter (Block 7).

Washed aluminum hydroxide is delivered for drying and calcining (Block 11) to obtain smelter-grade alumina. (Optionally, the secondary solution is subjected to finer de-ironing through sorption or extraction methods; not shown in the diagram).

Acid mother liquor, from which alum crystals have been separated (Block 5), undergoes the NewChem treatment (Block 10). Sulfuric acid solution, obtained at each cycle of the treatment, is returned to the head of the circular process, i.e. to the reagent solution preparation step (Block 1), while the NewChem filtrate, i.e. acid-free solution, is allowed to stand for iron hydroxide precipitation (and, optionally, oxidized with aerial oxygen, though it is not shown in the diagram), the hydroxide is separated, and the resulting solution is combined (Block 8) with the ammonium sulfate residual solution.

The combined solution is subjected to evaporation and ammonium sulfate crystallization (Block 8). Condensate from evaporation, i.e., desalted water, is used at the digestion step to produce the reagent solution (Block 1), to dissolve pure ammonium alum (Block 6), separated from the mother liquor, to wash aluminum hydroxide, and to desorb (rinse off) the acid solution at the NewChem Treatment step (Block 10) (herein, bottom residues of a brine after crystallization are processed by adding lime thereto, and sodium sulfate and potassium sulfate crystallization, as de25 scribed in Example 1, though not shown in the diagram).

The resulting ammonium sulfate is dehydrated, e.g., via centrifuging and drying, and is subjected to thermal decomposition (Block 9) to isolate ammonia to be used for ammoniation (Block 7) of the de-ironed alum solution (secondary solution) and ammonium hydrosulfate to be delivered to the head of the process, i.e., the digestion step, to prepare the reagent solution (Block 1).

Blocks 1 - 3 in Fig. 5 pertain to the second of the present inventions, related to a method for alumina-bearing raw material processing that concludes with separate collection of the alum mother liquor (termed “pri10 mary filtrate” in the diagram) and washwater resulting from the slurry separation into the liquid and solid phases. The relevant operations are shown in Fig. 1.

Example 1

A. “Pur.” grade chemical reagents or commercial “Tech.” grade chemicals are used to prepare 200 g of hot (75°C) reagent solution containing 96 g of ammonium hydrosulfate (59.4%), 4.3 g of aluminum and ammonia sulfate, and 1 g of 94% technical sulfuric acid (0.47 % of pure acid). (Softened mains water (61.7 ml) is used in the first process cycle. In subsequent cycles, a certain amount of washwater, resulting from the semi20 product, i.e., aluminum hydroxide, washing at the final stages, is used. See below, paragraph J.)

B. The reagent solution is poured into a reactor/autoclave, where 26.3 g of Timanian bauxites are further added, the bauxite being composed as follows (wt %):

A1₂O₃ - 47.7; Fe₂O₃ - 28.3; SiO₂ - 8.0; TiO₂ -2.8; K₂O - 0.63; MgO 0.39; Na₂O - 0.23; P₂O₅ - 0.22; SO₃ - 0.2; CaO - 0.17; MnO - 0.04;

3Ί

Cr₂O₃ - 0.04; V₂O₅ - 0.04; loss on calcining (LOC) represented mainly by water - 11.25.

The slurry is autoclaved during 3 hours at 150°C;

C. 1) The resulting mixture is vacuum filtered in a Buchner ther5 mostatted suction funnel with a water jet pump at 95°C. Herein, 131 g of primary filtrate is obtained and maintained at the temperature of 95°C. The filter is washed with 125 g (ml) of hot water with the temperature of 95°C. A cake remains at the filter. The washwater is evaporated in a flask with a refrigerator condenser down to the residual volume of 60 ml. Herein, 65 ml (65 g) of condensate is collected and kept for using in the next cycle.

2) The washwater residual volume is added with ammonium solution until pH = 3 is achieved (a total of 15 g of 24% aqueous ammonium solution is added, ref. to paragraph G.), the resulting slurry is filtered through the Buchner suction funnel without changing the filter with the above cake to obtain a combined precipitate (Precipitate 1) with the weight of 19.2 g, containing the undissolved portion of bauxites, hydrated iron, titanium, vanadium, chromium, and magnesium oxides, calcium sulfate (gypsum), and moisture water. As a result of the above procedure, 77 g of treated washwater filtrate is obtained, which is heated and combined with the primary filtrate to obtain 208 g of combined filtrate (Filtrate 1).

D. Hot Filtrate 1 is added with 2 g of 15% ammonium sulfite solution, and the resulting mixture is heated to 20°C. Herein, alum crystallization occurs, the crystals being separated from the mother liquor via filtering through the Buchner funnel with a jet water pump at room temperature.

The filtrate from the acid mother liquor (Filtrate 2, 97 g) is delivered to the next processing step (ref. to the next paragraph E). The obtained crystals are washed at 20°C with 200 g of 42.1% ammonium sulfate solution, the wash solution (also termed “Filtrate 4”, 201.2 g) is kept for using at an appropriate processing step (ref. to paragraph F). Potassium alum in the amount of 111 g with the moisture content of 20% (88.9 g of pure alum) is obtained. The obtained alum is mixed during heating (75°C) with 58 g of condensate (desalted) water to obtain a de-ironed alum solution. The solution is added with 42 g of 24 % ammonium solution (ref. to paragraph G). A slurry with the weight of 220 g is obtained, containing aluminum hydroxide in ammonium sulfate solution. The slurry is filtered through the

Buchner funnel with the water jet pump to obtain Filtrate 3 in the amount of 200 g and a precipitate in the amount of 19.13 g, containing aluminum hydroxide and moisture (20%). Filtrate 3 is used as a solution to wash alum in the subsequent process cycle. The precipitate is processed at the next steps (ref. to paragraph J).

E. Filtrate 2 (acid mother liquor) in the amount of 97 g, obtained according to paragraph D, is passed through a NewChem Column containing 45 ml of AV-17x8 strong-base anion exchange resin in a sulfate form and dodecane filling the free space. The passing rate is 50 ml/h. A supersaturated neutral solution of iron (II) and (III) hydroxides is obtained at the column outlet. Full conversion is achieved by blowing air through the obtained filtrate during 5 minutes by a mini-compressor (e.g., for aquariums). The obtained weak slurry is filtered, and a wet cake is separated in the amount of 0.6 g, which is a substantially pure iron hydroxide. The cake is sent for disposal or recycling. Herein, 95.5 g of filtrate (Filtrate 5) is ob25 tained, which is a concentrated ammonium sulfate solution admixed with soluble magnesium, potassium, and sodium sulfates. Filtrate 5 is kept for use at the next steps of the process (ref. to the next paragraph F).

F. Filtrate 4 (ref. to paragraph D) and Filtrate 5 (ref. to paragraph E) are combined to obtain a combined ammonium sulfate solution that is crystallized to a damp sand condition by evaporating the solution with the use of a refrigerator condenser. Residual brine in the amount of 1.8 g is fil5 tered by suction via the Buchner funnel with the water jet pump to obtain wet crystalline ammonium sulfate in the amount of 138 g with the moisture content of 20%. The resulting condensate (desalted water) in the amount of 1158.3 g is divided into three portions. One portion of 57.3 g is used in the next process cycle at the alum precipitation step according to paragraph

D. The second portion (66 g) is also used in the next cycle to make various process solutions according to paragraphs C, D, G, and Ή, and the third portion (35 g) is used in the subsequent cycles to prepare the reagent solution.

G. The crystalline ammonium sulfate is dried and calcined at

275°C to obtain 96 g of ammonium bisulfate (admixed with 4.2 g of dry alum). Herein, the evolving ammonia is passed through desalted (condensate) water (52.8 g, starting from the second cycle, returned from the ammonium sulfate crystallization and evaporation stage according to paragraph F) to obtain 57 g of 24 % ammonium aqueous solution for use at various steps of the subsequent process cycle: to achieve pH=3 of the evaporated solution as part of isolating the bulk of iron according to paragraph C (15 g) and to precipitate aluminum hydroxide according to paragraph D (42 g).

H. Condensate (or deionized) water in the amount of 6.2 ml (6.2 g) is passed through the NewChem Column after the acid has been retarded at it as described in paragraph E from the top downwards with the flow rate of 20 ml/h. As a result, 7.5 g of 17.1% of sulfuric acid solution is ob40 tained and delivered to the reagent solution preparation step in the next process cycle according to paragraph A.

I. The residual brine (1.8 g), obtained according to paragraph F, is added with 0.4 g of calcium hydroxide in the form of 22.2% lime milk (1.8 g). The evolving ammonia (0.2 g) is delivered to the step of reagent solution preparation according to paragraph A for use in the next process cycle. Herein, 3.6 g of slurry is obtained, which is evaporated, and the resulting residue is sent for disposal or recycling. As a result of this operation, a precipitate with the weight of 2.6 g and moisture content of 20 %, comprising magnesium and aluminum hydroxides, as well as gypsum and potassium and magnesium sulfates.

J. Wet aluminum hydroxide precipitate (19.13 g), obtained according to paragraph D, is washed at an ashless paper filter using the Buchner funnel with a vacuum pump and 50 ml of deionized water (or condensate).

The washwater (50 g) is kept for use in the next cycle to prepare the reagent solution according to paragraph A, and, partially, to make up for losses from evaporating washing solutions at the precipitate washing stage according to paragraph C. The washed precipitate, together with the paper filter, is dried in an oven at 120°C and calcined at a muffle furnace at

970°C during 2 hours. Pure aluminum oxide in the amount of 10.02 g is obtained with the content of primary components corresponding to the smelter-grade alumina (GO) quality. Herein, the total alumina recovery degree is 80.01%.

K. A next process cycle is carried out according to paragraphs A

- J, except that a portion of the desalted water (condensate) obtained in the previous cycle according to paragraph F, specifically, 35 g, is used to prepare the reagent solution; a further portion, specifically, 66 g, is used for: washing the precipitate resulting from bauxite decomposition and filtration (5.3 g, ref. to paragraph C); preparing the ammonium sulfite solution (1.7 g, ref. to paragraph D); preparing the ammonia aqueous solution (52.8 g, ref. to paragraph G); acid desorbing from the NewChem Column (6.2 g, ref. to paragraph Ή). Further, the washwater obtained in the previous cycle at the stage of semi-product, i.e., aluminum hydroxide, washing according to paragraph J, specifically, 50 g, is used for: washing the precipitate (19.7 g, ref. to paragraph C) resulting from bauxite de10 composition and filtration; preparing the reagent solution (28.9 g according to paragraph A); preparing the lime milk (1.4 g, according to paragraph I).

L. Any subsequent process cycles are carried out exactly in accordance with paragraph K. Herein, a circular process completely closed with respect to the main reagent, i.e. ammonium sulfate (and bisulfate). Component intake to obtain 10 g of smelter-grade alumina in each cycle is as follows: 2.63 g of bauxites (with the aluminum recovery degree of 85% at the raw material decomposition step and the total recovery degree of 80%); 1 g of technical (94%) sulfuric acid; 0.3 g of ammonium sulfite; 0.4 g of technical calcium hydroxide; 11.7 ml of softened mains water.

Example 2

The process is carried out according to Example 1, except that Filtrate 1 is added with 0.15 g of aluminum chips at the step described in paragraph ”D”.

At the final step, 7.15 g of pure aluminum oxide is obtained with the content of primary components corresponding to the smelter-grade alumina (GO) quality. Herein, the total alumina recovery degree is 75 %.

Example 3

The process is carried out according to Example 1, except that 70 ml (0.2 g) of sulfur dioxide gas is flown through Filtrate 1 with vigorous mixing at the step described in paragraph ”D”.

At the final step, 7.05 g of pure aluminum oxide is obtained with the content of primary components corresponding to the smelter-grade alumina (GO) quality. Herein, the total alumina recovery degree is 74 %.

Example 4

The process is carried out according to Example 1, except that:

- the reagent solution is prepared in the amount of 200 g, containing 45% of ammonium hydrosulfate and 1% of sulfuric acid;

- the raw material is 52 g of the Kaichak Deposit kaolin clay with the following composition (wt %):

A1₂O₃ -18.3; Fe₂O₃ - 2.7; SiO₂ - 64.2; TiO₂ -1.7; K₂O - 2.0; MgO -0.81; Na₂O- 2.1; P₂O₅ - 0.15; CaO - 0.87; MnO - 1.0; LOC -7.0.

- the raw material decomposition process is carried out during 5 hours in an open reactor at 98°C.

The recovery degree achieved through decomposition at the digestion step is 82%.

At the final step, 7.42 g of pure aluminum oxide is obtained with the content of primary components corresponding to the smelter-grade alumina (GOO) quality. Herein, the total alumina recovery degree is 78 %.

Example 5

The process is carried out according to Example 1, except that:

- the reagent solution is prepared in the amount of 200 g, containing 45% of ammonium hydrosulfate and 3% of sulfuric acid;

- the raw material is 35 g of the Kola Deposit nepheline concentrate with the following composition (wt %):

A1₂O₃ - 28.0; Fe₂O₃ - 2.4; SiO₂ - 44; TiO₂ - 0.55; K₂O - 7.6; MgO - 0.45;

Na₂O - 12.1; P₂O₅ - 0.17; CaO - 1.75; SrO - 0.11; MnO - 0.08; LOC -1.5.

- the raw material decomposition process is carried out during 5 hours in an open reactor at 98°C.

The recovery degree achieved through decomposition at the digestion step is 96%.

At the final step, 8.8 g of pure aluminum oxide is obtained with the content of primary components corresponding to the smelter-grade alumina (GOO) quality. Herein, the total alumina recovery degree is 90 %.

Examples 6-35

Alumina-bearing raw materials of the above three types are decomposed at different temperatures, but under the same conditions as follows: the reagent solution containing hydrosulfate in the amount of 40% and sulfuric acid in the amount of 1%; L:S = 10:1, decomposition duration is 3 hours. The nepheline concentrate and the bauxite material are then subjected to decomposing under the same conditions, but according to M. Buchner’s method (without adding the acid to the reagent solution). The obtained data on the degree of aluminum extraction from the raw materials are given in the below table:

Temperature (°C)	Nepheline concentrate, Buchner method	Nepheline concentrate, the present method	Kaolin clay, the present method	Bauxite, Buchner method	Bauxite, the present method
In a non-sealed reactor
85	48	66	52	18	55
90	72	95	59	44	58
98	85	99	59	48	60
In an autoclave
120	98	98	65	60	81
150	92	100	80	62	86
175	96	99	87	65	89

The obtained results demonstrate that in all of the above cases the present 10 methods are substantially more efficient than M. Buchner’s method, either overall or in so far as the alumina-bearing material digestion is concerned.

Example 36

The process is carried out according to Example 1, except that:

- the reagent solution is prepared in the amount of 200 g, containing 45% of ammonium hydrosulfate and 1% of sulfuric acid;

- the raw material is 35 g of red mud, i.e., alumina refining waste, with the following composition (wt %):

A1₂O₃ - 12.4; Fe₂O₃ - 44.3; SiO₂ - 9.3; TiO₂ - 4.4; K₂O - 0.1; MgO 0.93; Na₂O -2.9; P₂O₅ - 0.75; CaO - 12.3; SrO - 0.11; MnO - 0.52; LOC -7.5.

- the raw material decomposition process is carried out during 3 hours in an open reactor at 130°C.

The aluminum recovery degree achieved through decomposition at the digestion step is 75%.

Slurry separation and washing at the purifications step provide the following results:

- the primary filtrate contains substantially 90% of the recovered aluminum, the aluminum to iron weight ratio being 5:4;

- the washwater contains iron and aluminum at the ratio of 15:1. Herein, the total iron recovery degree is at or below 50%.

- more than 95% of silicon, titanium, and calcium remain in undecomposed residue.

Industrial Applicability

The provided description and example embodiments of the proposed methods demonstrate that the methods may be advantageously used for alumina-bearing raw material processing to obtain smelter-grade alu46 mina and by-products, and for the alumina-bearing raw material digestion prior to further processing thereof. The examples further demonstrate that the present method may be used even for waste, specifically, red mud, recycling. The washwater composition enables obtaining, via ammoniation, a mixture of easy to decompose aluminum and iron hydroxides at a ratio characteristic to high-grade bauxites, while iron oxide pigments or ore for the iron and steel industry may be obtained from the washwater.

References

1. Lainer, A.I., Yeremin N.I., Lainer Yu.A., Pevzner I.Z. Alumina Refining.

M., Metallurgy, 1978, 394 pages

2. Troitskiy I.A., Zheleznov V.A. Metallurgy of Aluminum. - M., Metallurgy, 1977, p. 42-116.

3. RF Patent No. RU 2360865, pubd. 10.07.2009.

4. RF Patent No. RU 2193525, pubd. 27.11.2002.

5. Matveev, V.A. Physical and Chemical and Technological Bases for Improving Efficiency of Nepheline-Bearing Raw Material Integrated Processing by Acid Methods. A Thesis for PhD in Engineering, Apatyty, 2009, 299 pages.

6. Isakov, E.A., Pikalevo Industrial Group “Glinozyom” in New Realities.

Non-Ferrous Metals, 1997, No.4, p. 8.

7. Vaitner, V.V., A Research into Nitric Acid Processing of Aluminosilicates to Produce Aluminum Oxide. A Thesis for Candidate in Engineering Sciences, Yekaterinburg, 2004, 146 pages.

8. RF Patent No.RU 2202516, pubd. 27.11.2002.

9. RF Patent No.RU 2215690, pubd. 10.11.2003.

10. RF Patent No.RU 2372290, pubd. 10.11.2009.

11. RF Patent No.RU 2460691, pubd. 10.09.2012.

12. Schwartzman, B.H., Acid-Based Methods for Alumina-Bearing Raw

Material Processing. M.: Tsvetmetinformatsia, 1964. 82 pages.

13. Pustilnik, G.F., Pevzner, I.Z., Acid-Based Methods for Processing LowGrade Aluminum-Bearing Raw Materials. M.: Tsvetmetinformatsia, 1978. 52 pages.

14. Herbert Weissenbaeck, Benedikt Nowak, Dieter Vogl, Horst Krenn,

Development of Chloride Based Metal Extraction Techniques - Advancements and Setbacks, Proceedings of Nickel-Cobalt-Copper Conference of AFTA-2013, 29 May - 1 June, 2013 Perth, WA., Melbourne, Australia, p. 360.

15. RF Patent No.RU 2471010, pubd. 27.12.2012.

16. Fainer, Yu.A., Integrated Processing of Aluminous Raw Materials by Acid-Based Methods. M.: Nauka, 1982, p. 208.

17. Zapolskiy, A.K., Sazhin, V.S., Zakharova, N.N., Crystallization of Basic Aluminum Sulfate Salts. Alumina Chemistry and Technology. Novosi20 birsk, Nauka, 1971, p. 430 - 438.

18. Zapolskiy, A.K., Sulfuric Acid Processing of High Silica Aluminous Raw Material. Kyiv, Naukova Dumka, 1981, p. 198 -200.

19. Paweena Numluk and Aphiruk Chaisena. Sulfuric Acid and Ammonium Sulfate Leaching of Alumina from Lampang Clay!I E-Joumal of Chemistry. 2012. V 9, No. 3. p. 1364-1372, http://www.ejchem.net.

20. RF Patent No.RU 2337877, pubd. 10.11.2008.

21. Sandler, E.M., Lainer, Yu.A., Lainer, A.I., Chizhikov, D.M. Product

De-Ironing in Processing Nephelines by Sulfuric Acid-Based Method. NonFerrous Metals Industry. News of Higher Education Institutions, 1962, No.2, p. 30 -33.

22. Funaki K. Sulfuric acid process for obtaining pure aluminum oxide from its oresH Bull, of the Tokyo Inst, of Technology, 1980, No.l.

23. RF Patent No.RU 2440296, pubd. 20.01.2012.

US Patent No.US 1,426,891, pubd. 22.08.1922.

USSR Inventor’s Certificate No. SU 42067, pubd. 31.03.1935.

26. Ullmann B. Encyklopadie der technischen Chemie, Auflage, Urban &

Schwarzenberg, Miinchen & Berlin. 1954, Bd. 3, 401 - 420.

27. Grim R.E. Applied Clay Mineralogy, McGraw-Hill, New York, 1962, p. 335-345.

28. G. Bayer, G. Kahr, and M. Mueller-Vonmoos. Reactions of ammonium sulphates with kaolinite and other silicate and oxide minerals, Clay Miner20 als, 1982, V.17,p. 271 -283.

29. British Patent No.GB 195,998, pubd. 12.04.1922; US Patent No.US 1,493,320, pubd. 06.05.1924; USSR patent No.SU 11489, pubd. 30.09.1929.

30. Encyclopedic Dictionary of Named Processes in Chemical Technology /Ed. Alan E. Cornyns Boca Raton: CRC Press LLC, 2000, 2-nd Ed., 312 p., p.19 (Aloton).

31. Bretsznajder St. Otrzymywanie estow kwasu ortokrzemowego w fazie 5 gazovey / St. retsznajder, W. Kawecki // Rocz. Chem,- 1955,- N° 29. - s.

287-299.

32. Bretsznajder St. Nova metoda otrzymywania hutniczego tlenku glinovego i innych zwiazkow glinu z glin // Przem. Chem.- 1963,- V.42, 12.

- s. 677-683.

33. O’Connor, D. J. Alumina Extraction from Non-bauxitic Materials,

Aluminium-Verlag, Diisseldorf, 1988, 159 p.

34. Khamizov, R.Kh., Krachak, A.N., Khamizov, S.Kh., Separation of ionic mixtures in columns with two liquid phases, Sorption and Chromatographic Processes, V.14, No.l, (2014), p. 14-23.

35. RF Patent No.RU 2434679, pubd. 27.11.2011.

36. Hatch MJ and Dillon JA, Acid retardation. A simple physical method of separation of strong acids from their salts. I&EC Process Design and Development. 2/4: 253-263 (1963).

37. US Patent No.US 4,673,507, pubd. 16.06.1987.

38. Sheedy M, Recoflo ion exchange technology. Proceedings of the TMS

Annual Meeting held in 1998 in San Antonio Texas (1998).

39. RF Patent No.RU 2153466, pubd. 27.07.2000.

Claims (19)

1. A method for alumina-bearing raw material processing via a circular process, comprising: a digestion step including preparing and heating a reagent solution containing ammonium hydrosulfate, decomposing the alumina-bearing raw material with the hot reagent solution to produce a slurry containing an ammonium alum solution together with solid decomposition residues, separating the slurry into a solid and a liquid phases to obtain undecomposed solid residues and alum mother liquor; a purification step where iron-free alum solution is obtained; a precipitation step where aluminum hydroxide, precipitated from the deironed alum solution, is obtained by treating the solution with ammonia; a precipitated aluminum hydroxide separation step where a semi-product in the form of such hydroxide is produced, while obtaining a residual solution of ammonium sulfate produced at the precipitation step; a solid ammonium sulfate production step; a solid ammonium sulfate thermal decomposition step, where ammonium hydrosulfate and ammonia are produced for using at the digestion step to prepare the reagent solution and the precipitation step, respectively, wherein the reagent solution preparation at the digestion step includes adding sulfuric acid thereto, and the slurry separation into the solid and the liquid phases at this step includes water washing of the undecomposed solid residue, the alum mother liquor and washwater being collected separately; at the purification step, at least the latter of the alum mother liquor and the washwater is de-ironed by precipitation, after which the heated washwater and the alum mother liquor are combined to obtain a pre-purified mother liquor, the pre-purified mother liquor is then processed via a sequence of operations, including: reducing the iron content of the liquor to ferrous iron, cooling the solution to crystallize ammonium alum, separating the alum crystals from the mother liquor and dissolving them in pure water to obtain a de-ironed alum liquor, the de-ironed alum liquor being delivered to the precipitation step; the

2015328791 12 Dec 2018 method further comprises recovering sulfuric acid from the pre-purified mother liquor, obtained at the purification step, from which ammonium alum crystals have been separated as part of the processing, to which end the liquor is flown through a column containing a strong-base anion exchange resin in a sulfate form, the sulfuric acid retarded at the anion exchange resin at the column is rinsed off with pure water, and the column continues in use, while the recovered sulfuric acid is returned to the head of the process to prepare the reagent solution at the digestion step, the mother liquor after processing, pre-purifying, ammonium alum crystal separating, and passing through the column, is combined with the ammonium sulfate residual solution obtained at the precipitated aluminum hydroxide separation step, the combined solution is then used as feedstock at the solid ammonium sulfate production step.

2. The method of claim 1, wherein the digestion step includes preparing a reagent solution containing 5 to 65 weight percent of ammonium hydrosulfate and adding sulfuric acid to the reagent solution until its concentration reaches 1 to 5%.

3. The method of claim 2, wherein the alumina-bearing raw material is decomposed with the hot reagent solution at the digestion step at a weight ratio of the reagent solution to the alumina-bearing raw material of 3 : 1 to 10 : 1.

4. The method according to each of claims 1-3, wherein de-ironing of the washwater or the washwater and the alum mother liquor by precipitation at the purification step, after the said separation of the slurry at the digestion step and water washing of the said solid residue, the alum mother liquor and washwater being collected separately, is done by adding ammonia to them and separating the resulting iron hydroxide.

5. The method of claim 4, wherein the said washwater de-ironing includes adding ammonia thereto until a pH of or below 4 is achieved.

6. The method of claim 5, wherein the said washwater de-ironing further includes boiling down thereof.

2015328791 12 Dec 2018

7. The method according to each of claims 1 - 3,5,6, wherein de-ironing of the washwater or the washwater and the alum mother liquor by precipitation at the purification step, after the said separation of the slurry at the digestion step and water washing of the said solid residue, the alum mother liquor and washwater being collected separately, is done until the aluminum to iron weight concentration ratio of at least 10:1 is achieved in the pre-purified mother liquor.

8. The method of claim 7, wherein the purification step of the pre-purified mother liquor processing includes an operation for the mother liquor iron content reduction to ferrous iron, the reducing agent being one of ammonium sulfite, sulfur dioxide gas, or metallic aluminum powder.

9. The method according to each of claims 1 - 3, 5, 6, 8, wherein the alumina-bearing raw material is processed, at the digestion step and further at the purification step through to and inclusive of the iron reduction to ferrous iron, at a temperature of 75 to 180°C.

10. The method of claim 9, wherein, at the purification step, the prepurified mother liquor after its iron content reduction to ferrous iron is cooled at a temperature of or below 20°C.

11. The method of claim 10, wherein the ammonium alum crystals, after separating thereof from the pre-purified mother liquor and before dissolving them in pure water to obtain a de-ironed alum solution, are washed with a concentrated ammonium sulfate solution cooled to a temperature of or below 20°C.

12. The method of claim 11, wherein the pure water, used for dissolving the said crystals to obtain a de-ironed alum solution, is desalted water obtained at the solid ammonium sulfate production step.

13. The method according to each of claims 1-3,5,6,8,10 -12 wherein the pure water for rinsing off sulfuric acid retarded at the anion-exchange resin when recovering the acid from the pre-purified mother liquor, obtained at the purification step after ammonium alum crystals were separated therefrom, as well

2015328791 12 Dec 2018 as the pure water for dissolving the said crystals to obtain the de-ironed alum solution, is desalted water produced at the solid ammonium sulfate production step.

14. The method of claim 1, wherein, during the pre-purified mother liquor processing after the ammonium alum crystals have been separated therefrom and before dissolving the crystals in pure water, the repetitive cycles of the circular process further include washing the crystals with a concentrated ammonium sulfate solution cooled to a temperature of or below 20°C and obtained at the separation step, and the solid ammonium sulfate production step includes using the said ammonium sulfate solution, after it has been used for the alum crystal washing, for combining with the mother liquor, from which sulfuric acid has been recovered.

15. The method of claim 14, wherein the processed pre-purified mother liquor, from which ammonium alum crystals were separated, after being flown through the column containing a strong-base anion exchange resin in a sulfate form, and before combining it with the ammonium sulfate residual solution, obtained at the precipitated aluminum hydroxide separation step and previously used for the ammonium alum crystal washing, is further oxidized by aerial oxygen and filtered to separate iron hydroxide therefrom.

16. A method for alumina-bearing raw material digestion as part of processing thereof, comprising: preparing and heating a reagent solution containing ammonium hydrosulfate, decomposing the alumina-bearing raw material with the reagent solution to produce a slurry containing an ammonium alum solution together with solid decomposition residues, and separating the slurry into a solid and a liquid phases to obtain undecomposed solid residues and alum mother liquor, wherein the reagent solution preparation includes adding sulfuric acid thereto, and the slurry separation into the solid and the liquid phases includes water washing of the undecomposed solid residue and separately collecting the alum mother liquor

2015328791 12 Dec 2018 and washwater for further use as solutions for subsequent extraction of aluminum compounds therefrom.

17. The method of claim 16, wherein a reagent solution containing 5 to 65 weight percent of ammonium hydrosulfate is prepared and added with sulfuric acid until its concentration reaches 1 to 5%.

18. The method of claim 16 or claim 17, wherein the alumina-bearing raw material is decomposed with the hot reagent solution at a weight ratio of the reagent solution to the alumina-bearing raw material of 3 : 1 to 10 : 1.

19. The method of claim 18, wherein it is carried out at a temperature of 75 to 180°C.