IT202100032825A1 - Process for the treatment of residues from the alumina refining industry, obtaining valorisable products and secondary raw materials, in accordance with circular economy strategies - Google Patents

Description

PROCEDURE FOR THE TREATMENT OF RESIDUES

OF THE ALUMINUM REFINING INDUSTRY, WITH OBTAINMENT OF VALUEABLE PRODUCTS AND SECONDARY RAW MATERIALS, IN ACCORDANCE WITH THE STRATEGIES OF THE ECONOMY

CIRCULAR

The present invention concerns a new process for the treatment of residues, in particular solid residues, of the alumina refining industry, to obtain valorisable products, secondary raw materials and a drastic reduction or elimination of the need for to dispose of such waste in landfill, in accordance with circular economy strategies.

In particular, the present invention has as its object the treatment of the residue deriving from the bauxite treatment process for the refining of alumina (Bayer process), more commonly known as red mud (or red mud or residual bauxite), through the coupling of pyrometallurgical and hydrometallurgical processes.

As ? known, the most industrial process? used to obtain alumina from bauxite? called the Bayer process. The main phases that characterize the process are:

- bauxite grinding;

- digestion of bauxite in 10% NaOH aqueous solution. In this phase, the hot solubilization of the alumina present in the bauxite takes place. The temperatures, in this phase, can range from approximately 150?C Celsius (solubilization of trihydrated alumina) up to 250?C and high reaction pressure (for the solubilization of monohydrated alumina);

- separation of unsolved substances (red mud) by decantation and filtration;

- precipitation of Al(OH)3 by lowering the temperature;

- calcination of Al(OH)3, to obtain anhydrous alumina, Al2O3.

Given that the exact composition of the bauxites used for the production of alumina with the Bayer process can? differ depending on the mine of origin, all of them contain Al, Ti, Fe and Si oxides in various compositions and percentages; furthermore, other elements such as zinc, vanadium and some rare metals and rare earths are present in traces. The extraction efficiency of the Bayer process is very low and there? involves the presence of large quantities of metallic elements in the treatment waste, generically called red mud.

In the following Table 1? indicated is a typical composition of a red mud produced by the Bayer process.

Table 1 - main components of red mud

The following Table 2 shows the trace elements and rare earths present in a NALCO red mud.

Table 2 ? trace elements and rare earths in red mud.

In addition to the characteristics reported above, red muds, if not neutralized, have a highly basic pH (around pH 12.5).

The red muds are diluted, to make them more pumpable? easily and are sent to a pressure filter, where some components are recovered; subsequently, in the form of sewage, they are pumped away from the plant to be discharged into landfill basins, similar to artificial lagoons. This practice has an important environmental impact as this waste is not disposed of effectively and has not yet found an industrial application capable of absorbing the huge quantity. of material produced each year.

Red mud, therefore, is potentially extremely impactful waste, the management of which still constitutes a serious problem today. Although red mud is currently managed in such a way as to minimize impacts, even so, it represents an enormous danger to human health; Furthermore, some sites are still affected by the legacy. of past incorrect management. The surface deposits where the red muds are stored must be built and managed with particular attention to avoid contamination of the underlying aquifers and surrounding land and to prevent powdery material from being dispersed into the air causing harmful effects on health; in fact these dusts are highly alkaline in nature and cause irritating effects on the skin, eyes and respiratory system.

In recent years, numerous studies and experiments have been promoted to identify an adequate treatment of this waste. In some cases the high aluminum content in the material has recommended modifying the production cycle in order to reduce the alkaline charge, in order to obtain an inert mud that can be used to fill exhausted mines as a substrate for the replanting of original vegetation or for other agricultural purposes or as landfill material for coastal areas. In the end ? The use of red mud has also been tested in the production of building materials.

On the other hand, in the bauxite used in the Bayer process, in addition to the main elements indicated above, there are also compounds containing rare metals and rare earths, potentially exploitable but whose concentration is too low to promote an economically sustainable extraction process.

Depending on the type of bauxite used, the type and concentration of rare metals and rare earths present varies, however, among those with the highest high concentration, include for example titanium and scandium. As a result of the extraction of alumina during the Bayer process, these rare metals and rare earths also become concentrated in the red muds produced, with the result of reaching concentration values, although still very low, sufficient to encourage an extraction process. The typical composition data of a red mud, in fact, are those reported in the previous Table 1 and Table 2. In particular, even the mere valorisation of the scandium and titanium present could, hypothetically, economically support an extraction process, however their concentration in the red mud, so? like that of many other metals, rare metals and rare earths present is not? still the optimal one to make their extraction economically sustainable. ? It is therefore necessary to develop enrichment processes that create ?concentrates? of the elements of interest, in order to overcome this qualitative deficit.

In conclusion, although possible alternative ways of managing red mud have been studied, these are still considered a waste with a high environmental impact, the treatment and/or disposal of which represents an enormous cost for society. In addition, the exhaustion of available landfill volumes constitutes a problem for the continuation of activities. industrial: due to the so-called ?NIMBY? effect (Not In My Back Yard), not only in Europe and the USA, but also in many of the emerging economic powers, landfill disposal? seen as the last option, after having implemented the so-called 3 R principle (Reduce, Reuse, Recycle). As a consequence of this, obtaining a concession for new landfill volumes is possible. more and more difficult.

There is, therefore, in the specific sector the need to manage red mud with a more efficient procedure. advantageous in both economic and ecological terms.

This need? satisfied by the process according to the present invention which also offers further advantages which will become clear later.

The solution according to the present invention fits into this context, which aims to treat the red mud through a combination of pyrometallurgical and hydrometallurgical processes, aimed at the extraction of aluminium, titanium and scandium. This procedure? characterized by a substantial reduction in the quantity? of material sent to landfill, from the production of compounds with high added value and with good market prospects and from the possibility? to recycle some of the by-products formed during this process according to the canons of the circular economy.

In particular, the process for the treatment of red mud according to the present invention allows the production of high purity alumina (HPA), starting from waste with a high environmental impact, creating a flexible process, capable of differentiating the product based on needs. contingent markets. In particular, the process for the treatment of red mud according to the present invention allows the alternative production of aluminum ammonium sulphate (AAS), which is also highly valorisable, or both HPA and AAS products.

In accordance with the dictates of the circular economy, the various lines surrounding the main process are designed to produce further valorisable products (TiO2, scandium oxide and, possibly, further rare earth oxides); Furthermore, the choice of process conditions? designed to be able to reuse (through internal regeneration) the main reagents directly in the various phases of the process.

These and other results are obtained according to the present invention by proposing a process for the treatment of residues, in particular solid residues, of the alumina refining industry, or for the treatment of red mud, which includes the following phases: reductive melting of the red mud or pre-reduction and subsequent treatment, with the production of a slag almost completely free of iron and enriched in aluminium, titanium and scandium; what's the waste? obtained, it undergoes digestion with hot sulfuric acid with consequent conversion of the oxides into water-soluble sulphates; the solid obtained from digestion is leached with water with solubilization of the desired elements and their separation from the rest of the solid matrix; the liquid solution containing mainly aluminium, titanium and scandium sulphates undergoes various phases of selective recovery of these elements; the first phase involves the addition of ammonium sulphate directly to the leaching solution, with the formation of a precipitate of aluminum ammonium sulphate; the residual solution continues towards a second treatment phase, in which the titanium is separated by hot hydrolysis; once the precipitate has been separated, the residual solution undergoes the recovery of the scandium through liquid-liquid extraction in the organic phase.

The aluminum ammonium sulphate (AAS) solid, deriving from the first phase of aluminum recovery, undergoes a series of purification sub-phases (generally five sub-phases) by washing with hot water and filtration. The solid is what? purified undergoes a heat treatment which, depending on the temperature, can? lead to the formation of various products with high added value, such as high purity aluminum sulphate and alumina.

The process for the treatment of red mud according to the present invention has the advantage not only of reducing the quantity of material to be sent to landfill, thus increasing the life of the landfill itself, but also of obtaining products with high added value such as HPA (High Purity Alumina), titanium oxide, scandium oxide and rare earth elements (hereinafter also indicated with the acronym REE, for rare earth elements), to promote the recovery of raw materials that can be recirculated, reduce the environmental impact, limit the overall energy consumption and encourage an improvement in economic gain.

Purpose of the present invention? therefore that of providing a process for the treatment of residues, in particular solid residues, of the alumina refining industry, or for the treatment of red mud, which allows to overcome the limits of the processes according to the known art and to obtain the technical results previously described.

Further purpose of the invention? that said process for the treatment of red mud can be carried out with substantially low costs and that it is simple, safe and reliable.

The specific object of the present invention is therefore a process for the treatment of residues, in particular solid residues, from the alumina refining industry, or for the treatment of red mud, as specified in claim 1.

Further characteristics of the process for the treatment of red mud according to the present invention are specified in the subsequent dependent claims.

Will this invention come? now described, by way of illustration, but not by way of limitation, according to one of its preferred embodiments, with particular reference to the attached figure, in which? shown is a block diagram of the process for treating red mud according to the present invention.

Referring to figure 1, the process for the treatment of red mud involves an initial thermal treatment (not shown) of the red mud by means of reductive melting or pre-reduction and subsequent treatment, with the generation of an iron enrichment and a complementary termed slag , indicated with the numerical reference 1, enriched in aluminium, titanium and scandium.

The slag 1 resulting from the heat treatment undergoes digestion 2 with sulfuric acid; this phase is possible? carried out using a sulfuric acid with a concentration between 6M (approximately corresponding to 30% m/m sulfuric acid) and 18.5M (approximately corresponding to 98% m/m sulfuric acid), at a temperature of 50 ?C ? 250?C, for a period of 30 minutes? 2 hours, in a ratio between 0.5 and ? Twice the value of the stoichiometric ratio compared to the composition of the slag itself and necessary for the conversion of the oxides into sulphates. In practice, since? in this step we want to transform the oxides of aluminium, titanium, scandium and rare earths (REE) into the corresponding soluble sulfates, the quantity? in moles of sulfuric acid to be added, compared to the stoichiometric value, is calculated taking as reference the moles of these elements present in slag 1, but expressed as sulphates.

At the end of acid digestion, the mass is like this? obtained contains aluminium, titanium, scandium and REE sulphates which are recovered by leaching 3 with water. This phase can be conducted at a temperature within the range of 25°C? 75?C, for a period of time that can? go from 5 minutes to 30 minutes; the liquid-solid ratio L/S used for water leaching? included in the range between 5/1 and 20/1, calculated with respect to the initial mass of red mud subjected to acid digestion.

The washing solution? separated from the residual solid 4 by filtration. The aluminum present in the solution is precipitated 5 in the form of aluminum ammonium sulphate (AAS) by adding ammonium sulphate (AS); specifically, is a quantity added? of AS equal to 1-1.5 times the quantity? stoichiometric compared to aluminium. The reaction takes place at room temperature, and the reaction time is ? included in the 5 minute interval? 60 minutes.

Is the AAS precipitate like this? obtained is separated from the solution by filtration and continues with the washing and purification phases 6. Specifically, the AAS undergoes a first washing, necessary to eliminate the residual mother liquor from the precipitation, retained in the solid by imbibition. In this subphase, H2SO4 is used with a concentration between 0.01 M? 0.5 M, at a temperature between 25?C and 75?C for a time ranging from 5 minutes? 30 minutes; the L/S liquid-solid ratio used? included in the range from 1/1 to 5/1. The solid is what? obtained ? separated from the washing solution by filtration and proceeds with the purification sub-phases. These sub-phases are at least three in number, and all consist of treating the AAS solid with water at a temperature in the range of 25°C? 75?C, for a time ranging from 5 minutes? 30 minutes, with a liquid solid L/S ratio ranging from 1/1 to 5/1.

At the end of each purification cycle, the solid is separated from the solution by filtration and recovered. At the end of the purification phases, the AAS solid is what it looks like? obtained can? undergo different thermal phases, based on the product you want to obtain. Specifically, a calcination 7 carried out in a temperature range between 250?C and 600?C allows the AAS to be converted into aluminum sulphate, releasing ammonia, sulfur oxides and water (which can then be recovered as reusable ammonium sulphate in the initial phase of aluminum recovery). By bringing the temperature to a range between 900?C and 1200?C the conversion of aluminum sulphate into alumina (Al2O3) is obtained, releasing sulfur oxide (which can instead be recovered as sulfuric acid and used at the top of the process in phase of slag 1) acid digestion 2). Yes ? obtained in this way high purity alumina (HPA), indicated with the numerical reference 8.

As regards the solution following the precipitation of the aluminium, the procedure continues with the recovery phases of the other valorisable elements.

The solution is heat treated at a temperature between 90?C and 120?C for a time between 1h and 5h, using a reaction pH in the range 0.5? 2.5; under these conditions, hydrolysis 9 and the consequent precipitation of titanium in the form of hydrated titanium oxide is obtained. Once separated from the liquid, the solid hydrated titanium oxide undergoes a calcination step 10 in a temperature range between 900 ?C and 1000 ?C; under these conditions, hydrated titanium oxide is converted to anhydrous titanium oxide.

The post-hydrolysis solution of titanium undergoes a further scandium recovery phase, through liquid/liquid extraction 11 and the use of an appropriate mixture of solvents. In particular, the organic phase used? consisting of an extracting agent DEHPA (di-2ethylhexyl phosphoric acid) and a synergizing agent TBP (tributyl phosphate), dissolved in a suitable solvent (in the specific example kerosene, but it can also be hexane or another low-medium organic solvent boiling). The organic/aqueous O/W ratio? included in the interval 1:5 ? 1:20 and contact time? included in the 5 minute interval? 30 minutes. After this phase, if separation by static decantation is used, the time for complete separation of the organic phase from the aqueous phase is ? between 15 minutes and 60 minutes.

The separated acidic aqueous phase can, based on the composition of the red mud subjected to the process, go to subsequent phases 12 of separation of any rare earths present or it can? be relaunched at the head of the extraction phase with water.

The organic phase, enriched with scandium, is subjected to a back-stripping process, which consists in extracting the scandium from the organic phase, bringing it into a new aqueous phase.

Prior to back-stripping, can it? a further pre-stripping phase be conducted; does this have the purpose of selectively removing any impurities? both co-extracted and dragged together with scandium; furthermore, with this operation, the formation of the so-called crud is avoided (i.e. a stable emulsion of the solvents with the impurities present, which causes the unwanted formation of layers interposed between the organic phase and the aqueous phase), allowing at the same time the achievement of greater purity of the subsequently separated scandium.

Pre-stripping is performed by washing the organic phase with a solution of hydrochloric acid at a concentration in the range of 0.01 M? 0.1 M, with an organic/aqueous O/W ratio in the 1:1 range? 1:10. The contact time used in pre-stripping? between 5 minutes and 30 minutes.

The back-stripping phase occurs by treating the scandium-enriched organic phase with a basic NaOH solution with a concentration in the 2M range. 7M, with an organic/aqueous O/W ratio of 1:1? 1:10; the contact time taken? between 5 minutes and 30 minutes. Subsequently, if separation by static decantation is used, the time for complete separation of the organic phase from the aqueous phase is ? between 15 minutes and 60 minutes. Under the reaction conditions described, scandium passes into the basic aqueous phase and, at the same time, forms insoluble scandium hydroxide.

The organic phase, after basic stripping, is reactivated and ready to be reused in the extraction phase, while the basic solution is microfiltered to collect the scandium precipitated in the form of hydroxide, generating an enriched product called cake. The cake is periodically collected and sent to the scandium purification section, by calcination 13, to obtain scandium oxide.

Even the microfiltered basic phase can be recirculated for a new stripping cycle.

The invention will come further described below by way of illustration, but not limitation, with particular reference to an illustrative example.

Example.

The example illustrates an application of the process of the invention in practice.

After the separation of the iron from the red mud, through a reducing reaction at high temperature, a slag depleted in iron compounds and enriched in aluminum compounds and other elements present in the initial red mud is obtained.

In the specific example, the slag used is was obtained by carbon-reductive melting using a thermal plasma reactor and, after cooling, is was subjected to grinding in order to increase the homogeneity? of the material.

In the following Table 3? The elementary analysis of the slag used for the test is reported.

Table 3

A portion of 150g of slag, obtained from red mud subjected to a carboreductive reaction at high temperature, is was added with a solution composed of 37.5mL of water and 150mL of 18.5M H2SO4, in a porcelain capsule. Once the mixture has been homogenised, the sample is was put to digestion in a thermostated stove at 200?C for a period of 1 hour. Once the digestion time has elapsed, the solid is been recovered and? was added to 1.5 L of water, preheated to 50?C, thus obtaining an L/S ratio equal to 10 compared to the slag initially used. The system ? was kept under stirring with a contact time of 5 minutes. Subsequently, yes? then proceeded to separate the residual solid from the aqueous solution by vacuum filtration.

In the following Table 4? the composition of the aqueous solution obtained at the end of the leaching with water of the aforementioned slag is illustrated. Table 4

Compared to the initial slag, following leaching with water it is obtained an extraction of aluminium, titanium and scandium equal to 94%, 84% and 82.5% respectively.

The solution like this? obtained, it undergoes three treatment phases which allow the selective extraction, in series, of aluminium, titanium and scandium respectively.

In the first phase, the selective precipitation of aluminum occurs in the form of aluminum ammonium sulphate (AAS). The precipitation reaction occurs by adding 90g of ammonium sulphate to the leaching solution, at room temperature in a stirred reactor, with a reaction time of 30 minutes. The precipitate obtained? separated from the solution by vacuum filtration.

The filtered solution, after the separation of the aluminium, has the composition shown in Table 5.

Table 5

If you evaluate the grams of aluminum present in 1.5 liters of extracting solution before and after precipitation you can? have a measurement of the precipitated aluminum. Before the addition of ammonium sulphate, 24588 mg of Al were present in 1.5 liters of solution (obtained with the following calculation: 16392 mg/L Al x 1.5 L = 24588 mg of Al).

After precipitation and filtration, 2859mg of Al are present in the extracting solution (obtained with the following calculation: 1906mg/L Al x 1.5 L = 2859mg of Al). Therefore, the precipitation of aluminum with ammonium sulphate had an extraction yield of approximately 88%. During the precipitation of aluminum ammonium sulphate (AAS), this brings with it some "contaminants", due to the co-precipitation of minority elements and the presence of the mother liquid, due to imbibition. To eliminate these minority elements present in the solid as impurities? ? It is therefore necessary to carry out purification phases of the AAS precipitated solid. Initially, ? A pre-washing of the precipitated AAS was carried out, using 0.01M H2SO4 in a 1:1 mass ratio, in a stirred reactor, at room temperature and with a contact time of 30 minutes. The solid? was then separated from the washing solution by vacuum filtration. After pre-washing with sulfuric acid, solid AAS undergoes another four actual purification sub-phases; each of the washing sub-phases is carried out by contacting the solid with ultra-pure water in a 1:1 mass ratio, in a stirred reactor, with a contact time of 30 minutes, at a temperature of 50°C. Following each purification sub-phase, the liquid is separated from the purified solid by vacuum filtration.

The composition of the AAS produced, before and after all the purification sub-phases, is ? shown in Table 6.

Table 6

The solid is what? got it already? the characteristics of a marketable product. However, to ensure flexibility? and increase profitability? of the process, ? It is possible to transform AAS into high purity alumina. To obtain this result purified AAS obtained ? was calcined in an electric muffle furnace, with a ramp of 5?C/minute, up to 1200?C; the material ? was then kept at this temperature for 3 hours. During the thermal ramp, between 250?C and 600?C, AAS decomposes, releasing ammonia and sulphates, which, in the industrial process, are captured and treated in order to regenerate the ammonium sulphate, which will be used in the precipitation phase of the AAS in a new cycle. An aluminum sulfate intermediate is also formed.

Between 900?C and 1200?C the conversion of aluminum sulphate into alumina (Al2O3, aluminum oxide) is obtained, releasing sulfur oxide (which can be industrially recovered as sulfuric acid and used at the head of the process in the acid digestion phase of the slag).

The aluminum oxide obtained is was further washed with water, thus eliminating? the salts still present and obtaining alumina with purity > 99.99%.

The solution obtained after the precipitation of the aluminum is was subsequently treated, in a dedicated phase, to obtain the selective extraction of titanium. This solution is sent to a hydrolysis reactor where, at a temperature of 110?C and pH equal to 1, with a reaction time of 1 hour, the hydrolysis and precipitation of the titanium in the form of hydrated titanium oxide takes place. Once brought back to the temperature of approximately 25°C, the solution is separated from the titanium precipitate by vacuum filtration.

The post-precipitation solution of titanium has the composition shown in Table 7.

Table 7

The precipitation yield of titanium? of approximately 81%.

The solid is what? obtained ? was then calcined in an electric muffle furnace with a ramp of 5?C/minute up to 1000?C and with a hold of 1 hour, to obtain the anhydrous titanium oxide.

The leaching solution, once aluminum and titanium have been extracted, is subjected to solvent extraction for the selective extraction of scandium. For this purpose, the solution is? was treated with an organic kerosene solution with a concentration of DEHPA equal to 0.04M and TBP equal to 0.04M, in a stirred reactor, with an organic/aqueous O/W ratio equal to 1:10, with a contact time of 15 minutes. Once the contact time has elapsed, the mixture is was transferred into a separating funnel for a residence time of 30 minutes, to allow complete separation between the two phases. Table 8 shows the concentrations of the majority elements after liquid/liquid extraction with solvent.

Table 8

The scandium extraction yield? turned out to be 99%.

The scandium-enriched organic phase contains quantities minor amounts of impurities, both co-extracted and entrained with aqueous micro-droplets, and which in the back-stripping phase can cause the formation of the so-called crud. For this reason, before the actual back-stripping phase, the workforce is was treated with a 0.01 M HCl solution, with an organic/aqueous O/W ratio equal to 1:1 and for a contact time of 15 minutes. Once the contact time has elapsed, the mixture is was transferred into a separating funnel for a residence time of 30 minutes, to allow complete separation between the two phases.

Did the extracted scandium remain in the purified organic phase? was then subjected to back-stripping, with a 5M NaOH solution, with an organic/aqueous O/A ratio equal to 1:1 and for a contact time of 20 minutes, in a stirred reactor. Also in this case, once the contact time has elapsed, the mixture is was transferred into a separating funnel for a residence time of 30 minutes, to allow complete separation between the two phases.

The stripped organic phase can be reused for another extraction phase, while the scandium hydroxide precipitate present in the basic solution is was recovered by filtration and washed and subsequently calcined to obtain scandium oxide.

The leaching solution, after the scandium extraction, is not was subjected to further extraction of rare earths because, on the sample used for the test, the concentration of these was not sufficient to justify a selective extraction. This invention? has been described by way of illustration, but not by way of limitation, according to its preferred embodiments, but? it is to be understood that variations and/or modifications may be made by experts in the field without thereby departing from the relevant scope of protection, as defined by the attached claims.

Claims (12)

1) Process for the treatment of residues, in particular solid residues, from the alumina refining industry, or for the treatment of red mud, comprising the following phases: - digestion with sulfuric acid, with concentration 6 - 18.5M, temperature 50-250?C, duration 0.5-2h, with conversion of aluminum, titanium and scandium oxides into water-soluble sulfates and obtaining an aqueous solution of aluminium, titanium and scandium sulphates and a solid residue; - leaching with water of the solid residue obtained from digestion, with temperature in the range 25-75?C, duration 5-30min and liquid/solid ratio from 5/1 to 20/1, calculated with respect to the initial mass of red mud subjected to acid digestion, with further solubilization of aluminium, titanium and scandium in said solution and separation of a solid residue; - added to said solution after leaching of ammonium sulphate in quantity? equal to 1-1.5 times the quantity? stoichiometric with respect to aluminium, reaction time 5-60 minutes, with precipitation of aluminum ammonium sulphate and its separation from the residual solution; - purification of said aluminum ammonium sulphate, through a plurality? of washing cycles with hot water and filtration. 2) Process according to claim 1, characterized by the fact that, in said digestion phase, the quantity? of sulfuric acid? included in the range between 0.5 ? 2 times the value of the stoichiometric ratio compared to the concentration in moles of aluminium, titanium, scandium and rare earths (REEs) expressed as sulphates. 3) Process according to claim 1 or 2, characterized in that said phase of purification of aluminum ammonium sulphate comprises a first washing with an aqueous solution of H2SO4 0.01 ? 0.5M, temperature 25-75?C, time 5? 30 minutes, with a liquid-solid ratio from 1/1 to 5/1 and, after filtration to separate the solid as follows? obtained from the washing solution, at least three subsequent washing cycles with water at a temperature of 25-75?C, time 5? 30 minutes, liquid-solid ratio from 1/1 to 5/1 and filtration to separate the solid so? obtained from the washing solution. 4) Process according to any of the previous claims, characterized in that, following said purification step, said aluminum ammonium sulphate is ? subjected to heat treatment. 5) Process according to claim 4, characterized in that said heat treatment of said aluminum ammonium sulphate is conducted at a temperature of 250-600?C, to convert said aluminum ammonium sulphate into aluminum sulphate, releasing ammonia, sulfur oxides and water. 6) Process according to claim 4 or 5, characterized in that said heat treatment of said aluminum ammonium sulphate is? conducted at a temperature of 900-1200?C, to convert said aluminum ammonium sulphate into alumina, releasing sulfur oxide. 7) Process according to any of the previous claims, characterized by the fact that it includes the following preliminary phase: - reductive melting of red mud or pre-reduction and subsequent treatment, with the production of a slag enriched in aluminium, titanium and scandium. 8) Process according to any of the previous claims, characterized by the fact that it comprises the following phase subsequent to said phase of precipitation and separation of aluminum ammonium sulphate: - hydrolysis of the residual solution at a temperature of 90-120?C for a time between 1h and 5h, pH of 0.5? 2.5, with precipitation of titanium in the form of hydrated titanium oxide and its separation from the residual solution; - calcination of said hydrated titanium oxide at a temperature of 900-1000 ?C, to obtain anhydrous titanium oxide. 9) Process according to claim 8, characterized by the fact that it includes the following phase following said titanium precipitation and separation phase: - liquid-liquid extraction in organic phase of scandium from said residual solution. 10) Process according to claim 9, characterized by the fact that said liquid-liquid extraction takes place by adding a quantity of from 1/5 to 1/20 with respect to said residual solution, of a solution of an extracting agent, preferably di-2ethylhexyl phosphoric acid (DEHPA) and of a synergistic agent, preferably tributyl phosphate (TBP), dissolved in a low organic solvent - medium boiling, preferably kerosene or hexane, with contact time of 5-30 minutes, followed by static decantation for 15-60 minutes, to separate an organic phase enriched in scandium from an aqueous phase containing rare earths. 11) Process according to claim 10, characterized by the fact that said scandium-enriched organic phase is subjected to a washing phase with a quantity of from 1:1 to 1:10, compared to the quantity? of said organic phase, of a solution of NaOH 2 ? 7M, for 5 -g 30 minutes, for the extraction of scandium from the organic phase and its passage into the basic aqueous phase where it forms insoluble scandium hydroxide, which is separated by filtration and subsequently calcined to obtain scandium oxide. 12) Process according to claim 11, characterized in that said washing phase is preceded by a washing phase with a quantity? from 1:1 to 1:10 compared to the quantity? of said organic phase, of an aqueous solution of 0.01? hydrochloric acid? 0.1 M, for 5 - 30 minutes, to remove any impurities.