Process for the treatment of residues of the alumina refining industry, obtaining valorizable products and secondary raw materials, in accordance with the strategies of the circulatory economy

Abstract

The present invention relates to a process for treating residues, in particular solid residues, of the alumina refining industry, i.e. for treating red mud, comprising the following steps: - reductive fusion of the red mud or pre-reduction and subsequent treatment, with the production of a scoria enriched in aluminium, titanium and scandium and with a low content of iron; - digestion of said scoria with sulphuric acid, at a concentration of 6 - 18.5M and temperature of 50-250°C, for a duration of 0.5-2h, with the conversion of the aluminium, titanium and scandium oxides into water-soluble sulphates and the obtainment of an aqueous solution of aluminium, titanium and scandium sulphates and a solid residue; - leaching with water of the solid residue obtained from the digestion, with a temperature in the range of 25-75°C, for a duration of 5-30min and with a liquid / solid ratio of from 5 / 1 to 20 / 1, calculated relative to the initial mass of red mud subjected to acid digestion, with a further solubilisation of aluminium, titanium and scandium in said solution and separation of a solid residue; - addition to said solution, after leaching, of ammonium sulphate in an amount equal to 1 -1.5 times the stoichiometric amount relative to the aluminium, with a reaction time of 5-60 minutes, precipitation of aluminium ammonium sulphate and the separation thereof from the residual solution; - purification of said aluminium ammonium sulphate, by means of a plurality of washing cycles with hot water and filtration.

Description

[0001] PROCESS FOR THE TREATMENT OF RESIDUES OF THE ALUMINA REFINING INDUSTRY, OBTAINING VALORIZABLE PRODUCTS AND SECONDARY RAW MATERIALS, IN ACCORDANCE WITH THE STRATEGIES OF THE CIRCULATORY ECONOMY

[0002] The present invention relates to a novel process for treating residues, in particular solid residues, of the alumina refining industry, to obtain exploitable products, secondary raw materials and a drastic reduction in or elimination of the need to dispose of such residues in landfills, in accordance with the strategies of the circular economy.

[0003] In particular, the object of the present invention is the treatment of residue deriving from the process of treating bauxite for the refinement of alumina (Bayer process), more commonly known as red mud (or bauxite residue), by coupling pyrometallurgical and hydrometallurgical processes.

[0004] As is well known, the industrial process most widely used to obtain alumina from bauxite is called the Bayer process. The main steps characterising the process are:

[0005] - grinding of the bauxite;

[0006] - digestion of the bauxite in a 10% NaOH aqueous solution. The hot solubilisation of the alumina present in the bauxite takes place in this step. The temperatures in this step can go from about 150°C Celsius (solubilisation of alumina trihydrate) up to 250°C with a high reaction pressure (for the solubilisation of alumina monohydrate);

[0007] - separation of the undissolved portions (red mud) by decantation and filtration;

[0008] - precipitation of AI(OH)s through lowering of the temperature;

[0009] - calcination of AI(OH)s in order to obtain anhydrous alumina, AI2O3.

[0010] Whereas the exact composition of the bauxites used to produce alumina with the Bayer process can differ according to the mine from which they originate, Al, Ti, Fe and Si oxides are present in all of them in various compositions and percentages; moreover, other elements such as zinc, vanadium and some rare metals and rare earths are present in trace amounts. The extraction efficiency of the Bayer process is very low, and this results in the presence of large amounts of metal elements in the waste from the treatment, generically called red mud. A typical composition of red mud produced by the Bayer process is indicated in Table 1 below.

[0011] Table 1 - main components of red mud The trace elements and rare earths in a NALCO red mud are shown in Table 2 below.

[0012] Table 2 - trace elements and rare earths in red mud.

[0013] In addition to the above-described characteristics, red mud, if not neutralised, has a strongly basic pH (around pH 12.5).

[0014] Red mud is diluted, in order that it may be more easily pumped, and is sent to a pressure filter, where some components are recovered; then it is pumped, in the form of slurry, far from the plant in order to be discharged into waste ponds, similar to artificial lagoons. This practice has a major environmental impact, since this waste is not effectively disposed of and has not yet found an industrial application capable of absorbing the huge amount of material produced every year. Therefore, red mud is waste with a potentially extreme impact, whose management still today constitutes a serious problem. Though at present red mud is managed in such a way as to minimise the impacts, even so it nonetheless represents an enormous human health hazard; furthermore, some sites still suffer from the legacy of a past improper management. The surface deposits where red mud is stored must be constructed and managed with particular care to avoid contaminating the underlying aquifers and the surrounding terrain and to prevent powdery material from being dispersed into the air, thereby causing effects harmful to health; in fact, these powders are of a strongly alkaline nature and cause irritating effects on the skin, eyes, and respiratory tract.

[0015] In recent years, numerous studies and trials have been promoted in order to identify a suitable treatment for this waste. In some cases, the high content of aluminium in the material has made it advisable to modify the production cycle to reduce the alkaline load thereof in order to obtain an inert sludge that can be used to refill exhausted mines as a substrate for replanting the original vegetation or for other agricultural purposes or as a filler material for coastal areas. Finally, the use of red mud also in the production of building materials has been experimented with.

[0016] Moreover, in the bauxite fed into the Bayer process, in addition to the main elements indicated previously, there are also compounds containing rare metals and rare earths, which are potentially exploitable, but whose concentration is too low to promote an economically sustainable extraction process.

[0017] Depending on the type of bauxite used, the type and concentration of the rare metals and rare earths present vary; however, those at a higher concentration include, for example, titanium and scandium. Because of the extraction of alumina during the Bayer process, these rare metals and rare earths as well end up being concentrated in the red mud produced, with the result that they reach values of concentration which, though still very low, are sufficient to encourage an extraction process. The typical composition data of red mud, in fact, are as shown above in Table 1 and Table 2. In particular, even the exploitation of the scandium and titanium present could, hypothetically, on its own economically sustain an extraction process; however, their concentration in red mud, like that of many other metals, rare metals and rare earths present, is not yet optimal for rendering their extraction economically sustainable. It is thus necessary to develop enrichment processes that create “concentrates” of the elements of interest, in order to overcome this qualitative deficiency.

[0018] By way of example, in document CN111842411 one acts directly on the red mud, optionally prewashed to recover sodium hydroxide. Subsequently, a reaction with concentrated sulphuric acid and leaching take place, which bring not only aluminium and iron, but also silicon, into solution, since the reaction renders silica soluble. Therefore, it is subsequently necessary to make the liquid obtained by leaching react with sodium hydroxide, in a reactor under pressure at 180°C, with a reaction time of up to 8 hours, in order to precipitate the silicon, thereby producing an insoluble sodium silicate plus other silicate-based waste. This step is necessary to eliminate the silica impurity from the aluminium-based products. In addition, said document describes a selective precipitation of ammonium alum, by means of liquid ammonia or ammonium hydroxide under specific conditions that do not allow the formation of insoluble iron salts. Furthermore, in said document, since iron is present in large quantities and could be exploited, after the precipitation of the alum, a part of the iron is precipitated as red iron oxide. As the remaining iron would be a strong interferent for the recovery of other elements with ion-exchange resins, and the silicon remaining in solution would inactivate the surface of the resins, in said document the precipitation of an FeSiAl2, by means of a pH correction with ammonia and ammonium hydroxide, to a specific pH of 5.0 -5-5 and 120 min of reaction time is promoted.

[0019] In conclusion, though possible alternative ways of managing red mud have been studied, the latter is still considered waste with a high environmental impact, the treatment and / or disposal of which represents an enormous cost for society. In addition, the depletion of available landfill volumes constitutes a problem for the continuation of industrial activities: because of the so-called “NIMBY” (Not in My Back Yard) effect, not only in Europe and in the USA, but also in many of the emerging economic powers, disposal in landfills is seen as the last option, after the so-called principle of the 3 Rs (Reduce, Reuse, Recycle) has been implemented. As a consequence of this, obtaining a concession for new landfill volumes is increasingly difficult.

[0020] Therefore, in this specific sector there exists a need to manage red mud with a process that is more advantageous in both economic and ecological terms.

[0021] This need is satisfied by the process according to the present invention, which offers, moreover, further advantages that will appear clear hereinbelow.

[0022] The solution according to the present invention fits into this context; it proposes a treatment of red mud through a combination of pyrometallurgical and hydrometallurgical processes aimed at the extraction of aluminium, titanium, and scandium. The process is characterised by a substantial reduction in the amount of material sent to landfills, by the production of compounds with a high added value and good market prospects and by the possibility of recycling some of the byproducts formed during the process according to the criteria of the circular economy.

[0023] In particular, the process for treating red mud according to the present invention enables high purity alumina (HPA) to be produced from waste with a high environmental impact, providing a flexible process, capable of differentiating the product based on contingent market needs. In particular, the process for treating red mud according to the present invention allows for producing alternatively aluminium ammonium sulphate (AAS), likewise highly exploitable, or both products, HPA and AAS.

[0024] In accordance with the precepts of the circular economy, the different lines accompanying the main process are conceived so as to produce further exploitable products (TiO2, scandium oxide and, possibly, further rare earth oxides); furthermore, the choice of the process conditions is conceived to enable the main reactants to be reused (by internal regeneration) directly in the various steps of the process.

[0025] These and other results are obtained according to the present invention by proposing a process for treating residues, in particular solid residues, of the alumina refining industry, i.e. for treating red mud, which comprises the following steps: reductive fusion of the red mud or pre-reduction and subsequent treatment, with production of a scoria almost completely devoid of iron and enriched in aluminium, titanium and scandium; the scoria thus obtained undergoes digestion with hot sulphuric acid with a consequent conversion of the oxides into water-soluble sulphates; the solid obtained from the digestion undergoes leaching with water with solubilisation of the desired elements and the separation thereof from the rest of the solid matrix; the liquid solution containing mainly aluminium, titanium, and scandium sulphates goes through various steps for the selective recovery of the elements concerned; the first step provides for the addition of ammonium sulphate directly to the leaching solution, with the formation of a precipitate of aluminium ammonium sulphate; the residual solution continues towards a second treatment step, wherein there is a separation of the titanium by hot hydrolysis; once the precipitate has been separated, the residual solution is subjected to liquid-liquid extraction in the organic phase to recover the scandium.

[0026] The solid aluminium ammonium sulphate (AAS), deriving from the first aluminium recovery step, undergoes a series of sub-steps of purification (generally five sub-steps) by means of washing with hot water and filtration. The solid thus purified undergoes a thermal treatment which, according to the temperature, can lead to the formation of various products with high added value, such as aluminium sulphate and high purity alumina.

[0027] The process for treating red mud according to the present invention has the advantage not only of reducing the amount of material to be sent to a landfill, thus increasing the life of the landfill itself, but also of obtaining products with a high added value, namely HPA (high purity alumina), titanium oxide, scandium oxide and rare earths (hereinafter also indicated by the acronym REEs, for rare earth elements), promoting the recovery of raw materials that can be recirculated, reducing the environmental impact, limiting overall energy consumption, and favouring an improvement in economic gain.

[0028] The aim of the present invention is thus to provide a process for treating residues, in particular solid residues, of the alumina refining industry, i.e. for treating red mud, which makes it possible to overcome the limits of the processes according to the prior art and to obtain the previously described technical results.

[0029] A further aim of the invention is that said process for treating red mud may be carried out at substantially low costs and is simple, safe, and reliable.

[0030] Therefore, a specific object of the present invention is a process for treating residues, in particular solid residues, of the alumina refining industry, i.e. for treating red mud, as specified in claim 1 .

[0031] Additional features of the process for treating red mud according to the present invention are specified in the subsequent dependent claims.

[0032] The present invention will now be described, by way of non-limiting illustration, according to a preferred embodiment thereof, with particular reference to the appended figure, wherein a block diagram of the process for treating red mud according to the present invention is shown.

[0033] Making reference to figure 1 , the process for treating red mud comprises an initial thermal treatment (not shown) of the red mud by reductive fusion or prereduction and subsequent treatment, with the generation of a product enriched in iron and a complementary product defined as scoria, indicated by the reference number 1 , which is enriched in aluminium, titanium and scandium. In particular, said scoria 1 deriving from the thermal treatment is low in iron. Said thermal treatment is a fundamental and preparatory step for the subsequent steps. In fact, the scoria will be subjected to the subsequent steps, which are possible under the conditions described also due to the low iron content obtained after the initial pretreatment.

[0034] In particular, the scoria 1 deriving from the thermal treatment undergoes digestion 2 with sulphuric acid; this step can be carried out using a sulphuric acid with a concentration comprised between 6M (approximately corresponding to 30% m / m sulphuric acid) and 18.5M (approximately corresponding to 98% m / m sulphuric acid), at a temperature of 50°C - 250°C, for a time of 30 minutes - 2 hours, in a ratio comprised in the range of between 0.5 - 2 times the value of the stoichiometric ratio with respect to the composition of the scoria itself and necessary for the conversion of the oxides into sulphates. In practice, as in this step it is desired to transform the oxides of aluminium, titanium, scandium, and rare earths (REEs) into the corresponding soluble sulphates, the amount in moles of sulphuric acid to be added, relative to the stoichiometric value, is calculated by taking as reference the moles of these elements present in the scoria 1 but expressed as sulphates. In particular, said molar ratio of the sulphuric acid with the aluminium compounds present in the treated solid enables a low solubility of silicon to be obtained.

[0035] At the end of the acid digestion, the mass thus obtained contains sulphates of aluminium, titanium, scandium and REEs, which are recovered by leaching 3 with water. This step can be carried out at a temperature comprised in the range of 25°C - 75°C, for a period of time that can go from 5 minutes to 30 minutes; the liquidsolid ratio L / S used for leaching with water is comprised in the range of between 5 / 1 and 20 / 1 , calculated relative to the initial mass of red mud subjected to acid digestion. In particular, said solid / liquid ratio and said reaction times enable a low solubility of silicon to be obtained.

[0036] Therefore, in said process according to the present invention, no silica is brought into the solution, except for a marginal fraction. This characteristic makes the subsequent steps of separation and purification of the aluminium-based products far easier.

[0037] Furthermore, in the process according to the present invention the pH value is not modified, because the combination of the effects of the molar ratio in the digestion step and the solid / liquid ratio in the leaching step leads the system to operate in a manner that is not dependent on a specific pH value. In fact, said system functions independently of the pH value.

[0038] Moreover, the process according to the present invention does not have the need to use liquid ammonia or ammonium hydroxide to adjust the pH and promote the precipitation of ammonium alum, as occurs in document CN111842411.

[0039] Furthermore, in the prior art it is evident that silica is, together with iron, one of the principal contaminants when it is desired to produce alumina with a high degree of purity from red mud. In the process according to the present invention, conditions are created under which the silica is not solubilised, except marginally. Therefore, in the process according to the present invention, it is not necessary to eliminate silica impurities from the aluminium-based products.

[0040] The washing solution is separated from the solid residue 4 by filtration. The aluminium present in the solution is precipitated 5 in the form of aluminium ammonium sulphate (AAS) by adding ammonium sulphate (AS); specifically, the amount of AS added is equal to 1 -1 .5 times the stoichiometric amount relative to the aluminium. The reaction takes place at room temperature, and the reaction time is comprised in the range of 5 minutes - 60 minutes.

[0041] The AAS precipitate thus obtained is separated from the solution by filtration and continues with the washing and purification steps 6. Specifically, the AAS undergoes a first washing, necessary to eliminate the residue of mother liquors from the precipitation, retained in the solid due to imbibition. In this sub-step use is made of H2SO4 with a concentration comprised from 0.01 M - 0.5 M, at a temperature comprised between 25°C and 75°C for a time comprised in the range of 5 minutes - 30 minutes; the liquid-solid ratio L / S used is comprised in the range of 1 / 1 to 5 / 1. The solid thus obtained is separated from the washing solution by filtration and proceeds with the purification sub-steps. These sub-steps are at least three in number, and they all consist in treating the solid AAS with water at a temperature comprised in the range of 25°C - 75°C, for a time comprised in the range of 5 minutes - 30 minutes, with a liquid-solid ratio L / S comprised in the range of 1 / 1 to 5 / 1. In particular, the aim of said first washing is to remove only the soluble contaminants and not bring the aluminium ammonium sulphate back into solution.

[0042] At the end of every purification cycle, the solid is separated from the solution by filtration and recovered. At the end of the purification steps the solid AAS thus obtained can go through different thermal steps, based on the product it is desired to obtain. Specifically, a calcination 7 carried out in a temperature range comprised between 250°C and 600°C allows the AAS to be converted into aluminium sulphate, while freeing ammonia, sulphur oxides and water (which can then be recovered as ammonium sulphate, reusable in the initial aluminium recovery step). By bringing the temperature to a range comprised between 900°C and 1200°C one obtains the conversion of aluminium sulphate into alumina (AI2O3), while freeing sulphur oxide (which can instead be recovered as sulphuric acid and used at the start of the process in the step of acid digestion 2 of the scoria 1 ). High purity alumina (HPA), indicated by the reference number 8, is obtained in this way.

[0043] As regards the solution following precipitation of the aluminium, the process continues with the steps of recovering the other exploitable elements.

[0044] The solution is treated thermally at a temperature comprised between 90°C and 120°C for a time comprised between 1 h and 5h, using a reaction pH comprised in the range of 0.5 - 2.5; under these conditions, one obtains hydrolysis 9 and the consequent precipitation of titanium in the form of titanium oxide hydrate. Once separated from the liquid, the solid titanium oxide hydrate undergoes a calcination step 10 in a temperature range comprised between 900 °C and 1000 °C; under these conditions, the titanium oxide hydrate is converted into anhydrous titanium oxide.

[0045] Following hydrolysis of the titanium, the solution goes through a further step to recover scandium by liquid / liquid extraction 11 using a suitable mixture of solvents. In particular, the organic phase used consists of an extractant agent DEHPA (di(2-ethylhexyl)phosphoric acid) and a synergising agent TBP (tributyl phosphate), dissolved in a suitable solvent (kerosene in the specific example, but it can also be hexane or another low-medium boiling point organic solvent). The organic / aqueous ratio O / A is comprised in the range of 1 :5 - 1 :20 and the contact time is comprised in the range of 5 minutes - 30 minutes. After this step, if one uses separation by static decantation, the time of complete separation of the organic phase from the aqueous phase is comprised in the range of between 15 minutes and 60 minutes.

[0046] The separated acidic aqueous phase can, based on the composition of the red mud subjected to the process, go to subsequent steps 12 for the separation of any rare earths present, or can be reintroduced at the start of the step of extraction with water.

[0047] 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.

[0048] Prior to the back-stripping, a further pre-stripping step can be carried out; this has the aim of selectively removing any impurities, either co-extracted or entrained together with the scandium; furthermore, with this operation, one avoids the formation of so-called crud (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), while making it possible at the same time to reach a greater purity of the scandium subsequently separated.

[0049] The pre-stripping is performed by washing the organic phase with a solution of hydrochloric acid at a concentration comprised in the range of 0.01 M - 0.1 M, with an organic / aqueous ratio O / A comprised in the range of 1 :1 - 1 :10. The contact time used in pre-stripping is comprised between 5 minutes and 30 minutes.

[0050] The back-stripping step takes place by treating the organic phase enriched with scandium with a basic solution of NaOH at a concentration comprised in the range of 2M - 7M, with an organic / aqueous ratio O / A of 1 :1 - 1 :10; the contact time used is comprised in the range of 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 comprised in the range of between 15 minutes and 60 minutes. Under the reaction conditions described, the scandium passes into the basic aqueous phase and simultaneously forms insoluble scandium hydroxide.

[0051] After the basic stripping, the organic phase is reactivated and ready to be reused in the extraction step, whilst the basic solution undergoes microfiltration to collect the scandium precipitated in the form of hydroxide, thereby generating an enriched product called a cake. The cake is periodically collected and sent to the section for scandium purification by calcination 13 in order to obtain scandium oxide.

[0052] The basic phase subjected to microfiltration can also be recirculated for a new stripping cycle.

[0053] The invention will be further described below by way of non-limiting illustration, with particular reference to an illustrative example.

[0054] Example.

[0055] The example illustrates in practice an application of the process of the invention.

[0056] After the separation of the iron from the red mud, by means of a high temperature reducing reaction, one obtains scoria that has a low content of iron compounds and is enriched in compounds of aluminium and other elements present in the initial red mud.

[0057] In the specific example, the scoria used was obtained by carboreduction fusion with a thermal plasma reactor and, after cooling, it was subjected to grinding so as to increase the homogeneity of the material.

[0058] Table 3 below shows the elemental analysis of the scoria used for the test.

[0059] Table 3

[0060] A 150g aliquot of scoria, obtained from red mud subjected to a high temperature carboreduction reaction, had a solution composed of 37.5 mL of water and 150 mL of 18.5M H2SO4 added to it in a porcelain capsule. Once the mixture was homogenised, the sample underwent digestion in a stove set at a temperature of 200°C for a time of 1 hour. After the digestion time elapsed, the solid was recovered and added to 1 .5 L of water, preheated to 50°C, so as to obtain an L / S ratio of 10 relative to the scoria used initially. The system was kept under stirring with a contact time of 5 minutes. Subsequently, the solid residue was separated from the aqueous solution by vacuum filtration. Table 4 below illustrates the composition of the aqueous solution obtained at the end of leaching of the aforesaid scoria with water.

[0061] Table 4

[0062] Compared to the initial scoria, following leaching with water an extraction of aluminium, titanium, and scandium equal respectively to 94%, 84%, and 82.5% was obtained. The solution thus obtained undergoes three treatment steps which enable a selective extraction, in series, of aluminium, titanium and scandium, respectively.

[0063] The selective precipitation of aluminium in the form of aluminium ammonium sulphate (AAS) takes place in the first step. The precipitation reaction takes place with the addition of 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 is separated from the solution by vacuum filtration.

[0064] After the separation of the aluminium, the filtered solution has the composition shown in Table 5.

[0065] Table 5

[0066] If one calculates the grams of aluminium present in 1.5 litres of extractant solution before and after precipitation, one can have a measure of the precipitated aluminium. Before the addition of ammonium sulphate, 24588 mg of Al were present in 1 .5 litres of solution (obtained with the following calculation: 16392 mg / L Al x 1 .5 L = 24588 mg of Al).

[0067] After precipitation and filtration, 2859 mg of Al were present in the extractant solution (obtained with the following calculation: 1906 mg / L Al x 1 .5 L = 2859 mg of Al). Therefore, the precipitation of aluminium with ammonium sulphate had an extraction yield of about 88%. During the precipitation of aluminium ammonium sulphate (AAS), the latter carries with it some “contaminants” because of the coprecipitation of minority elements and the presence of the mother liquor due to imbibition. To eliminate these minority elements present in the solid as impurities, it is thus necessary to perform steps to purify the solid precipitate of AAS. Initially, a pre-washing of the precipitated AAS was carried out using 0.01 M 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. Following pre-washing with sulphuric acid, the solid AAS undergoes another four actual purification sub-steps; each of the washing sub-steps is carried out by placing the solid in contact 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. After every purification sub-step, the liquid is separated from the purified solid by vacuum filtration.

[0068] The composition of the AAS produced, before and after all the subpurification steps, is shown in Table 6.

[0069] Table 6

[0070] The solid thus obtained already has the characteristics of a marketable product. However, to assure flexibility and increase the profitability of the process, it is possible to transform AAS into high purity alumina. To obtain this result, the 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 maintained at this temperature for 3 hours. During the heat ramp, between 250°C and 600°C, AAS is decomposed and releases 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 AAS precipitation step in a new cycle. Furthermore, an aluminium sulphate intermediate is formed.

[0071] Between 900°C and 1200°C one obtains the conversion of aluminium sulphate into alumina (AI2O3, aluminium oxide), with the release of sulphur oxide (which can be industrially recovered as sulphuric acid and used at the start of the process in the scoria acid digestion step).

[0072] The aluminium oxide obtained was further washed with water, thus eliminating the salts still present and obtaining alumina with a purity > 99.99%.

[0073] The solution obtained after precipitation of the aluminium was subsequently treated, in a dedicated step, to obtain a selective extraction of titanium. This solution is sent to a hydrolysis reactor where, at a temperature of 110°C and pH of 1 , with a reaction time of 1 hour, hydrolysis and precipitation of the titanium in the form of titanium oxide hydrate take place. Once it is brought back to a temperature of about 25°C, the solution is separated from the titanium precipitate by vacuum filtration.

[0074] After precipitation of the titanium, the solution has the composition shown in Table 7.

[0075] Table 7

[0076] The titanium precipitation yield is about 81 %.

[0077] The solid thus obtained was then calcined in an electric muffle furnace with a ramp of 5°C / minute up to 1000°C and a hold of 1 hour in order to obtain anhydrous titanium oxide.

[0078] The leaching solution, once the aluminium and titanium had been extracted, was subjected to extraction with solvent for the selective extraction of scandium. For this purpose, the solution was treated with an organic kerosene solution with a concentration of DEHPA of 0.04M and of TBP of 0.04M, in a stirred reactor, with an organic / aqueous ratio O / A of 1 :10 and a contact time of 15 minutes. Once the contact time had elapsed, the mixture was transferred into a separatory funnel for a dwell time of 30 minutes to enable a complete separation between the two phases. Table 8 shows the concentrations of the majority elements after the liquid / liquid extraction with solvent.

[0079] Table 8

[0080] The scandium extraction yield was 99%.

[0081] The organic phase enriched with scandium contains minor amounts of impurities, either co-extracted or entrained with aqueous micro droplets, and which in the back-stripping step can cause the formation of so-called crud. For this reason, before the actual back-stripping step, the organic phase was treated with a 0.01 M HCI solution, with an organic / aqueous ratio O / A of 1 :1 and for a contact time of 15 minutes. Once the contact time had elapsed, the mixture was transferred into a separatory funnel for a dwell time of 30 minutes to enable a complete separation between the two phases.

[0082] The extracted scandium remaining in the purified organic phase was then subjected to back-stripping, with a 5M NaOH solution, with an organic / aqueous ratio O / A of 1 :1 and for a contact time of 20 minutes, in a stirred reactor. In this case as well, once the contact time had elapsed, the mixture was transferred into a separatory funnel for a dwell time of 30 minutes to enable a complete separation between the two phases.

[0083] The stripped organic phase can be reused for another extraction step, whilst the scandium hydroxide precipitate present in the basic solution was recovered by filtration and washed and subsequently calcined to obtain scandium oxide.

[0084] After the extraction of scandium, the leaching solution was not subjected to further extraction of rare earths since, in the sample used for the test, the concentration thereof was not sufficient to justify a selective extraction.

[0085] The present invention has been described by way of non-limiting illustration, according to the preferred embodiments thereof, but it is understood that variations and / or modifications may be introduced by persons skilled in the art without going beyond the relevant scope of protection, as defined by the appended claims.

Claims

CLAIMS1 ) A process for treating residues, in particular solid residues, of the alumina refining industry, i.e. for treating red mud, comprising the following steps:- reductive fusion of the red mud or pre-reduction and subsequent treatment, with the production of a scoria enriched in aluminium, titanium and scandium and with a low iron content;- digestion of said scoria with sulphuric acid, with a concentration of 6 - 18.5M, temperature of 50-250°C, duration of 0.5-2h, with a conversion of the aluminium, titanium and scandium oxides into water-soluble sulphates and the obtainment of an aqueous solution of aluminium, titanium and scandium sulphates and a solid residue;- leaching with water of the solid residue obtained from the digestion, with a temperature in the range of 25-75°C, duration of 5-30 min and a liquid / solid ratio of from 5 / 1 to 20 / 1 , calculated relative to the initial mass of red mud subjected to acid digestion, with further solubilisation of aluminium, titanium and scandium in said solution and separation of a solid residue;- addition to said solution, after leaching, of ammonium sulphate in an amount equal to 1 -1.5 times the stoichiometric amount relative to the aluminium, with a reaction time of 5-60 minutes and precipitation of aluminium ammonium sulphate and the separation thereof from the residual solution;- purification of said aluminium ammonium sulphate by means of a plurality of washing cycles with hot water and filtration.2) The process according to claim 1 , characterised in that, in said digestion step, the amount of sulphuric acid is comprised in the range of between 0.5 - 2 times the value of the stoichiometric ratio relative to the concentration in moles of aluminium, titanium, scandium and rare earths (REE) expressed as sulphates.3) The process according to claim 1 or 2, characterised in that said step of purification of aluminium ammonium sulphate comprises a first washing with an aqueous solution of 0.01 - 0.5M H2SO4, temperature of 25-75°C, for a time of 5 - 30 minutes, with a liquid-solid ratio of from 1 / 1 to 5 / 1 and, after filtration to separate the solid thus obtained from the washing solution, at least three successive washing cycles with water at a temperature of 25-75°C, for a time of 5 - 30 minutes, with a liquid-solid ratio of from 1 / 1 to 5 / 1 and filtration to separate the solid thus obtained from the washing solution.4) The process according to any one of the preceding claims, characterised in that, following said purification step, said aluminium ammonium sulphate is subjected to a thermal treatment.5) The process according to claim 4, characterised in that said thermal treatment of said aluminium ammonium sulphate is carried out at a temperature of 250-600°C, in order to convert said aluminium ammonium sulphate into aluminium sulphate, releasing ammonia, sulphur oxides and water.6) The process according to claim 4 or 5, characterised in that said thermal treatment of said aluminium ammonium sulphate is carried out at a temperature of 900-1200°C, in order to convert said aluminium ammonium sulphate into alumina, releasing sulphur oxide.7) The process according to any one of the preceding claims, characterised in that it comprises the following step after said step of precipitation and separation of aluminium ammonium sulphate:- hydrolysis of the residual solution at a temperature of 90-120°C for a time comprised between 1 h and 5h, with a pH of 0.5 - 2.5, and with precipitation of the titanium in the form of titanium oxide hydrate and the separation thereof from the residual solution;- calcination of said titanium oxide hydrate at a temperature of 900-1000 °C, in order to obtain anhydrous titanium oxide.8) The process according to claim 7, characterised in that it comprises the following step after said step of precipitation and separation of titanium:- liquid-liquid extraction in the organic phase of scandium from said residual solution.9) The process according to claim 8, characterised in that said liquid-liquid extraction takes place by adding an amount of from 1 / 5 to 1 / 20, relative to said residual solution, of a solution of an extractant agent, preferably di(2- ethylhexyl)phosphoric acid (DEHPA), and a synergising agent, preferably tributyl phosphate (TBP), dissolved in an organic low-medium boiling point solvent, preferably kerosene or hexane, with a contact time of 5-30 minutes, followed by static decantation for 15-60 minutes, to separate an organic phase enriched with scandium from an aqueous phase containing rare earths.10) The process according to claim 9, characterised in that said organic phase enriched with scandium is subjected to a washing step with an amount offrom 1 :1 to 1 :10, relative to the amount of said organic phase, of a 7M NaOH 2 solution, for 5-30 minutes, for the extraction of the scandium from the organic phase and the passage thereof into the basic aqueous phase where it forms insoluble scandium hydroxide, which is separated by filtration and subsequently calcined to obtain scandium oxide.11 ) The process according to claim 10, characterised in that said washing step is preceded by a step of washing with an amount of from 1 :1 to 1 :10, relative to the amount of said organic phase, of a 0.01 - 0.1 M aqueous solution of hydrochloric acid, for 5 - 30 minutes, to remove any impurities.

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