OF SILICATES"

Application Field

The present invention relates to the mining field. Specifically, the present invention relates to a process for the concentration of manganese from tailings of a beneficiation plant.

Background of the Invention

The manganese ore can be processed by crushing, classification and washing to remove the fine fraction, which is discarded as tailing. However, with the exhaustion of high grade manganese ore, mining industries will face the challenge of benefiting more complex ores and reprocessing tailings of high manganese content ores.

Usually manganese ores beneficiation flowcharts consist primarily of fragmentation and particle size classification, by exploiting only the richest and coarse fractions, which are products that are called "granulated" and "sinter feed". The finer particle size fraction (below 0.150 mm) is currently discarded as tailing for not being noble and also due to the fact that the current equipment / beneficiation operations are not suitable for the recovery of these finer particle size fraction.

In this context it becomes relevant the development of alternative flowcharts and concentration routes for these wastes as a complement to current processes in order to increase the global recovery of manganese as well as to reduce the environmental impact of disposal of this finer particle size fraction.

According to the present invention, a new route of concentration of tailings from the Azul Mine is presented, through reverse flotation in pH> 10, with cationic collector and a polysaccharide, like Amide, as a depressor, with 20% solids, using stages of rougher, scavenger and cleaner flotation, whose mineral-ore consists of manganese oxides (cryptomelane-holandite) and the gangue mineral consists essentially of kaolinite.

The development of an appropriate technology for the concentration of fine manganese will enable the processing of millions of tons of tailings that have been discharged by processing plants, as well as prevent the continuity of such practice in the future. In addition to enhanced production, the recovery of fine manganese would also allow the reduction of the environmental impact of mining activity since it minimizes the disposal of waste.

When the industrial concentration circuit is fed by the lithologies

"Tabular Pelite" (PETB), Pelite Siltite (PEST), Detritic (DETR), Rich Pelite (PERC) or Metallurgical Bioxide (BXME) fine fractions (tailings) are produced, which are also called PETB, (PEST), (DETR), (PERC) and (BXME), respectively. Therefore, PETB, (PEST), (DETR), (PERC) and (BXME) should be understood herein with the aim of identifying the fine fractions which constitutes the tailings of the current processing circuits and which is also derived from the lithology of the same name.

Developing a process to recover (and concentrate) manganese from the samples/lithologies called PETB, (PEST), (DETR), (PERC) and (BXME) constitute the objective of the present invention. The invention is designed to concentrate manganese- carrying minerals existing in the materials called PETB, (PEST), (DETR), (PERC) and (BXME).

Summary of the Invention

The present invention relates to a process for concentrating manganese from the tailing of a beneficiation plant characterized by comprising the stages of removing coarse particle size fraction from the tailing, desliming and conducting an acid or a basic reverse cationic flotation. The manganese-carrying minerals of the present invention are usually minerals with low manganese content being preferred derived from the lithologies "Tabular Pelite" (or PETB), Pelite Siltite (or PEST), Detritic (or DETR), Rich Pelite (or PERC) and Metallurgical Bioxide (or BXME).

The present invention also relates to a reverse cationic flotation used to concentrate manganese which is floated using depressors agents and collectors agents as flotation reagents.

Brief Description of the Drawings

Figure 1 demonstrates a generic flowchart of PETB processing Figure 2 represents the configuration of a reverse cationic flotation circuit in basic medium.

Figure 3 represents the scheme adopted in the flotation assays with PETB.

Figure 4 represents the configuration of a reverse cationic flotation circuit in acid medium.

Figure 5 (panels A, B and C) demonstrates the global metallurgical balance of the reverse cationic flotation in basic medium.

Figure 6 demonstrates the metallurgical balance of the reverse cationic flotation in acid medium.

Figure 7 demonstrates the global metallurgical balance of the concentration process based on desliming followed by reverse cationic flotation in basic medium.

Figure 8 demonstrates the configuration of reverse cationic flotation circuit in basic medium.

Detailed Description of the Invention

The present invention relates to a process for concentrating manganese from the tailing of a beneficiation plant.

The fine fractions which constitutes the tailings of the current processing circuits and which are also derived from the lithologies "Tabular Petite" (PETB), Pelite Siltite (PEST), Detritic (DETR), Rich Pelite (PERC) or Metallurgical Bioxide (BXME) are known for their low manganese content. According to the state of the art for the ores containing manganese the direct anionic flotation process is preferred for the recovery (and concentration) of manganese. Nevertheless, until now no results in terms of adequate manganese liberation can be identified encouraging the persistence in the use or development of this concentration route. In fact, the direct anionic flotation in basic medium of the manganese minerals was not successful. In this way, the need for a better manganese recovery (and concentration) process still remains in the current state of the art.

Surprisingly, the present invention is designed to concentrate manganese minerals existing in the materials called PETB, (PEST), (DETR), (PERC) and (BXME) using a different route, a concentration process by flotation, but using reverse cationic flotation of gangue in basic or acid media. Through this process, instead of floating the manganese- containing ores, kaolinite, the main contaminant mineral is floated, being the manganese concentrated and recovered at the sunken products of the flotation process.

The manganese minerals of the present invention are usually minerals with low manganese content.

The process of the present invention is generally characterized by comprising the stages of:

a) Removal of the coarse particle size fraction (>210μιτι) of the tailing;

b) Desliming the finer fraction obtained in stage a) at ΙΟμηη, generating a fraction of slurries (overflow) and an underflow;

c) Joining the fraction removed in stage a) with the deslimed fraction greater than ΙΟμιη obtained in stage b;

d) Conducting acid or basic flotation of the product from stage c).

To be submitted to the concentration process by flotation, the tailing of the current processing circuit, which is derived from the typologies "Tabular Pelite" (PETB), "Pelite Siltite" (PEST), "Detritic" (DETR), "Rich Pelite" (PERC) or "Metallurgical Bioxide" (BXME), requires the general following procedures:

-> Removal of the coarse granulometry fraction (>210μιη), so that it does not cause blockages in the cyclones that will carry out the desliming at ΙΟμπι. The removed material, being very rich in Mn, should be incorporated to the deslimed product to compose the flotation feed;

-s> Desliming in cyclone at ΙΟμιη, generating a fraction of slurries (overflow) and an underflow which should compose the flotation feed.

In the reverse cationic flotation process of the present invention if a basic flotation is carried out the initial flotation feed is composed of 20% of solids. If an acid flotation is carried out the initial flotation feed is composed of 50% of solids. The procedures described above are carried out in batches, as illustrated in Figure 1. Adequate modifiers are used in order to improve the reverse cationic flotation selectivity. In the reverse cationic flotation process of the present invention depressors agents and collectors agents are used as flotation reagents. The depressor agent is usually a polysaccharide, preferably corn starch, and the cationic collector agent is usually an amine, preferably selected from the group consisting of amine ether and amide-amine.

The flotation process may be accomplished either in acid or basic media and one or more flotation stages (which may also be called cleaner stages) may be included in the flotation circuit configuration in order to achieve the desired manganese content in the concentrates.

Owing to the fact that the particles of kaolinite (main mineral of gangue) present a greater degree of liberation than the particles of the manganese minerals, the reverse cationic flotation of the gangue is more recommendable than direct flotation of the ore minerals. Indeed, direct anionic flotation in basic medium of the manganese minerals was not achieved successfully. In this way the purpose of the present invention is a process to recover (and concentrate) manganese from the tailing which is based on desliming followed by reverse cationic flotation.

In order to concentrate the tailing from the Azul Mine, for example, (typologies PETB, (PEST), (DETR), (PERC) and (BXME)) it is necessary to submit the materials to a single operation of desliming at ΙΟμιη, followed by flotation. The overflow constitutes the slurries and is discarded as tailing. The underflow should feed the flotation.

The reverse cationic flotation of gangue in basic medium of the present invention should be carried out with 20% solids, at 10<pH<10.3. Flotation reagents should be used for conditioning just like depressors and collectors. Figure 2 and Figure 3 represent the possible arrangements of reverse cationic flotation circuits in basic medium.

Examples of depressors, but not limiting the invention, are polysaccharides. Amide or the commercial product Fox Head G2241 (also not limiting the invention) will act as depressors of manganese minerals in the approximate concentration ranges of 200-500 mg/L or 900-2000 g/t.

Examples of collectors, but not limiting the invention, are amines. Amine ether (like the commercial product Lilaflot 811M) or amide-amine (like the commercial product Flotigam 5530) (also not limiting the invention) will act as collectors for kaolinite, or silicates in general, in the approximate concentration ranges of 1000-1500 mg/L or 3900-5900 g/t.

Depressors and collectors should be added in this order, being that the conditioning with depressors has to be conducted for at least 2.5 minutes and the conditioning with collectors has to be conducted for at least 1 minute.

After conditioning with the flotation reagents described, the rougher flotation should be carried out for 4-5 minutes. The foam produced (rougher tailing) should be mixed with water and submitted to a scavenger stage for 2-7 minutes, without adding reagents. The foam generated by the scavenger is considered to be tailing, whereas the sunken product should be mixed to the rougher sunken matter and together are considered to be concentrate, according to Figure 3.

Nevertheless, at this stage it is possible to realize that the manganese content obtained in the concentrate is still below the expectation, indicating the need of introducing in the process a cleaner stage into the flotation circuit configuration. In this case the foam generated by the first scavenger (scavenger-1) is considered to be tailing (Tailing-1), whereas the sunken product should be mixed to the rougher sunken matter and together should feed a 2nd stage composed of a Cleaner flotation stage, followed by a Scavenger-2 stage (according to Figure 2). The sunken products in the Rougher and Scavenger-1 stages should present a concentration of solids of 14-17%.

In this case, the pulp should be conditioned with depressor in the approximate concentration range of 90-120 mg/L or 500-650 g/t and with collector agent in the approximate concentration range of 350-500 mg/L or 2000-2650 g/t at 10<pH<10.3.

The cleaner flotation should be conducted for 2-4 minutes, producing a foam which will feed the Scavenger-2 stage. This should be carried out for 3-6 minutes, without adding reagents. According to Figure 2, the product floated in the Scavenger-2 stage constitutes Tailing-2, whereas the products which sank in the Cleaner and Scavenger-2 stages are mixed and considered to be the final concentrate.

Alternatively, the reverse cationic flotation in acid medium of the present invention should be conducted in accordance with the scheme illustrated in Figure 4. The best results are obtained by preparing the pulp with 50% solids, adding H ₂ SiF ₆ in an amount to correct the pH up to pH=3 and conditioning for at least 3 minutes. After that, NaP0 ₃ (1430 mg/L or 2000 g/t) is added as dispersant, followed by conditioning for at least 2 minutes. After conditioning, the pulp is diluted to approximately 30% solids, at the dosage of 3000 g/t (or 1360mg/L) of the collector agent is added and the conditioning is allowed for at least 1 minute. The rougher flotation is conducted for at least 6-7 minutes. The foam produced in the rougher stage fed a scavenger stage which is conducted for at least 10-11 minutes, in the absence of reagents.

Following the scheme illustrated in Figure 4, the sunken from the

Scavenger stage receive H ₂ SiF ₆ to correct the pH=3, conditioning it for at least 5 minutes. After this conditioning a collector agent is added and conditioned is allowed for at least 1 minute. The cleaner flotation is conducted at pH=3.2 for at least 5 minutes. The foam produced by the cleaner stage is considered to be tailing, whereas the sunken product is mixed with the rougher sunken compose the final concentrate.

For the processes of the present invention it is important to emphasize that the PETB, PEST, DETR and BXME ores are predominantly composed of kaolinite which, as well as other clayey minerals, have a notable capacity to alter the rheological properties of the flotation pulp, adversely affecting the mixture of the reagents and influencing the flotation kinetics. Said fact is less important for the BXME mineral, but much more relevant for other typologies of the Azul Mine (DETR, PEST and PETB). To solve the problem, the suggestion is to work with more diluted pulps, that is, with a percentage of solids lower than 25%. It is important to emphasize that the scavenger stage is necessary with the aim of eliminating the hydrodynamic drag of the fine particles of manganese minerals for the foam produced.

CHARACTERIZING THE SAMPLES

According to the present invention, before beginning the experiments designed to concentrate the manganese minerals present in the compounds PETB, PEST, DETR, PERC and BXME, a sample was submitted to characterization studies which were carried out at the Technological Characterization Laboratory (LCT) of the Mine Engineering and Petroleum Department at EPUSP. The most significant information for processing (mineralogy and degree of liberation) is presented in summary form ahead. The information refers to the compound PETB, but is analog for other typologies.

Chemical and mineralogical composition of the minerals - PETB

The particle size distribution of the material is as displayed in Table 1, where it is possible to note the major occurrence of material with very fine particles, since 45.5% of its mass present particle size lower than 0.010mm (ΙΟμιτι), whereas only 3.1% presents a size greater than 0.60mm.

Table 1 Particle Size distribution of the material.

Particle size Mass Retained (%) Content (%) fraction (mm) Simple Accumulated Mn Si0 ₂

+0.589 3.1 3.1 32.9 13.0

-0.589 +4.147 8.2 11.3 20.8 23.1

-0.147 +0.074 6.8 18.1 14.2 27.5

-0.074 +0.037 7.7 25.8 11.4 29.2

-0.037 +0.020 7.5 33.3 8.5 30.0

-0.020 +0.010 21.2 54.5 4.7 35.5

-0.010 45.5 100.0 2.0 39.6

Total (calculated) 100.0 - 7.1 34.2 The following is noted from Table 1:

-> The fraction retained in the sieve of 28# (opening of 0.589mm) is highly rich in manganese (32.9%). In fact, the results to be presented ahead inform that the typical concentrates from the flotation present content in this same range.

The content of Si0 ₂ rises with the decrease of the size of the particles, indicating that the finer fractions are the richest silicacarrying minerals.

In accordance with that informed in Table 2, the PETB sample is mostly composed of silica (34.2%) and alumina (29.7%), accompanied by high content of volatiles (12.5% loss in fire). The content of Mn, however, is only 7.1%, accompanied by 7.3% of Fe and 1.1% of Ti0 ₂ .

Table 2 Chemical composition of the PETB sample.

eralogical composition (Table 3 ahead) corroborates the chemical composition, since the sample in question is mostly made of kaolinite (71% in mass), accompanied by cryptomelane-hollandite (17%), goethite (3.7%) and bixbyite (3.1%).

From the information in Tables 1 and 3 and the characterization report, the following can be noted:

-> Cryptomelano-hollandite is the predominant manganese-carrying mineral (17% in mass) in the lithology PETB, with prominence also to the presence of manganese in bixbyite (3% in mass) and in lithiophorite (1% in mass), and the initial content of Mn of this lithology can be considered low if compared with other richer lithologies such as Rich Pelite (PERC - content of Mn: 23.3%) or Metallurgical Bioxide (BXME - content of Mn: 24.4%);

The content of manganese decreases considerably in the fraction of fines, with proportions situated between 11 and 33% above 0.037mm and in the range of 2.0 to 8.5% below 0.037mm;

The content of Si0 ₂ and Al ₂ 0 ₃ (Kaolinite, main mineral of gangue), presents a different behavior to content of manganese (cryptomelane), maintaining high concentration in all the particle size ranges which were analyzed, with a slight increase in the fine fraction below 0.010mm.

Table 3 Mineralogical composition of the fraction (-0.60 +0.010mm).

State of liberation of the cryptomelane-hollandite particles

Knowledge of the state of liberation of cryptomelane-hollandite particles by particle size range helps to choose the mesh of grinding to be adopted in developing the concentration process, and to predict the greater or lesser difficulty in obtaining concentrates with a content of Mn compatible with the market specification. According to the characterization studies of the present invention, the information related to the liberation of the particles of cryptomelane-hollandite which compose the PETB sample is demonstrated in Table 4.

Table 4 Liberation of the particles of cryptomelane-hollandite from the lithology PETB by particle size range (-0.60mm +0.010mm).

Liberation by Particle Size Range (%)

Total Liberation (%)

-0.60mm -0.15mm -0.074mm -0.037mm -0.020mm (*)

+0.15mm +0.074mm +0.037mm +0.02000 +0.010mm

45 21 31 39 59 82 (*) -0.60mm +0.010mm

Concerning the liberation of the particles of the main manganese mineral (Table 4), it is important to emphasize that:

- The degree of total liberation (GL) of the PETB sample is very low (GL=45 ). Therefore, it is unrealistic to expect to obtain flotation concentrates with a very high content of Mn;

-> In the course granulometry fractions (+0.037mm), GL assumes values below 40%, rising to GL=59% in the range of -0.037mm +0.020mm;

-> The degree of liberation only reaches higher values (GL=82%) in the finer particle size fraction (-0.020mm +0.010mm). However, according to common knowledge from the state or the art the flotation of fine particles is not very efficient.

Since the liberation of the main mineral (cryptomelane-hollandite) is deficient, it seems reasonable to make efforts in the reverse flotation of the main mineral of gangue (kaolinite). To carry out reverse flotation of kaolinite, it is necessary to know the degree of liberation (GL) of its particles, in accordance with the results shown in Table 5.

Table 5 Liberation of the kaolinite particles from the lithology PETB by particle size range (-0.60mm +0.010mm).

The results of Table 5 indicate that:

The degree of total liberation of the kaolinite particles is GL=88%. Said value is much higher than that of the main manganese mineral (GL=45%). Therefore, the reverse flotation of kaolinite demonstrates greater success than the direct flotation of the manganese oxides; -> In the coarse granulometry fraction (-0.60mm +0.020mm), the degree of liberation is in the range of: 68%_ GL _ 90%;

-> In the finer granulometry fraction (-0.020mm + 0.010mm), the degree of liberation reaches the amount of GL=95%, but said particle size range is already very near the limits of the flotation process, according to common knowledge from the state or the art.

PREPARING THE SAMPLES

The tailing from the typology "Tabular Pelite" (PETB) is dried in a stove at 40°C to withdraw the natural humidity. Once dried, the entire mass is homogenized and subsequently submitted to the preparation flowchart illustrated in Figure 1. The same procedure is carried out for the compounds PEST, DETR, PERC and BXME.

In accordance with the flowchart of Figure 1, the entire mass of PETB is classified in a sieve of 65# (opening of 0.21mm). This procedure is necessary to avoid blockage of the hydrocyclone on desliming.

Sieving the PETB generates two products:

A passing material (undersize) which is submitted to a single operation of desliming in a hydrocyclone (cycloning), seeking a cut at ΙΟμιη;

b) A material withheld in the sieve, which is subsequently mixed to the deslimed product to feed the flotation.

Following the preparation flowchart illustrated in Figure 1, a single cycloning operation is applied to the undersize of the sieve of 65# (opening of 0.21mm), generating two products:

a) An underflow, where the coarse particles are concentrated;

b) An overflow, where the slimes are concentrated.

Representative samples of the overflow and underflow from the hydrocyclone, still in pulp form, have their particle size analyzed, for example by laser beam diffraction technique. A summary of the results is displayed in Table 6, where it can be noted that 36% (in volume) of the deslimed product (underflow) corresponds to particles having a size less than ΙΟμιτι. Table 6 Particle Size distribution (in volume) of the overflow and underflow from desliming.

Regarding the particle size distribution of the slimes (Table 12), it is noted that 96.8% of its volume displays a size lower than ΙΟμιτι. Continuing the preparation flowchart described in Figure 1, the material collected in the underflow of the hydrocyclone (deslimed product) is coarse, dried at 40 ^e C, and finally homogenized in an elongated pile jointly with the oversize of the sieve of 65# (opening of 0.21mm), resulting the composition in a product that has been named "Flotation feed". From this pile of homogenization aliquots of 500 grams are withdrawn which are used in flotation assays. Particle size and chemical composition from the flotation feed - PETB

Particle size and chemical composition of the product called "Flotation feed" are presented in Table 7, where it is noted that 73% of its mass displays a size less than 0.020mm. It is important to emphasize that the flotation process loses efficiency when applied to particles fine. On the other hand, 10% of the mass that feeds the flotation presents a size greater than 0.21mm. The flotation process is also refractory to the recovery of coarse particles, according to common knowledge from the state or the art. It can be further noted in Table 7 that the manganese is concentrated in the coarse particle size fractions (withheld in 65#), whereby it is possible to calculate an average content of 34.0% of Mn. As the material gets finer, the manganese becomes impoverished and the contents of Si0 ₂ and AI ₂ C»3 become enriched, indicating that the content of kaolinite increases in the finer fractions. The distribution of the contents (Mn, Fe, P, Si0 ₂ , Al ₂ 0 ₃ , Ti0 ₂ , CaO, MgO,

K ₂ 0, BaO and PF) by particle size range of the product named "Flotation feed" found in Table 7, Panels A and B.

Table 7 Panel A: Particle size distribution of the "Flotation feed".

Table 7 Panel B: distribution and chemical composition of the "Flotation feed".

Concerning the distribution of the manganese in the "Flotation feed" (Table 7, Panel B), it is noted that:

a) 35% of the manganese is concentrated in the finest fraction, that is, that which passes through the sieve of 635# (opening of 0.020mm);

b) 30% of the manganese is concentrated in the coarse fraction, that is, that which is withheld in the mesh of 65# (opening of 0.21mm);

c) 35% of the manganese is distributed among the intermediate particle size fractions, that is, between 0.21mm and 0.020mm.

Regarding the distribution of silica and alumina in the "Flotation feed" (Table 7, Panel B), it is noted that 85% of the silica, and also of the alumina, is concentrated in the finest fraction (size less than 0.020mm), and the remaining 15% is distributed along the other particle size classes. Said behavior constitutes an indication of the distribution of the main mineral of gangue, kaolinite.

The density of the material named "Flotation feed" was determined in triplicate by pycnometry, resulting in a value of (2.51±0.01) g/cm ³ . Said low density is evidence of the predominance of the mineral kaolinite in the composition of this material.

The following examples are presented in order to improve the clarity of the present invention scope without limiting it.

Example 1. Flotation for "Tabular Pelite" (PETB) in basic medium

In order to concentrate the tailing from the Azul Mine (typology PETB) the material is formerly submitted to an operation of desliming at ΙΟμιτι, followed by flotation. The overflow constitutes the slurries and is discarded as tailing. The underflow feeds the flotation.

The reverse cationic flotation of gangue in basic medium is carried out with 20% solids, at 10<pH<10.3, after conditioning with flotation reagents: depressor (corn starch) and cationic collector, added in this order, after 2.5 minutes of conditioning with depressor and 1 minute of conditioning with cationic collector. Amide or Fox Head G2241 act as depressors of manganese minerals in the concentration of 227 mg/L or 900 g/t, whereas amine ether (Lilaflot 811M) or amide- amine (Flotigam 5530) act as collectors for kaolinite in the concentration of 1360 mg/L or 5333 g/t. After conditioning with the flotation reagents, the rougher flotation is carried out for 5-6 minutes. The foam produced (rougher tailing) is mixed with water and submitted to a scavenger-1 stage for 6 minutes, without adding reagents.

The foam generated by the scavenger-1 is considered to be tailing (Tailing-

1), whereas the sunken product is mixed to the rougher sunken matter and together feed a 2 ^nd stage composed of a cleaner flotation stage, followed by a scavenger-2 stage.

The products sunken in the rougher and scavenger-1 stages present a concentration of solids of 14-17%. Said pulp is than conditioned with a depressor agent (amide or Fox Head) in the concentration of ~90 mg/L or ~500 g/t and with a cationic collector (Flotigam 5530 or Lilaflot 811M) in the concentration of ~364 mg/L or ~2030 g/t at 10<pH<10.3. The cleaner flotation is conducted for 6 minutes, producing a foam which feeds the Scavenger-2 stage. This is carried out for 4 minutes, without adding reagents. According to Figure 2, the product floated in the Scavenger-2 stage constitutes Tailing-2, whereas the products which sank in the Cleaner and Scavenger-2 stages are mixed and considered to be the final concentrate.

The global metallurgical balance of the concentration process based on desliming followed by reverse cationic flotation in basic medium is summarized in Table 8 and illustrated in Figure 5 (panels A, B and C).

Table 8 Metallurgical balance of the process comprised of desliming + flotation in basic medium

Products Contents (%) Mass partition (%)

Mn Si0 ₂ Mass Mn Si0 ₂

Slurries 11.2 30.9 46.0 12.8 52.5

Flotation tailing 6.3 35.0 43.8 40.1 43.7

Concentrate 32.1 13.0 10.2 47.1 3.8

Feed (calculated) 6.9 35.1 100.0 100.0 100.0 Example 2. Flotation for "Tabular Pelite" (PETB) in acid medium

The reverse cationic flotation in acid medium is conducted in accordance with the scheme illustrated in Figure 4. The best results are obtained by preparing the pulp with 50% solids, adding H ₂ SiF ₆ to correct the pH=3 (930 mg/L or 1116 g/t), conditioning for 3 minutes, after which, NaP0 ₃ (1430 mg/L or 2000 g/t) is added as dispersant, followed by conditioning for 2 minutes. After conditioning, the pulp is diluted to 31% solids, at the dosage of 3000 g/t (or 1360mg/L) is added of the collector Flotigam 5530 which is conditioned for 1 minute. The Rougher flotation is conducted for 6-7 minutes. The foam produced in the Rougher stage fed a Scavenger stage which is conducted for 10-11 minutes, in the absence of reagents.

Following the scheme illustrated in Figure 4, the sunken from the Scavenger stage receive H ₂ SiF ₆ (255 mg/L) to correct the pH=3, conditioning it for 5 minutes. After this conditioning, Flotigam 5530 (455 mg/L) is added and conditioned for 1 minute. The cleaner flotation is conducted at pH=3.2 for 5 minutes. The foam produced by the Cleaner stage is considered to be tailing, whereas the sunken product is mixed with the rougher sunken to compose the final concentrate. The metallurgical balance of the concentration process comprised of desliming and reverse cationic flotation in acid medium is presented in Table 9 and illustrated in Figure 6.

Table 9 Metallurgical balance of the process comprised of desliming + flotation in acid medium

Products Contents (%) Mass partition (%)

Mn Si0 ₂ Mass Mn Si0 ₂

Slurries 11.2 30.9 46.0 12.8 51.6

Flotation tailing 7.0 35.6 45.7 46.5 45.7

Concentrate 33.7 11.7 8.3 40.7 2.7

Feed (calculated) 6.9 35.7 100.0 100.0 100.0 Example 3: Flotation for "Pelite Siltite" (PEST) in basic medium

To be submitted to the concentration process by flotation, the tailing of processing circuits derived from the typology "Pelite Siltite" (PEST) also requires the procedures of removal of the coarse granulometry fraction and desliming, according to the general procedure. The same concentration route adopted for typologies as PETB is followed, being carried out reverse cationic flotation of the silicates in basic medium (10.0<pH<10.3).

The reverse cationic flotation of the gangue in basic medium is carried out with 20% solids, at 10<pH<10.3, after conditioning with flotation reagents: depressor (corn starch) and cationic collector, which are added in this order, after 2.5 minutes of conditioning with depressor and 1 minute of conditioning with collector. Amide or Fox Head G2241 act as depressors of manganese minerals in the concentration of 230 mg/L or 900 g/t, whereas amide-amine (Flotigam 5530) act as collector for kaolinite in the concentration of 1360 mg/L or 5333 g/t. After conditioning with the flotation reagents described, the rougher flotation is carried out for 3.5 minutes. The foam produced (rougher tailing) is mixed with water and submitted to a scavenger stage for 7-8 minutes, without adding reagents. The foam generated by the scavenger is considered to be tailing, whereas the sunken product is mixed to the rougher sunken matter and together are considered to be concentrate.

The flowchart of the concentration process is illustrated in Figure 3. It is comprised by reverse cationic flotation in basic medium. Its metallurgical balance is summarized in Table 10, where it is noted that it is possible to obtain a concentrate containing 39% Mn and overall metallurgical recovery of 50%. The flotation tailing constitutes the main loss of Mn (34%) which can be justified by the deficient liberation of the Mn minerals. In the slurries, only 17% is lost.

Table 10 Metallurgical balance of the process comprised of desliming + flotation in basic medium Products Contents (%) Mass partition (%)

Mn Si0 ₂ Mass Mn Si0 ₂

Slurries 6.5 35.7 42.2 16.7 56.5

Flotation tailing 15.0 27.8 36.9 33.7 38.4

Concentrate 39.0 6.5 20.9 49.6 5.1

Feed (calculated) 16.4 26.7 100.0 100.0 100.0

Figure 7 shows the global metallurgical balance of the reverse cationic flotation for PEST in basic medium.

Example 4. Flotation for "Detritic" (DETR) in basic medium.

To be submitted to the concentration process by flotation, the tailing of processing circuits derived from the typology "Detritic" (DETR) also requires the procedures of removal of the coarse granulometry fraction and desliming, according to the former procedures. The same concentration route adopted for as typologies as PETB and PEST is followed, being carried out reverse cationic flotation of the silicates in basic medium (10.0<pH<10.3).

To concentrate the tailing from washing the Azul Mine (typology DETR) it is necessary to submit the material to an operation of desliming at ΙΟμιη, followed by flotation. The overflow constitutes the slurries and is discarded as tailing. The underflow feed the flotation.

Following the same strategy adopted for concentrating the other lithologies from the Azul mine, the reverse cationic flotation of the gangue (silicates) in basic medium is carried out with 20% solids, at 10<pH<10.3, after conditioning with flotation reagents: depressor (corn starch) and cationic collector, which are added in this order, after 2.5 minutes of conditioning with depressor and 1 minute of conditioning with collector. Corn starch (Fox Head G2241) act as depressor of manganese minerals, in the concentration of 300 mg/L (or 1183 g/t), whereas amide- amine (Flotigam 5530) act as collector for silicates, in the concentration of 1500 mg/L (or 5900 g/t). After conditioning with the flotation reagents described, the rougher flotation is carried out for 5.0 minutes. The foam produced (rougher tailing) is mixed with water and submitted to a scavenger stage for 5.5 minutes, without adding reagents. The foam generated by the scavenger is considered to be tailing (tailing-1, whereas the sunken product is mixed to the rougher sunken matter and together feed a 2 ^nd stage composed of a cleaner flotation stage, followed by a scavenger-2 stage (as illustrated in Figure 8).

The products sunken in the rougher and scavenger-1 stages present a concentration of solids of ~16 . Said pulp is conditioned with depressor (Fox Head G2241) in the concentration of ~120 mg/L or ~619 g/t and with collector (Flotigam 5530) in the concentration of ~500 mg/L or ~2609 g/t at 10<pH<10.3. The cleaner flotation is conducted for 3.5 minutes, producing a foam which feed the scavenger-2 stage. This is carried out 2.8 minutes, without adding reagents. According to Figure 8, the product floated in the scavenger-2 stage constitutes the tailing-2, whereas the products which sank in the cleaner and scavenger-2 stages are mixed and considered to be the final concentrate; the global metallurgical balance for processing the DETR typology is presented in Table 11 where it can be noted that:

-> By conducting the flotation in accordance with Example 4 it is possible to generate a concentrate with 22.3% of Mn and 21.2% of Si0 ₂ ;

-> The overall recovery of Mn from the process is 52.0%, and 14.9% is lost in desliming and 33.1% in tailing from the flotation process.

Table 11 Metallurgical balance of the process comprised of desliming + flotation in basic medium for typology DETR

Products Contents (%) Mass partition (%)

Mn Si0 ₂ Mass Mn Si0 ₂

Slurries 1.4 39.4 45.9 14.9 49.5

Flotation tailing 3.3 37.2 43.9 33.1 44.6

Concentrate 22.3 21.2 10.2 52.0 5.9

Feed (calculated) 4.4 36.6 100.0 100.0 100.0 Example 5: Flotation for "Rich Pelite" (PERC) in basic medium.

To be submitted to the concentration process by flotation, the tailing of processing circuits derived from the typology "Rich Pelite" (PERC) also requires the procedures of removal of the coarse granulometry fraction and desliming, according to the former procedures. The same concentration route adopted for typologies as PETB and PEST is followed, being carried out reverse cationic flotation of the silicates in basic medium (10.0<pH<10.3).

To concentrate the tailing from washing of the Azul Mine (typology PERC) it is necessary to submit the material to a single operation of desliming at ΙΟμηι, followed by flotation. The overflow constitutes the slurries and is discarded as tailing. The underflow should feed the flotation.

Following the same concentration route adopted for other typologies of the Azul Mine (PETB and PEST), the reverse cationic flotation of the gangue in basic medium is carried out with 20% solids, at 10<pH<10.3, after conditioning with flotation reagents: depressor (polysaccharides) and collector (fatty amines), which are added in this order, after 2.5 minutes of conditioning with depressor and 1 minute of conditioning with collector.

Following the same reasoning adopted for former processes, corn starch (Fox Head G2241) act as depressors of manganese minerals in the concentration of 300 mg/L (or 1183 g/t), whereas amide-amine (Flotigam 5530) act as collector for silicates in the concentration of 1200 mg/L (or 4717 g/t). After conditioning with the flotation reagents described, the rougher flotation is carried out for 3.4 minutes. The foam produced (tailing rougher) is mixed with water and submitted to a scavenger stage for 3.2 minutes, without adding reagents. The foam generated by the scavenger is considered to be tailing, whereas the sunken product is mixed to the rougher sunken matter and together are considered to be concentrate (Figure 3).

The overall metallurgical balance for processing the typology PERC is presented in Table 12 where it can be noted that: -> By conducting the flotation in accordance with the experiment conditions of Example 5, it is possible to generate a concentrate with 48.21% of Mn and 8.65% of Si02;

-> The overall recovery of Mn from the process is 63.9%, and 14.0% is lost in desliming and 22.1% in tailing from the flotation.

Table 12 Metallurgical balance of the process comprised of desliming + flotation in basic medium for typology PERC

Example 6: Flotation for "Metallurgical Bioxide" (BXME)

To be submitted to the concentration process by flotation, the tailing of processing circuits derived from the typology "Metallurgical Bioxide" (BXME) also requires the procedures of removal of the coarse granulometry fraction and desliming, according to the former procedures. The same concentration route adopted for as typologies as PETB and PEST is followed, being carried out reverse cationic flotation of the silicates in basic medium (10.0<pH<10.3). These procedures are carried out in batches, in accordance with that illustrated in the flowchart of Figure 1.

The reverse cationic flotation of the gangue in basic medium is carried out with 20% solids, at 10<pH<10.3, after conditioning with flotation reagents: depressor (polysaccharides) and collector (fatty amines), which are added in this order, after 2.5 minutes of conditioning with depressor and 1 minute of conditioning with collector. Following the same reasoning adopted for PETB and PEST, polysaccharides (Fox Head G2241) act as depressors of manganese minerals in the concentration of 500 mg/L (or 1967 g/t), whereas amide-amine (Flotigam 5530) act as collector for silicates in the concentration of 1000 mg/L (or 3933g/t).

After conditioning with the flotation reagents described, the rougher flotation is carried out for 6.0 minutes. The foam produced (rougher tailing) is mixed with water and submitted to a scavenger stage for 4.8 minutes, without adding reagents. The foam generated by the scavenger is considered to be tailing, whereas the sunken product is mixed to the rougher sunken matter and together are considered to be concentrate (as illustrated in Figure 3).

The global metallurgical balance for processing the typology BXME is presented in Table 13 where it can be noted that:

-> By conducting the flotation in accordance with the experiment conditions of example 6, it is possible to generate a concentrate with 47.99% of Mn and 5.03% of Si0 ₂ ;

-> The overall possible recovery from the process is 46.7%, and 15.80% is lost on desliming and 37.5% in tailing from the flotation.

Table 13 Metallurgical balance of the process comprised of desliming + flotation in basic medium for typology BXME

Once six examples of the preferred aspects of the present invention were presented above it is noteworthy to mention that the scope of protection conferred by the present document encompasses all other alternative forms appropriate for the implementation of the invention, which is defined and limited only by the content of the attached set of claims.