Method for separation of elemental sulphur and sulphide concentrate from intermediate products of sulphide concentrates hydrometallurgy processing

FIELD: chemistry.

SUBSTANCE: method for separation of the elemental sulphur and sulphide concentrate from the intermediate products of the sulphide concentrates hydrometallurgy processing includes sulphur-sulphide flotation of the intermediate original pulp in order to separate the sulphur and sulphides from intermediate oxide components with forming of the sulphur-sulphide concentrate. Then the autoclave disintegration of the sulphur-sulphide concentrate pulp is carried out in the presence of reagent - sulphide hydrophilisator and at the temperature higher than melting point of the elemental sulphur. After that the sulphur flotation is carried out in order to separate the disintegrated pulp into sulphur and sulphide concentrates. Before the sulphur-sulphide and sulphur flotations the pulp undergoes the additive agitation with the reagents, the disintegration is carried out at the temperature 140-150°C.

EFFECT: effective separation of the sulphur and sulphide concentrates.

5 cl, 3 dwg, 1 tbl

 

The invention relates to methods of processing residues autoclave leaching of sulfide materials of non-ferrous metallurgy and can be used to highlight formed on the leaching of elemental sulfur from the oxidized slurry with obtaining sulfur and sulfide concentrates.

A known way of separating elemental sulfur from the hydrate slurry autoclave oxidative leaching of pyrrhotite concentrate [Amineurin and other Method of obtaining sulfur and sulfide concentrates. Copyright certificate 1303549, MKI SW 17/027]. The method includes the following operations: collective flotation sulphur and sulphides in serosurvey concentrate, disintegration and sulphide flotation, the products of which are sulfur and sulfide concentrates. Sulphur concentrate is fed to autoclave smelting of sulfur and sulfide - on pyrometallurgical processing in conjunction with Nickel ore concentrate. The method involves the use of reagent-hydrophilization operations disintegration relatively cheap and safe lime (CaO) and Supplement organic surface-active agents.

The closest in technical essence is adopted as the prototype of a way of separating elemental sulfur and sulfide concentrate from the hydrate slurry autoclave oxidative videlacele the Oia pyrrhotite concentrate [Vinpearl, Angarov, Zierenberg, Washability. The development of technology autoclave oxidative leaching of pyrrhotite concentrates. Non-ferrous metals, 1983, No. 12, p.1-4].

The method is based on autoclave-flotation processing methods oxidized slurry autoclave leaching of pyrrhotite concentrate and includes the following operations: collective flotation sulphur and sulphides in serosurvey concentrate (SPC), the disintegration of the SCQ and sulphide flotation, the products of which are sulfur and sulfide concentrates. Sulphur concentrate is fed to autoclave smelting of sulfur and sulfide - on pyrometallurgical processing in conjunction with Nickel ore concentrate. Serosurvey flotation is carried out in acidic medium at pH~4. The tailings slurry is neutralized and pumped to the tailings pond. Disintegration serosorting concentrate is carried out at 115-130°C, pH 9÷10 filing of an alkaline reagent-hydrophilization Na2S. Final separation of elemental sulfur and sulfides occurs in sulfuric flotation disintegrated pulp, which is also carried out in alkaline medium at pH 9÷10.

The method provides a high level of sulphur in the sulphur concentrate (>70%) while the lowest content in sulphide residue (~3-5%).

The disadvantage of these methods is that they are tested for sulfide cheese is I, not having in its composition of pyrite FeS2. The main phases of such materials are hydrated form of iron (~60÷70%), the sulphides of non-ferrous metals (Ni, Cu, Co), pyrite and elemental sulfur. If the raw material contains pyrite (as, for example, in the Ural zinc concentrates), the separation of elemental sulfur and sulfides described method does not occur. In addition, it is associated with the use of expensive and toxic reagent-hydrophilization (sodium sulfide) on the operation of disintegration in large quantities (~35 kg/t raw material for cereulide flotation).

The difficulty of separation of pyrite and elemental sulfur due to the following facts. First, pyrite as well as elemental sulfur, has a natural high hydrophobicity and is easy to floatated sulfide minerals. Secondly, the hydrophobicity of pyrite increases with the processing of sulfuric acid [Shimaraev. Selective flotation. Moscow, 1958, 726 S.], as well as coverage of molten sulfur in the high wettability of it. This processing occurs at high temperature sulfuric acid leaching of zinc concentrate.

The objective of the invention is the separation of elemental sulfur from Chekov autoclave leaching of zinc production, the main phases are: elemental sulfur (~25-35%), pyrite (~15-25% and jarosite lead (up to 20%). The content of noble metals (BM) in this cake does not allow to consider it otvajnym. Return rich elemental sulfur product (cake autoclave leaching) in pyrometallurgical production does not meet the environmental requirements to reduce emissions of sulfur dioxide into the atmosphere, and also complicates the work of the dust collecting equipment and leads to its premature failure.

The technical result of the invention is the separation of sulfur concentrate suitable for sulfuric melting (that is, containing not less than 70% S°), and getting sulfide concentrate, concentrating in himself up to 80% of noble metals original cake leaching, with a minimum content of elemental sulfur (1÷10%). This material can be directed to the pyrometallurgical redistribution order to extract valuable components (copper, BM), without violating environmental requirements.

The problem is solved in that the selection elemetary sulfur from the oxidized slurry leaching of zinc concentrate at the expense of additional processing of the material with reagents before each flotation operation, as well as by increasing the temperature of disintegration.

According to research under the microscope Axioplan Zeiss principal amount (~60%) of elemental sulfur in the material is free the nom condition. The rest of it is associated with pyrite in the form of conglomerates of various sizes, from single grains of chalcopyrite. Grains of pyrite, destroyed in the process of leaching, covered with a border of elemental sulfur, in addition, elemental sulfur fills it cracks due to the high wettability of the molten sulfide sulfur. Only a small portion of pyrite is in a free state in the form of grains of various sizes. These features of the mineralogical structure of the source material in combination with a high content of pyrite determine the necessity of additional processing before flotation separation, as well as raise the temperature of disintegration.

It follows from the above that the main difficulty is the choice of modes of operations, ensuring the separation of elemental sulfur and pyrite.

From the practice of beneficiation of sulfide ores is known that the cheapest and most widely used depressant pyrite is lime. The time of agitation flotation pulp with lime before selective flotation can vary from one to several hours [Shimaraev. Selective flotation. Moscow, 1958, 726 S.].

For flotation of sulfur usually use hydrocarbon collectors: kerosene and other petroleum products. Despite the fact that when lime depression surface Piri is partially covered with plaster, the use of hydrocarbon collectors increases photoactivity this sulfide. Therefore, the use of specified types of collectors for elemental sulfur in this case ineffective.

The problem is solved by using a different reagent - depressant and hydrophilization sulfide minerals - sodium sulfide is used in the enrichment grey Chekov hydrometallurgy [Mantsevich M.I. development of theory and practice of flotation of Nickel-pyrrhotite raw materials in the combined and traditional schemes of its processing. The dissertation on competition of a scientific degree of the doctor of technical Sciences. M, 1995].

Consumption of sodium sulfide, corresponding to the pH of the slurry is greater 9.5 and its concentration in solution ~0,8÷1,5 g/l, provides the value of the redox potential of the pulp Eh=-560÷-360 mV relative to silver chloride electrode. Under these conditions, elemental sulfur saves flotation activity and enters the foam product without the use of hydrocarbon collectors, and sulphide and oxide components remain in the chamber product flotation. In contrast to the method-analogue, where disintegration is sodium sulfide (temperature 115-130°C), in the proposed method, this reagent is used only on cereulide flotation carried out at room temperature, and most importantly, in much smaller quantities: ~7 kg/t is of similar material, coming to serosurvey flotation.

Serosurvey concentrate is subject to additional separation of the sulfur and sulfide components. This material is subjected to alkaline treatment, called disintegration, at a temperature above the melting point of elemental sulfur. Adding an alkaline reagent in the heated slurry is Hydrophilidae sulphides, which creates conditions for further separating them from the elemental sulfur flotation.

High temperature and alkaline environment lead to some dissolution of sulfur reactions:

The most likely form of polysulfide sulfur is pentasulfide - CaS5.

Along with polysulfide (Sn2-) and thiosulfate (S2O32-) forms part of the sulphur is dissolved with the formation of monosulfide (S2-or HS-ions:

The presence of liquid phase sulfide ions promotes the wettability of the solid solution particles, represented by the remnants of sulphides. Elemental sulfur remains hydrophobic. All this creates conditions for further separation of sulphur and sulphides flotation.

However, for sulfuric flotation on the proposed method requires a pre-processing pulp reagents-depressors pyrite, as before cereulide flotation.

Thus, the technological scheme of the proposed method includes an additional operation: the agitation of the pulp with reagents before each flotation. The scheme is presented in figure 1.

The above is confirmed by the following examples.

Experiments on the implementation of the method of the prototype and the proposed method were performed on samples of the residue leaching of zinc concentrate containing, %: Zn 1,5-4,0; Cu of 1.4-2.3; Pb 2,4-3,6; Fe 17-21; S° 27-30, noble metals in the quantity g/t Au And 2.7-To 5.5; Ag 180-210.

The first flotation concentration (serosorting flotation) was conducted on a laboratory mechanical machines of the brand 237-FL removable cells in a volume of 1 l; 0.75 l and 0.5 l to maintain a certain density in flotation operations, as well as on the machines of the brand "Mekhanobr" with the volume of the chambers 3 litres.

Processing (agitation) pulp source material (solids content 50%) with reagents before flotation was carried out in a beaker with a stirrer, a speed which was 700 rpm Size used lime was 99% class - 0,074 mm Loading lime in a glass stirrer was carried out in a dry form, and sodium sulphide in the form of a 10%solution. The flow rate was determined by maintaining the pH at a given level.

The solid content in cereulide flotation was established within the 30-20%.

After conducting flotation experience of the products obtained was filtered, the solid was dried in a drying Cabinet to constant weight at a temperature of 80-85°C, the weight was recorded to calculate the outputs of each product. The prepared sample products were subjected to chemical analysis to elemental sulfur.

Scheme cereulide flotation presented in figure 2.

Studies of the process of disintegration was carried out in the laboratory of titanium autoclave American firm Parr volume of 7.5 liters when the ratio of the liquid and solid phases 2:1. As the alkaline reagent, hydrophilizing sulfides, used lime: activity 81% and a particle size of ~0,1÷1,0 mm Source material for disintegration served concentrate cereulide flotation with the content of elemental sulfur 42÷47%.

The experiments were conducted as follows. Open the autoclave was loaded a portion of the source material (serosurvey concentrate after flotation), the estimated quantity of lime (raspolojennom) and the required amount of the liquid phase. Then the autoclave was sealed, compressed and heated to the desired temperature (125÷145°C) with mixer, producing a reset abhasa (0 ATM) at a temperature of 90-95°C and turned off the mixer.

The time of reaching the set temperature was considered the beginning of the experience. At the end of the specified time autoclavation, the slurry was discharged into a sealed container (taking the sample solution) and sent for further processing sulphur flotation. The solution was analyzed for content monosulfide sulfur S2-and total sulphur.

Experiments on sulfur flotation was carried out successfully disintegrated pulp containing sulfide ion in the liquid phase. Scheme sulfuric flotation presented in figure 3.

Both agitation was carried out in two variants: with two reagents (Cao, Na2S) 40 minutes before the main flotation, 25 minutes before cleaning out; with one reagent (Cao) for 15 minutes each. The consumption of reagents was determined by the need to maintain the pH at a given level.

Method of preparation of products for chemical analysis was similar to that described above for the products cereulide flotation.

The results of experiments on the implementation of the prototype and the proposed method are presented in the table.

The table shows that cereulide flotation neotricula acid residue leaching without pre-treatment reagent material separation does not occur (op. CC1): not observed even forming a foam product. Does not improve the situation and washing the cake to a pH of 6.2: without pre-treatment with flotation reagents is not op. CC2).

Processing the washed cake alkaline reagents allows to increase the pH of the pulp. However, as the show is and experiments to achieve satisfactory performance flotation pH value floatated slurry should be at least 10.5. At pH< 10,0 flotation separation cereulide and oxide components does not occur (op. CC3). When the pH of 10.0 to 10.5 material is divided into two products, but removing sulfur serosurvey concentrate does not reach the required values (≥90%), and its content in the flotation tailings exceeds the permissible value is in the range of 9.0÷9,5% (op. CC4, 5).

At pH>10,5 received satisfactory performance of the flotation process (op. CC6-9): the extraction of sulfur in the concentrate ≥95% when the content of her in it >40%, and in the tails <5%.

Raising the pH to 13.0 seems inappropriate because it does not lead to a significant improvement process (op. CC9), but only to increased consumption of lime.

The results of experiments on the disintegration serosorting concentrate (od. D1-3) show that at a temperature of 125-135 mA°C do not develop reactions (2), (3). In the solution there are only thiosulfate sulfur forms (~10 g/l), sulfide ion (S2-)necessary for the successful progress of disintegration, is missing. In the absence of sulfide ion in the liquid phase particle surface sulphides is not wetted by the solution, sulfides remain hydrophobic and the subsequent flotation leave froth the place with gray.

When the temperature is raised to 145°C in the liquid phase appears sulfide ion (op. D4-7) in the amount of ~1.6 to 6 g/L. the Presence of sulfide ion leads to wetting of the surface of the solution sulphides and, consequently, to the subsequent flotation separation from them of elemental sulfur. The primary indicator of the presence of sulfide ion in the liquid phase is the smell of hydrogen sulfide during depressurization of the autoclave and a yellow stripe on the filter paper, placed in the disintegrated pulp. the pH of the slurry is in the range 8-9.

Optimal parameters for a successful process is the consumption of lime within 65÷70 kg/t S° in loading on the operation and temperature of 145°C. the Increase of the flow rate of the reagent (op. D4) and temperature increase (op. D) lead to a rising transition of sulfur in solution (~10-11%), resulting in increased loss of sulfur solutions. In addition, when the temperature of 145-150°C molten sulfur has a minimum viscosity. The temperature rise above 155°C leads to a sharp increase in the viscosity of liquid sulfur, which complicates the separation of sulphur and sulphides.

Thus, the disintegration is successful under conditions which ensure the presence of sulfide ion in solution in an amount of not less than 1.6 g/L.

Condition for the effective conduct of sulfuric flotation is a highly alkaline environment, as well as the presence of polished-ion battery. As in the case of cereulide flotation, before each operation requires contacting the pulp with reagents agitation. The condition for achieving satisfactory performance of the process is the pH value of the floatated pulp-defined reagent regime propaganda. On sulfuric flotation it should be not less than 11,5. At pH values<11,0 flotation separation of sulphur and sulphides does not occur (op. SF, 2). At pH 11.0 to 12.0 observed the formation of a foamy product (op. SF), but removing the elemental sulfur in the concentrate is low (~46%), and its content in the flotation tailings unacceptably high (29,9%).

When the pH to 12.0 (op. SF) obtained satisfactory performance of the flotation process: removing the sulfur in the concentrate ~ 80% when the content of her ~70%. The flotation tailings contain ~8% of sulfur. Despite the relatively high content of elemental sulfur in the middlings (32,9%), this material being mixed with tails sulfuric (8,1% S°) and cereulide (1,0% S°) flotation, forms a combined sulphide concentrate containing S° 9.7 per cent. The output of this material less than 25% of the cake leaching. The content of sulfur in the sulfide concentrate (~5-10%) allows you to send it in pyrometallurgical production. In this case, the process of sulfuric flotation can be simplified and implemented in a single operation similarly zeros Lydney flotation.

When processing the cake leaching of zinc concentrate for the proposed technology can extract up to 78% of sulfur in the sulfur concentrate, containing its at least 70%. Material with such content of elemental sulfur suitable for autoclave smelting of sulfur on the technology used for sulfur redistribution in Norilsk [Vigabatrin, Imminent, Vasilakos, Meanchevi, Asediko, Amerzon. Hydrometallurgical beneficiation of Nickel-pyrrhotine concentrate. Non-ferrous metals, 1974, No. 9, p.1-6; Maincause, Votewise. The technology of sulfur. - M.: Chemistry, 1985. - 328 S.].

Pyrometallurgy comes reduced by one third the content of elemental sulfur in it <10%, in contrast to 30% of S° in the original cake leaching. In addition, this material comprises more than 85% silver and 75% of the gold content of these metals in the cake leaching), which prevents their loss.

1. The method of separation of elemental sulfur and sulfide concentrate middlings hydrometallurgical processing of sulfide concentrates, including serosorting flotation original pulp of the concentrate to separate sulfur and sulfides from the oxide components of the concentrate with the formation of serosorting concentrate, autoclave disintegr is of pulp serosorting concentrate at a temperature above the melting point of elemental sulfur in the presence of reagent - hydrophilization sulphides, sulphur flotation to separate the disintegrated pulp sulfur and sulfide concentrates, characterized in that before serrasalminae and sulfuric florasiami the pulp is subjected to further agitation with reagents, and the disintegration is carried out at a temperature of 140-150°C.

2. The method according to claim 1, characterized in that the starting slurry middlings agitate first with lime, then with sodium sulfide.

3. The method according to claim 1, characterized in that the pH of the pulp entering serosurvey flotation support at least 10.5.

4. The method according to claim 1, characterized in that the pH of the pulp entering the sulfur flotation support not less than 11,5.

5. The method according to claim 1, characterized in that the resulting sulphur concentrate is separated into marketable sulfur and sulfide product that is sent to the turnover in the campaign preceding sulfuric flotation.



 

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18 cl, 14 tbl, 1 ex, 3 dwg

FIELD: metallurgy.

SUBSTANCE: invention relates to hydrometallurgy. Particularly it relates to method of extraction of nonferrous (Cu, Zn, Co, Ni and others), rare (U, rare earths, Y, Re, Ti, In and others) and precious (Au, Ag, Pt, Pd and others) metals from ores and materials. Method includes leaching of ores in two stages. At the first stage ore and materials treatment is implemented by the first spillage leaching solution with introduction of sulfuric acid and salts of ferric iron in amount, providing in the end of leaching in productive solution molar ratio ion concentration of ferric and ferrous iron no lower than 1:1. At the second stage ore and materials treatment is implemented by the second spillage leaching solution with introduction of sulfuric acid, thiocyanate salts and ferric iron in amount, providing in productive solution molar ratio of ion concentration of thiocyanate and ferric iron no higher than 2:1 and no lower 0.5:1, and ratio of concentration ferric and ferrous iron ions are also no lower than 1:1. Then it is implemented separate processing productive solutions of each stage by means of chemical deposition, sorption and/or electrolysis and spillage solutions return for corresponding stage.

EFFECT: increasing of extraction ratio of nonferrous, rare and precious metals.

5 cl, 5 tbl, 11 ex

FIELD: metallurgy.

SUBSTANCE: invention is related to noble metals metallurgy and can be used for technology of desilverisation and gold extraction from zinc-bearing golden-silver cyanic sediments with increased content of silver. Initial zinc-bearing golden-silver cyanic sediment is leached, at first, in nitric acid solution and then into received pulp excluding filtration it is added caustic soda solution till achieving the concentration NaOH, equal to 100-140 g/l. After it alkaline solution is separated from non-solved sediment. The latter is washed by alkaline solution, dried, molten with fluxes on golden-silver alloy. Received alloy is settled, slag is separated from silver gold-bearing alloy, which is directed to silver refining by means of electrolysis in nitro-acid electrolyte. Electrolysis products are refined cathodic silver and golden sludge, which is refined by well-known methods.

EFFECT: removing of detrimental impurities, essentially, zinc, selenium and tellurium made of initial cyanic sediment.

1 ex

FIELD: chemistry.

SUBSTANCE: invention refers to methods of gold and silver recovery from sulphide concentrates and industrial deposited concentrates. Method involves leaching gold-bearing and argentiferous concentrates with acid thiocarbamide liquors with the oxidiser added and extraction recovering noble metals from leaches. Extraction is preceded with adding thiocyanate ions to leaches in amount to ensure complete transferring thiocyanate gold and silver complexes into the organic phase. Extractant is mixed tributyl phosphate (TBP) and diphenylthiocarbamide (DPTC) in kerosene, containing TBP 1.5-2.0 mole/l and DPTC 0.015-0.022 mole/l. Gold and silver are re-extracted from the organic phase with the reducing agents precipitating noble metals within reduction process.

EFFECT: lower thiocarbamide loss at the stage of noble metal extraction from the leach.

9 cl, 1 tbl, 25 ex

FIELD: metallurgy.

SUBSTANCE: invention relates to method of metals heap leaching, notably, gold from ore. Method includes ore reduction, ore division into fractions, ore dump by uniform in fractions inclined layers with reduction of ore fineness from the low layer to top with separation of layers by perforated polymer film. Then it is implemented stack irrigation by leaching cyanide solution with concentration 0.2-0.8 g/l. Additionally after dump of each ore leaching it is implemented treatment of layer by cyanide solution with strengthen concentration and its standing. Concentration of cyanide solution and standing duration are reduced from low later to top from 2.0-4.0 g/l till 1.0-1.5 g/t and from 5-6 days till 2-3 days correspondingly.

EFFECT: leaching effectiveness increase.

1 dwg

The invention relates to methods for extracting sulfur from sulfur concentrates, in particular of sulphur sludge formed during interphase oxidative purification of hydrogen sulfide-containing gases in the oil industry
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