Method of gold and silver recovery from concentrates

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

 

The invention relates to hydrometallurgy of precious metals (BM) and can be used to extract gold and silver from sulfide concentrates and concentrates produced from placers.

Known application timesaving (thiocarbamide) solutions for the extraction of gold and silver from various types of raw materials (Panchenko A.F., Ladadika CENTURIES, Shalin L.A. Theoretical basis of the process of dissolution of gold, silver and their alloys in acidic solutions of thiocarbamide // X all-Union conference on the chemistry, analysis and technology of the precious metals. The abstracts. Part 1. Novosibirsk. 1976. P.14).

The wide industrial use of urea solutions instead of cyanide is constrained by the higher cost of thiocarbamide compared with sodium cyanide. In this connection there is a need to reduce the specific consumption of thiocarbamide in the process of extracting precious metals. The greatest loss of thiocarbamide associated with the stage of selection of gold and silver from the leaching solutions. So, for example, when selecting the cementation of precious metals required high temperature (not lower than 90C), which, on the one hand, leads to the decomposition of thiocarbamide, and on the other hand, solutions are contaminated with metal ions used for cementation, which hampers their use in circulation and also leads the leads to the loss of thiocarbamide. Sorption extraction of noble metals on activated charcoal leads to irreversible loss of thiocarbamide as a result of its adsorption. When the electrochemical extraction of gold and silver from thiocarbamide of leaching solutions is anodic oxidation of thiocarbamide.

The method for extracting gold from concentrates, including leaching of concentrates sulfuric acid solutions containing thiocarbamide, in the presence of the oxidizing agent is hydrogen peroxide. The leaching process is carried out in two stages. Recovery of gold from solutions after leaching is carried out in the cell with the plate titanium cathodes (extraction of gold cathode sludge 70-80%). After electrolysis gold doizvlekali sorption on active coal (Apparence, Viabilizem, Tab "Thiocarbamide the leaching of antimony concentrates non-ferrous metallurgy, 1987, No. 4, pp. 27-29).

In the work (Ladadika CENTURIES, Shamis L.A. and other Study of the kinetics of dissolution of silver in aqueous solutions of thiourea // Izvestiya vuzov. Non-ferrous metallurgy. 1975. No. 2. P.77-81) have shown the possibility for the recovery of silver in the same way.

The main disadvantage of the considered method for extracting gold and silver from concentrate is a multi-stage (two-stage leaching and two extraction stages BM race the thief leaching) and as a consequence, large volumes of recyclable solutions, as well as loss of thiourea in the stages of sorption and electrolysis.

There is also known a method of extracting silver from sulfide flotation concentrates, including leaching of raw materials sulfuric acid solution of thiocarbamide in the presence of an oxidant (oxygen) at a temperature of 50-90C. the Concentration of thiocarbamide ranged from 60-100 g/l, which is about the same 0,80-1,30 mol/l sulfuric acid concentration of 5-20 g/l (0.05 to 0.20 mol/l). The ratio of concentrate and leaching solution (T:W) is equal to 1:3, the leaching time was about 5 hours. From the leaching solution, combined with leaching waters, silver besieged by cementation on aluminum or iron plates at a temperature of 80-90C for 2.5-3 hours. The result has been the cement residue with a silver content of 82-95% (RF patent No. 2237092, publ. 27.09.2004). This method can be applied also for the extraction of gold from concentrates.

The main disadvantage of this method, as noted above, is the possibility of decomposition of thiourea on stage grouting. When conducting integrated experimental-industrial tests of this method, it was found that loss of thiocarbamide on stage cementation occur when the temperature of the solution to about 90C. in Addition, empirically fitted is but when the extraction of gold from concentrates in this way on stage grouting is necessary to maintain the temperature below 95C, which, in turn, leads to more significant losses of thiocarbamide.

The disadvantages of this method include the pollution solution of metal ions in the cementation process on aluminum or iron plates.

As the closest analogue to the technical nature and purpose of the selected method of extracting gold and silver from concentrates, including leaching of gold and silver sour thiocarbamide solution in the presence of an oxidising agent and then removing them from leaching solutions, for example, cementation (meretukov M.A., Orlov A.M. metallurgy precious metals, Foreign experience, M, metallurgy, 1991, s-203).

As in the method according to U.S. Pat. Of the Russian Federation No. 2237092, the disadvantages of the closest analogue is loss of thiourea on stage cementation due to its decomposition at temperatures of 90C and above, as well as contamination of the solution with metal ions in the cementation process on aluminum or iron plates.

The objective of the invention is to optimize the method for extracting gold and silver from concentrates by reducing consumption of thiocarbamide at the stage of extraction of precious metals from thiocarbamide of leaching solutions and exclusion the pollution solution extraneous ions of metals.

The problem is solved by the proposed method for extracting gold and silver from concentrates, including leaching of gold and silver sour thiocarbamide solution in the presence of an oxidising agent and then removing them from leaching solutions, which in contrast to the known method of extracting gold and silver from solutions of lead leaching extraction with subsequent reextracting, and before extraction in the leaching solutions enter the thiocyanate ions in an amount to provide a weight transfer thiocyanate complexes of gold and silver in the organic phase, as a solvent a mixture of tributyl phosphate (TBP) with diphenylthiourea (DFD) in kerosene, and reextraction gold and silver from the organic phase carry out a reducing agent capable of in the recovery process to precipitate the precious metals.

The method is as follows. Source sulfide concentrate or concentrates from placers containing gold and silver, is subjected to acid leaching solution containing from 0.8 to 1.30 mol/l of thiocarbamide, in the presence of an oxidant. The acid can be used sulfuric acid in a quantity of 0.05-0.2 mol/l hydrochloric acid in an amount of 0.1 to 0.4 mol/l, and as an oxidizer such well - known oxidants, as, n is the sample, the sodium persulfate or ammonium permanganate sodium or potassium ferric salt (chloride or sulfate) of manganese dioxide, hydrogen peroxide and others.

In the particular case of the invention, the leaching of lead using as oxidant chloride or ferric sulfate in the amount of 0.07 mol/L.

The leaching is carried out at a mass ratio of concentrate and leach solution (T:W)of 1:3, about five hours. At the end of the leaching slurry is filtered.

The resulting sludge (cake) washed first with a solution containing an initial concentration of acid and thiocarbamide, and then with water at the ratio of the weight of the precipitate and washing solutions of 1:1, at each stage of washing.

Further washing of the cake can be used for working capital solutions.

Before extraction in the leaching solution is injected thiocyanate ions in the form of thiocyanates of alkali metals, preferably of potassium, sodium or ammonium, in the amount of 0.3 to 0.5 moles per liter of solution, after which the leaching solution is directed to the stage of extraction.

The choice of the concentration of thiocyanate ions in the range of 0.3 to 0.5 mol/l provides high performance extraction of gold and silver. Increasing the concentration of thiocyanate ions above 0.5 mol/l impractical, since entails the increase of consumption, p the agents, but does not increase the degree of extraction of precious metals in the extract. Reduction of the concentration of the thiocyanate ions of less than 0.3 mol/l leads to a decrease in the degree of extraction of gold and silver in the extract. So, for example, ammonium thiocyanate shown that at a concentration of 0.15 mol/l, the extraction of silver by the extractant composition of 0.022 DFTC and 1,82 TBP (mol/l) in organic and aqueous phase (A:B)=1:2 is only 52%, and at lower concentrations of ammonium thiocyanate gold and silver no longer be extracted.

In the experimental studies found that thiocyanate complexes of gold and silver is moving from the leaching solution to extract with a high degree of extraction of the entire thiocarbamic remains in the leaching solution.

In addition, in the proposed method, due to the extraction of gold and silver from solution leaching extraction eliminates the possibility of contamination of the extract ions of other metals in contrast to the known method, in which gold and silver is removed by cementation on metallic plates.

As a solvent, a mixture containing 1.5 to 2.0 mol/l of tributyl phosphate (TBP) and 0.015-0,022 mol/l of diphenylthiourea (DFD) in kerosene.

Empirically it is shown that the ratio of TBP and DFD in kerosene obespechivayushchi the distribution coefficient of gold and silver in the extraction process and, accordingly, completion of the translation thiocyanate complexes of gold and silver in the organic phase.

Increasing TBP concentration above 2.0 mol/l impractical, since entails additional costs, but does not improve the recovery of precious metals. Reduction of the concentration of TBP below 1.5 mol/l leads to a decrease in the solubility DFTC and, consequently, to reduce the degree of extraction of precious metals.

The upper limit concentration DFTK - 0,022 mol/l due to its solubility. Use for extraction solutions with concentrations DFTC below 0.015 mol/l leads to a significant reduction in the extraction of precious metals in the organic phase.

The extraction is carried out at the optimal ratio of organic and aqueous phases, as measured by achieving a sufficiently high recovery of precious metals and rational consumption of reagents. In General the invention, the ratio A:b can be equal to 1:(1-6).

The proposed conditions for extraction of gold and silver in combination with the introduction of the thiocyanate ion in the leaching solutions before extraction provide high recovery of precious metals in the organic phase. Formed after separation of the phases, the aqueous phase containing the thiocarbamide, may be directed into circulation on the stage above is Oceania or washing of the cake.

Subsequent reextraction gold and silver from the organic phase is carried out using known reducing agents, which are in the process of recovering precipitated precious metals.

The specified requirements are met by such a strong reducing agents, as hydrides and borohydride alkali metals.

For the extraction of gold may be used and other well-known precipitants, such as oxalic acid, hydroquinone, sodium nitrite, formaldehyde, salts of hydrazine and other (Radelet. "The chemistry of gold". M.: Mir, 1982. 259). It should be noted that for the deposition of gold can not be applied reductants, the use of which leads to the formation of colloidal solutions of gold (sols), in particular a tin chloride (SnCl2).

The deposition of silver from the organic phase takes place most effectively when using sodium borohydride.

Precipitated gold and silver is filtered from the mixture, resulting in the target product in powder form.

This extractant is not destroyed and does not lose the ability to extract precious metals, it can be used after separation of the phases in circulation for subsequent extraction.

In the case of extracting gold and silver from man-made concentrates, which contain mercury leaching together with noble metals in the solution of veselaj is of almost completely and mercury. Mercury also sextrailers together with gold and silver, and is precipitated from the extract with a suitable reducing agent such as sodium borohydride. Next, from the obtained interfacial precipitate the mercury is separated from the gold and silver known methods, in particular, by washing the precipitate of noble metals with nitric acid.

Thus, the inventive method provides for the achievement of the technical result consists in a substantial decrease in the consumption of thiocarbamide by reducing its losses at the stage of extraction of precious metals from solution leaching extraction and to exclude contamination of the solution ions extraneous metals, which together with the reduction of energy consumption (at the expense of cementation) and use of reagents in circulation contributes to the optimization of the entire hydrometallurgical processing of gold - and silver-containing mineral raw materials in General.

The possibility of carrying out the invention with the achievement of the technical result is confirmed by the examples in which the extraction of silver was obtained from sulfide concentrate Coastal mining company "East", and gold from the concentrates from placers Sidetrack ore-placer node Primorsky Krai.

For all examples was calculated squozen the e extraction of gold and silver, defined as the product of the degree of extraction of the BM in the leaching solution and the degree of extraction of the BM in the organic phase, as well as loss of thiocarbamide as the difference between its concentration in the leaching solution before and after the recovery of precious metals from the leaching solution by extraction (by the present method) or cementation (by known means).

The conditions of examples of the proposed method and the results are tabulated in which examples 1-11 relate to the extraction of silver, 12-23 - extract the gold.

The experimental data clearly show that, at comparable rates of extraction of gold and silver from concentrates for a well-known and inventive ways of losing thiocarbamide during their extraction from the leaching solution by the present method at least 6 times lower than by a known method.

Example 1. 100 kg of flotation silver sulfide concentrate is leached with 300 l of a solution containing of 1.30 mol/l of thiocarbamide, 0.1 mol/l sulfuric acid, in the presence of an oxidant ions of trivalent iron (ferrous sulfate) at a concentration of 0.07 mol/l, at a mass ratio of concentrate and leaching solution, equal to 1:3, for five hours. The slurry is filtered. The cake washed twice, first with the initial solution, then with water, at mass with respect to the spacecraft and the leaching solution both times, equal to 1:1. Received 480 l combined with leaching water solution containing silver 0,330 mol/l; the extraction of silver in the leaching solution was 98.9%. The slurry is filtered. To the solution add ammonium thiocyanate 0.5 mol/l and direct to the stage of extraction. As extractant used solution in kerosene mixture of TBP (1.82 mol/l) and DFD (0,022 mol/l). The extraction is carried out at A:B=1:5. The degree of extraction of silver in the organic phase is 84%. After separation of the phases the aqueous phase is sent to the turnover on the re-leaching. From the organic phase silver extravert (precipitated) aqueous solution of 0.5 mol/l of sodium borohydride. The resulting mixture was filtered and the obtained silver powder. Through the selection of silver 83,1%, loss of thiocarbamide 4,50%.

Example 2. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following ratios: a:is 1:1. The extraction of silver in the leaching solution is 98.8%; the degree of extraction of silver in the organic phase 91,40%; end-to-end extraction of silver 90,35%; loss of thiocarbamide of 4.83%.

Example 3. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following indicators: the ratio is:B=1:2. The extraction of silver in the leaching solution is reached 98.9%; the degree of extraction of silver in the organic phase 86,11%; end-to-end extraction of silver 85,16%, loss of thiocarbamide 4,39%.

Example 4. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following indicators: as the thiocyanate ions enter NaCNS at a concentration of 0.3 mol/l, the Extraction of silver in the leaching solution is 93.0%; the degree of extraction of silver in the organic phase 84,25%; end-to-end extraction of silver 78,37%, loss of thiocarbamide br4.61%.

Example 5. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following ratios: A:B=1:10. The extraction of silver in the leaching solution is reached 98.9%; the degree of extraction of silver in the organic phase 22,45%; end-to-end extraction of silver 22,20%, loss of thiocarbamide 4,80%. Low extraction of silver in the organic phase is related to the ratio A:B, well beyond the optimal relationship 1:(1-6).

Example 6. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following indicators: extractant 1.82 mol/l of TBP and 0.007 mol/l DFTC. The extraction of silver in RA is creative leaching is 95,0%; the degree of extraction of silver in the organic phase 26,10%; end-to-end extraction of silver 24,79%, loss of thiocarbamide 4,55%. Low extraction of silver in the organic phase is related to the amount of extract DFTC, significantly less than the value of the lower limit of 0.015 mol/L.

Example 7. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following indicators: extractant 1.82 mol/l of TBP and 0.015 mol/l DFTC. The extraction of silver in the leaching solution is 98,7%; the degree of extraction of silver in the organic phase 81,65%; end-to-end extraction of silver 80,57%, loss of thiocarbamide of 4.13%.

Example 8. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following indicators: extractant of 0.91 mol/l of TBP and 0.010 mol/l DFTC. The extraction of silver in the leaching solution is reached 98.9%; the degree of extraction of silver in the organic phase 19,78%; end-to-end extraction of silver 19,57%, loss of thiocarbamide 4,84%. Low extraction of silver in the organic phase is associated with quantities in the extract TBP and DFTC, less than their value, the lower the optimal limits.

Example 9. Leaching of silver sulphide concentrate extraction and reextraction silver perform similarly note the 1 ru, except for the following indicators: extractant of 1.50 mol/l of TBP and 0,022 mol/l DFTC, as the thiocyanate ions enter the KCNS in a concentration of 0.5 mol/L. Extraction of silver in the leaching solution is 95.2 percent; the degree of extraction of silver in the organic phase 84,27%; end-to-end extraction of silver 80,19%, loss of thiocarbamide 4.32 percent.

Example 10. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following indicators: extractant 2.5 mol/l of TBP and 0,022 mol/l DFTC. The extraction of silver in the leaching solution is to 94.8%; the degree of extraction of silver in the organic phase 86,67%; end-to-end extraction of silver 82,19%; loss of thiocarbamide 4,30%. Increasing the concentration of TBP in the composition of the extractant is higher than the marginal value of the optimal interval (2.0 mol/l), practically does not lead to a noticeable increase in the recovery of silver from a solution of leaching for the extraction, and therefore impractical.

Example 11. Leaching of silver sulphide concentrate extraction and reextraction silver carried out analogously to example 1, except for the following indicators: extractant 2.0 mol/l of TBP and 0,022 mol/l DFD; a:=1:6. The extraction of silver in the leaching solution is 95,1%; the degree of extraction of silver in the organic phase 86,20%; end-to-end retrieval ser the bra 82,00%; loss of thiocarbamide 4,53%.

Example 12. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 1. Gold recovery in the leaching solution is 97,2%; the degree of extraction of gold in the organic phase 90,0%; gold recovery of 87.5%, a loss of thiocarbamide 4,20%.

Example 13. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 2. Gold recovery in the leaching solution is reached 98.9%; the degree of extraction of gold in the organic phase 99,12%; gold recovery 98,03%, loss of thiocarbamide of 4.75%.

Example 14. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 3. Gold recovery in the leaching solution is reached 98.9%; the degree of extraction of gold in the organic phase 95,39%; gold recovery 94,34%, loss of thiocarbamide 4,20%.

Example 15. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 4. The extraction of gold in solution leaching 96.7%; the degree of extraction of gold in the organic phase 89,67%; gold recovery 86,71%, loss of thiocarbamide 4,65%.

Example 16. Recovery of gold from the concentrate, leaching, extraction and reextraction, are analogues of the but example 5. Gold recovery in the leaching solution is reached 98.9%; the degree of extraction of gold in the organic phase 41,18%; gold recovery 40,72%, loss of thiocarbamide of 4.75%. Low gold extraction in the organic phase is associated with the ratio A:B=(1:10), well beyond the optimal relationship 1:(1-6).

Example 17. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 6. Gold recovery in the leaching solution is 95.6%; degree of extraction of gold in the organic phase 31,84%; gold recovery 30,43%; loss of thiocarbamide 4,47%. Low gold extraction in the organic phase is related to the amount of extract DFTC lower (0,007 mol/l)than the value of the lower limit of 0.015 mol/L.

Example 18. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 7. Gold recovery in the leaching solution is 98,7%; the degree of extraction of gold in the organic phase 85,33%; gold recovery 84,21%, loss of thiocarbamide to 4.23%.

Example 19. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 8. Gold recovery in the leaching solution is reached 98.9%; the degree of extraction of gold in the organic phase 24,31%; end-to-end extraction of gold is and 24,04%; loss of thiocarbamide 4,94%. Low gold extraction in the organic phase is associated with quantities in the extract TBP and DFTC, less than their value, the lower the optimal limits.

Example 20. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 9. Gold recovery in the leaching solution is to 98.2%; the degree of extraction of gold in the organic phase 88,73%; gold recovery 87,11%; loss of thiocarbamide was 4.42%.

Example 21. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 10. Gold recovery in the leaching solution is 97,6%; the degree of extraction of gold in the organic phase 91,40%; gold recovery 89,23%; loss of thiocarbamide 4,30%. Increasing the concentration of TBP in the composition of the extractant is higher than the marginal value of the optimal interval (2.0 mol/l), practically does not lead to a noticeable increase gold recovery from solution leaching for the extraction, and therefore impractical.

Example 22. Recovery of gold from the concentrate, leaching, extraction and reextraction is carried out analogously to example 11. Gold recovery in the leaching solution is 94,4%; the degree of extraction of gold in the organic phase 91,30%; gold recovery 86,20%; loss of thiocarbamide 4,36%.

Example 23. The implementation of the invention using a working solution, obtained after deposition of gold according to example 12.

Gold-bearing concentrate is poured thiocarbamide-thiocyanate working solution with a gold content of 5.710-5mol/l and conduct leaching, leaching and recovery of gold under the conditions of example 12. The extraction of gold at the stage of leaching were $ 97.6%; the degree of extraction of gold in the organic phase 94,85%; gold recovery 92,63%; loss of thiocarbamide of 4.9%.

Example 24. (Extraction of silver by the method prototype)

100 kg of sulfide flotation concentrate, containing silver, leached 300 l of a solution containing of 1.30 mol/l (~100 g/l) thiocarbamide, 0.1 mol/l sulfuric acid, in the presence of an oxidant at a temperature of 60C and the mass ratio of concentrate leaching solution, equal to 1:3, for five hours. The slurry is filtered. The cake washed twice, first with the initial solution, then with water, when the mass ratio of the cake and the leaching solution both times equal to 1:1. Received 480 l joint solution containing silver 0,328 mol/l, the extraction of silver in the solution amounted to 97.8 per cent. From the leaching solution, combined with leaching waters, silver besieged by cementation on iron plates at a temperature of 90C for 3 hours. During this time, silver is deposited from the region of the target with the release of 95%. The residual content of silver in the solution 0,0016 mol/l; silver content in the sediment of 82%. Through the extraction of silver from concentrate, 80.2 per cent. After cementation, the concentration of thiocarbamide in the circulating solution of 0.60 mol/l (45.6 g/l). Loss of thiocarbamide on stage grouting consists of 27.2%.

Example 25. (Extraction of gold by the method prototype)

1 kg of concentrate containing gold, leached 3 l of a solution containing of 1.30 mol/l of thiocarbamide, 0.1 mol/l sulfuric acid, in the presence of an oxidant at room temperature and the ratio of T:W=1:3 for five hours. The slurry is filtered. The cake washed twice, first with the initial solution, then water, both times at T:W=1:1. Received 4.8 l joint solution containing gold and 8.4*10-4mol/L. Extraction of gold in the solution amounted to 97.1%. From the leaching solution, combined with Promode, the gold is precipitated by cementation on iron plates at a temperature of 90C for 3 hours. During this time, the gold is deposited from a solution with a yield of 55%. Additional heating the solution to 95C can increase the degree of deposition of gold to 95,0%. Gold recovery from the concentrate of 92.2%. After cementation, the concentration of thiocarbamide in the circulating solution of 0.60 mol/l (45.6 g/l). Loss of thiocarbamide 26,7%.

# example The composition of the extracting solution mol/lThe ratio of organic and aqueous phasesThe concentration of thiocyanate ions mol/lThe degree of BM extract the organic phase %Through removing the CV %Loss of thiocarbamide %
The extraction of silver
11,82 TBP, 0,022 DFTC1:50,5 (NH4CNS)84,0083,14,50
21,82 TBP, 0,022 DFTC1:10,5 (NH4CNS)91,4090,35a 4.83
31,82 TBP, 0,022 DFTC1:20,5 (NH4CNS)86,1185,164,39
4 1,82 TBP, 0,022 DFTC1:50,3 (NaCNS)84,2578,37br4.61
51,82 TBP, 0,022 DFTC1:10,5 (NH4CNS)22,4522,204,80
61,82 TBP, 0,007 DFTC1:50,5 (NH4CNS)26,1024,794,55
71,82 TBP, 0,015 DFTC1:50,5 (NH4CNS)81,6580,574,13
80,91 TBP, 0,010 DFTC1:50,5 (NH4CNS)19,7819,574,84
91,50 TBP, 0,022 DFTC1:50,5 (KCNS)84,2780,194,32
102,50 TBP, 0,022 DFTC1:50,5 (NH4CNS)86,6782,194,30
112,0 TBP, 0,022 DFTC1:60,5 (NH4CNS)86,2082,04,53
121,82 TBP, 0,022 DFTC1:50,5 (NH4CNS)90,087,54,20
131,82 TBP 0,022 DFTC1:10,5 (NH4CNS)99,1298,03 4,75
141,82 TBP, 0,022 DFTC1:20,5 (NH4CNS)95,3994,344,20
151,82 TBP, 0,022 DFTC1:50,3 (NaCNS)89,6786,714,65
161,82 TBP, 0,022 DFTC1:100,5 (NH4CNS)41,1840,724,75
171,82 TBP, 0,007 DFTC1:50,5(NH4CNS)31,8430,434,47
181,82 TBP, 0,015 DFTC1:50,5 (NH4CNS) 85,3384,214,23
190,91 TBP, 0,010 DFTC1:50,5(NH4CNS)24,3124,044,94
201,5 TBP, 0,022 DFTC1:50,5 (KCNS)88,7387,114,42
212,5 TBP, 0,022 DFTC1:50,5 (NH4CNS)91,4089,234,30
222,0 TBP, 0,022 DFTC1:60,5 (NH4CNS)91,3086,204,36
231,82 TBP, 0,022 DFTC1:5 0,5 (NH4CNS)94,8592,634,90

1. The method of extracting gold and silver from concentrates, including leaching of gold and silver sour thiocarbamide solution in the presence of an oxidising agent and then removing them from leaching solutions, characterized in that the extraction of gold and silver from solutions of lead leaching extraction with subsequent reextracting, and before extraction in the leaching solutions enter the thiocyanate ions in an amount to provide a weight transfer thiocyanate complexes of gold and silver in the organic phase, as a solvent a mixture of tributyl phosphate (TBP) with diphenylthiourea (DFD) in kerosene, and reextraction gold and silver from the organic phase is performed by the reducing agents that can process recovery deposition of noble metals.

2. The method according to claim 1, characterized in that the leaching is carried out with acidic solution containing 0,80-1,30 mol/l of thiocarbamide.

3. The method according to claim 1 or 2, characterized in that the acid in the leaching solution using sulfuric acid, taken in an amount of 0.05-0.2 mol/l or chloride-hydrogen acid, taken in an amount of 0.1 to 0.4 mol/L.

4. The method according to claim 1, characterized in that wimalasiri who are using as oxidant chloride or ferric sulfate in the amount of 0.07 mol/L.

5. The method according to claim 1, characterized in that the leaching is carried out at a mass ratio of concentrate and leach solution, equal to 1:3.

6. The method according to claim 1, characterized in that as a solvent, a mixture containing 1.5 to 2.0 mol/l of TBP and 0.015-0,022 mol/l DFTC in kerosene.

7. The method according to claim 1, wherein the thiocyanate ions is introduced into the leaching solutions in the form of thiocyanates of alkali metals or ammonium in the amount of 0.3-0.5 mol/L.

8. The method according to claim 1, characterized in that as a reducing agent in the Stripping of gold and silver from the organic phase using borohydride sodium.

9. The method according to claim 1, wherein when extracting gold and silver from man-made concentrates containing mercury deposited from the organic phase gold and silver mixed with mercury is washed with nitric acid.



 

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FIELD: metallurgy.

SUBSTANCE: method of metal extraction recovery from solid-phase of raw-materials includes crushing of raw materials, leaching by means of direct microemulsion, consisting of aqueous phase and organic phase, containing kerosene in the capacity of extractant di-(2-ethylhexyl) sodium phosphate, separation of solid phase and re-extraction of recovered metals. At that leaching is implemented by microemulsion, consisting of 30-75% turn of aqueous phase and containing in organic phase di-(2- ethylhexyl) sodium phosphate in amount 1.0-2.0 mole/l and additionally introduced di-(2-ethylhexyl) phosphoric acid in amount 0.3-0.6 mole/l. Into compound of organic phase of microemulsion i can be additionally introduced aliphatic alcohols in amount till 5% turns.

EFFECT: ability of selective extraction of nonferrous and rare metals raw at leaching stage, without usage of concentrated acid and expensive organic solvent.

9 tbl, 7 ex

The invention relates to the separation of niobium from concentrated solutions containing niobium, tantalum and titanium

The invention relates to hydrometallurgy rare and scattered metals and can be used when removing India from solutions of zinc-lead production, in the processing of secondary raw materials

The invention relates to the extraction of substances extraction and can be used in ferrous and nonferrous metallurgy, as well as treatment of industrial and domestic wastewater

The invention relates to the extraction of substances organic extractants from aqueous solutions and can be used in ferrous and nonferrous metallurgy

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

 // 2350667

FIELD: technological processes; metallurgy.

SUBSTANCE: method for cuvette-heap leaching of metals is related to hydraulic metallurgy and may be used in leaching of non-ferrous, rare and precious metals ores. Method includes treatment of mineral mass with solution of leaching agent and metal extraction. Treatment of mineral mass with solution of leaching agent and metal extraction is carried out in two stages. Previously mineral mass is placed in cuvettes with hydraulically insulated walls and bottom. Then solution of initial reagent is introduced, and local portion activation of produced pulp is carried out with provision of secondary reagents. After activation pulp is exposed to fractioning with extraction of sludge-clay and sand fractions. Sand fraction is dehydrated. Productive solution and sludge-clay fraction produced as a result are exposed to sorption or electric sorption leaching. Stacks are formed from sand fraction, and material is exposed to heap leaching. Liquid phase that remained after leaching of sand and sludge-clay fraction is additionally strengthened and sent for heap leaching.

EFFECT: higher efficiency of process.

2 cl, 1 ex

FIELD: metallurgy.

SUBSTANCE: it is implemented treatment of copper-nickel sulfide concentrate by fluoride and/or ammonium bifluoride at the temperature 165-210°C during 2-3 hours with forming of hard fluorination products and emission of ammonia gas and water vapor. Fluoride and/or ammonium bifluoride are used in amount 1.0-1.2 from stoichiometric with respect to overall content in copper-nickel concentrate of silicates and pyrrhotine. Hard fluorination products exposed to water leaching at the temperature 40-60°C and S:L=1:5-6 during 1-2 hours with transformation to residue nickel sulfide, copper, cobalt, platinum metal and magnetite, and into solution - fluorine-ammonium silicon saline, magnesium, iron, aluminium and calcium. Residue is separated from solution and exposed to magnetic separation with magnetite extraction. Received solution is neutralised by ammonia water till providing of pH 8-10 with regeneration of fluoride and/or ammonium bifluoride and receiving of residue of magnesium, iron, aluminium, calcium and silica hydroxides. In the capacity of ammonia water for solution neutralisation can be used ammonia gas and water vapors from the stage of raw materials fluorination. Formed ammonium fluoride and/or bifluoride are returned to raw materials treatment stage.

EFFECT: energy content reduction and providing of selectivity sulfide crude ore treatment at less number of operations.

5 cl, 1 dwg, 4 ex

FIELD: metallurgy.

SUBSTANCE: invention concerns treatment of hard gold-arsenic ores. Particularly it concerns antimonous sulphide ores and concentrates. Method includes without oxidising melting in smelting chamber with receiving of matte and slag melts and treatment of melting products by metallic phase. At that without oxidising melting is implemented continuously in circulating melted slag with out of melting products into settling chamber to interphase boundary slag - matte. Before melting circulating melted slag is separated from operating gases. For circulating it is used maximum separated from matte slag. Treatment of matte by metallic phase is implemented in continuous operation. Furnace for processing of hard gold -arsenic ores and concentrates includes smelting chamber. Furthermore, it is outfitted by recycling contour, containing of gas-lift unit with tuyeres and descending and ascending channels of melted slag, gas separating and settling chambers. Gas separating chamber is communicated with smelting chamber through bleed blowhole by means of channel for separation of working gas of gas-lift unit and gas separating chamber from circulating melted slag. Smelting chamber immersed into settling chamber to interphase boundary slag - matte. Settling chamber contains gas flue for withdrawal of sublimates and low blowing melting products.

EFFECT: increasing of noble metals extraction into matte.

8 cl, 3 dwg

FIELD: metallurgy.

SUBSTANCE: invention refers to hydrometallurgy, particularly to facility for extraction of precious metals, notably, gold and silver out of tails and other materials with high contents of fine fractions. The facility consists of a solution proof base and a border and also of a drainage system. The drainage system is made in form of a perforated tube with a gate at a discharge end. On the solution proof base there are successively laid: a layer of crushed stone of 2-5 mm size and of 150-200 mm thickness, then a layer of crushed stone of 5-40 mm size and 300-600 mm thickness, in the central part of which the perforated tube is located, then there is a layer of fabric out of geotextiles and a layer of crushed stone of 2-5 mm size and of 200-300 mm thickness; also the solution proof base is made with counter incline of 3-5° from each side from two counter borders into the direction of the perforated tube and also with the incline to horizon of 2-4° into the direction of the discharge end of the perforated tube in a vertical plane passing through the central longitudinal axis of the perforated tube.

EFFECT: facilitating extraction of precious metals out of materials with high contents of fine fractions and also simplification and reduction in cost of facility and its operation.

3 dwg

FIELD: metallurgy.

SUBSTANCE: invention concerns method of golden-antimonial-arsenical sulphide concentrates processing. Method includes melting of concentrates into slag and matte with gold concentration in matte and with condensation of formed arsenious sublimate and gold extraction from melted matte by antimony melt. At that before gold extraction into matte after separation of slag it is introduced soda at stoichiometric ratio to sulphide of antimony 3:1. Gold extraction is implemented by means of feeding matte melt in disperse state under antimony melt.

EFFECT: increasing of complex extraction level of valuable components from concentrate, in increasing of gold extraction speed, in reduction of materials consumption and equipment level of sophistication.

4 cl, 1 ex

FIELD: metallurgy.

SUBSTANCE: invention concerns method of gold-bearing sulfide concentrates bacterial oxidation by hydrometallurgy method at gold receiving with usage of microorganisms. Method includes three-stage oxidation of concentrate by means of microorganisms' aggregation at speed of pulp passage, not exceeding reproduction rate of microorganisms' aggregation. Into concentrate it is added sulfuric acid in amount not more than 1:100 at chemical ratio of iron and arsenic (Fe/As) less than 2.9, from 1:100 till 2:100 at ratio Fe/As from 3.4 till 3.5, from 2:100 till 3:100 at ratio Fe/As from 3.6 till 3.9. At that not less than 70% of sulfuric acid of its total amount is added to sulfide concentrate at stage of preliminary preparation and not more than 30% of sulfuric acid - at the first stage of oxidation.

EFFECT: increasing of sulfides oxidation rate and expansion of manufacturing capabilities of the method.

1 tbl

FIELD: metallurgy.

SUBSTANCE: invention concerns hydrometallurgy of nonferrous and noble metals, mainly extraction of copper and gold from sulfur wastes which are wastes from sulfuric acid manufacturing, and can be used at dense, cuvet and percolation leaching. Method includes piling of sulfur wastes to antifilter basis, influence upon it of precipitations during its storage, collection of underspoil acid water, its additional if necessary acidation by sulfuric acid and copper leaching at values of pH and Eh of productional solutions 2.0-2.5 and 500-540 mV respectively. After reduction of copper concentration less than 100 mg/l there are reduced content of oxidant and acid in leaching solution, thiocarbamide is introduced in it, noble metals are leached, and mainly it is gold. Extracted in solution copper and gold are isolated by cementation.

EFFECT: reduction of reagent unit discharge, expenditures of energy and increasing of copper and gold extraction ratio.

4 cl, 6 tbl, 6 ex

FIELD: metallurgy, mining.

SUBSTANCE: method for gold extraction from cyanic pulp includes gold sorption by means of pulp directing to ion-exchange resin by backflow in cascade of appliances. At that sorption is implemented at the control of concentration of gold in pulp liquid phase of each appliance. At reduction of gold concentration till 0.9 mg/l it is implemented extraction of resin from appliance and its regeneration by solution of caustic soda and/or sodium chloride brine.

EFFECT: increasing of resin sorptive capacity and extraction increasing of solute gold.

2 cl, 2 tbl, 1 ex

FIELD: noble metal hydrometallurgy.

SUBSTANCE: invention relates to method for acid leaching of platinum method from secondary raw materials, in particular from ceramic support coated with platinum metal film. Target metals are leached with mixture of hydrochloric acid and alkali hypochlorite at mass ratio of OCl-/HCL = 0.22-0.25 and redox potential of 1350-1420 mV.

EFFECT: decreased leaching temperature, reduced cost, improved platinum metal yield.

2 ex

FIELD: metal recovery, in particular noble metals from technologically proof raw materials.

SUBSTANCE: method includes raw grinding to 0.2 mm; blending with batch containing halogen salts and/or oxygen-containing salts, and mixture opening: cake cooling, leaching with simultaneous reaction pulp agitation with hot water, and metal recovery from solution and insoluble residue. Opening is carried out in electrical furnace at 100-120oC preferably at redox potential of 1.8-2.6 V, by elevating of temperature up to 450-560oC at rate of 8-10oC/min and holding for 1-7 h at highest mixture redox potential. Opened and cooled cake is grinded and leached in opened agitator.

EFFECT: environmentally friendly method with increased yield; utilization of unconventional noble and non-iron metal sources.

1 cl, 2 tbl

FIELD: non-ferrous metallurgy; leaching-out polymetallic hard-to-open materials.

SUBSTANCE: proposed method includes treatment of material with chlorine in aqueous solution containing chlorine ions which is stirred in anode space of electrolyzer with separated anode and cathode spaces; as a result pulp is obtained; leaching-out operation is performed in anode space of electrolyzer separated from cathode space by cation-exchange membranes; operation is performed in aqueous solution containing hydrochloric acid in presence of nitrogen oxides at additional delivery of chlorine-containing gas by suction of this gas into rarefaction zone formed by impeller stirring the pulp. Gas formed during leaching-out process is combined with chlorine-containing gas; layer of finely-dispersed particles formed on surface of pulp is removed and is fed to stirring zone in lower part of anode space; productive leaching-out solution obtained after separation of it from insoluble residue is delivered to cathode space of electrolyzer; leaching-out process is performed at anode potential ensuring discharge of chlorine ions and cathode potential not exceeding the potential of discharge of hydrogen ions. Device proposed for realization of this method has housing with cover; interior of housing is divided into anode and cathode spaces with anodes and cathodes located inside them, units for loading the initial materials and discharging pulp formed during leaching-out process and units for mixing and feeding the chlorine-containing gas; anode space is made in form of chamber and cathode chambers are located on its opposite sides; anode chamber walls contain cation-exchange membranes; stirring unit is provided with impeller located below lower edge of anodes; vertical fins are provided on inner surface of anode chamber at level of impeller.

EFFECT: increased rate of extraction of beneficial components from initial material into solution for further extraction of them from solution.

17 cl, 3 dwg, 1 tbl, 8 ex

FIELD: mining art; hydro-metallurgical processing of ores and concentrates; extraction of beneficial components by underground leaching, heap leaching, vessel leaching and tank leaching.

SUBSTANCE: proposed method includes preparation of material for leaching-out process, delivery of leaching solution, discharge, collection and reworking of productive solution; intensification of leaching-out process is performed through ultrasonic treatment of material which is preliminarily saturated with solution of reagent (or water)inert to beneficial component and dissolving harmful admixtures. After discharge of leaching solution (or water), beneficial component is leached-out by leaching solution till reduction of its concentration in productive solution corresponding to maximum level obtained during standard leaching-out process. Then periodic ultrasonic treatment of material is performed again at contact with leaching solution till concentration of beneficial component in productive solution gets equal to permissible level for reworking of this solution in settling plant. Periodicity of ultrasonic treatment is determined by special relationship; radiators are mounted in cylindrical cavities (wells) or on surface of material.

EFFECT: enhanced intensification and efficiency due to increased rate of extraction of beneficial components; reduced consumption of reagents.

5 cl, 3 dwg,1 ex

FIELD: metallurgy; production of platinum and palladium concentrates and silver from platinum-containing raw materials.

SUBSTANCE: proposed method includes sulphatizing roasting and/or sulphatizing of platinum-containing raw material at temperature of 200-600 C for 1-17 hours. Cinder is molten with sodium chloride at mass ratio of 1:(1-10) at temperature of 600-900 C; fusion cake is leached-out with water at mass ratio of fusion cake to water equal to 1: (1-10) at temperature of 80-90 C. After filtration of this pulp filtrate and residue are obtained; filtrate settles separating the silver concentrate in form of sediment of insoluble salt of silver chloride from liquid phase. Silver concentrate is washed with solution of concentrated hydrochloric acid and water at mass ratio of 1:10; leaching residue is washed with solution of concentrated hydrochloric acid and water at ratio 1:1, after which it is washed with water. Washing water is mixed with liquid phase obtained after separation of silver chloride from filtrate and sediment in form of platinum-palladium concentrate is let to settle; this sediment is separated by filtration and is washed with water.

EFFECT: complete primary extraction of platinum metals from platinum-containing raw material; reduced toxicity; reduced duration of process; reduced power requirements.

1 tbl, 1 ex

FIELD: sludge recovery from surface depositions of chemical equipment.

SUBSTANCE: invention relates to method for recovery of sludge containing platinum-group metals from equipment using platinum metal-based catalysts. Method includes treatment with aqueous solution of active chemical agent (e.g. sodium-ammonium-substituted ethylenediaminetetraacetic salts) while controlling pH value and removing sludge retained on treated surface with diluted aqueous solution of mineral salts or mixture thereof. pH value is adjusted at 2-10, preferably at 3-9 by adding of organic acid selected from group containing citric, oxalic, maleic, phthalic, adipic, glutaric, succinic acids or basic agents selected from sodium hydroxide, potassium hydroxide, sodium carbonate or potassium carbonate, and hydrochloric acid, sulfuric acid or phosphoric acid is used as mineral acid.

EFFECT: recovery platinum-group metal with improved yield.

4 cl, 1 tbl, 12 ex

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