RU1786158C - Method of recovering noble and non-ferrous metals from sulfide ores and ore processing waste - Google Patents

Abstract

Usage: relates to the processing of sulfide ores, stale and current tailings of their enrichment with the extraction of precious and non-ferrous metals. Essence: The starting material is treated with a solution containing sodium chloride and sulfuric acid to obtain a solution containing non-ferrous metals. Leaching cakes are ground to a particle size of 80-85% of the 0.074 mm class. Sulphide flotation of the crushed product is carried out to recover 9% gold and 87% silver in the foam product. 3 tab.

Description

The invention relates to hydrometallurgy of precious and non-ferrous metals and can be used in the processing of sulfide ores, stale and current tailings of their enrichment, overburden and dump rocks of deposits of precious and non-ferrous metals.

A known method of flotation extraction of gold and silver from stale tailings (content of 1.2 and 7 g / t, respectively; copper and zinc content of 0.03%; 0.16% and 0.62%, respectively) after grinding to a particle size of 80% class - 0.074 mm. The recovery of gold and silver into the foam product by this method is 38.5 and 39.4%, respectively (recovery without preliminary grinding, 5.4% and 8.9%).

The disadvantage of this method is the low recovery of precious metals and the loss of oxidized forms of non-ferrous metals.

A known method for recovering gold from the tailings of gold recovery plants, including flotation and cyanidation of a chamber product, recovery by this method is 70% at an initial grade of 0.6 g / t.

The disadvantage of this method is the relatively low recovery of gold, as well as the use of a toxic reagent (cyanide).

Closest to the claimed method is developed for stale tailings containing 1.2 g / t gold; 7.0 g / t silver; 0.03% copper; 0.16% lead and 0.62% zinc. The method includes grinding the original tailings to a particle size of 85% of the class - 0.074 mm; cyanidation of crushed tails at, 5; duration 24 hours; the content of KCN in the initial solution of 0.02%; CaO content in the pulp 0.02; sulfide flotation of the cake of the first cyanidation without purging at

ate

with

VJ oo oh

ate

00

, duration 10 minutes, butyl xanthate consumption 120 g / t and T-80 blower 40-45 g / t; cyanidation of the sulfide flotation chamber product under the same conditions as the first cyanidation.

The method allows recovery of gold by 89.4% and silver by 66.0%. Its disadvantages are the relatively low silver recovery; loss of oxidized forms of lead, zinc, copper, which are not recoverable by this method; multi-stage technological scheme (one operation for grinding the initial tailings, two for cyanide leaching, three for filtering cyanide pulps and foam flotation product, one for flotation, one for processing solutions); the bulkiness of the hardware scheme (with a conditional installation capacity of 250 thousand tons of tailings per year, the capacity of the installed equipment for cyanide leaching should be 2500 m3); use of toxic reagent (potassium cyanide).

The aim of the invention is to increase the recovery of silver, the utilization of oxidized forms of non-ferrous metals, the simplification of the hardware-technological scheme and the reduction of toxicity of the chemicals used.

This goal is achieved by treating the initial stale tailings with enrichment with a HaSO-NaCI solution, grinding the treated tailings, subjecting them to sulfide flotation, removing non-ferrous metals from the solution after H2S04-NaCI, processing and reusing this solution at the processing stage.

Original tails containing,%: copper 0.05; lead 0.31; zinc 0.50; gold 1.6 g / t; silver 19.6 g / t, and having a particle size distribution: 25% of the class - 0.28 mm; 65% of the class - 0.28-0.10 mm, 2% of the class - 0.10-0.074 mm; 8% class -. 0.074 mm, treated with a solution of HaSCM-NaCI at

With H2 SO4

С№С | ISH

t °, C

g

out

40-50 g / dm 60-80 g / dm3 room 50-60 min 1: 3

stirring in an agitator.

During processing, the oxide films of lead, zinc and copper dissolve on the surface of sulfide minerals

Cu (Pb, Zn) (OH) 2 + H2S04 -

(Pb, Zn) S04 + 2H20 (1) Cu (Pb, Zn) C03 + H2S04 -

(Pb, Zn) S04 + H20 + C02 (2)

0 5 0

fifty

5

0 5

0

5

Pb (Zn) SCM + nNaCI -

- Nan-2PbCln2 n + Na2S04 (3)

As a result, the oxide forms of copper, lead, and zinc pass into solution in the form of chloride complexes of lead and zinc and copper sulfate. The surface of sulfide minerals is cleaned for flotation. With subsequent grinding, the number of sulfide particles available for flotation increases. The treatment of the tailings with hteSCM-NaCI solution after grinding can dissolve the oxide films of copper, lead and zinc by reactions (1) - (3) to the same extent as during processing before grinding, but with a higher consumption of sulfuric acid due to the opening of a larger areas of neutralizing acid alkaline agents such as calcite CaCO3 and siderite ReCO3.

Grinding of the processed tailings is carried out to a particle size of 80-85% of the class - 0.074 mm. Subsequent sulfide flotation allows the recovery of about 9% gold (as in the prototype) and 87% silver into the foam product. The recovery of silver in the foam product increases compared with the prototype by 21%. In the flotation of tailings subjected to grinding to a particle size of 80% class - 0.074 mm, but not treated either before or after grinding with HaSCM-NaCI solutions, the extraction of gold and silver into the foam product, as already noted, is 38.5 and 39.4 % respectively (1), which is significantly lower than by the proposed method. This indicates the removal of films from sulfide minerals with hteSCM-NaCI solutions not only from the surface of the particles of the initial tails, but also inside them (system of channels and microcracks).

Simplification of the hardware-technological scheme during the transition from the prototype to the proposed technology results in a decrease in the number of leaching operations on the entire mass of tailings from two to one, the number of filtrations under the same conditions from three to two, as well as a decrease in the amount of necessary equipment for leaching from 2500 to 100 m (with a conditional production capacity of 250 thousand tons of tailings per year).

The replacement of potassium cyanide as a leaching agent in the prototype scheme with sulfuric acid and NaCI in the proposed scheme dramatically improves the working conditions of personnel and reduces the risk of environmental pollution, since the mentioned substitute reagents are either non-toxic at all (NaCI) or significantly less dangerous. than KCN (H2S04).

Examples of the proposed method.

EXAMPLE 1. For the experiment, a test of stale tailings of enrichment was used, containing,%: copper 0.05; lead 0.31; zinc 0.50; gold 1.6 g / t; silver 19.6 g / t.

The granulometric composition of the sample is characterized by the following content of fineness classes: +0.28 mm grade 25%; class - 0.28 - +0.10 mm 65%; class - 0.10 - +0.074 mm 2% and class - 0.074 mm 8%. A sample of tailings weighing 500 g was loaded into the solution, washed with a solution with an initial content of H2S04 of 50 g / dm and NaCI of 80 g / dm3 with pHNax -0.05 to TIxx: 1: 3 and stirred with a stirrer for 60 min at room temperature the solution at the end of the treatment had a pH of 0.15. The solid phase after processing was filtered off, washed with water and subjected to wet grinding in a ball mill to a particle size of 80% of the class - 0.074 mm. The crushed solid was separated by filtration, repulped with water until a ratio was reached, soda was added in an amount of 2 kg per 1 ton of solid (up to pH 6.2), and sulfides were floated in an open cycle in the presence of butyl xanthogenate (120 g / t) and T blowing agent -80 (25 g / t) for 5 min. The foam product was not subjected to cleanings, valuable components were recovered from the chamber product, carried out control flotation at a consumption of butyl xanthate 80 g / t, T-80 foaming agent 15 g / t, duration 8 minutes. The pH of the pulp at the main flotation stage (after the introduction of flotation reagents) was 7.85. Foam and chamber products, filtrate and washings of the HaSC-NaCI treatment were subjected to chemical analysis. Table 1 illustrates the chemical composition of the final products of the experiment and the distribution of valuable components between them. Table 1 does not show data on industrial discharges, since they are taken into account in the filtrate (the contents of Cu, Pb, and Zn were recalculated taking into account the quantities transferred to the washes).

The chamber product obtained in the tested mode consists of tailings to be stored or disposed of. The foam product can be processed by connecting to a collective concentrate during the processing of polymetallic ores with the conversion of copper, lead and zinc sulfides to the same concentrates and the distribution of noble metals in them, or by connecting to the processing of refractory gold-containing sulfide ores and concentrates according to the gravity-cyanide scheme, and Cu , Pb, Zn-containing processing residues during the processing of polymetallic lead-copper-zinc ores.

Lead and zinc from tailings HteSC-NaCI solutions can be extracted by a known sorption method (sorption by LMP anion exchange resin in the Cl form with desorption by water and precipitation of lead and zinc from eluates by soda (3), and copper by carburizing on metallic iron.

PRI me R 2. The experiment was carried out on the same initial tails and by the same procedure as in the case of example 1, but with the same

the difference is that tailings were not ground after, but before they were treated with H2SCM-NaCI solution. This led to a change in some intermediate parameters (pH after treatment of tailings

HaSCM-NaCI solution was 0.85, pH at the stage of sulfide flotation 6.5) and to a certain distribution of valuable components in the final products of the experiment, Table 2.

An analysis of the data in Tables 1 and 2 shows that the variant with grinding of the tailings after their treatment with H2SCM-NaCI solution is preferable in comparison with the variant with grinding of tailings

because it provides less loss of lead, gold and silver. Loss of lead with dump chamber product in the version with grinding after processing amounted to 21.3%, and with grinding before processing at

the same consumption of sulfuric acid - 37.4%. The corresponding figures for gold are 10.2 and 19.2%; for silver, 12.6 and 15.7%. The tail finder option after processing deserves preference.

also because it requires less consumption of sulfuric acid (60 instead of 120 kg / t in the version with grinding before treatment with a solution of H2S04-NaCI).

Based on the conclusion received, everything

Further experiments were carried out in the tail finisher variant after treatment. In order to optimize the remaining parameters of the proposed scheme in further experiments, we monitored the behavior

only gold and silver. The experiments were carried out on the same initial tails as in the case of examples 1 and 2.

Examples 3-11. The experiments were carried out according to the same procedure as in the case of Example 1, but the contents of H2S04 and NaCI in the initial solution were varied before treatment of the tails and the duration of treatment (for the range of variations, see Table 3). The results are shown in table.3.

According to them, the optimal composition is h SCM-NaCI solution, in which C n25o4xx-40-50 g / dm3

  60-80 g / dm3.

The optimal duration of tail treatment with HaSCM-NaCI solutions is 50-60 minutes.

According to the data of Tables 1 and 3, the selected optimal mode is characterized by the following indicators.

1. The recovery of gold and silver is 89-90 and 86-87%, respectively. Noble metals are transferred exclusively to sulfide flotation concentrate.

2. Copper, lead and zinc are recovered in a sulfide flotation concentrate (52-57%) and in solution after treatment of the tailings (26-29%); total recovery reaches 79-86%.

A comparison of the above indicators of the proposed method and the prototype method shows that both methods provide approximately the same gold recovery (89-90%). The extraction of silver by the proposed method (86-87%) exceeds the similar indicator for the prototype (66%) by 20-21%, the extraction of copper, lead and zinc by 26-29% due to the utilization of their oxide forms.

PRI me R 12. An experiment was carried out, the procedure of which was completely similar to that described in example 1. As a feedstock, sulfide ore was used in it, containing,%: copper 0.45; lead 1.35; zinc 6.60; gold 0.8 g / t, silver 8 g / t, and according to particle size distribution

composition containing 80% of the class - 0.074 mm. The results obtained are similar in nature to those given in Table 1: gold and silver were extracted into the sulfide product on

90 and 85.7%; copper, lead and zinc - by 93% in total, including 77-83% in sulfide flocconcentrate and 12-17% in HaSCM-NaCI solution. The content of valuable components in the tails of the control flotation was,%: copper 0.07; lead 0.12; zinc 0.13; gold 0.08 g / t; silver 1 g / t.

PRI me R 13. An experiment was conducted, the procedure of which was completely similar to that described in example 1, but on the dump rock mass of the deposit containing 0.78% copper; 0.20% lead; 0.45% zinc; 0.5 g / t gold; 5 g / t silver with a fineness of 20% of the class +0.35 mm; 30% of the class - 0.35- + 0.28 mm; 45% of the class - 0.28 +0, 10 mm; 5% of the class - 0.10- +0.074 mm; 10% of the class - 0.074 mm.

The results obtained are similar to those given in table 1.

Claims (1)

The claims A method of extracting precious and non-ferrous metals from sulfide ores and waste from their processing, including grinding, processing with chemical reagents and flotation, characterized in that, in order to To increase the degree of silver recovery, utilization of oxidized forms of non-ferrous metals, simplify the process and reduce its toxicity, before grinding, an aqueous solution containing 60–80 g / dm sodium chloride and 40–50 g / dm3 sulfuric acid are treated for 50-60 minutes Data on the chemical composition of the final products of the experiment in example 1 and on the distribution of valuable components between them Note. The iron content of the solution after treatment was 2.1 mg / dm3 T a 6 l and 1 Data on the chemical composition of the final products of the experiment in example 2 and on the distribution of valuable components between them Note, The iron content in the final solution is different 12, | mg / dm3 The experimental results in the case of examples 3-11 Ta5litsa2 Table