Gold-cobalt-arsenic concentrate treatment method (2)

Note: The content of the raw calcined residue in the decomposition roasting is: 21.98S total; Sso 4 is not; 3.67As; the slag yield is 64.8%. The amount of sodium sulfate added was 10% by weight of the slag.

Cinder decomposed firing should be carried out in sulfuric acid roasting lined with clay brick muffle furnace. Sodium sulfate (Na 2 SO 4 ·10H 2 O) was added to the original product in an amount of 10% by weight of the slag. Sodium sulfate was added during firing can improve the non-ferrous metal and the degree of sulfation speed, and increasing the presence of sulfates does not decompose in the temperature range state.

The leaching of the sulphated calcined slag (grinding fineness of 95% - 0.074 mm) is carried out in two stages: starting with hot water and then leaching with a 10% sulfuric acid solution at a temperature of 60 to 70 °C. The leaching time per section is 2 hours. Mineral slurry: solid = 4:1. The reason for the need for two-stage leaching is that the presence of cobalt in the slag is different (sulfate, oxide, arsenate , etc.).

* The temperature of the I stage roasting is 450 ~ 500 ° C;

The temperature of the second stage roasting is 600 ~ 650 ° C

** 7.5% NaOH; temperature is 90 ° C; time is 24 hours.

The degree of sulfation of non-ferrous metals and the relationship between their entry into solution and the temperature and time of calcination, as well as the effect of the temperature of sulphation roasting on the degree of desulfurization have been studied (see Table 1). The results of the study prove that the temperature of the sulphation of the slag after decomposition roasting should not exceed 600 °C. The recovery rates of drill, copper and nickel in the combined solution (water and sulfuric acid solution) were 86.3, 87.6 and 59.9%, respectively, at a calcination temperature of 500 to 600 ° C for 3 hours.

Increasing the temperature of the sulphation roasting to 700 ° C results in decomposition of the sulphate and reduces the recovery of the drill and copper into the solution. However, under the optimal sulfation roasting system, the contents of cobalt, copper and nickel in the aqueous solution state were 51.6; 63 and 42% respectively, and the metal content in the oxide state in the sulfuric acid solution was 34.7% cobalt, respectively, 24.6. % copper and 17.4% nickel. The amount of iron dissolved in the sulfuric acid solution does not exceed 3% of the original iron content in the concentrate.

Since the concentrate has undergone two stages of calcination, the total desulfurization rate is 79.8%, which is only 6.81% in the decomposition roasting. In this case, the arsenic recovered from the concentrate was 98.6%, which was 93.22% at the time of decomposition roasting.

When the same calcination temperature, the most favorable sulfating agent consumption (H=SO 4 is 400 kg/ton) and the same amount of sodium sulfate added, the slag after oxidizing roasting is subjected to sulfation treatment, After two stages of leaching, the recovery of non-ferrous metals into the combined solution was 81.4% for cobalt and 85.3% for copper. This figure shows that the recovery of these metals into the solution is 4.9% and 1.3% lower than that of the calcined calcined calcined calcined, respectively.

When the calcination of the calcined slag is sulphated, if the amount of sodium sulphate is reduced from 10% to 5%, the recovery rate of non-ferrous metals transferred into the combined solution will decrease, and the cobalt will decrease by 8.7%. Decrease by 3%. When leaching the slag after calcination, the liquid-solid ratio of the slurry is increased from 4:1 to 10:1, which has no effect on the recovery rate of various metals.

The recovery of various metals from the combined sulphate solution (with a ratio of 1:1) was achieved using ammonia in accordance with known hydration procedures. The

The combined solution contained 2.17 g/L cobalt, 2.06 g/L copper, 9.5 mg/L nickel; 0.53 g/L iron, 1.04 g/L arsenic; pH=1.8. At the same time, it also obtained qualified sales products: copper products containing Cul 5.6%, recovery rate 82.6% 'cobalt products containing C024.3%, recovery rate 80.5%, precious metal concentrates, containing As0.29%, transferred to precious metal fines The gold and silver recovery rates in the mine were 98.6% and 89.1%, respectively.

After precipitating As, Fe, Cu, and Co, the sulfate solution contained: (mg/L) 1.2Cu; 17.1 Co; 2.0Fe; 0.023 As. The sulfate-ion and non-ferrous metals must be removed before they are discarded. To do this, it is first necessary to use a lime milk solution and then treat it with a sodium sulfide solution. When treated with sodium sulfide, the temperature should be 50 to 60 ° C, and it must be stirred for 2 to 3 hours. The amount of sodium sulfide should be 1.5 times or more the amount required to produce a non-ferrous metal sulfide calculated by a stoichiometric method.

The recovery of precious metals from various products obtained from the treatment of concentrates by cyanidation has been studied (Fig. 2). The cyanidation conditions were as follows: 50 g of the original product was weighed; the concentration of the reagent in the slurry was 0.1% NaCN, 0.02% CaO, 200 g/ton PbO; liquid: solid = 3:11 cyanidation time was 36 hours.

The data in Table 2 shows that the recovery of precious metals is low when the original concentrate and its various products are cyanidated. The reason is that the fine-grained dispersion gold in the slag of the oxidative roasting and the sulphation roasting, and the leaching slag of the sulphated roasting slag is present in the hematite granules in a very fine color, so that the gold granules are difficult to cyanide. The solution is in contact. When the decomposition calcination is carried out, the fine particle-dispersed gold is in the particles of the pyrrhotite, so that the cyanide solution cannot enter. One possible method of gold dissociation in hematite is the dissolution of hematite by inorganic acids. But this method is not economically viable.

According to the results of multiple tests and sulphide slag slag components (containing 9.98% Si 2 ; 5.28% CaO; 2.43% Al 2 O 3 ; 0.67% Mg; 80.0% Fe 2 O 3 ; 0.3% C; 0.23% Cu; 0.29% As) is considered to be the most suitable for the recovery of precious metals in non-ferrous metal smelters. The recommended process for the treatment of gold-cobalt-arsenic concentrate by pyrometallurgy-hydrometallurgical process is shown in Figure 2.

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