Recovery of gold, silver, copper and zinc from pyrite cinder by hydrometallurgical method

I. Introduction

Mineral Iberian pyrite containing 44% S and about 4% Cu + Zn + Pb plurality of metal sulfide ore. The main mineral pyrite, and a small amount of associated mineral brass. Sphalerite and galena, pyrite and arsenic, gold and silver. An economically important feature is the presence of minerals in a fine-grained symbiosis. Due to these characteristics, and their relatively high levels of base metals, it is difficult to treat them by flotation. However, sulfur in ore is an economic resource for the production of sulfuric acid, and the remaining base metals and precious metals in the residue or residue are also important potential resources.

Various processes, including pretreatment of the slag as a blast furnace feedstock, are typically based on chlorination roasting [eg Longmaid-Henderon (DKH) process or chlorination volatilization). The combination of the Mostedison Kowa-Seiko process forms the basis of the AIPSA process for recovering all valuable components from Spanish pyrite cinder. However, these pyrometallurgical methods are difficult to implement because they require a large amount of investment.

At present, in view of the crisis in the world steel industry, low-cost hydrometallurgical methods are in a favorable position for achieving more satisfactory economic effects. According to this situation, a method for recovering precious metals and base metals contained in the slag was developed on a laboratory scale and used for the design of a semi-industrial plant.

Second, the characteristics of burning slag

The slag obtained from a typical pyrite (Tharsis pyrite) in Spain contains an average of 1.4 g/t Au, 35 g/t Ag, 0.95% Cu, 2.8% Zn and 61% Fe. Other important features such as particle size, sulfur, arsenic and lead content, etc., depend on the calcination process. The three types of roasters used and their main features are summarized as follows:

(1) Multi-hearth furnace (Herreshoff)

The particle size range of the slag is 0 to 6 mm.

It can remove arsenic and lead to the utmost extent, and the arsenic and lead in the slag can be removed to As~0.08% and Pb~0.3%.

The sulphur content in the slag is very high, S ~ 4%.

Investment, maintenance and wages are high.

(2) Boiling roaster, Boliden-RT

The particle size range of the slag is 90%-0.5 mm.

It can remove arsenic very well, and lead can also be volatilized very well. The slag contains As~0.05%. Pb ~ 0.6%.

The sulfur content in the slag is low, S ~ 1.5%.

Investment costs are not too high, and maintenance and wages are relatively low.

(3) Boiling roaster, completely roasting

The particle size range of the slag is 90%-0.4 mm.

It is impossible to remove arsenic and lead, and the slag contains As~0.5% and Pb~1.3%.

The sulphur content in the slag is low, S < 1.5%.

The investment cost is the lowest, and the maintenance and salary costs are low.

In-depth mineralogical studies of various types of slag have found that their mineralogical composition is basically similar, but there are some variations in the number of calcinations used.

The slag is composed of a hematite aggregate and a microporous aggregate of magnetite as a second phase. Due to the magnetite containing, the sulphur content of the multi-hearth furnace slag is high. In this slag, pyrrhotite exists as a core of larger particles. In contrast, in the boiling calcined sample, there was only a trace amount of Fe _{1 -x} S.

It has been found that sulfate, oxide and iron sulfate solid solutions exist as oxidation products of colored metals. An important feature of fully calcined slag is the enrichment of ferrite, especially zinc ferrite (60% of total zinc).

In addition, during the calcination process, preferential oxidation of iron sulfide occurs, and as a result, a large amount of non-ferrous metals, particularly copper (50% of the total amount), can correspond to certain regions of the Cu-Zn-Fe-S system. The sulfide phase remains in the final slag. The local formation of these residual sulfides has a large effect on the methods described herein and also plays an important role in the leaching stage.

Third, the basic principles of the method

The method proposed in this paper is mainly based on chlorine water leaching, including the following simultaneous chemical reactions:

(1) Decomposition of sulfide residue:

MeS+4Cl ₂ +4H ₂ O→Me ^{2 +} +SO ₄ ^{2 -} +8Cl ^- +8H ⁺ (1)

(b) the dissolution of H ⁺ in the metal oxide produced in (1);

MeO+2H ⁺ →Me ^{2 +} +H ₂ O (2)

(3) The complex form reaction carried out by Cl ^- produced in (1):

Me ^{m +} +nCl ^- →(MeCl _n ) ^m-n (3)

(4) Dissolution of precious metals:

Au+3/2Cl ₂ +Cl ^- →AuCl ₄ ^- (4)

An important feature of the process is that the reaction of the base metal with the noble metal is selective with respect to the iron oxide, and the iron oxide is practically insoluble.

Fourth, the description of the method

The main steps of the method are shown in Figure 1.

Figure 1 Schematic diagram of the process

Step I. Grinding

Cl ₂ leaching depends on the dissociation of the reaction phase (sulfide). The results of the leaching test show that grinding to 100% - 90 microns (50% - 25 microns) ensures that these sulfides are fully dissociated. Although the slag has a microporous structure, the bismuth metal sulfide is still fine-grained, and the surface is covered with a dense iron oxide film which is not reactive.

The raw ore is ground with a final CaCl ₂ lean solution of pH 9.5 at a solid/liquid 1:1 (step VII), and the brittleness and porosity of the slag make the grinding easy.

Step II. Leaching

The ground slag is pulverized with the return solution of the step III, and chlorine gas is introduced into the produced slurry at normal temperature and normal pressure (solid/liquid is 1:2). Under these conditions, the decomposition of the residual sulfide is so rapid that the entire process is controlled by the rate of the gas Cl ₂ absorbed in the solution.

The consumption of Cl ₂ is related to the sulfur content in the form of sulfide in the slag, and can be calculated according to the reaction formula (1). In the CaCl ₂ medium, the generated gypsum remains in the leach residue. For boiling furnace slag (where S ≤ 1.5%), chlorine consumption ≤ 50 grams of Cl ₂ / kg of slag. In multi-hearth furnace slag, the residual sulfur content is high (~4% S). If these slags are directly treated, the consumption of chlorine is high, which is obviously uneconomical.

The recovery rate of precious metals is higher (80% to 90%), while the extraction rate of Cu and Zn is up to 80%, depending on the content of these base metals in the ferrite solid solution, and the ferrite solid solution is chlorine at room temperature. It is insoluble when it is processed.

When leached with Cl ₂ -CaCl ₂ at room temperature, the iron oxide does not react, but iron (<3% total iron) in the form of chalcopyrite, porphyrite and iron sphalerite solid solution dissolves. In addition, arsenic in the completely calcined slag is 50% leached.

Laboratory-scale leaching tests have been conducted and have been reported in the literature. Typical leaching results for the fully calcined sample (CROS. SA) are listed in Figures 2, 3 and 4.

Fig.2 Typical test curve of leaching pyrite cinder by Cl ₂ gas in CaCl ₂ medium

(300 g of slag / 600 ml solution, 25 ~ 30 ° C, 300 rpm, excess gas flow 1 g Cl / min)

Figure 3 Leaching of silver, copper, zinc and iron

Figure 4 Leaching of lead, arsenic and other minor elements

Step III. Precipitation in the slurry

After leaching, the resulting acid [formula (1)] is neutralized with pulverized limestone to control the pH between 2.5 and 3. During this operation, iron precipitates with more than 90% of As and Sb, and 50% of Bi. The CO ₂ produced in this step is capable of removing dissolved chlorine and allowing most of the lead which has been dissolved [possibly in the form of a Pb(IV) chloride) to precipitate. Typical solution analysis results are listed in Table 1.

Table 1 Typical analysis of the solution (CaCl2 approx. 200 g/l, sample: CROS SA)

Step IV. Filtration and washing

The leached slag is filtered together with the precipitated impurities, and washed with the CaCl ₂ lean solution obtained in the step VII. Since the neutralization is carried out at 25 to 35 ° C, the iron precipitates as Fe(OH) ₃ . However, the precipitation process is carried out in the presence of leached cinder, so the filtration performance of the precipitate is good. The washing liquid is returned to the leaching process.

Washed tailings containing 15% to 20% lean CaCl ₂ solution, this impregnation itself (Kowa-Seiko method) in a loop, CaCl ₂ content stabilized at about 200 g / liter, but also helps The leaching residue is treated by a chloride volatilization method, for example, as a pyrite cinder for iron making, which needs to be pretreated. Laboratory tests have shown that the filter cake is calcined to 1250 ° C. For arsenic-free Boliden-RT slag, it can produce slag suitable for ironmaking (63% Fe, 0.03% Cu, 0.08% Zn, 0.01% Pb). , 0.03% As).

Step V. Replacement precipitation

After the rich liquid obtained in the step IV is treated with iron filings, valuable metal products such as gold, silver and copper are obtained (Table 2).

Table 2 Typical analysis of solids (sample: CROS SA)

* The unit is gram / ton.

Step VI. Precipitation of zinc

The solution obtained upon displacement of the precipitate was adjusted to a pH of 9.5 with lime to precipitate zinc together with all the heavy metals present in the solution (Table 1).

Step VII. Filtration and washing

The zinc precipitate was filtered and washed with water. A portion of this lean liquid is returned to step IV after adjusting the pH, and the remainder is returned to the grinding process along with the washing solution.

Zinc can be recovered by industrial methods of solvent extraction and Zn electrowinning, or they can be used as raw materials for the production of high-grade zinc oxide. These projects will be studied in semi-industrial trials.

Step VIII. pH adjustment

The lean liquid used to wash the tailings, the pH is adjusted to 3-4 with HCl, and a small amount of chlorine or hypochlorite is added to maintain a high redox potential, thereby preventing the loss of gold during the washing process.

Five, semi-industrial trials

Based on the results of laboratory tests and batch tests conducted at the University of Barcelona and the Tharsis mine, a preliminary economic evaluation of this method has been carried out in order to determine the use of pyrite as raw material for the treatment of 230,000 tons of burnt slag per year (ie annual The economic benefits of industrial equipment producing 430,000 tons of sulfuric acid.

The post-sales revenue of copper, gold and silver recovered from typical Tharsis pyrite cinders was $29.27/ton. After studying the treatment of zinc, it will eventually increase the income of 7.3 US dollars / ton.

Production costs, including $4 per ton of slag, are estimated at $20.27 per ton, with chlorine being approximately 38%.

Using previous wet metallurgical engineering and cost studies, it is estimated that the investment in building such a plant in Spain is $3.34 million, but does not include post-treatment of the main zinc deposits (no zinc deposits at this stage) The assignment is evaluated).

Under current tax conditions, the return on investment is about 50%. Together with the consideration of the valuable components recovered in the 800,000 tons of slag obtained from the Huelva plant each year, it is proved that it is appropriate to construct a 1 ton/hour semi-industrial pilot plant in Tharsis (Huelva).

At the time of this writing, the pilot plant is under construction and is expected to go into production in the fall of 1984.

Conclusion

Using the method studied in this paper, it is possible to directly utilize the boiling roasting slag of pyrite as a raw material for recovering gold, silver, copper and zinc at a very low cost hydrometallurgical treatment. Selective leaching is carried out under normal pressure, all operations are carried out at room temperature, and the recovery rate of precious metals is also high.

The leached iron oxide can be removed under appropriate conditions (impregnated with a calcium chloride solution), and if the pyrite slag is economically treated, it can be treated by a chlorination method (Kowa-Seiko method).

Based on the economic evaluation of the laboratory scale study of each unit operation, it has been decided to build a semi-industrial pilot plant that processes one ton of pyrite cinder per hour.

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