Literature on the beneficiation of gold ores. Literature on mineral processing. Technological scheme for processing raw materials

The enrichment process is unified system, in which individual elements are interconnected. High results can only be achieved by taking into account systematic approach, which takes into account the interaction of system elements, that is, in this case, a full range of processes.

Gravity enrichment is undoubtedly one of the most known processes. It is to him that history owes the fact that gold was the first metal with which humanity became acquainted several thousand years BC. Nature itself took care of this, freeing gold deposits from the minerals that contained them in the beds of rivers and streams flowing through gold-bearing rocks, giving them such an attractiveness that our distant ancestors could not help but pay attention to. Mass extraction of gold from placers began with gravitational enrichment methods, after which these methods actively “stepped” into the factory technology of processing ores from primary deposits. Currently, gravitational concentration of gold is widely used at gold recovery factories (GRPs) in all countries of the world, including those that are the main producers of this metal.

In 2005, we analyzed the performance of more than two hundred gold processing plants from most gold-mining countries, including Russia and other republics of the former USSR. Based on the nature of the processed raw materials, these factories are divided into 3 groups.

Group I includes enterprises that extract gold and accompanying silver from relatively technologically simple quartz and quartz-sulfide ores containing noble metals mainly in cyanide-soluble form.

Group II includes mills that process cyanide-resistant pyrite and arsenic-pyrite ores with finely disseminated gold in sulfides, as well as ores containing sorption-active carbonaceous matter.

Finally, group III consists of enterprises for processing complex ores containing, along with gold and silver, heavy non-ferrous metals (copper, lead, zinc, antimony), as well as uranium.

Within each group, the number of enterprises using the processes of gravity, flotation enrichment and cyanidation is determined (Tables 1, 2).

Table 1. Extent of application of gravity, flotation and cyanidation at the mill

Table 2. Gravity enrichment of ores at the mill

Despite the fact that the list of gold processing plants presented in the tables is far from complete, it nevertheless fairly objectively reflects modern tendencies production of gold from ores of primary deposits and the role played by each of the above technological processes, including gravity.

From the tabular data it is clear that gravity enrichment is practiced by more than 1/3 of the analyzed processing plants, but gravity is almost never used without combination with other processes.

IN last years, great progress has been made in the technology of gravity enrichment of gold ore raw materials. This is manifested, first of all, in the creation of new devices capable of extracting not only large, but also very small particles of metallic gold, released during the grinding of ore.

Such devices, in particular, include centrifugal concentrators (Nelson, Falcon, Knudsen, from Russian analogues- Itomak) and centrifugal jigging machines (Kelsey, Russian), in which the intensity of separation of gold particles and other minerals with a lower grain density increases many times. Previously used gravitational devices have also been significantly improved: conventional jig machines with vertical pulsation, multi-deck concentration tables, screw separators, cone-type devices, etc. Optimal combinations of various gravitational devices have been determined to ensure maximum gold recovery with minimal operating costs. Developed and implemented in industrial scale new methods for processing gravity concentrates, including hydrometallurgical ones, based on the use of cyanide process.

It should be noted that in previous times, cyanidation of gravity concentrates containing large particles of gold and other heavy minerals (in particular sulfides) in tank-type devices (mechanical and pneumomechanical agitators) was considered unacceptable due to the low rate of dissolution of gold and the difficulties of maintaining the suspension in suspension. condition, which resulted in the settling of heavy fractions at the bottom of the apparatus. Currently, these problems are being solved through the use of horizontal drum mixers of the Gekko type, as well as devices with forced circulation of cyanide solutions of the Akatsiya type and Russian cone reactors designed by Irgiredmet. These devices allow cyanidation of gold-containing gravity concentrates with almost any granulometric characteristics. Thus, the traditional technology of gravitational concentration of gold with deep finishing of primary concentrates to rich “gold heads”, suitable for smelting into a gold-silver alloy (Dore metal), is complemented by an alternative method of hydrometallurgical processing of concentrates with moderate metal content, after their one- or two-time refining — concentration tables or other finishing devices. This makes it possible to achieve higher gold recovery rates in the gravity process. The effectiveness of this option increases even more if not only gravity concentrates are subjected to cyanidation, but also tailings from gravity enrichment of ore (using a “softer” leaching mode), since in this case it is possible to direct the solid residues of the “concentrate” cycle into the general hydrometallurgical process with ultimately obtaining a single commercial product.

It is important to emphasize that, in contrast to the processing of gold placers, gravity enrichment of ores from primary deposits is extremely rarely used as the only technological process.

Of the 239 gold processing plants presented in " Analytical review", only one uses this technology: the factory of the Sistine Tu Van Mine mine (USA, California). This enterprise, which is one of the oldest in the country and has been operating since 1896, processes rich quartz ores with native gold. Over 100 years of operation, more than 1 million ounces of gold (31.1 tons) were produced at the mine with a daily productivity of less than 120 tons of ore using “pure” gravity technology. Until 1997, the processing of ore mined underground was carried out according to a scheme including: coarse crushing in jaw and short-cone crushers to a particle size of minus 12 mm, grinding in a ball mill (operating in a closed cycle with a jigger and a mechanical classifier) ​​and gravitational concentration of gold from the classifier discharge on screw separators. The separator tailings were sent to the dump, and the concentrate was sent to finishing (concentration tables). The “golden head” obtained during finishing was subjected to smelting into Doré metal. The finishing middlings were returned to screw separators. Due to the increased losses of free gold from gravity concentration tailings (constituting about 30% of the initial metal content), new gravity devices were tested at the mill. Based on the test results, a decision was made to install at the mill outlet a Nelson KS-CD-20 centrifugal concentrator with a solids capacity of 14 t/h, equipped on top with a stationary screen with a 0.83 mm spat sieve. The screen product was returned back to the mill by means of a pump; the concentrate was subjected to smelting, and the tailings of the centrifugal concentrator were subjected to additional enrichment on Deister concentration tables. The reconstruction of the mill made it possible to increase the extraction of free gold from 70 to 90% while simultaneously increasing ore productivity by almost 3 times.

Experience shows that “purely” gravity enrichment schemes for gold ore raw materials can be applied to small-scale enterprises developing so-called “small” deposits. The construction of developed complexes with combined beneficiation and metallurgical technology, providing higher extraction of gold from ore, in small deposits seems to be economically infeasible, since a small amount of gold cannot pay for the construction of an expensive factory.

In a number of cases, as the experience of some Russian gold processing plants shows (Samartinskaya in Buryatia, Pervenets, Golets Vysochaishy in the Irkutsk region), it is advisable to use the gravity process with reduced gold recovery as the first stage of development of the enterprise, with the goal of accumulating funds for the subsequent expansion of production and transition to more complex technology. Subsequently, rich tailings from the first stages of field development can be successfully processed using, for example, heap leaching.

In the vast majority of cases (77 out of 78 enterprises presented in Table 2), gravity is used in combination with cyanidation, flotation, or both of these processes. For technologically simple ores (group I), the most typical schemes are gravity and gravity-flotation enrichment with cyanidation of flotation tailings, and in some cases, gravity concentrates. The main purpose of gravity in these options is to remove large free gold from the ore into products (concentrates) that are processed in a metallurgical cycle separate from the bulk of the ore.

In addition to increasing (usually 2-4% of total gold recovery), this makes it possible to prevent or at least significantly reduce the accumulation of gold in grinding and mixing apparatus.

Literature.

1. Gold extraction factories of the world: Analytical review/JSC “Irgiredmet”, V.V. Lodeyshchikov.-Irkutsk, 2005.-447 p.

2. Tsarkov V.A. Experience of gold mining enterprises in the world.-M.: Publishing House “Ore and Metals”, 2004.-112 p.

LITERATURE

1. Lodeyshchikov V.V. Technology of extraction of gold and silver from refractory ores: In 2 volumes. - Irkutsk: JSC “Irgiredmet”, 1999.- 775 p.

2.Innovation in Gold and Silver Recovety: Phase IY (Randol. -Colorado: Randol Intern.Ltd.1992. - Vol.5. Chart 17: Flotation.-P.2389-2898

Information about the raw material base of gold ores is presented; technological features of the material composition are considered various types refractory ores; the properties of mineral formations and associations of gold are described. The characteristics of ore beneficiation processes and apparatus, the main technological methods and methods of gold extraction using gravitational, magnetoelectric, flotation, hydrochemical and chemical-metallurgical methods are given. The features of domestic and foreign practices and gold enrichment schemes are noted; examples of hydrochemical technology of gold ores and materials are given; technological features of hydrometallurgical processing of pyrite, pyrrhotite, antimony, telluride, polymetallic, copper-zinc, clayey, carbonaceous and other ores and materials are highlighted. Physicochemical methods for leaching gold from refractory ores and methods for its extraction from solutions are considered; the technology of smelting gold-containing concentrates and products is described; technological schemes of gold extraction factories of Russian and foreign enterprises are given.
For scientists, specialists of gold mining enterprises, processing plants, chemical and metallurgical shops processing gold-containing raw materials, it may be useful for graduate students, students, higher education teachers, engineers, bachelors, masters.

Preface
1. Characteristics of the material composition of ores
1.1. Mineral composition
1.2. Technological classification
1.3. Development of gold extraction technology
1.4. Application noble metals
1.5. Requirements for the quality of raw materials
2. Physico-chemical characteristics of noble metals
2.1. Features of the material composition of ores
2.2. Physicochemical characteristics gold
2.3. Dissolving gold in solutions of chemical compounds
2.3.1. Gold sustainability criteria
2.3.2. Dissolution in acids
2.3.3. Dissolving gold in solutions of chlorine, iodine and bromine
2.3.4. Thiocarbamide and thiosulfate dissolution of gold
2.3.5. Cyanidation
2.3.6. Recovery (precipitation) of gold from leaching solutions
2.3.7. Technological properties that determine the choice of gold enrichment methods and devices
3. Preparation of mineral raw materials for enrichment
3.1. Revealing gold
3.2. Ore crushing
3.3. Ore grinding
3.4. Ore preparation and pulp conditioning
3.5. Physico-chemical methods for separating minerals
3.6. Characteristics of the main technological processes of gold enrichment
3.6.1. Gravitational processes
3.6.2. Magnetoelectric methods
3.6.3. Gold flotation
4. Gravity-flotation schemes for gold extraction
4.1. Basic ore preparation processes
4.2. Practice of gravity-flotation enrichment schemes

4.3. Features of gravity-flotation gold extraction schemes
5. Hydrometallurgical processes for processing ores and products
5.1. Acid leaching
5.2. Leaching in solutions of chlorine, iodine, bromine
5.3. Thiourea, thiosulfate and sulfite leaching
5.4. Dissolution of gold in sulfur-alkaline solutions
5.5. Leaching gold with alkaline cyanide solutions
5.6. Leaching by percolation
5.7. Heap leaching
5.8. Leaching by stirring
5.9. Sorptive leaching
5.10. Autoclave leaching
5.11. Biochemical leaching
5.12. Methods for precipitation of gold from solutions
5.13. Chemical enrichment practice
5.13.1. Processing of pyrite-containing ores and products
5.13.2. Processing of pyrrhotite-containing ores and products
5.13.3. Processing of antimony products
5.13.4. Processing of sulfur ores
5.13.5. Enrichment of clay ores
5.13.6. Enrichment of ferrous metal ores
5.14. Magnetic-electric pulse treatment
5.15. Electrochemical processing
5.16. Energy processing of accelerated electrons and electromagnetic pulses with high magnetic field strength
6. Pyrometallurgical methods for processing concentrates
6.1. Oxidative firing
6.2. Chlorination firing
6.3. Alkaline and sulfidizing firing
6.4. Vacuum firing
6.5. Smelting of copper concentrates
6.6. Smelting of zinc concentrates

6.7. Smelting of lead concentrates
6.8. Smelting of antimony products and sulphide concentrates
6.9. Smelting of concentrates using metal doré alloy
6.10. Fundamentals of refining processing of gold-containing materials
6.11. Chlorination and electrolytic refining of gold and silver-gold alloys
7. Neutralization of enrichment wastewater and leaching solutions of gold and non-ferrous metals
7.1. Treatment of waste cyanide water
7.2. Treatment of waste electrolytes and wastewater
7.3. Sampling and control at enterprises processing gold ores
Bibliography

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Methods for determining grindability
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Technology of conditioning and selective flotation of non-ferrous metal ores (V.A. Bocharov, M.Ya. Ryskin)
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Shinkorenko S.F. - Handbook on beneficiation of ferrous metal ores 1980
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A typical ore processing process is clearly divided into 3 technological stages:

• A) Mechanical ore enrichment (gravity, flotation, radiometric or magnetic separation, etc.), the purpose of which is to obtain products enriched in the content of a valuable component - concentrates and waste tailings that do not require additional processing. This goal, as a rule, is achieved without the use of processes that disrupt the crystal lattice of minerals, due to which the extracted valuable components are present in concentrates in the same mineral form as in the original ore.
• B) Metallurgical processing of ore concentrates using hydro (leaching of valuable components with aqueous solutions of acids, alkalis, salts) and pyrometallurgical (smelting) operations, the result of which is the production of crude metals.
• C) Refining of crude metals (refining) in order to clean them from foreign impurities and obtain final commercial products that meet market conditions.

The experience of the global gold mining industry indicates that smelting these materials is economically justified only if these materials contain (and in significant quantities) copper, lead, antimony and other metals that can act as an “internal” collector of precious metals during smelting, and besides, they themselves represent a certain industrial value. A reflection of this trend is the current practice of metallurgical processing of copper and other concentrates, in which gold is present as an associated valuable component and is extracted from the concentrates into independent commercial products at the stage of refining the resulting non-ferrous metals.

In principle, the smelting method can also be used to extract gold from certain categories of actual gold ores and concentrates that do not contain other non-ferrous metals. These may primarily include rich gravity concentrates or cinders, for which, along with classical methods of pyrometallurgical processing, the option of brushless smelting directly into rough gold or a gold-silver alloy is of interest. If a gold mining enterprise is located near existing pyrometallurgical plants, the use of gold ores (concentrates) as iron-containing fluxes in copper production also seems quite effective, provided that these ores (concentrates) in their composition satisfy the technical specifications for fluxes.

A special place in the global gold mining industry is occupied by the cyanidation process, based on the ability of metallic gold to dissolve in weak solutions of alkaline cyanides according to the reaction:

2Au + 4NaCN + 1/2O2 + H2O = 2NaAu(CN)2 + 2NaOH

The relative selectivity of the solvent (cyanide), the successful combination of the processes of dissolution and precipitation of noble metals from cyanide solutions (cementation with zinc dust, sorption on ion exchange resins and activated carbons, etc.), the simplicity of the equipment and other advantages of cyanidation make it very effective and productive, providing the possibility of applying this technology not only to mechanical concentration concentrates, but also to ordinary gold ores and even to concentration tailings containing 1-2 g/t gold and below.

Currently, cyanidation is used in the processing of 85% of gold ores in the world.

The advantages of the cyanide gold leaching process include its environmental friendliness.

An analysis of the current state of technology and technology for cyanidation of gold ores (concentrates), which covers the activities of the majority of existing enterprises, showed that the global gold mining industry has a large number of options for technological schemes and the use of cyanide process (Figure 2.2), which together provide a complete cycle of ore processing at place even for technologically resistant ores, with a sufficiently high end-to-end gold recovery.

The classical technology of cyanidation of gold ores (full sludge process) includes the following technological operations:

a) Grinding the ore to a size that ensures the necessary completeness of gold discovery;

b) Mixing crushed ore with alkaline cyanide solutions in agitator apparatus of mechanical, pneumomechanical and pneumatic types;

c) Separation of gold-containing solutions from the solid part of the pulp (discharged into a dump) by thickening and filtration methods;

d) Precipitation of gold from solutions by cementation on zinc dust;

e) Processing of gold-bearing sediments (leaching with acids, roasting, smelting) to obtain rough metallic gold, sent to refineries;

f) Chemical treatment of wastewater and hydrometallurgical process tailings from toxic cyanide compounds.

It is necessary to emphasize once again that all of the above operations by themselves do not provide commercial gold-containing products and, as a rule, play an auxiliary role in ore processing schemes, complementing and intensifying the cyanide technology of metal extraction.

A noticeable depressive effect on gold during cyanidation is exerted by minerals and chemical compounds of copper, the dissolution of which consumes from 2.3 to 3.4 kg of NaCN per 1 kg of copper present in the original ore (Table 1.1). At the same time, most copper-containing minerals do not exhibit reducing properties during cyanidation. At the same time, it has been established that an increase in the concentration of Cu in solutions can cause the formation of secondary chemical films on the surface of gold particles, inhibiting the process of subsequent dissolution of gold. It is assumed that the composition of these films is represented by complex compounds such as AuCu(CN) 2 and simple copper cyanide CuCN.

Table 1.1 - Dissolution reactions of copper minerals in aqueous solutions of sodium cyanide

INTRODUCTION

1. GENERAL PART

1.1 CHARACTERISTICS OF GOLD RAW MATERIALS AND METHODS OF ITS PROCESSING

1.2 TECHNOLOGICAL SCHEME FOR RAW MATERIALS PROCESSING

BRIEF DESCRIPTION OF THE MAIN STAGES OF RAW MATERIAL PROCESSING

PREPARATION OF ORES FOR GOLD AND SILVER EXTRACTION

CRUSHING AND GRINDING OF GOLD ORES

GRAVITY METHODS FOR INDIGENOUS ENRICHMENT

GOLD EXTRACTION BY AMALGAMATION

THICKENING

CYANIDATION OF GOLD ORES

SORPTION FROM PULPS (SORPTION LEACHING)

ELUTION OF GOLD AND SILVER AND REGENERATION OF SATURATED ANION EXCHANGES

1.3 ROLE OF THE PROCESS OF SEPARATION OF GOLD FROM THIOUREA ELUTES IN THE TECHNOLOGICAL SCHEME

2. OVERVIEW OF THE TECHNOLOGICAL PROCESS FOR ISOLATING GOLD FROM THIOUREA ELUTES

2.1 CHARACTERISTICS AND CHEMISTRY OF THE PROCESS

PRECIPITATION OF GOLD FROM THIOUREA SOLUTIONS

3. METALLURGICAL CALCULATIONS

INTRODUCTION

Gold is a yellow metal. It has a face-centered cubic lattice and is distinguished by exceptional malleability and ductility. From gold you can draw a wire with a diameter of 0.001 mm. The thermal and electrical conductivity of the metal is very high: gold is second only to copper and silver.

Physicochemical properties of gold:

Au is in the 1st group,

Atomic mass 197,

Density at 20°C 19.32 g/cm 3

Characteristic oxidation states are +1 and +3,

Normal electrode potentials are +1.88 and +1.5 V,

Melting point 1064.4 °C,

Boiling point 2877°C,

Heat capacity at 25°C 25.5 J/(mol K),

The heat of evaporation is 368 kJ/mol.

A distinctive feature of gold is its tendency to form complexes and the ease of reducing most of its compounds to metal.

Gold is a noble metal. Low chemical activity is an important and characteristic property of this metal. In air, even in the presence of moisture, gold remains virtually unchanged. Even at high temperatures, gold does not interact with hydrogen, oxygen, nitrogen, sulfur and carbon.

Gold combines with halogens, and with bromine the process it's already underway at room temperature, and with fluorine, chlorine and iodine - when heated.

The electrode potential of gold in aqueous solutions is very high:

Au®Au + + , j o ​​= +1.68 V;

Au®Au +3 +3, j o ​​= +1.58 V;

Therefore, gold does not dissolve either in alkalis or in acids such as sulfuric, nitric, hydrochloric, hydrofluoric, or organic.

However, in the presence of strong oxidizing agents, gold can dissolve in some mineral acids. So it dissolves in concentrated sulfuric acid in the presence of periodic acid H 5 IO 6, nitric acid, manganese dioxide, as well as in hot anhydrous selenic acid H 2 SeO 4, which is a very strong oxidizing agent.

Gold easily dissolves in aqua regia, chlorinated hydrochloric acid, and aqueous solutions of alkali and alkaline earth metals in the presence of oxygen. A good solvent for gold is an aqueous solution of thiourea containing ferrous chloride or sulfate (+3) as an oxidizing agent.

Other gold solvents include chlorine and bromine water, a solution of iodine in potassium iodide or hydroiodic acid. In all cases, the dissolution of gold is associated with the formation of complex compounds.

In its chemical compounds, gold can have an oxidation state of +1 and +3. All gold compounds are relatively weak and are easily restored to metal even by simple calcination.

The purpose of the course work is to review technologies for extracting gold from solutions of thiourea eluates, show the advantages and disadvantages of each of them, and also consider in detail the technology of electrolytic deposition of gold from thiourea eluates.

1.GENERAL PART

1.1 CHARACTERISTICS OF GOLD RAW MATERIALS AND METHODS OF ITS PROCESSING

Over the past two to three decades, the share of gold extracted from technologically simple gold ores has been steadily decreasing. At the same time, the share of gold extracted from such ores is increasing, the effective processing of which requires much more complex and developed schemes, including operations of gravity enrichment, flotation, roasting, smelting, leaching, etc. Gold ores and concentrates, the processing of which under conventional cyanide process conditions (in combination with gravitational and amalgamation methods for extracting large gold) does not provide a sufficiently high recovery of gold or is accompanied by increased costs for individual technological operations (crushing, cyanidation, dehydration, precipitation of gold from solutions, etc.), called persistent.

ORES WITH FINE GOLD AND METHODS OF THEIR PROCESSING

Fine dissemination of gold in rock-forming minerals is the most common reason for the persistence of gold ores.

Ores of this type are divided into two main categories: ores in which gold is associated with quartz; ores in which gold is associated with sulfides.

To extract gold from ores of the first category, fine or ultrafine grinding is used, which ensures sufficient exposure of the gold. For this purpose, schemes with three-stage grinding and preliminary classification of the material are used before II and III stages processing. Grinding ore according to this scheme ensures the production of a product with a particle size of 90-95% class - 0.04 mm.

Cyanidation of such finely ground material allows, as a rule, to obtain waste tailings with a low gold content. However, due to the high cost of fine grinding, the processing of finely disseminated gold ores is significantly more expensive compared to the processing of conventional ores. In addition, due to the increased content of secondary silts in the cyanide pulp formed during fine grinding, the productivity of the thickening and filtration cycle is noticeably reduced, which further increases the cost of extracting gold from such ores. As a result, when processing ores with finely disseminated gold, the specific costs for grinding and dewatering can reach 60% of the total cost of ore processing, while when processing ordinary ores they do not exceed 30-40%. In order to reduce the cost of grinding, in recent years, much work has been carried out to introduce a progressive method of ballless grinding (autogenous grinding) of gold ores.

Ores of the second category contain gold in the form of fine and emulsified dissemination in sulfides, mainly pyrite and arsenopyrite. The most common method for extracting gold from such ores is flotation, which allows the extraction of gold-containing sulfides and free gold into a concentrate. The concentrate can then be processed using various methods to extract gold from it.

If the grain size of the gold particles is not too small and allows the gold to be exposed by fine grinding, the flotation concentrate is further crushed and cyanidated.

The use of flotation in this case makes it possible to eliminate the expensive operation of fine grinding of the entire mass of the original ore and limit ourselves to additional grinding of a relatively small amount of flotation concentrate, the yield of which is usually 2-5% of the mass of the original ore.

Often, however, the dissemination of gold in sulfides is so small that fine and even ultrafine grinding of the material does not allow achieving the required degree of exposure. In this case, finely divided gold is exposed using oxidative roasting. During oxidative roasting of flotation concentrates, sulfides are oxidized and converted into a porous mass of oxides that is highly permeable to cyanide solutions. Subsequent leaching of the cinder allows the gold to be converted into a cyanide solution.

The oxidation of pyrite begins at a temperature of 450-500° C. The process proceeds with the formation of pyrrhotite FeS 2 + O 2 = FeS + SO 2 as an intermediate product, which is oxidized to magnetite 3FeS + 5O 2 = Fe 3 O 4 + 3SO 2 and further to hematite 2Fe 3 O 4 + ½ О 2 = ЗFe 2 О 3 .

At temperatures above 600 °C, the oxidation of pyrite is preceded by its dissociation with the formation of pyrrhotite 2FeS 2 = 2FeS + S 2, which is then oxidized to hematite.

The performance of oxidative firing depends on a number of parameters, of which temperature is the most important. If the firing temperature is not high enough (below 500° C), the rate of oxidation reactions is low, and a noticeable amount of incompletely oxidized pyrite particles may be present in the cinder. Cyanidation of such a cinder will be accompanied by significant losses of gold due to its insufficiently complete opening. With increasing firing temperature, pyrite oxidation occurs faster and more completely. However, at temperatures exceeding 900-950 ° C, partial melting of the cinder is possible due to the formation of relatively low-melting eutectic mixtures consisting of pyrrhotite and magnetite. The appearance of a melt leads to sintering of the material and the production of dense, low-porous cinders that are difficult to cyanide.

The oxygen concentration in the gas phase significantly affects the firing performance. At low oxygen concentrations, the rate of pyrite oxidation decreases, which can lead to insufficient gold recovery. At the same time, with an excessively high oxygen concentration, the speed of the process can become so high that, if heat exchange conditions are not good enough, the heat of exothermic reactions will not have time to dissipate in the environment and the temperature of the fired grains will exceed a dangerous limit (900-950 ° C). As a result, the cinder will melt and its structure will not be porous enough. It has been practically established that the optimal firing temperature for pyrite concentrates depends on their material composition and ranges from 500-700 ° C. Calculations and experimental studies show that due to “overheating” of the cinder, its temperature can exceed the temperature in the furnace by 300-400 hail The relationship between the rate of pyrite oxidation and the temperature of its grains indicates that in order to obtain a porous cinder, the rate of oxidation reactions must be regulated so that the temperature of the particles during firing does not exceed 900-950 ° C. To achieve this, it is necessary to reduce the amount of air supplied to oven, or reduce the oxygen concentration in the gas phase. At the same time, it is possible to reduce the “overheating” of the fired particles by improving the conditions of heat exchange between the material and environment. This way is more rational, since it allows you to maintain the optimal temperature of the material in the furnace without a corresponding reduction in the firing speed. The conditions for heat exchange between the fired concentrate and the environment are improved with intensive mixing of the material in the furnace. Therefore, carrying out the firing process on a hearth under conditions of relatively weak mixing of the material creates a significant danger of “overheating” of the cinder and its partial melting. Carrying out the process in fluidized bed furnaces, where due to intense mixing the heat exchange conditions are extremely favorable, makes it possible to maintain the firing temperature much more accurately, preventing the cinder from melting.

The behavior of arsenopyrite during oxidative roasting is in many ways similar to the behavior of pyrite. Intensive oxidation of arsenopyrite begins at a temperature of approximately 450 ° C and proceeds with the formation of pyrrhotite and magnetite as intermediate products:

2FeAsS + 1.5O 2 = 2FeS + As 2 O 3 (gas),

3FeS + 5O 2 = Fe 3 O 4 + 3SO 2,

2Fe 3 O 4 + 0.5O 2 = 3 Fe 2 O 3.

At temperatures above 600° C, the oxidation of arsenopyrite is preceded by its dissociation: 4FeAsS = 4 FeS + As 4 (gas).

Gaseous arsenic is oxidized to trioxide As 4 +3O 2 = 2Аs 2 О 3, and pyrrhotite to hematite.

The resulting arsenic trioxide is highly volatile. At a temperature of 465° C, the vapor pressure As 2 O 3 is 1 am. Therefore, arsenic, oxidized to As 2 O 3 , turns into gas

phase. However, with an excess of oxygen, arsenic trioxide can be oxidized to pentoxide: As 2 O 3 + O 2 == As 2 O 5.

Depending on the firing conditions and the material composition of the fired material, arsenic pentoxide can remain unchanged in the cinder or interact with iron oxides, forming divalent and trivalent iron arsenates Fe 3 (AsO 4)) 2 and FeAsO 4. Since arsenic pentoxide and iron arsenates are practically non-volatile, arsenic, oxidized to the pentavalent state, remains completely in the cinder. The latter circumstance is extremely undesirable, since during subsequent planning of the cinder, arsenic goes into solution and in some cases completely disrupts the precipitation of gold by zinc dust. The recycling of gold-free cyanide solutions becomes practically impossible. In addition, the presence of pentavalent arsenic compounds in the cinder leads to the formation of films on the surface of gold particles, which makes it difficult to dissolve them in a cyanide solution.

In this regard, when roasting concentrates containing arsenopyrite, arsenic must be transferred into the gas phase. For this purpose, roasting of arsenic concentrates should be carried out in a weakly oxidizing atmosphere, which promotes the formation of volatile trioxide and minimizes the oxidation of arsenic to the pentavalent state.

However, the weakly oxidizing atmosphere favorable for arsenic removal does not correspond to the conditions for maximum oxidation of sulfide sulfur, which requires a much more oxidizing atmosphere to remove. In this regard, the most rational way of oxidizing gold-arsenic concentrates is two-stage roasting. The first stage of firing, carried out under conditions of limited air access, is aimed at transferring arsenic in the form of As 2 O 3 into the gas phase. The resulting cinder enters the second stage, where, with a significant excess of oxygen, oxidation of sulfide sulfur occurs. This two-stage firing makes it possible to obtain a porous cinder favorable for cyanidation with a low content of sulfide sulfur and arsenic. Due to its advantages, two-stage roasting is used in gold mining factories that process gold-arsenic concentrates.

Approximately a similar effect can be achieved with single-stage firing, if you use the countercurrent principle, i.e., the movement of the material towards the firing gases. In this case, the initial sulfide concentrate during the first firing period will be in contact with already partially used gases, which therefore have a low oxygen concentration. This circumstance contributes to the fact that arsenic will be removed during the first firing period. As the material moves further in the furnace, it will come into contact with the gas,

increasingly enriched with oxygen, as a result of which, at the exit from the furnace, the cinder will be free not only of arsenic, but also of sulfur. The countercurrent principle is widely used in hearth roasting of gold-bearing sulphide concentrates.

Until 1946, concentrates were fired in all factories without exception in deck furnaces. This type of firing has not lost its importance to this day. In Australia alone, there are several dozen installations that roast concentrates on hearths. Of all existing types of hearth furnaces, Edwards furnaces are most suitable for firing gold-containing concentrates. The Edwards furnace is a mechanized reverberatory furnace with a rectangular cross-section. It consists of a metal casing lined with refractory bricks. With a high sulfur content in the concentrate (above 26%), roasting can proceed autogenously, that is, solely due to the heat released during the oxidation of sulfides. If there is a lack of sulfur, the furnace is heated with coal, fuel oil, gas or wood. For this purpose, one or two fireboxes are located at one end of the furnace. At the other end of the furnace there is a special hole in the roof through which the fired concentrate is loaded. To mix and move the material during firing, one or two rows of rotating paddles are installed along the length of the furnace, driven by a common shaft located above the furnace. The rotation of the paddles ensures repeated movement of the fired material from one wall of the furnace to another and its simultaneous advancement along the furnace. As a result, a sufficient duration of residence of the material in the oven is achieved (3-6 h) and conditions are created for its mixing.

In some cases, Edwards furnaces have special devices for changing the angle of inclination of the furnace, which allows you to adjust the speed of material passing through the furnace when the material composition of the fired concentrate changes. In enterprises with small productivity (up to 7-10 T concentrate per day) use furnaces with one row of strokes; at higher productivity (10-50 t/day) install furnaces with two rows of rows.

The following advantages contribute to the widespread use of Edwards furnaces:

1) minimal dust removal during firing of concentrates, not exceeding 0.5-1.0% of the weight of the loaded material. Low dust entrainment allows you to work without complex dust collection systems. At most gold mining factories that use hearth roasting, gases are purified from dust in cyclones or dust chambers;

2) relative cheapness, simplicity of design and ease of maintenance. Routine furnace repair operations, such as replacing the rake and rake holders, are carried out from outside the kiln without unloading or cooling it. Each furnace is serviced by one operator;

3) the ability to work in a wide range of temperatures and roast concentrates with different granulometric characteristics and different chemical compositions.

However, along with the advantages of the Edwards furnace, like all hearth-type furnaces, they have serious disadvantages, the main of which are the following:

1) relatively low specific productivity of approximately 0.25 m/(m 2 - day);

2) uneven temperature distribution over the mass of the fired material;

3) the difficulty of regulating temperature and oxygen conditions.

These disadvantages of hearth firing served as an impetus for the development of a much more progressive firing method - fluidized bed firing. Currently, fluidized bed roasting is used at gold mining enterprises in Canada, the USA and other countries. In Fig. A diagram of the fluidized bed roasting plant for flotation concentrate at the Dickenson Mines mine (Canada) is shown. The fluidized bed furnace is a vertical steel cylinder with a diameter of 2.5 m and height 5.5 m, lined with refractory bricks. Podina f ovens with an area of ​​3.14 m 2 made of refractory concrete. There are 116 nozzles in the hearth, through which air is supplied from the turbocharger. The concentrate is continuously fed into the furnace and in the form of a pulp using a pump. Entering the furnace, the concentrate particles are set in continuous upward motion.

air currents. The height of the fluidized bed is approximately 1.2 m. The temperature in the furnace is 700" C. The cinder is unloaded continuously through a special unloading pipe located at the level of the fluidized bed on the side opposite to the loading. Upon exiting the furnace, the cinder enters a bath of water, where it is cooled. The roasting gases are cleaned of dust in three sequentially located cyclones , after which it is released into the atmosphere through a chimney. Dust from the cyclones is unloaded into baths filled with water. The pulp, consisting of cinder and dust, is concentrated and sent for planning.

A further improvement in the technology of roasting gold-arsenic concentrates in a fluidized bed is to carry it out in two stages. Two-stage firing can be carried out in two connected kiln chambers or in separate kilns.

At the first stage, with a limited amount of air, arsenic is distilled off from the concentrate in the form of As 2 O 3. The second stage, carried out with excess air, serves to oxidize sulfide sulfur. In Fig. shows a diagram of a two-stage fluidized bed roasting plant at the Campbell Mine (Canada). Flotation concentrate in the form of a pulp with a solid content of 70-80"% is supplied to the first stage of firing. The cinder of the first stage is sent through the unloading pipe to the second stage. For better flow of material, a nozzle for supplying compressed air is installed in the unloading pipe. Gas from the first stage furnace is supplied into the intermediate cyclone and then into the above-layer space of the furnace of the second stage. Dust from the cyclone, together with the cinder, is sent to the second stage. The gas leaving the furnace of the second stage enters two parallel threads of cyclones (three cyclones in each) and is discharged through the chimney into atmosphere.The cinder and dust from stage II cyclones are cooled with water in a special bath and sent for cyanidation.

To carry out autogenous firing, the sulfur content in the fired material should not be less than 16-20%. At higher contents, there is a need to remove excess heat. In practice, this is accomplished by supplying additional water either to the furnace feed or directly to the fluidized bed.

Firing concentrates in fluidized bed furnaces is accompanied by large entrainment of dust (40-50% of the starting material). Therefore, thorough purification of gases from dust is one of the central problems. The use of cyclones alone often does not provide the required degree of gas purification. In these cases, the dust collection system is supplemented with electric precipitators. Some enterprises practice the extraction of arsenic trioxide from gases. For this purpose, the gases leaving the furnace are thoroughly cleaned of dust and cooled;

Condensed arsenic trioxide in the form of a fine powder is collected in bag filters. If necessary, gases from fluidized bed furnaces can be used to produce sulfuric acid.

Compared to hearth furnaces, fluidized bed furnaces are very effective devices for roasting gold-containing concentrates. Their main advantages are the following:

1) high specific productivity, amounting to about 5 t/(m 2 -day), which is approximately 20 times higher than the productivity of deck ovens;

2) higher quality of the resulting cinders, due to the possibility of precise control of the temperature and oxygen firing conditions.

However, along with the advantages, roasting in a fluidized bed has some disadvantages, the main one of which is large dust removal. This circumstance requires the construction of complex dust collection systems.

The considered scheme for processing sulfide gold-containing concentrates by oxidative roasting followed by planning the cinder is a very common, but not the only possible scheme for processing such products.

In some cases, flotation concentrates obtained at gold mining enterprises are sent to copper smelters, where they are smelted together with copper concentrates. In this case, gold goes into matte and is ultimately concentrated in the anode sludge, from where it is extracted using special methods (see p. 282). This method is not suitable for processing flotation concentrates with high arsenic content, since arsenic makes it difficult to produce pure copper. Therefore, gold-arsenic concentrates must be oxidized to remove arsenic before being sent to a copper smelter.

Oxidative roasting can also be used in the processing of arsenic-free pyrite concentrates to produce sulfuric acid.

The method of processing raw or roasted concentrates at copper smelters does not require large expenditures and makes it possible to extract gold even from such refractory materials, for which oxidative roasting followed by cyanidation of the cinder does not give positive results. The disadvantages of this method are increased transportation costs and losses of gold during transportation and smelting of the concentrate.

The method of processing flotation concentrates by oxidative roasting followed by cyanidation of the cinder has known disadvantages. The main one is increased losses of gold from cyanidation tailings. Despite all the measures taken, oxidative firing is inevitably accompanied by partial sintering of the material and the formation of films of low-melting compounds on the surface of the gold. As a result, a certain amount of gold becomes inaccessible to the action of cyanide solutions and is lost with cyanidation tails.

The desire to increase gold recovery from sulfide flotation concentrates has led to the development of a number of other methods: oxidative-chlorinating roasting; chloride sublimation; autoclave leaching.

Oxidative-chlorinating roasting is carried out with the aim of opening finely dispersed gold for subsequent cyanidation. The essence of this type of firing is that the material being processed is mixed with 5-20% sodium chloride and fired in an oxidizing atmosphere at a temperature of 500-600 ° C. The mechanism of the process boils down to the fact that sulfur dioxide and sulfur vapor formed during firing in the presence of oxygen react with sodium chloride, releasing free chlorine:

2NaCl + SO 2 + O 2 = Na 2 SO 4 + Cl 2

2NaCl + S + 2O 2 = Na 2 SO 4 + Cl 2

Possessing high chemical activity, chlorine interacts with iron sulfides and oxides, forming the chlorides FeCl 2 and FeCl 3. The latter are decomposed by atmospheric oxygen:

2 FeCl 3 + 1.5 O 2 = Fe 2 O 3 + 3 Cl 2.

The released free chlorine reacts again, etc. This process mechanism, associated with multiple diffusion of gaseous products through the mass of mineral grains, is the reason for the formation of porous hematite Fe 2 O 3, the structure of which is favorable for access of cyanide solutions even to the deepest and thinnest inclusions of gold. Due to this, when cyanidating the cinder of oxidative-chlorinating roasting, the recovery of gold in solution is higher compared to cyanidating the cinder of simple oxidative roasting. If non-ferrous metals are present in the source material, then during oxidative-chlorinating firing they turn into chlorides. To extract them, as well as to wash water-soluble sodium sulfate, unreacted sodium chloride and small amounts of unreacted iron chlorides, the cinder should be leached with water or a weak acid solution before cyanidation.

As can be seen from the above reactions, a necessary condition for successful oxidation-chlorination firing is the presence of sulfide sulfur in the fired material. At the same time, the high sulfur content in the source material leads to increased consumption of sodium chloride and thereby reduces the economic efficiency of the process. Therefore, before oxidative-chlorinating firing, it is advisable to subject high-sulfur materials to simple oxidative firing to produce cinders containing 3-5% sulfur.

Chloride sublimation, proposed by B. N. Lebedev, as well as oxidative-chlorinating roasting, consists in the fact that the gold-containing concentrate is mixed with sodium chloride and fired in an oxidizing atmosphere. However, unlike oxidative-chlorinating roasting, which is only a preparatory operation for planning, chloride sublimation involves the complete conversion of metallic gold into volatile chloride and its subsequent capture from gases in the form of a very metal-concentrated product. This effect is achieved only at high temperatures, approximately 900-1000 ° C. At the same time as gold, chlorides of silver, copper, lead and other metals are also sublimated. The mechanism of chloride sublimation is basically similar to the mechanism of oxidation-chlorination roasting.

To avoid sintering of sulfide concentrates at high temperatures, pre-fired materials with a sulfur content of 2-5% should be subjected to chloride sublimation. Lower sulfur content is also undesirable due to reduced gold recovery. The optimal consumption of NaCl is 10-15% by weight of the starting material. With a lack of NaCI, gold and its accompanying elements are not completely chlorinated and are partially lost with the cinder; Excess NaCI leads to melting and enlargement of cinder particles, which also impairs the extraction of metals. If these conditions are met, up to 99% Au, 98% Ag, 96% Cu, 90% Zn are transferred into sublimates. The gold content in cinders does not exceed 2 g/t.

Processing of sublimates involves leaching them with water and converting chloride salts of arsenic, iron, copper, lead, zinc, as well as sodium sulfate and chloride into a solution. In this case, gold is reduced to metal and, together with silver chloride, remains in an insoluble residue. The total content of precious metals in the residue after water leaching is several percent, which allows it to be directly smelted into crude metal. The chloride solution can be used to extract non-ferrous metals.

The chloride sublimation process is very versatile; it can be used to extract gold from concentrates of almost any composition. An important advantage of this process is the possibility of complex processing of concentrates with the extraction from them not only of gold and silver, but also of accompanying non-ferrous metals. The disadvantages of chloride sublimation include the complexity of the equipment for high-temperature firing and the capture of sublimates. For this reason, chloride sublimation has not yet found application in the gold mining industry.

Autoclave leaching of gold-containing concentrates with gold finely disseminated in sulfide minerals consists of their hydrometallurgical processing at elevated temperatures (100-200 ° C) and oxygen pressures (I-20 am). Autoclave technology for extracting finely disseminated gold can be carried out in two ways.

The first option involves opening thin gold for the purpose of its subsequent extraction using cyanidation. As studies by I. N. Maslenitsky, I. N. Plaksin, S. V. Khryashchev and others have shown, the release of gold associated with sulfides can be achieved by autoclave leaching of concentrates in water, solutions of sulfuric acid or caustic soda.

During the autoclave oxidation of pyrite and arsenopyrite concentrates in water and dilute solutions of sulfuric acid, the following chemical reactions occur:

2FeS 2 + 7.5O 2 + 4H 2 O = Fe 2 O 3 + 4H 2 SO 4

FeAsS + 3.5 O 2 + H 2 O – FeAsO 4 + H 2 SO 4

Due to the decomposition of sulfides and their transformation into a porous mass of iron oxide and iron arsenate, easily permeable to cyanide solutions, solid residues after autoclave leaching are a favorable product for the recovery of gold by cyanidation.

When leaching pyrite and arsenopyrite concentrates in caustic soda solution, the nature of the flow chemical reactions other:

2FeS 2 + 8NaOH + 7.5 O 2 = = Fe 2 O 3 + 4Na 2 SO 4 + 4H 2 O

2FeAsS + 10NaOH + 7O 2 = 2Na 3 AsO 4 + 2Na 2 SO 4 + = Fe 2 O 3 + 5H 2 O

In this case, not only sulfide sulfur passes into the solution, but also arsenic. This facilitates the subsequent cyanidation of the residues and the precipitation of gold by zinc dust. Caustic soda can be regenerated from autoclave leaching solutions by treating them with lime:

2Na 3 AsO 4 + 3Ca(OH) 2 =Ca 3 (AsO 4) 2 + 6NaOH

The insoluble calcium arsenate obtained along the way can be used in the chemical and woodworking industries.

According to the second option, studied by S.I. Sobol, I.N. Plaksin and other researchers, autoclave leaching of gold-containing concentrates is carried out in such a way that simultaneously with the opening of finely dispersed gold, its dissolution occurs. In this case, ammonia solutions are used to leaching gold-containing concentrates. The chemistry of the reactions occurring in this case is quite complex. In a simplified form, it boils down to the fact that during the autoclave oxidation of sulfides in ammonia solutions, a number of soluble sulfur compounds are formed, including the thiosulfate ion S 2 O 3 2-.

As already indicated, the S 2 O 3 2- ion forms a strong complex with gold, as a result of which the potential of gold shifts to the negative side and its oxidation by oxygen becomes possible. Therefore, during ammonia leaching of sulfide concentrates, not only gold is exposed, but also it goes into solution in the form of the Au(S 2 O 3) 3-2 anion. The main, still completely unresolved problem of processing refractory gold-containing concentrates by this method is the difficulty of extracting gold from ammonia solutions with a complex composition. The most promising in this regard is the use of ion exchange resins and activated carbon.

Research work shows that in some cases, autoclave technology for processing gold-containing concentrates allows one to achieve higher gold recovery compared to the oxidative roasting method. In addition, the use of autoclave technology eliminates the loss of gold through dust, eliminates the need to construct complex dust collection systems, and can significantly improve the working conditions of operating personnel. Currently, the autoclave leaching method has not yet found application in the gold mining industry. The main reason for this is the relatively high cost of high-pressure equipment.

1.2 TECHNOLOGICAL SCHEME FOR RAW MATERIALS PROCESSING

PREPARATION OF ORES FOR GOLD AND SILVER EXTRACTION

Currently, gold and silver are extracted from primary ores either using hydrometallurgical processes, or using combined schemes, in which enrichment techniques of various methods play a major role. Thus, the mined ore is presented in large pieces up to 500 mm, and sometimes larger, then it is first crushed and crushed.

CRUSHING AND GRINDING OF GOLD ORES

The purpose of these operations is the complete or partial opening of grains of gold-containing minerals, mainly particles of native gold, and bringing the ore into a state that ensures the successful completion of subsequent enrichment and hydrometallurgical processes. Crushing and especially fine grinding operations are energy-intensive and their costs account for a significant share (from 40 to 60%). Therefore, it must be borne in mind that grinding should always be completed at the stage when the noble metals are sufficiently exposed for their final extraction or for their intermediate concentration. Since the main method of extracting gold and silver for most ores is hydrometallurgical operations, the required degree of grinding should ensure the possibility of contact of solutions with open grains of gold and silver minerals. The sufficiency of exposure of these minerals for a given ore is usually determined by preliminary laboratory process tests for the extraction of precious metals. To do this, ore samples are subjected to technological processing after varying degrees of grinding while simultaneously determining the recovery of gold and accompanying silver. It is clear that the finer the inclusion of gold, the deeper the grinding should be carried out. For coarser gold ores, coarse grinding (90% grade -0.4) is usually sufficient mm). However, since most ores contain small gold along with large gold, That most often ores are crushed more finely (before-0,074 mm). In some cases it is necessary to subject it to even finer grinding (up to -0.043 mm).

However, the economically feasible degree of grinding is determined by a combination of a number of factors:

1, the degree of metal extraction from ore;

2, increased consumption of reagents with more intensive grinding;

3, the cost of additional grinding when bringing the ore to a given size;

4, an increase in sludge with finer grinding and the associated additional costs for dewatering operations (thickening, filtration).

Crushing and grinding schemes vary depending on the material composition of the ores and their physical properties. Usually. The ore is first subjected to coarse and medium crushing in jaw and cone crushers with test screening. Sometimes a third stage of fine crushing is used, carried out in short-cone grinders. After two-stage crushing, the material usually obtained is 20 mm, after the three-stage process, the material size sometimes decreases to -6 mm. The crushed material is fed to wet grinding, which is often carried out in ball and rod mills. Ores are usually crushed in two stages, with rod crushers being preferred for the first stage.

GRAVITY METHODS FOR ENRICHMENT OF BIG GOLD ORES

When extracting gold from bedrock ores, gravity enrichment is currently of significant importance.

The overwhelming majority of gold-bearing ores contain some amount of coarse gold (+0.246 mm and larger), which is poorly recovered not only by flotation concentration, but also during hydrometallurgical processing. Therefore, its preliminary separation by gravity enrichment at the beginning of the technological process can eliminate unnecessary losses of gold. In addition, the extraction of free gold at the beginning of the gold ore processing process makes it possible to quickly sell this part of the metal and reduce the loss of undissolved and unwashed gold.

Jigging machines are the most widely used primary gravitational device that catches gold at the drain of mills.

GOLD EXTRACTION BY AMALGAMATION

Application of amalgamation in schemes of gold recovery factories

In world practice, the amalgamation process has been widely used to extract gold from ores. Currently, it is rarely used, this is caused, firstly, by the constant change in the quality of gold-containing ores, as a result of which the content of gold associated with sulfides, having coating formations, as well as low-grade gold, i.e., forms that are not extracted by amalgamation; secondly, amalgamation is a labor-intensive process, always accompanied by losses of gold in the form of amalgam, which is not recovered in subsequent stages of the technological process; thirdly, due to the strong toxicity of mercury vapor, the use of large volumes creates a danger of mercury poisoning of people and the environment.

However, amalgamation has retained its importance for the recovery of free gold from gravity concentrates obtained from the processing of bedrock and placer ores. In this case, a small amount of rich material has to be processed, and the amalgamation process retains its main advantage - the cheap and quick sale of gold in metal form. This method, in particular, is used to process the bulk of gravity concentrates in South Africa.

THICKENING

Thickening is the next stage of pulp processing after grinding. It consists of partial dewatering of the pulp by settling - settling of solid particles to the bottom of the thickener vat and draining the clarified solution. In most cases, about 50% (by weight) water remains in the settled material. which corresponds to the ratio w:t = 1: 1. The concentration limit depends on the size, density and physicochemical properties of the crushed particles of the ore being processed.

The particles contained in the pulp usually vary greatly in size. Along with relatively large granular particles (over 0.1 mm) the pulp usually contains a significant amount of particles several microns in size and even smaller (less than 0.001 mm). Larger particles settle faster, while smaller ones remain suspended for a long time.

CYANIDATION OF GOLD ORES

The methods of gravity enrichment and amalgamation discussed above make it possible to extract only relatively large gold from ores. However, the vast majority of gold-bearing ores, along with large gold, contain a significant and sometimes predominant amount of fine gold, which is practically unrecoverable by these methods, as a result of which the tailings of gravity enrichment and amalgamation, as a rule, contain a significant amount of gold, represented by small gold particles. The main method for extracting fine gold is the cyanidation process.

The essence of this process is the leaching of precious metals using dilute solutions of cyanide salts of alkali or alkaline earth metals. The resulting gold-containing solutions are separated from the solid phase (waste tailings) by thickening or filtration and sent for the precipitation of noble metals with zinc metal. The precipitate of precious metals, after appropriate processing, is sent for refining to obtain pure gold and silver.

SORPTION FROM PULPS (SORPTION LEACHING)

A feature of the sorption process from pulps is a slightly lower process speed due to the increased viscosity of the pulps (at l: t = 1...2: 1) and the deposition of sludge covers on the surface of the ion exchanger particles, which impede the diffusion of ions. In addition, one should take into account the inevitability of increased losses of ion exchanger due to the destruction of its grains under the abrasive action of ore particles. Therefore, sorption from pulps into production conditions should be carried out when the size of ore particles is no more than 0.15 mm. The kinetics of sorption of gold and silver from cyanide pulp indicates that most of gold passes into the ion exchanger phase during the first 2 hours of mixing the pulp. Increasing the duration of contact has little effect due to the approach of the resin-pulp system to equilibrium: after 8 hours, the sorption of gold was only 68.4%. The extraction of silver into the resin phase is significantly reduced: 2% after 2 hours and 28% after 8 hours of contact of the resin with the pulp. Complete sorption of noble metals occurs with a significant increase in the amount of loaded resin.

The possibility of complete sorption of dissolved gold from pulps by anion exchangers has been demonstrated, and a method has been developed for using this process to determine the content of dissolved unwashed gold in filter cakes or waste pulp of the mill (I.D. Fridman). The use of the sorption method makes it possible to extract a larger amount of gold from samples, and therefore obtain greater accuracy of analysis compared to the usually used decantation washing method. Sorption of gold from pulps is used not only for the analysis of mill tailings, but directly in the technological process of cyanidation of ores and concentrates. In the latter case, sorption from the pulp is usually combined with the process of leaching gold and silver from ores, and the process is called “sorption leaching.” The first studies on sorption leaching of gold ores in our country were carried out by I.N. Plaksin and his colleagues. The study of this process was further developed in the works of B.N. Laskorin and his employees, who developed and introduced into production a countercurrent scheme for sorption leaching of gold ores. As a result of research and production work, it was established that sorption leaching leads to a significant acceleration of the process of gold dissolution and a reduction in the duration of cyanidation by 2-3 times. In addition, in some cases, the degree of gold recovery increases, and the loss of undissolved gold from cyanidation tailings is noticeably reduced. During sorption leaching of quartz ore crushed to 95.4% class -0.044 mm

(at l: t = 2: 1), in laboratory conditions, already in the first 4 hours, gold extraction was 85.5%, and in 8 hours it increased to 96.8% (Fig. 8). Under conventional cyanidation conditions, only 61.2% of gold passed into solution in 4 hours, and 96.0% in 24 hours of cyanidation. Thus, when combining the processes of leaching and sorption of dissolved gold, the rate of the cyanidation process increased 3 times (8 hours instead of 24 hours), while the loss of gold from tailings decreased from 1-1.2 to 0.8 g/t. In this case, the speed of the process during sorption leaching increased 3 times, since the maximum gold recovery of 94.9% was achieved in 3.5 hours; with conventional cyanidation, such extraction was obtained in 10.3 hours. The acceleration of the process during sorption leaching is explained by a shift in the equilibrium of the gold dissolution reaction towards the formation of an anion - with a decrease in its concentration in the solution due to sorption by an anion exchanger: 2Аu + 4CN- + ½O 2 + H 2 0= 2 - + 2OH - . Analysis of the kinetics of the process shows that an increase in the concentration gradient of the anion accelerates its diffusion removal from the reaction zone and the dissolution process as a whole. The increase in the rate of dissolution of gold and silver is also influenced by the removal of ions of accompanying base metals as a result of their sorption from the solution by the resin.

ELUTION OF GOLD AND SILVER AND REGENERATION OF SATURATED ANION EXCHANGES

In the process of sorption of noble metals from cyanide solutions and pulps, saturated anion exchangers are obtained containing sorbed complex cyanide anions of gold, silver and base metals and non-metallic anions - SCN -, CN -, OH -, etc. Saturated anion exchangers undergo a regeneration process for the purpose of desorption of sorbed anions and restoration of their sorption activity for recycling use in the sorption process. Desorption of sorbed compounds from the resin is carried out by elution (washing out) with solutions of appropriate reagents, and it is advisable to selectively extract gold and silver into a concentrated solution with their subsequent production in the form of a commercial product. Other desorbed components should also be used as fully as possible: copper, cyanide, etc. Desorption - . Testing for desorption of gold a number of conventional elution solutions - sodium chloride, ammonium chloride, hydrochloric and sulfuric acids, sodium and ammonium hydroxyl, sodium carbonate, sodium cyanide, etc. - turned out to be ineffective: gold is recovered only partially and non-selectively. English researchers have established the possibility of successful elution of the anion - with a number of organic solvents in a mixture with mineral acids, such as methyl or ethyl alcohol + 5-10% HC1 + 5% H 2 O, acetone + 5% HCl, ethyl acetate + 10% HNO 3 + 5 % H 2 O, etc. The best results were obtained using mixtures: acetone + 5% H 2 O or acetone + 5% HNO 3 + 5% H 2 O. When using a mixture of acetone with HCl, complete displacement of gold and copper is achieved, while while iron, zinc and silver elute in smaller quantities. The use of a mixture of acetone with HNO 3 gives complete, almost selective, extraction of gold; iron and copper are much less eluted by this solution. The elution of gold cyanide by this method is explained by the formation of a gold-containing complex with an organic solvent in the presence of a mineral acid, a covalent complex that is not retained by the anion exchanger. The method using organic solvents was tested at a large pilot plant in Rhodesia, but did not find industrial application. Its main disadvantages are its high cost, the flammability of organic reagents and the large volume of elution solution. A number of domestic and foreign studies have established that effective elution of the anion from anion exchangers is achieved by solutions of thiocyanate salts - KSCN, NaSCN, preferably NH 4 SCN, which contains more groups SCN per unit mass. For more complete and rapid desorption of gold, it is recommended to use concentrated alkaline solutions of NH 4 SCN - 3-5 N. (228-380 g/l) with NaOH content from 10 to 25 g/l. The process of gold desorption occurs via an anion exchange reaction:

RAu(CN) 2 +SCN - =RSCN+ -

The resin then transforms into rhodanium form. Elution curve of gold from AM5 resin n. solution of NH 4 SCN shows that for a sufficiently complete extraction of gold, 14 volumes of solution per 1 volume of resin are required, but most of the gold can be obtained in the first 6-8 volumes of a more concentrated solution. In addition to gold, thiocyanate alkalis desorb complex cyanides of silver, copper, nickel, cobalt and iron, free cyanide ions and hydroxyl ions. Rhodane compounds do not desorb zinc cyanide compounds, but the latter are extracted by alkali, usually present in the rhodanide solution. The main disadvantage of thiocyanate salts as eluent reagents is the transition of the resin to the thiocyanate form. The use of resin in this form during sorption is impractical both technologically (reducing the capacity of the resin for noble metals) and economically(losses with expensive reagent tails). As a result, there is a need to desorption of the thiocyanate ion from the resin and converting it into another form. However, the desorption of thiocyanate ion causes significant difficulties: a large volume of elution solutions is required - 15 or more volumes of 1-2 N per 1 volume of resin. solution of NaCI or NH 4 NO 3 The result is a large volume of solutions with a low content of thiocyanate, the regeneration of which has not been developed. Due to the noted disadvantages, the use of thiocyanate salts as desorbents under industrial conditions encounters significant difficulties. The most effective desorbent of dicyanaurate ion is weakly acidic solutions of thiourea (thiocarbamide). The elution ability of thiourea is explained by its high polarizability and complexation. When interacting with - in an acidic environment, it displaces the cyanide ion and binds gold through a pair of free electrons of sulfur into a cationic complex (according to Reynolds) Au 2, which is not able to be retained by an anion exchanger having positively charged ionogenic groups. The resin then turns into chlorine or sulfate form, and the released CN ions are bound into HCN. The process of gold elution proceeds according to the reaction:

RAu(CN) 2 +2SC(NH 2) 2 +2HCl=RCl+Au 2 Cl+2HCN.

A similar reaction occurs in the case of sulfuric acid, the use of which is preferable in practice. The completeness of gold elution increases with increasing concentration of thiourea (TM) in the solution up to a saturated concentration of 9.1%. When the concentration of HCl changes, a maximum is observed in the elution curve, corresponding to concentrations of 1.9-2.3% HCl. With a further increase in the concentration of HCl to 10%, thiourea decomposes with the release of elemental sulfur. In practice, solutions with a HM concentration of 90 g/l + 20-25 g/l sulfuric or hydrochloric acid are used for gold desorption.

The elution process extends to 10 volumes of solution per 1 volume of resin, but most of the gold is concentrated in the first 4-6 volumes of a solution that is richer than NH 4 SCN. The first 1-2 volumes of the solution have a low gold content, which is associated with the absorption of HM by the resin in the initial period, which can reach 10% of the anion exchanger mass. This circumstance has been used in practice to extract heavy metals from excess circulating solutions. In addition to the dicyanaurate ion, weakly acidic HM solutions desorb silver, copper, nickel, and much worse - zinc and iron; cobalt is almost not extracted. Acidic solutions of HMs desorb gold much more completely and quickly with an increase in temperature to 50-60 °C. Application more high temperature impractical due to the thermal instability of anion exchangers. Silver is desorbed faster than gold and mainly goes into the first fractions of the eluate. Selective elution of one silver is possible when more weak solution containing 8-10 g/l SC(NH) 2 and 2-3 g/l HCl or H 2 SO 4. Copper cyanide compounds are also well desorbed by HM solutions. However, with a high copper content in the resin (more than 5 mg/g), due to the formation of a precipitate of simple copper cyanide CuCN in the resin in an acidic environment, the elution process is lengthened, and the extraction of gold may not be complete enough. Nickel, usually contained in a saturated anion exchanger in a small amount (no more than 2-3 mg/g), is desorbed by the HM solution quite completely. In this regard, in order to avoid complications of the process and contamination of the gold- and silver-containing eluate with copper and nickel, it is advisable to first desorb them from the anion exchanger.

Desorption - . As noted above, the most effective desorbent of silver cyanide compounds is a weakly acidic solution of HM containing 8-10 g/l HM + 2-2.5 g/l HC1 or H 2 SO 4 . When interacting with silver, TM forms a cationic complex with the composition Ag 3 Cl + 2HCN.

According to the principle of ion exchange, silver cyanide anions are well desorbed by solutions of 75-225 g/l NH^SCN + 10-20 g/l NaOH according to the reaction:
RAg(CN) 2 +SCN-=RSCN+ - .

By the same principle, silver cyanides are quite completely desorbed by solutions:

a) 250 g/l NH^NO^; b) 100 g/l NaCN; c) to a lesser extent - with a solution of 150-200 g/l NaCl. Silver is almost not desorbed by NaOH solutions.

Desorption n-1. Copper cyanide anions are effectively desorbed by weakly acidic HM solutions of the same concentration as for gold, as well as by an alkaline solution of ammonium thiocyanate with a concentration of 50-75 g/l NH 4 SCN + 10-20 g/l NaOH. Copper anions are well desorbed by a solution of sodium cyanide with a concentration of 40-80 g/l NaCN + 0.1 g/l NaOH at a temperature of 50-60 °C according to the exchange reaction:

2R 2 Cu(CH) 3 +2SC-=2RCN+ 2 -

The cationic groups of the anion exchanger are neutralized by cyanide ions. The extraction of copper into the solution at a consumption of up to 10 volumes per 1 volume of resin is 70-90%. When treating resin containing complex copper cyanides with solutions of mineral acids (2% sulfuric or hydrochloric acid), the complexes decompose according to the reactions:

· *2RCu(CN) 2 +H 2 SO 4 =*R 2 SO 4 +2CuCN+2HCN;

· *R 2 Cu(CN) 3 +H 2 SO 4 =*R 2 SO 4 +CuCN+2HCN;

· *2R 3 Cu(CN) 4 + 3H 2 SO 4 = *3R 2 SO 4 + 2CuCN + 6HCN.

As a result of the decomposition of the complexes, a precipitate of simple copper cyanide CuCN is formed, which remains in the resin, and, therefore, the copper is not desorbed. At the same time, partial regeneration of cyanide occurs, released in the form of hydrocyanic acid HCN. Cyanide is regenerated from ions: - - 50%; [Cu(CN) 3 ] 2- - 66.6%; 3- - 75.0%. To elute copper, it is necessary to oxidize the copper cyanide precipitate, converting monovalent copper into the form of the Cu 2+ cation, which is not retained by the anion exchanger and leaves with the solution. When Fe 2 (SO 4) 3 (l.3%) is used as an oxidizing agent in a solution of H 2 SO 4, elution occurs according to the reaction

In this case, the cyanide is completely regenerated. If HCl is used, FeCl 3 should be used as an oxidizing agent.

Thus, the use of a solution of sulfuric or hydrochloric acids with an oxidizing agent makes it possible to elute copper with simultaneous complete regeneration of cyanide, which corresponds to the data of I.N. Plaksin and M.A. Kozhukhova obtained during the regeneration of cyanide solutions. The CuCN precipitate can also be extracted from the resin with a strong (40-50 g/l) NaCN solution. In addition to the mentioned reagents, desorption of copper cyanide anions can be carried out with NaCl solutions with a concentration of 150-175 g/l using an exchange reaction. The process proceeds more efficiently when the temperature rises to 50-60 °C.

Desorption 2- . Zinc elutes well from the resin with dilute solutions of H 2 SO 4 with a concentration of 20-25 g/l. In this case, the zinc cyanide complex decomposes, the zinc passes into the form of the Zn 2+ cation, which is not retained by the anion exchanger, and the cyanide is completely regenerated. When using HCl, it is necessary to use diluted solutions with a concentration of 0.1 N. (3-5 g/l) HCl, since at a concentration of 0.5 n. HCl and above, the cyanide complex transforms into the chloride complex ZnCI 4 2-, which is retained by the anion exchanger:

R 2 Zn(CN) 4 +4HCl=R 2 ZnCl 4 +4HCN.

In this case, zinc is removed from the resin upon subsequent washing with water due to the decomposition of the ZnCI 2 anion with the formation of a soluble ZnCI 2 salt, in which zinc is present in the form of the Zn 2+ cation, which is not retained by the anion exchanger.

If ferrocyanide 4- ions are present in the resin, zinc cations in an acidic environment form with them precipitates of Zn 2 Fe(CN) 6 and H 2 ZnFe(CN) 6, which remain in the resin. For this reason, the degree of acid desorption of zinc decreases. Zinc cyanide compounds are effectively eluted by NaOH solutions with a concentration of 40-50 g/l. In this case the following reactions occur:

· *R 2 Zn(CN) 4 +6NaOH=*2ROH+Na 2 Zn(OH) 4 +4NaCN;

· Zn 2 Fe(CN) 6 +4NaOH=2Zn(OH) 2 +Na 4 Fe(CN) 6;

· Zn(OH) 2 +2NaOH=Na 2 Zn(OH) 4 ;

· *4ROH+Na 4 Fe(CN) 6 =*R 4 Fe(CN) 6 +4NaOH.

The resulting zincate ion 2- is highly hydrated and passes into the aqueous phase. The ferrocyanide ion 4- can be partially sorbed by the anion exchanger. Complex zinc cyanide is weakly eluted by solutions of ammonium thiocyanate and sodium chloride, but the addition of NaOH in an amount of 20 g/l to these solutions sharply increases the extraction of zinc from the resin. The tetracyanozincate ion is also successfully desorbed by a solution of 250-400 g/l NH 4 NO 3 + 10 g/l NH 4 OH. Cyanide solutions do not desorb zinc compounds. Desorption 2- . Anions 2-, like copper anions, are actively desorbed by a weakly acidic solution of TM, as well as an alkaline solution of NH 4 SCN. In the latter case, desorption occurs via an exchange reaction:

*R 2 Ni(CN) 4 +2SCH-=*2RSCN+ 2-
Nickel is effectively eluted with dilute solutions of sulfuric (20-25 g/l) or hydrochloric (10-20 g/l) acid according to the reaction:

*R 2 Ni(CN) 4 +2H 2 SO 4 =*R 2 SO 4 +NiSO 4 +HCN.

The process proceeds with the formation of the Ni 2+ cation not retained by the anion exchanger, and complete regeneration of cyanide. A solution of NH 4 NO 3 with a concentration of 250 g/l desorbs nickel partially (about 40%). Solutions of NaCN, NaOH, NaCl practically do not desorb nickel cyanide.

Desorption of ferrocyanide ion 4-. Ferrocyanide ion is effectively desorbed from the anion exchanger by NaCN solutions with a concentration of 50-100 g/l, preferably at temperatures up to 50-60 °C. The process proceeds through an ion exchange reaction:

*R 4 Fe(CN) 6 +4NaCN=*4RCN+Na 4 Fe(CN) 6.

Anion 4- is also well desorbed by solutions of 2-3 N. NaCl (120-180 g/l) containing 0.25-0.5 N. NaOH (10-20 g/l), preferably at a temperature of 50-60 ° C, by exchange reaction with the C1 - ion. The 4- anion is eluted quite completely by solutions of NH 4 SCN with a concentration of 75-225 g/l with the transition of the resin to the thiocyanate form RSCN. After sulfuric acid treatment of the resin for desorption of zinc, nickel and cyanide ion and desorption of gold, silver and copper with a weakly acidic solution of TM, iron and copper residues can be eluted with a solution of the composition: 160 g/l NH 4 NO 3 +50 g/l NH 4 OH + 40 g/l NaOH at 25 °C. The consumption of the elution solution is 7 volumes per 1 volume of resin. After treatment, iron in the resin is in the form of sorbed ferrocyanide ion R 2 Fe(CN) 6 and in the form of precipitation of salts Ni 2 Fe(CN) 6, Zn 2 Fe(CN) 6, etc., insoluble in an acidic environment, copper - in the form of a precipitate of simple cyanide CuCN. In an alkaline environment, ferrocyanide salts with heavy metals decompose to form a precipitate of hydroxides Zn(OH) 2 and Ni(OH) 2 and ferrocyanide ion 4-. Copper cyanide and metal oxide hydrates dissolve in an ammonia solution according to reactions:

· CuCN+NH 4 NO 3 +2NH 4 OH=Cu(NH 3) 2 NO 3 +NH 4 CN+2H 2 O;

· Zn(OH) 2 +2NH 4 NO 3 +2NH 4 OH=Zn(NH 3) 4 (NO 3) 2 +4H 2 O;

· Ni(OH) 2 +2NH 4 NO 3 +3NH 4 OH=Ni(NH 3) 5 (NO 3) 2 +5H 2 O;

The resulting complex cations of copper, zinc and nickel pass into the eluate. Ferrocyanide ion is desorbed from the resin by the NO 3 - ion according to the exchange reaction:

*R 4 Fe(CN) 6 +4NO 3 - =*4RNO 3 + 4- .

Anion 4- is partially (up to 40%) eluted with solutions of 2-4 N. HNO, and is poorly eluted by solutions of H 2 SO 4 and NaOH.

Desorption 2-. Cobalt cyanide anion is usually contained in a saturated anion exchanger in small quantities (no more than 1 mg/g), but its desorption is difficult. 2- is most effectively eluted with a solution of 150-375 g/l NH 4 SCN, partially (up to 30-60%) - with solutions of 225-250 g/l NH 4 NO 3, 180 g/l NaCI + 20 g/l NaOH, 50 -100 g/l NaCN. Cobalt desorption increases with increasing temperature to 50-60 °C. Desorption of cyanide ion CN - . Cyanide ion is regenerated by solutions of sulfuric or hydrochloric acids with a concentration of 10-20 g/l according to the reaction:

*2RCN+H 2 SO 4 =*R 2 SO 4 +2HCN.

Hydrocyanic acid HCN is distilled from gold, absorbed by a solution of NaOH or Ca(OH) 2 and returned to the cyanidation process in the form of alkali cyanide NaCN or Ca(CN) 2.

The cyanide ion CN - is also desorbed by solutions of NH 4 SCN, NH 4 NO 3, NaCI, NaOH, etc., the anions of which replace it in the resin.

Desorption of impurity anions S 2 O 3 2-, SO 3 2-, SiO 3 2-, etc. is successfully carried out with NaOH solutions with a concentration of 40-50 g/l. In the process of regeneration of anion exchangers, it is necessary to achieve the most complete desorption of both noble metals and impurities. The impurities remaining on the resin when it is reused during the sorption process worsen the kinetics of the process, reduce the capacity of the resin for precious metals and increase the loss of dissolved gold in the liquid phase of the tailings. The more impurities remaining in the anion exchanger, the more significant the effect of incomplete regeneration is. As practice shows, the content of residual components in the anion exchanger after regeneration can be: gold - no more than 0.1-0.3 mg/g, impurities - no more than 3-5 mg/g of air-dry sorbent. When the value of residual impurities is more than 10-12 mg/g, a significant increase in the concentration of gold in the solution after sorption is observed, i.e. the loss of dissolved gold with tailings increases.

In Fig. 1. The technological scheme for processing gold-containing raw materials is shown.

Fig.1 TECHNOLOGICAL DIAGRAM FOR PROCESSING GOLD RAW MATERIALS

1.3 ROLE OF THE PROCESS OF SEPARATION OF GOLD FROM THIOUREA ELUTES IN THE TECHNOLOGICAL SCHEME

The extraction of gold from the acidic thiourea eluate is the final processing operation of almost any type of ore, if the thiourea leaching operation of gold was used in the technological scheme. Thiocabamide is the most prepared for industrial use in the processing of gold ores of technological type B.

The effectiveness of this solvent, in particular, can be judged from the experimental results given in Table. 1. Leaching of gold was carried out from ore mixtures composed of gold-bearing quartz (2 samples of ore with an Au content of 12.9 and 2.3 g/t, respectively) and various mineral additives introduced into it. The latter were used: bornite-chalcocite concentrate (Cu content 70, S 20%), ore stibnite (98% Sb 2 S 3), a mixture of realgar and orpiment. containing 53% As and 32% S, as well as OU grade activated carbon. The proportion of introduced additives was 1% by weight of the quartz base. From the table above it can be seen that the thiourea leaching process provides approximately the same gold extraction rates from simple quartz ores as cyanidation. .less sensitive to impurities (copper, antimony, arsenic), which allows us to consider it as one of the possible options for hydrometallurgical processing of ores belonging to technological type “B”.

Table 1

Results of comparative experiments on the extraction of gold from mineral mixtures with cyanide and thiourea solutions

Note: General terms leaching: temperature 20-25 °C, duration 6 hours; F:T=2:1: initial concentration of NaCN and solutions 2.5 g/l; having collected thiourea solutions (g/l): ThiO - 20; Fe 2 (SO 4) 3 (oxidizing agent) - 3.0; H 2 SO 4 (environment regulator) - 5.0.

In the period 1964-1984. Irgiredmet specialists conducted a complex of theoretical and experimental studies to study general patterns the process of dissolving gold and its chemical compounds in acidic solutions of thiocarbamide with the subsequent development of technology for the extraction of metals from ores based on thiocarbamide leaching, including the operations of deposition of gold and silver from solutions and neutralization of technological process wastewater. However, this operation is quite complex, since the process is poorly understood and difficult to implement since the ores are quite poor and it is not economically profitable to use this operation.

2. REVIEW OF THE TECHNOLOGICAL PROCESS FOR ISOLATING GOLD FROM THIOUREA ELUTES

2.1 CHARACTERISTICS AND CHEMISTRY OF THE PROCESS

Gold and silver can be deposited from a HM solution by a number of methods.

For solutions with intermediate gold and silver contents, as studies have shown, the use of the method of cementation of Au and Ag with electronegative metals (zinc, aluminum, lead, iron) seems quite promising.

Metallic lead is more preferable in this regard, allowing for the extraction of 99.5% Au and 99.9% Ag into the cement sludge in 10 minutes. However, the relatively high consumption of cementator (7 parts per 1 part of the sum of Au and Ag) and the associated low content of noble metals in the sediment (less than 12%) require careful assessment of the effectiveness of the cementation process in relation to specific raw materials.

1. Lead cementation gold production is carried out from hydrochloric acid HM eluates with subsequent cupellation of the sediment. When lead shavings or dust are loaded into a solution, the following gold cementation reaction occurs:

2AuCl+Pb=2Au+Pb 4 Cl 2.
The consumption of lead dust is 20-30 g per 1 g of gold. The disadvantages of the method are contamination of the solution with lead, which makes it difficult to use in circulation, and unsuitability for sulfuric acid HM eluates.

2 Zinc cementation. The soda-zinc method for deposition of gold and silver was developed and put into production at the Lebedinsk experimental ion exchange plant. The solution is first neutralized with soda to pH 6-7, then zinc dust is added in an amount of 3-4 g per 1 g of gold and mixed with the solution for 2-3 hours. Gold is deposited according to the reaction:

(Au 2 ) 2 SO 4 +Zn=2Au+Zn 4 SO 4 .
After the precipitation of gold and silver, soda is added to the solution until pH 9-10 for partial precipitation of zinc from the solution in the form of Zn(OH) 2. The filtered zinc sediment contains 6-10% gold and is processed with sulfuric or hydrochloric acid, followed by smelting of the sediment into rough gold. The solution is sent into circulation - for desorption. The disadvantages of the method include: low quality of the sediment, high consumption of reagents (soda, zinc dust, acids) and contamination of the circulating elution solution with zinc and sodium salts, which reduce the quality of the regenerated resin.

3. Aluminum cementation. Aluminum dust with a particle size of 95% class -0.074 mm is used. Precipitation occurs according to the reaction:

3 2 SO 4 +2Al=6Au+12TM+Al 2 (SO 4) 3.
Aluminum consumption is 3 g per 1 g of gold, contact duration is 4 hours at a temperature of 20 ° C. The gold content in the sediment is up to 25%, the residual gold content in the solution is 2-5 mg/l. The sediment is treated with a 5% NaOH solution for 3 hours to remove aluminum, after which the gold content in the sediment increases to 85%. This sediment is melted to obtain rough gold. The disadvantages of the method are the same as with zinc deposition.

4. Alkaline precipitation of gold. According to this method, the eluate is pre-blown with air for 4-6 hours to remove HCN in order to avoid the formation of NaCN upon the introduction of alkali, dissolution and incomplete precipitation of gold. A 40% NaOH solution is added to the solution at a temperature of 40-50 °C to pH 10-12. In this case, gold hydroxide is deposited according to the reaction:

Au 2 Cl + NaOH = Au (OH) + 2SC (NH 2) 2 + NaCl.

The duration of precipitation is 2-4 hours. It is of great importance to maintain the specified pH value of the solution, since with a lack of alkali, incomplete precipitation of gold is possible, and with an excess of it, the dissolution of gold hydrate. Together with gold, hydroxides of impurity metals are deposited, due to which the solution is freed from impurities that complicate the process of resin regeneration. Hydrated sediments with a gold content of 10-15% are filtered on a filter press, washed with hot water, blown with air and fired at a temperature of 300 ° C to remove sulfur. The cinders are then subjected to sulfuric acid treatment to dissolve impurities and, when the gold content is 30-45%, they are sent for refining. The disadvantages of the method are: poor filterability of hydrated sediments, additional acid consumption and relatively low quality of sediments.

5. Gold precipitation activated carbon . The extraction of gold and silver from acidic thiourea solutions with a relatively low concentration of noble metals (Au up to 50 mg/l) can be carried out by the method of sorption on activated carbons (SKT, OU, KAD, etc.).

The capabilities of this technological method can be judged by the results of gold sorption by powdered activated carbon of the SKT brand from solutions from flotation concentrate leaching Artemovskaya gold processing plant . containing, in addition to gold, a certain amount of copper (up to 0.2 g/l), iron (up to 1.2 g/l) and other impurity components. Depending on the initial concentration of gold in solutions (1-20 mg/l), the working capacity of coal for gold was 2-12%, with impurity content (%): iron 0.12, copper up to 2.7; calcium. magnesium, aluminum, zinc, nickel - ranging from 0.01-0.1. After ashing, the coal contained Au 40-50; Fe 5-10: Cu 10-15; SiO 2 5-8%, thus representing. a product sufficiently concentrated in gold, suitable for shipment to refineries.

The results of a study on the mechanism of gold adsorption from acidic thiocarbamide solutions by activated carbon were obtained. It has been established that sorption obeys the Freundlich and Langmuir equations and is an endothermic process. The activation energy of 3.5 kcal/mol indicates that

The limiting stage in the kinetics of adsorption is diffusion in the porous structure of activated carbon.

A sorption-flotation option for extracting gold from unfiltered thiourea pulps using activated carbon has been proposed. But in this option, the pulp is treated with powdered coal (50-60% of particles with a diameter of less than 40 microns), after which it is subjected to flotation. Sodium oleate is used as a flotation reagent. Duration 5 min. During this period, from a pulp (pH = 1.5) containing 20 mg/l gold, 400 mg/l thiocarbamide, 0.5 g/l coal powder, with a sodium oleate consumption of 100-120 mg/l, gold extraction is achieved in concentrate at 90%. Increasing the oleate consumption to 200 mg/l increases gold recovery to 95%.

The best results when testing various grades of coal were obtained when extracting gold with SKT coal. The sorption of gold by coal is quite selective; most of the impurities remain in solution. In a countercurrent process, in 4-5 steps, gold is completely extracted from solutions and coal sediments with a gold content of 15-20% are obtained. They must be burned and the ashes melted into rough metal. Disadvantages of the method: the difficulty of burning coal and large losses of heavy metals (10-15%), which are completely lost during the processing of coal sludge.

6. Gold extraction. Research by scientists has shown the fundamental possibility of extracting gold from hydrochloric acid solutions by extraction with tributyl phosphate. When the organic and aqueous phases come into contact for 5-10 minutes and the phase volume ratio O:B = 2:1, almost complete extraction of gold from the aqueous solution is achieved. Further research is needed on extraction from sulfuric acid eluates, stripping and processing of strippings.

7. Electrolytic deposition of gold and silver. This is the most effective method processing of acidic HM eluates, as it makes it possible to obtain metallic gold of high purity without the cost of reagents and additional operations. Graphite or platinum plates are used as anodes in the electrolytic deposition of gold and silver, and titanium plates or graphite batting are used as cathodes.

Thus, during the electrolysis of gold, the following electrochemical system takes place: Au with impurities (cathode) | 2 SO 4 ,TM,H 2 SO 4 ,H 2 O impurities | Ti (anode)

The following main reduction reactions occur at the cathode:

Au 2 + +e=Au+2SC(NH 2) 2

It is also possible to recover copper, lead and other impurities. At the anode, an electrolytic reaction takes place to form H + ions:

2H 2 O-4e=O 2 +4H +

In addition, processes of oxidation and decomposition of heavy metals with the release of sulfur according to the reaction are possible at the anode:

SC(NH 2) 2 -2e=CNNH 2 +S+2H +

Over time, cyanamide adds water and turns into urea.

CNNH 2 +H 2 O=CO(NH 2) 2

The overall process of anodic oxidation of heavy metals proceeds according to the reaction

SC(NH 2) 2 + H 2 O -2е= CO(NH 2) 2 +S+2H +

The oxidation of heavy metals especially increases with increasing current density > 15 A/m2. Anodic decomposition of heavy metals increases the consumption of this expensive reagent, and its decomposition products have a harmful effect on the electrolytic separation of metals from solution and the quality of the sediment. To avoid this, it is advisable to carry out the electrolysis process with the separation of the anode and cathode spaces by a diaphragm with an ion exchange membrane, i.e. with division of the electrolysis bath into anode and cathode chambers. The gold-containing HM solution is placed in the cathode space, the anolyte is a 0.5-1.0% solution of H 2 SO 4. Anion or cation exchange membranes are used as diaphragms, through which HM molecules do not pass into the anolyte. Complex HM cations pass into anolyte in small quantities: 3-4%. In the presence of diaphragms, the consumption of heavy metals during electrolysis is sharply reduced. The electrolysis process is carried out at a cathode current density Jc = 8...10 A/m2, bath voltage 3.0 V, electrolyte temperature 50-60 °C, potential value 0.3-0.4 V relative to normal silver chlorine electrode. The bulk of gold is deposited within 2-3 hours of electrolysis. After a concentration of 100-120 mg/l to obtain a residual gold content in the solution of 10-20 mg/l, the duration of deposition increases to 6-12 hours. The gold content in the cathode deposit is 70-85%, silver 10-25%, copper 0. 5-5%, zinc 0.1-0.2%, iron 0.1-0.4%. The bulk of base metal impurities remains in the electrolyte and is returned to the gold desorption process. The current efficiency for gold and silver is about 30%. Increasing the current density to more than 15-20 A/m2 is impractical, since this increases the deposition of impurities, in particular copper, the current efficiency decreases, the entrainment of the electrolyte with bubbles of released hydrogen increases, and the structure of the deposit changes.

8. Electroelution. The method of electroelution, or electrodesorption, is a combined process of desorption of gold and its electrodeposition from solution. The electroelution of gold using desorption of an NH 4 SCN solution was studied. For this purpose, 1.5 l of saturated anionite Deacidit N with 24% strong basic groups, after preliminary elution of nickel and copper from it with a NaCN solution, was mixed in an electrolysis bath with 10 l of 5 N. (380.65 g/l) NH 4 SCN solution (6.67 volumes of solution per 1 volume of resin). A carbon plate was used as an anode, and lead foil was used as a cathode. The cathode current density was 154 A/m2, voltage 1.5 V. Gold is extracted from the resin almost completely within 24 hours (up to a content of 0.066 g/l) and is quite completely deposited on the lead cathode (up to a content in the solution of 15 mg/l). At a voltage across the electrodes of 1.5-1.6 V, ammonium thiocyanate is not destroyed and can be used for electroelution many times. The destruction of NH 4 SCN occurred when the voltage on the electrodes was 1.8-1.86 V. In the case of non-selective electroelution 5 N. With a solution of NH 4 SCN at a cathode current density of 300 A/m2 and a voltage at the electrodes of 1.6 V, gold and copper were washed out of the resin and deposited on the cathode almost 100%, nickel was practically not eluted, but some of it (up to 9%) was deposited at the cathode. The duration of the process is 60 hours. The advantage of the method is the small volume of the elution solution and the acceleration of the desorption process, but the disadvantages inherent in rhodanium eluates remain. The electroelution method using weakly acidic TM solutions was developed by M.S. Girdasov. The setup diagram for electroelution is shown in Fig.

A saturated anion exchange resin and electrolyte are loaded into a hermetically sealed electrolyzer in an amount of 2-10 m3 per 1 m3 of resin. The electrolyte is a hydrochloric or sulfuric acid solution of HM with a concentration of 55-65 g/l HM and 18-20 g/l HC1 or 25 g/l H 2 SO 4. Platinized titanium is used as anodes, titanium (VT-1 grade) is used as cathodes. The resin is kept in suspension by air agitation. The released HCN is discharged with a current of air into absorption vessels filled with an alkali solution (NaOH or CaO). To avoid oxidation of heavy metals and contamination of the metal deposit with elemental sulfur, anodes should be placed in chambers with diaphragms made of cation exchange or anion exchange membranes. A 1-2% solution of H 2 SO 4 is poured into the anode chambers as an anolyte. The metals deposited on the cathode are protected from the mixed resin by a screen made of polyethylene mesh.

The electroelution process without anode diaphragms (according to M.S. Girdasov) is carried out at a current density of 10-20 A/m2, a voltage at the electrodes of 1-1.5 V for 24-30 hours and allows to obtain a residual gold content in the resin of 0.2 -0.4 mg/g. The use of anodic ion exchange membranes makes it possible to increase the current density to 100-300 A/m 2, as a result of which the speed of the electrodesorption process sharply increases and the duration of the electrodesorption process decreases. At the same time, the concentration of heavy metals in the elution solution decreases from 80-90 g/l in a dynamic process to 55-65 g/l during electroelution, which significantly reduces the loss of this reagent and the volume of equipment. These benefits offset some of the increase in power consumption as current density increases. Thus, when electroeluting an anion exchanger, previously purified from impurities of non-precious metals, using a cation exchanger anode membrane at a current density of 100-150 A/m 2, voltage at the electrodes 3 V, HM concentration - 55 g/l and HCl - 19 g/l, the residual the gold content in the resin was 0.2 - 0.4 mg/g obtained with a process duration of 6 hours. The degree of electrodesorption of gold was 95-97%, with deposition on the cathode 87.0 - 92.0%. The gold content in the cathode deposit is 80-85%. Compared to dynamic elution during electrodesorption of gold, the process accelerates by 3-4 times in the absence of anode diaphragms and a current density of 10-20 A/m2 and by 20-25 times in the presence of diaphragms and an increase in current density to 100-150 A/m2. The increase in the rate of electroelution is explained by a shift in the equilibrium of the gold desorption reaction and an increase in the rate of diffusion of ions of thiourea complexes of gold from the ion exchanger grains due to a decrease in the concentration of these ions in the solution during electrodeposition of the metal. The process of electrodesorption of impurities - copper, zinc and iron - with an alkaline solution of NaCI was developed by A.S. Stroganov. Recommended conditions: anode diaphragm - cation exchange membrane, current density 150-200 A/m2; elution solution: 75-85 g/l NaOH + 60-75 g/l NaCI; the amount of solution is 2 volumes per 1 volume of resin, the process duration is 6-10 hours. The residual total content of impurities in the resin is 1.5-3.5 mg/g. The main obstacle to the industrial implementation of the developed electroelution methods is the lack of a satisfactory electrolyzer design that could ensure stable implementation of the process under production conditions.

Possible options for extracting gold from thiourea solutions can also be used: ion flotation, extraction with TOA, hydrate (NaOH) precipitation with subsequent processing of the resulting gold-bearing sediments by cyanidation, reduction with hydrogen gas and other methods.