Sulfating Roasting-Selenium Distillation-Hydrometallurgical Treatment Process

Its main features are as follows:

(1) The copper-removed residue is processed via the "ammonia leaching - hydrazine hydrate reduction" route to produce silver powder;

(2) The silver-removed residue is treated using the "sodium cyanide leaching for gold - sulfur dioxide reduction" process to obtain gold powder;

(3) Lead is separated with nitric acid. This process addresses the issue of severe lead pollution.

Low-Temperature Oxidizing Roasting-Hydrometallurgical Treatment Process

Its main features are:

(1) Cu, Se and Te are separated with dilute sulfuric acid;

(2) Sodium sulfite (Na₂SO₃) leaching is adopted to replace ammonia leaching for silver extraction, optimizing the operating environment;

(3) Short production cycle and fast capital turnover;

(4) Elimination of lead hazards;

(5) High direct recovery rates of gold and silver, reaching 98.5% and 96% respectively.

Total Hydrometallurgical Treatment Process

This process uses dilute acid oxidative leaching to extract Se and Te. To prevent the dissolution of Au, Pt and Pd, the potential of the leaching process must be controlled. Finally, high-potential leaching is applied to extract Au, Pt and Pd. The chlorinated residue is leached with ammonia or sodium sulfite (Na₂SO₃) and then reduced to obtain crude silver powder. The crude gold and silver powders can be further purified by electrolysis.

Comparison and Analysis of Gold Purification Processes

Comparison of Advantages and Disadvantages

1. Automation Level

All-wet Process: High automation level, with PLC control, equipment fault alarm function, and operation record parameters remotely transmitted to the central control room.

Pyrometallurgy + Hydrometallurgy Process: Low automation level. Some procedures such as smelting slag formation and slag skimming involve poor operating environments.

2. Environmental Protection Investment

All-wet Process: Generates small amounts of waste gas, wastewater, and waste liquid.

Pyrometallurgy + Hydrometallurgy Process: Generates large amounts of waste gas, wastewater, and waste liquid, resulting in high environmental protection investment.

3. Product Purity

All-wet Process: Can produce products with relatively high purity (the purity of gold can reach 99%-99.99%).

Pyrometallurgy + Hydrometallurgy Process: Low product purity (the purity can only reach around 84-90% after preliminary purification).

4. Energy Consumption

All-wet Process: Low energy consumption.

Pyrometallurgy + Hydrometallurgy Process: High energy consumption.

Silver is a sulfur-loving element, and sulfur-containing extractants can be used for its extraction and purification.

Effective extractants for silver include diisooctyl sulfide, dialkyl sulfide, petroleum sulfide, etc. Diisooctyl sulfide has good antioxidant properties and can extract silver from nitric acid medium.

Currently, silver extraction technology is still in the experimental research stage. A factory in China has applied diisooctyl sulfide for silver extraction in small-scale production. In its precious metal refining process, 5-stage extraction is carried out using a centrifugal extractor, with an O/A phase ratio of (1~2):1 and an organic phase extraction capacity of approximately 70 g/L. The silver extraction rate exceeds 99.9%.

Pure silver powder is obtained by reducing the purified stripping solution with hydrazine hydrate at a reduction temperature of 50~60℃. The pure silver powder is then filtered, washed, dried, and cast to produce silver ingots.

For silver extraction using diisooctyl sulfide, the direct recovery rate is over 99%, the total silver recovery rate is more than 99.9%, and the produced silver ingots have a silver content of over 99.9%. Under certain conditions, extraction purification of silver is more economical and reasonable than electrolytic purification.

A Method for Selective Separation of Copper, Zinc and Iron from Low-Grade Complex Chalcopyrite

Relating to the technical field of metallurgical engineering

low-grade polymetallic complex chalcopyrite is finely ground and dried, then mixed with industrial-grade sodium lignosulfonate and ammonium chloride. After uniform mixing and slurry preparation with water, the mixture is heated to 160-180℃ in an autoclave. Industrial oxygen with a purity of 90% is introduced, and a hydrothermal reaction is carried out for 1.0-3.0 hours while maintaining an oxygen partial pressure of 1.0-1.4 MPa. After the reaction is completed, the mixture is cooled to below 80℃. After pressure relief, solid-liquid separation and washing are performed to obtain a solution rich in copper and zinc, and the filter cake is a conversion slag mainly composed of elemental sulfur, lead sulfate and Fe₂O₃.

During the leaching process, a composite catalytic additive composed of a surfactant and a catalyst in a certain ratio is added. Using water as the leaching agent under the pressurized oxygen hydrothermal condition, copper and zinc in the low-grade complex chalcopyrite are leached out, while impurity elements such as lead and iron are directionally converted and retained in the leaching residue.

The leaching process of chalcopyrite involves oxidative leaching of chalcopyrite under low temperature and atmospheric pressure, using silver salt as the catalyst and ammonium persulfate as the oxidant. By controlling the temperature at 70-95℃ and carrying out leaching for 5-10 hours, a copper sulfate solution can be obtained, with a leaching rate of over 96% and a copper recovery rate of over 97%.

Since high temperature and high pressure are not required, the requirements for equipment corrosion resistance and pressure during leaching are low, and no harm is caused to the environment. The silver salt used in the process incurs no loss, while ammonium persulfate can be recycled after regeneration, thus reducing production costs.

This is a chalcopyrite leaching method characterized by simple process flow, short production cycle, low production cost and high production efficiency.

copper wire (or copper chips) by heating displacement recovery of gold. Gold containing chlorinated waste liquid is generally recovered by heating and replacing with copper wire (or copper chips). 

In addition, electrolysis can also be used to recover gold from waste plating solutions with high gold content, while activated carbon adsorption and ion exchange adsorption can be used to recover gold from waste liquids and washing water with low gold content. 

Combined Beneficiation and Metallurgy Method for Mixed Copper Ore：

This method involves processing oxidized-sulfide mixed copper ores with an oxidation rate ranging from 10% to 80% via flotation to separate sulfide ores and obtain copper concentrate. The flotation tailings are then directly subjected to acid agitation leaching, followed by solvent extraction and electrowinning to produce cathode copper, thereby forming an integrated beneficiation and metallurgy process combining flotation and acid leaching.

It can significantly boost the copper recovery rate of oxidized-sulfide mixed copper ores, generally increasing the recovery rate by 10 to 40 percentage points. The process boasts such advantages as being concise and highly efficient, requiring low dosages of flotation reagents, having low fresh water consumption, cutting down on investment and production costs for the solid-liquid separation process of flotation tailings, preventing oxidized copper minerals from entering copper concentrate, featuring a rational flow, and achieving high recovery rates.

Raw materials for ruthenium purification are mainly crude ruthenium and ruthenium-containing salt solutions obtained by distillation separation. Crude ruthenium is usually converted into solution using the alkali fusion-water leaching method. The primary goal of ruthenium purification is to remove osmium, an impurity with similar properties.

Transfer the ruthenium solution into a clean glass or glass-lined distillation kettle, and connect the exhaust pipe to an osmium absorption bottle filled with a 20% NaOH-3% ethanol solution as the osmium absorbent. Heat to boil for 40-50 minutes; osmium volatilizes as OsO₄ (osmium tetroxide) and is absorbed until a thiourea cotton ball test shows no red color. Then add a certain amount of hydrogen peroxide to continue oxidizing and volatilizing residual osmium.

Concentrate the osmium-removed ruthenium solution by heating to a ruthenium content of approximately 30 g/L. Add solid ammonium chloride while the solution is hot to form a dark red precipitate of ammonium hexachlororuthenate ((NH₄)₂RuCl₆). After cooling and filtration, wash the filter cake with anhydrous ethanol until the filtrate becomes colorless. The washed cake is dried, calcined, and reduced with hydrogen to produce sponge ruthenium with a ruthenium content of 98%-99%.

Basic Principles of Rhodium Refining：Rhodium refining is entirely conducted in solution. Crude rhodium metal must first be dissolved, and the sodium bisulfate melting-dilute sulfuric acid leaching method is employed to prepare a relatively high-purity chlororhodic acid solution. Rhodium raw material is mixed and melted with sodium bisulfate; after the melt cools, it is leached with dilute sulfuric acid, and rhodium enters the solution as rhodium sulfate. 

The rhodium sulfate solution is neutralized and hydrolyzed with alkali solution to form rhodium hydroxide precipitate, which is then boiled and dissolved in hydrochloric acid to obtain chlororhodic acid solution. Chlororhodic acid solution reacts with sodium nitrite at high temperature to form rhodium nitro complex. Ammonium chloride is added to the filtered solution to precipitate ammonium sodium hexanitrorhodate; the precipitate is dissolved and boiled in hydrochloric acid, then hydrolyzed and reduced with formic acid to obtain rhodium black, and finally high-purity rhodium powder is obtained via hydrogen reduction at high temperature.