RU2298585C9 - Method of extraction of zinc - Google Patents

Abstract

FIELD: extraction of zinc from zinc ore by means of halogenide-based leaching-out solution; removal of manganese from halogenide solutions of zinc and other metals.

SUBSTANCE: extraction of metallic zinc from zinc mineral includes the following stages: leaching-out of zinc mineral with solution containing halogenide compound of two or more halogenides, electrolysis of solution for obtaining metallic zinc and generation of halogenide compound and return of electrolyzed solution containing halogenide compound to leaching-out stage. Part of electrolyzed solution may be removed in form of flow escaping from cathode section of electrolyte cell of electrolysis process and may be treated for removal of manganese in form of manganese dioxide by addition of limestone and halogenide compound from anode section of electrolysis process. In this case pH and Eh of solution may be controlled by method enhancing formation of sediment of manganese dioxide as compared with formation of zinc sediment. Proposed method is characterized by low operational and capital outlays as compared with well-known methods, extraction of noble metals from solutions, tolerance to low-grade and contaminated concentrates.

EFFECT: reduced power requirements; avoidance of emission of liquid and toxic gases; low temperature and atmospheric pressure in carrying-out operations.

20 cl, 4 dwg, 6 tbl, 9 ex

Description

Field of Invention

The present invention relates to a method for extracting zinc from zinc ore using a halide-based leach solution. The invention also relates to a method for removing manganese from solutions of zinc halide and another metal, and especially solutions of zinc and / or lead halide. In addition, the invention relates to a method for removing silver and / or mercury from metal halide solutions, especially zinc or copper (I) chloride solutions (for example, in a hydrometallurgical method for producing zinc or copper).

BACKGROUND OF THE INVENTION

An existing method for producing zinc involves calcining an ore containing zinc, followed by leaching with acid sulfate and then electrolytically separating zinc from the leach liquor. The calcination process occurs with the formation of sulfur-based contaminants that must be removed from the exhaust gases of the calcination furnace. In addition, if the ore contains high levels of impurities, this can have a detrimental effect on the electrolysis and purity of the resulting zinc. For example, many zinc ores contain significant amounts of manganese. The existing method is limited in its ability to process zinc mineral concentrates containing a significant amount of manganese, because Manganese is leached into the leach liquor with zinc and cannot be removed effectively. Manganese creates problems in the stage of electrolytic extraction of zinc by deposition on the anode in the form of MnO ₂ .

In Australian Patent 6,699,060, the present inventors have developed a multi-stage leaching process followed by electrolysis for copper recovery. Conversely, US Pat. No. 4,292,147 describes a method for recovering zinc using chloride leaching followed by electrolysis.

Applicants of the present invention also observed that in the case of a solution, manganese does not deposit on the anode during electrolytic extraction of the metal from the solution, especially solutions obtained by leaching of zinc and / or lead ores. This makes allowance for the removal of manganese from the solution in other ways.

In addition, in hydrometallurgical methods in which there is a high concentration of halides in solution (for example, from 200 to 300 grams per liter of NaCl and from 10 to 50 grams per liter of NaBr), it is desirable to remove harmful impurities before electrowinning various metals such as zinc or copper in the specified method. Many impurities can be removed from electrolyte solutions of these methods by gradually adjusting the pH and precipitation (for example, up to about pH 6), however, both silver and mercury are not easily removed.

US Pat. No. 4,124,379 describes the removal of silver from an electrolyte of copper (I) chloride, by which metallic copper is added to an electrolyte to reduce copper (II) ions to copper (I) ions, followed by contact with an amalgam that exchanges metal for silver. The amalgam is obtained from metallic mercury with one metal of copper, zinc or iron, preferably a copper fraction coated or associated with metallic mercury. Amalgam requires a separate preparation, requiring the physical association of mercury with copper, and all this makes the method difficult and complicated, as a result of which the cost of the method increases. In addition, the amalgam should then be introduced into the process and contacted with a silver-containing electrolyte, which increases the complexity and cost of the process.

SUMMARY OF THE INVENTION

In a first aspect, the present invention relates to a method for extracting metallic zinc from a zinc mineral, comprising the steps of:

- leaching of the zinc mineral in one step using a solution including halide compounds formed from two or more different halides to leach zinc into the solution;

- electrolysis of a zinc-containing solution to produce zinc metal and the generation of a halide compound and

- return of the electrolyzed solution, including the halide compound, to the leaching stage.

The inventors of the present invention unexpectedly found that a halide compound formed from two or more different halides has a sufficiently high oxidizing potential so that zinc (and other metals such as lead) can be directly leached into the solution in a single step without the need for a multi-stage leaching operation (for example, as described by the author of the invention in Australian patent 669906).

A halide compound is usually formed anodically (i.e., at the anode) in an electrolysis step. Typically, a halide compound is formed at a lower oxidation potential than the potential for the formation of many insoluble forms of impurities in solution. For example, a halide compound typically forms at a lower oxidation potential than the potential for manganese dioxide. This allows you to save impurities in the solution and then remove from the solution, either in the present line, or in a separate removal stage.

хотя может быть также образован или затем использован газообразный BrCl. Было указано, что такие соединения эффективно окисляют цинковый минерал (и другие минералы), так что требуется только одностадийный способ выщелачивания цинка.Two different halides or a large number of different halides are chlorine and bromine, and preferably the halide compound is a soluble halide complex formed at the anode, such as

although gaseous BrCl may also be formed or then used. It has been indicated that such compounds effectively oxidize the zinc mineral (and other minerals), so that only a one-step zinc leaching method is required.

The leaching of the mineral is preferably facilitated by a catalyst, which typically catalyzes the oxidation of the mineral with a halide compound. In this regard, the catalyst is preferably a metal catalyst, such as copper, which may be present in zinc ore or which is introduced into the leaching process (for example, in the form of copper particles in the first leaching stage).

The method preferably further includes one or more subsequent steps to ensure zinc leaching in one step and removing impurities in the subsequent step (s). For this, preferably, at least two subsequent stages are combined with a leaching step, so that the method includes a first zinc leaching step, a second aeration step and a third sulfate precipitation step.

Air is preferably introduced in a second step to oxidize and precipitate any iron present in the zinc mineral. When sulfur is present in the mineral, it is preferable to add limestone in the third stage to precipitate the sulfate formed as a result of the oxidation of sulfur in the first leaching stage as calcium sulfate.

The liquor from the leaching process preferably has these solid precipitates isolated from it, and before electrolysis, any metal of gold and platinum group metals (PGM's) present in it is preferably removed by passing liquor over activated carbon (usually a carbon column is used) for adsorption of gold and PGM's on coal.

Before electrolysis, the liquor is preferably then passed through a series of cementation processes in which zinc dust (usually using part of the zinc formed during the electrolysis) is added to the liquor for cementing any of the metals copper, silver, lead and other impurities present in the liquor ( as a result of their presence in the mineral).

Preferably, before electrolysis, and if necessary, the liquor is then passed to a step for further removal of iron (and residual metals), in which lime and a halide compound from the electrolysis step are added to oxidize and precipitate iron as iron oxide (III). In this same step, preferably, the lime and the halide compound will oxidize and precipitate at least a portion of the manganese present as manganese dioxide.

Part of the electrolyzed solution (usually spent catholyte) is preferably removed and treated to remove manganese from it. This part is preferably a stream exiting the cathodic compartment of the electrolytic cell for the electrolysis process, the outlet stream having added limestone and a halide compound from the anode compartment of the electrolysis process to precipitate manganese dioxide.

In the manganese removal step, the pH and Eh of the solution are preferably adjusted so that a preferred formation of a manganese dioxide precipitate occurs compared to the formation of a zinc precipitate.

The pH is preferably controlled by the incremental addition of limestone to increase the pH of the solution to a level at which Eh can be raised by the halide compound to a level at which the formation of MnO ₂ is more favorable, or to a higher level. The amount of added limestone is preferably less than the stoichiometric amount required for the formation of MnO ₂ , so less zinc is precipitated.

Before electrolyzed liquor is returned to the leaching process, if necessary, part of the catholyte (usually the catholyte from which manganese was removed) is preferably passed to the magnesium removal step, to which hydrated lime is added to remove any zinc (which is returned to the leaching process) first and then removing magnesium as a precipitate of magnesium oxide.

Therefore, in a second aspect, the present invention relates to a method for removing manganese from a metal halide solution, which is subjected to electrolysis to produce one or more metals, comprising the steps of:

- in electrolysis, the cathodic extraction of one or more metals, at the same time, the anodic formation of a halide compound from two or more different halides with an oxidation potential lower than the oxidation potential for the formation of manganese dioxide;

- removing part of the solution and processing it to remove manganese from it; and

- returning the treated portion to the metal halide solution.

The formation of a halide compound with an oxidation potential lower than the potential for the formation of manganese dioxide, manganese can be stored in solution and this then allows you to remove it from the solution, usually in a separate separation stage.

The metal halide solution is preferably a liquor obtained from a leaching process in which a mineral concentrate is leached in a halide-containing solution to leach one or more metals into a solution, most preferably the zinc halide liquor of the first aspect.

The liquor is preferably sent to the electrolysis stage as an electrolyte and, after electrolysis, is returned to the leaching process to further leach the mineral concentrate. In this regard, the manganese removal method of the present invention is carried out as part of a mineral leaching process and electrolytic extraction in a closed loop.

The part of the solution that is removed and treated to remove manganese from it is preferably a stream leaving the process of leaching the mineral and electrolytically extracting the metal in a closed loop, this stream is returned to the specified process after the removal of manganese.

The leaching of the mineral is preferably facilitated by the formation of a halide compound at the anode, which is preferably returned with the electrolyte to the leaching process.

хотя можно также использовать газообразный BrCl.Preferably, two or more different halides are chlorine and bromine, and a preferred halide compound is a soluble halide complex formed at the anode, such as

although gaseous BrCl can also be used.

The leaching of the mineral is preferably facilitated by a catalyst, which typically catalyzes the oxidation of the mineral with a halide compound. For this purpose, the catalyst is preferably a metal catalyst, such as copper, which is present in the leaching process or is introduced into it.

The mineral preferably includes zinc and / or lead, which are metals formed in the electrolysis step.

Manganese in this part of the solution is isolated from it, preferably by incremental addition of an alkaline reagent, preferably a reagent that causes the precipitation of manganese as MnO ₂ . The MnO ₂ precipitate can then be separated from a portion of the solution until it is returned to the metal halide solution. Part of the solution is preferably part of catholyte from the electrolysis step, usually spent by catholyte.

The alkaline reagent is preferably calcium carbonate. When the metal to be isolated also tends to precipitate when an alkaline reagent such as calcium carbonate (e.g. zinc) is added, the alkaline reagent is preferably added in an amount less than the stoichiometric amount to form MnO ₂ . To increase the oxidation potential of a portion of the solution to a level that favors the formation of MnO ₂ , a reagent with a high redox potential is also added. The preferred reagent with a high redox potential is the halide compound (or its gaseous form) formed at the anode in the electrolysis stage. Alternatively, the reagent with a high redox potential may be a hypochlorite or hypobromite salt (such as calcium hypochlorite), which is added to part of the solution along with an alkaline reagent.

Based on this, in a third aspect of the present invention, a method for removing manganese from a metal halide solution from which at least one metal can be obtained, comprising the steps of:

- adjusting the pH and Eh of the solution in a manner that favors the formation of a precipitate of manganese dioxide compared with the formation of a precipitate of at least one metal and

- removal of a precipitate of manganese dioxide from the solution.

The pH value is preferably controlled by incremental addition of an alkaline reagent to increase the pH of the solution to a level at which Eh can be raised to a level that favors the formation of MnO ₂ , or above such a level. The amount of alkaline reagent added is preferably less than the stoichiometric amount required to form MnO ₂ so that less deposition of at least one metal occurs.

The pH is preferably increased by the addition of calcium carbonate, and Eh is preferably increased by the addition of a reagent with a high redox potential according to the first aspect.

At least one metal is preferably zinc and / or lead.

The solution in the third aspect is preferably an effluent from the second aspect. The solution of the third aspect is preferably also a liquor from the leaching process of the first aspect.

In a fourth aspect of the present invention, there is provided a method for removing dissolved silver and / or mercury from a metal halide solution, wherein the metal is capable of forming cement with silver and / or mercury, the method comprising the steps of:

(a) reducing the Eh of the solution with a reducing agent to a level that causes metal deposition;

(b) adding an ionic compound to a solution that interacts with the deposited metal in a manner that causes the formation of silver cement and / or mercury with the deposited metal, and

(c) cement removal.

The formation of a metal precipitate and the addition of an ionic compound results in the carburization of silver and / or mercury, which is specific to the present invention, without the need for the formation and then introduction of a separate amalgam.

The method of the fourth aspect typically provides the combination removal of silver and mercury and has particular application in the processing of chloride solutions derived from mineral ores containing both silver and mercury (which are often found together in nature in many ores).

The most preferred application of the method relates to solutions of zinc and copper chlorides (such as solutions obtained by leaching zinc or copper ore). In this regard, the metal is zinc and / or copper and the halide is chlorine, and the method is preferably carried out when copper is in its form of copper (I). In at least preferred forms of the process, silver and mercury are desirably removed to deposit only a minimal amount of copper in the form of this metal.

Subsequently, the Eh values are presented with reference to the standard potential of the Ag / AgCl electrode. The eh of the solution is preferably reduced to below 0 mV. Eh can be reduced up to about -200 mV, but usually it is reduced to about 150 mV. A typical electrolyte of the copper (I) chloride method has an Eh of about +150 mV, and thus, when the Eh of the solution is reduced to -150 mV, an additional 300 mV of driving force is available to precipitate silver.

Eh is preferably reduced by adding to the solution a reducing agent selected from one or more reducing agents: metallic aluminum, metallic zinc, metallic iron or hydride, metal borohydride or dithionite. The most preferred reducing agent is aluminum metal, as it is cost effective, readily available and causes the formation of more easily filtered cement.

Sodium dithionite can also be used as a reducing agent to reduce Eh, with the final reduction in Eh being optionally achieved with sodium borohydride (i.e., at the end of the Eh reduction step). By using an immediate reducing agent, the removal of any mercury present in silver is not affected by fusion with large quantities of copper.

The added ionic compound is preferably ionic mercury in either or both valence states of monovalent mercury and divalent mercury (Hg (I) or Hg (II)). Another ionic compound includes ionic gold, etc., which is less cost effective. The use of an ionic compound, such as ionic mercury, is preferred over the use of metallic mercury due to the problems mentioned above in US Pat. No. 4,124,379.

Cemented mercury is usually removed from the solution by passing an inert gas through the solution so that at least some of the mercury is removed from it with the gas. For example, an inert gas can be bubbled into the solution; an inert gas such as nitrogen is usually used (i.e. because of its availability and low cost).

After removal from the solution, the inert gas is usually washed to remove mercury from it before it is recycled to the solution for further use. In this case, the inert gas can be washed with a solution of divalent copper (i.e. containing a divalent copper ion) to remove the released mercury from it. Other oxidizing agents, such as mercury (II) chloride, may be used to flush mercury from nitrogen. Carbon dioxide can also be used.

In addition, the removal of metallic mercury from the solution can be enhanced by heating the solution to further evaporate the metallic mercury. When an inert gas is used in the removal method, an increased amount of vaporized metallic mercury is thus captured in the gas.

Alternatively, metallic mercury can be removed from the solution by passing the solution over activated carbon, so that metallic mercury is adsorbed on the carbon. In the case of such an alternative, mercury can then be removed from activated carbon by passing a second solution over activated carbon, the second solution having an Eh value that is sufficient to convert the adsorbed mercury into ionic mercury, thereby dissolving it in the second solution.

This alternative method is suitable for easy implementation in line and, thus, is easily adapted for commercialization; for example, a column containing activated carbon can be used, and an electrolyte with a reduced Eh can be pumped continuously through the column using a pump (for example, at about -150 mV). After the limiting amount of mercury has been adsorbed on coal, automatic switching can be used to pump a second solution through the column. Dissolved mercury can then be easily recovered from the second solution.

A typical application of the method of the fourth aspect is observed in hydrometallurgical extraction methods with solutions of zinc chloride and / or copper (I). A typical value of Eh in hydrometallurgical processes using such solutions and up to the removal of silver and / or mercury is approximately +150 mV.

A brief description of the graphic material

Despite any other variations that may fall within the scope of the present invention, preferred embodiments of the invention will now be described by way of example and with reference also to the accompanying graphic material, in which:

figure 1 shows a schematic flow chart illustrating a method of leaching and electrolytic extraction of zinc in a closed cycle, including the stage of removal of manganese;

2 shows a Pourbaix diagram for a Mn-Cl-H ₂ O system at 25 ° C;

figure 3 is a graph of the removal of silver and mercury from the development of the process and

4 is a graph of mercury removal versus time from a copper (I) electrolyte using a nitrogen sparge.

Methods for recovering zinc according to the invention

A preferred zinc recovery method of the present invention is shown schematically in FIG. 1, and a description of this method will now be made with reference to FIG.

The method is intended for producing high-purity metallic zinc from concentrates of complex mixed zinc / lead sulfides obtained from zinc / lead ores. Lead and silver were obtained as cement by-products, while gold and platinum group metals (PGM's), when present, were obtained as an ingot.

Concentrates containing significant levels of iron were easily processed, with all leached iron transferred to the liquor residue in the form of hematite, while sulfide sulfur was converted to elemental sulfur. High levels of impurities such as manganese and arsenic were also easily adjusted, with arsenic passing into the liquor residue as environmentally stable iron arsenate, and manganese precipitated as a separate residue as manganese dioxide (described below).

(в дальнейшем "Halex"). Это соединение проявляло высокие характеристики выщелачивания, когда его рециркулировали для обработки поступающего концентратного сырья.The preferred method was based on electrolytic deposition at the cathode of high purity zinc from a purified electrolyte, sodium chloride - sodium bromide. During the electrochemical extraction step, a halide compound was generated in solution at the anode. Such a compound was a mixed halide compound such as soluble

(hereinafter "Halex"). This compound exhibited high leaching characteristics when it was recycled to treat incoming concentrate feed.

The preferred method includes three main stages of leaching, purification and electrochemical isolation, as shown in figure 1. The leaching process included one leaching stage (one reactor) and was combined with a number of subsequent stages (reactors). Zinc concentrate and oxidizing agent (halide compound) were sent to the leaching stage. The cleaning consisted of the stages of cementation and alkaline deposition. During electrochemical isolation, a series of electrolyzers with a diaphragm were used with the option of continuous removal of the product in dendritic form or the production of a conventional cathode with electrolytic zinc coating.

Zn electrolysis (electrolytic deposition)

Electrolytic deposition was an integral part of the preferred method. Zinc metal was precipitated by electrolysis from a purified electrolyte, which had 100 g / l zinc, 50 g / l sodium chloride (ordinary salt, NaCl), 50 g / l calcium chloride (CaCl ₂ ) and 110 g / l sodium bromide ( NaBr). All other components, including the "equilibrium" levels of many other elements (manganese, magnesium, etc.), were considered as impurities.

Electrolysis included passing an electric current at 500 A / m ² in the region of the electrode with the formation of very pure zinc on a negatively charged cathode. The initial zinc electrolyte concentrate, in turn, was depleted from a concentration of 100 to 50 g / l, which was the concentration of the steady state of the cell.

Oxidizing (leaching)

The spent catholyte continuously penetrated through the membrane of the nonwoven fabric M located in the electrolytic cell of the EU (figure 1). The catholyte penetrated a positively charged electrode (anode).

(Halex). Однако могли быть образованы другие галогенидные соединения. Это галогенидное соединение рассматривалось как молекула хлора, удерживаемая в растворе бромид-ионом, и было отмечено, что это соединение является очень сильным выщелачивающим агентом при окислительном потенциале (Eh) 1000 мВ (против Ag/AgCl). Образовавшийся, содержащий Halex раствор (электролит) из отделения анода (анолит) используют для выщелачивания концентрата сульфида цинка.Chloride (Cl ^- ) and bromide (Br ^- ) were preferably present in the solution, and therefore, the preferred formation of the halide compound

(Halex). However, other halide compounds could be formed. This halide compound was considered as a chlorine molecule held in solution by the bromide ion, and it was noted that this compound is a very strong leaching agent with an oxidation potential (Eh) of 1000 mV (against Ag / AgCl). The resulting Halex-containing solution (electrolyte) from the anode compartment (anolyte) is used to leach zinc sulfide concentrate.

Leaching stage

не был израсходован. В первом CSTR имело место значительное растворение железа и окисление сульфида.Zinc concentrate and Halex oxidizing agent from the electrolysis process were sent to the liquor of the first stage (first, continuously stirred reactor-tank (CSTR) in figure 1). CSTR operations were performed at atmospheric pressure and a solution (electrolyte) temperature of 85 ° C. The leaching (oxidation) reaction was first carried out without sparging with air (aeration (or air injection)) until the entire Halex oxidizing agent

was not used up. Significant iron dissolution and sulfide oxidation occurred in the first CSTR.

The second and third CSTR's were introduced into the process for further aeration to remove iron in the form of a precipitate of iron oxide (III). Any arsenic present was oxidized in the first CSTR and then precipitated as environmentally stable iron arsenate.

Recycled copper cement (i.e., cement from the subsequent Cu and Ag cementation step) was added to the first CSTR to increase both re-absorption of oxygen and metal extraction. Copper helped catalyze the oxidation of zinc concentrate with a halide compound.

When the leaching was completed, limestone was added to the third CSTR to drain (precipitate) any dissolved iron that remained and any sulfate present. Limestone also balanced the generation of sulfuric acid as a result of the oxidation of sulfur in the first CSTR. Optionally, a fourth CSTR was used to increase the residence time (reaction) and to allow maximum precipitation of iron and sulfate.

Purification stage

The leach residue is separated from the zinc-rich “rich” solution by filtration and washed before discharge to the collapse at the first solid-liquid separation point (S / L).

When gold or PGM's were present, they were extracted (oxidized) during leaching and recovered by feeding a zinc-rich solution through a column filled with activated carbon on which gold and PGM's were adsorbed. This coal can then be separately eluted with an eluent to isolate gold and other PGM's.

The zinc-enriched solution was then further purified by a series of cementitious reagent reactions with zinc dust. It was a two-stage, CSTR-facilitated operation with copper and silver, mainly removed in the first stage, which was carried out at a temperature of 85 ° C. In the second stage, excess zinc dust was added to remove remaining impurities such as cadmium, lead, nickel, cobalt, thallium, etc.

After two additional stages of the separation of the solid-liquid system, a small amount of Halex vapors from the anolyte of the electrolytic cell was already added to the solution loaded with relatively pure zinc in the CSTR until an Eh of 700 mV was reached. This ensured that any remaining iron was oxidized to the oxidation state of ferric iron (Fe ³⁺ ). Subsequent addition of crushed limestone to raise the pH to 4.5 precipitated most of the remaining impurities (e.g. Bi, Fe, In, Ge, etc.). These precipitates were removed by filtration and either discharged or processed again to recover economically viable by-products such as indium.

It was also apparent that at least some of the manganese was able to be removed from the metal halide solution as a precipitate of manganese dioxide if the pH was adjusted to higher than 3.2 at this stage of the CSTR. The pH was raised above this level by the addition of calcium carbonate (usually ground limestone). At this pH value and at a redox potential of 600-800 mV against Ag / AgCl (800-1000 mV against SHE), manganese entered the favorable stability region of the MnO ₂ diagram of the Pourbaix. The remainder of manganese dioxide was then removed with precipitated iron from the solution. Eh was also regulated (increased) by the addition of a Halex reagent with a high redox potential. The specific chemical reactions and reaction conditions for the removal of manganese are described in detail briefly in the following text for the subsequent manganese removal cycle (for spent catholyte effluent), which also uses limestone and Halex for this purpose. It should be noted that in the effluent to be described, the concentration of manganese in the solution is usually higher than it was in the present CSTR purification step.

It is noted that any impurities remaining now do not contaminate zinc during electrolysis and were removed either in the purification cycles of manganese or magnesium. In particular, since the halide compound is formed at the anode at an oxidation potential lower than the oxidation potential for the formation of manganese oxide, this metal does not electrolytically extract the extracted zinc.

The purified zinc solution was then electrolyzed as described above to produce high purity zinc and generate a leachate, a halide compound, for recycling to the leach stage. The zinc product was washed and dried in an inert atmosphere before melting in an oven. A certain amount of zinc dust formed from zinc in the furnace was used in the cementation stages (described above).

Removal of manganese and magnesium

These metal removal cycles included catholytic effluent treatment, where limestone and Halex were used in the manganese cycle, and hydrated lime was used in the magnesium cycle.

It is noted that this method is usually suitable in the processing of mineral concentrates containing manganese. In the present zinc recovery method, manganese leached into the liquor was removed in equal proportions (i.e., it was removed at least at the rate of leaching it into the liquor) to prevent it from accumulating in the process solution.

Up to a maximum of 20% of the total solution flowing through the EC electrolytic cell, was withdrawn in the form of an effluent BS from spent catholyte. This stream was treated with Halex vapor and limestone to precipitate manganese as manganese dioxide MnO ₂ .

It was noted that the vapor pressure of Halex above the electrolyte is temperature dependent and the vapor itself probably consists of gaseous BrCl.

The method for removing manganese is chemically depicted below in equations (1) to (3), where it is assumed that the Halex vapor is BrCl.

The electrolyte composition in spent catholyte was approximately as follows: 50 g / L NaCl, 110 g / L NaBr, 50 g / L CaCl _2, and 50 g / L Zn ²⁺ . However, the manganese removal method can also be carried out on the electrolyte of the method before electrolysis. When using the effluent from the spent catholyte of the zinc electrolysis cell, the level of manganese in the solution depends on the amount of leached into the solution and the amount of effluent reaching the removal of Mn.

The inventors of the present invention showed that manganese was completely removed from the effluent as manganese dioxide by adding Halex vapor to it. The electrolyte-pair reaction was found to be very fast in the presence of incrementally added small amounts of limestone.

The deposition of manganese as manganese dioxide is predicted by the Pourbaix diagram shown in FIG. 2. As can be seen, manganese does not precipitate at a neutral redox potential to a pH above 7.5. At a high redox potential of 600-800 mV against Ag / AgCl (800-1000 mV against SHE) and a pH of 3.2, however, manganese entered the stability region of MnO ₂ and was able to precipitate as MnO ₂

A small amount of zinc also precipitates during the removal of manganese. However, the authors of the present invention showed that less zinc was deposited at lower levels of aqueous zinc (e.g., 25 g / L Zn ²⁺ instead of 50 g / L Zn ²⁺ ) or when limestone was incrementally added in small amounts to maintain a pH of about 3.2 (as opposed to precipitation of manganese in the presence of a stoichiometric excess of limestone).

Thus, using spent catholyte containing 50 g / l Zn ²⁺ and 15 g / l Mn ²⁺ , with the slow addition of limestone, only 7.8% of the zinc in the solution was precipitated together with manganese. For an 8% catholytic effluent, this equals a loss of less than 0.7% of the total zinc production relative to the precipitate of the effluent.

Examples of successful manganese removal experiments are shown below. As can be seen from them, complete removal of manganese was performed in the first two experiments, while complete removal was not achieved in the third experiment. It should be noted that intermediate analyzes of zinc in solution were only an approximate guideline, since zinc deposited from solution was difficult to characterize due to experimental error associated with this analysis. Accurate determinations of zinc deposition were determined by direct analysis of the final residue.

The results of experiments 2 and 3, shown in tables 2 and 3, reflect the way in which this method was used for the method of separation of lead / zinc metals in which Halex was formed at the anode. The results showed that for 8% of the effluent from the spent catholyte, the concentration of recycled manganese was maintained at 15 g / l. Co-precipitation of zinc with precipitation of MnO ₂ represents a loss of approximately 0.6% of the total zinc production.

Manganese Removal Examples

Non-limiting examples of manganese removal methods will now be described.

     Example 1

An electrolyte comprising 200 g / L NaCl, 150 g / L NaBr, 30 g / L Ca ²⁺ , 25 g / L Zn ²⁺ and 5 g / L Mn ²⁺ was obtained as a typical lead / zinc recovery process . 25 g / L CaCO _{3 was} added to the electrolyte in the reaction vessel at the beginning of the removal method. Ca (OCl) _{2 was} added directly to the electrolyte to oxidize the solution. The temperature of the solution was maintained at a typical electrolyte temperature of the method from 60 to 65 ° C. The results are shown below in table 1.

     Example 2

An electrolyte comprising 50 g / L NaCl, 110 g / L NaBr, 50 g / L CaCl ₂ , 50 g / L Zn ²⁺ and 15 g / L Mn ²⁺ was obtained as a typical lead / zinc recovery process. 85 g / L CaCO _{3 was} added to the electrolyte in the reaction vessel at the beginning of the removal method. Halex vapors were generated externally and pumped into the reaction vessel. The temperature of the solution was maintained at a typical electrolyte temperature of the method from 60 to 65 ° C. The results are shown below in table 2.

     Example 3

An electrolyte comprising 50 g / L NaCl, 110 g / L NaBr, 50 g / L CaCl ₂ , 50 g / L Zn ²⁺ and 15 g / L Mn ^{2+ was} obtained as a typical lead / zinc recovery process. CaCO _{3 was} added to the electrolyte in the reaction vessel in small doses throughout the removal process. Halex vapors were generated externally and pumped into the reaction vessel. The temperature of the solution was maintained at a typical electrolyte temperature of the method from 60 to 65 ° C. The results are shown below in table 3.

     Example 4

10.0 g of the manganese precipitate from Example 2 was added to 1.0 liter of demineralized water in the reaction vessel. H ₂ SO ₄ was added. The results are shown below in table 4.

     Example 5

10.0 g of the precipitate from Example 3 was added to 1.0 liter of demineralized water in the reaction vessel. Halex vapors were generated externally and pumped into the reaction vessel. The results are shown below in table 5.

Manganese was thus completely removed from the effluent as manganese dioxide by the addition of Halex vapors. It was found that the reaction of the electrolyte pair proceeds very quickly in the presence of small amounts of limestone.

A small amount of zinc was also precipitated during the removal of manganese. However, less zinc was precipitated at lower levels of aqueous zinc (e.g. 25 g / L Zn ²⁺ instead of 50 g / L Zn ²⁺ ) or when limestone was added incrementally in small amounts to maintain a pH of about 3.2 (as opposed to precipitation of manganese in the presence of a stoichiometric excess of limestone).

Thus, through the use of spent catholyte containing 50 g / L Zn ²⁺ and 15 g / L Mn ²⁺ , with the slow addition of limestone, only 8.4% of the zinc in the solution was precipitated together with manganese. For an 8% catholyte stream, the output loss is less than 0.7% of the total zinc production in the precipitate of the exit stream.

In the magnesium removal process, spent catholyte from the manganese removal process had slaked (or caustic) lime in CSTR's series added to them to precipitate in the first two zinc CSTR's. This zinc was then returned to the leaching process, usually at the stage of adding limestone (as shown in figure 1). Then, in the next two CSTRs, slaked lime was added to precipitate magnesium oxide and thereby prevent magnesium from entering the process. This removal step was only required for periodic action, as it usually reduced magnesium levels in the zinc mineral. The treated spent catholyte was then returned to the leaching process, usually in the first leaching CSTR.

It should be noted that the economy of the preferred zinc recovery process in which Halex is formed at the anode is site specific. For the concentrate manufacturer, a loss of 0.6% of zinc during production as a result of the removal of impurities (especially manganese) represents a significant improvement with respect to losses in the prior art smelter, where approximately 4% of the zinc contained is subtracted to calculate the appropriate metal content. However, when higher zinc recoveries are desired by the method, a certain amount of zinc precipitated from the solution with manganese is recovered by re-suspending the precipitate of the effluent in water and acidifying it with sulfuric acid. Zinc was able to dissolve again with a selectivity of approximately 90%. By adjusting the higher proportion of the effluent required to partially re-dissolve the manganese, the loss of zinc with the precipitate of the effluent was reduced to 0.1% of the total zinc production.

An important feature of the preferred method was that all impurities, including manganese, mercury and arsenic, were either recovered as marketable by-products or stabilized for disposal. Another equally important sign was that heat was provided by exothermic leaching reactions. This, combined with the addition of air to the liquor, evaporated water and thereby maintained the water balance at a neutral value. As such, there was no liquid flow.

Removing silver and mercury

Now, preferred embodiments for the removal of silver and mercury will be described with reference to the lye of copper (I) chloride, but it should be understood that the invention can be used with solutions of other metal chlorides (including zinc chloride).

It was noted that during the electrolytic production of copper, the typical value of Eh of a solution of a copper (I) ion upon contact with metallic copper was +150 mV (rel. Ag / AgCl). Unexpectedly, a possible chemical decrease in the Eh of the copper (I) ion solution to -150 mV (Ag / AgCl) with minimal deposition of metallic copper was found. This allows a driving force value above 300 mV for silver deposition. In addition, the silver-mercury amalgam formation efficiency was not significantly reduced by fusion with Cu. It has been observed that a certain amount of Cu is precipitated together with the silver / mercury amalgam, but it is noted that the amount of Cu in the amalgam (residue) is 1/20 of the amount in the residue of the methods of the prior art, which thus makes the extraction of silver and mercury much easier .

It was found that after removing silver, traces of mercury can be removed from a solution of copper (I) ions by sparging nitrogen through it. After bubbling, the nitrogen is washed, for example, with a solution of copper (I) ions to remove trapped mercury from it. Nitrogen is then recycled to the solution to further remove mercury.

     Examples of silver / mercury removal

     Example 6

An improved method for removing silver and mercury from chloride solutions was developed using an advanced reducing agent, namely aluminum metal, which, as noted, is a cheap and readily available reducing agent compared to the reducing agents described below and compared to zinc dust or metallic iron.

In addition, the use of aluminum metal caused the formation of a sponge-like copper precipitate in copper chloride solutions, which after subsequent cementation of silver and / or mercury was easier to filter than a precipitate formed from other reducing agents. Further, aluminum was involved in this precipitate, while zinc dust was not involved in it, and aluminum could thus be removed from the mortar with cement. The outer coating of copper formed by metallic iron and its core did not react, which, therefore, made it less effective as a reducing agent.

In an improved method for removing silver and mercury from chloride solutions, aluminum metal is added in small amounts (0.05-0.3 grams per liter of electrolyte) to the electrolyte, which causes a decrease in Eh of the solution to a value of -120 to -150 mV (against Ag / AgCl) and high surface area copper deposition. It is noted that this copper is a relatively coarse, sponge-like precipitate. It was very reactive to the added mercury ions and was able to form an amalgam with a chemical composition in which there is approximately 7.5% silver, copper, the amount of which varies from 10% to 50%, the balance is the amount of mercury.

The amalgam formed by the addition of aluminum was also more compatible and had a more easily filterable particle size. The use of amalgam is an improvement over prior art methods for separating solid and liquid phases.

The levels of mercury in the solution exiting this step were similar to those described for Example 8, and the same mercury removal steps (nitrogen sparging or activated carbon adsorption) were used as described below.

The improved method was tested on a bench scale before testing during a 48-hour operation on a pilot scale at a continuous flow rate of 20 l / h. The pilot test was carried out using four 20-liter tanks through which nitrogen was sparged, with aluminum foil added to tank 1 at a dosing rate of 0.05 to 0.15 grams per liter of electrolyte (1-3 g / h) and ionic mercury ( for example, mercury (I) or mercury (II) nitrate) was added to tank 1 at a mass ratio of 3: 1 with silver starting material (1.5 g / h). The treated electrolyte had concentrations of 75 g / L Cu ⁺ , 280 g / L NaCl and 28 g / L NaBr at 35 ° C. Figure 3 shows the results of average silver and mercury and Eh contents during 48 hours of operation.

As can be seen in FIG. 3, the average silver concentration in tanks 3 and 4 throughout the 48 hours of operation was 1.5 mg / L with a minimum concentration of 1.2 mg / L. The observed mercury concentration ranges were as shown in table 6.

Table 6 Mercury concentration range through pilot plant - 48 hours experience Tank number Analysis number Average concentration (μg / L) Concentration range (μg / l) one 6 6311 815-24700 2 6 1293 337-4300 3 6 468 275-773 four 6 322 277-388

     Example 7

In this example, the reducing agent was sodium borohydride, but it was noted that other reducing agents may also be used, such as sodium dithionite and metal hydrides (e.g. sodium and calcium hydride, optionally in combination with sodium borohydride to reduce the amount of borohydride used). A more cost-effective method included the use of sodium dithionite to carry out a reduction in Eh in the mass (i.e., usually to the limit level), with the final reduction being carried out with sodium borohydride (i.e. until a preferred lower level of about -150 mV (Ag / AgCl)).

In contrast to the use of aluminum metal as a reducing agent, sodium borohydride precipitated ultrafine sponge copper and, when ionic mercury was added, formed an amalgam that was difficult and expensive to filter and which was less stable. In addition, the borohydride compound itself was both expensive and unstable in water. If the borohydride compound is not stored at a high pH, it will quickly hydrolyze, causing the loss of an expensive reagent. Similarly, the small amalgam formed after the addition of mercury and sodium borohydride must be completely protected from oxidation or it is quickly dissolved again, releasing silver and mercury ions back into the solution. Despite these difficulties, it was also possible to use sodium borohydride as a reducing agent with the addition of ionic mercury (Hg (I) or Hg (II), monovalent or divalent mercury, respectively), and the formation of amalgam of copper, mercury, and silver resulted in to remove both mercury and silver from the solution.

Silver removal

45 mg of NaBH _{4 was} added to 1.5 liters of reduced copper electrolyte, which contained 22 ppm. silver and caused a decrease in the Eh of the solution from +120 mV to -135 mV (Ag / AgCl). 65 mg of mercury in the form of mercury (II) nitrate was added to the electrolyte and thereby reduced the silver level in the solution from 22 ppm. up to 4.4 ppm in 40 minutes. Another 65 mg of mercury was added as mercury (II) nitrate, and the silver level dropped to 1 ppm over 20 minutes. The precipitated solid was filtered and dried, and it was found to contain 7.4% silver, and the copper content ranged from 10% to 47%, the rest being mercury.

The process was repeated using sodium dithionite to achieve Eh levels lower than 0 mV, with a final decrease in Eh achieved in small amounts of NaBH ₄ .

     Example 8

Mercury removal

The copper (I) electrolyte from Example 7 of silver removal was adjusted to -150 mV by adding a reducing agent, such as NaBH _{4, to it} . This shifts the usual equilibrium of the solution (i.e., the equilibrium between metallic mercury, mercury (I) ion and mercury (II) ion) towards the formation of metallic mercury. Nitrogen gas was bubbled into the electrolyte thus treated for 1 hour. It was noted that during this period, the mercury content of the electrolyte decreases from 0.29 ppm. up to 0.037 ppm It was noted that the removal of mercury is increased by heating the solution while sparging with nitrogen.

In the second test, the initial mercury concentration of 1600 μg / L was reduced to 150 μg / L in 60 minutes and 100 μg / L in 120 minutes. The results are shown in figure 4 below.

     Example 9

Alternative mercury removal

The Eh value of the copper (I) electrolyte from Example 2 of silver removal containing 500 μg / L of mercury was adjusted to approximately -150 mV (Ag / AgCl). This solution was then passed (for example, by pumping under pressure) through a column packed with activated carbon at a speed of 4 ml / min (for the first three hours) and then 20 ml / min (for the next 2 hours). Metallic mercury was adsorbed on activated carbon, thereby reducing its concentration in the solution to lower than 55 μg / L. The metallic mercury was then separated from the activated carbon by passing a second other high Eh solution through the column, thereby converting the metallic mercury into dissolved ionic mercury for passage from the column into the second solution. Mercury was then isolated from the second solution.

The above experimental procedures showed that a very effective removal of both silver and mercury could be achieved without the need for copper (according to the state of the art). In addition, these removal methods are easily performed. The simplicity of the method easily allowed to scale it up to a fully commercial and continuous implementation.

The zinc extraction method of the present invention has the following undoubted advantages in comparison with the method of roasting / leaching / electrowinning of zinc and other hydrometallurgical methods of the prior art:

- significantly lower operating and capital expenses;

- the selection of noble metals in the leaching cycle of zinc (for example, when cleaning non-zinc metals from a solution, when releasing the best-selling metals of copper, lead and silver (Cu, Ag, Pb and Fe in the stage of cementation and deposition of figure 1));

- tolerance to low quality and "contaminated" concentrates;

- low power consumption;

- lack of liquid emission;

- lack of formation of toxic gases;

- mild conditions for operations: low temperature and atmospheric pressure;

- no need for solvent extraction and

- lack of need for pure oxygen in the stages of aeration.

Although the invention has been described with reference to a number of preferred embodiments, it should be understood that the invention can be implemented in many other forms.

Claims (20)

1. A method of extracting zinc metal from a zinc mineral, comprising the following steps: leaching zinc mineral in one step using a solution comprising a halide compound formed from two or more different halides to leach zinc into a solution; electrolyzing a solution containing zinc to obtain a metallic zinc and generating a halide compound, and returning the electrolyzed solution including the halide compound to the leaching step. 2. The method according to claim 1, where the halide compound is formed at the anode in the electrolysis stage at an oxidizing potential lower than the oxidizing potential for the formation of insoluble forms of impurities in solution. 3. The method according to claim 1 or 2, where two or more different halides are chlorine and bromine, and the halide compound is a soluble halide complex, such as

and / or gaseous BrCl. 4. The method according to claim 1 or 2, where the leaching of the mineral with a halide compound is facilitated by a catalyst that catalyzes the oxidation of the mineral, where the catalyst is a metal catalyst, such as copper, which is either present in the zinc mineral or which is introduced into the leaching process. 5. The method according to claim 1 or 2, further comprising one or more subsequent steps for leaching zinc in one step and removing impurities in the subsequent step (s), wherein at least one of the subsequent steps is an aeration step and at least one of the subsequent steps, it is a sulfate precipitation step, where air is introduced in the aeration step to oxidize and precipitate any iron present in the mineral, and limestone is added in the sulfate precipitation step to precipitate the resulting sulfate sulfur oxidation in the leaching step to form calcium sulfate. 6. The method according to claim 1 or 2, where the liquor from the leaching process before electrolysis has gold and platinum group metals present in it, which are removed from it by passing liquor over activated carbon to adsorb gold and platinum group metals on carbon. 7. The method according to claim 1 or 2, where the liquor from the leaching process before electrolysis is passed through a series of cementation processes in which zinc dust is added to the liquor to cement any metal from copper, silver, lead and other impurities present in the liquor, and / or where the liquor from the leaching process before electrolysis is passed to the next iron removal step in which limestone and a halide compound from the electrolysis step are added to oxidize and precipitate iron as iron (III) oxide. 8. The method according to claim 1 or 2, where part of the electrolyzed solution is removed and processed to remove manganese from it, where this part is a stream exiting the cathode compartment of the electrolytic cell of the electrolysis process, the outlet stream having limestone and a halide compound added to it from the anode compartment of the electrolysis process to precipitate manganese dioxide. 9. The method of claim 8, wherein the pH and Eh of the solution are controlled in a manner that favors the formation of a precipitate of manganese dioxide compared to the formation of zinc precipitate, where the pH is controlled by incremental addition of limestone to increase the pH of the solution to a level at which Eh can be increased by a halide compound to a level at or above which Eh favors the formation of MnO ₂ , and where the amount of added limestone is less than the stoichiometric amount required for the formation of MnO ₂ , so less c Inca. 10. The method according to claim 1 or 2, where before returning the electrolyzed liquor to the leaching process, part of the electrolyzed liquor from the cathode compartment of the electrolysis process is passed to the magnesium removal step, in which hydrated lime is added in order to first remove any zinc that is returned to the process leaching, and then remove the magnesium as a precipitate of magnesium oxide. 11. The method according to claim 1, where the liquor from the leaching process before electrolysis is passed to the stage of iron removal, limestone and the halide compound are oxidized and precipitated at least part of the manganese present in the form of manganese dioxide. 12. A method of removing manganese from a solution of a metal halide, which is subjected to electrolysis to obtain one or more metals, comprising the following steps: during electrolysis, the cathodic extraction of one or more metals, at the same time, the anodic formation of a halide compound from two or more different halides with an oxidation potential less than the oxidation potential for the formation of manganese dioxide, removing part of the solution and processing it to remove manganese from it and returning the treated part to the halo solution metal genide. 13. The method according to item 12, where the manganese is removed in a separate stage and where the metal halide solution is a liquor obtained by the leaching process, in which the mineral concentrate is leached in a solution containing halide to leach one or more metals into the solution, and the liquor is directed to the electrolysis stage as an electrolyte and, after electrolysis, is returned to the leaching process for further leaching of the mineral concentrate. 14. The method according to item 12 or 13, where the removal of manganese is carried out as part of a closed loop of the leaching process of the mineral and electrolytic separation, and where the part of the solution that is removed and processed to remove from the solution of manganese is a stream exiting the closed loop leaching process mineral and electrolytic precipitation, which is returned to this process after removal of manganese, and where the leaching of the mineral is facilitated by the halide compound formed at the anode, which is returned with the electrolyte to Processes leaching. 15. The method according to item 12 or 13, where two or more different halides are chlorine and bromine, and the halide compound is a soluble halide complex formed at the anode, such as

16. The method according to item 12 or 13, where the leaching of the mineral is facilitated by a catalyst that catalyzes the oxidation of the mineral with a halide compound, the catalyst being a metal catalyst, such as copper, which is present in the leaching process or introduced into the leaching process. 17. The method according to item 12 or 13, where the mineral ore includes zinc and / or lead, which is the metal (s) formed in the electrolysis stage. 18. The method according to item 12 or 13, where the manganese in this part of the solution is separated from it by the incremental addition of an alkaline reagent such as calcium carbonate, which causes the precipitation of manganese in the form of MnO ₂ , and the precipitate MnO _{2 is} then separated from the part of the solution before how this part is returned to the metal halide solution, and so that when the metal to be isolated also tends to precipitate when an alkaline reagent is added, the alkaline reagent is added in an amount less than the stoichiometric amount to form MnO ₂ , and where a reagent with a high redox potential is also added to a portion of the solution to increase the oxidation potential of a portion of the solution to a level that favors the formation of MnO ₂ , wherein the reagent with a high redox potential is a halide compound or its gaseous form formed at the anode at the stage of electrolysis. 19. The method according to item 12 or 13, where part of the solution is part of catholyte from the electrolysis stage. 20. The metal obtained by the method according to claim 1 or 12.

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