KR930006088B1 - Hydrometallurgical recovery of metals and elemental sulphur from metallic sulphides - Google Patents

Abstract

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Description

Wet metallurgical recovery of metals and sulfur elements from metal sulfides

1 is a schematic of a good leaching and gas recirculation according to the invention.

The present invention relates to a method of oxidizing and leaching metal sulfides, such as zinc sulfide concentrates, in sulfuric acid solution in the presence of an oxidizing agent consisting of nitrogen oxides. In the present invention, a mixture of nitric acid / sulfuric acid is used.

The present invention relates to the recovery of oxidants, NO / NO ₂ and complete removal of residual oxides (mainly nitrates) of nitrogen from solution. The present invention also relates to a method of sufficiently lowering the nitrate concentration of a solution to be treated for recovery of metal by electrowinning.

Further treating zinc sulfide containing minerals or concentrates to be zinc soluble; Most of sulphide is converted to sulphur to recover relatively pure products; It is also advisable to recover all silver and lead present as commercially available lead / silver residues and to reduce the residual nitrate content in the solution to a level low enough to avoid reproblems in the electrolytic recovery of dissolved zinc. Purpose.

Conventional treatment processes for metal sulfides containing concentrates of zinc and minerals and other precious metals include calcination, leaching, solution purification and electrowinning steps.

During the calcination process, sulfur present as sulfide in minerals or concentrates is converted to sulfur dioxide and for economic and environmental reasons they are usually converted to sulfuric acid. The solid product by calcination is leached in dilute sulfuric acid solution recycled from the electrowinning step, called electrolytic waste.

After the impurity removal step, the solution passes through the electrolytic cell, where zinc metal is electrolyzed and sulfuric acid is recycled to the electrolytic waste and recycled for the next calcination reaction.

The main limitation of the conventional process is that the production of zinc metal is directly related to the production of sulfuric acid. Since it is impossible to store large amounts of sulfuric acid, for example, when used in fertilizer production, many difficulties arise in treating the generated acid value, so the only way is to limit the production of metal in proportion to the acid demand. All. The problem may be solved by developing a method for converting sulfide content in minerals or concentrates into sulfur elements. These products can be stored easily and safely and can be converted to sulfuric acid as needed.

Another disadvantage of the prior art process is that the calcination process, in particular the calcination process in a fluid bed calciner, relies on the supply of very high grade minerals or concentrates. Such materials are rapidly reduced and this trend is expected to continue as higher grade mineral precipitation occurs. As feedstock and quality decrease, more problems arise in the calcination process and in achieving high yield recovery of the contained materials. In one example, in a fluidized bed calciner, difficulties arise due to bed agglomeration upon calcination of zinc concentrates containing at least 5% copper and lead and silica.

Direct leaching of metalsulfide minerals or concentrates has a number of advantages over conventional processes when used in combination with conventional processes. Oxidative direct leaching is a method of dissolving a simple metal sulfide in an inorganic acid in the presence of a suitable oxidant to produce a metal salt solution and a sulfur element.

Many oxidative direct leaching methods have been patented and published, but only a few are commercially available. The main reason is a combination of boundary and process difficulties. Due to the high capital and fair prices, many processes did not attract attention. Wet metallurgy limitations such as metal recovery, impurity recovery, reagent consumption, by-products, energy requirements, and environmental issues have prevented many processes from being industrially accepted.

A Sherrett-Gordon pressurized leaching process involving autoclave treatment of zinc concentrates under oxygen pressure has recently been commercially accepted in two zinc plants (Trail and Timmins) in Canada. However, due to the formation of plumbozasite in autoplates, it is difficult to recover lead / silver residues to an acceptable grade without suspending either (a) additional processing or (b) autoclaving at high pH. The product contains a number of impurities. Another obvious disadvantage is the need for expensive pressurization and corrosion-resistant structural materials for autoclaves.

Nitric acid / sulfuric acid systems have been proposed as one method of direct leaching. In this method, MS and NO ₂ are reacted by the following reaction.

MS + NO ₂ + H ₂ SO ₄ → MSO ₄ + NO + S ⁰ + H ₂ O

3MS + 2HN0 ₃ + 3H ₂ S0 ₄ → 3MO ₄ + 3S ⁰ + 2N0 + 4H ₂ 0

MS + 2HN0 ₃ + H ₂ S0 ₄ → MS0 ₄ + S ⁰ + 2N0 ₂ + 2H ₂ 0

(M = metal, eg Cu, Zn).

The liberated NO can be oxidized by 0 ₂ if necessary to reuse and reabsorb to form HNO ₃ . Thus, the reaction is ideally catalytic to the oxidant. Unfortunately, large amounts of nitrogen dioxide enter the solution in the form of nitrate ions, and conventional methods do not completely recover even in the presence of excess concentrate.

The fact that the residual nitrogen oxides cannot be completely removed from the solution prior to electrolytic immersion and the high price of the oxidant are known to be serious problems. A recent patent by Connecott Copper Corporation of New York (US Pat. Nos. 4,189,461, 4,144,310, 4,132,758) states that nitric acid and nitrogen dioxide are effective oxidants for use in copper concentrates and are affordable. (Nitric-Sulphuric Leach process Improvements, DSDavis, et al, Mining Engineering, 1981, 1252-1258). However, the content of nitrate remaining in the solution after the multi-step leaching remains high at> 3.Og / L.

Ferric ions react with NO ₃ at 160-110 ° and under pressure as

+4H⁺→ 3Fe³⁺+N0+2H₂0.3Fe ²⁺ +

+ 4H ⁺ → 3Fe ³⁺ + N0 + 2H ₂ 0.

Connecoat suggested a method of autoclaving the leaching liquid to promote the reaction. However, the NO ₃ ⁻ content of -500 mg / L continued to remain (Nitric-Sulphuric Leach Process For recovery of copper from concentrate, HMBrennecke, et al, Mining Engineering August 1981, 1259). Additional nitrates were removed during selenium removal processes, including treatment with copper elements, and during the jaw site precipitation, but the final nitrate concentration remained at 40 mg / L. This method is not applicable to the treatment of zinc sulfide containing materials because the above concentration of nitrates has a serious effect on the electrolytic production of zinc metals.

In other methods, about 40% of the supplied oxidant may be lost. An example is the use of nitric acid to treat nickel / copper sulfide concentrates (Shukloc P.0., Et al, Hydrometallurgy 19783 (1) 55-64).

Most recently, detailed literature on the nitric acid treatment of gold-containing sulfide concentrates was published (G. Van Weert, K. J. Fair and J. C. Schneider, CIM Bulletin, Vol. 79 (895), p84, November 1986). A specially designed leaching reactor (J.C. Schneider, Australian patent, AD-A-5389 / 86) accelerates in situ reaction of free NO gasified oxygen and concentrate. The exhaust gas reacts with lime to form calcium nitrate, which is recycled to the reactor. Formed by attack on pyrites and arsenopyrite materials present in sulfate / sulfide-containing minerals other than sulfur elements. In another method of treatment of sulfide concentrates containing silver and copper, small amounts of nitric acid are used to enhance autoclave pressure leaching (JBAckerman and CSBucans, Minerals and Metallurgical Processsing, Feb. 1986. p.20). ).

High concentrations of nitric acid may be used to facilitate the attack on refractory gold and silver, but the reaction is again carried out in the autoclave. All sulfur sulphates are oxidized (R. Raudsepp. E. Peters and M. J. V. Beattie, US Chinese Patent No. 4,647,307, March 3, 1987).

A distinctive drawback to the general pressurized leaching of zinc concentrates is that iron in the fine concentrates rapidly dissolves into the oatclave and then precipitates as a jar site. The presence of ammonium ions or alkali metals in the electrolytic waste liquor used as the repulp solution of the autoclave facilitates precipitation of the jarosite material. As a result. Severe contamination of the lead / silver residues remaining after the dissolution of zinc occurs.

This adverse effect can be minimized by terminating the autoleaching reaction at the final high acidity to keep iron dissolved in solution. This section adds another treatment step to the pressurized leaching process because it precipitates in the next step.

Another problem that occurs during the pressure leaching process is that low sulfur is produced. Although sulfur contamination was somewhat influenced by the concentration of the fed composition, the yield of sulfur produced by Ceri-Gordon / Comminco in the pilot treatment of Canadian concentrates was 14ppm Se, 14ppm Te. 3 ppm As and 22 ppm Hg (E.G. Parker and s. Romanchuck, Proc. 109th AIME Annual Meeting, February 24-28, 1980, p.407).

Because of these impurities, the sulfur produced from the pressure leach process cannot be marketed directly and has no acceptable commercial viability compared to the high grade materials marketed directly.

It is an object of the present invention to overcome the technical difficulties encountered in existing processes and to achieve cost effective processes through recycling and regeneration of oxidants without the use of high capital autoclave, which requires a lot of capital. Another object of the present invention is as follows,

(a) treating zinc sulfide containing raw materials under atmospheric pressure to maximize solubility of the contained zinc and other useful metals such as copper and cadmium; (b) converting the sulfur part of the sulfide into a high purity sulfur element and into a form which can be easily separated from the leach solution and the unhazardous residues; (c) yielding almost all of the lead and silver residues originally contained in the raw material in a form which is readily separated from the sulfur element and all of the parent material and in a commercially available form due to improved lead and silver content; (d) maintain leaching conditions to prevent dissolved iron from settling into insoluble materials such as jarosite, to prevent contamination and degradation of lead- and silver-rich residues; (e) minimizing the loss of oxidant and lowering the nitrate content of the leach solution to an acceptable level, thereby treating it through conventional processes to recover useful metals containing the solution without adverse effects; Another object is to integrate the leaching solution in a suitable position in the existing electrolytic zinc plant circuit after lowering the content of residual nitrate salts without adversely affecting the circuit; (f) contact with oxygen or an oxygen-containing gas and absorb the generated NO ₂ gas into a suitable cleaning solution to produce a nitric acid-containing solution that can be recycled to one or more process steps by maximizing the recovery of NO _x gas generated during the process. It is.

Applicants have found that in the presence of oxidants such as HNO ₃ , NO ₂ gas, N0 / 0 ₂ gas mixtures or nitrates such as NaN0 ₃ , they contain large amounts of lead, a feedstock that is not readily processed by conventional fluid bed calcination methods. It has been found that commercial zinc sulfide concentrates can be effectively oxidized. In addition, the lead and silver present in the original concentrate can be recovered as solid residues, which are commercially available byproducts. Sulfur elements formed by the oxidation of sulfur in sulfides are very high in purity and commercially available and can be separated from lead residues by techniques known in the art.

Applicants have found that carrying out a two step process to leach zinc sulfide concentrates is particularly efficient in maximizing the extraction of zinc and minimizing residual nitrate concentrations. This is accomplished by performing a two stage backflow. High concentration zinc extraction is realized by reacting the partially leached concentrate from the second stage (and, if installed, the third stage) with a hot acid solution with a relatively high concentration of NO ₃ ⁻ ions during the first stage leaching. . After separation of the undissolved residue and the sulfur element, the first stage leach solution containing the high concentration of NO ₃ ^- ions is reacted with the excess fresh concentrate of the second stage to produce NO gas according to the following reaction:

_{^{3ZnS + 2NO 3 - + 8H +}} → 3Zn 2+ + 3S 0 + 4H 2 0 + 2NO

The generated NO gas is mixed with oxygen or oxygen-containing gas and converted to NO ₂ gas and recycled to the first stage leaching.

In such circumstances, it is desirable to add a third leaching step so that the fresh zinc sulfide concentrate reacts with the second leaching solution according to the following reaction:

ZnS + 2Fe ³⁺ → S ⁰ + Zn ²⁺ + 2Fe ²⁺

In this third stage, most of the dissolved iron is converted to the ferrous state. It is preferred that iron continues to be precipitated with getite or hematite according to known methods, but not with jarosite. Thus, whether or not to include a third leaching step in the process of the present invention depends on the method of removing the dissolved iron.

In the case of including the third stage, separation of the residual solid from the second stage leach liquid is unnecessary and the solid mixture from the second and third stages may be separated after the third stage and recycled to the first stage. Preferred leaching and gas recycle process diagrams are shown in FIG.

Applicants have found that extracts with a particularly high zinc content are produced when the first leaching step is carried out continuously using a stirred tank reactor connected in series. General results are described in Examples 1 and 2.

However, when the leaching of the second stage was carried out continuously, even if the zinc sulfide concentrate was present in excess, it was found that the nitrate concentration could not be lower than about 5 g / L (Example 4). This undesirable result is believed to be due to the increased concentration of ferrous ions, which are mixed with nitrate ions to form a stable fetosnitroside complex. As illustrated in Example 3, it can be seen that when the second step is carried out batchwise, a final salt nitrate content of 1 g / L or less is obtained without using excess zinc sulfide concentrate. Applicants therefore used a non-continuous, ie batch or semi-batch process in the second stage leaching. Applicants also conclude that if the acidity of the first stage leaching solution is terminated in the range of 80-100 g H ₂ SO ₄ / L and the temperature is increased to 90-95 ° C. before the concentrate is added batchwise or semi-batch, It was found that nitrate removal was optimized. It is a novel method and very important according to the invention to carry out the second stage leaching in batch or semibatch under special conditions.

Applicants have found that as injecting oxygen gas or air under the stirrer of the first stage reactor, NO gas formed as a by-product of the action of the zinc sulfide concentrate can be oxidized to NO ₂ and reused in the reactor. However, NO generated from the second stage and not reoxidized in the first stage must be returned to the recovery and process by using some means. Cannet is proposed to convert gas into nitric acid or to homogenize or separate oxidation of NO with oxygen and to wash NO _{2 back} to the reactor (Nitric-Sulphuric Leach Process Improvements, DSDavies et al Mining Engineering, 1981, 1252-1258) Applicants note that the first alternative is not effective at the reactor temperature required for the first leaching stage, and the second alternative, including the collection, cooling and reoxidation of large amounts of NO _x gas, is a serious process problem. Found out.

In contrast, it was found that the NO gas can be oxidized to oxygen gas and absorb the generated NO ₂ gas in the existing unfilled spray tower using a zinc plant electrolytic waste solution as a cleaning solution (Example 7). . More than 99% of the NO ₂ gas can be removed from the air stream and converted to a dilute nitric acid solution in the zinc plant electrolytic waste solution.

This result is only possible when there is enough oxygen in the air stream to oxidize all the NO present in the off-gas as well as the additional NO generated during the production of HNO ₃ by the reaction:

3N0 ₂ + H ₂ 0 → 2HNO ₃ + NO

It is very convenient to use an acidic electrolytic waste solution as a carrier for recycled NO ₂ because the electrolytic waste liquid generated during the recovery of zinc by electrolysis acts as a leaching agent of the zinc sulfide concentrate in the first stage reactor. Because it is added.

With the exception of small amounts of gaseous impurities introduced with oxygen, when using commercially available oxygen as a source of oxygen (usually containing 95% or more of 0 ₂ ), there are very few exhaust gases that need to be cleaned to suit environmental conditions. There is an advantage to losing. Final washing of the exhaust gas is easily carried out by a method of discharging the leaching solution of the third stage containing a high concentration of ferrous ions. In addition, a separate source of ferrous sulfate solution can be used for the removal of NO _x gas as described above (W.Kristof and P. Koehler, Ger, Offen. DE 3,406,085, 22 August 1985; S. Boslc et al. , Ind. Eng. Chem. Process Des. Dev., Vol. 24, No. I, 1985).

Although the use of commercially available oxygen as the oxidant of NO gas is most preferred, the process of the present invention is not limited to commercially available oxygen. Therefore, in some cases, it is desirable to use air instead of commercial oxygen, in which case it is preferable to pre-compress NO _x and air before cleaning the NO ₂ gas generated in the spray tower. In this case, the absorption tower needs to be operated above atmospheric pressure.

Residual solids from the first stage leaching contain both sulfur and lead / silver residues. These two components must be separated to be a commercially available product. In Examples 9 (a) and 9 (b), floating and hydrocyclone fractionation methods can be used to recover and separate the two fractions, while higher sulfur elements are produced by filtration and melting of the sulfur concentrate. . The purity of the produced sulfur element is higher than or equal to that of sulfur element produced through pressurized leaching.

Under the optimum process conditions in the second leaching step, the content of nitrate <1 g / L is obtained. However, in order to rule out problems during the electrolytic recovery of zinc, the content of nitrates should be less than about 1 mg / L. It has been found that this can be achieved by using a suitable reducing agent.

In neutral or acidic solutions, reduction occurs only in nitrite ions and is known to have been used to analyze nitrate ions in wastewater (JH Margeson, JCSuggs, MRMidgett, Anal. Chem. 1980, 52,1955-1957-Reduction) of Nitrate to Nitrite with Cadmium), but metal zinc is added in a solution with high ionic strength, such as a leach liquid produced by the process of the present invention. Almost complete reduction of the products ammonia, nitrogen and nitrates is achieved. Although powder carburizing or zinc powder containing zinc or copper-activated cadmium is a very effective reduction controller, zinc manipulation or zinc prills can also be used and are inexpensive. The use of various reducing agents is illustrated in Example 8.

Continuous pilot plant tests for leaching and individual recycle / recovery sections of the process chart were performed. During the process, the concentration of ammonium ions in the total waste liquid should be low and preferably maintained at <0.5 g NH ₃ / L. During reduction of nitrate by the concentrate during the leaching process, nitrite ions, which are produced as intermediates, react with ammonium ions to form nitrogen gas, indicating that nitrogen is unacceptably lost from the system.

Ammonium ion is present in the Risdon zinc plant circuit solution because ammonia is added to other sections of the main plant and precipitates into jarosite. For this purpose, however, ammonium ions can be easily replaced with sodium ions, which is common practice in many zinc plants around the world. The results of the pilot float test process using the deammonia solution are illustrated in Example 10. This example achieves at least 95% zinc extraction and is produced by further treatment of leach residues with acceptable residues of lead and sulfur elements. The residual nitrate content in the leaching solution did not exceed 1.0 g / L. After purification and removal of nitrate by the conventional method, the leaching solution has an electrolysis efficiency of 91%. In another embodiment more than 99.5% zinc extraction was achieved.

In order to minimize the required amount of reducing agent such as zinc powder, it is necessary to neutralize the acid present in the leaching liquid discharged in the second step (if there is a third step). In doing so, iron precipitates, as in the various methods described above, but the main difference is in the form of precipitated iron.

Depending on the form in which iron is precipitated, iron can be precipitated in the form of getite, hematite or jarosite, but the invention is not limited by these particular methods.

In two of the world's circuits, where iron is precipitated with getite or hematite, it is preferred that iron in the solution entering the iron precipitation step is present in the form of ferrous iron. Therefore, it is preferable that this determination includes a third leaching step. The solution discharged in such a step can be neutralized with a suitable neutralizer such as zinc oxide calcine to produce a low acid leaching solution in which most of the iron is present in the ferrous state. It is known to those skilled in the art that such a solution is a suitable source for the gettite or hematite precipitation step.

Another advantage of the two circuits is that, if desired, may include a known arsenic removal step prior to the iron precipitation step.

In addition, in order to precipitate the dissolved iron as a jarosite compound, a third leaching step is unnecessary and preferably includes a preneutralization step of neutralizing excessive acidity using zinc oxide calsine or a neutralizing agent prior to the precipitation step of jarosite. Do.

Even if an iron precipitation method is used, it is preferable to perform the citrate reduction step after the iron removal step. The process of the present invention preferably adds a zinc powder purification section in the electrolytic zinc plant main circuit to the calcining-leaching-electrolyzing step used as the nitrate reduction step.

Although the process of the present invention is particularly preferred for leaching higher zinc sulfide concentrates, the process of the present invention is limited by the specific arrangements and raw materials described above in the treatment of leaching solutions produced in electrolytic zinc plant circuits. no. Applicants have found, by way of example, that the process of the present invention is preferably used for the treatment of bulk concentrates and can also be used for finely divided minerals. In this case, however, the process conditions, the granularity of the feedstock, and the continuous treatment of the leaching solution to recover the metal should be adjusted to maximize recovery and extraction of the useful metals that are suitable for the particular raw material and dissolved.

Most comprehensively, the present invention relates to a process for recovering valuable metals from sulfide containing minerals or concentrates and simultaneously producing high quality residues and sulfur elements containing a large amount of useful metals such as lead, silver and gold. At a temperature below the boiling point of the solution, in the presence of an oxidizing agent containing an oxygen-containing gas and an oxide of nitrogen, the metal sulfide-containing material in the dilute sulfuric acid solution was counter-leached in at least two stages, and the generated gas was washed with the tankhouse electrolytic waste solution under atmospheric pressure. The first stage leaching was carried out by passing through a diluted nitric acid / sulfuric acid solution, and the insoluble residue and the reducing element were separated from the first stage leaching solution, and the first stage leaching solution was semi-batch at a temperature below the boiling point under atmospheric pressure. Contact with fresh concentrate to produce a hemorrhoid solution containing low concentrations of nitrate ions and partially Separating from the metal sulfide-containing minerals or concentrates leached with and then recycling to the first step leaching, the low concentration nitrate-containing solution is treated in a conventional manner to precipitate the dissolved iron, to separate the iron and contact with the metal reducing agent To produce a leaching solution in which the nitrate ions are almost removed, and then the useful metal dissolved in the conventional electrolytic process is recovered.

In a preferred embodiment, the oxidant used in the process of the present invention consists of nitrogen dioxide gas, nitric acid, a mixture of nitric acid and an oxygen containing gas or a salt of nitric acid or a mixture of these oxidants.

The invention may include the following preferred embodiments:

(1) The concentration of sulfuric acid in the tankhouse electrolytic waste liquid passed through the first stage leaching is 50-250 g / L;

/L이다 ;(2) The concentration of oxidant used in the leaching stage is 0.5-120g

/ L;

(3) The first stage leaching solution is partially leached, recycled from the second stage leaching, in a series connected closed stirred tank reactor operating continuously under atmospheric pressure, at a temperature below the boiling point, in the presence of oxidant for 1-12 hours. Contact with minerals or concentrates.

(4) the concentration of sulfuric acid from the first stage leaching is 80-11Og H ₂ S0 ₄ / L and the concentration of nitrate is 20-60g NO ₃ / L;

(5) Oxides of nitrogen gas generated from nitric acid solution added to this stage by injecting oxygen or oxygen-containing gas under the impeller in the first stage leaching reactor, oxidized with nitrogen dioxide and reabsorbed into solution to concentrate the metal sulfide React again with;

(6) The second stage leaching is carried out batchwise according to the following process sequence:

(b) partially fill the closed stirred tank reactor with the first stage leaching solution, but leave the sufficient space in the solution to allow for the formation of bubbles and heat the solution to a temperature above 85 ° C .;

(b) add excess concentrate calculated by various methods over 5-25 minutes, the rate of addition being controlled to maintain the production of gas but avoid the formation of excess foam;

(c) the stirred reaction mixture is maintained at a temperature of 85 ° C. or higher for 10-90 minutes before the reactor is emptied and the above process sequence is repeated;

(7) mixing nitrogen oxide produced from the second stage reactor with oxygen or an oxygen containing gas to produce nitrogen dioxide gas;

(8) Mixing the gas mixture from the first stage immersion reactor with the oxide / oxygen gas mixture of nitrogen derived from the second stage, and contacting the gas mixture with the tankhouse electrolytic waste in a series of spray towers operating in countercurrent operation. Nitrogen dioxide formed in the gas mixture is absorbed into the tankhouse electrolytic waste solution to form a dilute nitric acid / sulfuric acid solution and then passed through a first stage leaching reactor;

(9) separating the solid containing the partially leached metal sulfide from the second leaching solution and recycling the solid to the first leaching;

(10) separating the undissolved solid and the sulfur element from the first stage leaching solution and separating the solid from the sulfur element by known techniques such as flotation and hydrocyclone treatment;

(11) The precipitated iron is separated and precipitated by a known method, and then the leaching solution is contacted with a metallic reducing agent in the presence of the added acid to produce a leaching solution in which nitrate ions are almost freed. Recover useful metals.

In another preferred embodiment, the process of the present invention is carried out in an electrolytic zinc plant electrolytic waste solution containing dilute nitric acid formed by rinsing a nitrogen dioxide gas produced by oxidizing an oxide of nitrogen gas produced from a leaching reactor with a commercial grade oxygen gas. A method of leaching zinc sulfide concentrate, in which the loss of nitric acid is supplemented by the addition of nitrates such as sodium nitrate, and the dissolved iron is precipitated in the form of a jatosite compound to form a leaching solution and treated with zinc powder to nitrate By sufficiently lowering the content of the molten zinc, dissolved zinc can be recovered electrolytically with high efficiency in conventional electrolytic zinc plant circuits, while at the same time commercially available or used in other processes, higher lead / silver containing residues and sulfur elements It relates to a process for generating a.

EXAMPLE

The following describes the method of the present invention based on the examples, but the present invention is not limited thereto.

The concentrate used in this example is a zinc concentrate produced in the Elura mine of Electronic Zinc Company of Australia Limited (E.Z.).

All solutions used in this example are general zinc plant solutions collected at the Risdon plant in E.Z, unless otherwise noted.

Example 1

Stage 1 Leaching

This example relates to a zinc extract and a metal recovery which can be obtained by directly leaching zinc concentrate into a HNO ₃ / H ₂ SO ₄ solution in the first stage leaching.

In a subsequent first stage leaching, the partially leached zinc concentrate was oxidized in acidified electrolytic waste with the addition of HNO ₃ , (43 g / L). Solid load rate 75g / L, 85 ° C. temperature, solution feed rate 2L / hour A residence time of 4 hours in 4 stirred reactors was used. The following results were obtained.

* Back-calculated from the concentrate used in the second step.

Crop analysis was performed in Example 3.

* * 22% conversion of sulfide sulfur to sulfate (SO ₄ ^2- )

* * * No nav extraction into solution is considered.

(The solubility of PbSO ₄ is 0.44 g / L)

From the above results, it can be seen that a very practical lead / silver residue containing 58% Pb is produced by the complete separation of elemental sulfur from the residual solids.

Example 2

Stage 1 Leaching

This example illustrates the general solution composition change that occurs during the first stage leaching of a partially leached zinc sulfide concentrate using a series of four stirred reactors while injecting a commercial oxygen gas under the impeller.

The composition of the starting solution was as follows:

H ₂ SO ₄ ^* 202.4 g / L

Zn 45g / L

Fe ³⁺ 〈5mg / L

43g/L

43g / L

* With HNO ₃

This is a representative sample of the electrolytic waste liquor recovered from the electrolytic section of the electrolytic zinc plant of the EZ Company after being used to adsorb NO ₂ produced by oxidation of the NO gas liberated from the first and second stage reactors.

Leached under the following conditions:

Temperature 85 ℃

Solid load rate 75g / L

8 hours residence time

Total O ₂ Flow Rate 560 mL / min

Solution flow rate 2L / hour

Stirring Speed 800r.p.m.

A 4 × 4 L sized stirred tank reactor, undissolved solids and sulfur elements were separated, and a solution with the following composition was obtained from the fourth stirred tank reactor:

H ₂ SO ₄ ^* 170 g / L

Zn 70g / L

Fe ²⁺ 〈5mg / L

Fe ³⁺ 4 g / L

26g/L

26 g / L

* With HNO ₃

Example 3

Stage 2 Leaching

This example illustrates a second stage leaching where the solution from Example 2 is contacted with fresh concentrate in a batch reactor. The concentrate was added to the reactor over 10 minutes and after 80 minutes the partially leached solid was separated from the solution and circulated to the first stage. The NO / NO ₂ gas generated by the second stage leaching reacted with 0 ₂ and all oxides of nitrogen were completely converted to NO ₂ and then absorbed into the electrolytic waste solution.

Temperature 95 ℃

Solid load rate 100g / L

Stirring Speed 800r.p.m.

One 3L stirred tank reactor.

The analysis of solids and solutions was as follows:

From this example, it can be seen that a nitrate concentration of <1 g / L can be achieved by only partial dissolution with the added zinc sulfide concentrate.

Example 4

Stage 2 Leaching

This example illustrates the second stage leaching carried out continuously.

The following experimental conditions were used:

Temperature 95 ℃

Solid load rate 100g / L

Residence time 12 hours

Total O ₂ Flow Rate 2L / Min

Solution flow rate 500mL / min

Stirring Speed 800r.p.m.

3 x 2L stirred tank reactor in series.

Analysis of solutions and solids from the third reactor showed that the average was:

* With HNO ₃

From this example, the attack of the zinc sulfide concentrate sufficiently takes place. Even if the concentration of residual nitrate is found to be unacceptably high.

Example 5

Third stage leaching

This example illustrates a third stage leaching. The solution of Example 3 was contacted with fresh zinc concentrate to convert ferric ions to ferric ions according to the following formula:

In addition, a small amount of acid was consumed by the following reaction.

Zsn + 2Fe ³⁺ → ZnS0 ₄ + S ⁰ H ₂

The following experimental conditions were used:

Temperature 85-95 ℃

50g / L solid load

Residence time 6 hours

Total air flow 1.5L / min

Solution flow rate 2L / hour

Stirring Speed 800r.p.m.

3 x 4L reactor in series.

The analysis of solids and solutions from the third reactor was as follows:

This example shows that the nitrate concentration was negligibly reduced, the sulfuric acid concentration was slightly reduced and ferric iron was reduced to mostly ferrous ions.

Example 6

Extensive analysis was performed on the second stage leaching, and the two stages were performed batchwise to determine the trend of reduction of trace elements.

Nitrogen oxide and oxygen gas were used as oxidants and injected into a first stage reactor, a 5 L stirred tank reactor. The first step was performed at 85 ° C. for 4 hours using the solid recycled from the second step. The second stage leaching was carried out at 95 ° C. using a first stage leaching filtrate stock with a continuous flow of 150 g / L fresh zinc concentrate for 30 minutes.

The second stage leaching continued for 1 hour before the solids were separated by filtration.

A comprehensive analysis of the leached solution and solids follows:

First stage leaching solids

Second stage leaching solution

Second Stage Leaching Solid

Example 7

Recirculation of oxidant

The off-gas produced in the leaching reactor was simulated using pure NO gas from the cylinder. Commercial oxygen gas was also used.

The experimental setup consisted of a gas premix vessel (5 L) connected to two pore spray guns (20 L and 11 L volumes, respectively) in series. A constant volume (5 L and 3 L, respectively) of the electrolyte was individually recycled on each tower at a rate of 30-5 OL / min. The gas temperature of the premix vessel was 70 ° C. The temperature of the electrolytic waste liquid was 30 degreeC.

Each tower was individually recycled at a rate of oxygen gas (5.5 L / min). The gas temperature of the premix vessel was 70 ° C. The temperature of the electrolytic waste liquid was 30 degreeC.

Oxygen gas (5.5 L / min) and NO gas (7.5 L / min) were premixed and washed in two towers for 1 hour to bring the nitrate content to 238.1 g / L. This content is comparable to the theoretical concentration of 227.8 g / L that can be obtained when complete adsorption occurs within ± 5% experimental error.

Example 8

Removal of Nitrate from Leaching Solution

The leaching solution after removal of iron and other impurities, ie arsenic, etc., should be further purified before electrolysis for recovery of zinc. In addition to nitrate ions, other traces of metallic impurities, Cu and Cd, must also be removed.

(a) Test 1: In the following experiment, zinc powder was added to precipitate Cu in accordance with Australian Patent No. 536,376, and a sufficient amount of zinc powder and acid were added to facilitate removal of nitrate.

Purification was carried out in a 3 × 6 L continuous operating stirred tank reactor at a temperature of 80 ° C. and a solution flow rate of 18 L / hour. Acidity was added by adding an impurity solution containing Cu ²⁺ (320 mg / g), Cd ²⁺ (215 mg / L), Zn ²⁺ (150 g / L) and NO ₃ ⁻ (0.953 g / L) and additional sulfuric acid. Increased to 7.0 g H ₂ SO ₄ / L.

(이 농도는 U.V. 흡착분석 방법을 사용한 검출 하한치 보다 작은 것이다).The zinc powder was injected into the first oven with the solution at a flow rate of 150 g Zn / hour so that the ratio of zinc to nitrate was 8: 1. The analysis of the resultant solution after equilibration and filtration was as follows: 0.6mg Cu / L, 200mg Cd / L and <5mg

(This concentration is less than the lower limit of detection using the UV adsorption assay method).

(b) Test 2: In further experiments, zinc prills charged to the column reactors were found to be effective in removing Cu and Cd as well as nitrate ions.

/L과 7.0g H₂SO₄/L을 함유하는 불순물 용액을 80℃로 예열하고 10kg의 아연 프릴을 함유하는 예열된 4L 플라스틱 칼럼 반응기를 통해 4L/시간으로 계속하여 펌핑하였다. 48시간의 여과후, 다음 조성을 갖는 용액이 얻어졌다 : 2mg Cu/L, 3.4mg Cd/L 및 ＜5mg

/L(이 농도는 U.V. 흡착분석 방법을 사용하여 검출한

의 하한치보다 작은 것이다)Thus, 420mg Cu / L, 210mg cd / L, 1.21g

The impurity solution containing / L and 7.0 g H ₂ SO ₄ / L was preheated to 80 ° C. and pumped continuously at 4 L / hour through a preheated 4 L plastic column reactor containing 10 kg of zinc prills. After 48 hours of filtration, a solution with the following composition was obtained: 2 mg Cu / L, 3.4 mg Cd / L and <5 mg

/ L (this concentration was detected using the UV adsorption

Is less than the lower limit of

(c) Test 3: The change of nitrate removal rate according to zinc powder loading was measured.

의 도시의 초기 기울기로부터 얻어진 유사 1차 속도 상수(C_t는 시간 t에 대한 NO₃ ^-함량, C_o는 t=0일때의 NO₃ ^-함량). 초기의 용액 조성은 : Cu 225mg/L, Cd 300mg/L, Co 35mg/L, Zn 120g/L, NO₃ ^-300mg/L(NaNO₃로 도프됨).* for t

Pseudo-first-order rate constants obtained from the initial slope of the plot of (C _t is the NO ₃ ^- content for time t, and C _o is the NO ₃ ^- content at t = 0). The initial composition of the solution: Cu 225mg / L, Cd 300mg / L, Co 35mg / L, Zn 120g / L, NO 3 - ( being doped with NaNO ₃₎ 300mg / L.

An almost primary increase in rate constant with increasing zinc powder indicates that the rate of reduction depends on the effective area of the zinc powder. Therefore, the efficiency of a system using a high load zinc powder, such as a zinc powder or a prill-filled column, is expected to be very high, as can be seen in Test 2 above.

(d) Test 4: This example shows that cadmium acts as a very strong reducing agent in the presence of copper.

(a) copper-clad cadmium metal

The cadmium metal fill was washed with a zinc plant electrolytic waste solution, washed with water and stirred until the cadmium was blackened in a 10% CuS0 ₄ solution. The alloy of _{(3.5g / L) NaNO 3 (} 3OOmg NO 3 - / L) was added to the plant cell, the feed solution and the resulting dope was reacted for one hour at 85 ℃, the NO ₃ 98.4% ^as-is removed and a small amount of NO ₂ ⁻ was formed.

(b) copper / cadmium carburized

In many zinc plants, two stage zinc powder purification of the leach liquid is carried out to remove metallic impurities prior to electrolysis of the leach liquid. In one of the processes copper and cadmium are removed together and the carburized material is filtered prior to the second step (see Australian Patent Application No. 47932/72). Cu / cd carburates were found to be active reducing agents for nitrates, as demonstrated by the following test results:

* Composition: Cu 0.02mg / L, Cd 0.3mg / L, Co 10mg / L, Zn 120g / L NO 3 - ( being doped with NaNO ₃₎ 300g / L.

* * The stoichiometric amount of 120% zinc powder was added to a solution containing Cu (2.5g / L) and Cd (3.5g / L) and reacted for 1 hour. The resulting precipitate was filtered off and washed and used immediately to minimize oxidation in the air.

The removal rate of nitrates shown in this example, carried out in an acidic or neutral zinc sulfate solution, was totally unexpected in the prior art.

The exact mechanism of nitrate reduction seems complicated, but it is known that ammonia is the product and the reaction is by addition of acid. The reaction therefore takes at least partly the following form:

Nitrite is produced as an intermediate by the reaction of zinc powder or carburizing products with nitrates, the amount of which depends on the reducing agent used. If ammonium ions are present and the solution is acidic, nitrogen gas is produced through the reaction:

(J. H. Dusenbury and R. E. Powell, Reactions of Nitrous Acid 1.Ammonium Nitrite Decomposition, Journal American Chemical Society.Vol. 73, July 1951, p. 3266).

를 형성하고 50%는 N₂(기체)를 형성하는 것으로 나타났다.Nitrogen gas is detected in the gas phase. When acidic nitrate-doped zinc sulfate solution is treated with zinc powder, about 50% of the nitrite

And 50% form N ₂ (gas).

의 반응은 농축물 침출공정과 밀접한 관계가 있다. (하기 실시예 10 참조)NH ₄ ⁺ and

Reaction is closely related to the concentrate leaching process. (See Example 10 below)

Example 9a

Suspension of the First Stage Residual Solid

This practice illustrates a method for separating and recovering lead / silver residual fractions and sulfur elements from the first stage residual solids by flotation.

The washed and dried first step residual solid (225 g) was suspended in 2 L of water (1.5 L) in an experimental floating cell. Suspension was carried out at pH 5 without reagents and with an air inlet rate of 0.5 L / min. A concentrate of sulfur element, overflow from the cell, was collected, filtered and dried to give a lead / silver residue. Some of the concentrated elemental sulfur was melted at 140 ° C. and then vacuum filtered. The filtrate, consisting of purified sulfur, was collected and cooled to solidify. The following analysis was obtained:

Example 9b

Hydrocyclone Classification of Residual Solids

This example illustrates a method for separating and recovering elemental sulfur and lead / silver residues from a first stage solid by hydrocyclone fractionation.

The washed and dried first stage solid (160 g) was suspended in water with a small amount of detergent added and the pulp was introduced into five hydrocyclones in series. Underflow sulfur element concentrate and overflow lead / silver residue from each hydrocyclone were collected and dried. A concentrate of elemental sulfur corresponding to a particle size of about 20-30 μm was melted at 140 ° C. and concentrated in vacuo. The filtrate, consisting of purified sulfur, was collected and dried to solidify. The following analysis results were obtained.

Similar samples of lead / silver residues produced by hydrocyclone fractionation of the residual solids of the first stage were found to contain 55.8% Pb, 0.06% AS and 95 ppm Hg.

Example 10

Continuous Pilot Plant Test Process

A small pilot flange was installed and two tests were conducted, consisting of three leaching steps and a mechanical recirculation / absorption step.

The device was sized to treat 3.2 Kg of zinc concentrate per hour at an electrolyte waste flow rate of 20 L / h. The two leaching steps and the gas absorption step are briefly described as follows.

First stage leaching: 4 × 40 L stirred reactor.

Oxygen injection under the impeller.

Second stage leaching: 2 × 100 L stirred reactor.

(Batch, sequential operation)

Absorption tower: Unfilled 2xO.26cm spray tower equipped with liquid recirculation unit around each tower.

A 250 L vessel was used for the pulp / solution storage between the stages, and a 2 × 10 m 3 vessel was used for the storage of the product solution and the supply of the electrolytic waste solution.

The vessels of the first and second stages were provided with a gas exhaust device connected to the absorption tower, and each vessel was airtight and waterproof around the impeller shaft. Most of the pilot plant vessels were made of FRP with 316L stainless steel interior. The reactor was equipped with a coil through which steam or coolant can circulate.

The absorption tower operates countercurrent to the electrolyte and gas streams. A small fan on the exhaust gas kept a small negative pressure on the tower. The airflow from each reactor was measured and the oxygen inflow was adjusted as needed to maximize the oxidation of nitrogen oxides.

Liquid-solid separation between the first and second stages was carried out using a plate of 0.1mO 3 volume and flame filter press, followed by a 0.3 m diameter "Enviroclear" concentrator.

The pulp and solids were introduced into the reactor using a device designed to cause no gas loss from the system.

General control of the pilot plant was performed by Fox3 computer. Process conditions were selected based on the best results of the batch test (see examples 2, 3 and 4 and reference). Elura zinc concentrate was used as feed for the pilot plant, and electrolytic and synthetic electrolytes purchased from Risdon Zine Plant were used for individual tests.

When the Risdon Zine Plant electrolyte was used for gas chromatography / mass spectral analysis of the gas from the first stage reactor, a large amount of nitrogen gas was present. This was found to be due to the reaction of ammonium ions present in the plant solution with the nitrite produced as an intermediate as a result of the reaction between the concentrate and the leaching solution. Thus up to 20% of the nitrates entering the first stage were thus lost while the ammonium ion content was reduced from 4.3 g / L to about 0.5 g / L.

Next, a synthetic, ammonia-free electrolyte containing zinc sulfate and sulfuric acid was used. The general results from the pilot plant test for five consecutive days are summarized as follows:

Zinc concentrate composition (%): Zn 47.3 S0 ₄ / S 2.9

Fe 11.8

Pb 3.4

S / S 32.3

Composition of electrolytic waste liquid after NOx gas absorption (g / L): Zn 40

H ₂ SO ₄ 169.2

NO ₃ ^- 57.6

Stage 1 emissions:

Stage 2 emissions:

Total metal extract (%): Zn 96.2

Fe 96.3

Recovery rate of S ° from S / S (%): 87.0

Suspension test results of the second stage leach residue

Furtherance (%)

Composition of applied sulfur (%): S ° 99.9 Hg 12ppm

Hg 12ppm Zn 37ppm

Pb 28ppm Pe 19ppm

As <0.0001 Ag <5ppm

Representative product solutions from the pilot plant are mixed with the impure Rosdon plant solution at a ratio of 1: 2 and subjected to a series of purification steps prior to small electrolysis for recovery of zinc metals, followed by iron, nitrate, copper and cadmium. , Cobalt and nickel were removed.

Calcined at 75 ° C. under ventilation to remove the first ferric and ferric ions from the solution first, the first ferrous and ferric ions were removed from the solution, resulting in pH ₂₅ of 5.5 and the total content of iron in the solution was 10 Mg / L.

(＜2mg/L)을 함유하는 최종생성을 용액을 수득하였다.Removal of nitrates and other impurity metals is basically carried out in two steps using the EZ Zinc powder refining process (Australian Patent No. 536,376), but by increasing the zinc powder content in the first step most copper as well as Removed. Additional sulfuric acid was added to keep the pH ₂₅ of the reaction at 4-5 during the reaction. More than 98% of the copper precipitated and 97.9% of the nitrates degraded. After filtration, cobalt and cadmium were precipitated by the standard second step method, and Cu (0.09 mg / L), Cd (<2 mg / L), Co (0.18 mg / L), Ni (0.3 mg / L) and

A final product containing (<2 mg / L) was obtained with a solution.

Electrolysis of 100 L of this solution resulted in satisfactory zinc deposition at a current efficiency of 91%.

Claims (19)

Recovers useful metals from metal sulfide-containing minerals or concentrates, while producing high-purity residues and high-purity sulfur elements such as lead, silver and gold, and dissolving them such as dissolved iron and copper, cadmium and zinc A method of recovering a dissolved useful metal by treating a leaching solution containing a metal with sedimentation of dissolved iron and other impurities, separating the precipitated iron from the precipitated iron, and a known method including an electrolytic means; (a) containing a low concentration of nitrate ions by leaching a metal sulfide-containing mineral or concentrate in a dilute sulfuric acid solution in the presence of an oxidant containing one or more nitrogen oxides and an oxygen-containing gas at temperatures below the atmospheric boiling point; To produce a leaching solution; (b) separating insoluble residue and sulfur element from the leaching solution, treating the leaching solution to precipitate iron, and then removing the precipitate; (c) contacting the treated solution with a metal reducing agent in the presence of the added acid to produce a solution that is substantially free of nitrate salts and then separating all the carburized formed; (d) recovering the dissolved useful metal from the solution produced in step (c). The method of claim 1, wherein step (a) is counterflowed in at least two stages and the partially leached solid from the second and subsequent stages is recycled to the first stage. The method of claim 2 comprising collecting a gas containing an oxide of nitrogen generated from the leaching step and mixing it with an oxygen containing gas to promote the formation of nitrogen dioxide. 4. A process according to claim 2 or 3, wherein the first stage leachate containing the high concentration of nitrate ions is reacted with excess fresh metal sulfide containing minerals or concentrates in batch or semi-batch in the second stage. The method according to any one of claims 1 to 3, wherein the oxidant is nitrogen dioxide gas; Nitric acid; A salt of nitric acid, one of a mixture of nitrogen oxide gas and another oxygen containing gas, or a mixture thereof. The process according to any one of claims 1 to 3, wherein the concentration of sulfuric acid in the first leaching solution is 50-25Og H ₂ S0 ₄ / L. The process according to any one of claims 1 to 3, wherein the concentration of nitric acid in the first leaching solution is 10-120 g HNO ₃ / L. The oxidizing agent according to any one of claims 1 to 3, wherein the partially leached solid recycled from the second and subsequent steps is subjected to the oxidant for 2 to 12 hours at a temperature in the range of 80 ° C. to the boiling point under atmospheric pressure at the first step leaching. Leaching in sulfuric acid solution in the presence of. The process according to claim 1, wherein the solution exiting the first leaching step contains 80-11Og H ₂ SO ₄ / L and 20-60g HN0 ₃ / L. 4. The process according to any one of claims 1 to 3, wherein the oxygen or oxygen containing gas required to oxidize the oxide of nitrogen produced in the first leaching stage is added via a sparger below the impeller in the first leaching reactor. The first stage leaching solution according to any one of claims 1 to 3, wherein the second stage leaching is preheated to a temperature of 85 ° C. or higher at a rate controlled to maintain rapid generation of gas while excluding foam formation. Performing the analysis or semi-batch by adding the concentrate excess to, followed by maintaining the slurry in the suspension at 85 ° C. or higher for a sufficient time such that nitrate ions are almost completely removed. 4. A process according to any one of claims 1 to 3, comprising the step of admixing a gas containing an oxide of nitrogen and an oxygen or oxygen containing gas resulting from the second stage leaching to promote the formation of nitrogen dioxide gas. The process of claim 1, wherein the nitrogen oxide / oxygen gas mixture resulting from the second stage ash leaching is mixed with the gas mixture resulting from the first stage leaching and the gas mixture is operated in countercurrent. Contacting with sulfuric acid solution in a series of spray towers to adsorb nitrogen dioxide formed in the gas mixture to the sulfuric acid to produce a dilute nitrogen / sulfuric acid containing solution used as the first stage leaching solution. The process according to any one of claims 1 to 3, wherein the leaching solution discharged in the second or subsequent steps is treated in a known manner to precipitate the dissolved iron and to separate it from the precipitated iron in one or more steps. A method in which dissolved useful metal is recovered by conventional electrolytic means without adverse effects due to the presence of nitrate ions in solution by contacting with a metal reducing agent in the presence of an added acid to sufficiently lower the concentration of nitrate ions. The method of any one of claims 1 to 3, wherein the metal reducing agent is a powder, dust, particles, or prill. High purity zinc metal in dross or large particle form; Zinc dust carburizing products containing any one or a mixture of zinc alloys of aluminum, lead, magnesium or cadmium, copper, nickel, cobalt and cadmium in the above-described form; And one or a mixture thereof selected from the group consisting of cadmium metal in the presence of copper in the alloy or mixture state. The method according to claim 1, wherein the reducing agent is contacted with the leaching solution at a temperature of 20 ° C.-boiling point under atmospheric pressure for a time of up to 2 hours. The method according to any one of claims 1 to 3. Non-hazardous material mixed with elemental sulfur is separated from the first stage leaching solution to produce high purity elemental sulfur and commercial grade lead / silver residues by selective separation methods such as suspended or hydrocyclones. The method according to any one of claims 1 to 3, wherein the sulfuric acid solution is a tankhouse electrolytic waste solution of an electrolytic zinc plant circuit. Zinc plant recovers zinc from zinc sulfide and simultaneously produces high purity sulfur and commercial grade high concentration lead / silver residues A method of producing an electrolytic waste liquid, the method comprising: (a) recycling the concentrate to a first stage while recycling the partially leached solid from the second and subsequent steps to a temperature of 80 ° C. to a boiling point in the nitric acid-containing electrolytic waste liquid; Backflow leaching in at least two stages of operation, the nitric acid is formed by scrubbing a gas containing a mixture of nitrogen oxides and oxygen or oxygen-containing gas generated from the leaching step and some of the oxygen or oxygen-containing gas is passed to the sparger. Is added to the first leaching step and used as the leaching agent of the first leaching step; (b) separating the residue and sulfur element from the first stage leaching solution and reacting the first stage solution containing a sufficient concentration of nitrate ion with an excess of zinc sulfide concentrate at a temperature between 85 ° C. and boiling point in a batchwise manner. Producing a leaching solution containing low concentrations of nitrate ions; (C) separating the partially leached solid, precipitating the dissolved iron and then separating the iron-containing precipitate; (d) contacting the leaching solution containing less than 1 g / L nitrate ion with a metal reducing agent such as zinc dust in one or more steps in the presence of the added acid to produce a solution that is substantially free of nitrates, from which Recovering zinc by pirates.

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