CN114686684A - Method for recovering metal elements from molybdenum concentrate - Google Patents

Abstract

The invention relates to a method for recovering metal elements from molybdenum concentrate, belongs to the technical field of metal smelting, and solves the problem of using ammonia water or sodium hydroxide in the recovery process of molybdenum in the prior art; the process of using high-temperature roasting has high energy consumption, Difficulty in temperature control; the use of inorganic strong acid for leaching at high temperature requires a large amount of acid and high requirements for equipment corrosion resistance; the recovery of valuable metals in molybdenum concentrate is difficult and the process is cumbersome. In the method for recovering metal elements from molybdenum concentrate provided by the invention, the molybdenum concentrate and the leaching aid system are subjected to oxygen autoclaving under the conditions of oxygen pressure and water to realize the leaching of metals in the molybdenum concentrate; the leaching aid is: One or a combination of phosphoric acid and calcium phosphate. The efficient recovery and utilization of various metal elements in molybdenum concentrate is realized.

Description

A kind of method for recovering metal elements from molybdenum concentrate

technical field

The invention relates to the technical field of metal smelting, in particular to a method for recovering metal elements from molybdenum concentrate.

Background technique

Molybdenum concentrate is the ore obtained by flotation of molybdenite. Molybdenum concentrate is a typical sulfide ore, in which many valuable metals such as copper, bismuth, lead and rhenium are associated. However, the smelting process of molybdenum concentrate in the current state of the art mostly adopts the oxidative roasting-ammonia leaching process, and a large amount of sulfur dioxide flue gas is generated during the oxidative roasting process, which cannot effectively recycle the sulfur element in the molybdenum concentrate, and causes a large amount of air pollution. And in the process of oxidative roasting, the very valuable metal rhenium in the molybdenum concentrate is difficult to recover from the sublimation loss. At the same time, it is difficult to control the temperature in the roasting process. If the roasting temperature is too high, the sublimation loss of molybdenum will also be caused. If the roasting temperature is too low, the oxidation of molybdenum will be incomplete and the leaching difficulty will be increased.

In view of the above reasons, technicians began to seek methods for directly processing molybdenum concentrates, namely atmospheric pressure decomposition method and high pressure decomposition method. The atmospheric decomposition method uses strong oxidants such as sodium hypochlorite and potassium permanganate for oxidative leaching. Due to the large use of strong oxidants, the reaction process is difficult to control, and the danger of explosion is easy to occur, which is not convenient for large-scale application. The high-pressure decomposition method, also known as oxygen pressure cooking, is a promising process to directly oxidize and decompose molybdenum concentrate by introducing oxygen under alkaline or neutral high temperature and high pressure conditions.

However, in the process of decomposing molybdenum concentrate by the traditional pressure cooking method, most of the molybdenum forms molybdenum acid precipitation. In turn, the oxidation of molybdenum is affected, and the obtained molybdenum acid precipitation also requires a dissolving process of ammonia water or sodium hydroxide, and a small part of molybdenum remains in the leaching solution at the same time, which requires special molybdenum recovery equipment, resulting in very cumbersome subsequent recovery procedures.

SUMMARY OF THE INVENTION

In view of above-mentioned analysis, the present invention aims to provide a kind of method of reclaiming metal element from molybdenum concentrate, at least one of the following technical problems can be solved: (1) use ammonia water or sodium hydroxide in the recovery process of molybdenum; (2) ) The process of using high temperature roasting has high energy consumption and difficult temperature control; (3) Using inorganic strong acid for leaching at high temperature, the acid consumption is large, and the corrosion resistance of equipment is high; (4) There are valuable metals in molybdenum concentrate Recycling is difficult and the process is cumbersome.

The invention provides a method for recovering metal elements from molybdenum concentrate. Under the conditions of oxygen pressure and water, the molybdenum concentrate and a leaching aid system are subjected to oxygen pressure cooking to realize the leaching of metals in the molybdenum concentrate; The infusion agent is one of phosphoric acid and calcium phosphate or a combination thereof.

Further, the oxygen pressure is 0.8Mpa to 1.5Mpa.

Further, in the molybdenum concentrate and the leaching aid system, the amount of the leaching aid is 0.5 to 1.5 times the mass of the molybdenum concentrate.

Further, the oxygen autoclaving time is 2 hours to 5 hours.

Further, the temperature of the oxygen autoclave is 180°C to 210°C.

Further, after the molybdenum concentrate and the leaching aid system are subjected to oxygen autoclaving, a leaching solution containing molybdenum, rhenium and copper and a leaching residue enriched with bismuth are obtained to realize the separation of bismuth.

Further, the leaching rate of molybdenum is more than 99%, the leaching rate of rhenium is more than 95%, and the leaching rate of copper is more than 98%.

Further, after obtaining the leaching solution containing molybdenum, rhenium and copper, including:

The molybdenum is extracted from the leaching liquid by using an extractant to obtain a raffinate and an extract containing molybdenum to realize the separation of molybdenum.

Further, the extractant is a cationic extractant.

Further, the cationic extractant is one of P204 and P507.

Further, the concentration of the extractant is 10%-50% in terms of mass percentage.

Further, the extraction ratio of extracting molybdenum is O/A=4:1 to 2:1.

Further, the extraction method for extracting molybdenum is countercurrent extraction.

Further, the extraction stages for extracting molybdenum are 4 to 7 stages.

Further, after realizing the separation of molybdenum, including:

The ion exchange resin is used to adsorb rhenium in the raffinate to obtain the exchange residue and realize the separation of rhenium.

Further, the grade of the ion exchange resin is 201.

Further, after the separation of rhenium is realized, it includes:

Soluble sulfide is added to the exchange residue to precipitate and enrich copper ions to achieve copper separation.

Further, the described use of extractant to extract molybdenum from the leachate specifically includes:

Using a cationic extractant to extract molybdenum from the leachate to obtain a raffinate and a molybdenum-loaded cationic extractant;

The molybdenum-containing extract is obtained by back-extracting molybdenum from the molybdenum-loaded cationic extractant using a back-extracting agent.

Further, the stripping agent is hydrogen peroxide.

Further, the mass fraction of the hydrogen peroxide is 10% to 20%.

Further, the extraction ratio of back-extracted molybdenum is O/A=7:1 to 5:1.

Further, the extraction method of back-extracting molybdenum is countercurrent extraction.

Further, the extraction stages of back-extracted molybdenum are 2 to 4 stages.

Further, the raffinate is recovered by adding a neutralizing agent to obtain a leaching aid, which is recycled for oxygen autoclaving and leaching of molybdenum concentrate.

Further, the main components of the molybdenum concentrate are 45% to 55% molybdenum, 1% to 1.5% copper, 0.04% to 0.05% rhenium, and 2.3% to 2.7% bismuth.

Further, the recovery rate of molybdenum is more than 99%, the recovery rate of rhenium is more than 85%, and the recovery rate of copper is more than 90%.

Compared with the prior art, the present invention can achieve at least one of the following beneficial effects:

(1) compared with existing need to use roasting process technology in the processing of molybdenum concentrate, the present invention processes molybdenum concentrate simultaneously by oxygen pressure cooking and digestion leaching, avoids using the roasting process of high temperature and high energy consumption, avoids the use of roasting The effect of high or low temperature on the recovery of molybdenum in the process avoids the loss of rhenium in the roasting process.

(2) compared with existing use inorganic strong acid to carry out leaching technology, molybdenum concentrate of the present invention does not need to consume a large amount of inorganic strong acid through oxygen pressure cooking leaching process, by adding phosphoric acid or calcium phosphate as leaching aid, molybdenum concentrate is oxygenated Pressure cooking to achieve mild leaching and high leaching rate of molybdenum concentrate (the leaching rate of molybdenum is as high as 99%, the leaching rate of rhenium is greater than 95%, and the leaching rate of copper is greater than 98%).

(3) the present invention is to the molybdenum concentrate leaching process, molybdenum is converted into molybdenum acyl cation, rhenium is converted into rhenate ion, copper is converted into copper sulfate and enters the solution, and then the molybdenum, rhenium and copper in the leaching solution are successively carried out. Enrichment and recovery; bismuth is all enriched in the slag for recovery or direct sale, realizing the recovery of molybdenum and three valuable metals of rhenium, bismuth and copper in molybdenum concentrate.

(4) By using neutral phosphine extractant to extract molybdenum, using hydrogen peroxide to carry out back extraction of molybdenum, making full use of the different existence modes of molybdenum in different environments (phosphomolybdenum heteropolyacid, molybdenum acyl cation, peroxymolybdenum Molybdenum is efficiently recovered by simple extraction, avoiding the use of strong alkali and the discharge of ammonia nitrogen wastewater.

In the present invention, the above technical solutions can also be combined with each other to achieve more preferred combination solutions. Additional features and advantages of the invention will be set forth in the description which follows, and some of the advantages may become apparent from the description, or may be learned by practice of the invention. The objectives and other advantages of the invention may be realized and attained by means of the content particularly pointed out in the description.

Detailed ways

The invention provides a method for recovering metal elements from molybdenum concentrate. Under the conditions of oxygen pressure and water, the molybdenum concentrate and a leaching aid system are subjected to oxygen pressure cooking to realize the leaching of molybdenum, rhenium and copper. , bismuth is enriched in the leaching residue; the leaching aid is one of phosphoric acid and calcium phosphate or a combination thereof.

The present invention adopts a treatment method combining oxygen pressure cooking and leaching. Phosphoric acid and/or calcium phosphate are used as leaching aids in the system to assist in leaching and digesting ores. Molybdenum sulfide oxidation, in which the sulfur is oxidized to IV-valent sulfur, and further oxidized to VI-valent sulfuric acid. Under the action of sulfuric acid produced by the oxidation of sulfur in the molybdenum concentrate, the molybdenum in the molybdenum concentrate is converted into soluble molybdenum acyl cations Thus completely dissolved in the liquid phase to form a solution.

After experimental research, it was found that the molybdenum concentrate could not be digested by using the leaching aid phosphoric acid and/or calcium phosphate alone, and the ore could not be oxidized and digested by oxygen pressure cooking alone in a water environment. Therefore, under oxygen conditions, oxygen autoclaving with phosphoric acid and/or calcium phosphate as a leaching aid realizes the digestion and leaching of molybdenum concentrate ore, and realizes the leaching of molybdenum, rhenium and copper. The bismuth will not be leached, but is all enriched in the residual leaching residue, which can be used for recycling or directly sold.

As the oxygen autoclaving continues, under the combined action of oxygen and leaching aids, the molybdenum sulfide in the ore is continuously oxidized and combined with phosphate to convert into phosphomolybdenum heteropolyacid and sulfuric acid, and the acidity gradually increases. In a strongly acidic environment, the reaction balance of the mutual conversion of phosphomolybdic heteropolyacid and molybdate cations is constantly moving towards the direction of generating molybdate cations, and the molybdenum in the ore is dissolved and transformed into a state of molybdate cations with good solubility. Phosphoric acid or calcium phosphate assists the conversion of molybdenum into phosphomolybdenum heteropolyacid in the process of ore digestion and leaching, and is released during the conversion of phosphomolybdenum heteropolyacid into molybdenum acyl cation, which plays a role in the digestion and leaching process of ore. The effect of immersion.

Specifically, the oxygen pressure is 0.8Mpa to 1.5Mpa.

Oxygen is the key factor that plays an oxidizing role in the process of oxygen autoclaving. Since the leaching process of oxygen autoclaving is a heterogeneous reaction of gas-solid-liquid, the general gas participating in the reaction mainly relies on the heterogeneous interface of gas-solid and gas-liquid. On the contrary, the reaction rate of the interfacial reaction involving gas is slow, which seriously affects the efficiency of oxygen autoclaving leaching. Provide enough pressure for oxygen to better dissolve it in the liquid, and a gas-solid-liquid heterogeneous reaction occurs at the liquid-solid heterogeneous reaction interface, which greatly improves the efficiency of oxygen autoclaving leaching .

Specifically, in the molybdenum concentrate and the leaching aid system, the amount of the leaching aid is 0.5 to 1.5 times the mass of the molybdenum concentrate.

Specifically, the pressure cooking time is 2 hours to 5 hours.

Specifically, the leaching oxygen autoclaving temperature is 180°C to 210°C.

The longer the oxygen pressure cooking time is, the more sufficient the reaction will be, but at the same time, too long pressure cooking will cause higher energy consumption, so the pressure cooking time is selected to be 2 hours to 5 hours.

The higher the leaching oxygen autoclaving temperature, the faster the autoclaving rate, but it will also cause higher energy consumption. Therefore, considering the production efficiency and energy consumption of autoclaving, the leaching oxygen autoclaving temperature is determined to be 180°C according to the cost-effectiveness ratio. to 210°C.

Specifically, the leaching rate of molybdenum is more than 99%, the leaching rate of rhenium is more than 95%, and the leaching rate of copper is more than 98%.

Specifically, after oxygen autoclaving the molybdenum concentrate and the leaching aid system, a leaching solution containing molybdenum, rhenium and copper is obtained.

Specifically, after obtaining the leaching solution containing molybdenum, rhenium and copper, it includes:

Using an extractant to extract molybdenum from the leaching solution, obtain a raffinate and an extraction solution containing molybdenum, and realize the separation of molybdenum;

Use ion exchange resin to adsorb the rhenium in the raffinate to obtain the exchange residue and realize the separation of rhenium;

Soluble sulfide is added to the exchange residue to precipitate and enrich copper ions to achieve copper separation.

In a possible embodiment, the grade of the ion exchange resin is 201.

The method for recovering copper is the precipitation method, that is, adding negative divalent sulfur and copper to the exchange residue to form cuprous sulfide precipitation, thereby realizing the enrichment and recovery of copper. Wherein, the negative divalent sulfur is a soluble sulfide; in a possible embodiment, the soluble sulfide includes sodium sulfide, sodium sulfide and other sulfide salts; in another possible embodiment, the soluble sulfide can also include sulfide hydrogen gas.

Specifically, the use of an extractant to extract molybdenum from the leachate specifically includes:

Using a cationic extractant to extract molybdenum from the leachate to obtain a raffinate and a molybdenum-loaded cationic extractant;

The molybdenum-containing extract is obtained by back-extracting molybdenum from the molybdenum-loaded cationic extractant using a back-extracting agent.

Specifically, the cationic extraction agent is one of P204 and P507 or a combination thereof, and the back extraction agent is hydrogen peroxide.

Specifically, the concentration of the extractant is 10%-50% by mass.

Specifically, the extraction ratio of extracting molybdenum is O/A=4:1 to 2:1.

The extraction ratio is an important factor affecting the extraction. When the extraction ratio O/A is less than 2:1, the organic extractability will be insufficient, and part of the molybdenum cannot be transferred to the organic phase. Therefore, from the perspective of the recovery rate of molybdenum , the greater the O/A ratio, the more complete the transfer of molybdenum to the organic phase, and the less residual in the aqueous phase. But adding too much organic phase will also cause the cost of solvent and process to rise. After research, when O/A is greater than 4:1, the extraction efficiency and cost ratio is seriously reduced. Therefore, the extraction of molybdenum is selected to be compared to O/A=4: 1 to 2:1.

Specifically, the extraction method for extracting molybdenum is countercurrent extraction.

Specifically, the extraction stages for extracting molybdenum are 4 to 7 stages.

Specifically, the mass fraction of hydrogen peroxide is 10% to 20%.

Specifically, the extraction ratio of back-extracted molybdenum is O/A=7:1 to 5:1.

Specifically, the extraction method of back-extracting molybdenum is countercurrent extraction.

Specifically, the extraction stages of the back-extracted molybdenum are 2 to 4 stages.

The cationic extractant chemically extracts the leaching solution, and the hydrogen ions in the cationic extractant undergo cation exchange with the leached molybdate cations, and the molybdate cations are transferred to the organic phase.

There is a chemical equilibrium process in which a small amount of molybdate cations in the organic phase enter the aqueous phase and are converted into molybdate ions. Oxymolybdate anion, promotes the chemical balance to move towards the direction of the conversion of molybdate cation into molybdate ion, and converts it all into peroxymolybdic acid and the anion completely separates from the organic phase, so as to realize the removal of molybdenum from the molybdenum-loaded cation extractant. The reverse chemical extraction was carried out, and all of it was stripped back into the aqueous phase of the stripping agent.

Specifically, the raffinate is recovered by adding a neutralizing agent to obtain a leaching aid, which can be used again for oxygen autoclaving and leaching of molybdenum concentrate, and is recycled for oxygen autoclaving and leaching of molybdenum concentrate.

Specifically, the neutralizing agent is one or more of calcium oxide, calcium hydroxide, calcium carbonate, and calcium phosphate.

The neutralizing agent neutralizes the excess hydrogen ions in the system and retains the phosphoric acid or calcium phosphate in the system. Therefore, the leaching aid can be recovered after the neutralizing agent is neutralized and directly used in the leaching oxygen autoclaving process in step 1. Digestion and leaching of ore.

It should be noted that the addition of the neutralizer can also precipitate the enriched impurity iron in the system and remove it from the system to prevent Fe ³⁺ from interfering with the process of cationic extraction of molybdate cations in the system.

Specifically, the main components of the molybdenum concentrate are 45% to 55% molybdenum, 1% to 1.5% copper, 0.04% to 0.05% rhenium, and 2.3% to 2.7% bismuth.

Specifically, the recovery rate of molybdenum is over 99%, the recovery rate of rhenium is over 85%, and the recovery rate of copper is over 90%.

The preferred embodiments of the present invention are specifically described below to illustrate the principles of the present invention, but not to limit the scope of the present invention.

Example 1

A specific embodiment of the present invention discloses a method for recovering metal elements and copper from molybdenum concentrate.

The main components of molybdenum concentrate raw materials are 45.1% molybdenum, 1.3% copper, 0.04% rhenium and 2.59% bismuth.

Using phosphoric acid as the leaching aid, the dosage of the leaching aid is 0.5 times the mass of the molybdenum concentrate, the oxygen partial pressure is 0.8Mpa, and the temperature is 210 ℃.

After the oxygen pressure cooking leaching is completed, it is cooled to room temperature and filtered to obtain a leaching residue containing 16.1% bismuth and a leaching solution containing molybdenum, rhenium and copper.

Among them, the leaching rate of molybdenum is as high as 99%, the leaching rate of rhenium is 95%, and the leaching rate of copper is as high as 98%.

For the leaching solution containing molybdenum, rhenium and copper, molybdenum was first extracted with 50% cationic extractant P204 to obtain molybdenum-loaded P204. The extraction ratio was O/A=4:1, and the extraction stage was five countercurrent stages.

The P204 loaded with molybdenum is back extracted with 15% hydrogen peroxide to obtain back extraction liquid and raffinate; the back extraction ratio is O/A=7:1, and the back extraction stage is three countercurrent.

The total recovery rate of molybdenum in the extraction and stripping process is 99%.

The back extraction liquid is evaporated and crystallized to obtain molybdic acid, and then calcined to obtain molybdenum trioxide product.

201 resin is used to adsorb rhenium in the raffinate after molybdenum extraction, and the recovery rate of rhenium can reach 85%.

After the rhenium is extracted, the residual liquid is exchanged and added with sodium sulfide to precipitate copper to enrich copper and recover copper, and the recovery rate of copper reaches 90%.

Embodiment 2

A specific embodiment of the present invention discloses a method for recovering metal elements and copper from molybdenum concentrate.

The main components of molybdenum concentrate raw materials are 50.2% molybdenum, 1.2% copper, 0.05% rhenium and 2.42% bismuth.

Calcium phosphate is used as leaching aid, and the amount of leaching aid is 1 times the mass of molybdenum concentrate. Under the condition that the oxygen partial pressure is 1.5Mpa and the temperature is 180℃, the leaching is carried out in the autoclave for 2 hours.

After the oxygen autoclave leaching is completed, it is cooled to room temperature and filtered to obtain a leaching residue containing 16% bismuth and a leaching solution containing molybdenum, rhenium and copper.

Among them, the leaching rate of molybdenum is as high as 99%, the leaching rate of rhenium is 95%, and the leaching rate of copper is as high as 98%.

For the leaching solution containing molybdenum, rhenium and copper, molybdenum was first extracted with 40% cationic extractant P204 to obtain molybdenum-loaded P204. The extraction ratio was O/A=2:1, and the extraction stage was five countercurrent stages.

Then use 10% hydrogen peroxide to carry out back extraction on the molybdenum-loaded P204 to obtain back extraction liquid and raffinate;

The total recovery rate of molybdenum in the extraction and stripping process was 99.2%.

The back extraction liquid is evaporated and crystallized to obtain molybdic acid, and then calcined to obtain molybdenum trioxide product.

201 resin was used to adsorb rhenium in the raffinate after molybdenum extraction, and the recovery rate of rhenium could reach 86%.

After rhenium is extracted, the residual liquid is exchanged and passed into hydrogen sulfide to precipitate copper to enrich copper and recover copper, and the recovery rate of copper reaches 91%.

Embodiment 3

A specific embodiment of the present invention discloses a method for recovering metal elements from molybdenum concentrate.

The main components of molybdenum concentrate raw materials are molybdenum 48.2%, copper 1.4%, rhenium 0.045%, bismuth 2.69%.

Using phosphoric acid as the leaching aid, the dosage of the leaching aid is 1.5 times the mass of the molybdenum concentrate, the oxygen partial pressure is 1.0Mpa, and the temperature is 190 ℃.

After the oxygen autoclave leaching is completed, it is cooled to room temperature and filtered to obtain a leaching residue containing 17% bismuth and a leaching solution containing molybdenum, rhenium and copper.

Among them, the leaching rate of molybdenum is as high as 99.1%, the leaching rate of rhenium is 95.5%, and the leaching rate of copper is as high as 98.6%.

For the leaching solution containing molybdenum, rhenium and copper, molybdenum was first extracted with 30% cationic extractant P507 to obtain molybdenum-loaded P507. The extraction ratio was O/A=3:1, and the extraction stage was five countercurrent stages.

The P507 loaded with molybdenum was back-extracted with 15% hydrogen peroxide to obtain back-extraction liquid and raffinate;

The total recovery rate of molybdenum in the extraction and stripping process is 99%.

The back extraction liquid is evaporated and crystallized to obtain molybdic acid, and then calcined to obtain molybdenum trioxide product.

201 resin was used to adsorb rhenium in the raffinate after molybdenum extraction, and the recovery rate of rhenium could reach 85.4%.

After the rhenium is extracted, the residual liquid is exchanged and added with sodium sulfide to precipitate copper to enrich copper and recover copper, and the recovery rate of copper reaches 90.2%.

The above description is only a preferred embodiment of the present invention, but the protection scope of the present invention is not limited to this. Substitutions should be covered within the protection scope of the present invention.

Claims (10)

1. A method for recovering metal elements from molybdenum concentrate is characterized in that under the conditions of oxygen pressure and water, the molybdenum concentrate and a leaching aid system are subjected to oxygen pressure boiling to realize leaching of metals in the molybdenum concentrate; the leaching aid is one or the combination of phosphoric acid and calcium phosphate.

2. The method of recovering metal elements from molybdenum concentrate according to claim 1, wherein the oxygen pressure is 0.8Mpa to 1.5 Mpa.

3. The method of claim 1, wherein the leaching aid is used in an amount of 0.5 to 1.5 times the mass of the molybdenum concentrate.

4. The method of recovering metal elements from molybdenum concentrate according to claim 1, wherein the oxygen autoclaving time is between 2 hours and 5 hours.

5. The method for recovering metal elements from molybdenum concentrate according to claim 1, wherein the separation of bismuth is achieved by subjecting the molybdenum concentrate and the leaching aid system to oxygen pressure digestion to obtain a leaching solution containing molybdenum, rhenium and copper and a leaching residue rich in bismuth.

6. The method for recovering metal elements from molybdenum concentrate according to claim 1, wherein the method comprises the following steps after obtaining the leaching solution containing molybdenum, rhenium and copper:

extracting molybdenum from the leaching solution by using an extracting agent to obtain raffinate and molybdenum-containing extract, thereby realizing the separation of molybdenum.

7. The method of claim 6, wherein the separation of molybdenum is effected by:

and adsorbing rhenium in the raffinate by using ion exchange resin to obtain exchange raffinate, thereby realizing the separation of rhenium.

8. The method of claim 7, wherein the separation of rhenium is achieved by:

adding soluble sulfide into the exchange residual liquid to deposit and enrich copper ions, thereby realizing the separation of copper.

9. The method for recovering metallic elements from molybdenum concentrate according to claim 6, wherein the extracting molybdenum from the leachate with an extractant comprises:

extracting molybdenum from the leaching solution by using a cationic extractant to obtain raffinate and the cationic extractant loaded with molybdenum;

and (3) stripping molybdenum from the molybdenum-loaded cation extractant by using a stripping agent to obtain molybdenum-containing extract.

10. The method of claim 6, wherein the raffinate is recycled to the oxygen poaching and leaching of the molybdenum concentrate after being added to the neutralizer to obtain the leaching aid. .