US20060182674A1

US20060182674A1 - Reduction of copper content in the molybdenite concentrate

Info

Publication number: US20060182674A1
Application number: US11/331,406
Authority: US
Inventors: Javier Jara; Sylvester Zuttah
Original assignee: Individual
Current assignee: Individual
Priority date: 2005-02-02
Filing date: 2006-01-12
Publication date: 2006-08-17
Also published as: WO2006082484A2; US7794677B2; WO2006082484A3; US20100098605A1

Abstract

Methods and systems for removing copper minerals from a molybdenite concentrate. One embodiment provides leaching copper from the molybdenite concentrate with a leaching solution comprising ferric chloride, removing molybdenite from the leaching solution, introducing an acid into the leaching solution and introducing O₂, O₃, or a combination of both, into the leaching solution. A method for regenerating ferric chloride in a leaching solution is also provided. One embodiment provides adding a leaching solution comprising Fe(II) ions, Fe(III) ions, or a combination of both, to a mixture of mineral sulfides, introducing an acid into the leaching solution, and introducing O₂, O₃, or a combination of both, into the leaching solution.

Description

CROSS-REFERENCE TO RELATED APPLICATIONS

This application claims benefit of U.S. Provisional Patent Application Ser. No. 60/649,463, filed Feb. 2, 2005, entitled “Reduction of Copper Content in the Molybdenite Concentrate” which is herein incorporated by reference.

BACKGROUND

1. Field of the Invention
Embodiments of the present invention generally relate to a method for mineral purification and more particularly to a method of removing metal sulfides from a molybdenite concentrate.
2. Description of the Related Art
Copper ore deposits containing copper sulfide minerals, such as chalcopyrite (CuFeS₂), chalcocite (Cu₂S), and bornite (Cu₅FeS₄) may contain minor amounts of molybdenite (MoS₂). Recovery of the valuable molybdenite locked up in the ore is usually performed by a milling operation, followed by several flotation steps. The final molybdenite concentrate usually contains some sulfide minerals, and, to be commercial, the copper sulfide mineral content is typically reduced through a leaching step in which the copper sulfide minerals are dissolved by a leaching solution.
The leaching step is typically performed in a batch operation where the molybdenite concentrate is exposed to the leaching solution in a leaching vessel. After the leaching process, the leaching solution is separated from the molybdenite and is regenerated using chlorine gas. Due to the hazardous nature of the chlorine gas a batch operation with several safety procedures is required during the regenerating process, resulting in high labor, handling, and safety costs.
Accordingly, given the high costs associated with using chlorine gas, the batch operation using chlorine gas is suited to produce small amounts of material. For large amounts of products a continuous mode is normally more economical. In a continuous mode operation, leaching can be performed uninterrupted because the leaching solution is replenished as it is being used. Thus, in a continuous mode operation more molybdenite concentrate can be leached in the same amount of time it takes for leaching in a batch mode operation.
Therefore, a need exists for a method of regenerating a leaching solution for use in a leaching process that is more cost efficient, safer, and can be utilized in a continuous process as well as in a batch process.

SUMMARY

The embodiments of the present invention generally provide a method for removing copper minerals from a molybdenite concentrate. One embodiment of the invention provides a method for removing copper sulfides from a molybdenite concentrate by leaching the copper sulfides from the molybdenite concentrate with a leaching solution comprising ferric chloride, removing molybdenite from the leaching solution, introducing an acid into the leaching solution and introducing O₂, O₃, or a combination of both, into the leaching solution.
Another embodiment of the invention provides for obtaining commercial grade molybdenite from a copper ore. The method includes separating a molybdenite concentrate from the copper ore, leaching copper from the molybdenite concentrate with a leaching solution comprising ferric chloride, removing molybdenite from the leaching solution, introducing an acid into the leaching solution and introducing O₂, O₃, or a combination of both, into the leaching solution.
Further embodiments of the invention provide a method for removing copper minerals from a molybdenite concentrate. An exemplary method includes pumping molybdenite concentrate into an autoclave vessel, introducing a solution of Fe(II) ions, Fe(III) ions, or a combination of both, into the autoclave vessel, introducing an acid into the autoclave vessel, introducing O₂, O₃or a combination of both, into the autoclave vessel, and filtering the molybdenite from a stream exiting the autoclave vessel.
In another embodiment, the invention further provides a method for regenerating ferric chloride in a leaching solution. An exemplary method includes adding a leaching solution comprising Fe(II) ions, Fe(III) ions, or a combination of both, to a mixture of mineral sulfides, and introducing an acid and O₂, O₃or a combination of both, into the leaching solution.

BRIEF DESCRIPTION OF THE DRAWINGS

So that the manner in which the above recited features of the present invention can be understood in detail, a more particular description of the invention, briefly summarized above, may be had by reference to embodiments, some of which are illustrated in the appended drawings. It is to be noted, however, that the appended drawings illustrate only typical embodiments of this invention and are therefore not to be considered limiting of its scope, for the invention may admit to other equally effective embodiments.
FIG. 1 is a block diagram for the process of reducing copper content in a molybdenite concentrate at atmospheric pressure.
FIG. 2 is a block diagram for the process of reducing copper content in a molybdenite concentrate under pressure.

DESCRIPTION OF THE PREFERRED EMBODIMENTS

FIG. 1 is a block diagram of a system 100 for carrying out a first process, according to one embodiment of the invention. The system 100 includes introducing a molybdenite concentrate into a dissolution vessel 110. The molybdenite concentrate may be stored in a storage tank 120, and typically includes 3-4% w/w copper sulfide minerals, such as, chalcopyrite, chalcocite, bornite, etc. The dissolution vessel 110 is made from a material which will not 35 dissolve or etch in the conditions used during the dissolution process. In one embodiment, the material is glass. A solution of hydrochloric acid is introduced into the dissolution vessel through inlet 140. The concentration of hydrochloric acid is kept between about 0.7 M and about 4.0 M, and more preferably at about 4.0 M, throughout the dissolution process. A stream comprising ferrous chloride, ferric chloride, or a combination of the two, is introduced to the dissolution vessel through inlet 190. Oxygen, ozone, or a combination of the two, is introduced through inlet 130 into the slurry of the dissolution vessel so that gas bubbles are formed in the slurry and solution. The dissolution vessel is kept at temperatures above about 90° C., and more preferably between about 100° C. and about 120° C. The slurry is agitated by stirring methods, such as mechanical agitators which may include a motor, a shaft and an impeller. Depending on the operating parameters and the mineralogical copper species, the leach process is completed after about 10 minutes to about 120 minutes. A stream of the slurry, which has been leached, is then filtered at filter 150 and the filter cake is rinsed with hot water. The water used to rinse the slurry is heated to between about 60° C. and about 100° C., and preferably to about 60° C. The water used for rinsing may also be acidic. The solid separated from the filtrate is dried to a moisture content of less than about 5% w/w and contains molybdenite with a copper content of less than about 0.2% w/w. The filtrate comprising ferrous chloride, ferric chloride, or a combination of both, acid, and dissolved cupric chloride then go through a copper removal process 160 where the copper ions are precipitated as elemental copper out of solution by using iron scrap as a reductant, as shown in Equation 1:
Fe+2Cu²⁺=Fe²⁺+Cu Equation 1.
After decopperization, the filtrate contains a higher concentration of iron ions than may be desirable, and thus the filtrate may then go through an optional iron removal process 170 to keep the iron ion concentration around 100 g/L. One way to remove excess iron ions is by reducing the temperature of the filtrate which will decrease the solubility of the iron ions causing precipitation of excess iron chlorides. The filtrate is then heated in preheater 180, and reintroduced into the dissolution vessel 110 through inlet 190.
FIG. 1 describes the continuous process at atmospheric pressure. An alternative embodiment of FIG. 1 is a batch process. In a batch process oxygen, ozone, or a combination of the two, and hydrochloric acid are added to a predetermined volume of acid solution containing ferrous chloride, ferric chloride, or a combination of the two, in dissolution vessel 110. When the ferric chloride concentration reaches a desired level (between about 20 g/L and 100 g/L) molybdenite concentrate is added to vessel 110. During dissolution of copper, only hydrochloric acid is added to vessel 110 in order to maintain an acidity between 1 to 4 M, while ferric chloride concentration decreases with time due to production of ferrous chloride.
In yet an alternative embodiment of FIG. 1, the O₂/O₃and hydrochloric acid are added to the stream of ferrous chloride, ferric chloride, or a combination of the two, prior to entering the dissolution vessel 110.
FIG. 2 is a block diagram of a system 200 for carrying out a second process at pressures higher than atmospheric pressure, according to another embodiment of the invention. System 200 includes many of the same steps as system 100, and identical elements are numbered as they are in FIG. 1. Molybdenite concentrate in the form of an aqueous slurry is introduced from the storage tank 120 through a high pressure pump 205 into an autoclave 210. Oxygen, ozone, or a combination of the two, hydrochloric acid, and a solution of ferrous chloride, ferric chloride, or a combination of the two, are introduced into the autoclave 210 through inlets 130, 140, and 190, respectively, as in system 100. The conditions in autoclave 210 are similar to the conditions of dissolution vessel 110 of system 100. However, in system 200, oxygen, ozone, or a combination of the two, is introduced into autoclave 210 to elevate the pressure in the autoclave. The internal pressure of the autoclave is elevated to about 7 bar gauge compared to atmospheric pressure; however, other pressures are also contemplated, such as 20 to 30 bar. Additionally, autoclave 210 has an outlet 215 for the controlled removal of excess gas. A stream of the slurry which has been leached is then discharged into flash vessel 220 where the pressure of the slurry is reduced to atmospheric pressure, and part of the water evaporates as steam. The steam may be used to heat pre heater 280. The slurry and solution, at atmospheric pressure and about 50° C., are then filtered by filter 150. The solid separated from the solution is dried to a moisture content of less than about 5% w/w and contains molybdenite with a copper content of less than about 0.2% w/w. The filtrate comprising ferrous chloride, ferric chloride, or a combination of both, acid, and cupric chloride then go through a copper removal process 160 where copper is precipitated out of solution. The decopperized solution may then go through the optional iron removal process 170 before the solution is heated in pre heater 280. The filtration is then reintroduced into the autoclave vessel 210 through inlet 190.
An alternative embodiment of the process of FIG. 2 is a batch process wherein oxygen, ozone, or a combination of the two, and hydrochloric acid are added to a predetermined volume of acid solution containing ferrous chloride, ferric chloride, or a combination of the two, in autoclave 210. When the ferric chloride concentration reaches a desired level (between about 20 g/L and 100 g/L) molybdenite concentrate is added to autoclave 210. During dissolution of copper, only hydrochloric acid is added to autoclave 210 in order to maintain an acidity between 1 to 4 M, while ferric chloride concentration decreases with time due to production of ferrous chloride.
In yet an alternative embodiment of FIG. 2, the O₂/O₃and acid are added to the stream of ferrous chloride, ferric chloride, or a combination of the two, prior to entering the autoclave 210.
The first and second processes carried out in the systems 100 and 200, respectively, take advantage of the fact that Fe(III) in a solution of ferric chloride (FeCl₃) will dissolve copper containing sulfide minerals, such as chalcopyrite and bornite. The following equations show the copper dissolution of chalcopyrite (Equation 2) and bornite (Equation 3) in the presence of ferric chloride:
CuFeS₂+4FeCl₃=CuCl₂+5FeCl₂+2S Equation 2
Cu₅FeS₄+12FeCl₃=5CuCl₂+13FeCl₂+4S Equation 3
From the equations above it is seen that when ferric chloride reacts with the iron/copper sulfides, the Fe(III) of ferric chloride is reduced to Fe(II) (ferrous chloride). Ferrous chloride is not a strong enough oxidizer to dissolve the copper containing sulfide minerals, and regeneration of ferric chloride from ferrous chloride must take place for there to be any further leaching of the copper containing sulfide minerals.
By providing acid and an oxygen-containing gas, such as oxygen and/or ozone, to the iron chloride solution, ferrous chloride is oxidized to ferric chloride which can again be used to leach copper containing sulfides from the molybdenite concentrate. The oxidation of Fe²⁺ can be described by the following equations:
4Fe²⁺+O₂+4H⁺=4Fe³⁺+2H₂O Equation 4
2Fe²⁺+O₃+2H⁺=2Fe³⁺+H₂O+O₂ Equation 5
However, due to the high concentration of HCl and iron ions, significant changes in free acid, iron complexes and water activity make it difficult to follow the stoichiometry of the reactions. Thus, Equations 4 and 5 above represent an example of the stoichiometry that might occur, and not every stoichiometric possibility of the high acid concentration reactions.
Iron Ion Oxidation Employing Oxygen and/or Oxygen
The effectiveness of using oxygen or ozone with hydrochloric acid to oxidize ferrous chloride to ferric chloride is tested in several experiments in solutions of ferrous chloride (100 g/L), copper (0 to 10 g/L), and hydrochloric acid (0.7 to 4 M). As this test is for the determination of the feasibility of oxidizing ferrous chloride to ferric chloride, this experiment is performed in the absence of molybdenite concentrate. Oxygen or ozone is introduced into the solution through a glass fritted bubbler, and the oxidation rate is obtained at one hour during which time the unit gas consumption is measured. Table 1 shows the effect of the level of agitation and oxygen gas flow on the rate of iron oxidation and the corresponding consumption of gas.

TABLE 1

Effect of oxygen flow and level of agitation

at atmospheric pressure and 4 M HCl.

Oxygen Oxidation

Flow Temperature Rate Consumption

Test L/min ° C. Agitation g Fe³⁺/L/h g O₂/g Fe

15 0.2 72 600 rpm 28 0.65

16 0.4 72 600 rpm 39 0.84

6 1.0 82 Magnetic 69 3.34
Agitation is provided by either a mechanical agitator which includes a motor, a shaft and an impeller providing agitation at 600 rmp, or a magnetic stirrer which creates a lower level of agitation than the mechanical agitator. It can be seen that upon increase in oxygen flow, the oxidation rate also increases, even if the level of agitation is significantly reduced as when agitation is performed by a magnetic stirrer. The oxidation rates in the range of 28 to 69 g/L/h observed using oxygen and hydrochloric acid are significantly higher than the oxidation rates obtained using chlorine gas which are typically about 15 g/L/h.
The oxidation rate of Fe(II) to Fe(III) is increased when the oxygen pressure in the reaction vessel is increased. The effect of oxygen pressure on the oxidation rate is presented in Table 2. For the high pressure reaction, an oxygen pressure regulator is fixed to maintain an oxygen pressure of about 7 bar. The reaction vessel has a small opening at the exit valve in order to release excess pressure. Attached to the exit valve is a wet meter which measures the exhaust gas flow as the gas exits the reaction vessel. The exhaust gas flow measured is not uniform, indicating that the oxygen is introduced to the reaction vessel in pulses. In Table 2, it can be seen that the oxidation rate increases three to four times when the operating pressure is 7 bar gauge compared to atmospheric pressure, while the gas consumption is three to four times lower.

TABLE 2

Effect of pressure on oxidation rates and oxygen consumption

Pressure Oxygen Flow Oxidation Rate Consumption

Test Gauge Bar L/min g Fe³⁺/L/h g O₂/g Fe

14 7 By pulses 100 0.22

15 0 0.2 28 0.65

16 0 0.4 39 0.84
Additionally, the iron oxidation rate as a function of molar concentration of hydrochloric acid using ozone or oxygen is presented in Table 3. For the same rate of consumption of gas, the iron oxidation rates increase with increased HCl concentration, meaning that the gas is a more efficient oxidizer at higher HCl concentrations. Additionally, the presence of copper increases the oxidation rate by 33%, indicating that the presence of copper is a catalyst for iron oxidation. The odidation potential of ozone gas is higher than the oxidation potential of oxygen gas, and as seen in Table 3, the oxidation rate of iron is higher when using ozone gas instead of oxygen gas under the same conditions.

TABLE 3

Effect of HCl concentration and presence of copper

HCl Cu, Oxidation Rate, Consumption

Test Molarity Gas g/L g Fe³⁺/L/h g O₂/g Fe

21 0.7 O₃ 0 35 3.4

22 2 O₃ 0 47 3.4

23 4 O₃ 0 54 3.4

26 4 O₃ 10 72 3.4

24 4 O₂ 0 30 3.4

27 4 O₂ 10 56 3.4

Reduction of Copper Content in a Molybdenite Concentrate

The effectiveness of leaching copper from a molybdenite concentrate using ferric chloride is tested in several experiments using solutions of ferric chloride. The leaching is performed on a disk filter cake sample containing 3.2% w/w Cu, 1.7% w/w Fe, and 49.2% w/w Mo which is fed to the copper leaching reactors. A ferric chloride solution is added to the leaching reactor. The ferric chloride solution is prepared by oxidizing ferrous chloride to ferric chloride in the presence of oxygen. The molybdenite concentrate is agitated in the ferric chloride solution at atmospheric pressure and at 100° C. No oxygen is injected during these leaching tests. Table 4 shows the experimental conditions and results for several leaching times followed by filtration at 60° C. To assure an excess of Fe(²⁺) during the leaching process, the concentration of Fe(²⁺) is about 90 g/L.

TABLE 4


Dissolution of Copper Sulfide Followed by Filtration at 60° C.

		Initial	Final
		solution	solution	Residue	kg Fe(³⁺)
Time	%	g/L	g/L	%	per kg

Test

min

solids

Fe(³⁺)

Fe(²⁺)

Cu

Fe(³⁺)

Fe(²⁺)

Cu

Fe

Mo

concentrate

LHV-24	15	40	85.5	14.5	1.74	30	55	15.7	0.21	0.59	52.6	0.082
LHV-25	30	40	85.5	14.5	1.74	34	67	19	0.13	0.53	52.7	0.076
LHV-23	30	20	95	12	0.4	71.7	32.9	8.8	0.22	0.46	53.1	0.093
LHV-26	45	40	85.5	14.5	1.74	31.9	68.1	19.5	0.09	0.42	52.6	0.079
LHV-27	60	40	85.5	14.5	1.74	34	63	18.7	0.09	0.42	53.0	0.076
LHV-28	90	40	85.5	14.5	1.74	25.2	65.8	17.7	0.05	0.40	53.1	0.089
LHV-22	120	20	95	12	0.4	66.1	37.1	9.1	0.02	0.29	53.4	0.116

The residue left after leaching for 15 minutes has a copper content of 0.2% w/w after filtration, and the copper content continues to decrease to 0.05% w/w as the leaching time increases. The mass ratio of consumed ferric chloride to initial concentrate is almost constant at about 0.08 for solid concentrations of 40% w/w. The mass ratio is increased to about 0.09 to about 0.12 for solid concentrations of 20% w/w. The occurrence of a constant mass ratio of consumed ferric chloride to initial concentrate indicates that when the copper sulfide reaction is almost complete (when the concentration of copper is below 0.2% w/w), the consumption of ferric chloride is negligible.

Because the solubility of copper and iron ions decreases as water solvent temperature decreases and the solvent pH increases, the effect of temperature and acidity of the water used to rinse the residue after leaching is evaluated in Table 5. After filtration, the residue is divided into halves, and each half is rinsed with either tap water at 60° C. or with acidic water at 100° C. The higher temperature rinse results in a slightly higher dissolution of copper and iron compounds than at the lower temperature rinse. The filtration and rinse will often be performed at about 60° C. due to the nature of the filtration material.

TABLE 5


The Effect Temperature and Acidity of Rinsing Water in Filtration

	Filtration using	Filtration using 0.5 L tap water
	0.5 L tap water at 60° C.	at 100° C., pH = 2.0

		Rinsed		Rinsed
Leaching	Residue	solution	Residue	solution
Time	% w/w	g/L	% w/w	g/L

Test	Min	Cu	Fe	Mo	Cu	Fe	Cu	Fe	Mo	Cu	Fe

LHV-24	15	0.211	0.59	52.62	0.279	1.22	0.202	0.41	52.9	0.358	1.80
LHV-25	30	0.131	0.53	52.67	0.24	0.98	0.12	0.31	52.87	0.288	1.48
LHV-26	45	0.091	0.42	52.63	0.175	0.62	0.086	0.30	52.92	0.219	1.06
LHV-27	60	0.09	0.42	53.03	0.415	1.88	0.086	0.31	53.01	0.404	2.14
LHV-28	90	0.05	0.40	53.08	0.548	2.58	0.042	0.23	53.1	0.491	2.56

Evaluation of the dissolution of copper in a molybdenite concentrate in the presence of a continuous flow of oxygen is tested in a set of experiments as tabulated in Table 6. Hydrochloric acid is added to the dissolution vessel to keep the concentration constant at the molarities given in the Table 6. After 1 hour of leaching, the copper in the final concentrate is below 0.2% only for tests run in the presence of oxygen. For these tests, the final ferric chloride concentration is similar to its initial concentration. When the initial hydrochloric acid concentration is 0.5 M in the presence of oxygen, the final ferric ion concentration is zero and the iron in the molybdenite concentrate increases to 3.05% w/w, indicating that a significant iron precipitation occurs below 0.5 M HCl. In the absence of oxygen, there is no ferric chloride regeneration and, in these tests, the copper concentration in the final concentrate is above 0.2% w/w.

TABLE 6


Effect of Oxygen and HCL Concentration on Copper Dissolution

	Disk Filter	Final
	concen-	Concen-
Initial conditions	trate	trate	Final

	Temperature	Fe(³⁺)		HCl	Cu	Fe	Cu	Fe,	Fe(³⁺)
Test	° C.	g/L	Gas	M	%	%	%	%	g/L

LHV-16	73.4	18.7	No	2	3.77	1.84	0.41	0.33	0
LHV-17	73.4	18.7	0.5 L/min 02	2	3.77	1.84	0.11	0.34	11.3
LHV-18	73.4	18.7	0.5 L/min 02	2	3.77	1.84	0.09	0.34	14.6
LHV-19	64.4	26.8	0.5 L/min 02	0.5	3.77	1.84	0.50	3.05	0
LHV-20	74.8	25.1	No	1 to 3	0.42	0.75	0.29	0.47	0
LHV-21	74.8	25.1	No	1 to 3	3.77	1.84	0.35	0.32	0

Based on the above mentioned experimental results, an embodiment of the process carried out in system 100 shown in FIG. 1 is described for 1 metric ton of a molybdenite concentrate comprising about 3.8% w/w copper, about 1.8% w/w iron, and about 50% w/w molydenum (as in Table 6). Based on Table 4, the amount Fe³⁺ needed will be about 0.1 metric ton, or 0.1 kg of Fe³⁺ per 1 kg of concentrate (from 0.08 to 0.12, Table 4). To keep the percentage of solids at 20% of the total mass, 4 metric tons of water, or about 4000 L, is needed. Based on the values of 0.65 to 3.4 kg O₂/1 kg Fe in Tables 2 and 3, the amount of oxygen needed will be 2.5 kg of O₂per 1 kg of Fe, or 250 kg of O₂per metric ton of concentrate. However, in Table 6 the oxygen flow rate is 0.5 L O₂per minute for an hour in a 0.5 L solution. For a 4000 L solution such a flow rate yields 328 Kg O₂per metric ton concentrate over a period of an hour, which would provide an excess amount of O₂to react ferrous chloride to ferric chloride. The amount of HCl required to assist in the iron oxidation is based on 0.33 kg HCl per 1 kg of Fe. With 100 kg Fe present per metric ton of concentrate, the amount of HCl needed is 33 kg per metric ton concentrate introduced into the dissolution vessel over a period of 1 hour.
In an embodiment of the process carried out in the system 200 shown in FIG. 2, the amount of O₂consumed is less at higher pressures than at atmospheric pressures. Additionally, the oxidation rate of Fe is higher at the higher pressures. Therefore, less O₂gas is needed for the process to operate. At about 7 bar gauge, the consumption of O₂is 0.22 kg per kg of Fe (Table 2). This translates to 22 kg O₂per metric ton of concentrate over a period of 1 hour.
Also common for both process 100 and process 200 is that both processes can be performed in a batch operation mode and a continuous operation mode. A batch operation mode is very much like the experiments described above, where the concentrate is leached for a set amount of time, then filtered and the filtrate recycled into a new batch of concentrate. However, in a continuous operation mode, a constant flow of gas and HCl is added to keep the concentration of ferric chloride at a level which is efficient for continuous leaching of the molybdenite concentrate. Leached molybdenite concentrate can be removed as new unleached concentrate is introduced into the dissolution vessel. Additionally, a stream of iron chloride solution from the dissolution vessel can be removed to undergo dedecopperization and the optional iron removal before being recycled back into the dissolution vessel. This way, leaching can continue uninterrupted for an extended period of time.
Preferred processes and apparatus for practicing the present invention have been described. It will be understood and readily apparent to the skilled artisan that many changes and modifications may be made to the above-described embodiments without departing from the spirit and the scope of the present invention. The foregoing is illustrative only and that other embodiments of the integrated processes and apparatus may be employed without departing from the true scope of the invention defined in the following claims.

Claims

1. A method for removing copper sulfides from a molybdenite concentrate, the method comprising:

a) leaching the copper sulfides from the molybdenite concentrate with a leaching solution comprising ferric chloride;

b) introducing an acid into the leaching solution; and

c) introducing O₂, O₃or a combination of both, into the leaching solution.

2. The method of claim 1, further comprising separating the molybdenite concentrate from a copper ore prior to leaching the copper sulfides from the molybdenite concentrate.

3. The method of claim 1, further comprising removing molybdenite from the leaching solution by filtration.

4. The method of claim 1, wherein the leaching process is performed in a batch operation mode.

5. The method of claim 1, wherein the leaching process is performed in a continuous operation mode.

6. The method of claim 1, wherein the acid comprises hydrochloric acid.

7. The method of claim 6, wherein the hydrochloric acid is kept at a concentration between about 1.0 M and about 4.0 M.

8. The method of claim 1, wherein the leaching solution maintains a leached copper ion concentration above about 1 g/L.

9. The method of claim 1, wherein the leaching solution is maintained at a temperature above about 90° C.

10. The method of claim 1, wherein the leaching solution is maintained in a vessel at a pressure between about 0 and about 30 bar gauge compared to atmospheric pressure during the introducing O₂, O₃or a combination of both.

11. The method of claim 1 further comprising agitating the leaching solution for between about 10 minutes to about 120 minutes during the introducing O₂, O₃or a combination of both.

12. The method of claim 1, further comprising removing copper from the leaching solution by exposure to scrap iron.

13. The method of claim 1, wherein the molybdenite concentrate comprises copper in a concentration of less than 0.2% w/w after leaching copper from the molybdenite concentrate.

14. A method for obtaining commercial grade molybdenite from a copper ore, the method comprising:

a) separating a molybdenite concentrate from the copper ore;

b) leaching copper from the molybdenite concentrate with a leaching solution comprising ferric chloride;

c) removing molybdenite from the leaching solution;

d) introducing an acid into the leaching solution; and

e) introducing O₂, O₃or a combination of both, into the leaching solution.

15. The method of claim 14, wherein the removing of molybdenite from the leaching solution comprises filtration.

16. The method of claim 14, wherein the leaching process is performed in a batch operation mode.

17. The method of claim 14, wherein the leaching process is performed in a continuous operation mode.

18. The method of claim 14, wherein the acid comprises hydrochloric acid.

19. The method of claim 18, wherein the hydrochloric acid is kept at a concentration between about 1.0 M and about 4.0 M.

20. The method of claim 14, wherein the leaching solution maintains a leached copper ion concentration above about 1 g/L.

21. The method of claim 14, wherein the leaching solution is maintained at a temperature above about 90° C.

22. The method of claim 14, wherein the leaching solution is maintained in a vessel at a pressure between about 0 and about 30 bar gauge compared to atmospheric pressure during the introducing O₂, O₃, or a combination of both.

23. The method of claim 14 further comprising agitating the leaching solution for between about 10 minutes to about 120 minutes during the introducing O₂, O₃or a combination of both.

24. The method of claim 14, further comprising removing copper from the leaching solution by exposure to scrap iron.

25. The method of claim 14, wherein the molybdenite concentrate comprises copper in a concentration of less than 0.2% w/w after leaching copper from the molybdenite concentrate.

26. A method for removing copper sulfides from a molybdenite concentrate, the method comprising:

a) pumping the molybdenite concentrate into an autoclave vessel;

b) introducing a solution comprising Fe(II) ions, Fe(III) ions, or a combination of both, into the autoclave vessel;

c) introducing an acid into the autoclave vessel;

d) introducing O₂, O₃or a combination of both, into the autoclave vessel; and

e) filtering the molybdenite from a stream exiting the autoclave vessel.

27. The method of claim 26, wherein the autoclave vessel maintains a pressure between about 0 and about 30 bar gauge compared to atmospheric pressure.

28. The method of claim 26, wherein the Fe(II) ions comprise ferrous chloride and the Fe(III) ions comprise ferric chloride.

29. The method of claim 26, wherein the acid comprises hydrochloric acid.

30. The method of claim 29, wherein the hydrochloric acid is kept at a concentration between about 0.5 M and about 4.0 M in the autoclave vessel.

31. The method of claim 26, wherein the solution maintains a leached copper ion concentration above about 1 g/L.

32. The method of claim 26, wherein the solution and molybdenite concentrate are maintained at a temperature between about 100° C. and about 120° C. in the autoclave vessel.

33. The method of claim 26, further comprising providing agitation in the autoclave vessel for between about 10 minutes and about 120 minutes.

34. The method of claim 26 further comprising:

f) chemically reducing copper ions in the solution to elemental copper by exposing the solution to iron; and

g) moving elemental copper.

35. The method of claim 26, wherein the molybdenite concentrate comprises copper in a concentration of between about 3% w/w and 5% w/w before entering the autoclave vessel, and wherein the molybdenite comprises copper in a concentration of less than 0.2% w/w after the filtering of the molybdenite.

36. A method for regenerating ferric chloride in a leaching solution, the method comprising:

a) adding a leaching solution comprising Fe(II) ions, Fe(III) ions, or a combination of both, to a mixture of mineral sulfides;

b) introducing an acid into the leaching solution; and

c) introducing O₂, O₃or a combination of both, into the leaching solution; whereby ferric chloride is regenerated.

37. The method of claim 36, wherein the leaching process is performed in a batch operation mode.

38. The method of claim 37, wherein the leaching process is performed in a continuous operation mode.