AU2014225297B2

AU2014225297B2 - Metal extraction from fines

Info

Publication number: AU2014225297B2
Application number: AU2014225297A
Authority: AU
Inventors: Pavel Nicolaevich SPIRIDONOV
Original assignee: INNOVECO AUSTRALIA Pty Ltd
Current assignee: INNOVECO AUSTRALIA Pty Ltd
Priority date: 2013-03-08
Filing date: 2014-03-07
Publication date: 2018-07-12
Anticipated expiration: 2034-03-07
Also published as: WO2014134681A1; CL2015002463A1; AU2014225297A1; ZA201506636B

Abstract

A method of recovering copper and other metals from fine and low grade fractions of oxide or oxide/sulphide mixes utilises fine fractions to prepare a pulp, adjusting the pH of the pulp, leaching the metal ions from the pulp, absorbing the metals on ion exchange resin and desorbing the loaded resin to recover the metal. Milling and screening the mixes obtains particles up to 0.5mm, usually 0.3mm and preferably 0.1mm.

Description

The present invention relates to a hydrometallurgical method of extraction of copper and/or other metals from fine and/or low grade fractions of oxide or mixed oxide/sulphide ores by ion exchange technology used for sorbing copper and other metals.

[0003] In particular this method is intended to improve the efficiency of the existing leaching technologies such as heap leaching and its combination with solvent extraction. Also it can be \

used to process mining cut-off and tailings, metallurgical slag and other mineral wastes.

BACKGROUND [0004] Nowadays about 20% of total primary copper is produced by hydrometallurgical extraction methods. In absolute numbers it accounts for about 2.5 million tonnes of metallic copper per year. The essential hydrometallurgical technique used currently in mining is heap leaching.

[0005] It is known that copper extraction rate and the process efficiency improve with decreasing ore particle size. However, crushing below 10mm does not further improve copper extraction while crushing below 5mm, on the contrary, decreases heap permeability and therefore reduces the process efficiency.

[0006] The major reason for the poor permeability caused by fine particles is that due to their migration they clog the natural flow channels and form impermeable layers (lenses) within the heap that limit the percolation of leaching solution. Consequently, this leads to poor solution

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PCT/AU2014/000219 distribution and low metal recovery.

DESCRIPTION OF THE PRIOR ART [0007] Nowadays, agglomeration of the crushed minerals is used as a pre-treatment option for heap leaching operations. It is required for ores that contain excessive amount of clay or an excessive quantity of fines generated during mining and crushing. A mineral feed crushed to a nominal size of 19mm or finer is agglomerated with strong sulphuric acid and placed on the leach heap.

[0008] Although agglomeration is considered as possible insurance for good recovery in heap leach technology, an improper agglomeration could be one of the major causes for reduced recovery and higher costs associated with heap leach operations.

[0009] Another way to minimise the impact of the fines on the heap leaching efficiency is to separate them from the coarse fraction and treat them in separate circuit. Authors of AU 200071731 propose to leach the fines that have a P-80 particle size of 0.21mm with a ferric sulphate lixiviant at a ferric sulphate concentration of between 5 and 30g/l, at atmospheric pressure and a temperature of between about 50 and 80°C.

[0010] The fines fraction leaching process may be carried out in relatively small multi-stage cocurrent agitated leach tank reactors. Depending on mineralogy, recovery desired, particle size, leach temperature, ion concentration and net acid generation or consumption, it may take from about 8 to 16 hours to complete the leaching operation. The leach tank reactors are preferably insulated and heat exchangers may be utilized to facilitate heat balances in the leach circuit. Following liquid/solid separation of the leached tailings, the solvent extraction method is used to recover copper. The copper recovery rate of 85 to 95% is typically obtained, depending on mineralogy. A high PLS grade of between 5 and 15g/l is normally obtained.

[0011] The major drawback of this proposal is that it cannot be used for sorption of metals from pulps and requires the liquid/solid separation, which in turn leads to loss of some copper with solids, and therefore, results in a relatively low recovery rate.

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PCT/AU2014/000219 [0012] The patent WO 2007/088010 relates to the use of monodisperse, macroporous chelating resins in metal winning in hydrometallurgical processes, in particular in resin-in-pulp processes. According to the patent, the monodisperse, macroporous chelating resins have a mean bead diameter in the range 0.35-1.5mm, preferably 0.45-1.2mm and particularly preferably 0.551.0mm. The application also provides a process for preparing monodisperse, macroporous chelating resins having weak base groups, in particular picolinamino groups. The monodisperse chelating resins are able to remove, for example, nickel and cobalt ions in relatively large amounts from a leach suspension as heterodisperse chelating resins.

[0013] According to this invention the process for recovering metals from their ores by the resin-in-pulp principle proposes the following stages: a) roasting and milling a metal-containing ore down to the particles size of less than 0.5mm; b) leaching the ore with acids, preferably sulphuric acid, hydrochloric acid, nitric acid or mixtures thereof, to leach out the metals to be recovered; c) adjusting the pH of the suspension towards neutrality by means of a neutralizing agent; d) introduction of the chelating exchanger into the suspension; e), filtration of the metalladen chelating resin from the suspension by means of a screen after a certain contact time; f) elution of the metal from the chelating exchanger with mineral acids such as sulphuric acid or hydrochloric acid or with complexing solutions such as ammoniacal solutions.

[0014] The patentee teaches that the size of the resin beads (0.35-1.5mm) and the size of the ore particles (less than 0.5mm) may overlap. However, it makes it difficult or sometime impossible to separate the resin from the ore fines. The other drawback of this invention is the use of the monodisperse resins, which are characterised by a narrow particle size distribution (PSD), with a uniformity coefficient less than or equal to 1.2. The pulps are abrasive media that lead to attrition of resin beads and reduction of their diameter. When the beads become smaller than the size of the resin/pulp separation screen they are lost. Due to this fact a certain amount of fresh resin needs to be added to the system to compensate the loss of resin. In case of monodisperse beads the loss could happen suddenly, which under production conditions could cause problems and financial loss.

[0015] The patents WO 01/29276 and WO 2007/087698 relate to the direct selective recovery of nickel and cobalt from laterite ore or oxide ore by ion exchange methods, including resin in

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PCT/AU2014/000219 pulp technique. The process described in WO 01/29276 consists of pressure or atmospheric leaching conducted at higher temperature, neutralisation or removal of iron, aluminium and copper; mixing of an ion exchange resin with the slurry to load the first metal onto the resin, separation of the metal loaded ion exchange resin from the slurry and elution of metals from the metal loaded ion exchange resin with a mineral acid.

[0016] Being in general similar to the above patent, the patent WO 2007/087698 proposes that the neutralization step of the solution can be eliminated, preventing the losses of nickel in the precipitate through the co-precipitation. Instead the impurities (iron, copper and aluminium) are removed in an additional ion exchange circuit, while the nickel and cobalt remain in the solution.

[0017] Both patents WO 01/29276 and WO 2007/087698 just state a general possibility to use ion exchange resins for the nickel and cobalt recovery from pulps, but they do not provide sufficient details on how the processes are conducted.

SUMMARY OF INVENTION [0018] The method aspect of this invention provides a process for recovery of copper and/or other metals from fine and/or low grade fractions of oxide or mixed oxide/sulphide mineral materials, which comprises:

[0019] a. separation of the fine fraction of the minerals from coarse fraction;

[0020] b. preparation of pulp which includes mixing the fine fraction with water at certain proportion and pH level adjustment;

[0021] c. simultaneous leaching of copper or/and other metals from solids and sorption of dissolved metals from the pulp using selective ion exchange resins;

[0022] d. separation of the copper loaded resin from the barren pulp;

[0023] e. desorption of the loaded resin (elution process) and production of high purity high

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[0024] f. returning the desorbed resin to the sorption cycle.

[0025] The fine fraction may be separated from the coarse fraction and processed according to this invention. The coarse fraction may be processed with a traditional technology such as heap leaching.

[0026] The fine fraction may be separated from the coarse fraction by a screen with a mesh size of up to 0.5mm, usually 0.3mm and preferably 0.1mm by milling. While the process may recover copper, a second metal present in the ore may be recovered in a separate extraction circuit which precedes or follows the copper extraction stage.

[0027] Pulp may be prepared by mixing the fine fraction or a mixture of the fine fraction with the milled coarse fraction in one or more mixing tanks.

[0028] The pH of the pulp in the mixing tank may be maintained between pH 1.1-3.8 by the addition of a mineral acid or acids and if necessary an alkaline substance or base. The pH range is optimised using lab tests and mini pilot plant trials.

[0029] The mixing of solids and liquids conducted in pulp preparation tanks and sorption tanks is assisted by agitation with pressurised air streams or mechanical mixers, especially in pulp preparation tanks.

[0030] The metal-saturated ion exchange resin is separated from pulp and desorbed in a separate circuit.

[0031] The ion exchange resin may be selected from the group consisting of strong cation, weak cation or chelating resins in the hydrogen or sodium forms, or strong anion resins in chloride or hydroxide form, or weak anion resins in free base form. Preferably, the ion exchange resin has a polydisperse particle size distribution and the particle size of the ion exchange resin is greater than about 0.6mm, preferably greater than about 0.8mm.

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Advantageous Effects of Invention [0032] 1. The proposed process is able to recover copper and other metals from fines, clays and other minerals that are difficult or impossible to process with conventional methods.

[0033] 2. The proposed process is able to recover copper and other metals in a cost effective way from tailings, slag, cut-offs, low grade ores and other minerals that are currently considered as wastes in mining industry or processing of which is uneconomical.

[0034] 3. Several processes - leaching, extraction, separation and concentration of copper can be conducted in one ion exchange (IX) plant. It means the IX plant is able to replace the existing technologies such as a combination of both heap leaching and cementation or a combination of heap leaching and solvent extraction plants or a combination of atmospheric leaching and counter current decantation (CCD) or similar technologies.

[0035] 4. The proposed process enables the reduction of capital costs due to the minimisation of equipment size.

[0036] 5. The proposed process enables the reduction of operating costs due to the reduction of process time, better space utilisation, lower consumption of energy, chemicals and water.

[0037] 6. The proposed process is characterised by high copper extraction rate of up to about 99% from solids and high concentration, high purity product solution containing about 40 g/L of copper or more.

[0038] 7. The environmental advantages of using the proposed process include the absence of flammable or toxic chemicals, lower consumption of energy, chemicals and water, deeper extraction rate of valuable metals which reduces the possibility of their leaching into the environment.

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BRIEF DESCRIPTION OF DRAWINGS [0039] One example of the invention is now described'with reference to the accompanying drawings, in which:

[0040] Figure 1 is a process flow sheet for one example of the process showing the separation and processing of the fine fraction according to the present invention.

[0041] Figure 2 is a process flow sheet for another example of the process showing the separation of the fine fraction, milling of the coarse fraction and their joint processing according to the present invention.

[0042] Figure 3 is a process flow sheet for another example of the process showing the recovery of a second metal prior to the copper recovery according to the present invention.

[0043] Figure 4 is a process flow sheet for another example of the process showing the recovery of a third metal following the copper recovery according to the present invention.

, [0044] Figure 5 is a process flow sheet for another example of the process of the present invention showing more technological details.

[0045] The following examples illustrate, but do not limit, the present invention.

Example 1 [0046] This example illustrates the copper recovery from fines of a typical copper ore separated from coarse fraction according to the process of the present invention.

[0047] In this test a sample of the copper ore containing high amount of fine fraction was used. A part of the sample was sieved on a 425 micron screen to separate coarse and fine fractions.

The separated fractions were weighed and their percentage in the total weight was calculated.

The results are provided in Table 1. It was found that approximately a half of the provided

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PCT/AU2014/000219 sample is a fine fraction which passes through the 425 micron screen.

TABLE 1

Weight and Percentage of the Fine Fraction

	Total weight of the sample, g	Weight of fine fraction (<425pm), g	Percentage of the fine fraction, %
Fine fraction	870.4	425.4	48.87

[0048] 300g of each fraction were used in the tests. The fine fraction was used as it was sieved; the coarse fraction was milled down to 425 microns. The samples were mixed with water to form 35-40% pulps. Small portions of 30% sulphuric acid solution were added to the pulp during one hour to set pH at the level of 1.8. 30ml of commercial chelating resin (iminodiacetic type) was added to each pulp. The mixture was gently stirred with overhead mixer at ambient temperature. After three hours the metal saturated resin was separated from the pulp on the 425 micron sieve and a new portion of fresh resin was added. The sorption process with the resin was repeated three times for fine fraction and five times for coarse fraction.

[0049] The samples of the pulps were taken before and after the sorption tests and analysed for copper content in both solid and liquid phases. The results of the analysis are shown in Tables 2 and 3. Based on these results the copper extraction rate was calculated. It varies between 94.5% for the fine fraction and 98.7% for the coarse fraction. The copper extraction from liquids (Table 3) is very close to 100%, particularly for the coarse fraction.

TABLE 2

The results of copper extraction from solids with ion exchange

Initial copper content in solids (before the test), ppm

Residual copper content in solids (after the test), ppm

Copper extraction rate, %

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Fine fraction sieved below 425 microns	2200	120	94.5
Coarse fraction milled down to 425 microns	11000	190	98.3

TABLE 3

The results of copper extraction from liquid phase of pulps with ion exchange

	Initial copper content in liquid (before the test), ppm	Residual copper content in liquid (after the test), ppm	Copper extraction rate,%
Fine fraction sieved below 425 microns	1500	19	98.7
Coarse fraction milled down to 425 microns	3900	7.9	99.8

[0050] The samples of saturated resin were collected and analysed for the content of copper and other metals (impurities). Then they were desorbed with 10% solution of sulphuric acid in a burette to generate the product solution, which was also analysed for copper and other metals.

The results obtained for the resin saturated during the sorption from milled coarse fraction pulp are given in Table 4. The total content of impurities in the product solution does not exceed lg^L.

TABLE 4

Content of copper and other metals on saturated resin and in product (desorption) solution

Metals	Loading on resin, mg/L	Desorption solution, mg/L
Calcium	600	390
Magnesium	250	20
Aluminium	95	60

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Copper	31500	45000
Iron	120	300
Total Impurities	1065	770

Example 2 [0051] This example illustrates the copper recovery from a typical copper ore according to the process of the present invention.

[0052] Three batches of the mineral samples were used in this test. This test was designed to simulate the continuous counter current process of sorption leaching with ion exchange resin from pulp. Prior to the test they were milled down to particle size of below 90 microns. The weight of the samples varied from 4.6 to 6.2kg. They were mixed with a measured amount of water to prepare 30-40% pulps. Small portions of 30% solution of sulphuric acid were added to the pulp to achieve the pH level in the range of 1.7-2.3. The process of mixing and stirring was conducted in a 20L vessel with an overhead mixer during 1-2 hours at ambient temperature.

[0053] The process of sorption was carried out in a series of 5 mini reactors. The pulp was pumped into the reactor 1 and then flowed through all the reactors and collected from reactor 5. The fresh ion exchange resin was supplied into the reactor 5 and then transferred from one reactor into another. The loaded (saturated) resin was taken from the reactor 1 and placed in a column for desorption.

[0054] The samples of the pulps were taken before and after the sorption tests and analysed for copper content in both solid and liquid phases. The results of the analysis are shown in Tables 5 and 6. Based on these results the copper extraction rate was calculated. For the blank test (Table 5, Sample 3 blank) - leaching under the same conditions but without sorption with ion exchange resin - the copper extraction rate was 81.4%. Much higher results were achieved using the proposed method of copper recovery with ion exchange sorption from pulp - the copper extraction rate varied between 95.8% and 96.8% for different samples. The copper extraction from liquids (Table 6) varied from 97.7% to 98.8%.

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TABLE 5

Copper content in solids and extraction rate

	Cu content in solids - original and tailings, mg/kg	Extraction rate, %
Original	1	2	3	4	1	2	3	4	Average
1	7100	270	320	-	- ·	96.2	95.5	-	-	95.8
2	8800	340	350	310	280	96.1	96	96.5	96.8	96.4
3	7700	250	250	250	250	96.8	96.8	96.8	96.8	96.8
3 blank	7700	1300	1400	1600	-	83.1	81.8	79.2	-	81.4

TABLE 6

Copper extraction rate from liquid

Sample	Initial Cu content, mg/L	Residual Cu content, mg/L	Extraction rate, %
1	2600	31	98.8
2	2000	46	97.7

[0055] The samples of saturated resin were collected and analysed for the content of copper and other metals (impurities). The results are given in Table 7.

TABLE 7 . , '

Content of copper and other metals on saturated resin, mg/L

Metals	Sample of resin
1	2
Aluminium	100	180
Copper	43300	40000
Magnesium	320	200

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Iron	10	10
% impurities	0.98	0.95

[0056] The resin saturated with copper during the sorption from pulp process was placed in a column and desorbed with a dilute solution of sulphuric acid. This fyas done to simulate the continuous counter current process of desorption. A portion of saturated resin (200-400 ml) was placed on top of the column and approximately the same amount of the desorbed (regenerated) resin was taken from the bottom of the column. The period of loading/unloading resin varied from 30 min to 1 hour. The acidic solution was fed from the bottom of the column, and the product solution was collected from the top of the column.

[0057] The samples of the product solution from the desorption colpmn were taken and analysed. Table 8 represents the composition of two samples of the product solution. The copper concentration achieved in the column desorption varies from 36g/L to 41g/L of Cu. This concentration is sufficient for the copper cathode production by electro winning. The total content of impurities does not exceed lg/L.

Resorption column (mg/1)

The composition of the product (desorption) solution produced in a

TABLE 8

	Samples of product solution
1	i ²
Aluminium	380	390
Copper	36000	41000
Manganese	220	230
Iron	5.1	4.7
Calcium	310	310
Magnesium	³⁸	36
Total impurities, g/L	0.95	0.97

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Example 3 [0058] This example illustrates the copper recovery from old flotation tailings according to the process of the present invention.

[0059] Twelve samples of old floatation tailings were obtained from several spots of the tailing dam and homogenised. 99% of the particles go through a 106 micron screen, therefore separation of the coarse fraction and its milling are not required. Two types of tests were conducted with the homogenised compound - acidic leaching test in a mixing reactor and resin in pulp (RIP) test that combined leaching and sorption tests.

[0060] The leaching test was conducted in the following maimer. 400g of the mixed sample was used to prepare a 30% pulp by mixing it with water with the overhead mixer. A 30% solution of sulphuric acid was added to the pulp over time to maintain a pH level in the range of 1.6-1.9. The leaching process was run for eight hours. After the leaching completion the samples of the pulp were taken and analysed for total and soluble copper in the solids and total copper in liquid phase.

[0061] Table 9 demonstrates the copper content in the solids of the pulp samples before and after the leaching process. The content of total and acid soluble copper in the original tailing was 0.62% and 0.542% respectively. The residual total and soluble copper in solids of the pulp samples after 8-hour leaching was 0.154% and 0.077% respectively.

[0062] The second test was conducted to estimate how much copper can be recovered from the pulp with the ion exchange technology. 400g of the mixed tailings were used to prepare a 30% pulp. During a constant stirring with an overhead mixer, a 30% solution of sulphuric acid was added to the pulp to maintain the pH level in the range of 1.6-1.9. When the pH level was stabilised, a certain ambunt of ion exchange resin was added to the pulp. When the resin was saturated with copper, it was removed from the pulp and a new portion of resin was added. The test was repeated three times. After the sorption (resin in pulp - RIP) test a sample of the pulp was analysed for total and acid soluble copper in solids and copper in liquid phase. The results are given in Table 9. The content of total and acid soluble copper in the original tailing was

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0.635% and 0.568% respectively. The residual total and soluble copper in solids of the pulp samples was 0.099% and 0.039% respectively.

TABLE 9

Copper content in the original tailing and in solids after leaching and RIP test

	Leaching test (total time 8 hrs)	RIP test (total time 4 hrs)
Cu total	Cu soluble	Cu total	Cu soluble
Initial copper content, %	0.62	0.542	0.635	0.568
Residual copper content, %	0.154	0.077	0.099	0.039

[0063] Based on these data, the copper extraction rate was calculated (see Table 10). It can be seen that the extraction of total copper during the leaching and sorption tests was 75.2% and 84.4% respectively. The extraction of acid soluble copper during the leaching and sorption tests was 85.8% and 93.1% respectively. These results demonstrate that the resin in pulp (RIP) process that utilises ion exchange technology enables to increase the copper extraction rate by -7-8% while reducing the process time.

TABLE 10

Copper extraction from solids after leaching and RIP tests

Sample	Copper extraction from solids, %
Total copper	Soluble copper
After leaching	75.2	85.8
After RIP	84.4	93.1

[0064] An additional benefit of the IX technology is high recovery of dissolved copper from liquids (solutions). Table 11 shows that the initial copper content in solutions before sorption was 1745mg/L (ppm). The residual copper concentration in solution after the RIP test was less than lmg/L. Therefore, 99.96% of copper in the dissolved state was recovered by the ion

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TABLE 11 • Copper extraction from liquid phase with ion exchange

Copper in solution before RIP, mg/L	Copper in solution after RIP, mg/L	Copper extraction from liquid, %
1745	0.76	99.96

/

Example 4 [0065] This example illustrates the copper recovery from old copper smelter slag according to the process of the present invention.

[0066] In this example a sample of copper smelter slag was used. A large amount of copper slag was industrially milled and floated to extract sulphide copper. The flotation tailings still contain some sulphide and oxide copper.

[0067] A representative sample of slag flotation tailings was obtained from the mining company. 99% of the particles go through a 106 micron screen. Two types of tests were conducted with the samples - acidic leaching test in a mixing reactor and resin in pump (RIP) test that combined leaching and sorption tests.

[0068] The leaching test was conducted in the following manner. 300g of the slag flotation tailings was used to prepare a 30% pulp by mixing it with water with the overhead mixer. A 30% solution of sulphuric acid was added to the pulp over time to maintain a pH level in the range of 1.7-1.9. The leaching process was run for six hours. After the leaching completion, the samples of the pulp were taken and analysed for total and soluble copper in the solids and total copper in liquid phase.

[0069] Table 12 demonstrates the copper content in the solids of the pulp samples before and

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PCT/AU2014/000219 after the leaching process. The content of total ahd acid soluble copper in the original tailing was 0.472% and 0.263% respectively. The residual total and soluble copper in solids of the pulp samples after 6-hour leaching was 0.388% and 0.125% respectively.

[0070] The resin in pulp (RIP) test was conducted in the following manner. 300g of the tailings were used to prepare a 30% pulp. During a constant stirring with an overhead mixer a 30% solution of sulphuric acid was added to the pulp to maintain the pH level in the range of 1.7-1.9. When the pH level was stabilised, a certain amount of ion exchange resin Was added to the pulp. When the resin was saturated with copper, it was removed from the pulp and a new portion of resin was added. The test was repeated three times. After the sorption (RIP) test, a sample of the pulp was analysed for total and acid soluble copper in solids and copper in liquid phase. The results are given in Table 12. The residual content of total and soluble copper in solids of the pulp samples was 0.306% and 0.098% respectively.

[0071] Based on these data, the copper extraction rate was calculated (see Table 12). It can be seen that the extraction of total copper during the leaching and sorption tests was 17.8% and 52.5% respectively. The extraction of acid soluble copper during the leaching and sorption tests was 35.2% and 62.7% respectively. These results demonstrate that the resin in pulp (RIP) process that utilises ion exchange technology enables to increase the copper extraction rate by over 10% while reducing the process time.

TABLE 12

Copper content in solids and extraction rate from slag flotation tailings

	Copper content, %	Copper extraction rate, %
Total Cu	Soluble Cu	Total Cu	Soluble Cu
Original	0.472	0.263	-	-
Leach 6 hr	0.388	0.125	17.8	52.5
After RIP	0.306	0.098	35.2	62.7

[0072] The recovery of dissolved copper from liquids (solutions) with the IX technology was

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PCT/AU2014/000219 also determined. Table 13 shows that the initial copper content in solutions before sorption was 1086mg/L (ppm). The residual copper concentration in solution after the RIP test was 9.5mg/L. Therefore, 99.1% of copper in the dissolved state was recovered by the ion exchange process.

TABLE 13

Copper extraction from liquid phase with ion exchange

Copper in solution before RIP, mg/L	Copper in solution after RIP, mg/L	Copper extraction from liquid, %
1086	9.5	99.1

[0073] It is to be understood that the word “comprising” as used throughout the specification is to be interpreted in its inclusive form, ie. use of the word “comprising” does not exclude the addition of other elements.

[0074] It is to be understood that various modifications of and/or additions to the invention can be made without departing from the basic nature of the invention. These modifications and/or additions are therefore considered to fall within the scope of the invention.

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Claims

1. A process for recovery of copper and/or other materials from fine and/or low grade fractions of oxide or mixed oxide/sulphide mineral materials, which comprises separation of fine fraction of the minerals from coarse fraction, preparation of pulp which includes mixing the fine fraction with water at certain proportion and pH level adjustment, sorption of dissolved copper and/or other metals from the pulp using selective ion exchange resins, separation of the copper loaded resin from the barren pulp, desorption of the loaded resin and production of high purity, high concentration copper solution, and returning the desorbed resin to the sorption cycle.

2. A process as claimed in Claim 1, wherein the coarse fraction is separated from the fine fraction and subjected to conventional extraction technique.

3. A process as claimed in Claim 1 or 2, wherein the fine fraction is separated from the coarse fraction at the screen with an aperture below 0.5mm, preferably below 0.3mm and most preferably below 0.1 mm.

4. A process as claimed in any one of Claims 1-3, wherein the fine fraction is separated from the coarse fraction and the coarse fraction is milled down to the size of about 0.1-0.5mm and then processed together with fine fraction.

5. A process as claimed in Claim 1, wherein the starting material is ready to use fine material such as tailings which in the absence of a coarse fraction does not require separation from the coarse fraction.

6. A process as claimed in any one of Claims 1 -5, wherein only copper is recovered.

7. A process as claimed in any one of Claims 1-5, wherein in addition to copper a second metal is recovered in a separate extraction circuit that precedes the copper extraction circuit.

8 . A process as claimed in any one of Claims 1-5 and 7, wherein in addition to copper a third metal is recovered in a separate extraction circuit that follows the copper extraction circuit.

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9. A process as claimed in any one of Claims 1-8, wherein the pulp is prepared by mixing the fine fraction or a mixture of fine fraction with the milled coarse fraction in multiple mixing tanks.

10. A process as claimed in Claim 9, wherein the pH level of the pulp in the mixing tank(s) is maintained in range between about 1.1 and about 3.8 by adding either a mineral acid such as sulphuric acid, hydrochloric acid, nitric acid, or their mixtures, or a solution or pulp of hydroxide compound such as sodium hydroxide, calcium oxide, calcium hydroxide, calcium carbonate, lime, lime stone or their combinations.

11. A process as claimed in any one of Claims 8-10, wherein copper or the second metal or the third metal is sorbed in a number of sorption vessels, the number N of which is determined during the engineering design stage based on the data from the laboratory tests and mini pilot plant trials.

12. A process as claimed in Claim 10 or 11, wherein each pH level sorption vessel is maintained in the range between about 1.8 and about 7.0 by adding either a mineral acid such as sulphuric acid, hydrochloric acid, nitric acid, or their mixtures, or a solution or pulp of hydroxide compound such as sodium hydroxide, calcium oxide, calcium hydroxide, calcium carbonate, lime, lime stone or their combinations.

13. A process as claimed in any one of Claims 10-12, wherein the mixing and stirring in the pulp mixing tank or tanks and sorption vessels and transferring of pulp and resin between the sorption vessels is carried out by mechanical or pneumatic means.

14. A process as claimed in any one of Claims 7-13, wherein the second metal is selected from the group consisting of aluminium, iron, rare metals, rare earth metals and mixtures thereof.

15. A process as claimed in any one of Claims 8-13, wherein the third metal is selected from the group consisting of nickel, cobalt, manganese, magnesium, and mixtures thereof.

16. A process as claimed in any one of Claims 1-15, wherein the metal-saturated ion

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17. A process as claimed in any one of Claims 1-16, wherein the ion exchange resin is selected from the group consisting of strong cation, weak cation or chelating resins in the hydrogen or sodium forms, or strong anion in chloride or hydroxide form, or weak anion resins in free base form.

18. The process of Claim 17, wherein the ion exchange resin has a polydisperse particle size distribution and the particle size of the ion exchange resin is greater than about 0.6mm, preferably greater than about 0.8mm.

19. Metals when extracted by a process as claimed in any one of Claims 1-18.

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FIGURE 1

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Solids to tailings-► dam

FIGURE 2

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FIGURE 3

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FIGURE 4

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FIGURE 5