DIRECT ELECTROWINNING PROCESS WITH LEACHING SOLUTION

Abstract

The invention relates to a process that allows electrolytic copper cathodes to be produced, using the pregnant leach solution (PLS) directly in the electrowinning, avoiding the step of mineral concentration by solvent extraction. Furthermore, this process has a modular structure and the full process can be mobilised depending on the requirements of the process itself. The invention also relates to the system that operates with the previously described process.

Claims

1. A system of electrowinning without extraction by solvents based on a pregnant leach solution (PLS) with a copper content from 5 to 30 g/L, comprising: rotary pumps of PLS, intermediate lixiviation solution (ILS), water, and reagents; a PLS accumulation tank; an emergency pool, mixed solution pool, and ILS pool; a PLS heater; a heat exchanger; modules of electrolytic cells for lixiviation with direct electrowinning (EW-LED cells); piping for feeding the PLS to the EW-LED cells; and rectifying low current transformers and instruments that permit taking a census and measuring the variables of an electrowinning process.

2. The system of claim 1, wherein the PLS heater and the heat exchanger work together where the heat exchanger, using PLS/PLS plates, transfers the heat energy from the PLS and a raffinate solution leaving the EW-LED cells, towards recirculating/fresh incoming PLS, according to a PLS recirculation cycle, so as to power feedback the PLS in turn, and then thermally treat the PLS once again in the heat exchanger, using PLS/water plates operated with hot water as a heat-contributing fluid in an operation of countercurrent flows without direct contact, as in the heat energy transfer using the PLS/PLS plates, permitting maintaining a constant temperature of the PLS in a range between 20° C. and 60° C. upon entering the EW-LED cells.

3. The system of claim 1, wherein the PLS rotary pump, PLS accumulation tank, emergency pool, mixed solution pool, ILS pool and piping for feeding the PLS to the EW-LED cells are configured to operate sequentially to pumping onto dynamic pads of agglomerated material a solution of lixiviation raffinate originating from the spent solutions of the EW-LED cells and then recirculate ILS from the ILS pool, which is formed based on the ILS with a low concentration of copper, process water and also sulfuric acid from progressive enrichment of copper, via piping to the EW-LED cells as PLS that is heated and conditioned in line with the adding of additives, fresh sulfuric acid and process water, then from the EW-LED cells as a copper-poor solution towards a transfer tank, which derives the solution towards the mixed solution pool a specific number of times until the concentration of copper in the copper-poor solution from the EW-LED is lowered to a pre-established value and thereafter derive the copper-poor discharge solution to the mixed solution pool as a raffinate solution for its re-enrichment in copper through the dynamic pads and thus returns to the electrowinning process such that continuous volumes of electrolytic solution are handled.

4. The system of claim 1, wherein the instruments that permit taking a census and measuring the variables of the electrowinning process are integrated into the system, and also permit detecting and controlling continuous current, temperature, flow rates, pH and electric conductivity.

5. The system of claim 1, wherein rectifier transformers feed and control a density of load to the EW-LED cells.

6. The system of claim 1, wherein the entire system is modular and enlargeable depending on the offer of material.

7. The system of claim 1, wherein the entire system is mobile and can be installed on site in the field.

8. A method of electrowinning without extraction by solvents with pregnant leach solution PLS with a copper content from 5 to 30 g/L, comprising: a. crushing an ore until 100% of the crushed ore is below 1.27 cm in diameter; b. agglomerating the crushed ore with an acid and process water until an agglomerated material for lixiviation is obtained: c. lixiviating the agglomerated material through irrigation with a raffinate solution of a dynamic pad with the agglomerated material with a maximum concentration range of dissolved copper of 3 to 5 gr per liter, to obtain an ILS solution; d. recirculating the ILS solution passing by the dynamic pad until a PLS is formed having a concentration of copper in the range of 10 to 18 gr per liter; e. conditioning the PLS with process water and acid up to a concentration of acid of 120 gr per liter of conditioned PLS; f. thermally conditioning the PLS in a temperature range between 20° C. and 60° C., using an external heat exchanger that transfers part of the heat at the discharge of the PLS from an electrolytic cell for lixiviation with direct electrowinning (EW-LED cell) and part of the heat from another external source based on a boiler, to fresh PLS before entering the EW-LED cell; g. incorporating additives into the PLS to avoid jacketing of electrodes in the EW-LED cell before entering the EW-LED cell; h. flowing the PLS into the EW-LED cell with a volume of flow in a range of 0.1 to 10 m.sup.3 /hour, with the application of a current density controlled independently by bank and by modules, in a range of 100 to 450 A/m.sup.2, wherein a series of the EW-LED cells form a module and a series of the modules form a bank and a series of banks form a bay; i. circulating the PLS in the bay, from bank to bank through feeder tanks in series, where an outgoing solution contains a concentration of copper between 2 and 6 gr per liter, and is sent towards a transfer tank, to later be driven to the external heat exchanger to a sub-pool of PLS in recirculation; and j. removing copper cathodes obtained in the banks of EW-LED cells.

9. The method of claim 8, wherein the acid used in stage b) is sulfuric acid in a volume of flow between 0.01 and 5 ton/h.

10. The method of claim 9, wherein the volume of flow of the acid in stage b) is applied between 0.01 to 5 ton/h, preferably 2.11 ton/h, and its concentration must be within a range of between 1 to 100 kilograms per ton of ore, plus the process water.

11. The method of claim 8, wherein stage b) is executed until the agglomerated material has a humidity between 5% and 20%, and between 5 and 20 kilograms of acid per ton of agglomerated material.

12. The method of claim 8, wherein in stage c) the dynamic pad is irrigated with raffinate solution that has already been processed through the modules of EW-LED cells.

13. The method of claim 8, wherein in stage f) the optimum temperature is 45° C.

14. The method of claim 8, wherein the additives added in stage g) are fresh sulfuric acid, guar gum and cobalt sulfate.

15. The method of claim 8, wherein the outgoing solution in stage i) contains a concentration of copper of 4 gr per liter.

16. The system of claim 1, wherein: each of the modules of EW-LED cells included between 2 to 12 compact cells, where each cell is formed by a cathode and an anode; the distribution of electrical connections in the cells are connected in series between cathodes and anodes and integrated, in order to maintain an identical continuous current and with an equal current density in all the modules and inter-modules; an operative area of the cathode is in the range of 2 to 0.3 m.sup.2 ; the current density is regulated in the module in a range between 0 and 450 amperes, with an operative current density between 150 and 300 amperes per m.sup.2 of cathode; and piping also exists independent of the flow of PLS; and an independent control of the electric field.

Description

DESCRIPTION OF THE FIGURES

[0066] FIG. 1

[0067] This figure presents a simplified sketch of the traditional copper cathode production process (upper sketch) versus a simplified sketch of the process of this invention (lower sketch).

[0068] A: Crushing: at this stage, the material extracted is reduced to smaller and smaller and more compact portions.

[0069] B: Lixiviation (LX): Metallurgical technique that consists of watering the piles of mineralized material with a solution of water with sulfuric acid, dissolving the copper contained in the oxidized ores and forming a solution of copper sulfate, which is taken to the PLS (pregnant leaching solution) pools.

[0070] C: Extraction by solvent (SX): is a method of separating one or more substances from a mixture, using solvents, obtaining a solution rich in copper.

[0071] D: Electrowinning (EW): this is a process whereby the copper solution [electrolyte concentrated in copper, after the extraction by solvents (SX)] is taken to the electrowinning bay where there are a series of cells, that when a current is applied to them, the copper sulfate solution breaks down, and the copper becomes adhered to the cathodes.

[0072] E: Direct Electrowinning with Lixiviation (EW-LED) is a process whereby the LPS solution [Solution obtained after the lixiviation (LX) stage] is taken to the different modules of the modified electrowinning cells, after the incorporation of reagents such as guar gum and cobalt sulfate as additives for the cathodes, finally the copper becomes adhered to the cathodes.

[0073] FIG. 2

[0074] This figure represents a general diagram of the productive process of this invention.

[0075] The line of arrows of the upper part of the diagram shows the physical phenomena that the water suffers in the different positions of the movement of the ore: [0076] X: Impulse pumps [0077] : Flow controllers and temperature meters [0078] F1: Evaporation of the mixed pool [0079] F2: Evaporation of the ILS pool [0080] F3: Evaporation piles [0081] F4: Evaporation in EW [0082] F5: Decomposition of water by electrolysis [0083] F6: Washing water to discard [0084] F7: Production of copper cathode

[0085] The line immediately below the arrows of the upper part shows the behavior of the solid material in the different positions of the movement of the ore: [0086] G1: Agglomerated ore from the crusher-binder [0087] G2: Dynamic pile [0088] G3: Gravel to dump

[0089] In the following line of arrows, the handling of the acid is presented: [0090] H1: Sulfuric acid from trucks [0091] H2: Acid TK [0092] H3: Sulfuric acid to agglomeration.

[0093] In the following line of arrows, the handling of the process water is presented: [0094] I1: Process water from water supply [0095] I2 Service water [0096] I3: ILS pool [0097] I4: Mixed pool [0098] I5: Emergency pool

[0099] The last line presents the system's heating network: [0100] J1: Oil supply [0101] J2: Boiler [0102] J3: Water conditioning chamber [0103] J4: Heat exchangers

[0104] There are other parts associated to the adaptation and preparation of the LPS before the EW-LED: [0105] K1: Chemical product, concentrated Guar [0106] K2: Chemical product, concentrated cobalt sulfate [0107] K3: Guar TK, this is a tank where the Guar is diluted in water and is left at an optimum concentration to be applied to the PLS that is sent to the EW cells. [0108] K4: Cobalt TK, this is a tank where the cobalt sulfate is diluted in water and is left at an optimum concentration to be applied to the PLS that is sent to the EW cells. [0109] K5: TK Bank 2, this is a tank where the electrolyte in series is received when it has passed once through the first bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of the electrolyte reaches a range below 4 g/L.) [0110] K6: TK Bank 3, this is a tank where the electrolyte in series is received when it has passed once through the second bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of the electrolyte reaches a range below 4 g/L.) [0111] K7: TK Bank 4, this is a tank where the electrolyte in series is received when it has passed once through the third bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of copper in the PLS solution reaches a range below 4 g/L.) [0112] K8: TK transfer bank, this is a tank where the PLS is received, in series, that has passed once through the fourth bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of the PLS reaches a range below 4 g/L.)

[0113] The PLS used is transferred to the mixed pool. [0114] L1: EW-LED bank No 1 [0115] L2: EW-LED bank No 2 [0116] L3: EW-LED bank No 3 [0117] L4: EW-LED bank No 4

[0118] FIG. 3

[0119] This figure represents a specific sketch of the stages prior to the LED process.

[0120] FIG. 4

[0121] This figure represents a flow diagram of the lixiviation process.

[0122] In the arrows of the right part of the diagram, the physical phenomena that the water suffers in the different positions of the movement of the ore are presented: [0123] X: Impulse pumps [0124] : Flow controllers and temperature meters [0125] F1a: Evaporation of the mixed pool, raffinate section [0126] F1b: Evaporation in the mixed pool, PLS section [0127] F2: Evaporation ILS pool [0128] F3: Evaporation piles

[0129] In the upper line of arrows, the behavior of the solid material in the different positions of the ore movement is presented: [0130] G1: Agglomerated ore from the crusher-binder [0131] G2: Dynamic pile [0132] G3: Gravel to dump

[0133] In the following line of arrows, the handling of the PLS and one line is shown: [0134] M1: PLS in recirculation/raffinate from EW-LED [0135] M2: PLS/PLS recirculated to EW-LED

[0136] In the following line of arrows, the handling of the acid is shown: [0137] H1: Sulfuric acid from trucks [0138] H2: TK acid [0139] H3: Sulfuric acid to agglomeration [0140] H4: Sulfuric acid to EW-LED

[0141] In the following line of arrows, the handling of the process water is presented: [0142] I1: Process water from water supply [0143] I3: ILS pool [0144] I4: Mixed pool [0145] I5: Emergency pool [0146] I6: Process water to EW-LED [0147] I7: Emergency LX shower

[0148] Additionally, the following numbering shows: [0149] 1. Discharge of Raffinate from EW [0150] 2. Raffinate available from EW (by balance) [0151] 3. Raffinate to watering [0152] 4. ILS to pool [0153] 5. ILS in recirculation to watering [0154] 6. Process water to operations [0155] 7. Water to emergency LX service [0156] 8. Water to Mixed Pool (Raffinate section) [0157] 9. Water to ILS pool [0158] 10. Sulfuric acid to operations [0159] 11. Sulfuric acid to Agglomeration [0160] 12. Sulfuric acid to ILS pool [0161] 13. Process water to EW [0162] 14. Sulfuric acid to EW [0163] 15. Evaporation Mixed Pool (Raffinate section) [0164] 16. Evaporation ILS pool [0165] 17. Evaporation Mixed Pool (PLS section) [0166] 18. PLS to Mixed Pool (PLS section) [0167] 19. PLS available to EW (by balance) [0168] 20. Discharge of PLS to EW

[0169] FIG. 5

[0170] This figure represents a diagram of direct electrowinning process flows in series EW-LED.

[0171] In the line of arrows of the lower right hand part of the diagram, the physical phenomena that the water suffers in the different positions of the movement of the ore are presented: [0172] F5: Decomposition of water by electrolysis [0173] F6: Washing water to discard [0174] F7: Production of copper cathodes [0175] F8: Evaporation of the water by environment [0176] F9: Water for washing cathodes

[0177] In the following line of arrows, the handling of the acid is presented: [0178] X: Impulse pumps [0179] : Flow controllers and temperature meters [0180] H4: Sulfuric acid to EW-LED.

[0181] In the following line of arrows, the handling of the process water is presented: [0182] I6: Process water to EW-LED [0183] I7: LX emergency shower [0184] I8: Service water for human consumption

[0185] The final line presents the system's heating network: [0186] J1: Oil supply [0187] J2: Boiler [0188] J3: Water conditioning chamber [0189] J4: Heat exchangers

[0190] There are other parts associated to the adaptation and preparation of the LPS before the EW-LED. [0191] K1: Chemical product, concentrated Guar [0192] K2: Chemical product, concentrated cobalt sulfate [0193] K3: TK Guar, this is a tank where the Guar is diluted in water and left at an optimum concentration to be applied to the PLS that is sent to the EW cells. [0194] K4: TK Cobalt, this is a tank where the cobalt sulfate is diluted in water and is left at an optimum concentration to be applied to the PLS that is sent to the EW cells. [0195] K5: TK Bank 2, this is a tank where the PLS in series is received when it has passed once through the first bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of PLS reaches a range below 4 g/L.) [0196] K6: TK Bank 3, this is a tank where the PLS in series is received when it has passed once through the second bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of PLS reaches a range below 4 g/L.) [0197] K7: TK Bank 4, this is a tank where the PLS in series is received when it has passed once through the third bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of PLS reaches a range below 4 g/L.) [0198] K8: TK transfer bank, this is a tank where the PLS is received, in series, that has passed once through the fourth bank of the EW-LED system. (Without restricting the number of banks to be used except when the concentration of PLS reaches a range below 4 g/L.)

[0199] The PLS used is transferred to the mixed pool. [0200] L1: EW-LED No 1 bank [0201] L2: EW-LED No 2 bank [0202] L3: EW-LED No 3 bank [0203] L4: EW-LED No 4 bank

[0204] In the following line of arrows, the handling of the PLS and a line are presented: [0205] M3: PLS/PLS recirculated from LX [0206] M4: PLS recirculated/raffinate to LX

[0207] Additionally, the following numbering shows: [0208] 21: PLS to conditioning [0209] 22: Sulfuric acid to line [0210] 23: Process water to line [0211] 24: PLS to E/E heat exchanger [0212] 25: PLS to E/A heat exchanger [0213] 26: Cobalt sulfate to Cobalt sulfate TK [0214] 27: Guar to Guar TK [0215] 28: Solution of cobalt sulfate [0216] 29: Solution of guar to distribution [0217] 30: Guar solution to EW 1 Bank [0218] 31: Guar solution to EW 2 Bank [0219] 32: Guar solution to EW 3 Bank [0220] 33: Guar solution to EW 4 Bank [0221] 34: PLS to EW 1 Bank [0222] 35: PLS to EW 2 Bank [0223] 36: PLS to EW 3 Bank [0224] 37: PLS to EW 4 Bank [0225] 38: PLS to transfer TK [0226] 39: PLS in circulation to pool [0227] 40: Hot water from heater [0228] 41: Hot water to Cobalt Sulfate TK [0229] 42: Hot water to Guar TK [0230] 43: Hot water to heat exchanger [0231] 44: Hot water to cathode washing [0232] 45: Hot water in return [0233] 46: Process water to services and operation [0234] 47: Water to EW-LED emergency service [0235] 48: Water to human consumption service [0236] 49: Process water to replacement [0237] 50: Water to heater [0238] 51: Oil to heater [0239] 52: Evaporation of water in bay [0240] 53: Decomposition of water in bay [0241] 54: Cathodic copper [0242] 55: Propelling raffinate to LX

[0243] FIGS. 6A-6D

[0244] These figures present, in FIGS. 6A and 6B, the Iso-pH tests in material of Bella Andina.

[0245] FIGS. 6C and 6D present the Iso-pH tests in material from Chiapa.

[0246] FIGS. 7A-7B

[0247] In these figures, the sulfation tests on the ITE material (graph of FIG. 7B) and in the composite material (graph of FIG. 7A) are described.

[0248] FIGS. 8A-8B

[0249] In FIG. 8B, the graph presents the relationship between the percentage of copper extraction and the lixiviation ratio. It is clear that the extraction is not optimum with regard to the presentation of the PLS to the exposure to traditional EW, that achieves concentrations of approximately 48 grams per liter of copper in the electrolyte.

[0250] The graph of FIG. 8A presents the relationship between the net acid consumption and the lixiviation ratio.

[0251] FIG. 9

[0252] This figure presents a photograph of a copper cathode extracted by the direct EW technique, the object of this patent.

EXAMPLES OF APPLICATION

[0253] Experimentally, 2300 kg of crushed ore were prepared and screened until a product size below 1.27 cm or half an inch in size, with a granulometry of P80, was obtained.

[0254] This ore was stored to be used as the tests were developed, just as it happens in reality where a stock pile is stored. Once here, the material that was used for the tests was separated in sub-lots of 50 Kilograms. One of the sub-lots was used to execute a chemical characterization (500 g), iso-pH (1000 g) and metallurgical characterization.

[0255] A 5 kg sample of a sub-lot was crushed and shaken in a rotary drum so as to obtain representative 250 g samples.

[0256] The tests carried out with these samples are:

[0257] a. Chemical Analysis of Leach feed CuT (Total Cu), CuS (Soluble Cu), FeT (Total Fe), Al, Mg, Mn, Na, K, Cl and CAA (Consumption of Analytic Acid).

[0258] In order to quantify the contaminating elements in the ore sample, analysis by ICP (inductive coupling plasma) of 31 elements were carried out in the leach feed sample.

[0259] b. Preliminary Metallurgical Tests

[0260] Preliminary metallurgical tests were carried out with the object of obtaining results of metallurgical parameters of lixiviation.

[0261] c. Test of Iso-pH These tests are executed using 1 kg of ore with adequate granulometry, for a period of 48 hours and with a percentage of solids of 33%. The tests are carried out in a plastic reactor with a capacity of 10 liters, which rotates on a roller at 55 rpm, designed especially for doing this work.

[0262] The leaching solution is maintained always at a pH of 1.5. This situation is reached by the constant and permanent addition of the H.sub.2SO.sub.4 acid, which is reported as a net and gross consumption of acid.

[0263] With a frequency of 2, 4, 6, 8, 10, 24, 48 hours, samples of pulp (with a 10% humidity) are taken to carry out the follow-up of the kinetics of copper extraction and acid consumption. At the end of the lixiviation period, the pulp is filtered and washed, obtaining a rich solution and a washing solution.

[0264] The ore cake (lixiviation gravels) is dried, weighed, disintegrated and analyzed for Total Cu (CuT) and Total Fe (FeT), and finally, on the basis of weights, volumes of solutions and chemical analysis, execute a metallurgical balance.

[0265] d. Sulfation Test

[0266] The sulfation tests were used to determine the dose of acid to be employed in the column lixiviation tests. This protocol formed part of a set of four sulfation tests, that employed a rest period of 48 hours. Once the rest period was finished, the sample was washed to obtain the solution and characterization by Cu, acid free. From the results obtained in the four sulfation tests, it was determined that the dose of acid in curing, under the principle of employing the smallest amount of acid, on which the residual acid starts to remain. This dose of acid was employed to cure the ore prior to the column lixiviation stage.

[0267] e. Column Lixiviation Tests

[0268] The objective of the column lixiviation tests was to obtain PLS solutions with contents of weighted Cu between 8 to 10 g/L, because at this level, this is what is required for use in EW tests.

[0269] The metallurgical program carried out nine (9) lixiviation tests in columns of 20.32 cm (8 inches) in diameter and 3-5 meters of height. Each test used an acid rate in curing determined on the basis of the sulfation tests. The irrigation rate was of 10 L/h/m.sup.2. Moreover, the type of water constituted a variable, which is why six (6) columns employed a lixiviation solution made up of acidulated potable water and three (3) columns with acidulated seawater.

[0270] The lixiviation ratio was 2 m.sup.3/ton, or else until a concentration of 8-10 g/L of Cu was obtained.

[0271] The control of the experiments included a daily analysis for: Cu, FeT and H+, during the first five (5) days, to then continue with samplings every other day, until the end of the irrigation cycle.

[0272] The results of these experiences were as follows:

[0273] For the chemical characteristic, three ore samples originating in Peru were taken that correspond to the Bella Andina, ITE and Chiapa mines. These mineral deposits lie very close to Arica.

[0274] A composite was made with the samples from Bella Andina and Chiapa, which were characterized by means of atomic absorption techniques and ICP, whose results are shown in Tables I and II.

TABLE-US-00001 TABLE I CuT CuS FeT Al Mg Mn Na K Cl CAA Sample (%) (%) (%) (%) (%) (%) (%) (%) (%) (kg/t) Chiapa Mine 2.45 2.00 10.66 3.71 0.22 0.13 0.06 0.07 0.50 160.7 Bella Andina Mine 3.17 3.05 6.66 4.91 1.81 0.31 0.21 0.14 0.16 305.8 Composite (75; 25) 2.63 2.26 9.66 4.01 0.62 0.17 0.10 0.09 0.41 197.0

[0275] The results obtained show that both ores have a high ore content of Cu Total: 2.45 and 3.17% for Bella Andina and Chiapa, respectively, while the CuS was of 2.0 and 3.05% (Bella Andina and Chiapa), indicating a solubility ratio of 81.6% for (Chiapa Mine), while (Bella Andina) showed a solubility ratio of 96.2%. The samples showed high in FeT, 10.7 and 6.7% for Bella Andina and Chiapa, respectively. With regard to cations analyzed (Al, Mg, Na, K), these present normal levels, Al between 3.7 and 5%, Mg between 0.22 and 1.81%. Then, in concentrations below 1%, we have Na, K, Cl; these latter elements are indicative of the absence of soluble salts in both ores.

TABLE-US-00002 TABLE II Element Unit Bella Andina Chiapa ITE K (%) 0.02 0.05 0.11 Na (%) 0.03 0.10 0.09 Ti (%) 0.19 0.12 0.03 S (%) 0.34 2.53 0.05 Mg (%) 1.22 0.14 0.34 Al (%) 3.03 1.95 0.79 >15.0 Fe (%) 5.22 8.82 Ca (%) 10.86 12.68 0.20 >1000 >100 Cu (ppm) >10000 0 00 Zn (ppm) >10000 250 88 Mn (ppm) 4076 1416 218 As (ppm) 2899 49 10 P (ppm) 939 3713 385 Ni (ppm) 289 35 43 Pb (ppm) 260 126 45 Tl (ppm) 218 64 <10 Sr (ppm) 140 157 15 V (ppm) 133 25 84 Cd (ppm) 114 2 3 Ba (ppm) 66 <10 19 Zr (ppm) 50 13 8 Cr (ppm) 44 57 60 Co (ppm) 38 21 84 La (ppm) 38 <10 <10 Sb (ppm) 28 <5 <5 Li (ppm) 26 16 4 Y (ppm) 23 18 3 Ga (ppm) 18 16 22 Ta (ppm) 15 9 24 Ag (ppm) 11 19 <1 Sc (ppm) 10 4 <1 Bi (ppm) 9 <5 10 Mo (ppm) 9 8 20 Be (ppm) 3 <1 <1 Th (ppm) <5 <5 <5 Nb (ppm) <10 <10 <10 Se (ppm) <10 <10 <10 Sn (ppm) <10 <10 <10 Te (ppm) <10 11 20 U (ppm) <10 <10 <10 W (ppm) >10 <10 <10 Hg (ppm) <1 <1 <1

[0276] This table presents the results of the ICP

[0277] With regard to the elements analyzed by ICP, we draw attention to the Calcium content in both samples, indicative of a high consumption of acid.

[0278] One of the relevant aspects in the chemical characterization is the determination of the contents of soluble salts in the samples and their consumption of acid, for which ore lixiviation was carried out with a granulometry of 10#, with one (1) liter of hot water (boiling), which was shaken during one hour.

[0279] The solution obtained was analyzed for Na, Cl and K, whose results are shown in Table III below.

TABLE-US-00003 TABLE III Solution Analysis Dissolved Salts Na Cl— K Na Cl— K Washing solution (g/L) (g/L) (g/L) (g) (g) (g) Bella Andina 0.118 0.61 0.013 0.12 0.61 0.01 Chiapa 0.174 0.40 0.035 0.17 0.40 0.04

[0280] This table presents the dissolution of salts in hot water. It is also observed that the concentration of dissolved salts is low in the three cases; the possibility of significant concentrations of hydrochloric acid in situ does not exist.

[0281] The next step in the tests were the metallurgical tests in which Iso-pH, sulfation and column lixiviation tests were carried out.

[0282] The first test was carried out for the samples from Bella Andina (1940 g) and Chiapa (1946 g), a constant pH of 1.5.

[0283] The results of this test are seen in FIGS. 6A-6D and in the following Table IV.

TABLE-US-00004 TABLE IV RESULTS LEACH Acid FEED consumption CuT CuS Cu Extraction (%) (kg/t) Sample (%) (%) Analysis Calc. Earth Gross Net Bella 3.17 3.05 93.97 93.97 93.97 185.19 185.19 Andina Mine Chiapa 2.45 2.00 79.56 77.68 77.14  52.85  22.78 ITE 3.45 3.40 85.20 89.39 89.89  67.89  22.69

[0284] This table presents a summary of the results of the Iso-pH tests. It can also be seen that the samples present high extraction levels of Cu, of 94% for Bella Andina and 89% for Chiapa. On its part, the consumption of acid presented the highest levels for Bella Andina, 185 kg/t, while for the ITE and Chiapa samples, the net consumption was 23 kg/t.

[0285] On the other hand, the sulfation test was carried out based on the results obtained in the acid consumption in the iso-pH tests. The sulfation tests were executed with concentrations of 35, 45, 55 and 65 kg/t, whose results are reported in Table IV. FIGS. 7A-7B.

[0286] Based on the results presented in FIGS. 7A-7B, the acidity levels in the agglomerate were obtained from the sulfation tests.

[0287] As the next step, the column lixiviation tests were executed, which consisted in obtaining a PLS solution, with Cu contents between 8 to 10 g/L. The results of these tests are clearly evident in Table V and FIGS. 8A-8B.

[0288] According to the results obtained, a low Copper extraction was observed, whose maximum value was 58% (C-1 and C-2) for the 3-meter columns (bed height) and 45% for the 5-meter columns (C-3 and C-4), executed with a mineral composite (25% Bella Andina and 75% Chiapa). The columns with ITE mineral (C-10 and C-11) presented 47% of Cu extraction. These tests only reached a Lixiviation Ratio of 2 m.sup.3/t. All these results were inferior to the solubility ratio, that in all the cases was higher than 90%, which is why certain adjustments needed to be made in the experimental operation.

[0289] The low extraction level observed in all the experiences was the result of not operating under optimized conditions for a lixiviation by percolation based on: [0290] Granulometry: was employed au naturel and was obtained at the Industrial Plant in Arica. The material was segregated, with the component of fine material very low. [0291] The end condition of the experience was the concentration of Cu in PLS and not the level of maximum extraction (which was not sought by the subsequent EW process).

[0292] Based on the above, an additional column was loaded to produce the volume and concentration of Cu required to be able to carry out the EW test.

[0293] In the first days, the PLS solution presented concentrations high is Fe (30 to 40 g/L) which are not adequate for the EW process. This is why the PLS was recirculated in the same column, thereby reducing the iron intake as the consumption of acid.

[0294] With regard to the concentration of Cu, the columns made with the composite (Bella Andina and Chiapa) presented a concentration of 8 to 9 g/L, weighted, while the tests executed with the ITE mineral, the concentration was of 22 g/L. With mixtures of both solutions, we managed to obtain the PLS solution for the EW tests.

[0295] The consumption of net acid was 33 kg/ton for columns C-1 and C-2 (3m), while for C-3 and C-4 (5m) the consumption was 24 kg/ton, executed with the composite (Bella Andina and Chiapa). The C-10 and C-11 tests, executed with ITE mineral, presented low levels of net acid consumption of 16 kg/ton.

[0296] The tests made with seawater were started at different times than the tests with potable water, the columns C-5, C-6 and C-7, with composite sample (5m), only reaching the lixiviation ratio of 2 m.sup.3/ton. These presented low extraction levels of Cu (30%), the concentration of Cu (weighted) was of 8 g/L. On its part, the net acid consumption was of 17 kg/ton.

[0297] As the objective of the tests was to obtain PLS, the solutions were stored based on the concentration of Cu; therefore, they were placed in separate drums by concentrations between 10 and 20 g/L, 5 and 10 g/L and less than 5 g/L.

[0298] As can be seen, the results obtained in the nine experiences are the consequence of a process to obtain PLS solution of conditions required to be used in a direct EW process. Therefore, the metallurgical results obtained are not the consequence of an optimized process of lixiviation by percolation.

TABLE-US-00005 TABLE V Metallurgical Parameters Unit C-1 C-2 C-3 C-4 C-5 C-6 C-7 C-10 C-11 ID Sample Composite Composite Composite Composite Composite Composite Composite ITE ITE LEACH FEED Total Copper % 2.42 2.42 2.42 2.42 2.42 2.42 2.42 3.45 3.45 grade Soluble % 2.21 2.21 2.21 2.21 2.21 2.21 2.21 3.4 3.4 Copper grade Solubility % 90.95 90.95 90.95 90.95 90.95 90.95 90.95 98.64 98.64 ratio Granulometry inch P80 ½″ P80 ½″ P80 ½″ P80 ½″ P80 ½″ P80 ½″ P80 ½″ P80 ½″ P80 ½″ OPERATING CONDITION Height m 3 3 5 5 5 5 5 5 5 Diameter inch 8 8 8 8 8 8 8 8 8 Acid in Kg/t 25 25 25 25 20 20 20 30 30 curing Watering — Potable Potable Potable Potable Sea Sea Sea Potable Potable solution Acidity in g/L 15/20 15/20 15/20 15/20 15/20 15/20 15/20 15/20 15/20 watering solution Watering rate L/h/ 10 10 10 10 10 10 10 10 10 m.sup.2 Lixiviation days 44 44 44 44 37 37 37 39 39 time Lixiviation M.sup.3/t 2.31 2.22 2.27 2.31 1.91 2.01 1.95 2.0 1.97 ratio GRAVEL Total Copper % 1.11 1.11 1.24 1.34 1.64 1-58 1.63 1.37 1.44 Grade Weight loss % 4.42 4.52 2.46 3.18 2.69 2.41 2.36 4.08 2.7 RESULTS Metallurgical % 95 95 103 100 102 105 102 104 102 Accounting Copper in g/L 8.4 8.7 11.8 11.6 8.5 8.1 8.4 22.7 23.1 rich solution Copper % 61.9 61.6 47.5 46.4 29.9 30.5 29.7 48.4 46.8 extraction Gross acid Kg/t 56.3 55.0 40.7 40.5 27.9 28.3 29.6 44.8 44.5 Consumption Net acid Kg/t 33.2 32.0 23.0 23.2 17.6 18.1 17.7 16.5 16.5 consumption

[0299] On the other hand, electrolyte solutions were prepared to be exposed to the EW process. Electrolyte was prepared for 2 experiences (P1 and P2), of which the first was of 400 L for calibration and subsequently one of 800 L for the EW test itself, to which acid was added in the levels required (150 g/L). According to the latter, the solution was characterized via EAA for Cu and FeT, C1 for volumetric analysis and multi-element analysis via ICP.

[0300] Table VI shows the chemical characterization of both electrolytes at the entrance of the EW process, while Table VII shows an analysis of the impurities of the P2 solution.

TABLE-US-00006 TABLE VI Test N° Cu.sup.++ (g/L) H.sup.+ (g/L) FeT (g/L) Cl (g/L) P1 12.27 151.0 0.71 0.87 P2 14.78 149.5 0.57 0.69

TABLE-US-00007 TABLE VII Al Mn Ca Mg Na K SO.sub.4 Sample (g/L) (g/L) (g/L) (g/L) (g/L) (mg/L) (g/L) P2 0.62 0.17 0.42 0.61 0.42 47.64 142.53

[0301] On its part, Table VIII shows the multi-element composition of the PLS solutions (using potable water and seawater as a medium) before adding acid to transform it into a solution for electrowinning.

TABLE-US-00008 TABLE VIII Element Unit PLS Seawater PLS Potable Water Cu (mg/l) >500 >500 S (mg/l) >500 >500 Zn (mg/l) >500 >500 Na (mg/l) >2500 377 Mg (mg/l) 1942 661 Fe (mg/l) 1765 582 Al (mg/l) 879 562 Ca (mg/l) 468 456 K (mg/l) 308 31 Mn (mg/l) 283 207 As (mg/l) 162 8.15 P (mg/l) 149 16.6 Ni (mg/l) 37.9 22.9 Pb (mg/l) 23.4 18.8 Cd (mg/l) 16.1 8.52 Co (mg/l) 11.2 7.33 Sr (mg/l) 6.32 1.81 U (mg/l) 3.50 3.36 V (mg/l) 3.28 1.51 Cr (mg/l) 1.44 1.19 La (mg/l) 1.34 0.90 Li (mg/l) 1.23 0.72 Y (mg/l) 1.03 0.81 Tl (mg/l) 0.60 0.51 Bi (mg/l) 0.58 <0.25 Ta (mg/l) 0.51 <0.50 Ag (mg/l) 0.45 <0.05 Hg (mg/l) 0.32 0.16 Be (mg/l) 0.16 <0.10 Mo (mg/l) 0.14 0.17 Zr (mg/l) 0.13 <0.05 Sc (mg/l) 0.10 <0.05 Ti (mg/l) <5.00 <5.00 Ba (mg/l) <0.50 <0.50 Nb (mg/l) <0.50 <0.50 Sb (mg/l) <0.50 <0.50 Se (mg/l) <0.50 <0.50 Sn (mg/l) <0.50 <0.50 Te (mg/l) <0.50 <0.50 Th (mg/l) <0.50 <0.50 W (mg/l) <0.50 <0.50 Ga (mg/l) <0.05 <0.05

[0302] Finally, for the EW tests (2), their chemical characterization was executed by Cu.sup.++, H.sup.+, FeT and Cl.sup.−−, whose results are shown in Tables IX and X.

TABLE-US-00009 TABLE IX Test N° Time Cu.sup.++ (g/L) H.sup.+ (g/L) FeT (g/L) Cl (g/L) P1-0 16:20 12.27 151.0 0.71 0.87 P1-1 17:20 12.19 156.05 0.59 0.92 P1-2 18.20 11.63 152.95 0.68 1.15 P1-3 19.20 10.98 156.70 0.75 1.26 P1-4 20.20 10.33 157.21 0.78 0.82 P1-5 21.20 9.85 158.40 0.83 1.26 P1-6 22:20 9.37 158.44 0.86 0.82 P1-7 23:20 8.88 160.42 0.89 0.82 P1-8 00.20 8.56 161.02 0.92 1.03 P1-9 01:20 8.24 160.92 0.94 0.92 P1-10 02.20 7.75 160.91 0.97 1.26 P1-11 03:20 7.43 160.86 1.00 1.15 P1-12 04:20 — — — — P1-13 05:20 7.27 160.57 1.05 1.26 P1-14 06:20 6.78 160.75 0.97 1.47

TABLE-US-00010 Table X Test N° Time Cu.sup.++ (g/L) H.sup.+ (g/L) FeT (g/L) Cl (g/L) P2-0 14.78 149.5 0.57 0.69 P2-1 19:10 13.89 158.6 0.60 0.76 P2-2 21:10 13.37 157.9 0.61 0.87 P2-3 19.20 10.98 156.70 0.75 1.26 P2-4 22:10 13.37 158.9 0.64 0.98 P2-5 23.10 13.76 158.3 0.62 0.64 P2-6  0:10 13.11 159.4 0.64 0.42 P2-7  2:10 12.72 160.6 0.65 0.64 P2-8 00.20 8.56 162.5 0.62 0.30 P2-9  3:10 12.14 160.0 0.66 0.87 P2-10  4:10 12.07 162.4 0.63 0.76 P2-11  5:10 11.82 162.0 0.60 0.64 P2-12  6:10 11.37 161.9 0.65 0.53 P2-13  7:10 11.37 161.7 0.66 0.87 P2-14  8:10 10.53 162.8 0.60 0.64 P2-15  9:10 9.95 166.2 0.64 0.42 P2-16 10:10 10.14 170.5 0.63 0.42 P2-17 11:10 10.01 170.0 0.65 0.53 P2-18 12:10 9.82 170.1 0.65 0.53 P2-19 13:10 9.95 169.4 0.63 0.60 P2-20 14:10 9.56 166.2 0.61 0.26 P2-21 15:10 9.56 168.6 0.63 0.60 P2-22 16:10 9.11 167.0 0.62 0.48 P2-23 17:10 8.20 166.6 0.62 0.55 P2-24 18:10 9.04 172.3 0.60 0.37 P2-25 19:10 8.20 171.1 0.62 0.48 P2-26 20:10 8.01 172.5 0.63 0.55 P2-27 21:10 8.01 171.9 0.62 0.60 P2-28 22:10 7.81 173.8 0.64 0.62 P2-29 23:10 7.69 174.9 0.63 0.50 P2-30 M-1 23:40 7.43 177.0 0.63 0.26 P2-30 M-2 23:40 7.56 176.0 0.65 0.60 P2-30 M-3 23:40 8.11 177.0 0.62 0.62

[0303] In these tables one can clearly see that actually the concentration of the copper declined after the electrolyte solution, which is an indicator of the electrolytic deposit. In keeping with the latter, the concentration of acid increased.

[0304] The electric and volumetric variables applied in this test were the application of a current density of 250 amperes per square meter of cathode and with a flow rate of PLS per cell of 15 liters per minute per square meter of cathode.

[0305] After these experiences, two more tests, E1 and E2, were carried out that were conducted in the same way, with the difference that these tests had a duration of approximately 24 hours, and their final result was a cathode that could be detached and characterized. (As shown in FIG. 9.)

[0306] Because of the low mass, the cathode was not very thick and fell apart when it was detached, as indicated in the photograph (FIG. 9). Despite the latter, three samples were taken from the cathode, upper, middle and lower, for the respective chemical analyses, as presented in Table XI.

TABLE-US-00011 TABLE XI Limit Element Unit detected High Medium Low Cu (%) 0.1 99.800 99.810 99.790 Cu (%) 0 99.991 99.983 99.977 O (ppm) 2 216 210 142 C (ppm) 2 50 37 36 Cl (ppm) 5 19 20 20 Pb (ppm) 0.5 15 107 169 Ag (ppm) 1 6.0 6.0 6.0 S (ppm) 2 3.0 6.0 6.0 Ca (ppm) 1 1.0 1.0 1.0 As (ppm) 0.1 <0.1 <0.1 <0.1 Bi (ppm) 0.1 <0.1 <0.1 <0.1 Sb (ppm) 0.1 <0.1 <0.1 <0.1 Se (ppm) 1 <1 <1 <1 Sn (ppm) 1 <1 <1 <1 Te (ppm) 1 <1 <1 <1 Co (ppm) 1 <1 <1 <1 Cr (ppm) 1 <1 <1 <1 Fe (ppm) 1 <1 <1 <1 Mn (ppm) 1 <1 <1 <1 Ni (ppm) 1 <1 <1 <1 Cd (ppm) 1 <1 <1 <1 Zn (ppm) 1 <1 <1 <1 P (ppm) 1 <1 <1 <1 Si (ppm) 5 <5 <5 <5

[0307] Based on the results obtained (direct analysis), the purity of the cathode, determined by electrogravimetry, was an average of 99.80%; we can say that it was even as to depth, because the small difference observed could be related to an experimental error. With regard to the contaminant, principally Pb, it is clearly stratified, and its origin cannot be other than from the anode, as the PLS did not have this element in an amount to deposit equivalent to what was analyzed.

[0308] The conclusion of these tests was that starting with a low-grade ore or a PLS with initial concentrations of copper with a low marketability of between 8 and 22 g/L, if the PLS is conditioned adequately and the direct electrowinning system is operated, electrodes of top quality copper can be obtained without requiring the concentration of the copper with solvents (SX).

[0309] The final product of the EW-LED process was a cathodic copper in sheets of approximately 42 kg with an area of 1 m.sup.2 and a purity equal or higher than 98%.

Example of Application 2

[0310] In a real test in a pilot plant, 83.3 tons of ore are required daily, with a size 100% below 1.27 cm and it is accumulated in a stock pile (with an autonomy of 7 days and 583 tons of reserve).

[0311] Then this material, ground and sieved, is agglomerated until it has a 10% humidity with a proportioning of sulfuric acid of 10 kg/ton, generating a flow of damp ore of 92.5 ton/day. These data can be seen in the following table XII.

TABLE-US-00012 TABLE XII Balance of materials: Crushing and Agglomeration Mass flow Mass Flow Design condition t/d t/d A Crushed ore to Stock Pile-Agglomeration 83.3 6.94 B Sulfuric Acid to Agglomeration 0.85 0.07 C Process water to agglomeration 8.40 0.70 D Agglomerated ore to LX 92.5 7.71

[0312] In this process, the lixiviation is a continuous process, 24-hours a day, with an annual processing time of 358 days of ore. With a total cycle of 100 days of ore processing per pile. To be able to comply with this, a daily dry flow of ore of 83.3 ton/day is handled, with the same flow of gravel to the dump.

[0313] The real requirements of ore in the lixiviation, at the pilot plant, were the following according to table XIII

TABLE-US-00013 TABLE XIII Solution, Flow of Number Balance of materials: volume of solution [H.sub.2SO.sub.4] of cycles lixiviation Design flow t/h [Cu.sup.2+] (purity) # LX condition M.sup.3/h (L/h) (kg/h) g/L g/L (%) [EW]  1. Raffinate drive 30.2 38.6  5.22 132 from EW  2. Raffinate available 51.9 66.2  5.22 132  0 from EW (by balance) (*)  3. Raffinate to 52.4 63.9  4.66 132  0 watering (*)  4. ILS to pool 52.2 63.0 12.2  57.6 10 (**)  5. ILS in recirculation 52.4 63.9  5.40 125  1 to watering  6. Process waters to  1.93  1.93 operations  7. LX emergency  0.03  0.03 service water  8. Water to mixed  0.48  0.48 pool (Raffinate Sec.)  9. Water to ILS Pool  0.28  0.28 10. Sulfuric acid to  2.18 (98) operations 11. Sulfuric acid to  0.07 (98) Agglomeration 12. Sulfuric acid to  0.00 (98) ILS pool 13. Process water to  1.14  1.14 EW 14. Sulfuric acid to  2.11 (98) EW 15. Evaporation Mixed  0.05  0.05 Pool (Raffinate Sec.) 16. Evaporation ILS  0.10  0.10 Pool 17. Evaporation Mixed  0.05  0.05 Pool (PLS Sec.) 18. PLS to Mixed Pool 52.0 62.8 13.0  50.3 11 (PLS Sec.) (***) 19. PLS available to 51.9 62.7 13.0  50.4 11 EW (by balance) (***) 20. PLS drive to EW 30.2 36.5 13.0  50.4 11 (***) (*) Fresh raffinate from EW (**) Characterization of solution considering up to the cycle prior to final sending to EW. (***) Characterization of solution in quality of PLS.

[0314] Once the lixiviation process and preparation of the PLS for its electrowinning have been executed, this latter process requires a continuous operation of 24 hours per day, with the same annual ore processing time of 358 days. The cathodic cycles in the EW-LED process are of 5 days, with a gathering of 42 kilograms of copper per cathode with a purity above 97%, an operational current density of 300 A/m.sup.2. The PLS flow entering the bay is 30.2 m.sup.3/h, with six cycles of the volume of PLS entering the bay.

[0315] The data and characteristics for the operation of the EW-LED process are described in the following tables XIX and XX.

TABLE-US-00014 Table XIX Solution, Flow of Balance of materials: volume of solution [H.sub.2SO.sub.4] Number of Electrowinning-LED flow t/h (kg/h) [Cu.sup.2+] (purity) cycles Design condition M.sup.3/h (L/h) [t Cu/h] g/L g/L (%) # LX/[EW] 21. PLS to conditioning 30.2 36.5 50.4 101 (*) 22. Sulfuric acid to line 2.11 (98) 23. Process water to line (12.1) (12.1) 24. PLS to E/E heat 30.2 38.0 13.0 120 [0] (*) exchanger 25. PLS to E/A heat 30.2 38.0 13.0 120 [0] (*) exchanger 26. Cobalt sulfate to (21.6) Cobalt sulfate TK 27. Guar to Guar TK (0.01) 28. Solution of cobalt (433) (433) sulfate 29. Guar solution to (1.29) (1.29) distribution 30. Guar solution to EW 1 (0.32) (0.32) Bank 31. Guar solution to EW 2 (0.32) (0.32) Bank 32. Guar solution to EW 3 (0.32) (0.32) Bank 33. Guar solution to EW 4 (0.32) (0.32) Bank 34. PLS to EW 1 Bank 30.2 38.0 13.0 120 [0] (*) 35. PLS to EW 2 Bank 30.2 38.0 12.6 120 [1]    36. PLS to EW 3 Bank 30.2 38.0 12.3 121 [1]    37. PLS to EW 4 Bank 30.2 38.0 12.0 121 [1]    38. PLS to Transfer TK 30.2 38.0 11.7 122 [1]    39. PLS in recirculation to 30.2 38.6 6.51 130   [5] (**) pool 40. Hot water from heater 21.1 21.1 41. Hot water to Cobalt (411) (411) sulfate TK 42. Hot water to Guar TK (1.27) (1.27) Table XX Solution, Flow of Balance of materials: volume of solution [H.sub.2SO.sub.4] Number of Electrowinning-LED flow t/h (kg/h) [Cu.sup.2+] (purity) cycles Design condition M.sup.3/h (L/h) [t Cu/h] g/L g/L (%) # LX/[EW] 43. Hot water to heat 20.2 20.2 exchanger 44. Hot water to 0.50 0.50 cathode washing 45. Hot water runoff 20.2 20.2 46. Process water to 1.13 1.13 services and operation 47. Water to EW-LED 0.03 0.03 emergency service 48. Service water to 0.19 0.19 human consumption 49. Process water to 0.91 0.91 replenishment 50. Water to heater 21.2 21.1 51. Petroleum to heater 0.09 52. Evaporation of (1.00) (1.00) water in bay 53. Decomposition of (11.1) (11.1) water in bay 54. Cathodic copper (0.04) 55. Discharge of 30.2 38.6 5.22 132 [6] (***) Raffinate to LX (*) Fresh raffinate from LX (**) Characterization of solution considering up to the cycle prior to the final runoff to LX. (***) Characterization of solution as raffinate.

[0316] Under the conditions described previously, copper cathodes are obtained weighing 42 kilograms, high purity, above 97%, such as those seen in application example 1.