MXPA98001138A - Fluorocarburo fluids as gas carriers to help lixiviation operations of amounts of precious metals and not price - Google Patents

Abstract

This invention provides an improved method for leaching uncondensed minerals from precious and non-precious metals, from a mineral, which comprises contacting a mineral having metals or mineral concentrate with an oxidizing gas solution comprising a solution of an oxidizing gas dissolved in a perfluorinated liquid, said contact is made before or during the extraction of the precious or non-precious metal from the mineral. In still another aspect, this invention provides a method for destroying cyanide from a solution containing cyanide, comprising contacting said solution with a solution of a dissolved oxidant gas, in a perfluorous liquid.

Description

FLUOROCARBURO FLUIDS AS GAS CARRIERS FOR HELPING LIONIVIATION OPERATIONS OF AMOUNTS OF PRECIOUS AND NON-PRECIOUS METALS FIELD OF THE INVENTION This invention relates to a process for the recovery of precious metals and non-precious metals from low-grade minerals. More particularly, the present invention relates to the recovery of minerals and riches of precious and non-precious metals through the mining industry 8r leaching. In still another aspect, this invention relates to a process for decreasing the concentration of residual cyanide in the sterile leach leach. In still another aspect, this invention relates to a method for destroying cyanide from a solution containing cyanide, comprising contacting said solution with a solution of an oxidizing gas dissolved in perfluorinated liquid.

BACKGROUND OF THE INVENTION The leaching of clusters or heaps of ore has been a preferred method of recovery of REF: 26767 precious metals such as gold and silver, and non-precious metals such as copper, of their corresponding minerals. Sometimes also referred to as solution mining, the leaching of clumps or heaps involves the extraction of soluble metals or salts from ore or ore by distribution of solutions, or leaching, during a leaching cycle on a pile of open ore piled on top of a waterproof base. Heap leaching can also be done through tank leaching or tank leaching and agitation mining. Typically, diluted aqueous alkali metal cyanide is used as a leach for the extraction of gold and silver, and dilute aqueous sulfuric acid is used as a leach for the recovery of copper. Recovery of unconcentrated gold and silver ores from low grade ores using oxidative cyanidation is well known. See, for example, 11 Kir -Othmer Encyclopedia of Chemical Technology 972-92 (3rd ed., 1979). Such recovery methods have been used commercially in the late 1960s. Typically, in the leaching of heaps of gold ore, a dilute aqueous solution of sodium cyanide and limestone, which has a pH of between about 10.5 and 12.5, is distributed over the top of a pile of ore. The heaps of ore are generally on average approtely 100,000 to 500,000 tons in weight and contain ore parts in the range of less than 12.7 mm (1/2 inch) to more than 15.24 cm (6 inches) in diameter and are stacked on a waterproof base. The gold is dissolved in an aerated solution of cyanide according to the following two-step reaction mechanism: 1 ^ 2Au + 4CN ~ + 02 + 2H20 = > 2Au (CN) 2 ~ + 20H ~ + 2H202 II. 2Au + 4CN "+ H202 => 2Au (CN) 2" + 20H " J.B. Hiskey, Arizona Bureau of Geology and Mineral Technology Fieldnotes. Vol. 15, No. 4, Winter 1985. The gold formed in complex is then recovered from the pregnant leaching aqueous solution, usually by adsorption on activated carbon, and the complex is subsequently removed and converted to elemental gold by electrolytic extraction. The sterile cyanide solution is then recirculated to the pile for subsequent leaching, with some replenishment of cyanide, if necessary. The leaching of the silver is carried out analogously, forming the complex of Ag (CN) 2 ~ from which the elemental silver is usually recovered by replacing zinc metal powder. Currently, most of the copper produced by hydrometallurgical processing is recovered from the leaching of oxide or mineraj.es from secondary copper sulphide in heap and deposit leaching operations. Leaching is typically carried out by spraying a dilute solution of sulfuric acid on top of heaps of crushed ore, allowing the acid to trickle through the heaps and dissolve copper mineralization over a period of several weeks. or months Such hydrometallurgical recovery of copper through the leaching of primary sulphide deposits, is considered difficult and not economic due to the refractory nature of copper mineralization and does not lend itself to leaching by sulfuric acid, unless oxidative conditions are present during the leaching cycle. The recovery of unconcentrated copper ore from primary minerals and sulfides such as chalcopyrite is typically limited to the conventional processing of pyrometallurgical material by mining, crushing and rotating ore, followed by electrolytic copper melting and refining. According to Hiskey, supra, compared to conventional milling (for example, leaching by crushing, milling and stirring), the recovery of gold and silver by heap leaching offers several advantages, including low capital costs and of operation, shorter startup times, and lower environmental risks. Such advantages are, however, displaced to a certain extent by lower metal extractions. Typically, only 60 to 80 percent of the unconcentrated minerals of precious and available materials can be recovered using state-of-the-art heap leach techniques. Because many large pieces of ore in the piles are poorly wetted, they are poorly extracted. At the opposite end, when the larger pieces of ore are crushed to smaller pieces to improve extraction, fine powders are produced that can plug the pile and especially in its lower part, reducing the flow rate of the leachate through the pile .

The techniques of beneficiation are sometimes used to increase the recoveries of metals from minerals, over those obtained by conventional methods. The beneficiation techniques encompass many and varied processes, all with the design of concentrating the ore for subsequent processing and extraction. Shredding and agglomeration are the most widely used beneficiation technologies for the recovery of gold and silver. See, N.C. Wall, and collaborators, Gold Beneficiation. Mining Magazine (May 1987) (which describes the recent developments in beneficiation technology in gold mining). See also A.K. Biswas & W. G. Davenport, Extractive Metallurgy of Copper, Pergamon (3rd ed., 1994) (which describes concentration techniques for copper extraction). Oxidative treatments are also sometimes used to increase the recovery of metals from minerals that, due to their particular characteristics, show little recovery through conventional leaching processes. These so-called refractory minerals may, for example, contain significant concentrations of clays that prevent uniform percolation of the leach, or may contain other materials that consume the leach. The oxidizing agents can be used to alter the mineralization of the sulphide and carbonate gangue in such minerals, whereby the leaching channels are opened and / or the insoluble forms of the metals, such as sulfides, are converted to other forms. easily soluble in leaching agents, for example, oxides and chlorides. Oxidizing gases, including oxygen, ozone, chlorine, and chlorine dioxide, have been used as oxidizing agents, but due to their relatively high vapor pressure, such gaseous oxidizing agents require sealed and pressurized units or batch-type containers, limiting of volume. To provide management convenience and cost effectiveness, oxidizing reagents are most frequently applied from the solution -acuosa The widely used aqueous oxidizing solutions include hypochlorous acid (generated from sodium hypochlorite), hydrogen peroxide, and nitric acid. These aqueous solutions are, however, less reactive than their gaseous counterparts frequently requiring high temperatures and significant agitation to complete the oxidation process of the ore in a reasonably timely manner. The use at elevated temperatures can also lead to a decrease of the oxidant from the hydrolysis, which prevents the distribution of the oxidizing agent to the mineral particles in its most reactive and efficient form. As a further drawback to using aqueous oxidizing solutions, large volumes of corrosive and polluting aqueous effluent are created stream-downstream of the leaching process, which must be processed and treated, adding additional costs to the process of complete metal recovery. Attempts have been made in recent years to provide oxidative treatments comprising saturated aqueous solutions of oxidizing gases. US Patents Nos. 3,846,124, 4,038,362 and 4,259,107, all to Guay, explore the use of chlorine gas to increase the recovery of gold from gold-bearing sedimentary minerals by suspending the mineral with water and saturating the suspension with chlorine gas before of cyanidation In a similar vein, U.S. Patent No. 4,979,986 (Hill et al.) Describes a method for oxidizing the gold-bearing mineral by contacting an aqueous suspension of ore with gaseous chlorine or hypochlorite salt and attaching the suspension to high shear using a propeller. In addition, US Pat. No. 4,289,532 to Matson et al. Describes a process for the recovery of unconsolidated gold ores from carbonaceous ores by the formation of an aqueous alkaline mineral suspension., subjecting the suspension to oxidation with an oxygen-containing gas, intimately contacting the oxygenated suspension with a source of hypochlorite ions, and subsequently contacting the suspension with a complexed, cyanide agent. In recent years, several experimental techniques for oxidative recovery and copper leaching have been proposed, including: (a) the leaching of ferric and cupric chloride followed by solvent extraction and electrolytic extraction of copper powder; (b) pressure leaching with sulfuric acid-oxygen, followed by direct electrolytic extraction of the copper produced; (c) leaching of pressure agitation with ammonia-oxygen, followed by copper reduction and solvent extraction and electrolytic extraction (Escondida and Arbiter processes); and (d) oxidative calcination of the ore, followed by leaching by sulfuric acid. All these processes incorporate the concentration of the mineralization by crushing and flotation by mineral foam, to eliminate unwanted gangue minerals before leaching, to ensure adequate recoveries of copper with minimum reagent waste. BRIEF DESCRIPTION OF THE INVENTION In summary, in one aspect, this invention provides an improved method of leaching unconcentrated ores of precious metals and non-precious metals from ore, which comprises contacting a mineral having metal or a mineral concentrate with a solution. of oxidizing gas, comprising an oxidizing gas dissolved in a perchlorinated liquid, said contacting being carried out before or during the extraction of the precious or non-precious metal from the mineral. In yet another aspect, the present invention provides a method for the destruction of cyanide from the sterile cyanide-containing leachate, which comprises contacting the leaching solution with a solution of an oxidizing gas dissolved in a perfluorinated liquid. In another aspect, this invention relates to a method for destroying the cyanide of the cyanide-containing solution, comprising contacting said solution with a solution of an oxidizing gas dissolved in a perfluorinated liquid.

BRIEF DESCRIPTION OF THE DRAWINGS FIGURE 1 is a graph of the daily gold extractions recovered from two column tests of a precious metal mineral, a column leached only with a cyanide-containing leach and another column interrupted during the leaching cycle during the leaching cycle. days 53 and 54, for treatment with an oxidizing gas solution, according to this invention.

FIGURE 2 is a graph of the daily silver extractions, recovered from the same two tests on columns of a precious metal ore as presented in FIGURE 1, comparing the relative recoveries obtained with and without the oxidative treatment of according to this invention.

DETAILED DESCRIPTION OF THE INVENTION The recovery of precious metals, for example, gold and silver, and non-precious metals such as copper, is greatly improved in leaching mining applications by treating the ore that holds the metal or ore concentrate before or after during the leaching cycle, with a solution of an oxidizing gas dissolved in a perfluorinated fluid (hereinafter referred to as the "oxidizing gas solution"). The oxidizing gas solutions used in this invention provide a means for distributing a stable solution of an oxidizing gas in its more active non-hydrolyzed state. These oxidizing gas solutions offer the additional advantage of providing a medium of very low surface tension (generally in the order of approximately 15 dynes / cm), thereby making it possible for the oxidizing gas solution to make contact efficiently and penetrate perfectly by the mineral particles. The oxidizing gas solutions can be contacted with the mineral using any of the conventional processes currently used to distribute and disperse aqueous cyanide or aqueous sulfuric acid leaching solutions. Typically, these methods include pumping the oxidizing gas solution to the top of a heap or cluster of material, and allowing the solution to percolate slowly through the pile. This process can be performed under environmental conditions, and does not require special equipment. A conventional leaching cycle using a suitable leach is performed to recover the desired material, and the spent, non-miscible, oxidizing gas solution can be recycled and regenerated to cool the oxidizing gas solution by dissolving the spent solution, from a supply of oxidizing gas replenishment. The ore treatment can be carried out either before or during the leaching cycle, or extraction. The oxidation treatment can therefore constitute a pretreatment cycle performed on the ore before the first leaching cycle, or it can be carried out in stages in relation to the leaching cycle, to form one or more cycles of the alternate oxidation and leaching treatment . The leaching process may require the crushing or grinding of mineral particles before oxidation treatment. The crushing of minerals that possess metals is mainly required to release precious metals and non-precious metals, and the minerals that have such metals, to make them more suitable for extraction. The degree of comminution or grinding that may be required depends on many factors, including the size of the release of the metal, the size and nature of the host minerals and the method or methods to be applied for the recovery of the metal. The optimal distribution of particle size is dictated by economic considerations; a balance between the amount of gold to be recovered, the processing costs, and the crushing costs. For a review of these considerations see, for example, J. Marsden & I. House, "Chemistry of Gold Extraction", pp. 35, 105-06, Ellis Hor ood Ltd., 1992 ISBN 0-13-131517-X. The degree of crushing of the ore, required for crushing or pulverization, to achieve gold permeability, uniform distribution of the leaching solution, and satisfactory gold extraction, has a maximum effect on the economy of the projected heap leach operation . Any excessive proportion of clay in the ore or an excessive amount of fine dust generated by the crushing can slow down the rate of percolation of the leaching solution causing non-productive contact of the leach with the mineral in the heap. Agglomeration of the crushed ore is often required to overcome these difficulties, and to achieve a permeable and uniform feed throughout the entire ore pile. Crushing circuits and agglomeration systems can sometimes be employed, where their capital costs are justified by the economy of the entire process. Typical leaching operations of gold and silver heaps can incorporate multiple crushing stages, where justified, to make the ore more suitable for heap leaching and subsequent dissolution and recovery of valuable metals. The oxidizing gases useful in making the oxidizing gas solutions of the present invention include any gas capable of oxidizing the metal or metallic mineral, which is also readily soluble in a perfluorinated fluid. Such gases include, for example, chlorine, ozone, chlorine dioxide, and sulfur dioxide. Fluorinated fluids useful in this invention are compounds that contain a high level of fluorine attached to carbon, which are liquid at the operating conditions of the leaching process, for example, having a boiling point near or higher than room temperature, and have a freezing point below room temperature. These fluorinated fluids must be capable of dissolving a substantial amount of an oxidizing gas at operating conditions, typically in a temperature range from about 0 ° C to about 50 ° C. Preferred fluids will dissolve at least 500 ml of chlorine gas per 100 ml of fluid at 1 atmosphere, and 25 ° C. Most fluorinated fluids will dissolve at least 1200 ml of chlorine gas at 1 atmosphere and 25 ° C. Preferably, the oxidizing gas solutions used according to the methods of this invention will be saturated with the chosen oxidant gas. Fluorinert® fluids, product of the bulletin, 98-0211-8301-1 (65.05) R, distributed on 5/95, available from 3M Co. , St. Paul, Minn., Provides the solubility of these oxidizing gases in Fluorinert ™ Electronic Fluids.

Useful, specific fluorinated fluids include perfluoroaliphatic and perfluorocycloaliphatic compounds having from 4 to about 18 carbon atoms, preferably from 4 to 10 carbon atoms, which may optionally contain one or more catenary heteroatoms, such as divalent oxygen atoms or trivalent nitrogen . The term "perfluorinated fluid" as used herein, includes organic compounds in which all (or essentially all) of the hydrogen atoms are replaced with fluorine atoms. Representative perfluorinated liquids include perfluoroalkanes, cyclic and non-cyclic, cyclic and non-cyclic perfluoroamines, cyclic and non-cyclic perfluoroethers, cyclic and non-cyclic perfluoroaminoethers, and any mixtures thereof. Representative, specific perfluorinated liquids include the following: perfluoropentane, perfluorohexane, perfluoroheptane, perfluorooctane, perfluoromethyleryclohexane, perfluorotribuyamine, perfluorotriamylamine, perfluoro-N-methylmorpholine, perfluoro-N-ethylmorpholine, perfluoroisopropylmorpholine, perfluoro-N-methyl -pyrrolidine, perfluoro-1,2-bis (trifluoromethyl) hexafluorocyclobutane, perfluoro-2-butyltetrahydrofuran, perfluorotriethylamine, perfluorodibutyl ether, and mixtures of these and other perfluorinated liquids. Commercially available perfluorinated liquids that can be used in this invention include: Fluorinert ™, FCMR-43 Electronic Fluid, Fluorinert ™ FCMR-72 Electronic Fluid, Fluorinert ™ FCMR-77 Electronic Fluid, Fluorinert ™ FCMR-84 Electronic Fluid, Fluorinert ™ FCMR-87 Electronic Fluid, Performance FluidMR PF-5060, Performan.ee FluidMR PF-5070, and Performance FluidMR PF-5052. Some of these liquids are described in Fluorinert ™ Electronic Fluids, product bulletin 98-0211-6086 (212) NPI, distributed 2/91, available from 3M Co. , St. Paul, Minn. Other commercially available perfluorinated liquids which are considered useful in the present invention include perfluorinated liquids sold as Galden ™ LS fluids and Flutec ™ PP fluids. The oxidizing gas solutions of the present invention can be used to rapidly oxidize the residual cyanide in the sterile cyanide leach at a sufficiently low concentration to facilitate the release of the sterile leach into the environment. These oxidizing gas solutions may comprise any of the aforementioned cases dissolved in a perfluorinated liquid, or they may comprise a solution of oxygen dissolved in a fluorinated liquid. The destruction of cyanide with a chlorine oxidizing gas solution proceeds by the following reaction mechanism: (I) CN "+ Cl2 < > CNCl (g) + Cl" (II) CNCl (g) + 20H "< > CNO" + Cl "+ H20 (II) 2CNO" * 3C10"+ H20 < > N2 + 2C02 + 3C1"+ 20H" The residual cyanide in a waste leaching stream can be reduced or eliminated by contacting the sterile leach stream with fresh oxidizing gas solution, for example, comprising chlorine gas dissolved in a fluorinated liquid proceeding according to the above mechanism or which comprises dissolved oxygen gas in a fluorinated fluid, proceeding by means of a mechanism -analog. The oxidizing gas solution can also be used to rinse the leached ore, thereby reducing the amount of residual cyanide remaining in the ore that is contacted. These techniques can be used to reduce the concentration of cyanide in the sterile leach stream to levels below 0.20 parts per million. In still another aspect, this invention relates to a method for destroying cyanide from a solution containing cyanide, comprising contacting said solution with a solution of an oxidizing gas dissolved in a perfluorinated liquid.

Examples The following examples are offered to help better understand the present invention. These examples do not have to be considered as an exhaustive compilation of all the modalities of the present invention, and are not unnecessarily considered as limiting the scope thereof.

Comparative Example Cl For comparison purposes, a column study was conducted for a period of 79 days to develop a baseline for the extraction of gold and silver from low grade ore, leached by an aqueous alkaline cyanide leach. For Comparative Example 1, approximately 1360 kg (3000 pounds) of leached free material was obtained from Coeur Rochester, Inc., a heap leach operation in Nevada. The ore was mixed without subsequent crushing, then sieved through sieves of four different mesh sizes: 1.3 cm (1/2 inch), 1.0 cm (3/8 inch), 0.6 cm (1/4 inch) and mesh 10 using a Gilson Test Master mesh device. The sieve analysis for each size fraction for the column test feed was then calculated based on the relative weight percent of each size fraction from the sieve analysis of the test feed. A test load of 90 Kg (200 pounds) was reconstituted by weight and size distribution, and was placed in a 200 L (55 gallon) steel drum. A composite sample was also produced for the main test and the individual analysis of the gold and silver sieve fraction, by means of an ignition test and atomic absorption methods. A surfactant solution was also prepared, containing a surfactant of the structure C5FnO (CF2) 5COO "H4NA for addition to the mineral." The fluoroaliphatic surfactant was prepared as follows: 118.2 g (1.0 mol) of hexan-1 were stirred at reflux. 6-diol, 4.4 g of quaternary ammonium salt Adogen ™ 464 (available from Witco Corp.), 80.0 g (2.0 mol) of sodium hydroxide and 250 ml of tetrahydrofuran, 80 ml of deionized water were added to this mixture to facilitate After about 20 minutes, 151 g (1.0 mol) of n-pentyl bromide was added in 30 minutes, and the contents were stirred overnight at reflux.The reaction mixture was purified using a rotary evaporator. The resulting purified product was added with 100 ml of chloroform, then 150 ml of acetyl chloride was added dropwise and the mixture was subsequently heated to reflux for 4 hours.The solvent was removed to produce the crude product. C5HnO (CH2) 6OC (0) CH3 was distilled at 125 ° C (3 torr) and the distillate was fluorinated by direct fluorination as described, for example, in US Patent No. 4,894,484 (Lago et al.). The fluorinated ester was treated with an aqueous solution of sodium hydroxide at 23% by weight, and acidified with 50% by weight aqueous sulfuric acid. The addition of 3M Fluorinert R FCMR-75 electronic fluid, a 3M available perfluorinated fluid consisting mainly of a mixture of perfluoro (2-butyltetrahydrofuran) and perfluorooctane, and mixing produced a clear 2-phase system. The lower phase was purified and distilled to a substantially pure acid product, C5FnO (CF2) 5C02H, which boils at 90-110 ° C at 0.4 torr. The fluorinated carboxylic acid was treated with an excess of dilute aqueous ammonia, to form the ammonium salt which was dehydrated by lyophilization to a solid and dissolved in water to form a stock solution of 5% by weight solids. The reserve surfactant solution was then diluted with water to form 10 liters of a surfactant solution at 250 ppm. The drum containing the previously described mineral was placed on an inclined roll mixer, 90 g of calcium oxide (limestone) was added to each drum, and 4 L of the surfactant solution was added to the drum, slowly during mixing , to achieve the uniform distribution of the solution. The drum was agitated for about 5 minutes, then the agglomerated ore was discharged into a column 1.8 m (6 ft.) High by 20 cm (8 in.) In diameter with the sides constructed from a pipe section. polyvinyl chloride (PVC) and the bottom constructed from a circular PVC sheet with a hole, with a piece of PVC tubing 2.54 cm (1 inch) inserted tightly through the hole in the bottom from the outside, and connected to the other end through a hole in the cap of a plastic cylinder with narrow mouth 3.8 L (1 gallon) covered. The leach was prepared by loading the following ingredients in a 200 L (55 gallon) drum. First, approximately 120 L of water was added to the drum, followed by 120 g of calcium oxide (1 g of CaO / L H20). The solution was allowed to mix for 3 hours, then the exact amount of water charged to each drum was calculated by grinding the hydroxide ion produced by the reaction of the calcium oxide with the water. Sodium cyanide was added to 0.5 g of NaCN / L of water, the surfactant buffer solution was added to give the desired ppm level, and the resulting solution was mixed again for approximately 2 hours at ambient conditions (approximately 95 ° F). or 35 ° C). The final pH of the leach was in the range of 10-11.5. For each day of the leaching study, fresh leach from each drum was used. The concentrations of each ingredient in the leach were determined before leaching each day, due to possible degradation under aerobic conditions of high pH. The leach was crushed for cyanide and limestone, and the appropriate amount for replenishment was calculated and added. As a daily procedure for the leaching test, the leach was applied to the top of the ore column at a rate of 210 ml / min / m2 (0.005 gal / min / foot) for each of the first 15 days, followed by 42 ml / min / m2 (0.001 gal / min / ft2) for each of the final 66 days. The solutions were collected daily from the edge of the column and subjected to the analysis for gold and silver using conventional methods of atomic absorption. Gold and silver concentrations were reported in parts per million (ppm). The total amount of gold and silver present in each daily pregnant leach solution (μg) was calculated by multiplying the volume of the pregnant lixiviate (L) times the metal concentration (ppm). FIGURE 1 graphically represents the results of this study for the recovery of gold in direct comparison to Example 1, where the column was subjected to an oxidation treatment using an oxidizing gas solution of this invention during the leaching cycle. FIGURE 2 presents the analogous results for the recovery of silver.

Example 1 For Example 1, a column study using an aqueous alkaline leach to leach the gold and silver from the ore was run in the same manner as Comparative Example Cl, except that the flow of leach was interrupted during days 53 and 54. of the test to treat the column with an oxidizing gas solution according to the method described by this invention. The oxidizing gas solution prepared by bubbling chlorine in 4 liters of Fluorinert ™ Electronic Fluid FCMR-75 until the fluid was saturated with chlorine gas (for example, when the solution no longer turned dark yellow-green and the chlorine gas remained on the surface of the solution). The oxidizing gas solution was then introduced into the upper part of the column, allowing the solution to come into contact and percolate through the mineral particles in the column, and come out through the bottom of the column, from the same way that the pregnant leaching solution comes out. The oxidizing gas solution was added in a period of 2 days (for example, during days 53 and 54 of the leaching test), at a rate of 42 ml / min / m2 (0.001 gal / min / ft2) for a total volume of 1.5 L added every day. The oxidizing solution that entered the column was yellow, due to the presence of chlorine.

The oxidant solution that came out of the column was white as water indicating the absence of chlorine (for example, essentially all of the chlorine had reacted with the mineral in the column). The drainage of the alkaline cyanide leach liquid and the oxidizing gas solution were collected from the bottom of the column and their respective volumes were determined by the use of a separatory funnel and a graduated cylinder. After two days of treatment with the oxidizing gas solution, leaching was resumed with the alkaline cyanide leach at day 55 and continued until the end of the test at day 79. FIGURE 1 graphically presents the results of this study for the recovery of gold in direct comparison to Example 1, where the column was not subjected to an oxidation treatment. FIGURE 2 presents the analogous results for the recovery of silver.

The data in FIGURES 1 and 2 show that, compared to the results of baseline leaching from this mineral without oxidation treatment, the daily gold and silver recoveries in the pregnant leach solution rose immediately and remained unchanged. elevated on days 55-79 after the mineral in the column was treated with the oxidizing gas solution at days 53-54 of the test. Table 1 shows the daily average recoveries of precious metals of Example 1, before and after the ore was treated with the oxidizing gas solution (for example, day 31 to day 52 and day 57 to day 79 respectively). The data presented in parentheses correspond to the average daily recoveries of precious metal of Comparative Example Cl, where the ore was not treated with the oxidizing gas solution.

Table 1 Table 1 (continued) The data in Table 1 show that the average daily gold recovery increased after the ore was treated with oxidizing gas solution, while the average silver recovery was only slightly lower than before the treatment of the column. In contrast, the ore that was not treated with the oxidizing gas solution showed much lower daily gold and silver recovery from day 57 to day 79, compared to day 31 to day 52.

Example 2 In Example 2, an experiment was run to demonstrate the use of an oxidizing gas solution to destroy by oxidation the residual cyanide left on the column of the leach test. Using the same procedure as described in Comparative Example Cl, a column containing mineral particles was leached., with aqueous cyanide leach for 81 days. By titration, the concentration of the cyanide in the pregnant solution was measured, on day 81, as approximately 0.5 g / L (500 ppm). For the next 48 hours (days 82 and 83 leaching), the mineral column was rinsed with a saturated solution of chlorine in FluorinertMR Electronic Fluid FCMR-75 at a rate of 42 ml / min / m2 (0.001 gal / min / feet2). After rinsing with the oxidizing gas solution, the column was rinsed daily with water at 42 ml / min / m2 (0.001 gal / min / ft2) and the rinse solution was titrated daily for the cyanide concentration, with the objective of reduce the concentration of cyanide in the rinse, to less than 0.20 parts per million. The aqueous rinse that came out of this column contained less than the objective concentration of 0.20 ppm of cyanide on day 107 of leaching.

Comparative Example C2 In Comparative Example C2, the same experiment as in Example 2 was run, except that instead of rinsing the column for 48 hours during leaching days 82 and 83 with the oxidizing gas solution, the column was rinsed with water. In other words, only water was used during the complete rinse test. The aqueous rinse leaving this column contained less than the target concentration of 0.20 ppm of cyanide on day 114 of leaching, one full week after when the oxidizing gas solution was used prior to rinsing with water.

Example 3 In Example 3, a mineral column was oxidized twice during the leaching cycle with an oxidant solution consisting of a saturated chlorine solution in the Fluorinert ™ Electronic Fluid FCMR-77 (a perfluorinated fluid available from 3M and consisting primarily of a mixture of perfluoro (2-butyltetrahydrofuran) and perfluorooctane) and the effect on the daily recovery of gold and silver was measured. The mineral used for this experiment was a spent gold-silver ore obtained from the Coeur Rochester mine near Lovelock, Nevada, which had previously been leached with cyanide leach for a period of approximately 2 years having approximately 50% of its silver and approximately 5% of your gold remaining. The complete sample of the spent mineral was sieved on sieves of four different sizes of mesh: 1.3 cm (1/2 inch), 1.0 cm (3/8 inch), 0.6 cm (1/4 inch) and 10 mesh using an apparatus of Gilson Test Master mesh. The test loads weighing 270 kg were reconstituted based on the particle size distribution of the complete sample. A representative portion of each individual fraction of the sieve was separated by rake, pulverized and subjected to analysis for gold and the silver by the ignition test. A composite mineral sample was prepared from pulverized material and subjected to gold and silver analysis by the ignition test. A 270 kg ore test load was placed in a drum agglomeration apparatus as described in Comparative Example Cl and agglomerated with Water from the tap of the City of Tucson. The agglomerated test load was loaded on a polyvinyl chloride column 6.1 m high by 20 cm in diameter, with a drain pipe inserted tightly through its bottom and leached at an irrigation rate of 105 ml / min. / m2 (0.0025 gal / min / ft2) with an aqueous leaching feed solution containing 1 kg (2.0 pounds) of sodium cyanide per ton of solution (at 0.10% by weight) and enough calcium oxide to maintain the pH of the solution in the range of 10.5 to 11.5. After 6 days of leaching, a saturated, limestone, aqueous leaching solution containing 1 kg of sodium cyanide (2 pounds) per ton of solution (0.10% by weight) and having an aqueous leaching solution was introduced into the column. pH above 12.0, to raise the pH of the aqueous pregnant solution, which had moved below 10. This aqueous leaching feed solution saturated with limestone was continued for the following 36 days (for example, days 7). -42) until the pH of the pregnant solution was above 10.0. The pregnant solutions were also analyzed for gold and silver using atomic absorption. On day 42, 4.5 L of an Fluorinert ™ FCMR-77 Fluidine Oxidizing Gas solution solution, saturated with chlorine, was introduced into the mineral column in place of the leaching solution and at the same flow rate. The process with the oxidizing gas solution was repeated on day 43, resulting in a total of 9.0 L of added oxidant gas solution in the two-day period. The volumes of pregnant solution with cyanide and spent oxidant gas solution were determined for leaching days after chlorination, by using a separating funnel and a graduated cylinder. For the next 27 days (day 44 to day 70), the column was irrigated with normal cyanide leaching feed solution until day 70, then for the next 4 days with the aqueous leaching feed solution, saturated with limestone (day 70 to day 74) to raise the pH of the pregnant cyanide solution. On day 74, the leaching was discontinued and the column allowed to drain for 48 hours. On days 76 and 77, the procedure was repeated as it was performed on days 42 and 43, for example, by contacting the mineral in the column with the oxidizing gas solution. After the second oxidation of the ore, the column was left for a rest period of 72 hours (day 78 to day 80), then the cyanide leaching was resumed on day 81 with the normal cyanide leaching feed solution.

After six days, the column was irrigated with a feed solution saturated with limestone, in order to increase the pH of the daily pregnant solutions. On day 93, the irrigation rate was increased to 210 ml / min / m2 (0.005 gal / min / ft2) and this was continued until day 111. The proportion or flow rate of leaching was again reduced to 105 ml / min / m2 (0.0025 gal / min / ft2) on day 112. On day 113 the column was rinsed with water for 48 hours and allowed to drain for an additional 48 hours. Simultaneous to this experiment, a second analogous experiment was run with a control mineral column where the ore was not oxidized. Instead of the first oxidation of the ore, the control column was maintained under leaching with cyanide during days 42 and 43. Instead of the second oxidation of the ore, the column was treated with FluorinertMR FCMR-77 alone (for example, not contained chlorine) during days 76 and 77, at the same flow rate and total volume as with the oxidizing gas solution. The gold and silver contents (in μg) of the pregnant solutions were determined daily from the volume of pregnant solution and the metal tests (atomic absorption) for the columns of oxidized mineral with chlorine and control during 4 different periods of time: three weeks before the first oxidation of the mineral (day 22 to day 42), three weeks after the first oxidation of the mineral (day 44 to day 64), three weeks before the second oxidation of the mineral (day 56 to day 76), and three weeks after the second oxidation of the mineral (day 81 to day 101). For the three weeks before the first oxidation of the ore, the pregnant solution from the column of ore oxidized with chlorine had an average of 783 μg per day lower in silver content and no difference in gold content compared to the pregnant solution from the control mineral column. For the three weeks following the first oxidation of ore, the pregnant solution from the column of ore oxidized with chlorine, had an average of 474 μg and 3 μg per day of highest content of gold and silver respectively, compared to the pregnant solution coming from the control mineral column. In this way, the pregnant solution from the column of ore oxidized with chlorine showed a net average daily gain of 1267 μg and 3 μg per day of silver and gold respectively, after the first oxidation compared to the pregnant solution from the Control mineral column. For the three weeks before the second oxidation of the ore, the pregnant solution coming from the column of oxidized mineral with chlorine had an average of 358 μg per day plus silver content and no difference in the gold content, compared to the solution pregnant from the control mineral column. For the three weeks following the second oxidation of the ore, the pregnant solution from the column of ore oxidized with chlorine had an average of 579 μg per day plus chlorine content and no difference in gold content compared to the solution pregnant from the control mineral column. In this way, the pregnant solution from the column of ore oxidized with chlorine showed an average daily gain, net of 221 μg per day of silver, and no difference in gold, respectively after oxidation when compared to the pregnant solution, coming from the control mineral column. The pH of the pregnant solution decreased to less than 8.0 after the second oxidation of the mineral, resulting in lower concentration of alkaline cyanide, which may have masked the potential benefits of the treatment.

Example 4 In Example 4, a small-scale laboratory study was run to show the oxidation utility of the copper sulfide ore, with low recovery with an oxidizing gas solution of this invention, before extraction of the unconcentrated ores of copper metal with aqueous leaching solution of sulfuric acid. A sample of low recovery sulfide ore was obtained from the Phelps-Dodge Tyrone Mine located near Silver City, New Mexico. The dominant copper mineralizations of the mineral sample were chalcopyrite, covelite and chalcocite. The mineral sample had a total copper content of 0.41% with a recoverable value of leaching of 43% of the total copper value. The ore sample was crushed to a nominal size so that 100% ore passed through a 2.54 cm (1 inch) screen. Using a series of mesh sieves with a diameter of approximately 20 cm (8 inches), the sample of crushed material was segregated into 5 fractions of discrete size: -10 mesh, +10 mesh to -7 mesh, +7 mesh mesh - 4, +4 mesh to -2 mesh, and +2 mesh to 2.54 cm (1 inch). This separation by size selection allowed a consistent ratio of surface area to weight of the ore load, when samples are reconstituted during the subsequent copper ore oxidation, and leaching experiments. A load of 600 g of ore comprised of a mixture of 200 g of the mesh fraction +10 to mesh -7, 200 g of the mesh fraction +7 to -4 mesh, and- 200 g of the mesh fraction + 4 a -2 mesh was placed in a 1 L NalgeneMR narrow-mouthed polyethylene bottle. 600 g of an oxidizing gas solution consisting of FluorinertMR FCMR-77 Electronic Fluid saturated with chlorine was added to the bottle containing the mineral , and the bottle was hermetically sealed with the lid. Initially, the perfluorocarbon saturated with chlorine had a distinctive greenish-yellow tint. The sealed container was placed on a roller mill and allowed to spin for two hours at a speed of 4 rpm, to allow gentle mixing and continuous contact of the ore with the oxidizing gas solution. During the oxidation process, the temperature of the oxidizing gas solution increased markedly and after the two-hour treatment, the solution no longer had any visible color and had a slight chlorine odor. The vessel was then removed from the mill, the spent oxidant gas solution was drained of the ore, and the ore was removed from the bottle and left to dry overnight. The dry mineral charge was then returned to the polyethylene bottle, 300 g of an aqueous leaching solution containing 25 g / L of sulfuric acid was added, the bottle was sealed and then placed again on the rotating or roller mill at 4 rpm for a period of 72 hours. During this leaching cycle, 5 ml aliquots of the aqueous leaching solution were periodically removed to analyze the level of soluble copper using atomic absorption spectroscopy. After each withdrawal, a fresh 5 ml aliquot of the aqueous sulfuric acid leaching solution was added to the bottle to keep the total volume of the solution at its original level. The concentration of the soluble copper in the aqueous sulfuric acid leaching solution as a function of the leaching time is presented in Table 2. Comparative Example C3 In Comparative Example C3, the same experiment was run as in Example 4, except that the ore oxidation step was omitted and only the leaching step was carried out with aqueous sulfuric acid. The concentration of the soluble copper in the aqueous sulfuric acid leaching solution, as a function of the leaching time, is presented in Table 2.

Table 2 The data in Table 2 show that, at all times, 1-a aqueous sulfuric acid leach solution contained more than 50% copper when the copper ore sample was oxidized with the oxidizing gas solution prior to leaching with acid.

Example 5 600 g of the copper ore fraction +4 mesh to -2 mesh of Example 4 and 600 g of an oxidant gas solution consisting of Fluorinert ™ FCMR-77 Electronic Fluid saturated with chlorine were placed in a Nalgene ™ 500 polyethylene bottle ml. The loaded bottle was placed on a roller mill at 4 rpm for two hours. Subsequently, the oxidizing gas solution was drained from the bottle and the oxidized ore was removed and allowed to dry overnight. The ore was then returned to the original bottle, and the leaching was performed with aqueous sulfuric acid as described in Example 4, as was the periodic determination for the concentration of soluble copper using atomic absorption spectroscopy, and replenishing the solution of leaching of aqueous sulfuric acid.

The concentration of the soluble copper in the aqueous sulfuric acid leaching solution as a function of the leaching time is presented in Table 3.

Comparative Example C4 In Comparative Example C4, the same Experiment 5 was run except that 300 g of 5.25% aqueous sodium hypochlorite (by weight) was used, instead of the 600 g of FCMR-77 saturated with chlorine, during the oxidation step of the mineral. The concentration of the soluble copper in the aqueous sulfuric acid leaching solution, as a function of the leaching time, is presented in Table 3.

Comparative Example C5 In Comparative Example C5, the same experiment as in Example 5 was run, except that the oxidation step of the ore was omitted, and only the leaching step was carried out with aqueous sulfuric acid.

The concentration of the soluble copper in the leaching solution of the aqueous sulfuric acid as a function of the leaching time is presented in Table 3.

Table 3 The data in Table 3 show that the oxidation gas solution clearly exceeds the performance of aqueous sodium hypochlorite solution as an oxidizing agent for copper ore, prior to mineral extraction with aqueous sulfuric acid leaching solution . In fact, the aqueous solution of sodium hypochlorite produced slightly lower soluble chlorine in the sulfuric acid leaching solution than when the oxidation treatment of the mineral was not used.

It is noted that in relation to this date, the best method known to the applicant to carry out the aforementioned invention, is that which is clear from the present description of the invention.

Having described the invention as above, property is claimed as contained in the following:

Claims (10)

1. A method to leach unconcentrated minerals from precious metals and non-precious metals from ore and ore concentrates, characterized in the method because it comprises the steps of: contacting a mineral that has metal or ore concentrate, with a solution of oxidant gaß comprising a solution of an oxidizing gas dissolved in a perfluorinated liquid; and the extraction of the precious or non-precious metal from the ore that has metal.

2. The method according to claim 1, characterized in that the contacting step and the extraction step are performed intermittently to form a stepwise leaching process comprising the alternating contact and extraction stages.

3. The method according to claim 1 or 2, further characterized in that it comprises the concentration of the mineral by agglomeration or spraying before being placed in contact with the oxidizing gas solution.

4. The method according to claim 1, 2 or 3, characterized in that the leaching is carried out by mining heap leaching, in vat or by agitation.

5. The method according to claim 1, 2, 3 or 4, characterized in that the precious metal to be recovered is gold, characterized in that the extraction comprises contacting the mineral with an aqueous solution comprising alkali cyanide, and wherein the method further comprises the step of recovering the elemental gold by adsorption on activated carbon.

6. The method according to claim 1, 2, 3 or 4, wherein the precious metal to be recovered is silver, characterized in that the extraction comprises contacting the mineral with an aqueous solution comprising alkali cyanide, and where the method further comprises the step of recovering elemental silver and elemental gold by precipitation with zinc metal.

7. The method according to claim 1, 2, 3 or 4, characterized in that the non-precious metal to be recovered is copper, and wherein the extraction comprises contacting the mineral with an aqueous solution comprising sulfuric acid.

8. A method for destroying cyanide from a solution containing cyanide, characterized in that the method comprises contacting said solution with a solution of an oxidizing gas dissolved in a perfluorinated liquid.

9. The method according to any of claims 1 to 8, characterized in that the oxidation gas is selected from the group consisting of chlorine, oxygen, ozone, chlorine dioxide, and sulfur dioxide, and wherein the perfluorinated liquid is selected of the group consisting of cyclic and non-cyclic perfluoroalkanes, cyclic and non-cyclic perfluoroamines, cyclic and non-cyclic perfluoroethers, and cyclic and non-cyclic perfluoroaminoethers.

10. The method according to any one of claims 1 to 8, characterized in that the perfluorinated liquid has a boiling point close to or higher than room temperature, and has a freezing point below the ambient temperature, and where the solution Oxidizing gas is saturated.