METHOD OF RECOVERING GOLD FROM GOLD ORES CONTAINING PYRITE TECHNICAL FIELD [0001] The present invention relates to a method of recovering gold from gold ores containing pyrite. RELATED ART [0002] For a method of recovering gold from a substance containing gold, a method employing a wet process is known. Traditionally, leaching gold from sulfide ores into a solution has been conducted by use of chemicals such as a cyan, a thiourea, thiosulfuric acid and a halogen gas. Recently, for a leaching agent of lower toxicity, as mentioned in JP-A-2009-235525 (Patent Document 1), a gold leaching liquid using chloride ions, iron ions, copper ions and bromide ions has been proposed. [0003] For pretreatment in order to facilitate the gold leaching from sulfide ores, a method of oxidative roasting the sulfide ores is known, and recently a pretreatment of combining the oxidative roasting with the other process has been proposed. For example, in JP-A-2010-235999 (Patent Document 2), the copper sulfide ores is leached at the melting point of sulfur or less, fine particles of sulfur from the resulting residue and sulfide particles which is not leached and remains are floated by use of the difference in the hydrophobicity to the others such as iron oxide and veinstone 1 content. On the other hand, gold in the leached residue is concentrated by precipitating the iron oxide and the veinstone content, or by separating as precipitates. Thereafter, the content containing the concentrated gold is subjected to sulfur removal and then oxidative roasting to changing the iron content to iron oxides (hematite), so that the resultant is dissolved with sulfuric acid to thereby recover the gold concentrated in residue. PRIOR ART PATENT LITERATURE [0004] Patent Document 1: JP-A-2009-235525 Patent Document 2: JP-A-2010-235999 SUMMARY OF THE INVENTION PROBLEM TO BE SOLVED BY THE INVENTION [0005] A method as described in JP-A-2009-235525 (Patent Document 1) certainly is extremely useful for leaching gold in sulfide ores, since gold is allowed to be leached easily without using high toxic chemicals such as a cyan, a thiourea, thiosulfuric acid and a halogen gas. However, when this is applied to leaching pyrites, the gold leaching rate is not sufficient, so that there is a room to be improved. Therefore, as mentioned in JP-A-2010 235999 (Patent Document 2), it may be considered employing the method of raising the gold leaching rate by conducting pretreatment by employing the oxidative roasting with supplying 2 oxygen. [0006] When employing the oxidative roasting of sulfide ores including a method described in Patent Document 2, chemical reactions as for 2CuS + 202 - 2CuO + S02, 2CuFeS 2 + 602 - CuO + 4S02 + Fe 2 0 3 , and 4FeS 2 + 1102 - 2Fe 2 0 3 + 8S02 t ake place preferentially, so that the problem that sulfur dioxide (SO 2 ) which is known as an environmental pollutant is generated should be hardly avoided. Therefore, it is desirable that the pretreatment for raising the gold leaching rate should lead to reducing the sulfur dioxide which is generated during the ore treatment process, to raise safety, and reduce the influence on the environment from the viewpoint of safety and environment. [0007] In a case of wet processing pyrites, gold which is an accompaniment is leached with halogen bath after being separated and concentrated beforehand in the residue, or at the time in the later stage of leaching copper as a main component in halogen bath. A gold complex with a halide ligand still remains in the post-leaching solution. When recovering gold by adsorbing the gold complexes on activated carbon, the larger the adsorption amount thereof is, the larger the yield of gold should be. In particular, in the case of incineration of activated carbon, the adsorption amount per unit weight of activated carbon have a great influence directly to production costs. Therefore, although the development of a method for increasing the unit adsorption is desired, none of the Patent Documents 1 and 2 even make a consideration on 3 increasing the adsorption amount of gold to the activated carbon. Further, an appropriate method is still unknown in general since there are too many matters to be considered on such as the type of the activated carbon and the contaminant from the post leaching solution. [0008] The present invention has been made in view of the above circumstances, a method of recovering gold from gold ores containing pyrite, in which the gold leaching rate may increase without using highly toxic chemicals such as a cyan, a thiourea, thiosulfate, a halogen gas, while inhibiting the generation of sulfur dioxide, and which the amount of adsorption of gold to activated carbon may be raised, is provided. MEANS FOR SOLVING THE PROBLEM [0009] The present inventors have conducted intensive studies in order to solve the above problem, and found that, the pretreatment for pyrolysis of pyrite to iron(II) sulfide under an inert atmosphere, and then the gold leaching is conducted with a gold leaching liquid containing chloride ions, bromide ions and trivalent iron ions, so that while inhibiting the generation of sulfur oxide, the gold leaching rate may increase dramatically. Further, the present inventors identified that a competitive adsorption to the activated carbon is monovalent copper ions in the post-gold-leaching solution, when adsorbing gold in the post gold-leaching solution by the gold leaching to the activated 4 carbon and recovering gold, and found that such monovalent copper ions are subjected to the pretreatment for reducing before a step for adsorbing gold to the activated carbon to significantly improve the adsorption amount of gold to the activated carbon. [0010] The present invention was accomplished based on the above findings. In one aspect, the present invention is directed to a method of recovering gold from gold ores containing pyrite comprising: Pretreatment comprising Step 1 for preparing gold ores containing pyrite, and Step 2 for heating the gold ores under an inert atomsphere to 450 degrees C or higher, to pyrolyze the pyrite in the gold ores to iron(II) sulfide and elemental sulfur, and not comprising oxidative roasting process; Step 3 for contacting the gold ores thus obtained by the pretreatment process with a gold leaching liquid containing chloride ions, bromide ions and iron ions with supplying an oxidizing agent to leach the gold content in the ores; Step 4 for adding copper(I) chloride to the post-gold-leaching solution thus obtained in Step 3, and then adding an oxidizing agent to adjust the oxidation-reduction potential to 520 mV or greater, to thereby reduce monovalent copper ions in the post-gold-leaching solution; and Step 5 for adsorbing gold in the post-gold-leaching solution thus obtained in Step 4 to activated carbon. [0011] In one embodiment of the method of recovering gold from gold ores containing pyrite according to the invention, the Step 4 comprises adjusting the oxidation-reduction potential (reference 5 electrode: silver/silver chloride) to 520 mV to 570 mV. [0012] In one embodiment of the method of recovering gold from gold ores containing pyrite according to the invention, the Step 4 comprises the oxidation-reduction potential is adjusted by blowing air. EFFECT OF THE INVENTION [0013] The present invention may provide a method of recovering gold from gold ores containing pyrite, in which the dramatically improved gold leaching rate may be obtained while controlling the generation of toxic sulfur oxide, and it is possible to increase the adsorption amount per unit weight of gold, by leaching gold with a specific gold leaching liquid after conducting a pretreatment method of the invention, and by processing for reducing monovalent copper ions from the gold-leaching solution, which is an inhibitor for adsorption of gold to the activated carbon. BRIEF DESCRIPTION OF THE DRAWINGS [0014] FIG. 1 is a graph illustrating a relationship between the leaching time and Au content in the residue with respect to examples and comparative examples; FIG. 2 is a TG/DTA curve obtained by heat analysis on the fine grinded pyrite concentrate used in Example 1 under nitrogen 6 atmosphere; FIG.3 is a graph illustrating a relationship between the oxidation-reduction potential of the post-gold-leaching solution and the concentration of gold of the post-adsorption solution; and FIG. 4 is a graph illustrating a relationship between changes in the oxidation-reduction potential and the concentration of gold, provided that the leaching solution is continuously fed into a column filled with an activated carbon and that adding CuCl and flowing air is conducted. MODE FOR CARRYING OUT THE INVENTION [0015] The present invention will be illustrated below. [0016] 1. Pretreatment step One embodiment of a method of pretreating gold ores of the invention contains Step 1 for preparing gold ores containing pyrite; and Step 2 for heating the gold ores under inert atmosphere to 450 degrees C or higher, to pyrolyze the pyrite in the gold ores to iron(II) sulfide and elemental sulfur, and does not contain oxidative roasting process. [0017] (1) Step 1 In Step 1, gold ores containing pyrite are prepared, since the purpose of the invention is to improve gold leaching rate in pyrite, which is poorly soluble so that gold leaching rate should be small. However, any other requirements that, for example, the 7 small or large of the concentration of gold in the ores does not make any difference. Gold ore to be processed by the present invention can also be one that has been treated by the conventional ore dressing, such as flotation and gravity concentration. The ore may be pulverized and grinded to reduce the particle size of the ore, so that the gold-leaching liquid may be facilitated to contact gold inside of the ore. The gold concentration in the gold ores is typically about 0.1 to 100 ppm by mass and more typically about 1 to 20 ppm by mass. [0018] The gold ores contains pyrite, and may contain other than that, such as copper pyrites, galena, sphalerite, arsenopyrite, stibnite and pyrrhotite. In a typical embodiment of the invention, a gold ores containing 3 mass% or more of pyrite may be used. Further, in a more typical embodiment, a gold ores containing 30 mass% or more of pyrite may be used. By using such gold ores, the effect of the pretreatment of the invention may remarkably exhibit. There should be no upper limit of the pyrite content of the gold ores, it may be 100 mass%, more preferably 80 mass% or less. [0019] In the present invention, it is also one of characteristics not to include the oxidative roasting process. In the conventional technique, oxidative roasting process has been conducted under the presence of oxygen or air, so that sulfur in the sulfide ores and oxygen may be coupled to generate sulfur oxides. In the present invention, such oxidative roasting process should not be 8 conducted. [0020] (2) Step 2 In Step 2, the gold ores are heated under inert atmosphere to 450 degrees C or higher, to pyrolyze the pyrite in the gold ores to iron(II) sulfide and elemental sulfur. The chemical reaction at this time may be expressed as FeS 2 - FeS + S. No sulfur oxides may be raised in theory. The gold ores treated by pyrolysis exhibit the remarkable improvement of the solubility to the gold-leaching liquid to be described later. As compared with the case with no pyrolysis treatments, the gold leaching rate may increase by about 10 times. In the pyrolysis method conducted in the invention, the pyrite (FeS 2 ) should not be changed to hematite (Fe 2
O
3 ), so that it is expected that the gold leaching rate should be insufficient. Therefore, it is very surprising to obtain such a result. [0021] The inert atmosphere in which the pyrolysis would be conducted includes the noble gas atmosphere such as argon or helium, and nitrogen atmosphere. - Otherwise, the exhausted gas used for pyrolysis and circulated may be used. If the oxygen is contained in the atmosphere, gold ore is roasted and oxidized to generate sulfur dioxide, so that the influence on the environment is concerned. Therefore, oxidative roasting is not adopted in the present invention. [0022] At pyrolysis, the temperature of the gold ores should need 9 to be 450 degrees C or higher because at the temperature of less than 450 degrees C, it is difficult to progress pyrolysis of the pyrite. Pyrolysis may be preferably conducted with holding the temperature of the pyrite at 550 degrees C or higher, and more preferably conducted with holding 650 degrees C or higher. In pyrolysis, the holding temperature is preferably continued for 5 minutes or longer, more preferably 30 minutes or longer. This is because the pyrolysis reaction should be significantly progressed. However, since there is a risk the large energy and processing time should be required for elevating temperature when the temperature of the gold ores is excessively high, the holding temperature is preferably 800 degrees C or lower, and more preferably 750 degrees C or lower. As similar, the time for holding the temperature may be preferably 120 minutes or shorter, more preferably 60 minutes or shorter. [0023] No particular limitation on the type of furnace for carrying out pyrolysis, but can be used, for example a tubular furnace, a rotary kiln. [0024] Elemental sulfur which is generated by pyrolysis may be subject to a solid-gas separation from the gold ore since such sulfur is gasified in high-temperature furnace. Then, it is possible to send it to the exhaust system together with the atmosphere gas. However, when sending the elemental sulfur in the exhaust system, the sulfur would be precipitated with a decrease in temperature to cause a problem of clogging of the gas canal, so 10 that it is desirable to recover by wet scrubber and the like. Alternatively, the gasified elemental sulfur is cooled together with iron(II) sulfide, which is generated in Step 2, to collect them together in a solid form, so that it is possible to send them to the gold leaching step together. The elemental sulfur may be separated as a leaching residue in the gold leaching step without inhibiting the gold leaching. In this case, it is economically advantage since wet scrubber should be needed. [0025] 2. Gold leaching step (1) Step 3 In an embodiment according to the present invention, Step 3 is conducted, in which the resulting gold ores after the pretreatment process is contacted with a gold-leaching liquid containing chloride ions, bromide ions and iron ions with the supply for an oxidizing agent, to leach gold in the gold ores. [0026] Gold leaching would make progress by reacting the leached gold with chloride ions or bromide ions to generate a chloride complex of gold or a bromide complex of gold. Using bromide ions together, the complex may be generated at lower potential, so that the gold leaching efficiency may be improved. Further, with respect to iron ions, trivalent iron ions formed by being oxidized with the supply for an oxidizing agent or ions being originally trivalent would serve to oxidize gold. The post-gold-leaching solution contains preferably copper ions. Although copper ions do not involve the reaction directly, the rate of oxidation by iron 11 ions may be faster in the presence of copper ions. [0027] A manner of contacting an leaching liquid with gold ores includes, without special limitation, spraying and dipping. From the viewpoint of the reaction efficiency, it is preferable to dip the residue in the leaching liquid and to stir it. [0028] A source of supply for chloride ions includes, without special limitation, hydrogen chloride, hydrochloric acid, metal chloride and chlorine gas. From the viewpoint of economy and safety, chloride ions is preferably supplied in a form of metal chloride. Such metal chloride includes, for instance, copper chloride (copper(I) chloride, copper(II) chloride), and iron chloride (iron(I) chloride, iron(II) chloride), and a chloride of alkaline metal such as lithium, sodium, potassium, rubidium, cesium, francium, and a chloride of alkaline earth metal such as beryllium, magnesium, calcium, strontium, barium, radium. From the viewpoint of economy and availability, sodium chloride is preferable. Copper chloride and iron chloride are also preferable since those chlorides may be used as a source of supply for copper ions and iron ions. [0029] A source of supply for bromide ions includes, without special limitation, hydrogen bromide, hydrobromic acid, metal bromide and bromine. From the viewpoint of economy and safety, bromide ions is preferably supplied in a form of metal bromide. Such metal bromide includes, for instance, copper bromide 12 (copper(I) bromide, copper(II) bromide), and iron bromide (iron(I) bromide, iron(II) bromide), and a bromide of alkaline metal such as lithium, sodium, potassium, rubidium, cesium, francium, and a bromide of alkaline earth metal such as beryllium, magnesium, calcium, strontium, barium, radium. From the viewpoint of economy and availability, sodium bromide is preferable. Copper bromide and iron bromide are also preferable since those bromides may be used as a source of supply for copper ions and iron ions. [0030] Copper ions and iron ions are usually fed in a form of salt, and for example may be supplied in a form of a halogenated salt. Copper ions are preferably supplied in a form of copper chloride and/or copper bromide, and iron ions are preferably supplied in a form of iron chloride and/or iron bromide, from the viewpoint which they also could be used as a source of supply for chloride ions and/or bromide ions. Copper chloride and iron chloride are copper(II) chloride (CuCl 2 ) and copper(I) chloride (CuCl), and iron(II) chloride (FeCl 3 ) and iron(I) chloride (FeCl 2 ) [0031] The chloride ion concentration in the gold-leaching liquid as used in Step 3 is more preferably 40 g/L to 200 g/L. The bromide ion concentration in the gold-leaching liquid as used in Step 3 is preferably 20 g/L to 100 g/L from the viewpoint of reaction rate and solubility. Iron ion concentration in the gold leaching liquid is preferably 0.01 g/L to 10 g/L. From the viewpoint of the gold leaching efficiency, the ratio of mass concentration of bromide ions to chloride ions in the gold 13 leaching liquid is preferably 1 or greater. [0032] The oxidation-reduction potential (vs Ag/AgCl) of the leaching liquid at starting Step 3 is preferably 550 mV or greater, more preferably 600 mV or greater from the viewpoint of promoting gold leaching. Further, from the viewpoint of raising the gold leaching rate, it is preferable to maintain the pH of the gold-leaching liquid at 2.0 or less. However, the oxidation rate of iron is promoted at higher pH, so that it is more preferable to maintain the pH of the gold-leaching liquid at 0.5 to 1.9. Further, from the viewpoint of raising the gold leaching rate, it is preferable to maintain the temperature of the gold-leaching liquid at 45 degrees C or higher, more preferably 60 degrees C or higher. Further, when the temperature is too high, the laeching liquid is evaporated and the heating cost increases, so that it is prefreable to maintain the temperature at 95 degrees C or less, more preferable at 85 degrees C or less. [0033] In a suitable embodiment of the invention, a mixture of at least one of hydrochloric acid and hydrobromic acid, and at least one of copper(II) chloride and copper(II) bromide, and at least one of iron(II) chloride and iron(II) bromide, and at least one of sodium chloride and sodium bromide may be used as a gold-leaching liquid of Step 3, as far as both of chloride ions and bromide ions are included therein. [0034] The gold leaching step of Step 3 is conducted with supplying 14 an oxidizing agent to manage the oxidation-reduction potential. If the oxidizing agent is not added, the oxidation-reduction potential gets lower in the middle of the gold leaching step, so that the leaching reaction is not progressed. Such oxidizing agent includes, without special limitation, oxygen, air, chlorine, bromine and hydrogen peroxide. An oxidizing agent having the extremely high oxidation-reduction potential is not needed, and air should be sufficient. [0035] (2) Step 4 The oxidation-reduction potential of the post-gold-leaching solution obtained after the gold leaching is fully conducted is appropriately 500 to 520 mV. CuCl is further added to the post gold-leaching solution and stir it to reduce the oxidation reduction potential to 520 mV or less, preferably 500 mV or less, and then an oxidizing agent is added to re-adjust the ORP over 520 mV. As a result, monovalent copper ions in the post-gold-leaching solution, which inhibits the adsorption of gold to the activated carbon, is oxidized to divalent copper ions to reduce the amount of monovalent copper, so that the amount of competitive adsorption to the activated carbon in the post-gold-leaching solution. Accordingly, the adsorption rate of gold on the activated carbon is improved. [0036] For the oxidizing agent, without special limitation, air is used from the viewpoint of cost. Further, for the solution temperature, without special limitation, from the viewpoint that 15 the gold leaching is a heat-leaching, and the viewpoint of the oxidation efficiency, it is preferable to keep the solution temperature of the post-gold-leaching solution at 45 degrees C or higher, more preferable 50 degrees C or higher. [0037] The increase of the ORP indicates reducing the amount of monovalent copper in the post-gold-leaching solution. Monovalent copper is known as a quite soft element, so that it has high affinity for the activated carbon, and then compete against a gold complex in the adsorption to the activated carbon. By reducing the amount of the monovalent copper, the more activated sites for the adsorption in the activated carbon will open to gold, so that the selectivity by gold increases, and the effective recovery of gold is achieved. [0038] By adjusting the ORP to 520 mV or greater, the concetration of the monovalent copper is reduced, so that the adsorption rate of gold to the actiated carbon may be improved. For the upper limit of the ORP, without special limitation, from the viewpoint of a period requried to adjust the ORP and reduction efficiency of the amount of monovalent copper, the ORP is adjusted preferably to 570 mV or less, more preferably to 530 to 560 mV. [0039] 3. Gold recovery (Step 5) Step 5 is conducted for recovering gold by adsorption to the activated carbon from the gold solution obtained by the solid liquid separation after the gold leaching reaction. The contact of 16 gold to the activated carbon may be conducted by batchwise manner, or continuous flow the acidic leaching solution through the adsorption column filled with the activated carbon. [0040] In case of batchwise manner, the stirring rate is not designed. The activated carbon is filled to the amount of 50 times to 10000 times of the mass amount of gold. [0041] In case of continuous flow, flow rate is not limited (in general, SV 1 to 25). On the other hand, when the amount of gold adsorption at a unit mass of the activated carbon reaches 20000 to 30000 g/t, such activated carbon does not meet the required capability. Therefore, strip of gold from the activated carbon or recovering is conducted based on the above mentioned amount of gold adsorption. Regeneration of the activated carbon is conducted by, without special limitation, well known method with sulfur compound, nitrogen compound or acid. [0042] 4. Others After conducting Step 2 and before conducting Step 3, it may be possible to conduct a variety of treatment for removing impurities in the gold ores. For example, elemental sulfur may be removed by heating the pretreated gold ores to the temperature at which the elemental sulfur is fully melting, and by filtering and separating gold and the elemental sulfur. Iron sulfide (FeS) may be removed by leaching iron content by contacting the pretreated gold ores with a variety of mineral acid such as sulfuric acid and 17 hydrochloric acid, or a aqueous solution of Fe 3 - salt such as sulfate iron and iron chloride or the like, and then by subjecting the resultant to the solid-liquid separation. EXAMPLES [0043] Hereinafter, more detailed description of the present invention should be illustrated through examples. However, the present invention is not limited thereto. Incidentally, the analysis method of the metal used in the examples was conducted by ICP-AES. Further, in the analysis of gold, after precipitating gold in the sample by cupellation, quantitative analysis was conducted by ICP-AES. [0044] (Comparative Example 1) A pyrite concentrate (Lihir ore produced in Papua New Guinea) was prepared. By conducting XRD and chemical analyses of the pyrite content in the pyrite concentrate, the content was 17 mass%. The pyrite concentrate (Lihir ore) was pulverized and grinded by ball mill, so that the particle size (d80) in accumulated weight 80% in the distribution curve of the accumulated weight particle size was adjusted to 24 pm. d80 was obtained as the average value by measuring three times by laser diffraction-type particle size distribution measuring equipment (manufactured by Shimadzu Corporation: SALD2100) . Then, a hydrochloric acidic gold-leaching liquid having the composition shown in Table 1 was used for the ground pyrite concentrate 18 (200g), to a pulp concentration of 100 g/L, and then a leaching treatment was conducted for 90 hours at a liquid temperature of 85 degrees C. During the leaching, keeping blowing air (0.1 L/min based on 1 L of the concentrate) and stirring, the oxidation reduction potential (ORP: vs Ag/AgCl) was maintained at 500 mV. Further, during leaching, hydrochloric acid was added appropriately so that the pH of the gold-leaching liquid was maintained at 1.0 to 1.1. [0045] [Table 1] Gold-leaching liquid FeCl3-6H 2 O(g/L) 10 CuCl2-2H 2 O(g/L) 48 NaCl(g/L) 25 NaBr(g/L) 103 Total chloride ions (g/L 40 Total bromide ions (g/L) 80 ORP(mV) (vs Ag/AgCl) 717 pH 1.52 [0046] During leaching test, samples of the leach residue were taken periodically to measure Au content in the residue. FIG. 1 shows a relationship obtained from the result of the test between leaching time and Au content in the residue (in FIG. 1, refer to a plot indicated as "FeS2 without pyrolysis"). From the result, it is recognized that it took 90 hours that Au content in the residue which was 6 g/t initially was lowered to 0.9 g/t. [0047] <Example 1> 19 A ground pyrite concentrate which was the same as Comparative Example 1 (1.5 kg) was introduced into a tubular furnace, and the temperature was elevated to 700 degrees C over one hour (elevating rate: 10 degrees C/min) under nitrogen atmosphere, to further heat for one hour. Allowing to cool to room temperature, it was confirmed by comparison of XRD analysis before and after heat treatment that the peak of FeS 2 contained in the original ores disappeared, and that the peak of FeS arose. The elemental sulfur generated by heat treatment was not subjected to removing procedure in particular. Then, a hydrochloric acidic gold-leaching liquid having the same composition as Comparative Example 1 was used for the heat treated pyrite concentrate, to a pulp concentration of 100 g/L, and then a leaching treatment was conducted for 18 hours at a liquid temperature of 85 degrees C. During leaching, keeping blowing air (0.1 L/min based on 1 L of the concentrate) and stirring, the oxidation-reduction potential (ORP: vs Ag/AgCl) was maintained at 400 mV or greater. Further, during leaching, hydrochloric acid was added appropriately so that the pH of the gold-leaching liquid was maintained at 1.0 to 1.1. [0048] During leaching test, samples of the leach residue were taken periodically to measure Au content in the residue. FIG. 1 shows a relationship obtained from the result of the test between leaching time and Au content in the residue (in FIG. 1, refer to a plot indicated as "FeS2 with pyrolysis") . From the result, it is recognized that Au content in the residue which was 6 g/t 20 initially was lowered to 0.6 g/t for only 12 hours. [0049] <Change of peaks of FeS 2 and FeS in XRD affected by the condition of pyrolysis> A ground pyrite concentrate which was used in Example 1 (1.5 kg) was subjected to research the change in diffraction intensity of FeS 2 and FeS under the XRD analysis with varying the holding temperature and holding time as shown in Table 1. The experiments were conducted by use of a tubular furnace under nitrogen atmosphere. The elemental sulfur generated by pyrolysis was evaporated to be removed by a nitrogen stream. The elevating rate was all set to 10 degrees C/min. Cooling was conducted to allow to room temperature. XRD analysis was conducted by use of RINT2200 ultimate manufactured by Rigaku Corporation. FeS 2 has characteristic peaks at 20 of 32.98 degrees and 56.15 degrees, and FeS has characteristic peaks at 20 of 43.67 degrees and 33.78 degrees, so that those incidence angles were focused. The result was shown in Table 2. [0050] [Table 2] Heating Condition FeS 2 Intensity FeS Intencity (CPS) (CPS) Holding Holding 32.980 56.150 43.670 33.780 temperature time(min) (degrees C) Before heat treatment 250 170 ND ND 550 60 270 250 ND ND 550 120 60 60 ND ND 600 5 ND ND 100 120 600 30 ND ND 150 100 21 600 60 ND ND 120 130 650 60 ND ND 180 130 700 60 ND ND 350 190 ND: indicating lower limit of quantitation or below [0051] Based on the results in Table 2, heating at 600 degrees C or higher, it is recognized that the peaks from the pyrite should surely disappear. This fact indicates that the crystalline pyrite was pyrolyzed. Furthermore, the condition of the holding temperature and holding time is most preferably 650 degrees C or higher and 60 minutes or longer, respectively, as confirmed that a clear peak of FeS appears. [0052] <Example 2> A ground pyrite concentrate which was used in Example 1 was subjected to research the change in weight at each of temperature and the endothermic-exothermic analysis by use of the thermal analysis (TG/DTA6300 manufactured by Seiko Instruments Inc.) under nitrogen atmosphere. The elevating rate was set to 20 degrees C/min. The results are shown in FIG 2. It is recognized that the mass starts decreasing at 450 degrees C, and that exothermic phenomenon was observed at the same time, so that the pyrite starts degrading. Degradation of the pyrite does not take place until elevating to at least 450 degrees C, under nitrogen atmosphere. However, from the viewpoint of the results of the above mentioned XRD analysis, it is considered that it requires a long time for pyrolysis at about 450 degrees C, so that it is preferable to conduct heat treatment at 600 degrees C or higher. 22 [0053] <Example 3> Gold in a post-gold-leaching solution obtained after the gold leaching step was leached by use of a gold-leaching liquid containing 50 g/L of chloride ions, 80 g/L of bromide ions, 18 g/L of copper, and 0.2 g/L of iron. The post-gold-leaching solution contained NaCl: 84 g/L, NaBr: 103 g/L, Cu: 20 g/L, Fe: 2 g/L, and Au: 8 mg/L, and had the pH of 1.2. The ORP was adjusted to 510 mV by adding CuCl. Then, the post-leaching solution was heated to 55 degrees C, and stirred while blowing 0.4 L per minute of air. The resulting post-gold-leaching solution was passed through a glass column filled with about 14 ml of the activated carbon derived from coconut shell (Yashicoal MC, manufactured by Taihei Chemical Industrial Co., Ltd.), to adsorb gold to the activated carbon. The column had the size of 11 mm of diameter and 150 mm of height. The feeding rate of the liquid was 11.9 ml/minute, and the space velocity thereof was 50 (1/h) . The eluted gold in the post adsorption solution was diluted by hydrochloric acid to be determined the quantity thereof by ICP-AES. FIG. 3 shows the relationship between the ORP and the concentration of gold in the post-adsorption solution. [0054] It is recognized that the gold concentration contained in the post-adsorption solution remarkably decreased in case of adjusting the ORP of 520 mV of greater. It is also recognized that although the upper limit of the ORP should not be provided, the gold concentration of the post-adsorption solution should not 23 dramatically decrease when extremely raising the voltage, and that it is enough to oxidize the solution to at least 520 mV, and however it should not be prohibited to extremely oxidize it. [0055] <Example 4> While continuously supplying liquid used in Example 3 by used of the column filled with the activated carbon, the gold concentration of the post-adsorption solution was determined by varying the ORP by the addition of CuCl and blowing air. The result is shown in FIG. 4. [0056] It is also clear based on FIG. 4 that there is a relationship between the ORP and the adsorption of gold to the activated carbon. It is possible to favorably recover gold by contacting the post-gold-leaching solution with the activated carbon at the ORP of 520 mV or greater. Further, it is understood that what has an influence on the ORP should be Cu(I). [0057] Although Cu(I) is easily oxidized in an aqueous solution to become Cu(II), in an aqueous solution containing halide at high concentration as in a system of the invention, it may exist in a considerably stable state. Therefore, although it is estimated to obtain the same effect by oxidizing Cu(I) with an oxidizing agent such as hydrogen peroxide and hypochlorous acid other than by blowing air, blowing air may be preferable from the viewpoint of the cost and the convenience of handling. 24