JP5888780B2 - Copper concentrate processing method - Google Patents
Copper concentrate processing method Download PDFInfo
- Publication number
- JP5888780B2 JP5888780B2 JP2012081308A JP2012081308A JP5888780B2 JP 5888780 B2 JP5888780 B2 JP 5888780B2 JP 2012081308 A JP2012081308 A JP 2012081308A JP 2012081308 A JP2012081308 A JP 2012081308A JP 5888780 B2 JP5888780 B2 JP 5888780B2
- Authority
- JP
- Japan
- Prior art keywords
- copper
- concentrate
- sulfide
- particles
- flotation
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Active
Links
- 239000010949 copper Substances 0.000 title claims description 199
- 229910052802 copper Inorganic materials 0.000 title claims description 144
- 239000012141 concentrate Substances 0.000 title claims description 141
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims description 139
- 238000003672 processing method Methods 0.000 title claims description 13
- 239000002245 particle Substances 0.000 claims description 70
- 238000000034 method Methods 0.000 claims description 61
- 238000005188 flotation Methods 0.000 claims description 60
- 230000008569 process Effects 0.000 claims description 45
- 238000006243 chemical reaction Methods 0.000 claims description 44
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 claims description 39
- COHYTHOBJLSHDF-UHFFFAOYSA-N indigo powder Natural products N1C2=CC=CC=C2C(=O)C1=C1C(=O)C2=CC=CC=C2N1 COHYTHOBJLSHDF-UHFFFAOYSA-N 0.000 claims description 33
- 235000000177 Indigofera tinctoria Nutrition 0.000 claims description 31
- 229940097275 indigo Drugs 0.000 claims description 31
- 229910052683 pyrite Inorganic materials 0.000 claims description 30
- 239000011028 pyrite Substances 0.000 claims description 30
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 claims description 30
- 238000005486 sulfidation Methods 0.000 claims description 24
- DVRDHUBQLOKMHZ-UHFFFAOYSA-N chalcopyrite Chemical compound [S-2].[S-2].[Fe+2].[Cu+2] DVRDHUBQLOKMHZ-UHFFFAOYSA-N 0.000 claims description 22
- 229910052951 chalcopyrite Inorganic materials 0.000 claims description 22
- 238000000227 grinding Methods 0.000 claims description 17
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims description 14
- 229910052717 sulfur Inorganic materials 0.000 claims description 12
- 239000011593 sulfur Substances 0.000 claims description 12
- 238000003801 milling Methods 0.000 claims description 7
- 239000012298 atmosphere Substances 0.000 claims description 5
- 229910001779 copper mineral Inorganic materials 0.000 claims description 2
- 239000011261 inert gas Substances 0.000 claims description 2
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 35
- 238000011084 recovery Methods 0.000 description 19
- 238000010438 heat treatment Methods 0.000 description 13
- 229910052742 iron Inorganic materials 0.000 description 12
- 238000000926 separation method Methods 0.000 description 12
- 229910052500 inorganic mineral Inorganic materials 0.000 description 11
- 239000011707 mineral Substances 0.000 description 11
- 238000003723 Smelting Methods 0.000 description 9
- 238000002441 X-ray diffraction Methods 0.000 description 8
- 239000002893 slag Substances 0.000 description 7
- TUZCOAQWCRRVIP-UHFFFAOYSA-N butoxymethanedithioic acid Chemical compound CCCCOC(S)=S TUZCOAQWCRRVIP-UHFFFAOYSA-N 0.000 description 6
- 230000000694 effects Effects 0.000 description 6
- 238000012545 processing Methods 0.000 description 6
- WVYWICLMDOOCFB-UHFFFAOYSA-N 4-methyl-2-pentanol Chemical compound CC(C)CC(C)O WVYWICLMDOOCFB-UHFFFAOYSA-N 0.000 description 5
- 230000007423 decrease Effects 0.000 description 5
- BUGICWZUDIWQRQ-UHFFFAOYSA-N copper iron sulfane Chemical compound S.[Fe].[Cu] BUGICWZUDIWQRQ-UHFFFAOYSA-N 0.000 description 4
- 239000000203 mixture Substances 0.000 description 4
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 3
- NFMAZVUSKIJEIH-UHFFFAOYSA-N bis(sulfanylidene)iron Chemical compound S=[Fe]=S NFMAZVUSKIJEIH-UHFFFAOYSA-N 0.000 description 3
- 238000004453 electron probe microanalysis Methods 0.000 description 3
- 239000004088 foaming agent Substances 0.000 description 3
- 238000011835 investigation Methods 0.000 description 3
- 239000003002 pH adjusting agent Substances 0.000 description 3
- 238000002203 pretreatment Methods 0.000 description 3
- 230000009467 reduction Effects 0.000 description 3
- 239000007858 starting material Substances 0.000 description 3
- 101000993059 Homo sapiens Hereditary hemochromatosis protein Proteins 0.000 description 2
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 2
- 230000002411 adverse Effects 0.000 description 2
- 230000015572 biosynthetic process Effects 0.000 description 2
- 229910052947 chalcocite Inorganic materials 0.000 description 2
- 239000003795 chemical substances by application Substances 0.000 description 2
- 238000010586 diagram Methods 0.000 description 2
- ZOOODBUHSVUZEM-UHFFFAOYSA-N ethoxymethanedithioic acid Chemical compound CCOC(S)=S ZOOODBUHSVUZEM-UHFFFAOYSA-N 0.000 description 2
- 230000006872 improvement Effects 0.000 description 2
- 239000012535 impurity Substances 0.000 description 2
- QWENMOXLTHDKDL-UHFFFAOYSA-N pentoxymethanedithioic acid Chemical compound CCCCCOC(S)=S QWENMOXLTHDKDL-UHFFFAOYSA-N 0.000 description 2
- 239000002002 slurry Substances 0.000 description 2
- 241000894007 species Species 0.000 description 2
- 238000012360 testing method Methods 0.000 description 2
- IJGRMHOSHXDMSA-UHFFFAOYSA-N Atomic nitrogen Chemical compound N#N IJGRMHOSHXDMSA-UHFFFAOYSA-N 0.000 description 1
- 229910004298 SiO 2 Inorganic materials 0.000 description 1
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 1
- 230000003750 conditioning effect Effects 0.000 description 1
- 238000011109 contamination Methods 0.000 description 1
- 150000001879 copper Chemical class 0.000 description 1
- BWFPGXWASODCHM-UHFFFAOYSA-N copper monosulfide Chemical group [Cu]=S BWFPGXWASODCHM-UHFFFAOYSA-N 0.000 description 1
- 229910001873 dinitrogen Inorganic materials 0.000 description 1
- 238000001035 drying Methods 0.000 description 1
- 230000002349 favourable effect Effects 0.000 description 1
- 238000007561 laser diffraction method Methods 0.000 description 1
- 239000012633 leachable Substances 0.000 description 1
- 238000002386 leaching Methods 0.000 description 1
- 238000005259 measurement Methods 0.000 description 1
- 238000010907 mechanical stirring Methods 0.000 description 1
- 238000002156 mixing Methods 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 239000012299 nitrogen atmosphere Substances 0.000 description 1
- 239000010665 pine oil Substances 0.000 description 1
- 238000003825 pressing Methods 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- 230000009257 reactivity Effects 0.000 description 1
- 239000012047 saturated solution Substances 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
Images
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Description
本発明は銅精鉱の処理方法に関する。 The present invention relates to a method for treating copper concentrate.
銅鉱山で産出される銅鉱石は、主に硫化鉱である。硫化鉱を大別すると、輝銅鉱(Ｃｕ２Ｓ)、銅藍(ＣｕＳ)などの鉱物を主体とした比較的高銅品位の二次硫化銅鉱と、黄銅鉱（ＣｕＦｅＳ２）を主体とする初生硫化鉱とに分けられる。近年、銅鉱山で採取される銅鉱石は、後者主体となっている。その結果、鉄、硫黄などの不純物が増加し、銅品位は低下傾向にある。このことは、鉱山で銅製錬向けに生産する銅精鉱の銅品位の低下、鉄分の増加などの要因となる。 Copper ore produced in copper mines is mainly sulfide ore. Roughly categorizing sulfide ores, relatively high copper grade secondary copper sulfides mainly composed of minerals such as chalcocite (Cu 2 S) and copper indigo (CuS), and the first generation mainly composed of chalcopyrite (CuFeS 2 ) Divided into sulfide ores. In recent years, copper ores collected at copper mines are mainly the latter. As a result, impurities such as iron and sulfur increase, and the copper quality tends to decrease. This causes factors such as a decrease in copper quality and an increase in iron content of copper concentrates produced for copper smelting in the mine.
銅鉱山における選鉱処理では、原料鉱石(粗鉱)のＣｕ品位が低下すると、製品銅精鉱のＣｕ品位も低下する。一般的に、製品銅精鉱中のＣｕ品位を高めようとするとＣｕ回収率が低下するため、とりわけ高銅価の昨今では大きな収益ロスの要因となる。銅精鉱のＣｕ品位およびＣｕ回収率を維持するためには、粉砕、摩鉱、浮遊選鉱などの一連の工程の段数増加などの何らかの追加手段が必要となり、コスト増加は避けられない。 In the beneficiation process in the copper mine, when the Cu quality of the raw ore (coarse ore) decreases, the Cu quality of the product copper concentrate also decreases. Generally, when the Cu quality in the product copper concentrate is increased, the Cu recovery rate is lowered, and this causes a large profit loss especially in recent years with a high copper value. In order to maintain the Cu quality and the Cu recovery rate of the copper concentrate, some additional means such as an increase in the number of stages in a series of processes such as grinding, grinding, and flotation is necessary, and an increase in cost is inevitable.
銅精鉱を処理する乾式銅製錬所においては、一般に、銅は製品電気銅として、鉄分はスラグとして、硫黄分は硫酸として回収される。銅精鉱の低品位化は、銅製錬プロセスにおいて処理コストの上昇を招き、事業収益を慢性的に圧迫しているスラグの需給をさらに悪化させる要因である。すなわち、銅精鉱の銅品位低下および鉄分の増加は、銅製錬業の重大な懸念点の一つである。この問題緩和のために、銅精鉱中の鉄含有量を低減するための効率の良い手段が望まれる。 In a dry copper smelter that processes copper concentrate, copper is generally recovered as product electrolytic copper, iron as slag, and sulfur as sulfuric acid. The lower grade of copper concentrate causes an increase in processing costs in the copper smelting process, and is a factor that further deteriorates the supply and demand of slag, which is chronically pressing business profits. That is, the copper grade reduction of copper concentrate and the increase in iron content are one of the major concerns of the copper smelting industry. In order to alleviate this problem, an efficient means for reducing the iron content in the copper concentrate is desired.
通常の選鉱処理とは別に、この問題を解決するための一手段として、銅精鉱の予備処理法の応用がある。予備処理法とは、黄銅鉱(ＣｕＦｅＳ２)を主体とする銅精鉱粒子を硫黄（Ｓ）とともに所定の温度で反応させ、銅藍(ＣｕＳ)と黄鉄鉱(ＦｅＳ２)とで構成される銅精鉱粒子に硫化変換する処理のことである。本変換反応は、一般的に難浸出性の黄銅鉱を、比較的浸出が容易な形態にするという意味で、湿式製錬の前処理法として知られているが、予備処理から湿式製錬までのトータルコストに問題があり、現状普及していないプロセスである。上記問題を解決する他の手段として、予備処理(硫化変換反応)後の銅藍と黄鉄鉱とを選別分離し、銅藍主体の高銅品位精鉱として乾式製錬に供する方法がある。 Apart from the usual beneficiation treatment, one method for solving this problem is the application of a copper concentrate pretreatment method. The pretreatment method is a method in which copper concentrate particles mainly composed of chalcopyrite (CuFeS 2 ) are reacted with sulfur (S) at a predetermined temperature, and copper composed of copper indigo (CuS) and pyrite (FeS 2 ). It is a process of sulfidizing into concentrate particles. This conversion reaction is generally known as a pretreatment method for wet smelting in the sense that it forms a hardly leachable chalcopyrite in a relatively easy leaching form. There is a problem with the total cost, and this is a process that is not widely used. As another means for solving the above-mentioned problem, there is a method of selectively separating copper indigo and pyrite after preliminary treatment (sulfidation conversion reaction) and subjecting them to dry smelting as a high copper grade concentrate mainly composed of copper indigo.
この硫化変換反応とその応用について述べているものに、特許文献１が挙げられる。特許文献１によると、硫化変換プロセスは、硫化変換後の銅藍と黄鉄鉱とから銅藍を選別回収し、乾式製錬または湿式製錬処理に供するために実施される。特許文献１では、銅藍と黄鉄鉱との選別において、静電的方法、重力的方法、磁気的方法、風力的方法、粒径的方法、ハイドロサイクロン法、浮遊選鉱あるいはこれらの組み合わせにより行うことが開示されている。 Patent document 1 is mentioned as what describes this sulfidation conversion reaction and its application. According to Patent Document 1, the sulfidation conversion process is carried out in order to selectively collect copper indigo from the copper indigo and pyrite after sulfidation conversion, and to provide them for dry smelting or hydrometallurgical treatment. In Patent Document 1, selection of copper indigo and pyrite can be performed by an electrostatic method, a gravitational method, a magnetic method, a wind method, a particle size method, a hydrocyclone method, a flotation process, or a combination thereof. It is disclosed.
しかしながら、特許文献１では、銅藍と黄鉄鉱とを選別する具体的な方法については記述されていない。 However, Patent Document 1 does not describe a specific method for selecting copper indigo and pyrite.
本発明は上記の課題に鑑み、Ｃｕ品位の高い銅精鉱を効率良くかつ経済的に回収することができる銅精鉱の処理方法を提供することを目的とする。 An object of this invention is to provide the processing method of the copper concentrate which can collect | recover copper concentrate with high Cu quality efficiently and economically in view of said subject.
本発明に係る銅精鉱の処理方法は、少なくとも黄銅鉱および黄鉄鉱を含み、前記黄銅鉱に対して重量比で０以上０．５未満の高銅品位鉱を含む銅精鉱粒子を不活性ガス雰囲気において３３０℃〜４１０℃で硫黄と反応させることによって硫化精鉱粒子を得る硫化変換工程と、前記硫化精鉱粒子を摩鉱する摩鉱工程と、前記摩鉱工程で得られる摩鉱精鉱粒子に対して浮遊選鉱処理する浮遊選鉱工程と、を含むことを特徴とする。本発明に係る銅精鉱の処理方法によれば、Ｃｕ品位の高い銅精鉱を効率良くかつ経済的に回収することができる。 The copper concentrate treatment method according to the present invention includes at least chalcopyrite and pyrite, and copper concentrate particles containing high-grade copper ore having a weight ratio of 0 to less than 0.5 with respect to the chalcopyrite. A sulfide conversion step for obtaining sulfide concentrate particles by reacting with sulfur at 330 ° C. to 410 ° C. in an atmosphere; a grinding step for grinding the sulfide concentrate particles; and a mill concentrate obtained by the grinding step And a flotation process in which flotation processing is performed on the particles. According to the copper concentrate processing method of the present invention, it is possible to efficiently and economically recover copper concentrate with high Cu quality.
前記硫化変換工程における温度範囲を、前記硫化精鉱粒子中の銅鉱物に、黄鉄鉱と銅藍とが共存する温度範囲、または、黄鉄鉱と銅藍とヌクンダマイトとが共存する温度範囲としてもよい。前記硫化変換工程における温度範囲を、３５０℃〜４１０℃としてもよい。前記摩鉱工程において、５０％通過粒子径（Ｐ５０）が５μｍ〜１７μｍとなるように、前記硫化精鉱粒子を摩鉱してもよい。前記摩鉱工程において、５０％通過粒子径（Ｐ５０）が５μｍ〜１０μｍとなるように、前記硫化精鉱粒子を摩鉱してもよい。前記摩鉱工程において、ボールミル、ジェットミル、アトリッションミル、またはチューブミルにより摩鉱してもよい。前記硫化変換工程における温度範囲を、３３０℃〜３８５℃としてもよい。前記硫化変換工程における温度範囲を、３３０℃〜３５０℃としてもよい。 The temperature range in the sulfide conversion step may be a temperature range in which pyrite and copper indigo coexist in the copper mineral in the sulfide concentrate particle, or a temperature range in which pyrite, copper indigo and nukundamite coexist. It is good also considering the temperature range in the said sulfidation conversion process as 350 to 410 degreeC. In the grinding step, the sulfide concentrate particles may be ground so that the 50% passing particle diameter (P 50 ) is 5 μm to 17 μm. In the milling step, the sulfide concentrate particles may be milled so that the 50% passing particle diameter (P 50 ) is 5 μm to 10 μm. In the grinding process, grinding may be performed by a ball mill, a jet mill, an attrition mill, or a tube mill. It is good also considering the temperature range in the said sulfidation conversion process as 330 to 385 degreeC. It is good also considering the temperature range in the said sulfidation conversion process as 330 to 350 degreeC.
本発明によれば、Ｃｕ品位の高い銅精鉱を効率良くかつ経済的に回収することができる。 According to the present invention, copper concentrate with high Cu quality can be recovered efficiently and economically.
以下、本発明を実施するための実施形態について説明する。 Hereinafter, an embodiment for carrying out the present invention will be described.
（実施形態）
本実施形態は、少なくとも黄銅鉱（ＣｕＦｅＳ２）および黄鉄鉱（ＦｅＳ２）を含み、黄銅鉱に対して重量比で０以上１未満の高銅品位鉱を含む銅精鉱粒子を不活性ガス雰囲気において３３０℃〜４５０℃で硫黄と反応させることによって硫化精鉱粒子を得て、硫化精鉱粒子を摩鉱し、摩鉱で得られる摩鉱精鉱粒子に対して浮遊選鉱処理する方法を開示する。この銅精鉱の処理方法によれば、黄鉄鉱主体のＣｕ品位の低い尾鉱を分離し、高いＣｕ回収率でＣｕ品位を高めた銅精鉱を効率良くかつ経済的に回収することで、銅鉱山において、安価に銅精鉱のＣｕ品位を高めることができる。また、銅精鉱に含まれる鉄量を低減し、銅製錬プロセスのコスト上昇防止とスラグ発生削減による銅製錬事業採算の改善を可能とする。
(Embodiment)
In this embodiment, copper concentrate particles containing at least chalcopyrite (CuFeS 2 ) and pyrite (FeS 2 ) and containing high copper grade ore having a weight ratio of 0 to less than 1 with respect to chalcopyrite in an inert gas atmosphere Disclosed is a method of obtaining sulfide concentrate particles by reacting with sulfur at 330 ° C. to 450 ° C., grinding the sulfide concentrate particles, and subjecting the concentrate concentrate particles obtained in the mill to a flotation process. . According to this copper concentrate treatment method, the tailings of low pyrite mainly composed of pyrite are separated, and the copper concentrate with high Cu recovery rate and high Cu grade is recovered efficiently and economically, In the mountains, the Cu quality of the copper concentrate can be increased at a low cost. In addition, the amount of iron contained in copper concentrate will be reduced, and it will be possible to improve the profitability of the copper smelting business by preventing the increase in costs of the copper smelting process and reducing slag generation.
本実施形態に係る処理方法が対象とする出発原料は、銅精鉱である。当該銅精鉱は、黄銅鉱および黄鉄鉱を含んでいる。たとえば、当該銅精鉱は、重量比にて、黄銅鉱１に対して黄鉄鉱を０．２５〜１．７含んでいる。また、当該銅精鉱は、黄銅鉱に対して重量比で０以上１未満の高銅品位鉱を含んでいる。高銅品位鉱は、銅藍（ＣｕＳ）、輝銅鉱（Ｃｕ２Ｓ）、ダイジェナイト（Ｃｕ２−ｘＳ（ｘ＝０．４５〜１））、ハン銅鉱（Ｃｕ５ＦｅＳ４）、およびアイダ鉱（Ｃｕ５ＦｅＳ６）の少なくともいずれかを含む。上記鉱物以外に微量に検出される成分も、不純物として本実施形態に係る銅精鉱に含まれる。当該銅精鉱は、例えば、Ｃｕを１５ｍａｓｓ％〜３０ｍａｓｓ％、Ｆｅを２０ｍａｓｓ％〜３５ｍａｓｓ％含有する。このような銅精鉱は、鉄を多く含むため、製錬工程において、多量のスラグ発生をもたらす。 The starting material targeted by the treatment method according to the present embodiment is copper concentrate. The copper concentrate contains chalcopyrite and pyrite. For example, the copper concentrate contains 0.25 to 1.7 pyrite with respect to chalcopyrite 1 by weight ratio. Moreover, the said copper concentrate contains the high copper grade ore of 0 or more and less than 1 by weight ratio with respect to chalcopyrite. High copper grades include copper indigo (CuS), chalcocite (Cu 2 S), digenite (Cu 2-x S (x = 0.45-1)), han copper ore (Cu 5 FeS 4 ), and Ida It contains at least one of ore (Cu 5 FeS 6 ). In addition to the minerals, components detected in trace amounts are also included in the copper concentrate according to the present embodiment as impurities. The copper concentrate contains, for example, 15 mass% to 30 mass% of Cu and 20 mass% to 35 mass% of Fe. Since such a copper concentrate contains a lot of iron, a large amount of slag is generated in the smelting process.
図１は、本実施形態に係る銅精鉱の処理方法の一例を示す工程図である。図１を参照して、まず、銅精鉱に対して、硫化変換工程を実施する。例えば、銅精鉱中の銅（Ｃｕ）に対して、硫黄（Ｓ）を１．０から２．０のモル当量比で添加する。添加硫黄量を増してもよいが、反応性の向上は確認できず、連続処理する際の試料の流動性低下、変換銅精鉱への単体硫黄の残存などの弊害が多くなる。銅精鉱に対して、単体硫黄を混合することによって供給してもよく、別の容器で加熱して得た硫黄蒸気を供給してもよい。 FIG. 1 is a process diagram showing an example of a copper concentrate processing method according to this embodiment. With reference to FIG. 1, first, a sulfidation conversion step is performed on copper concentrate. For example, sulfur (S) is added at a molar equivalent ratio of 1.0 to 2.0 with respect to copper (Cu) in the copper concentrate. Although the amount of added sulfur may be increased, an improvement in reactivity cannot be confirmed, and adverse effects such as a decrease in fluidity of the sample during continuous processing and residual single element sulfur in the converted copper concentrate increase. The copper concentrate may be supplied by mixing simple sulfur, or sulfur vapor obtained by heating in a separate container may be supplied.
次に、硫黄を添加した銅精鉱に対して熱処理を実施することによって、黄鉄鉱と、銅藍粒子および／またはヌクンダマイト(Ｃｕ４−ｘＦｅｘＳ４（ｘ＝０．３３〜０．６２）)とを含む硫化精鉱粒子を得る。上記熱処理においては、硫黄を添加した銅精鉱に対して、不活性雰囲気において所定の温度および所定の時間で熱処理を施す。この熱処理は、例えば、ロータリキルンなどを用いて行うことができる。例えば、不活性雰囲気として、窒素ガスを用いることができる。また、熱処理時間を３０分〜６０分とすることが好ましい。未反応黄銅鉱の残存量を低下させることができるからである。熱処理温度の詳細については後述する。 Next, heat treatment is performed on the copper concentrate to which sulfur has been added, so that pyrite, copper indigo particles and / or nukundamite (Cu 4-x Fe x S 4 (x = 0.33 to 0.62)) ). In the heat treatment, heat treatment is performed on the copper concentrate to which sulfur is added at a predetermined temperature and a predetermined time in an inert atmosphere. This heat treatment can be performed using, for example, a rotary kiln. For example, nitrogen gas can be used as the inert atmosphere. Moreover, it is preferable that heat processing time shall be 30 minutes-60 minutes. This is because the remaining amount of unreacted chalcopyrite can be reduced. Details of the heat treatment temperature will be described later.
図２は、銅精鉱に対する３８５℃での硫化変換工程後における硫化精鉱粒子の電子線マイクロアナライザ（ＥＰＭＡ）組成像である。図２に示すように、銅精鉱中の黄銅鉱が消失し、銅藍と黄鉄鉱とが硫化精鉱粒子に含まれている。銅精鉱中の黄銅鉱粒子は内殻として黄鉄鉱が存在し、黄鉄鉱を銅藍が外殻として覆う粒子に変換される。 FIG. 2 is an electron microanalyzer (EPMA) composition image of sulfide concentrate particles after the sulfide conversion step at 385 ° C. for copper concentrate. As shown in FIG. 2, the chalcopyrite in the copper concentrate disappears, and the copper indigo and the pyrite are contained in the sulfide concentrate particles. The chalcopyrite particles in the copper concentrate have pyrite as the inner shell, and the pyrite is converted into particles that cover the pyrite with copper indigo as the outer shell.
このような硫化精鉱粒子から高Ｃｕ品位鉱物を主体として回収するためには、各硫化精鉱粒子を高Ｃｕ品位鉱物と、黄鉄鉱とに単体分離できることが好ましい。本発明者らが鋭意試験・調査した結果、単体分離に好ましい粒子径は、概ね１０μｍより小さい粒子であることがわかった。そこで、硫化変換工程によって得られた硫化精鉱粒子に対して、摩鉱工程を実施する。摩鉱に用いる粉砕機は、ボールミル、ジェットミル、アトリッションミル、チューブミル等である。摩鉱精鉱粒子の５０％通過粒子径（Ｐ５０）が５μｍ〜１７μｍ程度の範囲に入るように摩鉱度を調整できるものであれば、湿式および乾式を問わず、粉砕機の種類は問われない。また、５０％通過粒子径は５〜１０μｍ程度がより好ましい。 In order to recover mainly the high Cu grade mineral from such sulfide concentrate particles, it is preferable that each sulfide concentrate particle can be separated into a high Cu grade mineral and pyrite. As a result of intensive studies and investigations by the present inventors, it was found that the preferable particle size for single-unit separation is generally smaller than 10 μm. Then, a grinding process is implemented with respect to the sulfide concentrate particle | grains obtained by the sulfide conversion process. The pulverizer used for grinding is a ball mill, a jet mill, an attrition mill, a tube mill or the like. As long as the milling degree can be adjusted so that the 50% passing particle diameter (P 50 ) of the milling concentrate particles falls within the range of about 5 μm to 17 μm, regardless of the wet type or the dry type, the type of pulverizer is not questioned. I will not. The 50% passing particle diameter is more preferably about 5 to 10 μm.
次に、摩鉱工程で得られた摩鉱精鉱粒子に対して、浮遊選鉱工程を実施する。浮遊選鉱工程においては、空気供給式浮選機、空気吸込式浮選機、機械攪拌式浮選機、あるいはこれらの組み合わせを用いることができる。浮遊選鉱工程においてｐＨ調整剤としてＣａ（ＯＨ）２を用い、捕収剤として銅藍およびヌクンダマイトを優先的に捕収するブチルザンセート（ＢＸ）を用いることで、Ｃｕ品位の高い浮選精鉱とＦｅ品位の高い浮選尾鉱との分離が容易となる。なお、浮遊選鉱工程におけるｐＨ調整剤および捕収剤はこれに限られる訳ではない。ｐＨ調整剤は、例えば、ＮａＯＨを用いることもできる。捕収剤は、高Ｃｕ品位鉱および高Ｆｅ品位鉱のいずれか一方を優先的に捕収するものであればよく、例えば、アミルザンセート（ＡＸ）やエチルザンセート（ＥＸ）を用いることもできる。 Next, a flotation process is performed on the mill concentrate particles obtained in the mill process. In the flotation process, an air supply type flotation machine, an air suction type flotation machine, a mechanical stirring type flotation machine, or a combination thereof can be used. In the flotation process, Ca (OH) 2 is used as a pH adjuster, and butyl xanthate (BX) that preferentially collects copper indigo and nukundamite is used as a collection agent. Can be easily separated from flotation tailings with high Fe grade. Note that the pH adjusting agent and the collecting agent in the flotation process are not limited thereto. For example, NaOH may be used as the pH adjuster. The collector may be any one that preferentially collects either high Cu grade or high Fe grade ore, and for example, amyl xanthate (AX) or ethyl xanthate (EX) can also be used.
浮遊選鉱工程における起泡剤は、特に限定されるものではない。起泡剤の一例として、メチルイソブチルカルビノール（ＭＩＢＣ）、パイン油などを用いることができる。浮遊選鉱工程の条件は、選別精鉱のＣｕ品位、浮遊選鉱工程におけるＣｕ回収率、処理コストなどに応じて、任意に変更可能である。また、Ｃｕ品位およびＣｕ回収率のさらなる向上を狙う場合は、浮遊選鉱工程を多段にわたって実施すればよい。または一旦、摩鉱精鉱粒子の５０％通過粒子径（Ｐ５０）を１０〜１７μｍで浮遊選鉱し、回収した浮選精鉱と浮選尾鉱とに分けた後、５０％通過粒子径（Ｐ５０）を５μｍ〜１０μｍに再摩鉱して浮遊選鉱工程を再度実施すればよい。 The foaming agent in the flotation process is not particularly limited. As an example of the foaming agent, methyl isobutyl carbinol (MIBC), pine oil, or the like can be used. The conditions of the flotation process can be arbitrarily changed according to the Cu quality of the selected concentrate, the Cu recovery rate in the flotation process, the processing cost, and the like. Moreover, what is necessary is just to implement a flotation process in multiple steps, when aiming at the further improvement of Cu quality and Cu collection | recovery rate. Alternatively, after 50% passing particle diameter (P 50 ) of the fine ore concentrate particles is float-separated at 10 to 17 μm and divided into the recovered flotation concentrate and the flotation tailing, 50% passing particle diameter ( P 50 ) may be re-milled to 5 μm to 10 μm and the flotation process may be performed again.
浮遊選鉱工程の実施によって、摩鉱精鉱粒子は、浮遊する浮選精鉱と沈降する浮選尾鉱とに分離する。捕収剤にブチルザンセート等を用いることで、捕収剤によって銅藍およびヌクンダマイトが優先的に捕収され、浮選精鉱には銅藍およびヌクンダマイトが比較的多く含まれ、浮選尾鉱には黄鉄鉱が比較的多く含まれる。すなわち、浮選精鉱にはＣｕ品位の高い鉱物が比較的多く含まれ、浮選尾鉱にはＦｅ品位の高い鉱物が比較的多く含まれる。したがって、浮遊選鉱工程によって得られた浮選精鉱を回収することによって、Ｃｕ品位の高い銅精鉱を効率よくかつ経済的に回収することができる。 By performing the flotation process, the ore concentrate particles are separated into a flotation concentrate that floats and a flotation tailing that settles. By using butyl xanthate or the like as the collector, copper indigo and nukundamite are preferentially collected by the collector, and the flotation concentrate contains a relatively large amount of copper indigo and nukudamite. Contains a relatively large amount of pyrite. That is, the flotation concentrate contains a relatively high amount of high-quality Cu minerals, and the flotation tailing contains a relatively high amount of high-quality Fe minerals. Accordingly, by recovering the flotation concentrate obtained by the flotation process, it is possible to efficiently and economically recover the copper concentrate having a high Cu quality.
得られた浮選精鉱は銅精鉱からのＣｕロスが少なく、Ｆｅ分や石英（ＳｉＯ２）などの脈石成分が除去され、Ｃｕ品位が向上している。銅鉱山においては銅精鉱のＣｕ品位改善のための選鉱処理設備・コストの低減が期待できる。また、銅製錬所においてはスラグ発生量の少ない銅製錬を行うことから、スラグ販売での損益の改善、スラグ取扱設備の工作費の低減、銅精鉱の取扱量減少に伴う輸送・乾燥設備の工作費やエネルギーコストの低減などの効果が期待できる。 The obtained flotation concentrate has less Cu loss from the copper concentrate, removes gangue components such as Fe and quartz (SiO 2 ), and improves the Cu quality. In the copper mine, reduction of the beneficiation processing equipment and cost for improving the copper quality of the copper concentrate can be expected. In addition, because copper smelters produce less slag, copper smelters improve slag profits and losses, reduce slag handling equipment costs, and reduce transportation and drying facilities associated with reduced copper concentrate handling. Effects such as reduction of work costs and energy costs can be expected.
ここで、硫化変換工程の熱処理温度について説明する。硫化変換において、高銅品位鉱は、以下のような反応式によって黄鉄鉱および黄銅鉱と反応すると推定される。
２Ｃｕ２Ｓ ＋ ＣｕＦｅＳ２ → Ｃｕ５ＦｅＳ４ （１）
５Ｃｕ２Ｓ ＋ ２ＦｅＳ２ → ２Ｃｕ５ＦｅＳ４ ＋ Ｓ （２）
Ｃｕ５ＦｅＳ４ ＋ ３Ｓ → ５ＣｕＳ ＋ ＦｅＳ２ （３）
Here, the heat treatment temperature in the sulfidation conversion step will be described. In sulfide conversion, high copper grade ore is estimated to react with pyrite and chalcopyrite according to the following reaction formula.
2Cu 2 S + CuFeS 2 → Cu 5 FeS 4 (1)
5Cu 2 S + 2FeS 2 → 2Cu 5 FeS 4 + S (2)
Cu 5 FeS 4 + 3S → 5CuS + FeS 2 (3)
これらの反応により、内殻をＦｅＳ２、外殻をＣｕＳとする精鉱に変換され、あるいはハン銅鉱、ヌクンダマイト等のＣｕ−Ｆｅ−Ｓ三元系高銅品位鉱に変換される。これらの反応は温度が高いほど促進されるが、特にＣｕ−Ｆｅ−Ｓ三元系高銅品位鉱に変換される例えば（１）式および（２）式のような反応は、より温度が高い場合に促進される。また、この反応に伴い、元々単独で存在していた鉱物同士が結合することにより粗大な変換精鉱粒子が形成される効果もある。これにより次の摩鉱工程にて鉱物種ごとに単体に分離した粒子が形成されやすくなる。その結果、浮選工程にてＣｕとＦｅとの分離性を向上させることができる。 By these reactions, the inner shell is converted to concentrate with FeS 2 and the outer shell is CuS, or is converted to Cu—Fe—S ternary high copper grade ore such as han copper ore and nukundamite. These reactions are promoted as the temperature increases. In particular, the reactions such as the formulas (1) and (2) that are converted into Cu—Fe—S ternary high copper grade ore have higher temperatures. If promoted. Moreover, accompanying this reaction, there is also an effect that coarse converted concentrate particles are formed by combining minerals originally present alone. This facilitates the formation of particles separated into individual mineral species in the next milling process. As a result, the separation between Cu and Fe can be improved in the flotation process.
一方で、黄銅鉱粒子単独でも硫化変換時に以下の反応を起こすと考えられる。
ＣｕＦｅＳ２ ＋ Ｓ → ＣｕＳ ＋ ＦｅＳ２ （４）
５ＣｕＦｅＳ２ ＋ ２Ｓ → Ｃｕ５ＦｅＳ４ ＋ ４ＦｅＳ２ （５）
また、より温度が高いほど（４）式の反応に加えて（５）式の反応が促進される。黄銅鉱粒子が単独で硫化変換する場合、ＣｕＳ及びＣｕ−Ｆｅ−Ｓ三元系高銅品位鉱の粒子が粗大化する効果は無いため、上述したような浮遊選鉱工程におけるＣｕとＦｅとの分離性向上の効果は得られない。むしろ、ＦｅがＣｕ−Ｆｅ−Ｓ三元系高銅品位鉱として浮鉱となり精鉱中に回収されることによりＣｕ品位を低下させる悪影響の方が大きくなる。したがって、高銅品位鉱が黄銅鉱よりも重量比で少ない場合には、黄銅鉱粒子の（５）式による反応が起こりにくい比較的低い温度において硫化変換することが好ましい。
On the other hand, even the chalcopyrite particles alone are considered to cause the following reaction during sulfidation conversion.
CuFeS 2 + S → CuS + FeS 2 (4)
5CuFeS 2 + 2S → Cu 5 FeS 4 + 4FeS 2 (5)
Further, the higher the temperature, the more the reaction of the formula (5) is promoted in addition to the reaction of the formula (4). When chalcopyrite particles undergo sulfidation conversion alone, CuS and Cu-Fe-S ternary high copper grade ore particles have no effect of coarsening, so separation of Cu and Fe in the flotation process as described above The effect of improving the properties cannot be obtained. Rather, Fe is floated as a Cu—Fe—S ternary high copper grade ore and recovered in the concentrate, and the adverse effect of lowering the Cu grade becomes greater. Therefore, when the high copper grade ore is smaller in weight ratio than chalcopyrite, it is preferable to perform sulfidation conversion at a relatively low temperature at which the reaction according to the formula (5) of the chalcopyrite particles hardly occurs.
ただし、熱処理温度が低すぎると、大粒の硫化精鉱粒子に未反応黄銅鉱が残存することや、変換粒子の外側に生成する銅藍層が未発達で薄く、摩鉱時に、銅藍が細かくなり易く、浮鉱への銅藍の回収率の低下、尾鉱への混入が生じる。本発明者らが鋭意試験・調査した結果、硫化変換工程における好ましい熱処理温度は、３３０℃〜４５０℃である。なお、出発原料において、黄銅鉱に対する高銅品位鉱の重量比は、０以上０．５未満としてもよく、０以上０．４未満としてもよい。 However, if the heat treatment temperature is too low, unreacted chalcopyrite remains on large sulfide concentrate particles, or the copper indigo layer formed on the outside of the conversion particles is undeveloped and thin, and copper indigo is fine during grinding. It tends to occur, and the recovery rate of copper indigo into floating ore and contamination into tailings occur. As a result of intensive studies and investigations by the present inventors, a preferable heat treatment temperature in the sulfidation conversion step is 330 ° C. to 450 ° C. In the starting material, the weight ratio of the high copper grade ore to chalcopyrite may be 0 or more and less than 0.5, or may be 0 or more and less than 0.4.
上記熱処理温度は、硫化精鉱粒子に黄鉄鉱と銅藍とが共存する温度範囲、または、黄鉄鉱と銅藍と少量のヌクンダマイトとが共存する温度範囲であることが好ましい。本発明者らが鋭意試験・調査した結果、黄鉄鉱と銅藍と少量のヌクンダマイトとが共存する温度範囲は、３５０℃〜４１０℃である。この温度範囲により銅精鉱を処理することで、硫化精鉱粒子中の未反応黄銅鉱の残存、また、銅藍よりＦｅ品位の低いヌクンダマイトの生成を抑制でき、浮遊選鉱工程におけるＣｕとＦｅとの分離性が向上する。 The heat treatment temperature is preferably in the temperature range in which pyrite and copper indigo coexist with sulfide concentrate particles, or in the temperature range in which pyrite, copper indigo and a small amount of nukundamite coexist. As a result of intensive studies and investigations by the present inventors, the temperature range in which pyrite, copper indigo blue and a small amount of nukundamite coexist is 350 ° C. to 410 ° C. By treating the copper concentrate in this temperature range, it is possible to suppress the remaining unreacted chalcopyrite in the sulfide concentrate particles and the formation of nukundamite having a lower Fe grade than copper indigo, and Cu and Fe in the flotation process The separability is improved.
以下、上記実施形態に係る処理方法に従って、銅精鉱を処理した。 Hereinafter, the copper concentrate was processed according to the processing method according to the above embodiment.
（実施例１）
実施例１の試験に供した銅精鉱（元精鉱）のＣｕ品位は２１ｍａｓｓ％であり、Ｆｅ品位は３２ｍａｓｓ％であり、Ｓ品位は４２ｍａｓｓ％であった。図３に、Ｘ線回折(ＸＲＤ)結果を示す。図３の結果および電子線マイクロアナライザ(ＥＰＭＡ)によって特定された鉱物種と銅精鉱品位とから、鉱物組成は、黄銅鉱(ＣｕＦｅＳ２)４３ｍａｓｓ％、黄鉄鉱(ＦｅＳ２)４１ｍａｓｓ％、銅藍（Ｃｕ２Ｓ）１５ｍａｓｓ％、脈石成分(ＳｉＯ２等)が２ｍａｓｓ％であった。すなわち、元精鉱においては、黄銅鉱に対する高銅品位鉱の重量比が０以上１未満であった。また、粒度分布測定（レーザー回析法）から、銅精鉱の５０％通過粒子径（Ｐ５０）は３７μｍであった。
Example 1
In the copper concentrate (original concentrate) subjected to the test of Example 1, the Cu quality was 21 mass%, the Fe quality was 32 mass%, and the S quality was 42 mass%. FIG. 3 shows the X-ray diffraction (XRD) results. From the results of FIG. 3 and the mineral species and copper concentrate grade specified by the electron microanalyzer (EPMA), the mineral composition is 43 mass% of chalcopyrite (CuFeS 2 ), 41 mass% of pyrite (FeS 2 ), copper indigo ( Cu 2 S) was 15 mass%, and the gangue component (SiO 2 or the like) was 2 mass%. That is, in the former concentrate, the weight ratio of the high copper grade ore to chalcopyrite was 0 or more and less than 1. From the particle size distribution measurement (laser diffraction method), the 50% passing particle diameter (P 50 ) of the copper concentrate was 37 μm.
硫化変換工程においては、銅精鉱と単体硫黄とを、モル比で銅精鉱中Ｃｕ：Ｓ＝１:１．９で混合し、窒素雰囲気中において、３５０℃で６０分処理することで、銅藍および黄鉄鉱を含む硫化精鉱粒子に変換した。図４は、硫化精鉱粒子のＸＲＤ解析結果を示す図である。図３および図４から、銅精鉱中の黄銅鉱は硫化変換工程により銅藍、黄鉄鉱に変化していることがわかる。 In the sulfidation conversion step, copper concentrate and elemental sulfur are mixed at a molar ratio of Cu: S = 1: 1.9 in the copper concentrate and treated at 350 ° C. for 60 minutes in a nitrogen atmosphere. It was converted into sulfide concentrate particles containing copper indigo and pyrite. FIG. 4 is a diagram showing a result of XRD analysis of sulfide concentrate particles. 3 and 4, it can be seen that the chalcopyrite in the copper concentrate is changed to copper indigo and pyrite by the sulfide conversion process.
摩鉱工程においては、硫化精鉱粒子(Ｃｕ品位＝１９．５ｍａｓｓ％、Ｆｅ品位＝３０．７ｍａｓｓ％)を、湿式ボールミルを用いて摩鉱した。５０％通過粒子径（Ｐ５０）が５．２μｍ，１１μｍ，１４μｍ，３０μｍである摩鉱精鉱粒子に対して浮遊選鉱工程を実施した。浮遊選鉱工程においては、京大式アジテア型試験浮選機を用いた。パルプ濃度１００ｇ／ｌの摩鉱精鉱粒子のスラリーを、浮選セル内に入れ、Ｃａ(ＯＨ)２飽和溶液添加によりｐＨ１２.５とし、摩鉱精鉱粒子あたり１００ｇ／ｔに相当する量の捕収剤ブチルザンセート(ＢＸ)を添加し、コンディショニングとして１０分間攪拌した。その後、起泡剤としてメチルイソブチルカルビノール(ＭＩＢＣ)を２０μｌ／ｌ添加し、浮選機に空気を供給し、浮鉱(フロス)を回収した。浮鉱回収においては、鉱物の付着した安定的な気泡がなくなるまで回収し、その後、初期供給精鉱に対し、ＢＸを１００ｇ／ｔ、ＭＩＢＣを７μｌ／ｌずつ回分添加し、同様の操作を１０回繰り返した。この浮遊選鉱により回収した各浮鉱と、浮選機セル内のスラリーに最後まで残存した尾鉱のＣｕ品位、およびＦｅ品位を分析した。Ｃｕ品位に応じて、浮選精鉱と浮選尾鉱とに仕分けした。浮選精鉱Ｃｕ回収率が９０％以上の結果を表１に示す。表1中の総合分離効率は(６)式で示される。総合分離効率は、浮選における変換精鉱中のＣｕ分とＦｅ分との分離性を示す。浮選精鉱のＣｕ回収率が高く、さらに、浮選尾鉱のＦｅ回収率が高ければ、総合分離効率は高い数値となり、本発明の目的に対し、好ましい結果となる。 In the milling process, sulfide concentrate particles (Cu grade = 19.5 mass%, Fe grade = 30.7 mass%) were milled using a wet ball mill. The flotation process was carried out on the ore concentrate particles having 50% passing particle diameter (P 50 ) of 5.2 μm, 11 μm, 14 μm, and 30 μm. In the flotation process, a Kyoto University agitator type test flotation machine was used. A slurry of fine concentrate particles having a pulp concentration of 100 g / l is placed in a flotation cell, adjusted to pH 12.5 by adding a Ca (OH) 2 saturated solution, and an amount corresponding to 100 g / t of fine concentrate particles. A collector butyl xanthate (BX) was added, and the mixture was stirred for 10 minutes for conditioning. Thereafter, 20 μl / l of methyl isobutyl carbinol (MIBC) was added as a foaming agent, air was supplied to the flotation machine, and flotation was recovered. In the floatation recovery, recovery is performed until there are no stable bubbles with minerals attached, and then 100 g / t of BX and 7 μl / l of MIBC are added in batches to the initially supplied concentrate, and the same operation is performed 10 times. Repeated times. The Cu grade and Fe grade of each float ore collected by this flotation and the tailings remaining in the slurry in the flotation cell until the end were analyzed. According to Cu grade, it sorted into flotation concentrate and flotation tailings. Table 1 shows the results of the flotation concentrate Cu recovery rate of 90% or more. The total separation efficiency in Table 1 is expressed by equation (6). The total separation efficiency indicates the separation between the Cu content and the Fe content in the converted concentrate in flotation. If the Cu recovery rate of the flotation concentrate is high and the Fe recovery rate of the flotation tailing is high, the total separation efficiency becomes a high value, which is a favorable result for the purpose of the present invention.
摩鉱精鉱粒子が５μｍ＜Ｐ５０≦１７μｍの範囲において、Ｐ５０＝３０μｍの場合に比べ、浮選精鉱Ｃｕ品位、浮選尾鉱Ｃｕ品位、およびＦｅ回収率が良好な結果であり、分離性を示す総合分離効率も高い結果を得た。摩鉱精鉱粒子を５μｍ＜Ｐ５０≦１７μｍの範囲にすることで、銅藍と黄鉄鉱との単体割合が多くなり、分離性が向上した。また、Ｐ５０＝５．２μｍにおける総合分離効率が最も高く、より好ましい結果である。硫化精鉱粒子を５μｍ＜Ｐ５０≦１７μｍ、より好ましくは５〜１０μｍに摩鉱することで、Ｃｕ品位の高い銅精鉱を回収できる。 In the range of 5 μm <P 50 ≦ 17 μm for the fine concentrate particles, the flotation concentrate Cu quality, the flotation tail Cu quality, and the Fe recovery rate are better than those in the case of P 50 = 30 μm. A high overall separation efficiency showing separability was also obtained. By setting the fine ore concentrate particles in the range of 5 μm <P 50 ≦ 17 μm, the simple substance ratio of copper indigo and pyrite increased and the separability improved. Moreover, the total separation efficiency at P 50 = 5.2 μm is the highest, which is a more preferable result. By grinding the sulfide concentrate particles to 5 μm <P 50 ≦ 17 μm, more preferably 5 to 10 μm, copper concentrate with high Cu quality can be recovered.
(実施例２)
実施例１の出発原料の銅精鉱を、反応時間６０分で、３５０℃、３８５℃、４００℃、４２５℃の４水準の熱処理温度で硫化変換工程を実施した。硫化変換工程後のそれぞれのＸＲＤ解析結果を図４、図５、図６および図７に示した。図４および図５に示すように、３５０℃および３８５℃の硫化変換後の硫化精鉱粒子においては、銅藍および黄鉄鉱のピークのみ確認できた。図６に示すように、４００℃の硫化変換後の硫化精鉱粒子においては、銅藍および黄鉄鉱の一部がヌクンダマイト(Ｃｕ４−ｘＦｅｘＳ４)となっていた。また、図７に示すように、４２５℃の硫化変換後の硫化精鉱粒子においては、ヌクンダマイトの強度が増加した。
(Example 2)
The copper concentrate as a starting material of Example 1 was subjected to a sulfidation conversion step at a heat treatment temperature of four levels of 350 ° C., 385 ° C., 400 ° C., and 425 ° C. with a reaction time of 60 minutes. The respective XRD analysis results after the sulfidation conversion step are shown in FIG. 4, FIG. 5, FIG. 6, and FIG. As shown in FIGS. 4 and 5, only the peaks of copper indigo and pyrite were confirmed in the sulfide concentrate particles after sulfidation at 350 ° C. and 385 ° C. As shown in FIG. 6, in the sulfide concentrate particles after sulfidation conversion at 400 ° C., a part of copper indigo and pyrite was nukundamite (Cu 4−x Fe x S 4 ). Moreover, as shown in FIG. 7, in the sulfide concentrate particles after sulfidation conversion at 425 ° C., the strength of nukudamite increased.
硫化変換工程で得られた硫化精鉱粒子を、実施例１と同様に５０％通過粒子径（Ｐ５０）が１４μｍ〜１７μｍになるように摩鉱し、浮遊選鉱処理を実施した。浮選精鉱Ｃｕ回収率が９０％以上の結果を表２に示す。
熱処理温度が３３０〜４５０℃の温度範囲において、浮選精鉱は９０％以上のＣｕ回収率で、元銅精鉱よりＣｕ品位が向上した。また、浮選尾鉱としてＦｅ分を３０％以上除去できる結果を得た。３５０〜４１０℃の温度範囲においては、４２５℃の結果と比べ、浮選精鉱のＣｕ品位、浮選尾鉱のＦｅ回収率、およびＣｕとＦｅとの分離性を示す総合分離効率において良好な結果を得た。硫化変換工程の熱処理温度は３３０〜４５０℃、より好ましくは３５０〜４１０℃にすることで、Ｃｕ品位の高い銅精鉱を回収できる。
In the temperature range where the heat treatment temperature was 330 to 450 ° C., the flotation concentrate had a Cu recovery rate of 90% or more, and the Cu quality was improved over the original copper concentrate. Moreover, the result which can remove
以上、本発明の実施例について詳述したが、本発明は係る特定の実施例に限定されるものではなく、特許請求の範囲に記載された本発明の要旨の範囲内において、種々の変形・変更が可能である。 Although the embodiments of the present invention have been described in detail above, the present invention is not limited to such specific embodiments, and various modifications and changes can be made within the scope of the gist of the present invention described in the claims. It can be changed.
Claims (8)
前記硫化精鉱粒子を摩鉱する摩鉱工程と、
前記摩鉱工程で得られる摩鉱精鉱粒子に対して浮遊選鉱処理する浮遊選鉱工程と、を含むことを特徴とする銅精鉱の処理方法。 Copper concentrate particles containing at least chalcopyrite and pyrite and containing high copper grade ore having a weight ratio of 0 to less than 0.5 with respect to the chalcopyrite are reacted with sulfur at 330 ° C. to 410 ° C. in an inert gas atmosphere. A sulfide conversion step to obtain sulfide concentrate particles by
A grinding process for grinding the sulfide concentrate particles;
A method of treating a copper concentrate, comprising: a flotation process for flotation treatment of the mill concentrate particles obtained in the milling process.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP2012081308A JP5888780B2 (en) | 2012-03-30 | 2012-03-30 | Copper concentrate processing method |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP2012081308A JP5888780B2 (en) | 2012-03-30 | 2012-03-30 | Copper concentrate processing method |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| JP2013209718A JP2013209718A (en) | 2013-10-10 |
| JP5888780B2 true JP5888780B2 (en) | 2016-03-22 |
Family
ID=49527810
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| JP2012081308A Active JP5888780B2 (en) | 2012-03-30 | 2012-03-30 | Copper concentrate processing method |
Country Status (1)
| Country | Link |
|---|---|
| JP (1) | JP5888780B2 (en) |
Cited By (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN111515028A (en) * | 2020-04-13 | 2020-08-11 | 西部矿业股份有限公司 | Complex copper-lead-zinc polymetallic ore stage grinding and beneficiation method and collecting agent thereof |
Families Citing this family (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN109158219A (en) * | 2018-06-27 | 2019-01-08 | 昆明理工大学 | A kind of carrier flotation method containing uytenbogaardtite |
Family Cites Families (5)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JP4274079B2 (en) * | 2004-08-19 | 2009-06-03 | 住友金属鉱山株式会社 | Method for recovering gold concentrate from copper concentrate leach residue |
| US20100024601A1 (en) * | 2006-12-18 | 2010-02-04 | Alexander Beckmann | Method for obtaining copper and precious metals from copper-iron sulphide ores or ore concentrates |
| JP2010229542A (en) * | 2009-03-04 | 2010-10-14 | Sumitomo Metal Mining Co Ltd | Method of separating pyrite from copper-containing material |
| JP5385235B2 (en) * | 2010-09-07 | 2014-01-08 | Ｊｘ日鉱日石金属株式会社 | Copper concentrate processing method |
| JP2013155426A (en) * | 2012-01-31 | 2013-08-15 | Jx Nippon Mining & Metals Corp | Method for treating copper concentrate |
-
2012
- 2012-03-30 JP JP2012081308A patent/JP5888780B2/en active Active
Cited By (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN111515028A (en) * | 2020-04-13 | 2020-08-11 | 西部矿业股份有限公司 | Complex copper-lead-zinc polymetallic ore stage grinding and beneficiation method and collecting agent thereof |
| CN111515028B (en) * | 2020-04-13 | 2022-06-07 | 西部矿业股份有限公司 | Complex copper-lead-zinc polymetallic ore stage grinding and beneficiation method and collecting agent thereof |
Also Published As
| Publication number | Publication date |
|---|---|
| JP2013209718A (en) | 2013-10-10 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| Schlesinger et al. | Extractive metallurgy of copper | |
| Rashchi et al. | Anglesite flotation: a study for lead recovery from zinc leach residue | |
| Sasaki et al. | Spectroscopic study on oxidative dissolution of chalcopyrite, enargite and tennantite at different pH values | |
| US8960444B2 (en) | Method for separating arsenic mineral from copper-bearing material with high arsenic grade | |
| Ivanik | Flotation extraction of elemental sulfur from gold-bearing cakes | |
| JP4572703B2 (en) | Separation of arsenic minerals from copper concentrate | |
| JP5502006B2 (en) | Copper concentrate processing method | |
| JP2017202481A (en) | Beneficiation method | |
| Skandrani et al. | Desulfurization of aged gold-bearing mine tailings | |
| JP2013155426A (en) | Method for treating copper concentrate | |
| JP5888780B2 (en) | Copper concentrate processing method | |
| JP5497723B2 (en) | Copper concentrate processing method | |
| JP2012201920A (en) | Method for treating copper concentrate | |
| JP5385235B2 (en) | Copper concentrate processing method | |
| JP2018034128A (en) | Method for separating molybdenum concentrate | |
| JP7299592B2 (en) | beneficiation method | |
| WO2017110462A1 (en) | Mineral dressing method | |
| JP5641952B2 (en) | Copper concentrate processing method | |
| JP2012201922A (en) | Method for treating copper concentrate | |
| WO2007115377A1 (en) | Process for recovery of antimony and metal values from antimony- and metal value-bearing materials | |
| US8931642B2 (en) | Activated flotation circuit for processing combined oxide and sulfide ores | |
| Rabatho et al. | Investigation of a flotation process with de-sliming and attrition to upgrade and recover Cu and Mo from a Cu-Mo flotation tailing | |
| JP6442636B1 (en) | Beneficiation method | |
| Allahkarami et al. | Studies of Grinding Media Corrosion from Galvanic Interaction on Galena Flotation | |
| Talan | Beneficiation of oxide lead and zinc minerals by selective flotation and ammonia leaching |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| A621 | Written request for application examination |
Free format text: JAPANESE INTERMEDIATE CODE: A621 Effective date: 20140917 |
|
| A977 | Report on retrieval |
Free format text: JAPANESE INTERMEDIATE CODE: A971007 Effective date: 20151005 |
|
| A131 | Notification of reasons for refusal |
Free format text: JAPANESE INTERMEDIATE CODE: A131 Effective date: 20151013 |
|
| A521 | Request for written amendment filed |
Free format text: JAPANESE INTERMEDIATE CODE: A523 Effective date: 20151207 |
|
| TRDD | Decision of grant or rejection written | ||
| A01 | Written decision to grant a patent or to grant a registration (utility model) |
Free format text: JAPANESE INTERMEDIATE CODE: A01 Effective date: 20160209 |
|
| A61 | First payment of annual fees (during grant procedure) |
Free format text: JAPANESE INTERMEDIATE CODE: A61 Effective date: 20160212 |
|
| R150 | Certificate of patent or registration of utility model |
Ref document number: 5888780 Country of ref document: JP Free format text: JAPANESE INTERMEDIATE CODE: R150 |
|
| R250 | Receipt of annual fees |
Free format text: JAPANESE INTERMEDIATE CODE: R250 |
|
| R250 | Receipt of annual fees |
Free format text: JAPANESE INTERMEDIATE CODE: R250 |
|
| S531 | Written request for registration of change of domicile |
Free format text: JAPANESE INTERMEDIATE CODE: R313531 |
|
| R350 | Written notification of registration of transfer |
Free format text: JAPANESE INTERMEDIATE CODE: R350 |
|
| R250 | Receipt of annual fees |
Free format text: JAPANESE INTERMEDIATE CODE: R250 |
|
| R250 | Receipt of annual fees |
Free format text: JAPANESE INTERMEDIATE CODE: R250 |
|
| R250 | Receipt of annual fees |
Free format text: JAPANESE INTERMEDIATE CODE: R250 |