Patent Application
20160158767
The present invention relates to the field of mineral processing technology, and particularly to a method and process for separating chalcopyrite from talc, pyrite and pyrrhotite by flotation.
Copper is an extremely important metal material in the national economic construction. With the rapid development of national economy, the copper resources which are easy to separate are getting increasingly scarce, more and more attention has been paid to the development and utilization of refractory copper resources. With the aim of realizing the efficient development and utilization of refractory chalcopyrite resources, extensive studies have been conducted by a number of research units and institutes. CN Pat. App. No. 200710180591.5 discloses a method for flotation separation of copper-sulfur, in which copper-sulfur separation of bulk concentrate is achieved at a pH of between 12.5 and 13, with the addition of combined depressor including calcium oxide, NaSO3 and sodium cyanide. However, sodium cyanide is an extremely toxic reagent, the application of which will be detrimental to the sequential tailing treatment and the environment. CN Pat. App. No. 201010539277.3 discloses a method for copper-sulfur separations, in which sodium silicate or sodium sulfide acting as depressor is employed to accomplish the separation of copper from sulfur. CN Pat. App. No. 200710035782.4 discloses an efficient and environmentally friendly method for the separation of complex sulfide ores. Sulfide flotation is performed, one or more of oxalic acid, sodium carbonate, ammonium bicarbonate, ammonium sulfate, ammonium bisulfate and ferrous sulfate acting as activator, yellow collector (xanthate), black collector (aerofloat), white collector (thiocarbanilide) and thionocarbamate acting as collector of sulfide ore, and BC acting as frother are added and agitated to disperse uniformly, in which case sulfide concentrate is recovered. Such processes are performed to achieve copper-sulfur separations on sulfide concentrate generally by using combined depressor or activator, and flotation reagents with high degree of selectivity.
Some of refractory copper ores have multiple natural types, such as chalcopyrite ore containing pyrrhotite, chalcopyrite ore containing pyrrhotite, talc and serpentine, chalcopyrite ore containing pyrite, copper skarn and so on. The chalcopyrite ore containing pyrrhotite, talc, and serpentine, such as copper ore in Dongguashan is one kind of typical refractory ore. Based on variable characteristics of Dongguashan copper ores, there are alternative separation processes. For chalcopyrite ores which are free of pyrrhotite, talc and serpentine, they are treated using combined depressor and highly effective collector in the presence of xanthate, in which case sulfide ores are floated off to separate copper from other sulfide minerals. While chalcopyrite ores containing pyrrhotite, talc and serpentine are treated using combined depressor and highly effective collector to realize copper-sulfur separation. The processes have three disadvantages as follows: Firstly, although online analysis and inspection system are adopted in the flotation circuits, there is a certain lag between the changes occurred in differential separation processes, additionally, the separation processes often fail to reach steady state in time, causing it hard to obtain a steady metallurgical performance, both of which lead to lower separation efficiency and greater copper losses especially when changes in ores type are frequent. Secondly, talc exhibits excellent floatability, therefore increasing talc pre-flotation can effectively reduce the content of silicon and magnesium in the copper concentrate, whereas chalcopyrite is also readily floatable, and a portion of chalcopyrite is floated into the talc concentrate during the talc pre-flotation, which causes copper losses and lowers the recovery of chalcopyrite. Thirdly, even though the highly selective collector or combined depressor is employed, the metallurgical performance is still not very satisfactory, and the cost of flotation reagents is remarkably high.
The present invention provides a mineral processing technology and method for efficient separation of refractory chalcopyrite ores in view of the multiple natural types of copper ores containing chalcopyrite, pyrrhotite, talc and serpentine. Specific methods are carried out according to the following steps:
(1) performing a primary grinding and classification I on the crushed chalcopyrite ore, wherein 60% of the product size of the primary classification I is −0.074 mm; carrying out a first size grading I on the product (overflows) obtained from the primary classification I, wherein the cut-point of the size grading I is 0.04 mm; conducting a second grinding and classification II on the mineral which is coarser than 0.04 mm, wherein 95% of the product size of the classification II is −0.074 mm; thereafter conducting the secondary size grading II on the product (underflows) obtained from the classification II, wherein the cut-point of the size grading II is 0.04 mm;
(2) performing the coarse-grain flotation circuits on the coarse-grain product which is coarser than 0.04 mm obtained from the size grading II; undertaking a one-stage rougher (flotation) on the coarse-grain product from the size grading II, in which calcium oxide is used to adjust the pH value, sodium carboxymethyl cellulose (CMC) used as depressor, xanthate used as collector and pine camphor oil used as frother are added to the rougher cell sequentially; conducting a two-stage scavenger (flotation) on the rougher tailings, wherein the types of reagents used in the two-stage scavenger are the same as those used in the rougher; recycling the first scavenger concentrate back to the rougher feed; performing a second scavenger on the first scavenger tailings, wherein the second scavenger tailings are the final tailings I; conducting a two-stage cleaner (flotation) on the rougher concentrate, wherein the first cleaner tailings are recirculated to the rougher feed and the first cleaner concentrate is reported to the second cleaner feed, and wherein the second cleaner tailings are recirculated to the first cleaner feed and the second cleaner concentrate is the copper concentrate I;
(3) carrying out fine-grain flotation circuits on the fine-grain product which is smaller than 0.04 mm obtained from the size grading I and the size grading II; firstly, conducting a two-stage rougher on the fine-grain product, wherein kerosene or diesel used as collector and pine camphor oil used as frother are added to the two rougher cells sequentially; performing a two-stage scavenger on the second rougher tailings, wherein kerosene or diesel oil used as collector and pine camphor oil used as frother are also added to the two scavenger cells sequentially; returning the first scavenger concentrate to the first rougher feed; recycling the second scavenger concentrate to the first scavenger feed, wherein the second scavenger tailings are the final tailings II; undertaking the pre-cleaner on the concentrates from the two-stage rougher in the absence of any reagents; subsequently recirculating the pre-cleaner tailings to the first scavenger feed; conducting a two-stage cleaning rougher on the pre-cleaner concentrate, in which calcium oxide is used to adjust the pH value, sodium carboxymethyl cellulose(CMC) used as depressor, xanthate used as collector and pine camphor oil used as frother are added to the two cleaning rougher cells sequentially; conducting a second cleaning rougher on the first cleaning rougher tailings, and conducting a two-stage cleaning scavenger on the second cleaning rougher tailings, wherein the types of reagents used in the two-stage cleaning scavenger are the same as those used in the cleaning rougher cells; performing a first cleaning scavenger on the second cleaning rougher tailings; recirculating the first cleaning scavenger concentrate to the second cleaning rougher feed and recirculating the second cleaning scavenger concentrate to the first cleaning scavenger feed, wherein the second cleaning scavenger tailings are final tailings III; performing a re-cleaner on the concentrates from the two-stage cleaning rougher; and recirculating the re-cleaner tailings to the first cleaning rougher feed, wherein the re-cleaner concentrate is the copper concentrate II.
According to the above method in step (2), the pH value in the rougher is between 11.4 and 11.6, and between 65 and 75 g·t−1 of depressor, between 60 and 90 g·t−1 of collector and between 20 and 25 g·t−1 of frother are used in the rougher. Whereas the pH value in the scavenger is also between 11.4 and 11.6, and the amounts of the depressor, collector and frother in the first scavenger are half of those in the rougher respectively, and the amounts of the depressor, collector and frother in the second scavenger are one third of those in the first rougher respectively.
According to the above method in step (3), between 35 and 40 g·t−1 of collector and between 6 and 10 g·t−1 of frother are used in the first rougher, whereas the amounts of the collector and frother in the second rougher are the same as those in the first rougher respectively. The amounts of the collector and frother in the first scavenger are half of those in the first rougher respectively. The amounts of the collector and frother in the second scavenger are one third of those in the rougher respectively.
According to the above method in step (3), the pH value in the first cleaning rougher is between 11.8 and 12.0, and between 50 and 70 g·t−1 of depressor, between 50 and 75 g·t−1 of collector and between 18 and 24 g·t−1 of frother are used in the first cleaning rougher, whereas the pH values in the second cleaning rougher, and the first and second cleaning scavengers are all between 11.8 and 12.0. The pH values in two-stage scavengers are also between 11.8 and 12.0, wherein the amounts of said depressor, collector and frother used in the first cleaning scavenger are half of those in the first cleaning rougher respectively, and the amounts of said depressor and frother used in the second cleaning scavenger are one third of those in the first cleaning rougher respectively.
In the present invention which adopts the above technical scheme, size grading not only reduces the influence of talc, pyrite and pyrrhotite on chalcopyrite ore containing pyrrhotite, talc and serpentine, but also reduces the influence of fine size fraction mineral on coarse mineral, and reduces the consumption of reagents. In the flotation of coarse chalcopyrite ore containing pyrrhotite, talc and serpentine, CMC is added to depress talc, and calcium oxide is used to control the pH and to depress pyrite and pyrrhotite. Without any influence from the fine particles, high grade copper concentrate can be obtained consequently. The separation of fine-sized chalcopyrite from pyrite, pyrrhotite and talc remains an issue. Merely using CMC to depress talc, calcium oxide to control the pH and to depress pyrite and pyrrhotite, and xanthate to collect chalcopyrite, the metallurgical performance of chalcopyrite is not efficient at all. In order to achieve effective selective separation of fine chalcopyrite, preconcentration is conducted using hydrocarbon oil collector such as kerosene or diesel oil on the rougher concentrate to separate chalcopyrite from pyrite and pyrrhotite, in which case the influence of pyrite and pyrrhotite on the flotation process is reduced, thereafter conducting cleaner on the preconcentration concentrate using CMC to depress talc and calcium oxide to control the pH and depress pyrite and pyrrhotite. In the present invention, ores are divided into fine size fraction and coarse size fraction by means of size grading and are processed separately. Flotation separation of such refractory chalcopyrite ore can be achieved by treating ores with conventional sulfide collectors and depressors, and treating ores in fine size fraction with kerosene and diesel oil, employed as collector either alone or in combination with each other. Therefore, the cost of flotation reagents is cut down greatly. In the case that the chalcopyrite ore is free of pyrrhotite, talc and serpentine, this process will not influence the final metallurgical performance. Therefore, the present process can realize the flotation separation of chalcopyrite from multiple natural types of chalcopyrite ores containing pyrrhotite, talc and serpentine.
The present invention will now be described, by way of examples, with reference to the accompanying
Primary materials of the present invention: copper ore from Dongguashan, calcium oxide, CMC, sodium ethylxanthate, kerosere, pine camphor oil.
The chalcopyrite primary ore used in the present invention mainly contained 0.94% Cu, 23.20% TFe, 14.38% S, 3.9% MgO, 28.03% SiO2, and the principal mineralogical composition were sulfide minerals (including 24.75% pyrhotite, 10.71% pyrite, 3.07% chalcopyrite, 0.01% galena, etc), 8.07% iron oxide (mainly comprising magnetite), and gangue mineral (consisting predominantly of 7.49% quartz, 4.34% feldspar, 1.5% talc and 6.22% carbonate, etc). The dissemination of chalcopyrite and pyrrhotite was relatively fine in the ore, whereas pyrite was mainly disseminated in the coarse fraction, and the relationship amongst the dissemination of these three minerals was very complex, so that optimum liberation was seldom achieved, which has significant impact on grinding stage and the sequential floating circuits.
The crushed chalcopyrite ore was subjected to a primary grinding and classification I, wherein 60% of the product size of the primary classification I was −0.074 mm. Then, the product (overflows) obtained from the classification I was subjected to a primary size grading I, wherein the cut-point of size grading I was 0.04 mm. Thereafter, mineral larger than 0.04 mm was subjected to a secondary grinding and classification II, wherein 95% of the product size of the classification II was −0.074 mm. After that, the product (underflows) obtained from the classification II was subjected to a secondary size grading II, wherein the cut-point size of the size grading II was 0.04 mm.
The coarse-grain product which was larger than 0.04 mm produced from the size grading II was subjected to coarse-grain flotation circuits. A one-stage rougher was conducted on the coarse-grain product, in which calcium oxide was used to adjust the pH value. Sodium carboxymethyl cellulose used as depressor, xanthate and pine camphor oil were added to the rougher cell sequentially and the flotation time was 5 min. Then, a first scavenger was conducted on the rougher underflows, wherein calcium oxide was employed to adjust the pH value and the flotation time was the same as the rougher time. Then, the first scavenger concentrate was recycled back to the rougher feed. A second scavenger was conducted on the first scavenger underflows, wherein calcium oxide was employed to adjust the pH value and the flotation time was the same as the rougher time. Then, the second scavenger concentrate was recycled back to the first scavenger feed. Thereafter, a first cleaner was conducted on the rougher concentrate and the flotation time was 3 min. Then, the first cleaner tailings were returned back to the rougher feed, and a second cleaner was conducted on the first cleaner concentrate, wherein the flotation time was 3 min. At last, the second cleaner tailings were recirculated to the first cleaner feed. The second cleaner concentrate was the copper concentrate I, and the second scavenger tailings were the final tailings I.
The fine-grain product which was smaller than 0.04 mm obtained from the size grading I and the size grading II was subjected to fine-grain flotation circuits. A two-stage rougher was conducted on the fine-grain product, in which kerosene used as collector and pine camphor oil used as frother were added to the two-stage rougher cells sequentially. A two-stage scavenger was conducted on the second rougher tailings, in which kerosene used as collector and pine camphor oil used as frother were added to the two-stage scavenger cells sequentially. The first scavenger concentrate was returned to the first rougher feed, and the second scavenger concentrate was recycled to the first scavenger feed. The second scavenger tailings were the final tailings II. The concentrates from two-stage rougher were subjected to a pre-cleaner in the cleaning operation, in the absence of reagents. The pre-cleaner tailings were recirculated to the first scavenger feed. A two-stage cleaning rougher was conducted on the pre-cleaner concentrate, in which calcium oxide was used to adjust the pH value. Sodium carboxymethyl cellulose (CMC) used as depressor, xanthate used as collector and pine camphor oil used as frother were added to the two-stage cleaning rougher cells sequentially. A second cleaning rougher was conducted on the first cleaning rougher tailings, and a two-stage cleaning scavenger was conducted on the second cleaning rougher tailings. The types of reagents used in the two-stage cleaning scavenger are the same as those used in the cleaning rougher cells. A first cleaning scavenger was conducted on the second cleaning rougher tailings. The first cleaning scavenger concentrate was recirculated to the second cleaning rougher feed, and the second cleaning rougher concentrate was recirculated to the first cleaning scavenger feed. The second cleaning scavenger tailings were the final tailings III. A re-cleaner was performed on the concentrates from the two-stage cleaning rougher. The re-cleaner tailings were recirculated to the first cleaning rougher feed. The re-cleaner concentrate was the copper concentrate II.
The reagent amounts and mixing time of each operation are presented in Table 1.
After the above flotation, copper concentrates (combining copper concentrate I and copper concentrate II) grading 21.28% Cu was obtained at recovery and yield of 89.79% and 3.97% respectively.
The operations were similar to those in example 1. The reagent amounts and mixing time of each operation are shown in Table 2 below.
After the flotation, copper concentrates (combining copper concentrate I and copper concentrate II) grading 21.12% Cu was obtained at recovery and yield of 91.3% and 3.89% respectively.
The operations were similar to those in example 1 with the exception that diesel oil was used as collector in the step (3) fine-grain flotation circuits.
The reagent amounts and mixing time of each operation are shown in Table 3 below.
After the flotation, copper concentrates (combining copper concentrate I and copper concentrate II) grading 21.03% Cu was obtained at recovery and yield of 90.42% and 4.12% respectively.
| Filing Document | Filing Date | Country | Kind |
|---|---|---|---|
| PCT/CN2013/087823 | 11/26/2013 | WO | 00 |
