KR101295157B1 - Treating method of cobalt ore - Google Patents

Abstract

PURPOSE: A cobalt concentrate processing method is provided to effectively recover cobalt via a melting mat grinding process and a melting mat leaching process after reduce-melting cobalt concentrate. CONSTITUTION: A cobalt concentrate processing method includes the following steps of: manufacturing slag and an alloy phase by reduce-melting cobalt concentrate; manufacturing a mat by adding sulfur to the alloy phase and melting the same; obtaining ground materials of the mat by grinding the mat; leaching the cobalt and copper from the ground materials; and manufacturing a mineral ball by melting the slag and spraying high-pressured air. [Reference numerals] (AA) Slag; (S100) Cobalt concentrate; (S200) Reduction-melting; (S300) Atomizing; (S400) Mat manufacture; (S500) Mat grinding; (S600) Metal leaching; (S700) Metal returning

Description

Treatment Method of Cobalt Ore

The present invention relates to a method for treating cobalt concentrate that can efficiently recover cobalt from cobalt concentrate.

Cobalt is used in various fields such as secondary battery raw materials, cemented carbide, superalloy, catalyst, magnet, pigment, and the like. However, cobalt is produced only in some countries, such as the Democratic Republic of Congo, and Korea requires more than 5,000 tons of annual imports from overseas, so it is necessary to secure independent raw materials and develop smelting technology.

In the case of cobalt without domestic resources, it is necessary to jump into overseas resource development and compete with foreign advanced companies. However, it is necessary to develop high-efficiency, eco-friendly smelting technology because it is difficult to secure raw materials with existing general-purpose technology. The development of new smelting technologies is needed to enable the development of refractory, low quality cobalt light.

Republic of Korea Patent Publication 2007-0086330 (August 27, 2007 published)

The present invention is to solve the above problems, to provide a method for treating cobalt concentrate that can efficiently recover cobalt from cobalt concentrate.

The object of the present invention is a method for treating cobalt concentrates comprising cobalt, copper and iron, comprising: reducing and melting the cobalt concentrate to obtain an alloy phase and slag; Adding sulfur to the alloy and melting to obtain a mat; Grinding the mat to obtain a pulverized product; It is achieved by a method of treating cobalt concentrate comprising leaching cobalt and copper from the milled product.

Melting the slag and injecting a high pressure air may further comprise the step of producing a mineral ball.

The reduction melting may be performed after adding a flux and a reducing agent to the cobalt concentrate.

The flux may include at least one of alumina and calcium oxide, and the reducing agent may include coke.

The temperature of the reducing melt may be 1400 ℃ to 1550 ℃.

The content of sulfur in the mat may be 12 to 25% by weight.

The content of sulfur in the mat may be 20 to 25% by weight.

The leaching may be performed at a temperature of 120 ° C. to 200 ° C. at an oxygen partial pressure of 12 to 20 atm using a sulfuric acid solution.

In the leaching, the concentration of sulfuric acid solution may be 1.5% by volume to 3% by volume.

In the leaching step, cobalt and copper may be leached over 80%, and iron may be leached below 10%.

An object of the present invention is a method of treating cobalt concentrates comprising cobalt, copper and iron, comprising: obtaining an alloy phase comprising cobalt, copper and iron from cobalt concentrate; Adding sulfur to the alloy and melting to obtain a mat having a sulfur content of 20 to 25% by weight; Leaching the mat using a sulfuric acid solution at a temperature of 120 ° C. to 200 ° C. at an oxygen partial pressure of 12 to 20 atm. In the leaching step, cobalt and copper are leached at least 80% and iron is less than 10%. It can be achieved by a method of treating leached cobalt concentrate.

The concentration of the sulfuric acid solution may be 1.5% by volume to 3% by volume.

According to the present invention, there is provided a method for treating cobalt concentrate that can efficiently recover cobalt from cobalt concentrate.

1 is a flowchart of a cobalt concentrate processing method according to the present invention;
Figure 2 shows the XRD analysis of the cobalt concentrate,
Figure 3 shows the concentrate pellets prepared for reduction melting,
Figure 4 shows the alloy phase obtained through the reduction melting in the arc furnace,
Figure 5 shows the slag obtained through the reduction melting in the arc furnace,
6 shows an alloy phase and slag obtained through reduction melting in an induction furnace,
7 shows a mineral ball obtained by atomizing slag,
8 shows a mat prepared by adding sulfur,
Figure 9 shows the leaching behavior for each metal on the mat 1 in 2% by volume sulfuric acid,
Figure 10 shows the leaching behavior for each metal on the mat 2 in 2% by volume sulfuric acid,
Figure 11 shows the leaching behavior for each metal on the mat 1 in 1% by volume sulfuric acid,
Figure 12 shows the leaching behavior for each metal on the mat 2 in 1% by volume sulfuric acid.

Hereinafter, with reference to the accompanying drawings, a cobalt concentrate processing method according to the present invention will be described in detail.

1 is a flowchart of a cobalt concentrate processing method according to the present invention.

First, a cobalt concentrate is prepared (S100). Cobalt concentrate is available from the Democratic Republic of Congo. Cobalt concentrate contains various metals such as copper, iron, and aluminum in addition to cobalt, and the main valuable metals are cobalt, copper, and iron.

Cobalt in the cobalt concentrate can be 5 to 15% by weight, copper to 10 to 30% by weight, iron to 1 to 7% by weight.

Cobalt concentrate is divided into slag and alloy phase through reduction melting (S200). Reduction melting can be carried out using an arc furnace or induction furnace. When the concentrate powder is directly used, the powder may be lost due to the heat of reduction, so that the concentrate powder may be pelletized (eg, spherical) for reduction melting.

Flux and reducing agent are used for reducing melting in the arc furnace. As the flux, alumina and / or calcium oxide may be used, and a weight of 5% to 20% of the concentrate weight may be used. When both alumina and calcium oxide are used, the ratio of calcium oxide / alumina may be 1.5 to 3. Coke may be used as the reducing agent, and a weight of 5% to 20% of the concentrate weight may be used. The reducing melting temperature in the arc furnace may be 1400 ℃ to 1550 ℃, the reducing melting time may be 30 minutes to 150 minutes or 80 minutes to 100 minutes.

Flux and reducing agent are also used for reducing melting in induction furnaces. As the flux, alumina and / or calcium oxide may be used, and a weight of 10% to 30% of the concentrate weight may be used. When both alumina and calcium oxide are used, the ratio of alumina / calcium oxide may be 0.005 to 0.1. Coke may be used as the reducing agent, and a weight of 30% to 50% of the concentrate weight may be used. The reduction melting temperature in the induction furnace may be 1400 ℃ to 1550 ℃ or 1480 ℃ to 1520 ℃, the reduction melting time may be 20 minutes to 60 minutes.

For slag obtained in the reduced melting through the atomizing (S300) to produce a useful mineral ball. Atomizing refers to a process in which high-pressure air is injected into molten slag, separated into fine droplets, and rapidly cooled with injection air to obtain mineral balls. The melting temperature may be 1200 ° C to 1500 ° C as a temperature at which the slag is sufficiently melted, the nozzle angle (the atomizer angle) may be 25 ° to 35 °, and the air volume may be 50m / s to 80m / s.

The components of the obtained mineral ball may be silica, alumina, calcium oxide, magnesium oxide, or the like, and the silica may be 45 to 55% by weight, and the alumina, calcium oxide, and magnesium oxide may be 10 to 20% by weight, respectively. In addition, the trace amount may further include Fe, MnO, TiO 2, and the like.

Mineral balls may be used as abrasives, waste water refining agents, sound insulation materials, radiation shielding materials, aggregates, flooring materials and the like. Useful mineral balls are obtained from slag, which has previously been a problem with the treatment method.

For the alloy phase obtained in the reduced melting to produce a mat (S400). The mat can be obtained by melting the alloy phase with sulfur and then cooling. Melting may be performed using an induction furnace, the temperature may be 1200 ° C to 1400 ° C, and the melting time may be 10 minutes to 60 minutes. Sulfur may be lost in the mat manufacturing process, so that sulfur may be added in excess of the loss, for example, by adding 10% to 30% of the required amount.

The content of sulfur in the final mat obtained may be 12 to 25% by weight, and may be 20 to 25% by weight. The content of sulfur affects the leaching efficiency.

After that, the obtained mat is pulverized to obtain a pulverized product (S500). Grinding may be carried out using a jaw crusher and theta box, etc., the size of the mill may be less than 140 mesh.

For the pulverized product, the metal is leached to leach valuable metals (S600). Metal leaching is performed using sulfuric acid solution under high pressure and high pressure conditions. The oxygen partial pressure may be 12 to 20 atm, and the temperature may be 120 ° C. to 200 ° C. The sulfuric acid solution may be 0.5% to 3% by volume, and may be 1.5% to 3% by volume. The concentration of sulfuric acid solution affects the leaching efficiency.

The solids ratio in the leaching process may be 0.05 to 0.2 and the leaching time may be 1 to 5 hours.

In the leaching, cobalt and copper in the mill may leach more than 80% or more than 90% and iron may leach less than 10% or less than 5%. Or cobalt may be leached at least 95%, copper at least 50% or 60% and iron at less than 5%.

Finally, the metal is recovered from the leaching solution (S700). The recovery of the metal may be a variety of techniques for separating and recovering cobalt, such as the method of recovering copper first using an extractant.

Hereinafter, the present invention will be described in detail through experimental examples.

Cobalt concentrate

Valuable metal content of cobalt concentrate in the Democratic Republic of Congo was analyzed by ICP analysis, and the results are shown in Table 1. The main valuable metals in the cobalt concentrate in the Democratic Republic of Congo were copper, cobalt and iron, and their contents were 16.1 wt% copper, 7.53 wt% cobalt, and 2.97 wt% iron.

Analysis of Valuable Metals in Cobalt Concentrates of the Democratic Republic of Congo (weight%) ingredient Cu Co Fe Al Ca Cr Mg content 16.1 7.53 2.97 2.51 0.44 <0.004 3.79 ingredient Mn Ni Pb Zn SiO ₂ K S content 0.45 0.01 0.0085 0.074 29.2 0.44 2.66

XRD analysis (FIG. 2) was performed to determine the compound-containing form of copper, cobalt and iron in the concentrate. As a result, it was difficult to identify special crystal forms in the case of cobalt and iron, and in the case of copper, Cu2 (CO3) (OH) 2 (malachite) existed. The peak of (Chlorite) was observed.

Reduction Melting-Sample Preparation

When the powdery concentrate sample is directly melted and reduced, the sample is supported by the reduction heat and loss is generated. Thus, the powder sample was prepared in the form of pellets having a spherical shape as shown in FIG.

Reduction Melting-Arc

In order to observe the reduction melting smelting behavior, reduction melting experiments were performed using an arc furnace for 30 kg concentrate. In order to reduce and melt 30 kg of concentrate sample, 1.2 kg of alumina (Al2O3) and 2.5 kg of calcium oxide (CaO) were mixed with a flux, and 3.1 kg of coke was used as a reducing agent, and the reactor was operated at an input voltage of 380V and an input current of 173A. The reaction temperature was about 1450 ° C.

While performing the arc furnace experiment, the weight change of the alloy phase and slag was measured while changing the holding time after melting (Table 2).

Weight and ratio of alloy phase and slag obtained after arc melting reduction 40 minutes 90 minutes 110 minutes Weight (kg) ratio (%) Weight (kg) ratio (%) Weight (kg) ratio (%) sample 30 30 30 Flux Al ₂ O ₃ 1.2 1.2 1.2 CaO 2.5 2.5 2.5 synthesis 3.7 3.7 3.7 cokes 3.1 3.1 3.1 Alloy 4.52 19.2 6.42 27.3 7.22 33.9 Slag 19.2 80.8 17.7 72.7 14.6 66.1 Weight loss 30.1 30.3 36.8

As shown in Table 2, it was confirmed that the fluidity of the molten metal was increased when the holding time increased during tapping according to the change of the holding time through the arc furnace melting after melting.

4 and 5 show the appearance of the alloy phase and slag produced by the arc melting reduction process and the reduction melting.

Reduction Melting-Induction

The experiment was carried out using the flux diagram of the Cu-Co-Fe composition for the reduction melting experiments in which an induction furnace was concentrated to a metal phase of cobalt concentrate.

Among the components contained in the concentrate, Al2O3, SiO2, CaO, MgO, etc. will be the main targets for generating slag in the melt reduction process. Reduced melting experiments were carried out with a temperature change of 1450 ° C. to 1550 ° C. to form a melting point of slag in the vicinity of 1300 ° C. based on the Al 2 O 3 —SiO 2 —CaO—MgO quaternary state diagram.

The reduced melting experiment was performed while varying the temperature at 1450 ° C. to 1550 ° C., and the alloy phase and slag prepared at 1550 ° C. are shown in FIG. 6. The left side of Fig. 6 is an alloy phase and the right side is slag. Table 5 shows the weight ratios of the alloy phase and slag obtained after reduction melting in induction furnace.

Weight and ratio of alloy phase and slag obtained after reduction melting in induction furnace 1450 ^o C, 30 minutes 1500 ^o C, 30 minutes 1550 ^o C, 30 minutes Weight (g) ratio (%) Weight (g) ratio (%) Weight (g) ratio (%) sample 431 431 431 Flux Al ₂ O ₃ 0.5 0.5 0.5 CaO 68.5 68.5 68.5 synthesis 69 69 69 cokes 175.5 175.5 175.5 Alloy 119 34.4 111.8 32.6 117 34.1 Slag 227 65.6 231 67.4 226 65.9 Weight loss 30.8  31.4 31.4

Atomizing

The existing slag atomizing technology (SAT) process is optimized to process oxidized slag in steelmaking process, and it affects the temperature, viscosity, and fluidity of molten slag, which are important factors for atomizing when the slag composition is different. In terms of the conditions, it is difficult to produce slag balls (mineral balls) of similar quality and yield. Therefore, using the pilot plant manufactured on the same principle as the slag atomizing plant (SAP), the atomizing conditions suitable for the target slag were sought.

At the end of the atomizer, a nozzle was installed, and the size and angle of the atomizer were freely changed so that the nozzle was detachable, and a water jet facility was installed to enable additional injection.

Experiments were carried out using the pilot plant according to the following conditions to reduce molten cobalt concentrate slag, the experimental conditions are shown in Table 4. Airflow conditions were changed by the inverter control, and chemical composition, heavy metal dissolution, and hardness were analyzed for atomized mineral balls.

Atomizing Experiment Conditions Experimental factor level Metrics Melting temperature 1,500 ℃ (Max) ㅇ Recovery rate and particle size distribution
ㅇ Shape
ㅇ Dry density
ㅇ Chemical characterization (oxide)
ㅇ Other use characteristics (abrasive materials) Atomizer angle 30 degrees Air flow 65 m / s

Mineral ball manufacturing experiments using cobalt concentrate slag reduced melt in the arc furnace was prepared 20kg. In order to confirm whether the slag can be atomized, it was heated and melted to about 1,500 ° C., which is a commercial temperature of the melting furnace, and the total melting time was 4 hours.

As a result of atomization using cobalt concentrate slag, a mineral ball having a spherical granular shape as shown in FIG. 7 was prepared. In addition to the spherical particles, some glassy fiber shapes appeared, which is believed to be due to the spherical grains formed by the surface tension of slag scattered during the atomizing process.

Vickers hardness value (HV0.5) of the prepared mineral ball was found to be 684, the chemical composition of the mineral ball is shown in Table 5.

Chemical Composition of Mineral Balls Item unit Slag ball SiO ₂ weight% 53.6 Al ₂ O ₃ 14.3 CaO 12.3 MgO 12.2 Fe 0.84 MnO 1.25 Cr ₂ O ₃ 0.07 S 0.17 TiO ₂ 1.12 P ₂ O ₅ Not detected CuO 0.22 Sum 96.07

A heavy metal dissolution test was performed on the mineral balls. As shown in Table 6, it was found to satisfy the standard content of hazardous substances specified in the Waste Management Act.

Result of harmful ingredient in mineral ball  Test Items unit KS standard Mineral ball Remarks Pb mg / L 3 Not detected standard
satisfied Cu 3 Not detected As 1.5 Not detected Hg 0.005 Not detected CN- One Not detected Cr (VI) 1.5 Not detected CD 0.3 Not detected Oil % 5 0.01

Mat manufacturing

An experiment was conducted to prepare a mat by inputting two kinds of sulfur in an induction furnace for a simulated sample (Co, Cu, Fe). At this time, the temperature of the induction furnace was carried out at 1300 ℃ and maintained for about 20 minutes after melting and then poured into a mold for natural cooling. The weight change was confirmed by comparing the weight of the mock sample and the sulfur with the weight of the manufactured mat, and 15% of the required amount was added in consideration of the loss of sulfur. Table 7 shows the contents of the simulated sample and sulfur added during the manufacture of the mat.

Amount of simulated sample added during mat manufacturing (g) number Cu Co Fe S Sum One 175.4 138.7 39.3 162.7 516.1 2 174.9 138.8 38.8 128.4 480.9

The state of the mat manufactured in Figure 8 is shown.

The weight of the mat produced after the production was measured and the results of the weight change after the injection of sulfur and the sample are shown in Table 8.

Investigation of sulfur content by weight change after mat manufacturing number Simulated sample weight
(g) Sulfur input
(g) Mat weight
(g) S amount lost when sulfur is added (%) One 353.4 162.7 454.5 37.8 2 352.5 128.4 451.0 23.3

As shown in Table 8, the reason that the amount of loss is irregular in the input of sulfur is that it is difficult to add and control the desired amount of sulfur according to the input interval, the input method, and the input speed.

As a result of performing ICP analysis on the prepared mat, the main elements Cu, Co, and Fe were 32.5 to 37.8% by weight, 28.3 to 37.1% by weight, and 8.85 to 8.99% by weight, as shown in Table 9, respectively. It was found that sulfur contained 22.5% by weight and 16.0% by weight of each mat (since 22.5% of sulfur containing mat 1 'mat 1', 16.0% containing mat 'Matte 2' ).

Component analysis results by weight (% by weight) Mat 1 Mat 2 Cu 37.8 32.5 Co 28.3 37.1 Fe 8.85 8.99 Ni 0.024 0.40 Al 0.52 0.54 Ca 0.16 0.12 Cr <0.004 <0.004 Mg 0.029 0.024 Mn 0.0076 <0.004 Pb 0.021 0.0048 Zn <0.004 <0.004 K 0.03 0.023 Na 0.087 0.052 SiO ₂ 1.60 3.52 S 22.5 16.0

Crushing

For the mat prepared, primary crushing was performed with a jaw crusher and secondary crushing was performed using a theta box. After that, the particles were separated into particles of 140 mesh or less.

Metal leaching

The high temperature and high pressure leaching was carried out on the sulfuric acid base, and the leaching behavior of the mat was examined with varying the sulfuric acid concentration to 1% by volume and 2% by volume. At this time, the solid-liquid ratio was 0.1 (mat 70g, 700 ml of solution), atmospheric pressure (PO2) was 15 atm, stirring was 700 rpm, reaction time 2 hours, and temperature was 150 degreeC. In addition, 5 to 10 ml of the sample was taken every 15, 30, 60, 90, and 120 minutes, and the components were analyzed using an ICP analyzer.

As a result of performing high temperature and high pressure leaching in sulfuric acid 2% by volume, it was found that in the mat 1, cobalt and copper were leached 100% from the reaction time of 15 minutes while iron was almost not leached. Such leaching behavior is shown in FIG. 9.

As shown in FIG. 10, in the case of the high-temperature and high-pressure leaching of 2 vol% sulfuric acid, the leaching rate of cobalt was about 95% at 15 minutes of reaction time, and the leaching rate gradually increased with increasing time, and then the reaction time was 120 minutes. While leaching at 100%, copper was leached about 20% at 15 minutes of reaction time, and the leaching rate remained at about 60% even at 120 minutes of reaction time. At this time, Fe was hardly leached.

As a result of the above experiments in which the mat 1 has a better leaching rate than the mat 2, it can be seen that the sulfur content in the mat has a major influence on the high temperature and high pressure leaching of the mat. A minimum reaction time of 30 minutes at volume% led to the conclusion that 100% cobalt and copper were leached. The concentrations of cobalt, copper and iron in the solution were 33,900 ppm, 37,100 ppm and 73.8 ppm, respectively. Table 10 shows the concentration of each metal that varies with sampling time.

Metal leaching by ppm 1 reaction time in 2% by volume sulfuric acid (ppm) cobalt Copper iron 15 minutes 24,500 25,800 160 30 minutes 33,900 37,100 73.8 60 minutes 34,700 39,700 43.7 90 minutes 32,900 38,300 38.3 120 minutes 33,400 39,500 42.7

The high temperature and high pressure leaching of the mat 1 in 1% by volume of sulfuric acid was performed and the results are shown in FIG. 11. In the case of Mat 1, cobalt had a leaching rate of 100% from 15 minutes of reaction time, while copper had a low leaching rate of about 16% at 15 minutes of reaction time and about 45% at 120 minutes of reaction time.

As shown in FIG. 12, in the case of Matt 2 at 1% by volume of sulfuric acid, cobalt showed a leaching rate of about 77% to 87% according to the reaction time, and copper showed a leaching rate of about 5% until the reaction time of 30 minutes. From the reaction time of 60 minutes, the leaching rate increased rapidly, and the leaching rate of about 35% was obtained at the reaction time of 120 minutes.

The leaching rate of mat 1 was better than that of mat 2 even under 1% by volume of sulfuric acid, and it was confirmed that the sulfur content in the mat had a major influence on the high temperature and high pressure leaching of the mat. In addition, in the case of 1% by volume sulfuric acid, the leaching rate of cobalt and copper is significantly lower than that of 2% by volume sulfuric acid. It is considered an efficient leaching process.

Metal recovery

The recovery of the metal may be a variety of techniques for separating and recovering cobalt only, such as recovering the copper first using an extractant.

While the present invention has been particularly shown and described with reference to exemplary embodiments thereof, it is clearly understood that the same is by way of illustration and example only and is not to be taken by way of limitation and that those skilled in the art will recognize that various modifications and equivalent arrangements may be made therein. It will be possible. Accordingly, the true scope of protection of the present invention should be determined only by the appended claims.

Claims (12)

In the method of processing cobalt concentrate containing cobalt, copper and iron,
Reducing melting the cobalt concentrate to obtain an alloy phase and slag;
Adding sulfur to the alloy and melting to obtain a mat;
Grinding the mat to obtain a pulverized product;
A method of treating cobalt concentrate comprising leaching cobalt and copper from the pulverized product.
The method of claim 1,
Melting the slag and injecting high-pressure air to produce a mineral ball, characterized in that it further comprises the step of producing a mineral ball.
The method of claim 1,
The reduction melting,
The cobalt concentrate processing method characterized in that it is carried out after adding a flux and a reducing agent to the cobalt concentrate.
The method of claim 3,
The flux comprises at least one of alumina and calcium oxide,
The reducing agent is cobalt concentrate treatment method comprising the coke.
The method of claim 1,
The reduced melting temperature is 1400 ℃ to 1550 ℃ processing method of cobalt concentrate. The method of claim 1,
The content of sulfur in the mat is cobalt concentrate treatment method, characterized in that 12 to 25% by weight.
The method of claim 1,
The content of sulfur in the mat is cobalt concentrate processing method, characterized in that 20 to 25% by weight.
The method of claim 1,
The leaching is,
A method of treating cobalt concentrate, wherein the sulfuric acid solution is carried out at a temperature of 120 ° C. to 200 ° C. at an oxygen partial pressure of 12 to 20 atm.
9. The method of claim 8,
In the leaching,
The concentration of the sulfuric acid solution is 1.5 vol% to 3 vol% treatment method of the cobalt concentrate.
The method of claim 1,
In the leaching step cobalt and copper is leached more than 80%, iron is less than 10% leaching method of cobalt concentrate.
In the method of processing cobalt concentrate containing cobalt, copper and iron,
Obtaining an alloy phase comprising cobalt, copper and iron from the cobalt concentrate;
Adding sulfur to the alloy and melting to obtain a mat having a sulfur content of 20 to 25% by weight;
Leaching the mat with a sulfuric acid solution at a temperature of 120 ° C. to 200 ° C. at an oxygen partial pressure of 12 to 20 atm,
In the leaching step cobalt and copper is leached more than 80%, iron is less than 10% cobalt concentrate treatment method.
12. The method of claim 11,
The concentration of the sulfuric acid solution is a cobalt concentrate treatment method, characterized in that 1.5 to 3% by volume.

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