LATERITE ORES

Technical Field

The invention relates to the field of non-ferrous metallurgy, in particular to a method for recovery of nickel and cobalt from laterite ores.

Background Art

About 70% of the world nickel reserves are concentrated in laterite ores, which, according to their chemical and mineral composition, are divided into ferruginous (limonite) and magnesium silicate (saproiite) ores. The fraction of ferruginous ores accounts for about 70% of the nickel and of saproiite ores about 30%.

Magnesium silicate nickel ores are processed by pyrometallurgical methods, and ferruginous ores by hydrometallurgical methods. Hydrometallurgical processing of laterite ores is carried out in two variants (Reznik I.D., Emiakov G.P., Shneerson Ya.M.“Nickel (Current state of hydrometallurgy of oxidized nickel ores)”. - Moscow: Science and Technology, 2001, vol.2, pp. 364-464); Dr.Sc. Ashok D. Dalvi; Dr. W. Gordon Bacon; Mr. Robert C. Osborne.“The Past and the Future of Nickel Laterites”// PDAC 2004 International Convention, Trade Show & Investors Exchange. March 7-10, 2004, Inco Limited, 2060 Flavelle Boulevard, Sheridan Park, Mississauga, Ontario, L5K 1Z9 Canada]: sulfuric acid high pressure leaching at high temperatures (HPAL process) and ammonium-carbonate leaching after preliminary reduction roasting at 650-780°C (Caron process). Other sulfuric acid technologies have been developed and proposed: atmospheric, heap leaching and sulfation of ore with subsequent aqueous leaching. In known sulfuric acid technologies, the recovery of Ni and Co is not selective, during leaching occurs almost complete decomposition of ferruginous ore minerals, which requires a large consumption of acid (up to 500-600 kg per 1 ton of ore and more). During neutralization this leads to the formation of a large volume of ferruginous gypsum wastes (up to 2 tons/ ton of ore), which creates a serious environmental burden. Carrying out the process in autoclaves allows to reduce the consumption of acid somewhat (up to 230-300 kg/t), however, along with the use of expensive equipment, other problems arise which adversely affect the technical and economic parameters of the process. For example, the increased content of magnesium in the ore leads to a significant increase in the consumption of sulfuric acid.

Of the above hydrometallurgical technologies, the most interesting is the ammonium- carbonate technology, which is based on the principle of selective recovery of Ni and Co from the ore. The technology includes the following main steps: drying the ore, grinding, reduction of the grinded ore under conditions that ensure selective reduction of nickel and cobalt, cooling the reduced ore in an inert atmosphere, and leaching the ore in an ammonium- carbonate solution by air aeration with transfer of nickel and cobalt into a solution, and iron into an insoluble constituent. The preliminary reduction roasting of the ore ensures selectivity of the dissolution of nickel and cobalt during the leaching of the cinder. The cobalt and nickel are then recovered from the solution using a variety of methods. The process provides for the regeneration of the main ammonium reagent and, in part, carbon dioxide.

A reduction roasting in industrial conditions is carried out in the Herreshoff multiple- hearth furnaces in a temperature range of 650-750°C. A reducing atmosphere in the furnace is created by sub-stoichiometric combustion of fuel. At a temperature below 650°C, the rate of reduction noticeably decreases, and the temperature increasing to 800°C and higher leads to formation of forsterite, which adversely affects the results of nickel recovery during leaching. To ensure the selectivity of Ni and Co reduction, the process is carried out with a minimum degree of reduction of iron in its oxides to a metal state, with its content in the reduced ore in the range 2.0-2.5%. The increased content of metallic iron leads to a deterioration in the recovery of nickel and cobalt in the subsequent ammonium carbonate leaching of the cinder. To intensify the process of reducing nickel and cobalt at low temperatures (up to 650°C), before grinding, the ground ore is mixed with heavy oil, in particular with fuel oil, in an amount of about 2.5% of the ore mass.

After reduction roasting, the cinder is cooled to 150-250°C in an inert atmosphere in water refrigerators (multi-disk type), then the cinder from the refrigerator is unloaded into a chute, into which an ammonium solution containing 65-85 g/1 of N¾ and 35-45 g/1 of C0 ₂ is fed. The leaching of the pulp in the ammonium-carbonate solution is carried out by air aeration at the temperature of about 40°C and liquid-to-solid (L:S) ratio = (6.0-6.5):l. The total leaching time is 2.0-2.5 hours.

During ammonium-carbonate leaching with air supply, along with nickel and cobalt, dissolution of metallic iron occurs. Iron in ammonium solutions forms metastable compounds, which, with an oxidizer, easily decompose with the release of iron hydroxide Fe(OH) ₃ . This is accompanied by co-precipitation of nickel and especially cobalt, which leads to a significant reduction in the degree of their recovery. The presence of manganese and sulfur increases the loss of cobalt with a solid residue. In practice, the recovery rate of nickel is 60-85%, and cobalt - 20-40%, in rare cases reaches 60%. It should be noted that such low rates for the recovery of metals in ammonium-carbonate technology are associated with the characteristics of the leaching process.

After leaching, the nickel- and cobalt-containing solution is separated from the ore tailings and sent to recover nickel and cobalt by known methods, in particular by distillation process or in its combination with other methods such as liquid extraction, ion exchange and precipitation. The distillation process is carried out in conventional distillation multi-sectional columns by heating the solution with an acute steam to 100-1 l0°C. At the same time, the sublimable ammonium and carbon dioxide are absorbed by water and returned to the leaching process as a working solution. The process is simple to operate, but it is very sensitive to changes in the basic leaching parameters (temperature, ammonium concentration in the solution, etc.), requires high energy costs and is environmentally unsafe.

An increase in the recovery of nickel and cobalt is possible if the following conditions are met: increasing the leaching temperature; increasing of concentrations of N¾ and C0 ₂ in the ammonium-carbonate solution, increasing of the ratio of the liquid phase to the solid phase, i.e. increasing of the specific volume of the ammonium-carbonate solution with respect to the roasted ore; exclusion or minimization of sulfur content in the process scheme.

However, an increase in the leaching temperature from 40-45°C leads to an increase in the partial pressure of ammonium in the system, resulting in increased losses. To reduce ammonium losses, the concentration of ammonium in the solution is usually maintained at TOSS g/1, and the NH3/CO2 mass ratio is 1.7-2.

The decrease the ratio L:S below 6:1, during the leaching, complicates the conditions for maintaining the thermal conditions of the process, and also leads to a loss of cobalt.

The presence of sulfur in the fuel and reducing agent (fuel oil) leads to loss of metals in the form of sulfides in the solid phase (ore tailings) and to increase in ammonium consumption due to the formation of ammonium sulfate.

Thus, the set of the above-mentioned disadvantages does not offer the possibility of significant improving of parameters of ammonium-carbonate technology.

There are known works on the improvement of ammonium-carbonate technology. However, all these works are mainly devoted to the improvement of the stage of preliminary reduction roasting of laterite ores. Considering that in the ammonium-carbonate technology the low recovery of nickel and cobalt is more closely related to the characteristics of the leaching process in ammonium solutions, these known methods do not eliminate the main disadvantages inherent in this technology. In patent GB 1348031, various roasting furnaces (a fluidized bed furnace, a multiple- hearth furnace and a rotary kiln by indirect or direct heating) are offered based on laboratory tests to effect the reduction roasting of laterite ore, with the addition of a petroleum fuel reductant of 3 to 6%, more preferably 4.5% by dry weight of ore. With such expenses of petroleum fuel, the economic efficiency of the method is significantly reduced or becomes unprofitable.

US Pat. No. 3,656,934 proposes a selective reduction of limonite ore containing from 38 to 46% Fe total and approximately 1.5% Ni in a rotary furnace, followed by leaching of the reduced ore with an ammonium-carbonate solution. Reduction roasting of the ore is carried out with additives of up to 5% coal and 1% pyrite (sulfur-containing additive) when a gaseous reducing agent is supplied in several places along the furnace. During leaching is achieved an 80% recovery of nickel.

The closest in terms of technical essence is the method of recovery of nickel and cobalt from laterite ores (patent RU 2333972). According to this method, the dried and ground laterite ore consisting of either limonite ore or saprolite ore, or a mixture of these two ores is fed to a rotary furnace which is heated by sub-stoichiometric combustion of fuel (oil fuel or gas) to create a reducing atmosphere in the furnace. The reducing atmosphere in the rotary kiln is maintained by controlling the hydrogen and carbon monoxide content in the gas mixture. The temperature in the reduction zone of the roasting furnace can be varied within the range of 600-850°C, preferably 700-8 l0°C. The total residence time of the ore in the furnace can vary from 65 to 260 minutes, depending on the rotational speed of the furnace, and the residence time in the reduction zone at a temperature above 600°C can vary from 13 to 52 minutes. It is allowed to use up to 2.5% (by weight of ore) reducing agent, in particular heavy oil, as additives in grinded ore to intensify the reduction of nickel and cobalt at lower temperatures (600-650°C).

In the method, it is noted that when using a rotary furnace for reduction roasting, a significant cost reduction can be achieved, since a single rotary kiln can replace 12 or more multiple-hearth furnaces requiring large repair and maintenance costs and high energy consumption due to the high expense of expensive fuel, which accounts for the bulk of the technological costs.

After roasting, the reduced ore is cooled to a temperature of 150 to 300°C in a nonoxidizing atmosphere, then passivated with an ammonium-carbonate solution. To recover nickel and cobalt in solution, the reduced ore with air aeration is leached in an ammonium- carbonate solution containing 70-150 g/l of NH ₃ and 50-100 g/l of C0 ₂ , at atmospheric pressure and temperatures from 35 to 60°C. However, in the examples, for leaching was used a solution only with a concentration of 100 g/l of NH ₃ and 80 g/l of CO2, and the leaching temperature was not indicated. After leaching, the solution is separated from the ore tailings and sent to the recovery of nickel and cobalt using known methods. It was noted that in the optimal conditions for the recovery of limonite ore and ammonium-carbonate leaching, the degree of nickel recovery can reach 92-93%, and cobalt - 66-69%.

Although the implementation of the process of reduction roasting of ore in a rotary kiln contributes to some improvement in the technical and economic parameters of the process, in particular to an increase in the energy efficiency of roasting, but the main shortcomings inherent in ammonium carbonate technology persist.

So to achieve a high degree of recovery of nickel and cobalt, the reduced ore is leached in an ammonium-carbonate solution with a high concentration of NH ₃ (100 g/l) in the temperature range 35-60°C at atmospheric pressure. As noted above, with ammonium- carbonate leaching with air aeration at atmospheric pressure, an increase in the ammonium concentration above 85 g/l is considered impractical, due to a significant increase in ammonium loss even at relatively low temperatures of 30-40°C. To limit the loss of ammonium, especially at elevated temperatures (50-60°C), it is required to carry out the process under excess air pressure (2-4 atm.) in an autoclave.

In addition, in this method there are no data on the ratio of liquid to solid phase (L:S) in leaching. At the same time, it is known that an increase in the L:S ratio significantly improves the recovery of nickel and cobalt, but is considered economically unprofitable due to a significant increase in material flows in the technological cycle and energy costs when distilling the solution for stripping of ammonium and carbon dioxide.

In a known method, as well as in other analogues related to the ammonium-carbonate technology of recovery of nickel and cobalt from laterite ores, another disadvantage is the sensitivity of the technology to sulfur, the presence of which in the cinder leads to a reduction in the recovery of cobalt, as well as loss of ammonium due to formation of ammonium sulfate (NH ₄ ) ₂ S0 ₄ during leaching.

The combination of the above factors substantially increases the cost of the recovery of nickel and cobalt by a known method, and also increases the environmental burden. In this regard, in order to increase the efficiency of recovery of nickel and cobalt from ferruginous laterite ores, especially in the leaching stage, other approaches are required for hydrometallurgical treatment of the pre- reduced ore.

Disclosure of invention

Technical problem

The invention solves the problem of providing with a method for recovering nickel and cobalt from laterite ores with high technical and economic parameters (high recovery rate, reducing reagent consumption, environmental safety) by replacing the ammonium-carbonate leaching technology with the sulfuric acid leaching of the reduced ore.

Solution of the problem

The inventors carried out numerous diverse studies and found out that nickel and cobalt can be leached with sulfuric acid from laterite ore cinder produced after reduction roasting of the ore. To this end, in a method including ore grinding, roasting of ore in a reducing atmosphere for the selective reduction of nickel and cobalt, cooling the reduced ore in a nonoxidizing atmosphere to a temperature of l50-300°C, and leaching the cooled cinder with transferring nickel and cobalt into the solution, the roasting the ore in a reducing atmosphere is carried out in the temperature range 700-800°C, preferably at 725-775°C, and the cooled cinder is leached with sulfuric acid in a weakly acidic medium at a pH of 1.4-3.5, preferably at a pH of 2.5-3.0, at a temperature of 30-l00°C, preferably at 85-95°C for 45-120 minutes, preferably 60-90 minutes.

Before the roasting, the ore is preferably mixed with a reducing agent, for which is used brown (subbituminous) coal in an amount of 3-3.5% of the mass of the ore.

In addition, petroleum fuel, such as fuel oil, can be used as a reducing agent.

The ore roasting process is carried out in a tubular rotary furnace or in a vertical multiple-hearth furnace.

For selective reduction of nickel and cobalt during roasting, a reducing atmosphere can be created by controlling the H ₂ and CO content in the gas atmosphere in the furnace.

Effect of the invention

In accordance with the invention, a method has been developed for recovering nickel and cobalt from ferruginous laterite ores with a high degree of recovery of nickel and cobalt, a reduced consumption of reagents, providing economic and environmental safety.

Description of Embodiments

The essence of the proposed method is as follows.

Reduction roasting of laterite ore is a well-known process in the industry.

Roasting of laterite ore in reducing atmosphere (reduction roasting) at lower temperatures of 700-800°C practically eliminates formation of the phase of wustite FeO, (Fe ₂ 0 ₃ is reduced to Fe30 ₄ ), which is readily soluble in weakly acidic solutions of the phase, and this process substantially limits the consumption of sulfuric acid in the leaching. For example, if during the leaching of ore recovered at a temperature of 750°C, with 85-99% recovery of nickel, the specific consumption of sulfuric acid is 90-130 kg per 1 ton of ore at pH 2.5-3.0, then at a temperature of 850°C and 900°C consumption of acid increases to 180 kg and 215 kg, respectively. With reduction in the roasting temperature below 725°C, the reduction rate is markedly reduced, and in order to achieve the required degree of reduction of nickel and cobalt, it is necessary to increase the duration of the process, which is economically unprofitable due to an increase in energy costs.

In the case of using brown coal, the step of mixing ore with a reducing agent before roasting is simplified. The use of 3-3.5% of brown coal additives during reduction roasting of laterite ore significantly facilitates the regulation of the necessary reducing atmosphere in the reduction zone in a rotary kiln over a wide temperature range from 400 to 800°C. Reducing atmosphere is created by controlling the content of H ₂ and CO, which are produced during the combustion of fuel (methane) in the gas atmosphere of the furnace. Carrying out roasting without reducing agent additives requires creating conditions for additional supply of the gas reducing agent to the low-temperature (400-650°C) zone of the rotary kiln, which significantly complicates its design.

The total residence time of the ore in the reduction zone of the rotary kiln, where the temperature is above 400°C, can be from 60 to 120 minutes, and above 600°C from 30 to 60 minutes, depending on the speed of rotation of the kiln, the types and composition of laterite ores used and the roasting temperature.

Reduction roasting of laterite ore can be carried out in multiple-hearth furnaces. Although these furnaces do not always ensure sufficient conversion efficiency of nickel and cobalt due to poor contact of gas and solids, internal“breakthrough” of solids and reducing gas, and also due to poor temperature control, it is easier to regulate the reducing atmosphere in multiple-hearth furnaces than in rotating kilns.

Leaching after cooling of laterite ore cinder, reduced under optimal conditions, with sulfuric acid in a weakly acidic medium at pH of 1.5-3.0 permits, firstly, to achieve a high degree of selective extraction of nickel (up to 87-99%) and cobalt (up to 75 -80%) to the solution, and secondly, to significantly improve the conditions of the hydrometallurgical processing: carry out the leaching at sufficiently low ratios of L:S (2.5-3): 1 and without air aeration, remove the hard restriction of the upper limit of the leaching temperature and expand it temperature range from 30-40 to 30-l00°C, to simplify the separation of leached solution from ore tailings and the operations of separation of metals from the solution, and, at the same time, to minimize the environmental load due to elimination of ammonium from the process cycle. In addition, the proposed method eliminates the negative effect of sulfur in the reduced cinder on the recovery of cobalt and other parameters of the leaching process, i.e. there are no special requirements for the reducing agent and fuel for the sulfur content in them.

The advantage of using weak solutions of sulfuric acid for leaching of the reduced laterite ore cinder is due to the following main factors. First, unlike ammonium or ammonium- carbonate solutions, diluted solutions of sulfuric acid are practical for leaching in a wide temperature range from 30 to 100°C due to the absence of the release of aggressive and environmentally harmful gases, and second, in sulfuric acid solutions, iron, nickel and cobalt ions are more stable than in ammonium-carbonate solutions, which allows leaching to be carried out at essentially low ratios of L:S = (2.5-3): 1 (2-2.5 times less than during ammonium-carbonate leaching in the known analogues) without loss of nickel and cobalt. Therefore, in the method of the invention, when leaching the reduced ore in the weak sulfuric acid solutions under the aforementioned conditions, it becomes possible to achieve high metal recovery, up to 87-99% of Ni and up to 75-80% of Co, and also to reduce by 2-2.5 times the material flows in comparison with analogues using ammonium carbonate leaching.

In the case of sulfuric acid leaching of the reduced laterite ore at pH of 1.5-3.0, simultaneously with the dissolution of metals (nickel and cobalt), a small part of the iron oxide-magnetite is dissolved by reaction with formation of sulfate salts of Fe (II) and Fe (III):

Fe ₃ 0 ₄ + 4H ₂ S0 ₄ = FeS0 ₄ + Fe ₂ (S0 ) ₃ + 4H ₂ 0

The appearance in the solution of Fe ³⁺ ions substantially accelerates dissolution of metallic nickel and cobalt during leaching of the reduced ore in weakly acidic solutions according to the reactions:

Ni + Fe ₂ (S0 ) ₃ = NiS0 + 2FeS0 ₄

Co + Fe ₂ (S0 ₄ ) ₃ = CoS0 ₄ + 2FeS0 ₄

As a result, there is no need for aeration of the slurry with air, which is mandatory in the known method and other analogues based on ammonium-carbonate leaching.

The amount of iron that passes into the solution essentially depends on pH and leaching temperature, it increases with increasing pH and increasing temperature. Therefore, to limit the dissolution of magnetite, leaching is carried out with a certain combination of these parameters. It was found that when the pH is raised from 1.5 to 2.5-3.0, the optimum leaching temperature is from 30 to 85-95°C. A decrease in pH below 1.5 results in an increase in the amount of dissolved iron in the solution and in the consumption of sulfuric acid. At pH above 3.0, hydrolysis of Fe ₂ (S0 ₄ ) ₃ in the solution can occur with the release of iron hydroxide, which is accompanied by co-precipitation of nickel and cobalt from the solution. In the proposed method leaching at higher temperatures (about 90°C) is more preferable. This choice is related to the conditions for passivation of the reduced ore by a recycled solution and with the specialty of the leaching with sulfuric acid. First, during the passivation of the cinder cooled to l50-300°C by recycled solution, especially at low ratios L:S=(2.5-3):l, the temperature of the solution is substantially increased. Second, a noticeable increase in the temperature of the solution occurs during leaching of cinder in recycled solution with constant addition of sulfuric acid in automatic mode to maintain the pH at a certain level. In industrial conditions it is more efficient and expedient to use concentrated sulfuric acid, which inevitably leads to a substantial increase in the temperature of the solution. Third, the reactions of dissolution of metals and iron oxide in solutions of sulfuric acid are also exothermic, i.e. occurs with the release of heat. Given these circumstances, it is more preferable to leach at a temperature of about 90°C at elevated pH, preferably at pH of 2.5-3.0, to limit the dissolution of the iron oxide and reduce the acid consumption. At lower pH values (about 1.5), constant slurry cooling is required, which is not advisable.

As for the ratio of L:S in leaching, the most preferable ratio is in the range (2.5-3): 1. At L:S above 3, the results on recovery of metals practically do not change, but the product flows increase, which is considered undesirable in industrial conditions, and at L:S below 2.5: 1, the concentration of iron in the solution increases, which may promote hydrolysis of ferric sulfate under leaching conditions, particularly at an elevated pH (at pH-3).

In the method of the invention, in order to achieve the maximum recovery of nickel and cobalt, the sulfuric acid leaching of the reduced ferruginous laterite ore at specified pH values is carried out for 45-120 minutes, preferably 60-90 minutes. Decreasing of the leaching time, although it allows a little decrease in the consumption of sulfuric acid, does not always ensure the maximum degree of recovery of metals into the solution. Increasing the leaching time for more than 90 minutes does not lead to a noticeable improvement in the recovery results, but the consumption of acid slightly increases, which is not very acceptable.

The method of recovery of nickel and cobalt can be used for both ferruginous (limonite) and mixed (mixture of ferruginous and magnesium-silicate) laterite ores. However, the presence of silicate ore in the mixture leads to an increase in the specific consumption of acid during leaching due to transition of magnesium to solution. For example, under the same conditions of roasting and leaching, if using limonite ore, specific consumption of sulfuric acid at pH 3 does not exceed 100 kg per ton of ore, in the case of magnesium-silicate ore, the acid consumption reaches 250 kg/ton. In this regard, it is preferable to use laterite ores with a limited content (no more than 30%) of silicate ore.

Various types of laterite ores (ferruginous and silicate) were used to develop the proposed method. The chemical compositions of these ores are given in Table 1. Before reduction roasting, all ore samples were ground to the size of 100 pm. Brown coal (volatile components 51%, ash content 6.3%) was used as a reducing agent additive. Reduction roasting of ore with additives of 3.5% coal was carried out in a laboratory rotary kiln in a continuous mode using a gas reducing agent.

Table 1

Chemical composition of laterite nickel ore samples

*i.l. - Loss on ignition

An embodiment of the invention is given below, not excluding other embodiments within the claims.

Taking into account that the conditions for reduction roasting of the ore (the type of reducing agent, the conditions for creating a reducing atmosphere, the implementation of the process in rotary kilns and in multiple-hearth furnaces, etc.) are well known in the art, examples of the invention have been carried out for various sulfuric acid leaching conditions for cinder of reduction roasting of laterite ores.

EXAMPLE

3.5% of brown coal is added to the laterite ore, thoroughly mixed, the mixture is subjected to reduction roasting at a temperature of 725-775°C, the reduced ores are cooled in an inert atmosphere (without air access) to a temperature of 250°C and then passivated (cooled) in cold water. After passivation, the cinder is leached in a weakly acidic medium at pH 2.5-3.0, temperature 90-95°C, L:S = (2.5-3.0): 1 for 60-90 min, with the supply of a dosed amount of sulfuric acid in automatic mode.

Table 2 shows examples of the implementation of the method with various process parameters. Examples with optimal recovery rates of nickel and cobalt are shown in bold in the table.

Table 2

Results of testwork on sulfuric acid leaching of the products of the reduction roasting of laterite nickel ore

The use of the inventive method for recovery of nickel and cobalt from ferruginous laterite ores offers a means to: - ensure a high degree of recovery of nickel (up to 99%) and cobalt (up to 83%) from laterite ores;

- reduce product flows in the hydrometallurgical conversion by 2-2.5 times;

- reduce the leaching time for a reduced ore to 1-1.5 hours (in the ammonium-carbonate technology, the leaching period is usually 2-2.5 hours);

- minimize the environmental pressure due to the elimination of ammonium from the technological cycle.

Industrial Applicability

The invention can be applied to the processing of ferruginous laterite ores and their mixture with magnesium silicate laterite (saprolite) ores of various deposits for the rational use of natural resources.