KR20070053211A - Method for nikel and cobalt recovery from laterite ores by reaction with concentrated acid water leaching - Google Patents

Abstract

Two-step process for leaching laterite ore, including limonite and saprolite, the first step consists of mixing and reacting the ore with a concentrated inorganic acid and the second step preparing a slurry of acid / ore mixture in water And leaching the mixture to dissolve nickel and cobalt. Iron is effectively separated from nickel and cobalt in solid leach residues, mainly in the form of oxides or hydroxides of iron ions (II) rather than iron alum.

Description

METHODS FOR NIKEL AND COBALT RECOVERY FROM LATERITE ORES BY REACTION WITH CONCENTRATED ACID WATER LEACHING}

Cross Reference of Related Application

This application claims the benefit of US Provisional Application No. 60/583243, filed June 28, 2004, the disclosure of which is incorporated herein by reference herein.

FIELD OF THE INVENTION

FIELD OF THE INVENTION The present invention relates to the wet smelting of nickel-ferrous laterite ores and in particular to the acid leaching of both the brown iron and the saprolite portions by a single process.

Laterite ores are formed by ultramafic rocks containing nickel near the earth's surface or near the surface of the tropics by acidic meteoric water, which is inherently acidic over the geological age. It consists of various mud, oxide and silicate minerals, nickel and / or cobalt-rich portions, thereby distinguishing them from the majority of other nickel and sulfide mineral species. The latter typically consists of the sulfides of iron, nickel and cobalt, often copper and a small amount of precious metals, and are associated with the mafic-ultramafic magnetic intrusion in the earth's crust. .

Weathering processes typically produce layered deposits of weathering processes that occur near the entire or very broad surface, with deeper layers becoming the result of weathering to a lesser extent and ultimately un weathered at any deeper depth. Ends with rocks The highly weathered layer mainly contains most of the nickel contained in the microscopically distributed finely divided goethite. Goethite is an oxyhydroxide of iron ion (II) having the formula FeOOH. This layer was named limonite, which typically contains a large amount of iron.

Cobalt is generally associated with a layer of iron ore and is also commonly referred to as absolane or manganese wad, oxidized manganese minerals (Mn (III) and / or Mn (IV including oxides and hydroxides). Significantly related to)).

The less weathered layer (s) contain an increasing proportion of nickel contained in various magnesium silicate minerals, such as, for example, serpentine. The serpentine is a silicate mineral of magnesium with the formula 3 MgO * 2SiO ₂ * 2H ₂ O. It is believed that nickel replaces some magnesium in the serpentine. Magnesium will also be substituted with other kinds of divalent metals, for example iron ions (Fe ²⁺ ). There may be many other silicate minerals containing nickel in incompletely weathered areas. The partially weathered, high magnesium-containing zone is often referred to as saprolite or garnierite. ("Gyu-nickel stone" is also used to describe the magnesium-nickel silicate mineral in its characteristic, apple-green, yellow-green, variable composition.)

Some sediments have other zones, typically containing nontronite clay, typically located between amonite and saprolite, which may be magnesium, iron and aluminum, and also nickel-iron (nikelferous). Magnesium silicate of those that are. Most of the sediments located on the (prefecture) return line lack mainly nontronite zones.

It should also be noted that the weathered zone (s) is not uniform in mineralogy or chemical composition, and the boundaries between the zones are not parallel to the earth's surface. However, there is a fairly pronounced transition from much iron and relatively low magnesium containing ores, which generally occurs over a distance of 1 to 3 m vertically in laterite deposits, to minerals with a relatively high magnesium content and a changeable but lower iron content. have.

For illustrative purposes only, the chemical composition of typical Limonite and Saprolite is as follows.

Limonite: Ni 1.0-1.8%, Co 0.05-0.3%, Fe 35-50%, Mg 0.2-3.5%

Saprolite: 1.2 to 3.5% Ni, 0.02 to 0.07% Co, 7 to 20% Fe, 10 to 20% Mg

Each zone also typically contains significant concentrations of aluminum, manganese, and chromium, as well as trace heavy concentrations of other heavy metals such as copper and zinc in various other minerals.

The problem to be solved in the recovery of nickel from laterite ores is that the nickel values are generally almost concentrated by a physical means called the ore dressing technique before the chemical process to separate the metals. It can't be. This makes laterite treatment expensive, and means have been sought for decades to lower the cost of laterite treatment.

In addition, due to the completely different mineralogy and chemical composition of the ore and saprolite ores, these ores are often difficult to process with the same treatment techniques.

One known laterate acid leaching process is called high pressure acid leaching (HPAL). (For example, see Joseph R. Boldt, published by Longmans Canada Ltd. in 1967, “The Victory of Nickel: Its Geology, Mining and Extraction Metallography. It's Geology, Mining, and Extractive Metallurgy. ”This process was first introduced in Cuba's Moa Bay in the late 1950s, with an additional plant in Western Australia in the late 1990s. Was built in.

This process utilizes sulfuric acid leaching at high temperatures, typically 250 ° C., and at high pressures (related vapor pressure is approximately 570 psi at 250 ° C.). At this temperature, the nickel-iron minerals in the ore are almost completely dissolved. Molten iron quickly precipitates in hematite (Fe ₂ O ₃ ) at the high temperature applied, since the compound is mainly insoluble in solutions which are only slightly acidic at this temperature. After the nickel remains in solution and cooled, the leaching residue containing iron is concentrated by a series of wash thickners called the counter-current decantation circuit (CCD). It is separated from the solution containing nickel. Thus, separation of iron from nickel, which is the main purpose in the leaching process, is achieved.

The main disadvantage of the high pressure acid leaching (HPAL) process is that it requires complex high temperature autoclaves and connected devices that are both expensive to install and maintain. In addition, the high-pressure acid leaching step, a high temperature, most of the sulphate ions are sulfate ions (bisulfate, HSO ₄ ^-) of which is provided by sulfuric acid, because it is enclosed in, to melt the metal compound other than iron ore that required stoichiometrically Consumes more sulfuric acid. In other words, sulfuric acid (H ₂ SO ₄ ) removes only one hydrogen ion (H ⁺ ) at high temperature. During the cooling and neutralization of the leach liquid, bisulfate ions decompose into sulfates (SO ₄ ^{2 −} ) and another hydrogen ion. The latter hydrogen ions (acids) are therefore not fully utilized for leaching, for example, as excess sulfuric acid to be neutralized with limestone (stone).

Another drawback of the high pressure acid leaching process is the handling of predominantly ironite type infusions, since the presence of saprolite leads to a large and often uneconomical increase in sulfuric acid consumption, which results from leaching magnesium from saprolite. It is limited. This further exacerbates the problem of “change” of the biacid salt at the high temperatures described above.

U.S. Pat. No. At 4,097,575 roasters were calcined to harden the ore at about 820 ° C. or lower to speed up the reaction to sulfuric acid, and then to neutralize excess acid released from the autoclave where pressure leaching of the iron ore occurs. The development of a high pressure acid leaching process consisting of using roaster calcine) is described. Nickel in the saprolite ore is mostly dissolved during the neutralization process. The reported advantage of this process is that it makes better use of sulfuric acid added during the pressure leaching of the iron ore, reducing the consumption of limestone or other expensive neutralizing agents for treating liquids emitted by the autoclave, and the typical nickel laterite orebody. Has achieved the ability to process both the limestone and saprolite portions of. The disadvantage of this process is that it still requires an expensive autoclave to leach the iron, and both capital and operating costs require a hard heating of the expensive saprolite ore.

U.S. Pat. No. Further developments to the process described in 4,097,575 are described in U.S. Pat. Pat. No. 6,379,636 B2, in which the step of hard heating of the saprolite has been eliminated and the raw form of saprolite is used to neutralize the excess acid in the exit solution of the autoclave. In addition, more acid can be added to be used to increase the amount of saprolite that can be leached. However, this process still requires the use of expensive autoclaves.

Several processes have also been described which use acid leaching under atmospheric pressure, eliminating the disadvantages of pressure leaching described above. U.S. Patent No. 3,793,432 reacts ore with sulfuric acid at or below its boiling point, and atmospheric leaching for laterite ores, where precipitation of dissolved iron is achieved by the addition of precipitating reagents such as ammonium, sodium, potassium or lithium ions. ) Process is described. Although not explicitly described, as demonstrated by the high iron content and low magnesium content of the implant ore, all examples mentioned in the specification used amonite ore as a sample. Although this process overcomes the disadvantages of pressure leaching, it has other disadvantages. First, precipitation of iron will become a jarosite compound, a thermodynamically unstable compound of iron that releases sulfuric acid over time, thus causing environmental problems. (Although iron is not explicitly mentioned, it is obvious to the person skilled in the art that iron is to be precipitated under the conditions of the examples.) Iron is contained 2 moles of sulfate every 3 moles of iron, and therefore This compound exhibits a significant excess consumption of sulfuric acid to provide the required sulfate ions.

Secondly, the extraction of nickel from the ore is clearly relatively low. Although extraction is not explicitly mentioned, the nickel content of the residue, and that the iron is found to be heavier than the raw ore because the iron is lower in percent of iron than the original ore and all iron remains in the residue substantially. In fact, nickel extraction is generally in the range of 60-65%. Thirdly, very long leaching times of 4 to 5 days are required. Fourth, it is necessary to add relatively expensive iron flocculants such as potassium carbonate and sodium carbonate.

U.S. Patent Nos. 6,261,527 B1 and 6,680,035 B2 are the first result of the leaching of ore into the strong acid “completely” leaching into the acid, ie both nickel and iron are sufficiently melted from the goethite, and then the iron is simultaneously precipitated into the iron ore by the addition of the iron alum. Another atmospheric leaching process is described, in which saprolite ore is leached in a typical limonite leaching slurry. This process also has the disadvantage that iron ore is produced, which requires separate mining and preparation of the ore and the saprolite portion of the ore, and is limited only to a narrow range of the saprolite ratio to the ore. The latter is due to the fact that the amount of saprolite that can be effectively leached is determined by the amount of iron (II) ions in the limonite leaching solution.

WO 03/093517 A1 describes the development of this process consisting of the removal of ionic additions which lead to the formation of iron alum, such as sodium, potassium and ammonium and the precipitation of iron with compounds such as goethite rather than iron alum. This process overcomes the shortcomings of iron ore, but the consumption of sulfuric acid is 0.59 to 0.87 tonnes per tonne ore in the examples mentioned, and more than 0.72 tonnes for 1 tonne ore in nine of the 11 examples mentioned. to be.

U.S. Patent Nos. The processes described in 6,261,527 B1 and 6,680,035 B2 and WO 03/093517 A1 were based on the fact that goethite is more difficult to handle in acid leaching than typical saprolite minerals such as serpentine. This was demonstrated by other researchers. (See, for example, John H. Canterford, “Leaching of some Australian Nikeliferous laterites with Sulfuricacid at atmospheric pressure”, Proc. Australasian Inst. Min. Metall., P. 265 (1978) .19-26; “Leaking of laterite nickel ores in hydrochloric acid,” by NM Rice and LW Strong, Canadian Metallurgical Quarterly. 13 (3), 485-493; and Figure 5 of US Patent No. 5,571,308.) Therefore, only saprolite can be effectively used in the second stage of leaching, in which iron precipitation occurs simultaneously. This is because the acidity of the solution must be relatively low in order to enable precipitation of iron ore, and lower in order to enable precipitation of goethite or other hydration products of iron ions (II). The goethite contained in the limestone will leach very slowly under these conditions. Thus, galmonite (primarily goethite) is leached at the beginning with relatively high concentrations of acid and both iron and nickel as solvents.

U.S. Pat. No. 3,244,513 are mainly used to mix laterite ores of the Limonite species (defined as iron> 25%) with concentrated sulfuric acid in the presence of limited water, and then to allow nickel, cobalt, magnesium and manganese to be selectively sulfated for iron. The process of heating a mixture vigorously at about 500-725 degreeC is described. Subsequent water leaching results in high extraction of nickel and cobalt and low extraction of iron into the solution. The advantage of this process is that it does not require expensive autoclaves to affect leaching. The main disadvantage is the need for expensive, hard heating steps.

U.S. Pat. No. 4,125,588 eliminates the hard heating step, and mixes the ore and the concentrated sulfuric acid in a carefully controlled manner, in which the ore first mixes with the concentrated sulfuric acid in the absence of a significant amount of water, and then adjusts the amount in order to initiate the sulfidation of the ore. US except for adding water and then finally leaching the mixture with an additional amount of water Pat. No. A process similar to that described in 3,244,513 is described. The advantage of this process is that this is U.S. Nos. 3,244,513 eliminates the hard heating step required. However, this process also has significant disadvantages.

One disadvantage is that the ore used must contain up to 1% moisture, which means that in most cases the ore must be dried, since the moisture content of laterite ores is 20% or more. A second disadvantage is that this process does not provide for selective dissolution of nickel to iron, as exemplified in all the examples mentioned in the patent (> 90% iron extraction). Separation of nickel from iron in solution generally results in the loss of additional nickel.

In addition, this process is only applicable to ores containing “high amounts of magnesia and silica”, ie saprolite or silica. The exact composition of the ore was not mentioned in all examples, but the information that contains 3-4 times more magnesium than iron clearly indicates that no iron ore minerals (iron / magnesium ratio by weight approximately 10 to 90) were taken into account. In addition, the acid consumption in this process is very high, ranging from 0.9 to 1.1 tonnes of sulfuric acid per tonne of ore.

U.S. Patent No. 3,093,559 describes another process for treating laterite ores with relatively concentrated sulfuric acid (from approximately 25-50% sulfuric acid). In this process, a sufficient amount of acid is added to cause sulfation of most or all of the metal content, including iron. The leached solution is then evaporated to dryness to separate iron from nickel, and the resulting salt is heated vigorously at 975 to 1,050 ° F. to convert iron to hematite. Subsequent leaching of calcine causes nickel to come out of solution and iron to remain in the residue. An important disadvantage of the process is that it requires a high temperature hard heating process, such as the various processes described above.

U.S. Patent No. 2,899,300 processes moisture laterite ores with concentrated sulfuric acid and then bakes and drys the mixture between 100 ° C. and 150 ° C., preferably at 125 ° C., prior to water leaching to dissolve the metals in the ores. This is described. The baking step is a major drawback because it requires a significant amount of thermal energy to evaporate the water contained in the ore / acid mixture. Furthermore, as exemplified in the examples given in the specification, with ~ 77% and 53% respectively, nickel dissolution is relatively low while iron dissolution is relatively high. Nickel extraction can be reportedly increased by adding additional sulfuric acid to the residue in the first step, followed by a second step of sulfation and water leaching, but this adds complexity to the process and It will not improve the iron / nickel separation achieved.

US Patent No. RE 37,251 describes a pressure leaching process for ores and concentrates, not copper (I), including nickel laterite ores, using an acidic solution of bisulfate and sulfate ions with halogen ions such as chlorine ions and oxygen. It is. According to the specification required temperature and pressure is 225 ℃, 450 psig O ₂ . If the vapor pressure at 225 ° C. is approximately 370 psi, the total pressure will be in the range of 820 psig. These conditions are quite similar to the high pressure acid leaching conditions and would therefore require the use of the expensive high pressure autoclave system described above.

It is an object of the present invention to obviate and mitigate the disadvantages of the known process. Produced by bulk mining of typical laterite ore bodies at atmospheric or low pressure conditions, without subsequent classification of any ore types, while achieving high extraction of nickel and cobalt and very low ultimate iron extraction. It is also an object to provide an acid leaching process of a mixture of limonite and saprolite ore.

Brief summary of the invention

The present invention comprises the first step consisting of reacting ore with concentrated mineral acid and the second step consisting of preparing a slurry mixture of acid / ore in water and leaching the mixture to separate nickel and cobalt. The stepping process provides an efficient leaching process of nickel and cobalt from amonite and saprolite ores, which can be produced from the bulk mining of typical nickel laterite deposits. Iron is effectively separated from nickel and cobalt in solid leach residues, mainly as oxides or hydroxides of iron ions, but not iron-platite.

This process may also include curing of the mixed ore and acid prior to water leaching, and the ore is first broken before mixing and reacting with the acid. Preferably the saprolite is broken apart and then mixed with limonite. More preferably, the acid is first mixed with limonite and subsequently saprolite is added.

Leaching with water is advantageously carried out at somewhat elevated temperatures. For leaching under atmospheric pressure the temperature is preferably in the range of 90-150 ° C. Alternatively faster leaching times may be obtained by leaching in an autoclave at temperatures up to about 150 ° C. The corresponding pressure is up to about 70 psia and is approximately equal to the saturated vapor pressure at the leaching temperature.

The acid used is preferably selected from sulfuric acid, hydrochloric acid or nitric acid, more preferably sulfuric acid. Mixtures of any of these acid (s) can also be used.

Advantageously, cobalt elution is increased during water leaching by reducing agents such as sulfur dioxide, hydrogen sulfide, soluble bisulfite and sulfite compounds, or soluble iron (II) ionic compounds. To be added.

The process of the present invention avoids costly high pressure autoclaves and reduces the production of albite compounds. In some embodiments, this also eliminates the need for mining separately and processing the ore types of amonite and saprolite separately, which makes it possible to treat the ore of the saprolite ratio for a wide range of amonite. .

It has been found that the present invention can achieve at least about 80% nickel extraction and 95% or more cobalt extraction with up to 15% iron extraction.

Detailed description of the invention

The present invention provides an improved process for extracting nickel and cobalt from nickel-iron laterite ores, except for most of the iron contained in the ores in the solid leach residues. The process does not require preferential separation of the iron ore ore from the saprolite ore and can treat a mixture of the two types of ores. However, if it is convenient to mine the iron and the saprolite ores separately, high nickel extraction can be achieved by first reacting the acid directly with the iron ore.

Referring to FIG. 1, laterite ore in mining consisting of a mixture of amonite and saprolite is first broken into large sizes of about 5 to 10 mm. The broken ore is then mixed with a concentrated inorganic acid selected from sulfuric acid, hydrochloric acid, nitric acid, or mixtures thereof in a suitable apparatus such as a pug mill. It is not necessary to dry the ore, which typically contains a significant amount of free water in the state of mining, before adding the concentrated acid.

The amount of acid added is at least a stoichiometric amount of soluble iron-non-metals in the ore, ie most nickel, cobalt, magnesium, aluminum, copper, zinc and small amounts of chromium (but not iron) in the ore during subsequent leaching steps. It is the amount needed to dissolve it theoretically. A slight excess of acid is added to provide some free acidity in the subsequent leaching step to leach a small fraction of iron and to ensure maximum leaching of nickel and cobalt. The addition of acid is limited to ensuring that the final iron leaching from the ore slurry is minimized. In some cases some of the magnesium and aluminum may not melt, which must be taken into account in determining the correct addition of acid.

The addition of water during the acid mixing process is minimized to the extent that it is possible to provide the highest concentration of acid for reaction with the ore minerals. The addition of water is only necessary if the mixture of ore / acid hardens and cannot be mixed thoroughly or easily handled. The addition of concentrated acids to moist ores generates a significant amount of heat that raises the temperature of the mixture to or above the boiling point of the water and causes significant evaporation of the water. To adjust the concentration of the acid / ore mixture, such as the liquid paste, additional water may be added during the mixing process if necessary. If no water is added at all, depending on the moisture content of the ore and the total amount of heat generated, it may form a dry powdered reaction product (a more difficult, more difficult, stiff, intermediate between liquid paste and dry, powdery material). Mixtures such as toffees may be produced.)

Ideally, the mixture will have a concentration of liquid-paste or dry powder that allows for easy handling. The resulting acid / ore mixture or “pug” can be “cure” for a sufficient time at ambient temperature, allowing sufficient reaction of the mineral acid with the mineral components of the ore. This is accomplished by stockpiling the beaten soil onto a prepared non-permeable pad and placing it on the pad for several days before the material is recovered and taken to the next stage of the process.

In order to facilitate stockpiling and recovery, it may be desirable to form a small, separate portion of the beaten soil, for example extruding or pelletizing the material prior to stockpiling. However, it has been found that satisfactory results will be achieved with minimal or no curing time. In this case the acid / ore mixture will be mixed directly with water. Subsequent grinding would not be necessary if the ore was ground before adding the acid. Longer curing in subsequent water leaching steps will result in slightly better nickel extraction and lower iron extraction.

After curing, the beaten soil is ground to a particle size sufficient for leaching in an agitated tank. The water needed to form the leach slurry can be added prior to milling, so this step can proceed in the wet state. The beaten soil must be crushed to a particle size that allows solid suspension to the bottom of most of the particles in the slurry so as not to require excessive stirring power injection. In order to minimize the need for water and to produce a full leaching solution, including the most concentrated nickel and cobalt, it is desirable to make the slurry as dense as possible so that it is consistently well mixed during leaching.

If the process is to be continuous, the resulting leach slurry is heated as needed in a batch leach reactor or series of leach reactors. The temperature of the leach slurry is preferably maintained at or near the normal boiling point of the leach solution, typically 95-100 ° C. Instant steam injection or other means of adding thermal energy can be used for this purpose. Some necessary heat is provided from the heat of the solution of the metal salt formed during the mixing of excess acid and acid. In addition, the more necessary amount of heat can be obtained during the soil-winning process itself, either by preheating the water by cooling the reactor which beats the soil with water to be used for leaching, or by recovering steam formed during the reaction of acid and ore. .

The leach mixture is allowed to react until most of the nickel and cobalt in the ore is eluted and most of the iron is in the solid leach residue. The stirring leaching retention time is on the order of 12 to 48 hours. Satisfactory results are typically obtained at about 15 hours or less. In some cases, it may be desirable to add a controlled amount of reducing agent to the leaching slurry to increase the leaching of oxidized manganese minerals in the ore, ie cobalt associated with asbolane. As will be apparent to those skilled in the art, for example, a controlled addition of sulfur dioxide can be added to improve cobalt (and manganese leaching). Other reducing agents capable of reducing trivalent and tetravalent manganese oxides to divalent may also be used.

Using a temperature above the boiling point can greatly facilitate the leaching process. For example, if the leaching slurry is heated to about 150 ° C. in the autoclave, the leaching time can be shortened to about 1 hour. Increasing the temperature further may be beneficial, but the object of the present invention is to eliminate the complexity associated with high pressure leaching. It should be noted that the pressure required to leach the winning soil and water slurry at 150 ° C. is only 70 psia (vapor pressure at 150 ° C.). At this pressure, the slurry can be pumped into the autoclave with a conventional centrifugal pump and the pressure drop system can be a valve or choke, one very simple step. Suitable autoclave devices to be used in the present invention are completely different from those requiring high pressure leaching at temperatures around 250 ° C. and pressures around 450-500 psi, as will be apparent to those skilled in the art.

The use of a temperature above the boiling point will also give a higher nickel / iron ratio in the solution, which is advantageous with regard to the latter process of the leach solution. This is because virtually all iron must be removed from the solution in most cases before it affects the recovery of nickel and cobalt. Generally, iron remaining in the solution is removed by adding a base, such as calcium carbonate, to the leaching slurry and precipitating the iron oxyhydroxide compound. Some nickel may precipitate with iron, resulting in loss of spent metal, and in addition, neutralizing agent means additional operating costs of the process. A further advantage of using high temperatures is that higher settling rates are achieved at higher leaching temperatures, thereby improving the solid / liquid separation properties of the final leaching slurry.

In some cases, it may be advantageous to use both a series of leach stages under atmospheric pressure and subsequent leach stages of normal pressure. Sulfur dioxide, or other reducing agent, for the leaching of oxidized manganese minerals and cobalt and nickel contained is most conveniently added in the leaching step under atmospheric pressure. Subsequent pressure leaching steps can still be used to achieve the advantages mentioned in the preceding paragraph.

Nickel extraction of at least 80% and at most 90-95% can be achieved with the present invention, with the extraction of the corresponding at least 5-10% of iron. Moreover, the leaching residue typically contains at least about 1-2% sulfur, indicating that the iron compounds in the residue are not originally of the iron-plated stone type.

When sulfuric acid is used as a lixiviant, the leaching step of the present invention is carried out in the prior art, for example US Pat. It differs from 3,244,513, the latter being different in that it employs a hard heating step after mixing the acid to convert any iron sulfate formed during acid mixing into iron oxide. Sulfur trioxide gas (SO ₃ ) is released during this conversion, which reacts with oxides of iron-non-metals, namely NiO and MgO, to form non-ferrous sulfate. In the prior art processes, therefore, the iron-non-metal is preferentially sulfated.

Surprisingly, however, the cost of capital is enormous, and the hard heating step, which requires a significant amount of expensive fuel energy to achieve the desired heating temperature, i. It has been found that after achieving and thus leaching the ore / acid mixture with water, it is not necessary to achieve a full leaching solution containing most of the soluble iron-non-metals, in particular nickel and cobalt, and a relatively small amount of iron. It is all that is necessary to provide sufficient time and temperature during the water leaching to influence the desired reaction. Thus eliminating the hard heating step is a major advantage of the prior art.

After water leaching, the leaching slurry is solid / filtered by filtration or thickening to produce a full leach solution containing most of the nickel and cobalt in the ore and a solid residue containing most of the iron in the ore. The liquid is separated. Advantageously, the increment is a method called counter-current decantation (CCD) to flush out most of the droplet-associated metals from the leach residues, with a series of increments along with the flow of effluent and the leach slurry. Performed by a thicker. The metal value is overwhelmingly reported by the overflow of the first thickener, which is a full leaching solution.

Filled leaching solutions are known to those skilled in the art, such as sulfide precipitation, hydrogen sulfide or hydroxide precipitation using solution extraction, ion exchange, sulfiding reagents, for example using magnesia as precipitant. By the method, nickel and cobalt are recovered.

Nickel and cobalt can also be recovered from the leach slurry using a resin-in-pulp process without prior solid / liquid separation. In this process, nickel and, if possible, ion exchange resins which extract cobalt are added directly to the leach slurry. After extraction is complete, the resin is separated by screening from the leaching slurry depleted of nickel. After washing the resin to remove solids, nickel can be separated from the resin by dissolving it in a fresh acid solution.

Prior to or during the recovery of nickel and cobalt by one of the above-mentioned methods, the leach solution may be prepared such as calcium carbonate, magnesium oxide, sodium carbonate or the like to neutralize the excess acid remaining in the leaching process. It will be neutralized with a neutralizer and small amounts of iron ions (II), aluminum and chromium will be precipitated with minimal precipitation of nickel and cobalt. This process will be performed in one or several steps, by separation of the intermediate solid / liquid.

In one aspect of the invention, the first step of neutralization will be performed prior to separating the leach residue from the leach solution. The combined leach and neutral residue will then be separated from the partially neutralized leach solution by filtration or increment as described above. The second step of neutralization would still be desirable according to the method chosen for the recovery of nickel and cobalt from the full leaching solution. After the second step of neutralization, the resulting neutralization residue will be separated from the neutralized leach solution by filtration or increment. Ideally, the second stage of neutralization residue is returned to the first stage of neutralization to redissolve any co-precipitated nickel and cobalt.

Surprisingly, it has been found that there is no need to separate the limestone in laterite ore from saprolite for handling individually and separately in the leaching process. Moreover, the process is not limited to any saprolite / saltstone ratio as long as the minimum saprolite / saltstone ratio is met. The amount of acid added to the ore will be adjusted to the ratio range of saprolite / monite based on the content of the non-iron metal in the saprolite / monite ore mixture.

Surprisingly, patent number U.S. As in 3,793,432, it has been found that no iron precipitant needs to be added to the leach slurry. The addition of iron precipitants, such as sodium, potassium, and ammonium ions, is disadvantageous in that it promotes the formation of iron-containing alumite compounds, which are thermodynamically unstable and slowly decompose over time to release sulfuric acid. This can lead to re-dissolution of metal impurities present in the leach residues, potentially leading to environmental pollution. In addition, additional sulfuric acid is needed to meet the needs of sulfate ions needed to form iron ore during leaching.

In the process of the present invention, for example, sulfuric acid is added to the ore, nickel, cobalt and iron as well as other non-ferrous metals in the ore to become sulfates during the non-selective, i.e., mixing acid / ore. Sulfates dissolve well in water during subsequent leaching steps. However, the extraction of nickel and cobalt is incomplete at first because the amount of acid initially added is not sufficient to convert all metals, including iron, to sulfates. Thus, a significant amount of ions enter the solution together with nickel, cobalt, magnesium, aluminum, chromium, copper and zinc. However, by continuing the leaching process after the initial sulphate dissolution step, the iron content of the solution is reduced and additional nickel and cobalt are extracted from the solid phase. The exact iron extraction and recrystallization reactions are unknown, but the sulfur content of the final residues is predominantly below 2%, indicating that no significant amounts of iron ore have been formed.

One preferred embodiment of the present invention is a leaching slurry of iron-containing “seed” material at the start of water leaching to accelerate precipitation of dissolved iron ions and extraction from the remaining solid phases of nickel and cobalt. The surface of the seed particles provides low activation energy sites for hydrolysis and precipitation of iron, for example ferric hydroxide, goethite or hematite, ideally the seed material is ideally precipitated Part of the final leach residue itself, which contains the ferrous iron compound, one method of carrying out the process with seed recycling is shown in FIG.

As another aspect of the invention, the limonite portion of the ore is first mixed with concentrated sulfuric acid by calculating the amount of acid in the manner as discussed above and then the saprolite portion of the ore is added to the limonite / acid mixture. The saprolite ore will be added to the mixture of limonite / acid as broken saprolite or crushed saprolite. In the former case the ore / acid mixture will be ground before water leaching. Grinding will be performed without the addition of water. Water is added to the ore / acid mixture and wet grinding is advantageously employed prior to water leaching under atmospheric pressure. In the latter case the saprolite ore can be crushed in the dry state and added with water to the acid / monite mixture, or it can be crushed in the wet state and added as a slurry or filtercake to the limonite / acid mixture. The resulting mixed ore and acid are then water leached as previously described above. The method offers the greatest opportunity to sulfate the goethite component of laterite ores. A flowchart for carrying out this embodiment of the invention is shown in FIG. 3.

1 is a flow sheet showing one embodiment of the process of the present invention in a simplified form.

Figure 2 shows another embodiment of the process of the present invention in which some leach residues are recycled to provide a seed for the precipitation of iron.

Figure 3 shows a third embodiment of the process of the present invention in which all the required acids are first mixed with amonite ore, and then the saprolite ore is mixed in with the resulting limestone / acid mixture.

4 shows that all the required acid is first mixed with amonite ore and then the resulting saprolite ore and water broken into the resulting ironite / acid mixture, and then the resulting mixture is ground and then leached under atmospheric pressure. The fourth aspect of the process is shown.

The following examples illustrate the method of the present invention. The ores used in these examples come from Central America laterite deposits (layers). The limonite and saprolite portions of the ores used in Examples 1-3 have the compositions shown in Table 1. Saprolite ore is ground to about -6.4 mm before being used in the experiment.

Table 1

% Ni % Fe % Mg %moisture Limonite Ore 1.41 47.7 0.67 34.5 Saprolite Ore 3.17 8.7 17.8 21.3

Example  One

Approximately 1 kg of saprolite ore was ground wet in a ball mill, with approximately 100% passing through 100 mesh. The milled slurry was filtered to form a milled saprolite containing 41% moisture. 425.2 g of watery ground crushed saprolite was mixed with 381.7 g of watery limonite ore to produce 500 g of leach feed material with a limonite / saprolite ratio of 1: 1 in the dry state.

The ore mixture was placed in a 4.5 liter narrow neck vial. The bottles were rolled on a bottle rolling device at about 47-48 rpm at slightly tilted angles on the horizon. 312.5 g of 96% sulfuric acid was added to the bottled ore mixture over a period of 30 minutes. After all the acid was added, the ore and acid were mixed for approximately 30 minutes. Upon completion of the ore and acid mixing, a semi-liquid mass was formed and the temperature rose to between about 70 and 100 ° C.

The bottle was removed from the rollers and the mixture of acid and ore was allowed to cure for approximately 72 hours at ambient temperature. 622 ml of water was then added to the cured mass and the mixture was stirred to form a uniform leaching slurry. The leach slurry was transferred to a 2-liter cylindrical reaction kettle equipped with a mechanical lid, four plastic baffles and a tight lid with an open water condenser open to the atmosphere. The reaction kettle was heated externally with an electrical heating mantle.

While stirring vigorously, the leaching slurry was heated and maintained at a temperature of approximately 96-99 ° C. for a period of 48 hours. After 5 hours of leaching, 129 g of finely ground good grade hematite was added into the reactor to serve as a “seed” for iron precipitation. Leach solution samples were taken periodically for chemical analysis. At the end of the 48 hour leaching period the entire slurry was filtered. The filter cake (filtercake) was double repulping (repulp) fresh water to wash away the metal entrained. The cake was then dried and weighed. The dried solids, the filterate and the combined wash water were examined respectively.

The results of this experiment are shown in Tables 2 and 3 below.

TABLE 2

Hours (h) [Ni] g / L [Fe] g / L [Mg] g / L [H ₂ SO ₄ ] g / L One 8.8 70 27 <0.5 2 8.0 49 27 <0.5 5 9.5 24 35 <0.5 24 8.7 14 35 8 32 8.3 12 32 9 48 8.8 10 28 16

TABLE 3

Ni (% or g / L) Fe (% or g / L) Mg (% or g / L) S (%) Final solution 9.3 12.8 39 - Final residue 0.34 42.9 0.52 0.98 Calculated Extraction * 85.2 11.7 94.9 -

Nickel, iron and magnesium extraction were based on the weight, volume and inspection of the final residue and solution.

The results show that high extraction of nickel was achieved while extracting very little iron present in the ore. In addition, the sulfur content of the residue was very low and the Fe / S ratio was approximately 44: 1. At least 5% of the sulfate contained in sulfuric acid was carried to the residue. If all the iron in the ore formed iron ore, the Fe / S ratio of the final residue would be approximately 4.3. In addition, X-ray diffraction analysis of the residue revealed the presence of spinel phases originally present in hematite, goethite and ore.

Kinetic data indicate that it is necessary to leach a little longer than 24-32 hours to approach the final iron concentration in solution. Residual iron can be neutralized with a base, for example limestone, and removed from the solution without losing a significant amount of nickel in the solid phase.

Example  2

A sample of 381.7 g of limestone ore was mixed with 95 g of water and 317.7 g of pressed crushed saprolite ore in a 2 liter flat bottom glass beaker. Since the samples were somewhat dry compared to the original conditions, water was added to simulate the water content expected of the run-of-mine ore. 312.5 g of 96% sulfuric acid was added to the beaker over approximately 30 minutes. The acid was mixed with the ore using a stirrer that returned to about 60 rpm. The addition of acid was sufficient to give the ratio of acid to ore of about 600 kg H ₂ SO ₄ per tonne of ore (dry).

Ore and acid formed semi-liquid masses which were poured into a narrow pan to cure for about 72 hours at ambient temperature. Some of the mixture was not recovered in this operation. Based on the weight recovery, it was estimated that about 13% of the mixture was not recovered.

After curing of this period, the significantly hardened acid / ore mixture was broken into pieces and transferred to a small grinding mill. 300 g of water were added to the grinder and the mixture was ground with the rock medium to reduce the maximum particle size to about 100 mesh for approximately 1 hour. Additional water was added to the grinder to wash the slurry into a 2 liter leaching reactor. 1858 g of leach slurry were prepared in this way. The reactor was stirred with a temperature of 95-105 ° C. and the leaching was allowed to continue for 44 hours. After 5 hours of leaching, 129 g of finely ground good grade hematite was added to the reactor to act as a “seed” for iron precipitation. Samples of leach liquids were taken periodically for chemical analysis. At the end of the 48 hour leaching period, the entire slurry was filtered. The filtercake was repulsed twice in fresh water to flush out the entrained metal. Then the cake was dried and weighed. The dried solids, the filterate and the combined wash water were examined respectively.

The results of this experiment are given in Tables 4 and 5 below.

Table 4

Hours (h) [Ni] g / L [Fe] g / L [H ₂ SO ₄ ] g / L One 4.6 25.6 <0.5 2 5.36 29.7 <0.5 5 6.15 15.5 5 24 4.73 4.74 12 44 5.49 7.55 16

Table 5

Ni (% or g / L) Fe (% or g / L) Mg (% or g / L) S (%) Final solution 5.51 8.7 22.0 - Final residue 0.41 42.8 0.47 0.83 Calculated Extract * 81.1 11.3 94.7 -

Extraction of nickel, iron and magnesium was based on final precipitate and solution weight, volume and inspection.

The experiment simulates one of the preferred embodiments of the present invention in which the saprolite ore is not pulverized before being mixed with galmonite and acid is added. Instead, the mixture of acid / ore was crushed after curing.

The results of this experiment were similar to those of Example 1, although the extraction of nickel was slightly lower.

Example  3

This experiment was carried out in a manner similar to that of Example 2, with the following exceptions. The amounts of limonite ore, crushed saprolite ore, water, acid and hematite seeds used during mixing with the acid and subsequent processes were 336.9g, 280.3g, 84g, 275.8g and 113g, respectively. The ratio of ore, water and acid is the same as that of Example 2.

After the ore was mixed with the acid, the mixture was transferred to a narrow pan and allowed to cure for only 1 hour before the transfer and subsequent water leaching to the wet grinding mill. After an hour of curing of the mixture, the mixture was still liquid and could be poured into the grinder. Exactly 1776 grams of leach slurry was prepared and leached as in Example 2. The recovery of the acid / ore mixture was approximately 96%.

The results of this experiment are given in Tables 6 and 7 below.

Table 6

Hours (h) [Ni] g / L [Fe] g / L [H ₂ SO ₄ ] g / L One 5.53 22.7 <0.5 2 6.37 24.2 <0.5 5 7.59 16.7 4 24 7.4 9.14 13 44 7.12 10.8 16

TABLE 7

[Ni] (% or g / L) [Fe] (% or g / L) [H ₂ SO ₄ ] (% or g / L) S (%) Final solution 6.97 10.2 34.0 - Final residue 0.37 43.0 0.45 0.81 Calculated Extract 84.2 10.8 95.2 -

Nickel, iron and magnesium extractions were based on final residue and solution weight, volume, and inspection.

The results of this experiment, similar to the previous example, demonstrate that there is no need to cure the ore / acid mixture for a long period of time prior to leaching.

Example 4

The compositions of the ores used in the experiments below are given in Table 8.

Table 8

 Ni% Co% Fe% Mg% Si% moisture% Limonite Ore 1.31 0.2 47.2 0.63 2.67 41.9 Saprolite Ore 3.13 0.01 6.0 20.0 18.8 14.4

409.6 g (when wet) of the iron ore was placed in a small magnetic ball mill filled with a porcelain ball of 0.6-1.9 cm. Before adding the ore, the mill and grinding media were preheated to approximately 100 ° C. This was done to simulate the thermal conditions that would be present if the process was performed on a continuous basis. The grinder has a hole in the center, which is fitted with a plastic lid to add sulfuric acid. The grinder was rotated on the same roller as used in Example 1. 338.5 g of 96% sulfuric acid was added to the grinder over a period of approximately 1.5 minutes and allowed to react with the iron ore and acid for a period of 15 minutes. The temperature in the grinder is monitored during the reaction with a laser thermometer, which is taken by hand. The measured temperature varied from 97 to 121 ° C. with the highest temperature recorded about 1 minute after the acid addition was complete. At the end of the reaction period, the grinder is removed from the roller, the lid is removed and 306.1 g (when wet) of the broken saprolite ore with 510 g of 30 g / L Mg aqueous solution (such as MgSO ₄ ) in a limonite / acid mixture. Added. MgSO ₄ The solution was used to simulate the effect of using a poor solution, such as water, that constitutes the process remaining after purification and nickel / cobalt recovery. A tight magnetic lid was used to close the grinder and the entire ore / acid mixture was ground for approximately 60 minutes, until most of the solid particles present had a size of 100 mesh or less.

The total saprolite / monite ratio based on the weight and moisture content of the ore sample was 1.1 at dry condition (dry ore and 100% H ₂ SO ₄ Based on total acid / ore ratio was 0.65.

The contents of the mill were then discharged into the same reaction kettle as used in Example 1. A coarse screen was used to hold the grinding media and 764 g of 30 g / L Mg solution was used to rinse the ball mill and grinding media. Rinse solution was added to the reaction kettle.

The reaction kettle was stirred and heated to 95-100 ° C. Leaching was performed for 24 hours. During the first 5 hours of leaching, sulfur dioxide gas was bubbled into the leaching slurry to control the oxidation reduction potential (relative to saturated Ag / AgCl standard electrode) between 540 and 600 mV. Samples of the slurry were taken throughout the entire leaching period, the solids were washed, and the solution and solids were examined.

At the end of the leaching period, 20% limestone slurry was added to the leaching slurry over approximately 2 hours to neutralize the leaching slurry at 95 ° C. to pH 3.0. The neutralized leaching slurry was taken and the whole slurry filtered.

The results of the leaching period are given in Table 9.

Table 9

Hours (h) [Ni] g / L [Co] g / L [Fe] g / L [H ₂ SO ₄ ] g / L One 4.92 0.229 60.8 <0.5 2 5.33 0.248 50.0 <0.5 4 6.22 0.290 33.9 <0.5 5 6.52 0.303 26.5 <0.5 24 7.36 0.34 14.4 5.0

The solid upon completion of the leaching step showed 0.21% Ni, 0.006% Co, 31.3% Fe, 0.65% Mg, 14.9% Si and 1.39% sulfur. Calculate the leach residue weight using the silicon detection of the ore and residue, calculated using a “silicon tie” (assuming zero silicon extraction, and the weight of the ore and residue to calculate the extraction And detection), Ni and Co extraction were approximately 93.1% and 95.9%, respectively. Iron extraction was approximately 9%.

After neutralization to pH 3, the solution showed test results containing 6.23 g / L Ni, 0.29 g / L Co and 3.24 g / L Fe. The solids showed 0.22% Ni, 0.009% Co, 26.5% Fe and 11.7% Si. Approximately 74% of the iron in the leaching solution precipitated during neutralization; Therefore, the net iron extraction after neutralization was only about 2%. Approximately 2% nickel and 4% cobalt precipitated with iron during neutralization.

The above examples illustrate some important features of one of the preferred embodiments of the invention. Extraction of high nickel can be achieved by first reacting sulfuric acid directly with amonite ore, and then adding saprolite ore; The addition of a reducing agent, in this case sulfur dioxide gas, to the leach slurry improves the extraction of cobalt. It is not necessary to cure the acid / ore reaction mixture to achieve high extraction; The reaction of sulfuric acid with ore is very fast and very exothermic, so that very high temperatures can be achieved during the acid / ore reaction, thus increasing the reaction rate and making the apparatus necessary for carrying out the sulfate step very small. ; And partial neutralization of the leaching slurry following leaching can be used to remove most of the small amount of iron, which is leached with very low amounts of nickel and cobalt.

It will also be appreciated by those skilled in the art that many variations of the process will be possible within the broad scope of the invention. Those skilled in the art will recognize that the invention based on this specification may be used in other aspects.

The process of the present invention is broadly applied to nickel laterite ore bodies, which include both limonite and saprolite, including most such ore bodies. The saprolite / monite ratio in the process varies widely as long as the minimum constant saprolite / monite ratio used in the process is used. For example, the minimum ratio when using sulfuric acid is to be sulfated by sulfuric acid, which can be produced by hydrolysis and precipitation of the sulfated iron content of maneite and saprolite ore with an oxide or hydrate of iron (II) ions. Can be roughly determined by calculating the amount of non-iron components of the saprolite ore. If the addition of acid is calculated as described above, i.e. sufficient acid to react with the non-sulphurable iron components of the iron ore and the saprolite ore plus a little excess of acid to complete the reaction and also to complete the reaction If calculated, the ratio will be any value equal to or greater than this minimum value.

The above disclosure is for the purpose of description and the scope of the invention is defined by the following claims.

Claims (16)

Leaching method of laterite ore containing amonite and saprolite comprising the following steps: (a) mixing amonite and saprolite with sufficient concentrated inorganic acid to form a salt with a non-soluble iron component of the ore; (b) at a constant temperature, for a time sufficient to cause hydrolysis of the dissolved iron (II) ions, at least a significant amount of either nickel or cobalt is dissolved into the leaching solution, so that iron-containing precipitates are formed (a) Water leaching the mixture of and; (c) recovering at least either nickel or cobalt compounds from the leach solution. The process of claim 1, wherein step (a) is carried out by first mixing the acid with limonite and subsequently adding saprolite. The method of claim 1 or 2, wherein step (a) at least a portion of the ore is broken to promote intimate mixing of the acid and the ore. The method of claim 2, wherein the saprolite is separately milled and then added to the mixture of limonite and acid. The method of claim 1, wherein the combined limonite and saprolite are pressed and crushed prior to mixing with the acid. The process of claim 1, 2, 4 or 5, wherein the water leaching of step (b) is carried out in an autoclave at temperatures up to about 150 ° C. The method of claim 6, wherein the water leaching of step (b) is carried out at a temperature in the range of about 95-105 ° C. 8. The process of claim 1, 2, 4 or 5, wherein the water leaching of step (b) is carried out at a temperature in the range of about 95-105 ° C. and a pressure in the range of about 15-70 psia. The process according to claim 4 or 5, wherein the water leaching of step (b) is carried out in two stages: the first stage carried out at atmospheric pressure and the boiling point temperature of the leaching solution and the second stage carried out at about 150 ° C. in the autoclave. Way. The method of claim 4 or 5, wherein the acid is selected from sulfuric acid, hydrochloric acid, and nitric acid or mixtures thereof. The method of claim 8, wherein the acid is a salt comprising sulfuric acid and the sulfate formed in step (a). The method according to claim 1, 2, 4 or 5, wherein recovering at least one of nickel or cobalt compounds from the leaching solution of step (c) comprises adding an ion exchange resin to the leaching solution without separating the solid / liquid in advance. And comprising one. The process according to claim 1, 2, 4 or 5, wherein the leaching solution is first separated from the precipitate prior to recovering at least one of the nickel or cobalt compounds from the leaching solution of step (c). The process according to claim 1, 2, 4 or 5, wherein a reducing agent is added during the water leaching step of step (b) to promote the dissolution of cobalt from the ore. The method of claim 14, wherein the reducing agent is selected from the group consisting of sulfur dioxide, hydrogen sulfide, soluble bisulfite and sulfite compounds or is a soluble iron (II) ionic compound. The process according to claim 1 or 2, further comprising curing the mixture of step (a) prior to the water leaching step of step (b).

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