Patent Publication Number: US-2006002835-A1

Title: Method for nickel and cobalt recovery from laterite ores by reaction with concentrated acid and water leaching

Description:
CROSS REFERENCE TO RELATED APPLICATION  
      This application claims the benefit of U.S. provisional patent application No. 60/583243 filed Jun. 28, 2004, the disclosure of which is hereby incorporated herein by reference. 
    
    
     FIELD OF THE INVENTION  
      The present invention relates to the hydrometallurgical processing of nickeliferous laterite ore, and in particular to a method for acid leaching both the limonite fraction and the saprolite fraction of such ores in a single process.  
     BACKGROUND OF THE INVENTION  
      Laterite ores are formed by the in-situ weathering of nickel-bearing ultramafic rocks near or at the surface of the earth in tropical environments by the action of naturally acidic meteoric waters over geologic time. They consist of a variety of clay, oxide and silicate minerals, some enriched in nickel and/or cobalt, and this distinguishes them from the other major class of nickel ores, the sulfide ores. The latter consist typically of sulfide minerals of iron, nickel and cobalt, often with copper and minor precious metals, and are associated with mafic-ultramafic magmatic intrusions in the earth&#39;s crust.  
      The weathering process typically creates a layered deposit, with the products of complete or most extensive weathering occurring near the surface and these grading into the products of lesser degrees of weathering as depth is increased and finally terminating in unweathered rock at some greater depth. The highly weathered layer usually contains most of its contained nickel microscopically distributed within very finely divided goethite particles. Goethite is an oxyhydroxide of ferric iron with the chemical formula FeOOH. This layer is usually given the name limonite, and typically contains a high proportion of iron.  
      Cobalt is usually associated with the limonite layer and is usually predominantly associated with oxidized manganese minerals (Mn(III) and/or Mn(IV) containing oxides and hydroxides), often called asbolane or manganese wad.  
      The lesser weathered layers typically contain increasing proportions of their contained nickel in various magnesium silicate minerals, such as, for example, serpentine. Serpentine is a silicate mineral of magnesium which has the chemical formula 3MgO*2SiO 2 *2H 2 O. It is believed that nickel substitutes for some of the magnesium in serpentine. Magnesium may also be substituted by other divalent metals, for example, ferrous iron (Fe 2+ ). There may be many other silicate minerals that also host nickel in the incompletely weathered zones. The partially weathered, high-magnesium bearing zone is often given the name saprolite, or garnierite. (“Garnierite” is also used to describe a particular apple-green colored magnesium-nickel silicate mineral of variable composition.)  
      In some deposits there is another zone typically located between the limonite and saprolite that consists predominantly of nontronite clays; these are silicates of magnesium, iron and aluminum that may also be nickeliferous. In most deposits located in the (current) tropics, the nontronite zone is largely absent.  
      It should be noted also that none of the weathering zones are homogeneous in mineralogical or chemical composition, nor is the boundary between the zones parallel to the earth&#39;s surface. However, there is usually a fairly sharp transition from ore of high iron and relatively low magnesium contents to ore of a relatively high magnesium content and lower, although variable, iron content, which occurs over vertical distances of 1 to 3 m within a laterite deposit.  
      For illustration purposes only, typical chemical compositions of limonite and saprolite are as follows: Limonite: 1.0-1.8% Ni, 0.05-0.3% Co, 35-50% Fe, 0.2-3.5% Mg Saprolite: 1.2-3.5% Ni, 0.02-0.07% Co, 7-20% Fe, 10-20% Mg  
      Each zone also contains typically significant concentrations of aluminum, manganese and chromium, as well as trace concentrations of other heavy metals such as copper and zinc in a variety of other minerals.  
      A challenging aspect of nickel recovery from laterite ores is that the nickel values typically can not be concentrated substantially by physical means, that is, so-called ore dressing techniques, prior to chemical processing to separate the metal values. This renders the processing of laterites expensive, and means to lower the costs of processing laterites have been sought for many decades.  
      Also, because of the distinct mineralogical and chemical composition of limonite and saprolite ores, these ores usually are not amenable to processing by the same process technique.  
      One known acid leaching process for nickel laterites is the so-called High Pressure Acid Leaching (HPAL) process (see, for example pages 437-453 in “The Winning of Nickel Its Geology, Mining and Extractive Metallurgy,” by Joseph R. Boldt, Longmans Canada Ltd. 1967). This process was first employed at Moa Bay in Cuba in the late 1950s and additional plants were constructed in Western Australia in the late 1990s.  
      The process utilizes sulfuric acid leaching at high temperature, typically 250° C., and high pressure; the associated steam pressure at 250° C. is approx. 570 psi. At this temperature, the nickeliferous minerals in the ore are nearly completely solubilized. The dissolved iron is rapidly precipitated as hematite (Fe 2 O 3 ) at the high temperature employed because this compound is largely insoluble even in slightly acidic solutions at this temperature. The nickel remains in solution and after cooling, the leach residue containing iron is separated from the nickel-bearing solution by thickening in a series of wash thickeners, a so-called counter-current decantation (CCD) circuit. Thus, the primary objective of the leaching process, which is separation of nickel from iron, is achieved.  
      A major disadvantage of the HPAL process is that it requires sophisticated high-temperature autoclaves and associated equipment which are expensive, both to install and to maintain. In addition, the HPAL process also consumes more sulfuric acid than is required to stoichiometrically dissolve the non-ferrous metals content of the ore because at high temperature most of the sulfate ions provided by sulfuric acid are tied up as bisulfate ions (HSO 4   − ). In other words, sulfuric acid (H 2 SO 4 ) only dissociates to release a single proton (H + ) at high temperature. On cooling and neutralization of the leach liquor the bisulfate ions decompose to sulfate (SO 4   2   − ) and another proton. The latter proton (acid) is therefore not utilized fully for leaching and results in excess sulfuric acid which must be neutralized, for example with limestone.  
      Another disadvantage of the HPAL process is that it is limited to treating largely limonite-type feeds because the presence of saprolite will cause a large, and often uneconomic, increase in sulfuric acid consumption due to the leaching of magnesium from saprolite. This is exacerbated by the bisulfate “shift” problem at high temperature, which is described above.  
      U.S. Pat. No. 4,097,575 describes an improvement to the HPAL process which constitutes roasting saprolite ore below about 820° C. in order to render the ore more reactive with sulfuric acid and then using the roaster calcine to neutralize excess acid in the discharge of an autoclave wherein pressure leaching of limonite ore occurs. Nickel contained in the saprolite ore is largely dissolved during this neutralization. The reported advantages of this process are that it better utilizes the sulfuric acid added during pressure leaching of limonite, it reduces the consumption of limestone or other costly neutralizing agents to treat the autoclave discharge liquor, and it achieves the capability of treating both the limonite and saprolite fractions of a typical nickel laterite orebody. Disadvantages of the process are that it still requires the use of expensive autoclaves for leaching limonite, and it requires a roasting process for saprolite ore, which is expensive both in capital and operating cost terms.  
      U.S. Pat. No. 6,379,636 B2 describes a further improvement to the process described in U.S. Pat. No. 4,097,575 wherein the saprolite roasting step is eliminated and the saprolite in “raw” form is used to neutralize the excess acid in the autoclave discharge solution. In addition, more acid could be added to the discharge to increase the amount of saprolite that could be leached. However, this process still requires the use of expensive autoclaves.  
      Several processes have also been described that utilize acid leaching at atmospheric pressure only, eliminating the disadvantages of pressure leaching described above. U.S. Pat. No. 3,793,432 describes an atmospheric leaching process for laterite ore, in which the ore is reacted with sulfuric acid at or below the boiling point and the precipitation of dissolved iron is achieved by the addition of an iron precipitating agent such as ammonium, sodium, potassium or lithium ions. Although not stated explicitly, all of the examples cited in the specification employed limonite ore samples, as evidenced by the high iron content and low magnesium content of the feed ore. While this process overcomes the disadvantages of pressure leaching, it has other disadvantages. First, the precipitation of iron would be as a jarosite compound, which is a thermodynamically unstable compound of iron that decomposes over time to release sulfuric acid, thus causing environmental problems. (Although jarosite is not stated explicitly it would be apparent to one skilled in the art that jarosite will precipitate at the conditions outlined in the examples.) Jarosite contains two moles of sulfate for every three moles of iron and thus this compound represents substantial excess consumption of sulfuric acid to provide the necessary sulfate ions.  
      Second, the nickel extractions from the ore were apparently relatively low. While extractions were not stated explicitly, based on the nickel content of the residue and the fact that the residue weight must be more than the weight of the original ore because jarosite contains a lower percentage of iron than the original ore and virtually all of the iron remained in the residue, nickel extractions were usually in the 60-65% range. Third, there is a requirement for very long leach times, of the order of 4-5 days. Fourth, there is a need to add relatively expensive iron precipitating agents such as potassium carbonate, sodium carbonate or the like.  
      U.S. Pat. Nos. 6,261,527 B1 and 6,680,035 B2 describe another atmospheric leaching process in which limonite ore is first “totally” leached with strong sulfuric acid, i.e. both nickel and iron are substantially dissolved from goethite, and then saprolite ore is leached in the resulting limonite leach slurry while simultaneously precipitating iron as jarosite by the addition of a jarosite precipitating agent. This process also has the disadvantages of producing jarosite, requiring separate mining and preparation of the limonite and saprolite fractions of the ore, and being limited to a narrow range of saprolite to limonite ratios. The latter disadvantage is due to the fact that the quantity of saprolite that can be leached effectively is determined by the quantity of ferric iron in the limonite leach solution.  
      WO 03/093517 A1 describes an improvement to this process, which constitutes eliminating the addition of a jarosite-forming ion such as sodium, potassium and ammonium, and causing the iron to precipitate as a compound other than jarosite, such as goethite. The process overcomes the disadvantages of jarosite, but sulfuric acid consumption was 0.59 to 0.87 tonnes per tonne of ore in the examples cited, and was over 0.72 tonnes per tonne of ore in nine of the eleven examples cited.  
      The processes described in U.S. Pat. Nos. 6,261,527 B1 and 6,680,035 B2 and WO 03/093517 A1 are based on the fact that goethite is more refractory to acid leaching than typical saprolite minerals, such as serpentine. This has been demonstrated by other researchers (see, for example: John H. Canterford, “Leaching of Some Australian Nickeliferous Laterites with Sulfuric Acid at Atmospheric Pressure,” Proc. Australasian Inst. Min. Metall., 265 (1978), 19-26; N. M. Rice and L. W. Strong, “The Leaching of Lateritic Nickel Ores in Hydrochloric Acid,” Canadian Metallurgical Quarterly, 13(3)(1974), 485-493; and FIG. 5 of U.S. Pat. No. 5,571,308). Thus, only saprolite can be used effectively in the second stage of leaching where iron precipitation occurs simultaneously. This is because the acidity of the solution must be relatively low to enable the precipitation of jarosite and even lower to enable the precipitation of goethite or other hydrolysis products of ferric iron. The goethite contained in limonite would leach very slowly under these conditions. Limonite (principally goethite) is thus leached in an initial stage with a relatively high acid concentration and both-iron and nickel are brought into solution.  
      U.S. Pat. No. 3,244,513 describes a process in which laterite ore, largely of limonitic type (defined as &gt;25% iron), is mixed with concentrated sulfuric acid in the presence of limited water, then the mixture is roasted at temperatures from approx. 500 to 725° C. in order to cause preferential sulfation of the nickel, cobalt, magnesium and manganese values over iron. Subsequent water leaching results in high extraction of nickel and cobalt and low extraction of iron to solution. The advantages of this process are that it does not require expensive autoclaves to effect leaching. The primary disadvantage is that it requires an expensive roasting step.  
      U.S. Pat. No. 4,125,588 describes a similar process as that described in U.S. Pat. No. 3,244,513, except that the roasting step is omitted and the mixing of ore with concentrated acid is done in a carefully controlled fashion wherein ore is first mixed with concentrated sulfuric acid in the absence of significant water and then water is added in controlled amounts to initiate sulfation of the ore, and then finally the mixture is leached with further amounts of water. The advantage of this process is that it eliminates the roasting step required in U.S. Pat. No. 3,244,513. However, the process also has significant disadvantages.  
      One disadvantage is that the ore used should contain no more than 1% moisture, which means that in most cases the ore must be dried, as in-situ moisture contents of laterite ores are usually 20% or more. A second disadvantage is that the process does not provide selective dissolution of nickel over iron, as illustrated in all of the examples cited in the patent (&gt;90% iron extraction). The separation of nickel from iron in solution usually results in additional nickel losses.  
      In addition, this process is only applicable to ores containing “large amounts of magnesia and silica,” i.e. the saprolitic or garnieritic ores. While the precise ore composition is not cited in all of the examples, it is evident from the data given that the ore contained from 3 to 4 times as much magnesium as iron, clearly indicating that limonitic ores (iron/magnesium ratios of from approx. 10 to 90 by weight) were not considered. Moreover, acid consumption is very high in this process, from about 0.9 to 1.1 tonnes of sulfuric acid per tonne of ore.  
      U.S. Pat. No. 3,093,559 describes another process for treatment of lateritic ores with relatively concentrated sulfuric acid (from about 25 to 50% sulfuric acid). In this process sufficient acid is added to cause sulfation of most or all of the metal values including iron. Iron is then separated from nickel by evaporating the leach solution to dryness and roasting the resulting salts at 975 to 1050° F. in order to convert the iron to hematite. Subsequent releaching of the calcine brings nickel into solution, leaving iron in the residue. As with various processes described above, the requirement for a high temperature roasting step is a significant disadvantage in this process.  
      U.S. Pat. No. 2,899,300 describes a process in which moist laterite ore is treated with concentrated sulfuric acid and the mixture is then dried by baking at a temperature between 100° C. and 150° C., preferably 125° C., before water leaching to dissolve the metal values in the ore. The baking step is a significant disadvantage because it requires substantial heat energy to evaporate the contained water in the ore/acid mixture. Furthermore, as illustrated by the example given in the specification, nickel dissolution is relatively low, while iron dissolution is relatively high, ˜77% and 53%, respectively. Nickel extraction can reportedly be increased by adding additional sulfuric acid to the first stage residue and carrying out a second sulfation and water leaching step, but this adds complexity to the process and will not improve the iron/nickel separation achieved in the process.  
      U.S. Pat. No. RE37,251 describes a process for the pressure leaching of non-cuprous ores and concentrates, including nickel laterite ores, using an acidic solution of bisulfate and sulfate ions along with halogen ions, e.g. chloride ions, and oxygen. According to the specification, the temperature and pressure required are 225° C. and 450 psig O 2 . Given that the steam pressure at 225° C. is approximately 370 psi, the total pressure would be in the range of 820 psig. These conditions are fairly similar to the high pressure acid leach conditions and would therefore require the use of expensive high pressure autoclave systems, as described above.  
      The object of the present invention is to obviate or mitigate the disadvantages of these known processes. It is a further object to provide a process to acid leach a mixture of limonite and saprolite ores, such as would be produced by bulk mining of a typical laterite orebody without any subsequent separation of ore types, under atmospheric or low pressure conditions, while achieving high extractions of nickel and cobalt and very low ultimate extraction of iron.  
     BRIEF SUMMARY OF THE INVENTION  
      The present invention provides a process for the efficient leaching of nickel and cobalt from limonite and saprolite ore, such as could be produced by bulk mining of a typical nickel laterite deposit, in a two stage process, the first stage consisting of mixing and reacting the ore with concentrated mineral acid, and the second stage consisting of preparing a slurry of the acid/ore mixture in water and leaching the mixture to dissolve nickel and cobalt. Iron is efficiently separated from nickel and cobalt in the solid leach residue primarily as an oxide or hydroxide of ferric iron other than jarosite.  
      The process may also include curing of the mixed ore and acid prior to water leaching, and the ore is first crushed before it is mixed and reacted with the acid. Preferably, the saprolite is crushed separately and then blended with the limonite. More preferably, the acid is mixed first with the limonite, and the saprolite is added subsequently.  
      The water leaching is advantageously carried out at a modestly elevated temperature. For atmospheric leaching, the temperature is preferably within the range 95-105° C. Alternatively, faster leach times may be obtained by leaching in an autoclave with a temperature up to about 150° C. The corresponding pressure is in the range of up to about 70 psia and is approximately equal to the saturated steam pressure at the leaching temperature.  
      The acid used is preferably selected from sulfuric, hydrochloric, or nitric acid, and is more preferably sulfuric acid. A mixture of any of these acids may also be used.  
      Advantageously, a reductant, such as sulfur dioxide, hydrogen sulfide, soluble bisulfite and sulfite compounds, or soluble ferrous iron compounds, is added during the water leaching to enhance dissolution of cobalt.  
      The process of the present invention avoids the high cost of high-pressure autoclaves and reduces the production of jarosite compounds. In some embodiments, it also avoids the need to separately mine and separately treat limonite and saprolite ore types, and it may allow processing of ore over a wide range of saprolite to limonite ratios.  
      It has been found that the present invention can achieve at least about 80% nickel extraction and as much as 95% or more cobalt extraction, with less than about 15% iron extraction.  
    
    
     BRIEF DESCRIPTION OF THE DRAWINGS  
       FIG. 1  is a flow sheet showing in simplified form one embodiment of the process of the present invention.  
       FIG. 2  shows another embodiment of the process of the present invention in which some of the leach residue is recycled in order to provide seed for iron precipitation.  
       FIG. 3  shows a third embodiment of the process of the present invention in which all of the required acid is first mixed with the limonite ore and then the saprolite ore is mixed with the resulting limonite/acid mixture.  
       FIG. 4  shows a fourth embodiment of the process of the present invention in which all of the required acid is first mixed with the limonite ore, and then crushed saprolite ore and water are added to the resulting limonite/acid mixture, after which the resulting mixture is ground and then subjected to atmospheric leaching. 
    
    
     DETAILED DESCRIPTION OF THE INVENTION  
      The present invention provides an improved process for the extraction of nickel and cobalt from nickeliferous laterite ore, while rejecting most of the iron contained in the ore in a solid leach residue. The process does not require the prior separation of the limonite ore from the saprolite ore and can treat a mixture of the two ore types. However, where it is convenient to mine limonite and saprolite ore separately, high nickel extraction can be achieved by first reacting acid directly with the limonite ore and then adding the saprolite ore.  
      Referring to  FIG. 1 , run-of-mine laterite ore, consisting of a mixture of limonite and saprolite, is first crushed to approximately 5 to 10 mm top size. The crushed ore is then mixed with a concentrated mineral acid, chosen from the group sulfuric, hydrochloric and nitric, or a mixture of these, in a suitable device such as a pug mill. It is not necessary to dry the ore, which typically contains substantial free moisture in the run-of-mine condition, prior to the addition of the concentrated acid.  
      The quantity of acid added is at least that required to stoichiometrically dissolve the soluble non-ferrous metals in the ore (but not the iron), i.e. most of the nickel, cobalt, magnesium, aluminum, copper, zinc, and a small portion of the chromium, in the ore during the subsequent leaching stage. A small excess of acid is added to provide some free acidity in the subsequent leaching stage, so as to leach a small proportion of the iron, and to ensure maximum extraction of nickel and cobalt. Acid addition is limited to ensure that final iron extraction from the ore slurry is minimized. In some cases, some of the magnesium and aluminum may be insoluble and this should be taken into account to determine the precise acid addition.  
      Water addition during the acid mixing process is minimized to the extent possible to provide the highest concentration of acid for reaction with the ore minerals. Water addition is only necessary if the ore/acid mixture stiffens so much that it can not be blended completely or easily handled otherwise. The addition of concentrated acid to moist ore results in the generation of substantial heat, raising the temperature of the mixture as high as, and even above, the boiling point of water and causing significant water evaporation. Additional water can be added during the blending process, if desired, to control the consistency of the acid/ore mixture to that of a fluid paste. If no water is added, depending on the moisture content of the ore and the net amount of heat generated, it may also be possible to form a dry, powdery reaction product. (Intermediate between the conditions of a fluid paste and a dry, powdery material, a stiff, toffee-like mixture may be formed, which may be more difficult to handle.)  
      Ideally, the mixture will have either a fluid-paste consistency or a dry, powdery consistency to permit easy handling. The resulting acid/ore blend, or “pug”, may be allowed to “cure” at ambient temperature for a sufficient time as to allow substantial reaction between the mineral acid and the mineral constituents of the ore. This can be done by stockpiling the pug on a prepared impermeable pad and allowing the pug to stay on the pad for as much as several days before reclaiming the material and subjecting it to the next phase of the process.  
      To facilitate stockpiling and reclaiming, it may be desirable to form small, discrete portions of the pug by, for example, extruding or pelletizing the material before stockpiling. However, it has been found that satisfactory results may be achieved with minimum or no curing time. In this case, the acid/ore mixture may be directly blended with water. Subsequent grinding may not be necessary if the ore was ground prior to adding acid. Longer curing may give slightly better nickel extraction and lower iron extraction in the subsequent water leaching step.  
      After curing, the pug is ground to a particle size sufficient for leaching in agitated tanks. Conveniently, the water required to form a leach slurry can be added prior to grinding so that this step can be carried out in the wet mode. The pug should be ground to a particle size sufficient to enable off-bottom solids suspension of most of the particles in the slurry without requiring excessive agitation power input. It is desirable to make the slurry as dense as possible consistent with good mixing during leaching in order to minimize water requirements and produce the most concentrated nickel and cobalt-bearing pregnant leach solution.  
      The resulting leach slurry is heated, as required, in a batch leach reactor or series of leach reactors if the process is to be carried out continuously. The temperature of the leach slurry is preferably maintained at or near the normal boiling point of the leach liquor, typically 95-100° C. Live steam injection, or other means of adding thermal energy, can be used for this purpose. Some of the heat required is provided by the heat of solution of excess acid and metal salts formed during the acid blending step. In addition, further amounts of heat may be recovered during the pugging process itself by cooling the pugging reactor with the water to be used in leaching, thus preheating the water, or by recovering steam generated during the acid ore reaction.  
      The leach mixture is allowed to react until most of the nickel and cobalt in the ore have dissolved and most of the iron is present in the solid leach residue. The agitated leach retention time is of the order of 12 to 48 hours. Satisfactory results are typically obtained in about 15 hours or less. In some cases, it may be desirable to add a reducing agent in controlled amounts to the leach slurry in order to enhance the leaching of cobalt associated with oxidized manganese minerals, e.g. asbolane, in the ore. For example, controlled additions of sulfur dioxide can be added to improve cobalt (and manganese leaching), as will be apparent to those skilled in the art. Other reductants capable of reducing trivalent and tetravalent manganese oxides to the divalent state may also be used.  
      The leaching process can be accelerated significantly by employing a temperature above the boiling point. For example, by heating the leach slurry to about 150° C. in an autoclave, the leach time can be reduced to about 1 hour duration. Further increases in temperature may be beneficial but it is an aim of this invention to eliminate the complexity associated with high pressure leaching. It should be noted that the pressure required to leach the pug plus water slurry at 150° C. is only about 70 psia (the steam pressure at 150° C.). At this pressure, the slurry can be pumped to the autoclave with conventional centrifugal pumps and the pressure letdown system can be a very simple single stage valve or choke. Suitable autoclave equipment to be used in the present invention is vastly different from that required for high pressure leaching at temperatures of around 250° C. and pressures of around 450-500 psia, as will be apparent to those skilled in the art.  
      The use of a temperature above the boiling point may also provide a higher nickel/iron ratio in solution, which is advantageous with respect to downstream processing of the leach solution. This is because in most cases virtually all iron must be removed from solution before effecting nickel and cobalt recovery. Usually, the residual iron in solution is removed by adding a base, for example calcium carbonate, to the leach slurry and precipitating iron oxyhydroxide compounds. Some nickel may co-precipitate with the iron resulting in losses of the pay metal and in addition the neutralizing agent represents an additional operating cost of the process. A further advantage of the use of higher temperature is an improvement in the solid/liquid separation properties of the final leach slurry, with higher settling rates being achieved with a higher leaching temperature.  
      In some instances, it may be advantageous to employ both an atmospheric pressure leaching stage and a subsequent moderate pressure leaching stage, in series. Sulfur dioxide, or another reductant, for the leaching of oxidized manganese minerals and contained cobalt and nickel values is most conveniently added during the atmospheric leaching stage. A subsequent pressure leaching stage can still be used to achieve the advantages noted in the preceding paragraph.  
      Nickel extractions of at least 80% and as high as 90-95%, with corresponding iron extractions as low as 5-10%, are obtained with the current invention. In addition, the leach residue typically contains no more than about 1-2% of sulfur, indicating that the iron compounds in the residue are primarily not of the jarosite type.  
      This leaching process of the present invention, in the case where sulfuric acid is used as the lixiviant, differs from the prior art, for example U.S. Pat. No. 3,244,513, in that the latter employs a roasting step after acid blending in order to convert any iron sulfate formed during acid blending to iron oxide. Sulfur trioxide (SO 3 ) gas is released during this conversion and this reacts with non-ferrous metal oxides, e.g. NiO and MgO, to form non-ferrous sulfate salts. Thus, the non-ferrous metals are preferentially sulfated in the process of the prior art.  
      However, is has surprisingly now been found that the roasting step, which is likely to be capital cost intensive and require substantial, expensive fuel energy for achieving the desired roasting temperatures, e.g. 500-700° C., is not necessary to achieve selective sulfation of the non-ferrous metals, thus resulting in a pregnant leach solution after water leaching of the ore/acid blend, which contains most of the soluble non-ferrous metals, particularly nickel and cobalt, and relatively little of the iron. All that is required is to provide sufficient time and temperature during water leaching to effect the desired reactions. Elimination of the roasting step thus represents a major advantage over the prior art.  
      After water leaching, the leach slurry is subjected to solid/liquid separation by filtration or thickening to produce a pregnant leach solution containing most of the nickel and cobalt contained in the ore and a solid residue containing most of the iron in the ore. Advantageously, thickening is carried out in a series of thickeners with counter-current flow of a wash water stream and the leach slurry in order to wash most of the entrained metal values out of the leach residue, a method called counter-current decantation (CCD). The metal values report preponderantly to the thickener overflow of the first thickener, which is the pregnant leach solution.  
      The pregnant leach solution proceeds to nickel and cobalt recovery by methods known to those skilled in the art, such as solvent extraction, ion exchange, sulfide precipitation using sulfiding agents, e.g. hydrogen sulfide, or hydroxide precipitation, using for example magnesia as the precipitating agent.  
      The nickel and cobalt can also be recovered from the leach slurry without prior solid/liquid separation, using the resin-in-pulp process. In this process, an ion exchange resin which extracts nickel and possibly cobalt is added directly to the leach slurry. After the extraction is complete, the resin is separated from the nickel-depleted leach slurry by screening. After washing the resin to remove solids, the nickel can be eluted from the resin with a fresh acid solution.  
      Prior to or during nickel and cobalt recovery by one of the aforementioned methods, the leach solution may be neutralized with a base, such as calcium carbonate, magnesium oxide, sodium carbonate or the like, to neutralize free acidity remaining from the leach process and precipitate small amounts of ferric iron, aluminum, and chromium, while minimizing co-precipitation of nickel and cobalt. This process may be carried out in a single or multiple steps separated by intermediate solid/liquid separations.  
      In one embodiment of the invention, the first stage of neutralization may be carried out prior to separating the leach residue from the leach solution. The combined leach and neutralization residue may then be separated from the partially neutralized leach solution by filtration or thickening, as described above. A second stage of neutralization may then still be desirable, depending on the method selected for nickel and cobalt recovery from the pregnant leach solution. After this second stage of neutralization, the resulting neutralization residue may be separated from the neutralized leach solution by filtration or thickening. This second-stage neutralization residue is ideally returned to the first stage neutralization to re-dissolve any co-precipitated nickel and cobalt.  
      It has surprisingly been found that it is not necessary to separate the limonite from the saprolite in the laterite ore in order to treat them separately and distinctly in the leach process. Further, the process is not limited to any particular saprolite/limonite ratio, as long as at least a certain minimum saprolite/limonite ratio is met. The quantity of acid added to the ore may be adjusted for a range of saprolite/limonite ratios based on the non-ferrous metals content of the saprolite/limonite ore mixture.  
      It has surprisingly also been found that it is not necessary to add any iron precipitating agents to the leach slurry, as in U.S. Pat. No. 3,793,432. Addition of iron precipitating agents such as sodium, potassium and ammonium ions is a disadvantage in that they promote the formation of iron-bearing jarosite compounds, which are thermodynamically unstable, slowly decomposing over time and releasing sulfuric acid, which can cause the re-dissolution of metal impurities present in the leach residue and leading potentially to contamination of the environment. In addition, extra sulfuric acid is required during leaching to satisfy the requirement for sulfate ions in the formation of jarosite.  
      In the process of the present invention, where sulfuric acid, for example, is added to the ore, nickel, cobalt and iron, as well as other non-ferrous metals in the ore, are unselectively sulfated, that is, converted to sulfate salts, during the acid/ore blending step. The sulfates dissolve easily in water during the subsequent leach step. However, nickel and cobalt extraction are incomplete initially because the quantity of acid added initially is not sufficient to convert all of the metals, including iron, to the sulfate salts. Thus, considerable iron goes into solution along with nickel, cobalt, magnesium, aluminum, chromium, copper and zinc. However, by continuing the leach process after the initial sulfate salt dissolution phase, the iron content of the solution decreases and further nickel and cobalt are extracted from the solid phase. The exact iron leaching and re-precipitation reactions are unknown, but the sulfur content of the final residue is low, usually less than 2%, indicating that jarosite does not form to a significant extent.  
      In one preferred embodiment of the invention, an iron-bearing “seed” material is added to the leach slurry at the start of the water leach, in order to accelerate the precipitation of dissolved ferric iron and the extraction of remaining nickel and cobalt from the solid phases. The surfaces of the seed particles provide low-activation energy sites for hydrolysis and precipitation of iron, for example as ferric hydroxide, goethite, or hematite. This seed material is ideally a portion of the final leach residue itself, which contains precipitated iron compounds. One method of carrying out the process with seed recycle is shown in  FIG. 2 .  
      In a further embodiment of the invention, the limonite portion of the ore is first blended with the concentrated acid, the amount of acid being calculated in the same manner as discussed previously, following which the saprolite portion of the ore is added to the limonite/acid mixture. The saprolite ore may be added to the limonite/acid mixture either as crushed saprolite or as ground saprolite. In the former case the ore/acid mixture is ground prior to water leaching. The grinding may be carried out without the addition of water. Advantageously, water is added to the ore/acid mixture and wet grinding is employed prior to atmospheric water leaching. In the latter case, the saprolite ore can be ground in the dry condition and added to the acid/limonite mixture along with water, or ground wet and added to the limonite/acid mixture as a slurry or filtercake. The resulting blended ore and acid is then water leached as described previously. This method provides the maximum opportunity to sulfate the goethite component of the laterite ore. A flowsheet for carrying out this embodiment of the invention is shown in  FIG. 3 .  
      The following examples illustrate the method of the present invention. The ore used in these examples came from a Central American laterite deposit. The limonite and saprolite fractions of the ore used in Examples 1-3 had the compositions given in Table 1. The saprolite ore was crushed to approximately −6.4 mm before use in the tests.  
                                       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 1  
      Approximately 1 kg of saprolite ore was wet ground in a ball mill to approx. 100% passing 100 mesh. The ground slurry was filtered to produce ground saprolite at 41% moisture content. 425.2 g of the moist ground saprolite was blended with 381.7 g of moist limonite ore to produce 500 g of leach feed material with a limonite/saprolite ratio of 1:1 on a dry basis.  
      The ore mixture was placed in a 4.5 liter narrow-necked glass bottle. The bottle was rolled at an angle inclined slightly to the horizontal at approximately 47 to 48 rpm on a bottle rolling device. 312.5 g of 96% sulfuric acid was added to the ore mixture in the bottle over a period of about 30 minutes. The ore and acid were blended for approximately 30 min after all of the acid had been added. At the completion of blending the ore and acid had formed a semi-fluid mass and the temperature had risen to between about 70 and 100° C.  
      The bottle was removed from the rolling device and the blend of acid and ore was allowed to cure at ambient temperature for approx. 72 hours. 622 mL of water were then added to the cured mass and the mixture was stirred until a uniform leach slurry was formed. The leach slurry was transferred to a 2-liter cylindrical reaction kettle equipped with a mechanical stirrer, 4 plastic baffles, and a tight-fitting lid equipped with a water condenser open to the atmosphere. The reaction kettle was heated by an external, electrical heating mantle.  
      The leach slurry was heated to and held at a temperature of approximately 96-99° C. for a period of 48 hours while stirring vigorously. After 5 hours of leaching, 129 g of finely ground, technical grade hematite was added to the reactor to act as a “seed” for the precipitation of iron. Samples of the leach liquor were taken periodically for chemical analysis. At the end of the 48-hour leaching period, the entire slurry was filtered. The filtercake was repulped twice in fresh water to wash out the entrained metal values. The cake was then dried and weighed. The dry solids, filtrate and the combined wash water were assayed separately.  
      The results of this test are given in Tables 2 and 3 below.  
                               TABLE 2                       Time (h)   [Ni] g/L   [Fe] g/L   [Mg] g/L   [H 2 SO 4 ] g/L                                                    1   8.8   70   27   &lt;0.5       2   8.0   49   27   &lt;0.5       5   9.5   24   35   &lt;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 
                 85.2 
                 11.7 
                 94.9 
                 — 
               
               
                 Extractions* 
               
               
                   
               
               
                   *Ni, Fe and Mg extractions based on final residue and solution weights, volumes and assays.    
               
            
           
         
       
     
      The results show that a high extraction of nickel was achieved while extracting very little of the iron present in the ore. In addition, the sulfur content of the residue was quite low and the Fe/S ratio was approximately 44:1. Less than 5% of the sulfate contained in the sulfuric acid deported to the residue. If all of the iron in the ore had formed jarosite, the Fe/S ratio in the final residue would have been approximately 4.3. In addition, X-Ray diffraction analysis of the residue indicated the presence of only hematite, goethite and a spinel phase originally present in the ore.  
      The kinetic data indicate that slightly more than 24-32 hours of leaching was required to approach the terminal iron concentration in the solution. The residual iron could be removed from solution by neutralization with a base, for example, limestone, without siqnificant losses of nickel to the solid phase.  
     EXAMPLE 2  
      A sample of 381.7 g of limonite ore was blended with 317.7 g of crushed saprolite ore and 95 g of water in a 2-liter flat-bottomed glass beaker. The water was added to simulate the expected moisture content of run-of-mine ore, recognizing that the samples had dried somewhat compared to their in-situ condition. 312.5 g of 96% sulfuric acid was added to the beaker over approximately 30 minutes. The acid was blended with the ore using a stirrer rotating at about 60 rpm. The acid addition was sufficient to give an acid to ore ratio of about 600 kg H 2 SO 4  per tonne of ore (dry basis).  
      The ore and acid formed a semi-fluid mass which was poured into a shallow pan for curing at ambient temperature for approximately 72 hours. Some of the mixture was not recovered during this operation. It was estimated that about 13% of the mixture was not recovered, based on weight recovery.  
      After this period of curing, the acid/ore mixture, which had hardened considerably, was broken into pieces and transferred to a small grinding mill. 300 g of water were added to the mill and the mixture was ground with stone media for approximately 1 hour to reduce the maximum particle size to about 100 mesh. Additional water was added to the mill to wash the slurry to a 2-litre leach reactor. 1,858 g of leach slurry was prepared in this way. The reactor was heated to a temperature of 95 to 105° C., while stirring, and leaching was allowed to continue for 44 hours. After 5 hours of leaching, 129 g of finely ground, technical grade hematite were added to the reactor to act as a “seed” for the precipitation of iron. Samples of the leach liquor were taken periodically for chemical analysis. At the end of the 48-hour leaching period, the entire slurry was filtered. The filtercake was repulped twice in fresh water to wash out the entrained metal values. The cake was then dried and weighed. The dry solids, filtrate and the combined fresh water were assayed separately.  
      The results of this test are given in Tables 4 and 5 below.  
                                   TABLE 4                                   Time (h)   [Ni] g/L   [Fe] g/L   [H 2 SO 4 ] g/L                                                            1   4.6   25.6   &lt;0.5           2   5.36   29.7   &lt;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 
                 81.1 
                 11.3 
                 94.7 
                 — 
               
               
                 Extractions* 
               
               
                   
               
               
                   *Ni, Fe and Mg extractions based on final residue and solution weights, volumes and assays.    
               
            
           
         
       
     
      This test simulated one of the preferred embodiments of the current invention, in that saprolite ore was not ground before blending with limonite and adding acid. Instead, the acid/ore mixture was ground after curing.  
      The results of this test were similar to those of Example 1, although nickel extraction was slightly lower.  
     EXAMPLE 3  
      This test was carried out in a fashion similar to that of Example 2, with the following exceptions. The amounts of limonite ore, crushed saprolite ore, water, acid, and hematite seed used during the acid blending and subsequent leaching process were 336.9 g, 280.3 g, 84 g, 275.8 g, and 113 g, respectively. The proportions of ore, water and acid were the same as in Example 2.  
      After blending the ore with acid, the mixture was transferred to a shallow pan and allowed to cure for only one hour before transfer to a grinding mill for wet grinding and subsequent water leaching. After one hour of curing the mixture was still fluid and could be poured into the mill. 1,776 g of leach slurry were prepared and leached exactly as in Example 2. Recovery of the acid/ore mixture was approximately 96%.  
      The results of this test are given in Tables 6 and 7 below.  
                                   TABLE 6                                   Time (h)   [Ni] g/L   [Fe] g/L   [H 2 SO 4 ] g/L                                                            1   5.53   22.7   &lt;0.5           2   6.37   24.2   &lt;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) 
                 Mg (% or g/L) 
                 S (%) 
               
               
                   
               
             
            
               
                   
               
            
           
           
               
               
               
               
               
            
               
                 Final Solution 
                 6.97 
                 10.2 
                 34.0 
                 — 
               
               
                 Final Residue 
                 0.37 
                 43.0 
                 0.45 
                 0.81 
               
               
                 Calculated 
                 84.2 
                 10.8 
                 95.2 
                 — 
               
               
                 Extractions* 
               
               
                   
               
               
                   *Ni, Fe and Mg extractions based on final residue and solution weights, volumes and assays.    
               
            
           
         
       
     
      The results of this test, which are similar to the previous examples, demonstrate that it is not necessary to cure the ore/acid mixture for a long period of time prior to leaching.  
     EXAMPLE 4  
      The ore used in this test had the compositions given in Table 8.  
                                           TABLE 8                                   % Ni   % Co   % Fe   % Mg   % Si   % Moisture                                                                Limonite   1.31   0.2   47.2   0.63   2.67   41.9       Ore       Saprolite   3.13   0.01   6.0   20.0   18.8   14.4       Ore                  
 
      409.6 g (wet basis) of the limonite ore was placed in a small porcelain ball mill charged with 0.6-1.9 cm porcelain balls. Prior to adding the ore, the mill and grinding media were preheated to approx. 100° C. This was done to simulate the thermal conditions that would be present if the process were carried out on a continuous basis. The mill was fitted with a plastic lid with a hole in its centre through which sulfuric acid could be added. The mill was rotated on the same rollers used in Example 1. 338.5 g of 96% sulfuric acid were added to the mill over a period of approximately 1.5 minutes and the limonite ore and acid were allowed to react for a period of 15 minutes. The temperature inside the mill was monitored with a hand-held laser thermometer during the reaction period. The temperature measured varied from 97 to 121° C., with the highest temperature recorded about 1 minute after the acid addition was completed. At the completion of the reaction period, the mill was removed from the rollers, the lid was removed and 306.1 g (wet basis) of crushed saprolite ore was added to the limonite/acid mixture along with 510 g of a 30 g/L Mg (as MgSO 4 ) aqueous solution. MgSO 4  solution was used to simulate the effect of using the barren solution remaining after purification and nickel/cobalt recovery as make-up water in the process. A tight-fitting porcelain lid was used to close the mill and the total ore/acid mixture was ground for approximately 60 minutes, until most of the solid particles present were less than 100 mesh in size.  
      Based on the weights and moisture contents of the ore samples, the overall saprolite/limonite ratio was 1.1 on a dry basis and the overall acid/ore ratio was 0.65 (based on dry ore and 100% H 2 SO 4 ).  
      The contents of the mill were then discharged into a reaction kettle identical to that used in Example 1. A coarse screen was used to capture the grinding media and 764 g of the 30 g/L Mg solution was used to rinse the ball mill and the grinding media. The rinse solution was added to the reaction kettle.  
      The reaction kettle was heated to 95-100° C. while stirring. Leaching was carried out for 24 hours. During the first 5 hours of leaching, sulfur dioxide gas was bubbled into the leach slurry to control the oxidation reduction potential of the slurry at between 540 and 600 mV (versus a saturated Ag/AgCl reference electrode). Samples of slurry were taken throughout the leaching period, filtered, the solids washed, and the solution and solids assayed.  
      At the completion of the leaching period, 20% limestone slurry was added to the leach slurry over approximately 2 hours to neutralize the leach slurry to a pH of 3.0 at 95° C. The neutralized leach slurry was sampled, then the entire slurry was filtered.  
      The results of the leaching period are given in Table 9.  
                               TABLE 9                       Time (h)   [Ni] g/L   [Co] g/L   [Fe] g/L   [H 2 SO 4 ] g/L                                                    1   4.92   0.229   60.8   &lt;0.5       2   5.33   0.248   50.0   &lt;0.5       4   6.22   0.290   33.9   &lt;0.5       5   6.52   0.303   26.5   &lt;0.5       24   7.36   0.34   14.4   5.0                  
 
      The solids at the completion of the leaching period assayed 0.21% Ni, 0.006% Co, 31.3% Fe, 0.65% Mg, 14.9% Si and 1.39% S. The Ni and Co extractions, calculated using a “silicon tie” (assuming zero silicon leaching, calculating leach residue weight using silicon assays of ore and residue, and using ore and residue weights and assays to calculate extractions), were approximately 93.1% and 95.5%, respectively. The iron extraction was roughly 9%.  
      After neutralization to pH 3, the solution assayed 6.23 g/L Ni, 0.29 g/L Co, and 3.24 g/L Fe. The solids assayed 0.22% Ni, 0.009% Co, 26.5% Fe and 11.7% Si. Approximately 74% of the iron in the leach solution was precipitated during neutralization; thus net iron extraction after neutralization was only about 2%. Approximately 2% of the nickel and 4% of the cobalt were co-precipitated with iron during neutralization.  
      This example illustrates several key features of one of the preferred embodiments of the invention: high nickel extraction can be achieved by first reacting sulfuric acid directly with the limonite ore and then adding the saprolite ore; the addition of a reductant to the leach slurry, in this case sulfur dioxide gas, enhances cobalt extraction; there is no need to cure the acid/ore reaction mixture to achieve high extractions; the reaction between sulfuric acid and ore is extremely rapid and extremely exothermic, such that very high temperatures can be achieved during the acid/ore reaction, thereby enhancing the reaction kinetics and permitting the equipment required to carry out the sulfation step to be very compact in size; and partial neutralization of the leach slurry following leaching can be used to eliminate most of the small proportion of iron which leaches with minimal nickel and cobalt losses.  
      It will of course be appreciated by those skilled in the art that many variations of the process would be possible within the broad scope of the present invention. Those skilled in the art will appreciate that the invention upon which the description is based may be utilized in other embodiments that carry out the purposes and fulfill the objects of the present invention.  
      The process of the current invention is widely applicable to nickel laterite ore bodies containing both limonite and saprolite, which includes the majority of such ore bodies. The saprolite/limonite ratio in the process may vary widely as long as a certain minimum saprolite/limonite ratio is used in the process. When using sulfuric acid, for example, the minimum ratio is determined approximately by calculating the amount of non-ferrous components of the saprolite ore which can be sulfated by the sulfuric acid that would be generated by hyudrolyzing and precipitating the “sulfatable” iron content of the limonite and saprolite ore as an oxide or hydroxide of ferric iron. The ratio may be any value equal to or greater than this minimum, providing that the acid addition is calculated as described previously, that is, sufficient acid to react with the sulfatable, non-ferrous components of the linionite and saprolite ore, plus a slight excess to account for a small amount of soluble iron and to drive the reactions to completion.  
      The above disclosure is intended to be illustrative while the scope of the invention is defined by the following claims.