METAL RECOVERY SYSTEM AS APPLIED TO THE HIGH PRESSURE LEACHING OF LIMONITIC NICKEL LATERITE ORES

CROSS REFERENCE TO RELATED APPLICATION This patent is related to, and claims priority from, United States provisional patent application Serial No. 60/817,707 filed on July 3, 2006, entitled MODIFICATIONS AND IMPROVEMENTS TO THE METAL RECOVERY SYSTEM AS APPLIED TO THE HIGH PRESSURE LEACHING OF LIMONITIC NICKEL LATERITE ORES, filed in English, which is incorporated herein by reference in its entirety.

FIELD OF INVENTION

The present invention is directed at a unique non-conventional method of recovering nickel and cobalt from leach solutions produced by pressure acid leaching limonitic nickel laterite ores in a manner such that the amount of "fresh" water required is minimized and virtually no used process waste waters are discharged to the external environment thereby protecting the external environment from intrusion of extraneous chemical elements and compounds.

BACKGROUND OF THE INVENTION

In recent decades there has been a strong shift in metallurgical technologies as applied to the extraction of metals from ores, wherein pyrometallurgical processes are being replaced in whole or in part by hydrometallurgical processes. This is particularly true with regard to nickel laterite ores which in their natural state can contain as much as 45% of free moisture. Pyrometallurgy would require drying of such ores in preparation for furnacing, while hydrometallurgy would take advantage of the fact that a good portion of the water that is required for leaching comes with the ore itself. The net result is that the energy required to extract the nickel and cobalt from laterite ores by hydrometallurgical processes is drastically reduced as compared to pyrometallurgical processing. Conventional thinking also assumed that hydrometallurgy can be

conducted in a more environmentally sound manner as compared to pyrometallurgy.

However, the conventional acid pressure leaching methodologies which require large quantities of "fresh" water in preparing feed pulps for the autoclave reactors, also have large quantities of spent process waste waters, i.e., large quantities of excess "barren" solutions to dispose of. The so-called "barren" solutions are not in fact barren but contain elements and compounds leached from the ores that need to be disposed of, after the recovery of the nickel and cobalt values. The problem stems from the use of limestone (calcium carbonate) as the most practical and economical means of carrying out the primary, partial neutralization for the removal of iron and other elements such as aluminium and chromium from the leach solutions emanating from the autoclaves. As a result, the solutions in the conventional counter current decantation (CCD) system employed for the recovery of the nickel and cobalt bearing solutions, and their separation from the leached tailings, and in particular the barren solutions, are laden with dissolved calcium sulphate. Such barren solutions cannot be used for preparation of feed pulps for the autoclaves because, upon heating, solid calcium sulphate (gypsum) would precipitate out of such solutions and clog up the feed preheating system ahead of the autoclave.

Accordingly, the conventional acid pressure leaching circuits rely on the use of large quantities of fresh water for make-up water in the preparation of the feed pulp; and of necessity have correspondingly large quantities of so-called barren solutions that need to be disposed of. There is a growing objection in many jurisdictions to the disposal of such solutions to the external environment. In certain cases where the hydrometallurgical process is being carried out in proximity of the sea, the practitioner has been permitted by the host country to dispose of its process waste waters by discharging into the sea; however, in today's strong focus on protecting the external environment, such practice is not permitted by many jurisdictions. Furthermore, there are many locations in the world where the mineral deposit is virtually land-locked, and disposal of process waste waters becomes a critical factor in the viability of launching a commercial operation.

Another disadvantage of the conventional CCD circuit is that barren solution used for washing out of the nickel-cobalt values, must be acidified with acid to maintain a pH low enough so that nickel and cobalt do not precipitate out of solution and be lost to the solid tailings. This excess acid must eventually be neutralized with further limestone additions.

United States Patent No. 6,391 , 089, issued to Curlook (2002), teaches that by recirculation of a portion of the "mother" liquor emanating from the autoclaves, back to feed preparation that the use of fresh water can be largely, if not completely, replaced by the addition of mother liquor. However, this patent employs conventional CCD circuitry for separating out a pregnant solution carrying the dissolved nickel and cobalt values, see Figures 2 and 3 of US Patent 6,391 , 089.

In normal practice, the conventional CCD circuit employs at least seven thickener separators to maximize the recovery of nickel and cobalt pregnant solutions at levels of at least 98%. Furthermore, the number of thickeners (solid-liquid separators) required is more or less independent of the throughput rate if one strives for the same degree of nickel and cobalt recovery. Higher throughput rates require either larger thickeners or additional parallel lines. The final tailings and barren solutions need to be neutralized first with limestone followed by treatment with lime to insure low levels of base metal traces in these waters going firstly to a tailings disposal pond, and subsequently the excess process waste waters need to be discharged to the external environment. Construction of tailings ponds themselves present potential environmental hazards as they necessitate construction of earth dams that face a history of failures around the world, with drastic and severe consequences including environmental damage. The ideal system would best meet jurisdictional regulations by not having any excess used water being emitted to the external environment. Therefore it would be very advantageous to provide a method of recovering nickel and cobalt from leach solutions produced by pressure acid leaching limonitic nickel laterite ores in a manner that the amount of "fresh" water required is minimized and no waste water produced in the recovery of nickel and cobalt is discharged to the external environment

thereby protecting the external environment from intrusion of extraneous chemical elements and compounds.

There are two common misconceptions with regard to the removal of iron from pregnant solutions derived from leaching of nickel laterite ores. One such misconception, held even by some who are skilled in the art

(and also prevalent in some published textbooks), is that much of the iron present in saprolitic minerals of nickel laterite ores occurs as ferrous iron; and the other misconception, professed even by many scholars, is that precipitates of iron hydroxides will bring down certain quantities of nickel and cobalt if present, as "co-precipitants". The present invention proves otherwise.

SUMMARY OF THE INVENTION

The present invention relies on the "free" moisture that is inherent in the wet raw limonitic nickel laterite ores which usually contain between about 35% and about 45% H ₂ O by weight of the wet ore, and on the use of recycled "mother" liquor emanating from the autoclaves to provide the total make-up water required for pulping the laterite ores. The mother liquor is produced in a dedicated first thickener that separates out a portion of the mother liquor for recycle, while the remainder of the mother liquor and the leached tailings, (together with reverted filtrate and a portion of thickener unflow recycle discussed hereunder), pass onto primary partial neutralization with limestone pulp to pH of about 4. The excess acid in the leach solution is reacted with the limestone to produce gypsum, the iron is precipitated as ferric hydroxide and the aluminum and chromium are also precipitated as hydroxides.

This partially neutralized pulp containing the extracted nickel and cobalt in solution, the leached tailings and the precipitates, passes on to a tailings thickener from whence a first portion of pregnant solution is decanted, and thickened solids containing a second portion of pregnant solution pass on to vacuum filtration to yield a filter cake for "dry" placement at mined out sites; while the filtrate is reverted to the partial neutralization stage. (Direct placement of filter cake is commonly referred to, in the industry, as "dry" placement). A portion of the thickener

underflow is recycled to the partial neutralization reactors to provide "seed" materials for the freshly formed precipitates, in order to improve settling and densification of the solids and to enhance subsequent filtration rates. The present invention, by employing single-stage filtration instead of a multi-stage CCD circuit, yields a washed tailings filter cake containing all the solid waste compounds, i.e., the leached ore, the iron hydroxide precipitate, the gypsum precipitate and other hydroxide wastes. This composite tailings can be placed in impoundment areas but preferably in mined out areas, with the only outflows of waters being the normal rainfall which can be collected/captured and returned to processing as wash water. Surprisingly, the pregnant solutions with their nickel and cobalt values can be recovered by this single-stage filtration at levels equivalent to those achieved by the conventional seven-stage CCD circuits.

In producing an intermediate nickel-cobalt product for subsequent shipping to a refinery, it is generally advantageous to employ pressure filtration at this final stage of processing to minimize the quantity of water accompanying the nickel-cobalt cake. The filtrate from the metal recovery circuit at a pH of between 8.0 and 8.5 serves as an ideal wash water for washing the final filtered tailings. Thus, in an embodiment of the invention there is provided process of leaching a nickel and cobalt containing predominantly limonitic portion of a laterite ore profile, comprising the steps of: a) preparing a feedstock of a predominantly limonitic portion of a laterite ore containing nickel and cobalt; b) pulping said feedstock with a liquid to produce a pulped ore; c) adding an effective amount of sulphuric acid to the pulped ore to produce a sulphuric acid solution, agitating and leaching said feedstock in said sulphuric acid solution at an elevated temperature under pressure for a selected period of time whereby metal oxides are leached from said ore to produce a leach pulp; d) separating said leach pulp into a mother liquor solution and a first thickened leach pulp, wherein said liquid includes a selected amount of said mother liquor solution;

e) partially neutralizing said first thickened leach pulp by mixing it with a limestone (calcium carbonate) pulp and a liquid filtrate containing residual nickel and cobalt values produced in step j) to form a partially neutralized first leach pulp; f) separating said partially neutralized first leach pulp into a pregnant solution containing sulphates of all extracted metal values and a second thickened pulp wherein said second thickened pulp includes a tailings fraction which includes leached ore, precipitated gypsum and precipitated hydroxides of iron and other impurity elements present and some pregnant solution; g) treating a selected amount of said pregnant solution with precipitating agents to yield barren solution and a third thickened pulp containing precipitated metal values and residual barren solution; h) separating said third thickened pulp from said barren solution; i) filtering said third thickened pulp to produce an intermediate nickel-cobalt product separated from said residual barren solution present in said third thickened pulp; j) filtering said second thickened pulp to produce a tailings filter cake which includes said leached ore, precipitated gypsum and precipitated hydroxides of iron and other impurity elements present in addition to a portion of the pregnant solution, washing said tailings filter cake with said barren solution produced in steps h) and i) to produce said liquid filtrate containing residual nickel and cobalt values and separating the washed tailings filter cake from said filtrate which is mixed with said first thickened leach pulp and the limestone (calcium carbonate) pulp in step e), thereby obviating the need to dispose of any excess barren solution to the external environment; and k) impounding said washed filter cake.

A further understanding of the functional and advantageous aspects of the invention can be realized by reference to the following detailed description and drawings.

BRIEF DESCRIPTION OF THE DRAWINGS

The process for acid leaching of nickel and cobalt containing laterite ores in accordance with the present invention will now be described, by way of example only, reference being had to the accompanying drawings, in which:

Figure 1 is a block flow diagram of the overall process in accordance with the present invention; and

Figure 2 is a more detailed mechanical flow diagram showing the steps of the present method; and Figure 3 is a plot of % nickel recovery versus the number of displacements of cake water (gypsum + iron hydroxide cake) demonstrating the efficacy of the vacuum filtratration mode of recovering nickel values as virtually iron-free pregnant solutions, wherein wash water in amounts of 2 displacements of the liquid in the filter cake, is used.

DETAILED DESCRIPTION OF THE INVENTION The methods described herein are directed, in general, to embodiments for the metal recovery system as applied to the high pressure leaching of limonitic nickel laterite ores. Although embodiments of the present invention are disclosed herein, the disclosed embodiments are merely exemplary and it should be understood that the invention relates to many alternative forms, including different shapes and sizes. Furthermore, the Figures are not drawn to scale and some features may be exaggerated or minimized to show details of particular features while related elements may have been eliminated to prevent obscuring novel aspects. Therefore, specific structural and functional details disclosed herein are not to be interpreted as limiting but merely as a basis for the claims and as a representative basis for enabling someone skilled in the art to employ the present invention in a variety of manner. For purposes of instruction and not limitation, the illustrated embodiments are all directed to embodiments for the metal recovery system as applied to the high pressure leaching of limonitic nickel laterite ores.

As used herein, the term "about", when used in conjunction with ranges of dimensions of particles or other physical properties or

characteristics such as concentrations, temperatures etc., is meant to cover slight variations that may exist in the upper and lower limits of the ranges of dimensions so as to not exclude embodiments where on average most of the dimensions are satisfied but where statistically dimensions may exist outside this region. It is not the intention to exclude embodiments such as these from the present invention.

The advantages of the present invention , and the salient differences from conventional metals recovery systems employed in acid pressure leaching of nickel laterite ores, can better be appreciated by reference to Figure 1 and Figure 2. The following is a description of each of the steps in the present process and the numbers in square brackets below correspond to the process steps shown in circled numbers in Figures 1 and 2.

STEP [I]

Feed from the Mine

The raw run-of-mine ore is typically passed over a grizzly to reject +25 cm and greater rock boulders, the less than -25 cm ore is then crushed to about -10 cm size and screened at about 6 cm to reject further low-nickel rock material and ore less than 6 cm is passed as feed to the processing plant. This less than 6 cm ore can be trucked, sluiced or pumped to the processing plant site. In the latter two cases a dewatering system would need to be provided with the drained water recycled to the mine site for reuse.

STEP [2]

Feed Preparation

The feed preparation in step [2] of Figure 1 could include a further crushing stage ahead of the pulping, screening, grinding and pumping to feed storage tanks ahead of the autoclaves. A feature at this stage is the addition of the recycled mother liquor to make-up the liquid necessary for pulping of the screened and ground ore. This mother liquor is produced in step [4] of the process It is desirable to maintain a pulp density as high as is practical, preferably over 35% and at around 40% but below about

45% solids, in order that the pulp density through the autoclave is maximized for maximum ore throughput rates, recognizing that further dilution will be effected before the ore pulp reaches the autoclave due to water pick up in the direct heat exchange with steam in the feed preheaters. Thus the amount of mother liquor solution used for pulping said feedstock should be sufficient to give the pulped ore a pulp density in a range from about 35% to about 45%.

All of the equipment from the pulper onwards must be acid resistant as the mother liquor is highly acidic and also hot, arriving at about 95°C at the pulper. With the addition of the raw ore, the temperature in the pulper is expected to be between 40 ⁰ C and 50 ⁰ C.

The pulper could effectively be a rubber lined mill, while the following equipment in the feed preparation could be of stainless steel construction. The multi deck screen should pass all minus 48 mesh material onwards to the feed storage tank, the +1 cm coarse could be discarded if it analyzes below about 0.9%Ni, and the minus 1 cm plus 48 mesh intermediate fraction would pass through a hammer mill on the way to the feed storage tank. If the +1 cm fraction assays l% or higher in nickel then it too could be passed through the hammer mill and accepted as feed. (These cut-off levels of 0.9% and 1.0% could vary depending on detailed economic evaluations.) A small quantity of fresh water could be effectively employed to wash the oversized rock fraction on the screen.

As a result of the pulping, screening and crushing steps, the resulting feed should be, largely, -100 mesh size and essentially all -48 mesh size. It is self evident to those skilled in the art, that some leaching, particularly of the minor portion of magnesium silicate minerals present in the predominantly limonitic ores, will take place in the pulping stage; and it is further self evident to those skilled in the art that other devices other than a mill can be employed for pugging and pulping of the ore.

STEP [3] Pressure Leaching

The acid pressure leaching step in step [3] of Figure 1 may be achieved using the autoclave depicted in Figure 2, which may be a

conventional six compartment unit designed to operate at temperatures between 250 ⁰ C and 27O ⁰ C. The pressure letdown at the discharge end of the autoclave is effected in flash tanks, in three stages shown, although two stages may be adequate, and that the steam thereby produced at three (or two) different temperatures is returned to the feed end of the autoclave for preheating the feed pulp by direct contact and heat exchange with the three (or two) steam streams, in three (or two) separate preheat tanks. While a significant large quantity of water is converted to steam in the letdown flash tanks, the bulk of this steam is condensed and recovered as water in the preheating heat exchangers and is available for diluting the feed pulp. For 1000 tonnes of ore feed (dry wt. basis) some 450 to 500 tonnes of water could be picked up by the feed pulp on passage through the preheaters, and following the addition of the sulphuric acid to the autoclave the solids content of the pulp could be lowered substantially from the original pulp density after addition of the mother liquor.

Upon reaction of the sulphuric acid with the laterite minerals, the major limonite components are dissolved and reconstituted as hematite, while the minor saprolite components are largely decomposed by the dissolution of the magnesia and liberation of free silica, and in both cases the water of crystallization present in the natural unreacted minerals, which could vary between about 10% and as much as 15%, is also liberated causing further dilution of the pulp in the autoclave. Other solid minerals hosting the nickel, cobalt and manganese are also dissolved along with a high proportion of the alumina and a smaller proportion of the chromite.

The net result of all of these reactions is the shrinking of the solids fraction and increase of the liquid fraction causing yet further dilution of the leached pulp to around 30% solids.

Thus, in summary, the acid pressure leaching step involves adding an effective amount of sulphuric acid to the pulped ore to produce a sulphuric acid solution, agitating and leaching the feedstock in the sulphuric acid solution at an elevated temperature under pressure for a selected period of time whereby metal oxides are leached from said ore to produce a leach pulp.

STEP [4] Mother Thickener

In the solid-liquid separation step in step [4] of Figure 1, the "mother" thickener receives the incoming reacted leach pulp at about 30% solids; but after the decanting of a fraction of the mother liquor for recycle back to step [2] for feed preparation, the remaining portion of mother liquor along with the leached tailings pass as thickener underflow at a pulp density preferably between about 35% to about 40% solids, depending on the amount of mother liquor that is required for pulping of the feed, directly to partial neutralization. Thus, this step involves separating the leach pulp into the mother liquor solution and a first thickened leach pulp, wherein the first thickened leach pulp will still include some of the mother liquor solution.

STEPS [5] and [6] Partial Neutralization

The first thickened leach pulp arriving from the mother thickener in step [4] is too dense to accommodate the extra solid products such as gypsum, ferric hydroxide and alumina and chromium hydroxides to be produced upon neutralization. Accordingly, a very significant dilution of the pulp must be effected, and this is accomplished in a number of ways. Firstly, the neutralizing reagent, limestone (calcium carbonate), is added as a relatively dilute pulp of about 25% solids, secondly, a large volume of filtrate containing diluted quantities of nickel and cobalt is reverted from the tailings vacuum filters back to the neutralization reactors for recovery of their metal values shown by the dotted line 20 (from step [7] to back to step [5] in Figure 1), and thirdly, a significant proportion of the tailings thickener underflow is also recycled back to the neutralizing reactors used in step [5] as "seed" material for the fresh precipitates being formed during the partial neutralization, as indicated by solid line 22 on Figure 1. As a result of these dilutions, the resulting neutralized pulp reaching the tailings thickener will have been diluted to densities between about 10 % and 15% solids.

Thus, this step involves separating the partially neutralized first leach pulp into a pregnant solution containing sulphates of all extracted metal values and a second thickened pulp in which the second thickened pulp includes a tailings fraction which includes leached ore, precipitated gypsum and precipitated hydroxides of iron and other impurity elements present and pregnant solution, some of which is recycled back as mentioned above (line 22) to be mixed with the first thickened leach pulp limestone pulp and liquid filtrate from step [7] in the partial neutralization step [5]. Embodiments of the present process involve using some of the barren solutions of pH of around 8.5 produced in steps [9] and [10] discussed below for producing the limestone (calcium carbonate) pulp for use in this partial neutralization step.

The number of neutralizing reactors employed preferably will usually be four (4) in series, to insure efficient consumption of the limestone reagent and to control the pH within narrow limits, of around pH 4.0, so as to insure maximum precipitation of the undesirable dissolved iron while suppressing the precipitation and loss of any nickel and cobalt values to tailings.

STEP [7]

Vacuum Filtration of Tailings

As mentioned above, the solids in the tailings fraction of the second thickened leach pulp are comprised of the leached ore residues, the gypsum, the ferric hydroxide, as well as the hydroxides of aluminium and chromium and the second thickened leach pulp also includes some residual pregnant solution.

The present process includes washing the tailings filter cake with the barren solutions produced in steps [9] and [10] (see broken lines 24 and 26 in Figure 1) to produce the liquid filtrate containing residual nickel and cobalt values and separating the washed tailings filter cake from the liquid filtrate which is then recycled back and mixed with the first thickened leach pulp and the limestone (calcium carbonate) pulp in partial neutralization step [5] (indicated by the broken line 20 in Figure 1),

thereby virtually obviating the need to dispose of any excess barren solution to the external environment.

The liquid phase in the washed tailings filter cake contains dissolved magnesium sulphate and only trace amounts of the base metals such a nickel and cobalt but a significant proportion of the manganese that arrived with the raw ore. Based on laboratory drying tests the solids/liquid contents of the filter cakes usually hover around 50%/50%.

It is common in the metals industry, when filter cake is being placed in permanent disposal sites (rather than discharging thickened pulps in storage ponds), to speak of "dry" placement. It is the consensus of many soil experts that the moisture/the liquid phase associated with "dry" tailings will rest there on a virtually permanent basis. For example, the manganese that arrived with the ore and was leached out along with the nickel and cobalt but ended up in the barren solutions, is essentially concentrated in the "dry" filter cake and is placed back in the ground whence it came from. The magnesium sulphate that is produced by leaching out of the magnesia in the magnesium silicate minerals is an important component of the liquid phase of the tailings.

Also, the tailings filter cake is a permanent host for the other leached components that had been leached and precipitated. Although a drum vacuum filter is depicted in Figure 2, other filters such as a pan filter may prove to be more efficient particularly with the washing.

The method of "placing" the tailings will depend largely on the location and layout of the mining and processing facilities and on the location of the impoundment areas. Figure 2 depicts pugging and pumping of a dense/thick paste. However, trucking or using conveyor belts could prove to be better methods depending on local circumstances.

As seen in Figures 1 and 2, there are essentially no excess process waste waters to dispose of in this unique flowsheet that starts out employing recycled mother liquor for pulping of the feed ores, since little or no fresh water is required for processing. The environmental advantage rendered by the present invention, with regard to protection of the external environment, becomes very evident.

STEPS [8] and [9] Metal Recovery

These steps involve treating the pregnant solution from step [6] (Figure 1) with precipitating agents to yield barren solution and a third thickened pulp containing precipitated metal values and residual barren solution, followed by separating the third thickened pulp from the barren solution.

More particularly, the metal recovery circuit, while receiving and treating large volumes of clear pregnant liquid with precipitating agents, yields but a small amount of solid precipitates. This results in dilute feed to the product thickener usually of less than about 5% solids, and a thickened underflow pulp usually containing around 10% or less of solids. In order to enhance subsequent filtration rates, a portion of the thickener underflow is recycled back to the product precipitation reactors to serve as "seed" materials for the fresh precipitates, see solid line 28 in Figure 1.

The number of precipitating reactors is usually greater than two (2), with normally four (4) employed in series. The addition of a flocculent to the product thickener dilute pulp feed, is extremely useful at this stage, to flocculate the relatively small amount of suspended and dispersed solids. Depending on the particular type of intermediate nickel-cobalt product that is desired, several different precipitating agents may be used for recovering the nickel and cobalt values from the pregnant solutions, each being discussed herebelow. The pregnant solution streams pass onto nickel-cobalt precipitation which can render final neutralization of the leach solutions in the cases wherein lime or soda ash are employed as the reagents. When employing magnesia as the precipitating reagent, after separation of the nickel-cobalt hydroxide as a pulp from the treated feed liquor, it is generally desirable/necessary to proceed with the said treated feed liquor to a final scavenging neutralization step with lime to remove the last traces of nickel and cobalt. Also, when using H ₂ S gas or aqueous

NaHS, further neutralization needs to be carried out on the treated feed liquor, after separation of the nickel-cobalt sulphides as a pulp from the treated feed liquor, to neutralize its acid content and to remove any last traces of nickel and cobalt. Use of sodium sulphide (Na ₂ S) would require

a minimal amount of further scavenging neutralizing. The scavanged solids containing residual nickel and cobalt values would normally be recycled to the front end of the partial neutralization reactors in step [5].

When employing lime (CaO pulp at a density of about 15% solids), for the neutralizing/precipitation agent for the pregnant feed liquor, solid gypsum is produced along with the precipitated base metal hydroxides of nickel and cobalt as well as of the impurity base metals of zinc and copper; and a significant proportion of the manganese is brought down with the nickel and cobalt. As a result, usually over half of the product precipitate is CaSO ₄ with the nickel content usually in the range from about 15 % to about 25% Ni. The reaction with lime is rapid and it is easy to arrive at a pH of say 8.5, or higher if necessary to assure only trace quantities of nickel, cobalt and manganese in the barren solutions. Depending on particular circumstances, a "polishing" filter (not shown in Figure 2) may be used to insure complete removal of particulate matter from the barren solutions.

Embodiments of the present process involve using some of the barren solutions of pH of around 8.5 for producing the lime pulp for use in this partial neutralization step. When employing a magnesia pulp as the neutralizing/precipitation agent, the reaction is much slower and pH's of 8 and higher are difficult to achieve efficiently. The magnesium sulphate stays largely in solution, and nickel assays of around 35% or slightly higher can be achieved. However, after the separation and recovery of the nickel cobalt hydroxides as a pulp from the treated pregnant feed liquor it is generally desirable and may be preferable to subject the treated feed liquor to a final scavenging neutralization step with lime to remove the last traces of nickel and cobalt. In such a circumstance, the final neutralized barren solution of pH 8.5 or higher could proceed to a clarifier for recovery of the scavenged nickel cobalt hydroxides for recycle to step [5], producing a clarified barren solution for wash water at step [7]. (The auxiliary equipment for this scavenging step is not shown in Figure 2). Embodiments of the present process involve using some of the barren solutions of pH of around 8.5 for producing the magnesia pulp for use in this partial neutralization step.

With a soda ash (NaCO ₃ ) pulp being used as the neutralization/precipitating agent, similar higher grades of precipitates of 35%Ni or higher, can be produced, as the sodium sulphate produced stays largely in solution. As with the lime neutralization/precipitation, clarification of the barren solution by employment of a "polishing" filter may be desirable. Embodiments of the present process involve using some of the barren solutions of pH of around 8.5 for producing soda ash (NaCO ₃ ) pulp for use in this partial neutralization step.

When precipitating with H ₂ S gas, the flowsheet is more complicated than shown in Figure 2, as H ₂ S must usually first be manufactured and the precipitation is carried out in an autoclave slightly above atmospheric pressure and acid is produced which must eventually be neutralized; but a sulphide precipitate with about 55%Ni can be produced with the bulk of the manganese staying behind in the treated barren_solution. With aqueous NaHS, a simpler circuit can be employed and similar grades of sulphide precipitates can be produced low in manganese, and only half as much acid is produced. In either case, the acid must first be neutralized; and a final scavenging neutralization step must be carried out to raise the pH high enough to remove the last traces of nickel and cobalt from the barren solution, as in the case of precipitation with magnesia. Sodium sulphide reagent would require the simplest circuit, and the higher grade sulphide precipitates could be produced without concomitant production of acid; and final neutralization of the filtrate from the sulphiding circuit, to achieve a pH of 8.5 (or higher) in final barren solutions, can readily be produced with a small addition of lime. The big difference between the three sulphiding techniques are technology/equipment and costs.

Addition of flocculent to the products from the precipitation reactors is usually preferred to produce a clear natant solution at the product thickener as shown in Figure 1 with flocculent being added in steps [6] and [9].

STEP [10] Product Filtration

The particle size of the product nickel-cobalt hydroxides is very fine, and the filter cake that would be produced by vacuum filtration could contain as much as 70% moisture. However, by employing pressure filters the moisture can be drastically reduced to yield cake of around 30% or less of moisture. Clearly this is a major improvement in product quality where the product is to be shipped to an outside nickel refinery whether it be a pyrometallurgical, hydrometallurgical or vapometallurgical refinery. The school of thought that in partial neutralization of base metal solutions containing nickel, cobalt and manganese as well as iron, to pH levels of around four (4), to precipitate and remove the iron, that (1 ) there is significant co-precipitation of the other base metals such as of the nickel, cobalt and manganese, (2) that much of the iron present in nickel- ferrous magnesium silicate minerals is in the bivalent state, and (3) that it is extremely difficult to filter the iron hydroxide-containing pulps without undue losses of nickel and cobalt values, is misguided It was discovered that the greatest part of the iron dissolved from the nickel laterite minerals is in fact in the trivalent state, that no significant amount of co-precipitation takes place and that filtration can be much improved by recirculation of "seed" material. However, the most important finding was that by employing adequate/appropriate amounts of barren wash water to wash the hydroxide-containing filter cake preferably in amounts that would result in at least about two (2) volume displacements of the liquid in the filter cake, that recoveries of as high as 99% of the dissolved nickel and cobalt values can be achieved. Thus, by employing single-stage vacuum filters with appropriate washing, recovery of the nickel and cobalt values into virtually iron-free pregnant solutions can be as high as those obtained by CCD circuits employing as many as seven thickeners in series.

EXAMPLE Leach solutions with pH's of between about 1 and 1.5, obtained by leaching nickel laterite ore samples from three different ore regions in New Caledonia, were partially neutralized to pH levels ranging from around 3.5 to around 4.3 with the bulk between 3.8 and 4.0, by the use of limestone (CaCO ₃ ) slurries of around 25% solids. The pregnant solutions contained

between 5 and 6 gpl of Ni, between 0.1 and 0.6 gpl of Co, between 0.3 and 0.5 gpl of Mn, and between 15 and 20 gpl of Fe before partial neutralization, and the iron levels were lowered to between about 1 and 0.1 gpl after the partial neutralization. The entire de-watering and washing was accomplished on a vacuum filter, in one stage, to yield discard tailings of gypsum, iron hydroxide and other hydroxides such as those of aluminium and chromium, in a manner as depicted in the non-conventional flowsheet of Figure 2.

Figure 3 is a plot of % nickel recovery versus the number of displacements of cake water (gypsum + iron hydroxide cake) demonstrating the efficacy of the vacuum filtratration mode of recovering the soluble nickel values as virtually iron-free pregnant solutions, wherein wash water in amounts of two (2) displacements of the liquid in the filter cake, is used. Such a vivid demonstration that the gypsum plus iron hydroxide and other hydroxide precipitates could be filtered and washed so thoroughly, led to the next step and that was to subject the first thickened leach pulp (liquid and solids) directly to the partial neutralization with limestone, with the production of a single composite solids tailings fraction containing the leached ore tailings, the gypsum, the iron hydroxide and the other minor hydroxides. This then led to the simplified flowsheet depicted in Figures 1 and 2 incorporating single-stage filtration using appropriate yet practical quantities of wash water.

Thus the present method disclosed herein solves a major problem in this type of metal value recovery in that it greatly reduces the amount of "fresh" water required in preparing the feed pulp from which the metal values are extracted, and obviates the need to dispose of process waste waters to the external environment.

As used herein, the terms "comprises", "comprising", "includes" and "including" are to be construed as being inclusive and open ended, and not exclusive. Specifically, when used in this specification including claims, the terms "comprises", "comprising", "includes" and "including" and variations thereof mean the specified features, steps or components are included. These terms are not to be interpreted to exclude the presence of other features, steps or components.

The foregoing description of the preferred embodiments of the invention has been presented to illustrate the principles of the invention and not to limit the invention to the particular embodiment illustrated. It is intended that the scope of the invention be defined by all of the embodiments encompassed within the following claims and their equivalents.

References Cited:

United States Patent No. 6,391 ,089 B1 5/2002 Curlook United States Patent No. 6,379,637 B1 4/2002 Curlook