WO2015039177A1

WO2015039177A1 - Method for ammoniacal leaching of zinc from carbonate-hosted ores

Info

Publication number: WO2015039177A1
Application number: PCT/AU2014/000927
Authority: WO
Inventors: Matthew Leslie Sutcliffe; Garry Mervyn Johnston; Nicholas James Welham
Original assignee: Metaleach Limited
Priority date: 2013-09-20
Filing date: 2014-09-19
Publication date: 2015-03-26
Also published as: AU2018264020A1; MX2016003686A; AU2014324082A1; AU2018264020B2

Abstract

In accordance with the present invention there is provided a method for leaching zinc from a carbonate-hosted ore, the method comprising the method steps of: applying an aqueous solution of a curing agent to the carbonate-hoste ore, producing a cured ore; forming a leach solution by applying an ammoniacal solution that has an ammonium carbonate content of less than 5 g/L to the cured ore, producing a pregnant leach solution; and passing the pregnant leach solution to a means for recovering zinc.

Description

Method for Ammoniacal Leaching of zinc from Carbonate-Hosted Ores Field of the Invention

[0001] The present invention relates to a method for leaching zinc. More particularly, the present invention relates to a method for leaching zinc from carbonate-hosted ores using an ammoniacal medium.

Background Art

[0002] Some prior art reports of ammoniacal leaching have used tank or vat leaching, in which the ore or concentrate is simply immersed directly into ammoniacal solution.

[0003] For zinc, the Schnabel process was used for a number of years before being superseded by the more environmentally friendly, acid-based roast-leach- electrowin process. The Schnabel process feed was typically roasted sphalerite flotation concentrate but selective mining also allowed processing of zinc oxide ores. The Schnabel process is complex (as is evident from the summary by Harvey (Mineral Processing & Extractive Metallurgy Review, volume: 27, pages: 231-279, 2006), and it is perhaps unsurprising that there are currently few, if any, Schnabel process plants in operation.

[0004] Although many methods for ammoniacal leaching are described in the literature, including the patent literature, all require one or more of multiple leaching stages, energy-intensive pre-treatment of the ore (for example roasting or grinding) and/or elevated ieach temperatures and/or pressures. These requirements necessitate complicated engineering and significant energy consumption, so it is little surprise that none have been widely adopted.

[0005] Further, many ammoniacal leaching methods either suffer from significant ammonia loss, or require complicated chemistry or engineering to overcome such, providing a further impediment to the widespread adoption of ammoniacal leaching methods. [0006] Other ammoniaca! leaching methods have previously utilised ammoniacai leaching methods involving curing steps. However, these methods require the ammoniacai leach solution to have an ammonium carbonate concentration of at least 5 g/L. This requirement adds to operation costs as additional ammonium carbonate is often required to be added to the leach solution.

[0007] The method of leaching of the present invention has as one object thereof to overcome the abovementioned problems associated with the prior art, or to at least provide a useful alternative thereto.

[0008] Throughout this specification, unless the context requires otherwise, the word "comprise", or variations such as "comprises" or "comprising", will be understood to imply the inclusion of a stated integer or group of integers but not the exclusion of any other integer or group of integers.

[0009] The discussion of the background art is included exclusively for the purpose of providing a context for the present invention. It should be appreciated that the discussion is not an acknowledgement or admission that any of the material referred to was common general knowledge in the field relevant to the present invention in Australia or elsewhere before the priority date.

Disclosure of the Invention

[0010] In accordance with one aspect of the present invention there is provided a method for leaching zinc from a carbonate-hosted ore, the method comprising the method steps of:

applying an aqueous solution of a curing agent to the carbonate-hosted ore, producing a cured ore;

applying an ammoniacai solution that has an ammonium carbonate content of less than 5 g/L to the cured ore, producing a pregnant leach solution; and

passing the pregnant leach solution to a means for recovering zinc. [00 1] In accordance with one aspect of the present invention there is provided a method for leaching zinc from a carbonate-hosted ore, the method comprising the method steps of:

forming a leach solution by applying an ammoniacal solution that has an ammonium carbonate content of less than 5 g/L to the cured ore, producing a pregnant leach solution; and

passing the pregnant leach solution to a means for recovering zinc.

[0012] in accordance with a further aspect of the present invention there is provided a method for leaching zinc from a carbonate-hosted ore, the method comprising the method steps of;

applying an ammoniacal solution that has an ammonium carbonate content of less than 5 g/L to the cured ore, wherein the leach solution maintains a pH between 6 and 13 during the leach process

producing a pregnant leach solution; and

passing the pregnant leach solution to a means for recovering zinc.

[0013] In accordance with a further aspect of the present invention there is provided a method for leaching zinc from a carbonate-hosted ore, the method comprising the method steps of:

applying an ammoniacal solution that has an ammonium carbonate content of less than 5 g/L to the cured ore, wherein the leach solution self- buffers to a pH between 8 and 11 during the leach process;

producing a pregnant leach solution; and

passing the pregnant leach solution to a means for recovering zinc. [0014] Throughout this specification, unless the context requires otherwise, the term "ore" or variations thereof, will be understood to include, for example, the product of one or more pre-treatment steps, such as a roast or calcination steps, or one or more concentration steps, but is not limited thereto. The term "ore" or variations thereof, will also be understood to include, for example, the zinc- bearing waste product of steeimaking processes, such as electric arc furnace (EAF) wastes.

[0015] Throughout this specification, unless the context requires otherwise, the term "carbonate-hosted ore" or variations thereof, means target ores that are surrounded by and / or retained within, carbonate compounds. The term "carbonate-hosted ore" or variations thereof, will also be understood to include, for example, ores hosted within carbonate (limestone, marl, dolomite) formations and which share a common genetic origin.

[0016] As would be recognised by those skilled in the art, this process may form part of a larger flowsheet for processing zinc ores. Where there is an economic proportion of the zinc (or zinc in combination with other valuable metals, e.g. lead) present as sulphides, the outlined process may be used to treat the tailings of a prior flotation process to recover the sulhpide minerals. Equally, it may be more economic to use the outlined process prior to recovery of the sulphide minerals by flotation.

[00 7] Preferably, the carbonate-hosted ore has an acid neutralisation capacity of at least 5 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 10 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 15 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 20 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 25 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 50 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 75 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 00 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 150 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 200 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 250 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 300 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 400 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 500 kg of sulphuric acid per tonne, when measured at pH 1.5. Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 750 kg of sulphuric acid per tonne, when measured at pH 1.5, Preferably still, the carbonate-hosted ore has an acid neutralisation capacity of at least 1000 kg of sulphuric acid per tonne, when measured at pH 1.5.

[0018] In a highly preferred form of the present invention, the carbonate-hosted ore has an acid neutralisation capacity of between 50 kg and 400 kg of sulphuric acid per tonne, when measured at pH 1.5.

[0019] As would be understood by those skilled in the art, the procedure for determining the acid neutralisation capacity is to add a known mass of ground ore (e.g. 50g) to a large volume (e.g. 500mL) of sulphuric acid solution at the required pH (e.g. 1.5). The mixture is stirred and periodically measured volumes of known concentration sulphuric acid (e.g. 100g/L) is added to return the slurry to the starting pH. This procedure continues until the pH is stable for more than 1 hour. The acid neutralisation capacity is calculated from the measured volume of acid added and the strength of acid divided by the mass of ore used. For example, if 75mL of 100g/L H2SO4 were used to maintain pH1.5 for a 50g ore sample the acid neutralisation capacity would be (75/1000)*{100/1000)/(50/1000000) = 150kg/t.

[0020] Preferably, the zinc present as sulphide minerals is less than 90% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 80% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 70% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 60% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 50% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 40% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 30% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 25% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 20% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 15% of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 10%» of the total zinc in the ore. Preferably still, the zinc present as sulphide minerals is less than 5% of the total zinc in the ore.

[0021] In a highly preferred form of the present invention, the zinc present as sulphide minerals is less than 25% of the total zinc in the ore.

[0022] As would be recognised by those skilled in the art, deposits which have been derived from the oxidation of lead-zinc sulphides and are suitable for the outlined method are also likely to contain silver in potentially economic amounts. For such deposits the silver may be recovered either simultaneously with the zinc by the addition of thiosuiphate (S₂O₃ ^2") or cyanide (CN^") ions. It is also possible to run a two stage process whereby the first stage is the outlined one where zinc is recovered, the second stage is a conventional alkaline leach containing cyanide or thiosuiphate ions which selectively recover the silver. [0023] in one form of the present invention, where the ore contains silver, the step of:

passing the pregnant leach solution to a means for recovering zinc,

[0024] more specifically comprises the step of;

passing the pregnant leach solution to a means for recovering zinc and silver recovery, either simultaneously or concurrently.

[0025] In comparison to the present invention, using an acid process for the recovery of zinc would make simultaneous recovery essentially impossible due to the instability of the thiosulphate ions in acid and the generation of hydrocyanic acid on addition of cyanide. The subsequent use of either reagent would also prove difficult as the residue would need to be completely neutralised in order to prevent the problems with these reagents in acid systems. Attempting to recover the silver prior to the zinc would also prove difficult as the cyanide ions would also dissolve the zinc leading to separation difficulties and a high consumption of cyanide. In this case, the residue would need to be thoroughly washed to remove any cyanide ions to prevent hydrocyanic acid evolution on addition of acid.

[0026] In one form of the present invention, the carbonate hosted ore comprises one or more minerals selected from the group comprising; calcite, galena, hydrozincite, goethite, calcite, hydrozincite, cerussite, smithsonite, sphalerite, and hemimorphite. Preferably, the ore comprises one or more minerals selected from the group comprising; calcite, hydrozincite, galena, goethite.

Curing

[0027] The inventors have discovered that when an ammonia leach solution is applied to carbonate-hosted ores containing zinc that have undergone a curing step, there is no requirement for the leach solution to have a high ammonium carbonate content. In particular, the inventors have discovered that the carbonate-hosted ores may be viably leached in ammoniacal solutions with an ammonium carbonate content of less than 5 g/L. Previous ammoniacal leach methods have had a requirement of utilising an ammoniacal solution with a high ammonium carbonate content, which often led to the need for additional ammonium carbonate to be added to the leach solution, resulting in additional operating costs, in many cases, the improved economics afforded by removing the requirement of a certain ammonium-carbonate content will at least improve the commercial viability of the ore.

[0028] As would be understood by a person skilled in the art, the term curing is fundamentally distinct from leaching. Leaching describes a process by which a solution containing a leaching agent is contacted with an ore, the solution recovered and valuable metals extracted therefrom. The curing step of the present invention renders the ore to be leached more amenable to the leaching process, improving both the extent and rate of recovery of the zinc. Without wishing to be bound by theory, this may arise from one or more of the oxidation or reduction of the zinc or otherwise refractory ores containing the zinc, the complexation of metal (target or non-target) and the mobilisation of metal (target or non target),

[0029] The scope of the present invention encompasses methods where the aqueous solution of the curing agent is collected after the step of curing the ore to be leached through the application of an aqueous solution of a curing agent, and metal values recovered therefrom. However, conventional aqueous leaching solutions do not fall within the meaning of aqueous solution of a curing agent, as they do not render the ore to be leached more amenable to the subsequent leaching process. For example, two stage ammoniacal leaching processes differ markedly from the method of the present invention as there is no enhancement of the second leaching stage by performance of the first.

[0030] However, in a preferred form of the invention, the step of curing the carbonate-hosted ore through the application of an aqueous solution of a curing agent more specifically comprises substantially retaining the curing agent in contact with the ore to be leached when the ammoniacal solution is added. As would be understood by a person skilled in the art, in many applications, it is virtually impossible to completely retain a solution in contact with the ore to be leached. For example, in a heap leaching context, it is virtually impossible to stop drainage from the ore.

Curing agents: general

[0031] The nature and concentration of the curing agent depends on the mineralogy of the carbonate-hosted ore, the texture of the carbonate-hosted or ore and the pore volume of the carbonate-hosted or ore.

[0032] As would be understood by a person skilled in the art, the term texture describes the manner in which the minerals are arranged in the ore (e.g. goethite coating zinc oxide minerals is a very different proposition to a zinc oxide ore with associated goethite).

[0033] Throughout this specification, unless the context requires otherwise, the phrases "pore space" and "pore volume" refer to the space comprising the pores within the ore particles, as opposed to inter-particle pores created by any stacking process.

Curing conditions: general

[0034] Further, and similarly, the most desirable conditions under which the ore is cured vary as the composition, mineralogy and texture of the ore varies. For example, the nature and concentration of the curing agent, the temperature at which the curing step occurs, the pH at which the curing step occurs and the time for which the ore is exposed to the curing agent may all be varied in response to the composition, mineralogy, texture and pore volume of the ore (with low pore volumes necessitating higher concentrations). In general however, the higher the concentration of the targeted mineral in the ore, the higher the concentration of the solution used in the curing step. [0035] The curing agent may be an agent previously associated with leaching the metal of interest from the ore. in such cases, the concentration of the curing agent in the aqueous solution of the curing agent will typically be substantially higher than the concentrations conventionally associated with leaching the metal of interest from the ore to render the ore to be leached more amenable to the leaching process.

[0036] in a highly preferred form of the invention, the step of curing the ore to be leached takes place at atmospheric pressure.

[0037] in a highly preferred form of the invention, the step of curing the ore to be leached takes place at ambient temperature.

[0038] The volume of the aqueous solution of the curing agent applied to the ore is a function of a number of parameters including, but not limited to, the texture of the ore, the residence time (the time for which the ore is exposed to the curing agent prior to the leaching step), the concentration of the curing agent and the leach conditions. However, in preferred embodiments of the invention, the volumes of aqueous solutions of curing agent exposed to the ore are as low as practicable. That is, preferred forms of the invention utilise low volumes of an aqueous solution of a curing agent of high concentration, and preferred methods for curing the ore to be leached through the application of an aqueous solution of a curing agent are those adapted to utilise low volumes of aqueous solution of the curing agent.

[0039] Although the identity of the curing agent, the volume and concentration of the aqueous solution of the curing agent, and the conditions under which the curing step takes place wili differ for different ores, the present invention identifies a principle enabling the economical recovery of zinc from a wide range of carbonate-hosted ores, in that conventional, energy-intensive physical pre- treatment techniques such as grinding or roasting, used successfully or otherwise in ammoniacal leaching, can be replaced or enhanced by using chemical curing techniques, where combinations of the activity of the curing agent and elevated concentrations of the curing agent render the ore amenable to the subsequent atmospheric ammoniacal leaching.

[0040] Aqueous solutions of curing agents are preferably low volumes of high concentration solutions. In a preferred form of the invention, the aqueous solution of the curing agent is at least 0% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 20% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 30% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 40% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 50% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 60% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 70% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 80% of the saturated concentration of the curing agent under the prevailing conditions. Preferably still, the aqueous solution of the curing agent is at least 90% of the saturated concentration of the curing agent under the prevailing conditions.

Curing conditions: application of curing agent to the ore

[0041] The aqueous solution of the curing agent may be generated in situ, such as by electrolytic means.

[0042] In one form of the invention, the step of curing the ore to be leached through the application of an aqueous solution of a curing agent more specifically comprises: spraying the aqueous solution of the curing agent onto the ore prior to the step of leaching the cured ore at atmospheric pressure through the application of an ammoniacaf solution.

[0043] The method of the present invention may include the step of:

reducing the size of the ore to be leached by grinding.

[0044] In a particular form of the invention, the method of the present invention includes the step of:

reducing the size of the ore to be leached by wet grinding, wherein the ore is ground in contact with water or an aqueous grinding solution.

[0045] Where the invention comprises reducing the size of the ore to be treated by wet grinding, wherein the ore is ground in contact with water or a grinding aqueous solution, the aqueous grinding solution may be provided in the form of the aqueous solution of the curing agent.

[0046] In one form of the invention, the method comprises the steps of:

grinding the ore in an aqueous solution of a curing agent, thereby curing the ore to be leached through the application of an aqueous solution of a curing agent;

resting the ore for a predetermined period; then

forming a leach solution by applying of an ammoniacal solution to the cured ore; and

passing the pregnant leach solution to a means for metals recovery.

[0047] The method of the present invention may include a step of:

reducing the size of the ore to be leached by crushing.

[0048] In a particular form of the invention, the method of the present invention includes the step of:

reducing the size of the ore to be leached by wet crushing, wherein the ore is crushed in contact with water or an aqueous crushing solution. [0049] Where the invention comprises reducing the size of the ore to be treated by wet crushing, wherein the ore is crushed in contact with water or an aqueous crushing solution, the aqueous crushing solution may be provided in the form of the aqueous solution of the curing agent.

[0050] in one form of the invention, the method comprises the steps of:

crushing the ore in an aqueous crushing solution of a curing agent thereby curing the ore to be leached through the application of an aqueous solution of a curing agent;

resting the ore for a predetermined period; then

passing the pregnant leach solution to a means for metals recovery.

[0051] In one form of the invention, the step of curing the ore to be !eached through the application of an aqueous solution of a curing agent more specifically comprises:

stacking the ore to form a heap;

irrigating the surface of the heap with the aqueous solution of the curing agent such that the aqueous solution of the curing agent percolates down through the heap;

resting the ore for a predetermined period, prior to the step of forming a leach solution by applying of an ammoniacal solution to the cured ore.

[0052] In one form of the invention, the step of curing the ore to be leached through the application of an aq eous solution of a curing agent more specifically comprises:

immersing the ore in an aqueous solution of the curing agent for a sufficient time such that the curing solution infiltrates a desired fraction of pore volume;

resting the ore for a predetermined period prior to the step of forming a leach solution by applying of an ammoniacal solution to the cured ore. [0053] The method of the present invention may include a step of:

agglomerating the ore to be leached.

[0054] In a particular form of the invention, the method of the present invention includes the step of:

agglomerating the ore to be leached by contacting the ore with water or an aqueous solution of an agglomerating agent.

[0055] In one form of the invention, the aqueous solution of the curing agent is also the aqueous solution of the agglomerating agent. That is, the aqueous solution contains both a curing agent and an agglomerating agent. In one form of the invention, the curing agent is an agglomerating agent, such that the step of curing the ore to be leached through the application of an aqueous solution of a curing agent more specifically comprises:

agglomerating the ore with the aqueous solution of the curing agent;

resting the ore for a predetermined period prior to the step of forming a leach solution by applying of an ammoniacal solution to the cured ore.

[0056] In one form of the invention, the step of curing the ore to be leached through the application of an aqueous solution of a curing agent more specifically comprises:

spraying the aqueous solution of the curing agent onto the ore; and resting the ore for a predetermined residence time, prior to the step of forming a leach solution by applying of an ammoniacal solution to the cured ore.

[0057] In a particular form of the invention, the method of the present invention comprises the step of:

agglomerating the ore to be treated by contacting the ore with a >500g/L sulphuric acid solution. [0058] The predetermined time for which the ore is rested prior to the step of ammoniacai solution forming a leach solution by applying of an ammoniacal solution to the cured ore wiil be a function of a number of parameters including, but not limited to the particle size of the ore, the concentration of the curing agent and the texture of the ore.

[0059] In preferred forms of the invention, the predetermined period is between 5 minutes and twenty eight days. Preferably still, the predetermined period is between 1 day and 7 days. A person skilled in the art will realise that the curing time will be a function of particle size, small particles requiring substantially less curing time than large particles. A person skilled in the art will realise that a high concentration of curing agent will require a shorter resting time than a low concentration of curing agent.

[0060] The ideal extent of saturation of the pore space of the ore with the aqueous solution of the curing agent wiil depend largely on the texture of the ore. Preferably the step of curing the ore to be leached through the application of an aqueous solution of a curing agent saturates at least 50% of the pore space with solution. Preferably still, the step of curing the ore to be leached through the application of an aqueous solution of a curing agent saturates at least 60% of the pore space with solution. Preferably still, the step of curing the ore to be leached through the application of an aqueous solution of a curing agent saturates at least 70% of the pore space with solution. Preferably still, the step of curing the ore to be leached through the application of an aqueous solution of a curing agent saturates at least 80% of the pore space with solution. Preferably still, the step of curing the ore to be leached through the application of an aqueous solution of a curing agent saturates at least 90% of the pore space with solution.

Leaching: pressure and temperature

[0061] The most desirable conditions under which the cured ore is leached will vary as the conditions under which the ore is cured vary. For example, weaker cure solutions may require more aggressive leach solutions. [0062] Methods for leaching ore at atmospheric pressure are well known to persons skilled in the art, and include heap leaching, vat leaching, tank leaching and dump leaching. In preferred forms of the invention, the step of ammoniacal solution forming a leach solution by applying of an ammoniacal solution to the cured ore, producing a pregnant leach solution takes place at ambient temperatures. Atmospheric leaching, particularly at ambient temperatures, is one of the least energy-intensive leaching techniques available. A curing step that is not energy intensive and that renders an ore amenable to an ammoniacal leaching step that is also not energy intensive has clear advantages over prior art methods.

Leaching: ammoniacal solution

[0063] Advantageously, the Applicant has discovered that the leaching of carbonate-hosted ores does not require the addition of any buffering agents, such as ammonium carbonate, during the leach process. The carbonate content fixes the operating pH to a relatively narrow range and is self-regulating as the carbonate content acts as a self-buffer. Importantly, the pH range buffered by the carbonate content is maintained at a range in which zinc is soluble. A second advantage of carbonate systems is that there is iess prospect of gypsum scaling as the sulphate level is always too low for precipitation to occur. The calcium ievel will also be low as the precipitation of CaC0₃ will occur whenever calcium ions are released into solution.

[0064] Preferably, the concentration of ammonium carbonate in the ammoniacal solution is less than 5 g/L. Preferably still, the concentration of ammonium carbonate in the ammoniacal solution is less than 4 g/L. Preferably still, the concentration of ammonium carbonate in the ammoniacal solution is less than 3 g/L. Preferably still, the concentration of ammonium carbonate in the ammoniacal solution is less than 2 g/L. Preferably still, the concentration of ammonium carbonate in the ammoniacal solution is less than 1 g/L. Preferably still, the concentration of ammonium carbonate in the ammoniacal solution is less than 0.75 g/L. Preferably still, the concentration of ammonium carbonate in the ammoniacal solution is less than 0.5 g/L. Preferably still, the concentration of ammonium carbonate in the ammoniacal solution is less than 0.25 g/L. Preferably still, the ammoniacal solution contains no ammonium carbonate.

Ammonical Solution

[0065] The ammonia of the ammoniacal solution may be generated in situ, such as by hydrolysis of urea.

[0066] The free ammonia concentration of the ammoniacal solution may be taiiored to the rate at which the zinc is leached from the cured ore, thereby minimising excess free ammonia and thus minimising ammonia losses due to evaporation. Specifically, the resulting pregnant leach solution preferably contains only a slight excess of free ammonia over that necessary to retain the zinc in solution. As there is little free ammonia in the pregnant leach solution, ammonia losses due to evaporation are low. This is one of the major potential advantages of the present invention.

[0067] A person skilled in the art will readily be able to calculate the free ammonia concentration required to retain zinc in solution at a desired concentration. The conditions under which ammoniacal complexes of zinc form are readily calculable based on data contained in NIST Standard Reference Database 46, NIST Critically Selected Stability Constants of Metal Complexes: Version 6.0, the contents of which are incorporated by reference.

[0068] For example, where the ore contained 18.2% zinc in a rapidly leaching form ammoniacal solution comprises about 30-70 g/L ammonia.

[0069] As would be realised by a person skilled in the art the level of ammonia in the solution applied in step (b) would be matched to the level of zinc in the ore and the rate at which it leaches. A low grade ore where the zinc leaches slowly would require a lower concentration of ammonia than a high grade ore where the leaching is rapid.

[0070] Selection of a suitable free ammonia concentration can be used to maintain a specific concentration of zinc in the leach solution. If the concentration of free ammonia is less. than the value necessary to achieve maximum dissolution the solution will become saturated with zinc when all of the free ammonia is complexed. Leaching will therefore need to be performed for a longer period to achieve the same zinc recovery. Whilst this may seen counter productive the constant concentration of zinc in the leach solution makes it very substantially simpler to control the subsequent processes as the feed concentration will be invariant. Typical leaching plants have to deal with constantly varying leach solution concentrations which requires the process to be continually optimised. This is especially problematical in heap leach operations where the solution concentration decreases continually with time as the metal of interest is leached. In this process the quantity of zinc is constant over time until the ore is essentially leached out. The capability to control the maximum solution tenor is a major advantage of the present process.

Curing agents

[0071] As stated in the context of the preceding general discussion of curing agents, generally above, the nature of the curing agent depends on the mineralogy of the carbonate-hosted ore, the texture of the carbonate-hosted or ore and the pore volume of the carbonate-hosted or ore.

[0072] A person skilled in the art will use either single curing agents or a suite of curing agents within the same cure solution according to the mineralogy of the ore, the texture of the ore and the pore volume of the ore. Some curing agents are chemically compatible, and may be applied simultaneously. Others may necessitate sequential curing. That is, the step of curing the ore to be leached through the application of an aqueous solution of a curing agent, producing a cured ore may comprise the steps of: curing the ore to be leached through the application of a first aqueous solution of a first curing agent; then

further curing the ore to be leached through the application of a second aqueous solution of a second curing agent.

[0073] Examples of chemically incompatible curing agents include acids and bases. Specifically, in one form of the invention, the first curing agent may be provided in the form of an acid, and the second curing agent in the form of ammonia.

[0074] Further, the curing agent of the present invention need not act directly on the zinc. For example, where the ore is a zinc silicate the curing agent may be provided in the form of an aqueous fluoride solution which may complex the silica.

[0075] Advantageously, as zinc only has a single valence under normal conditions oxidising cures and reductants are not required.

Curing agent

[0076] In a preferred form of the invention, where the ore is an oxide ore, the curing agent is selected from the group: metal complexing agent, acid, base and combinations thereof. In a preferred form of the invention, where the ore is an oxide ore, the curing agent is selected from the group: metal complexing agent and combinations thereof. In a preferred form of the invention, where the ore is an oxide ore, the curing agent is a metal complexing agent.

[0077] Where the ore is an oxide of zinc in the form of smithsonite the curing agent is preferably provided in the form of ammonia. Preferably, the concentration of the ammonia in the aqueous solution of the curing agent is at least 30 g/L. However, some ores may be advantageously cured by ammonia concentrations as low as 1g/L. [0078] Where the ore is an oxide of zinc in the form of hemimorphite, the curing agent preferably comprises ammonia and CO2. Preferably, the concentration of the ammonia is at least 30 g/L. Preferably, the concentration of the C0₂ is at least 20 g/L. However, some ores may be advantageously cured by ammonia concentrations as low as 1g/L and / or C0₂ concentrations as low as 1g/L.

[0079] Where the ore is an oxide of zinc in the form of hemimorphite, the curing agent preferably comprises a solution of an acid at between pH 2 and 7.

[0080] Where the ore is an oxide of zinc in the form of zincite and/or hydrozincite, , the curing agent is preferably provided in the form of an ammonia solution. Preferably, the concentration of the ammonia solution is at least 30 g/L. However, some ores may be advantageously cured by ammonia concentrations as low as 1g/L.

Means for metal recovery

[0081] The means for metal recovery of the present invention may comprise one or more of the following: solvent extraction, ion exchange, precipitation and cementation.

Combined leach and cure solutions

[0082] In one form of the invention, the step of curing the ore to be leached through the application of a curing agent more particularly comprises curing the ore to be leached through the simultaneous application of a curing agent and a free ammonia solution. In a highly convenient form of the invention, the curing agent is provided in the form of an aqueous free ammonia solution. Preferably still, the free ammonia concentration of the curing solution exceeds the free ammonia concentration of the ammoniacal solution of the leaching step.

[0083] Without wishing to be bound by theory, where the curing step involves simultaneous application of a free ammonia solution, the high ammonia concentration used solubilises zinc within the pores and transports it towards the surface. Even after curing is concluded, the pores will contain a higher ammonia concentration that the leaching solution thereby giving enhanced diffusion of zinc out of the ore during the step of leaching the cured ore through the application of an ammoniacal solution.

[0084] In one form of the invention, after the step of curing the ore to be leached through the application of a curing agent and before the step of leaching the cured ore through the application of an ammoniacal solution, the method of the present invention comprises the step of: allowing the mixture of ore and curing agent to rest for a predetermined period.

[0085] in a preferred form of the invention, the predetermined period is at least one day.

Brief Description of the Drawings

[0086] The present invention wili now be described, by way of example only, with reference to one embodiment thereof and the accompanying drawing, in which:-

Figure 1 is a schematic flow sheet of a method for leaching zinc from an ore in accordance with the present invention;

Figure 2 is the results of the analysis of leached samples of ore showing recovery of zinc as function of ammonium carbonate concentration.

Best Mode(s) for Carrying Out the Invention

[0087] A method for leaching zinc from an ore in accordance with one embodiment of the present invention is now described. A zinc oxide ore is used as the basis for this disclosure, metal recovery is by solvent extraction and electrowinning. The flowsheet is shown in Figure 1. [0088] The ore 11 is mixed with the cure 10 and heaped in order to allow the mixture to rest 12, After an appropriate time the rested ore 13 is irrigated in the leach step 14 using zinc-depleted raffinate 15 from the solvent extraction unit 23. The pregnant leach solution 16 is sent to a storage pond 22 from which it I sent to solvent extraction 23 where the zinc is selectively removed using, for example, di-(2-ethylhexyl)phosphoric acid (DEHPA). The zinc is stripped into an acid solution 24 which passes to electrowinning 25 where zinc cathodes 26 are produced. The leached solids 17 are irrigated with water 19 in order to recover any entrained zinc and ammonia, the solution 20 being fed into the pregnant leach solution (PLS) pond 22. The barren solids 21 are disposed of in an appropriate manner.

[0089] Modifications and variations such as would be apparent to the skilled addressee are considered to fail within the scope of this invention.

EXAMPLE 1

[0090] Two different zinc ores, A and B containing 9.56 and 11.23% zinc respectively, were obtained. In order to obtain samples, ore A was initially crushed and ground. The ore was subjected to a sulphide flotation to remove the galena and the sphalerite as a concentrate leaving behind a residue, B. The major element analyses of A and B are given in Table 1. As can be seen, the flotation has eliminated the galena with both lead and sulphur decreasing. A small proportion of the zinc has been removed as the increase in zinc assay is less than that of calcium which is unaffected by flotation.

Table 1

A B

Ag 65 21 ppm

Al 3.08 2.74 %

Ca 6.37 8.15 %

Fe 19.93 29.81 % κ 0,74 0.77 %

Mg 0.36 0.43 %

Na 1.20 1.35 %

Pb 10.24 2.32 %

S 1.20 0.39 %

Si 1.10 1.09 %

Zn 9.56 11.23 %

[0091] X-ray diffraction analyses of the two samples showed the minerals present in each sample were as given in Table 2.

Table 2

[0092] The acid neutralisation capacities of the samples at pH2.0 were 240 and 300 kg H2SO4 / 1 for A and B respectively confirming the high carbonate content in the two samples. The higher value for B is a reflection of the higher proportion of carbonate remaining after removal of the mass associated with the lead and zinc sulphides.

[0093] Solutions containing 70g/L of free ammonia were made up with increasing concentrations of ammonium carbonate. The different samples were leached in all of the different solutions under identical conditions. After leaching, the solutions were sampled and analysed for zinc so that recoveries could be determined. The recovery of zinc as function of ammonium carbonate concentration is shown in Figure 2. [0094] The data at 0.1 g/L ammonium carbonate is for a run without any ammonium carbonate. As can be seen there is no obvious increase in zinc dissolution for the runs made with >5g/L ammonium carbonate added compared with those where the addition was <=5g/L.

[0095] As can be seen from the above results, the addition of ammonium carbonate is not required to maintain the pH of solution to that where zinc remains dissolved in the solution.

EXAMPLE 2

[0096] Fourteen different zinc ore samples were obtained from different areas within a single deposit. X-ray diffraction of the samples indicated that the zinc was present almost entirely as smithsonite, ZnC03 and the main gangue mineral was dolomite (Ca,Mg)C03, zinc-bearing goethite was also present in lesser amounts. The headgrade ranged from 2.2 to 28.0% Zn. The ore also contained 0.01-0.28% lead in the form of cerrusite, PbC03. In several samples, zinc was also present as hydrozincite and hemimorphite. The carbonate content of all samples was >20% by weight and this was confirmed by the Acid Neutralisation Capacity at pH 2.0 being >50kg H2SO4 / 1 for all samples.

[0097] The samples were crushed and ground in an identical manner then subjected to a 2h cure in 30g/L free ammonia followed immediately by a leach in either 30g/L free ammonia (30FA) and an ammoniacal ammonium carbonate solution (AAC) comprised of 68g/L free ammonia + 108g/L ammonium carbonate. All steps were performed under ambient temperature and pressure.

[0098] After 24h leaching time the solution was analysed to determine the recovery of zinc and lead. The zinc recovery is shown in Figure 3 and the lead recovery in Figure 4. To aid comparison, the recovery in AAC is plotted against the recovery in 30FA and the line shows a 1:1 ratio.

[0099] As can be seen, the zinc recovery in AAC is only slightly better than that in 30FA. This is despite there being more than double the amount of free ammonia and the presence of 108g/L ammonium carbonate. As indicated by the results, the effect of ammonium carbonate is small. Further analysis of the data showed that those samples where the recovery in AAC was notably higher than in 30FA were those where the zinc headgrade was highest.

[00100] As would be understood by those skilled in the art, the maximum solution concentration of zinc is controlled by the concentration of free ammonia as the tetra-amine zincate (Zn(NH3)4²⁺) complex is formed. Thus, the inventors believe that for the highest zinc headgrades there was insufficient free ammonia in the 30FA solution to dissolve the high quantity of zinc present in the sample. This was confirmed by increased zinc recovery in 30FA after leaching the samples at lower slurry density than shown here.

[00101] Broadly similar zinc leaching results were obtained after curing using 500g/L H2SO4 solution instead of the 30g/L free ammonia solution.

[00102] The extent of dissolution of lead is shown in Figure 4. To aid comparison, the recovery in AAC is plotted against the recovery in 30 FA and the line shows a 1 :1 ratio. As is clear, leaching in SAC resulted in significantly higher dissolution of lead than in 30FA.

[00103] In a process where the aim is to recover high grade zinc, the presence of lead will present additional problems, not only with separation of the lead but also with the need to recover and dispose of the lead in a suitable manner. The low level leached is unlikely to be sufficient for there to be a commercial recovery process for lead. Clearly, the presence of a high starting carbonate concentration in the leaching solution is going to lead to a requirement to remove a higher quantity of lead than from a solution without any carbonate present thereby increasing process costs. EXAMPLE 3

[00104] Three samples from different deposits, A, B and C were obtained, these had zinc headgrades of 10.9 %, 1 .2 % and 25.7 % Zn respectively. The zinc was present in samples A and C as predominantly smithsonite and hydrozincite, whilst sample B had a mix of smithsonite and hemimorphite. For all three samples the major gangue mineral was dolomite. All three ores had carbonate contents of >20wt % and Acid Neutralisation Capacities of >1 OOkg/t at pH1.5.

[00105] Samples were crushed, ground and cured for 24h using 30g/L free ammonia then subjected to leaching in solutions of 70g/L free ammonia (70FA) with 0-50g/L of ammonium carbonate (AC). After 24h the solutions were analysed to determine the extent of zinc dissolution which is shown in Figure 5. All steps were performed under ambient temperature and pressure.

[00106] As can be seen from the results, the increasing concentration of ammonium carbonate in the leach is largely ineffective at providing increased recoveries of zinc. Indeed, for sample A the effect is to reduce the recovery from ~93% to -75%. Sample B showed no clear increase in the presence of ammonium carbonate, sample C had a slight increase from -38 to -42%. The inventors believe that this suggests that a high carbonate concentration may be advantageous when there is substantial hemimorphite.

Claims

The Claims Defining the Invention are as Follows:

1. A method for leaching zinc from a carbonate-hosted ore, the method comprising the method steps of:

applying an aqueous solution of a curing agent to the carbonate-hosted ore, producing a cured ore; forming a leach solution by applying an ammoniacal solution that has an ammonium carbonate content of less than 5 g/L to the cured ore, producing a pregnant leach solution; and passing the pregnant leach solution to a means for recovering zinc.

2. A method according to claim 1 , wherein the leach solution maintains a pH between 6 and 13 during the leach process.

3. A method according to claim 1 , wherein the leach solution self-buffers to a pH between 8 and 1 during the leach process.

4. A method according to any one of the preceding claims, wherein the carbonate- hosted ore has an acid neutralisation capacity of at least 5 kg of sulphuric acid per tonne, when measured at pH 1.5.

5. A method according to any one of claims 1 to 3, wherein the carbonate-hosted ore has an acid neutralisation capacity of between 50 kg and 400 kg of sulphuric acid per tonne, when measured at phi 1.5.

6. A method according to any one of the preceding claims, wherein the zinc present as sulphide minerals is less than 25% of the total zinc in the ore.

7. A method according to any one of the preceding claims, wherein the ore comprises one or more minerals selected from the group comprising; calcite, galena, hydrozincite, goethite, calcite, hydrozincite, cerussite, smithsonite, sphalerite, and hemimorphite.

8. A method according any one of the preceding claims, wherein curing the ore to be leached through the application of an aqueous solution of a curing agent comprises substantially retaining the curing agent in contact with the ore to be leached when the ammonium carbonate solution containing free ammonia is applied.

9. A method according to any one of the preceding claims, wherein curing the ore to be leached through the application of an aqueous solution of a curing agent saturates at least 50% of a pore space of the ore with the aqueous solution of the curing agent.

10. A method according to any one of the preceding claims, wherein an ammonium carbonate concentration of the ammonium carbonate solution containing free ammonia is at least 5 g/L.

11. A method according to any one of the preceding claims, wherein a free ammonia concentration of the pregnant leach solution contains only a slight excess of free ammonia over that necessary to retain the target metal in solution.

12. A method according to any one of the preceding claims, wherein curing the ore to be leached through the application of an aqueous solution of a curing agent to produce a cured ore comprises: curing the ore to be leached through the application of a first aqueous solution of a first curing agent; and further curing the ore to be leached through the application of a second aqueous solution of a second curing agent.

13. A method according to any one of the preceding claims, wherein a volume of the aqueous curing agent applied to the ore to be leached is based on: a texture of the ore, a residence time; a concentration of the aqueous curing agent, or one or more leach conditions.

14. A method according to any one of the preceding claims, wherein a volume of the aqueous solution of the curing agent applied to the ore is sufficient only to moisten the ore.

15. A method according any one of the preceding claims, wherein the curing agent is a zinc complexing agent.

16. A method according to claim 16, wherein the curing agent comprises ammonia.

17. A method according to any one of claims 1 to 15, the curing agent is a reducing agent.