WO2024098097A1 - An improved process for recovery of metals from refractory ores - Google Patents

Abstract

A method of recovering a target metal such as gold, from a mineral source, includes the step of treating the mineral source under acidic conditions in the presence of a nitrogen oxide such as nitric acid at a temperature no greater than about 100°C and at about atmospheric pressure to thereby recover the target metal from the mineral source. An aqueous leach solution is produced comprising jarosite that comprises one or more other target metals that have been exogenously added to the aqueous leach solution and/or originally present in the mineral source. Treating the mineral source under acidic conditions in the presence of nitric acid may be performed using a pipe reactor.

Description

TITLE AN IMPROVED PROCESS FOR RECOVERY OF METALS FROM REFRACTORY ORES TECHNICAL FIELD THIS INVENTION relates to mineral processing. More particularly, this invention relates to provides recovery of refractory metals such as gold and other metals of commercial value from metal sulfide ores. BACKGROUND Gold in the earth’s crust is scarce and typically at concentrations of only three parts per billion. As a result, gold ores carrying more than 5 parts per million are potentially valuable even prior to extraction. This equates to 5 grams per tonne in mass terms. However, with the rapid depletion of “easy-to-treat” gold deposits and for gold mining companies to remain commercially viable, increased attention is being directed to the technology of recovering gold from refractory deposits. Refractory gold typically consists of extremely fine-grained gold particles either physically encapsulated within, or chemically bound (often described as “in- solid-solution”) within one or more other minerals. For example, refractory gold may occur as a component of sulfides (e.g. pyrite, arsenopyrite, and chalcopyrite), silicates, carbonates and/or oxide minerals. In these ‘solid-solution’ forms gold is physically difficult to find even with scanning electron microscopy (SEM), for example. Thus, these refractory gold deposits are metallurgically complex and are often difficult and expensive to process and recover the gold. This complexity has resulted in a growing proportion of refractory gold resources relative to overall (global) gold resources, particularly evident in the tailings of many thousands of gold mines. Refractory gold ore is highly resistant to extraction by conventional chemical processing. Such ore responds poorly to traditional gold cyanidation techniques such as Carbon-in-Leach (CIL), often yielding less than 80% of the contained gold. Historically when the gold price was less than US$300 per ounce (generally up until 1980), the loss of 20% in poor recoveries, whilst problematic, was unable to justify the added cost of an ‘improved’ process and was frequently ignored as a result. Improved processes historically, generally involved stronger oxidants, higher pressures and higher costs. As the prices of precious metals have risen over the last decade and a half, greater attention has been given to improved processes for the treatment of refractory gold ores, with particular emphasis on well-established techniques such as Roasting, Pressure Oxidation and the developing technology of Biological Oxidation.

However, given the relative scarcity of gold containing particles within the ores, these processes are relatively expensive and their requirement for aggressive oxidation results in high corrosion rates and hence significantly greater investment in plant and equipment. This equates to higher CAPEX and OPEX. For example in these processes the need for efficient gas scrubbing to capture evolved gases for recycle has been identified as one of the hurdles to broader implementation.

The developing interest in low-cost biological processes has as yet been unable to deliver the throughput required or expected for a commercial process plant.

Smelting of gold is a secondary metallurgical process with significant costs associated with the high costs of thermal energy. Within tailings and particularly waste, by-product and refractory gold deposits, commercial viability usually requires a lower cost form of pre-treatment, of the gold containing minerals, before smelting - due to the low grade of gold in the ore. (Tailings can often contain less than 1 g/t of gold or 1 part per million by weight).

Common pre-treatment processes involving water are classified as hydro- metallurgical processes. Those commonly used for low-grade gold ores and tailings include: gravity separation such as jigs, hydro-cyclones and wet sluicing; froth flotation: and microbial oxidation, particularly associated with 'heap' leaching.

Within polymetallic deposits, froth flotation has grown to dominate the primary metallurgical processes and the products of this process are concentrates or 'cons'. It is industry practice that such concentrates are shipped to specialised secondary smelters who rely on high temperature thermal processes to separate the respective metals. There are commercial and historical reasons for the dominance of this approach, principally the practice of the smelter paying for contained precious metals after deducting a relatively minor processing fee. Given the high CAPEX of the smelting process, this 'toll based' approach has been extremely successful in attracting smaller mining companies since it avoids the need to duplicate resources.

Silver is generally associated with gold however also has an affinity for lead and zinc containing minerals. In particular, silver associates with sulfide minerals such as galena and sphalerite but also the iron sulfides pyrite, pyrrhotite and marcasite, frequently found together in important zinc ore bodies. Gravity separation is less effective in separating blende, sphalerite and galena from pyrite because the density or 'specific gravity' (SG) of pyrite is typically too similar. However this poses an important challenge for the industry because the presence of iron in zinc ores acts as a reductant and is considered undesirable for metallurgical reasons.

Most metals occur in nature in their ores in some form of oxidized metal and must be reduced to their metallic forms. After processing into some form of concentrate by gravity or flotation, the ore is often dissolved into an aqueous electrolyte and the resulting solution is purified to enable recovery of the dissolved metals. There are many ways to recover the metals from solution.

In the gold industry, the use of activated carbon to separate gold from cyanide leaching solutions has become an industry standard for efficient gold recovery in Carbon-in-Leach (CIL), Carbon-in-Pulp (CIP) and Carbon-in-Column (CIC) operations. Activated carbon, a material produced from carbon-rich sources, offers an incredibly porous surface structure, which creates a vast surface area (about an acre per % teaspoon) on which to adsorb materials. This porous structure, in combination with attraction forces, allows activated carbon to capture material components and hold on to them for later recovery.

The gold can then be desorbed from the loaded carbon in a process known as elution (also referred to as stripping), which produces a high gold concentrate solution from which gold can be electrowon using electrolysis. The metal in this case is deposited on the cathode.

Several metals are naturally present as metal sulfides these include copper, lead, zinc, nickel, cadmium, cobalt, molybdenum and silver. In addition, gold and platinum group metals are often associated with sulfide ores. Metal sulfides or their salts, are generally electrically conductive enabling electrochemical redox reactions to efficiently occur in either the molten state or in aqueous solutions. For this reason, electrolytic cells can use anodes comprising mixed metal sulfides however the functionality of this operation leaves much to be desired. SUMMARY In a broad form, the invention provides a method, apparatus and/or system for recovering a target metal from a mineral source. In one broad embodiment, acid treatment of the mineral source under acid conditions at relatively low temperature and pressure facilitates recovery of one or metals, such as gold, from the mineral source. In a further embodiment, one or more other target metals may subsequently be recovered after forming a target metal-jarosite complex. In one broad aspect, the invention provides a method of recovering a target metal from a mineral source, said method including the step of treating the mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to thereby recover the target metal from the mineral source. In a particular aspect, the invention provides a method of recovering a target metal from a mineral source, said method including the step of treating the mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to thereby recover the target metal from the mineral source and produce an aqueous leach solution comprising jarosite, or a jarosite-related mineral, that comprises an extraneous other target metal added to the aqueous leach solution and the said other target metal originally present in the mineral source. Suitably, the nitrogen oxide is of general formula NxOy, wherein X is 1 or 2 and Y is 1, 2, 3 or 4. Non-limiting examples of nitrogen oxides include NO, NO₂, NO₃, N₂O₄, inclusive of acids such as HNO3 and metal (e.g alkali metal, alkaline earth metal and transition metal) salts such as NaNO3, KNO3, NaNO2, NH4NO3, Ca(NO3)2, Mg(NO3)2 and Fe(NO₃)₃. Suitably, initiation of the treatment is by addition of nitric acid (HNO3) to the mineral source. Suitably, in some embodiments further treatment occurs in the presence of one or more other nitrogen oxides as hereinbefore described. In this context, it will be appreciated that the method of treatment is preferably a “continuous flow” method. In one embodiment initiation with nitric acid is followed by generation of gaseous phase nitrogen oxides such as NO, NO2, N2O4 enabling the subsequent use of one or more other nitrogen oxide , such as those at higher valence states (+3, +4) including NaNO₃, KNO₃, NaNO₂, NH₄NO₃, Ca(NO₃)₂, Mg(NO3)2 or Fe(NO3)3. In an embodiment, the other nitrogen oxide is calcium nitrate. Thus, consumption of nitric acid is minimized, in which case nitric acid is essentially a catalyst, as it is not substantially consumed during treatment. Typically, the mineral source is mining waste, ground electronic waste, mining ores, tailings etc. Suitable mineral sources comprise, metal sulfide ores such as pyrite, arsenopyrite, chalcopyrite, sphalerite, galena, tetrahedrite, argentite and/or other ores typically of volcanogenic metal sulfide (VMS) origin. In one embodiment, the target metal is extracted as a sulfur-metal complex. In an embodiment, the target metal is a precious or noble metal such as gold. According to this embodiment, the sulfur-metal complex is a gold-disulfide complex. Suitably, the method of this aspect produces an aqueous solution, referred to herein as an “aqueous leach liquor” or “aqueous leach solution” which comprises one or more other target metals originally present in the mineral source. In an embodiment, the method includes the step of recovering the one or more other target metals from the aqueous leach solution. It will be appreciated that the method described herein includes production of a jarosite or jarosite-related mineral comprising the one or more other target metals. In one particular embodiment, jarosite or jarosite-related mineral is produced by adding an extraneous or exogenous target metal to the aqueous solution under conditions that facilitate formation of jarosite comprising the extraneous or exogenous target metal. Furthermore, the jarosite may comprise the other target metal(s )originally from the mineral source and present in the aqueous leach solution. In one particular embodiment, the extraneous or exogenous target metal is added as a water-soluble salt. In another particular embodiment, the extraneous target metal is added electrochemically, such as by the use of a sacrificial anode comprising the extraneous or exogenous target metal. In some embodiments, the one or more other target metals include lead, silver, cobalt, zinc, copper, cadmium, chromium and nickel, although without limitation thereto. In a further embodiment, the method includes treating the jarosite or jarosite-related mineral to thereby produce an iron oxide. Non-limiting examples of iron oxides include ferrous and ferric oxides such as magnetite (Fe3O4) and maghemite ( ^Fe₂O₃). Suitably, the step of treating the mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure is performed in a reactor vessel of a generally tubular configuration, such as a pipe reactor. It will be appreciated that the reactor vessel operates in a closed system with loading and unloading provisions well known to those in the art, such that minimal or substantially no gases can escape without deliberate venting . Another aspect of the invention provides a target metal recovery system or apparatus comprising: a reactor vessel for treating the mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to facilitate recovery of the target metal from the mineral source. Preferably, the reactor vessel is of a generally tubular configuration, such as a pipe reactor. As previously described, the reactor vessel operates in a closed system with loading and unloading provisions well known to those in the art, such that minimal or substantially no gases can escape without deliberate venting. In further embodiment, the system or apparatus further comprises a vessel for producing jarosite or a jarosite-related mineral from the aqueous leach solution. Suitably, the jarosite is substantially crystallized jarosite. In yet another embodiment, the system or apparatus further comprises a seeding vessel for producing jarosite or a jarosite-related mineral comprising one or more other target metals. In particular embodiments, the system and/or apparatus may include further components such as one or more slurry tanks, grind mills, flotation tanks, magnetic separators and pelletizers, as will be described in more detail hereinafter. Throughout this specification, unless otherwise indicated, "comprise", "comprises" and "comprising" are used inclusively rather than exclusively, so that a stated integer or group of integers may include one or more other non-stated integers or groups of integers. It will also be appreciated that the indefinite articles "a" and "an" are not to be read as singular indefinite articles or as otherwise excluding more than one or more than a single subject to which the indefinite article refers. For example, "a" metal includes one metal, one or more metals or a plurality of metals. The term “about” is used herein to refer to a tolerance or variation in a stated amount. The tolerance or variation may be no more than ± 10%, ± 9%, ± 8%, ± 7%, ± 6%,± 5%, ± 4%, ± 3%, ± 2%, ± 1% of a stated amount. BRIEF DESCRIPTION OF THE DRAWINGS Reference is made to non-limiting embodiments described in the drawings wherein: FIG. 1 shows a schematic overview of an embodiment of a target metal recovery system comprising a pipe reactor for acid treatment of a mineral source; FIG. 2 shows a schematic overview of an embodiment of a system for recovering one or more other target metals subsequent to acid treatment; FIG.3 shows an example of leach temperatures and oxidation reduction potential (ORP) achieved during acid leach in a continuous reactor; FIG.4 shows another example of leach temperatures and ORP achieved during acid leach in a continuous reactor; FIG.5 shows an extraction profile for target metals from the leach in FIG.3; FIG.6 shows an extraction profile for target metals from the leach in FIG.4; FIG. 7 shows another example of leach temperatures and ORP achieved during acid leach in a continuous reactor; FIG.8 shows an extraction profile for target metals from the leach in FIG.7; FIG. 9 shows another example of leach temperatures and ORP achieved during acid leach in a continuous reactor (10% w/w, 11% nitric acid, 1 hr residence time); FIG.10 shows an extraction profile for target metals from the leach in FIG.9; and FIG.11 shows Laser Size Analyses cumulative size distribution on a mineral feed sample. DETAILED DESCRIPTION This invention is at least partly predicated on the discovery that target metals, such as gold, present in “solid solution” in sulfide mineral sources such as crushed mining ores, tailings, concentrates or crushed, ground electronic waste can be extracted from the mineral source as a gold sulfide compound following acidic nitrogen oxide treatment at atmospheric pressure and relatively low temperatures below 100^oC. In an additional step, other target metals such as silver and lead may be extracted by forming jarosite, or a jarosite-related mineral, following nitric acid treatment and “seeding” the jarosite with extraneous target metal, such as by adding extraneous target metal salt and/or by way of a sacrificial anode comprising extraneous target metal or target metal salt. This method also facilitates the extraction of iron in the form of iron oxides such as magnetite and maghemite. A particular feature of the invention is that the consumption of expensive nitric acid can be relatively minimized. Furthermore, the weak acid, low temperature and low-pressure conditions reduce the likelihood that the reaction vessel used for nitric acid treatment will corrode and also enable suitable plastics to be used for the reaction vessel. In a broad form, the invention provides a method of recovering a target metal from a mineral source, said method including the step of treating the mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to thereby facilitate the target metal from the mineral source. In a particular aspect, the invention provides a method of recovering a target metal from a mineral source, said method including the step of treating the mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to thereby recover the target metal from the mineral source and produce an aqueous leach solution comprising jarosite that comprises an extraneous other target metal added to the aqueous leach solution and the said other target metal originally present in the mineral source. As used herein, “acidic conditions” refers to a pH below about pH 4 or 3.5, preferably below about pH 3, more preferably below about pH 2.5 and advantageously at about pH 1.5 -2, inclusive of about pH 1.6, 1.7, 1.8, 1.9 and 2.0. Suitably, the nitrogen oxide is of general formula N_xO_y, wherein X is 1 or 2 and Y is 1, 2, 3 or 4. Non-limiting examples of nitrogen oxides include NO, NO2, NO3, N2O4, inclusive of acids such as HNO₃ and metal (e.g transition metal, alkali earth metal and alkali metal) salts such as NaNO₃, KNO₃, NaNO₂, NH₄NO₃, Ca(NO₃)₂, Mg(NO₃)₂ and Fe(NO3)3. Suitably, initiation of the treatment is by addition of nitric acid (HNO3) to the mineral source. Suitably, further treatment occurs in the presence of one or more other nitrogen oxides as hereinbefore described. In this context, it will be appreciated that the method of treatment is preferably a “continuous flow” method whereby initiation with nitric acid is followed by generation of gaseous phase nitrogen oxides such as NO, NO₂, N₂O₄ and, optionally, subsequent addition of one or more other nitrogen oxides at higher valence states (+3, +4) including NaNO3, KNO3, NaNO2, NH4NO3, Ca(NO3)2, Mg(NO3)2 or Fe(NO₃)₃. In an embodiment, the other nitrogen oxide is calcium nitrate. Thus, consumption of nitric acid is substantially minimized., . Although not wishing to be bound by theory, in this context nitric acid behaves essentially as a catalyst, as it is substantially not consumed during treatment. By “substantially minimized” or “substantially not consumed” is meant that no more than about 10%, no more than about 9%, no more than about 8% no more than about 7% no more than about 6%, no more than about 5%, no more than about 4% or no more than about 1-3% of the initiating nitric acid is consumed. In this context and by way of example, nitric acid oxidation of pyrite generally occurs as described in Equation 1: 2FeS2 + 10HNO3 = H₂SO₄ + Fe₂(SO₄)3 + 10NO + 4H₂O Nitric oxide gas produced is further oxidized by oxygen shown in Equation 2: 2NO + O₂ = 2NO₂ Accordingly, a source of O2 is provided to facilitate this reaction. Nitrogen dioxide can then be absorbed in water to regenerate nitric acid as shown in Equation 3: 3NO2 +H₂O =2HNO3 + NO The overall reaction is shown in Equation 4: 4FeS₂ + 15O₂ + 2H₂O = 2H₂SO₄ + 2Fe₂(SO₄)₃ As described in the equations above, nitric acid can be mostly recovered in the process, and when regenerated or recycled efficiently and not substantially consumed in the method, can be considered to be a catalyst By way of comparison, typical prior art methods have a stoichiometric requirement of about 5 moles nitric acid in the reaction per mole of pyrite, whereas the present invention achieves complete decomposition of pyrite with 1/10th of the typical stoichiometric requirement (i.e.0.5 moles). Suitably, the treatment is performed at atmospheric pressure and temperatures not exceeding 100^oC. By keeping temperatures and pressure relatively low, the method may avoid the use of metallic materials in favour of reinforced advanced plastics and could also at least partly minimized corrosion that occurs in high pressure reaction vessels. Suitably, the temperature is less than 100^oC. Non-limiting examples include 30^oC, 35^oC, 40^oC, 45^oC, 50^oC, 65^oC, 70^oC, 75^oC, 80^oC, 85^oC, 90^oC, 95^oC inclusive of ranges between any of these stated values. In one broad form, the invention provides acid oxidation of the mineral source at atmospheric pressure. By “atmospheric pressure” is meant about 1 atmosphere (1 atm) which is equivalent to 101,325 Pa (1,013.25 hPa), 1013.25 millibars, 760 mm Hg, 29.9212 inches Hg, or 14.696 psi.1 atm unit is roughly equivalent to the mean sea-level atmospheric pressure on Earth. Operationally, this definition means that no external source of pressure is used to increase or decrease the internal pressure of the reaction vessel (e.g. pipe reactor) during acid treatment. Suitable reaction times, defined as the time it takes for an ore particle of ore to travel from the start to the end of the reaction vessel, where the particle is being subjected to aggressive leaching to solubilize and contained minerals, may be in the range of about ten (10) minutes to about six (6) hrs. By way of example, for arsenopyrite and pyrite, the residence time can vary from about 30 minutes to about 2 hrs (120 minutes) or longer depending on (1) the acid concentration (2) the slurry density; and (3) the ease of oxidation of the various contained minerals. As used herein, a “target metal” is any metal of economic or commercial value. Typically, metals include precious metals, noble metals and/or transition metals including, but not limited to, iron, zinc, lead, nickel, cobalt, platinum, palladium, manganese, copper, silver and gold and other metals such as aluminium. The target metal may be in atomic or ionic form (e.g as a salt), or a combination of these in the mineral source. Typically, the mineral source is, or comprises, or is present in mining waste, ground electronic waste, mining ores, tailings etc. Suitable, the mineral source is, or comprises, metal sulfide ores such as pyrite, arsenopyrite, chalcopyrite, sphalerite, galena, tetrahedrite, argentite and/or other ores typically of volcanogenic metal sulfide (VMS) origin. Mineral source particle size is generally in the range of 1 micron to 75 microns and more typically in the 10-40 micron particle size. Frequently gold amenable to this process is found within the finer fractions such as for example that passing a 200 mesh per inch Tyler screen (74 microns) and even more preferably in the minus 400 mesh (37 micron) fraction. Bond Energy, also known as average bond enthalpy or simply bond enthalpy, is a quantity that reflects the strength of a chemical bond. The bond energy of a chemical bond in a given compound can be visualized as the average amount of energy required to break one such chemical bond. In this context, sulfur ions, and particularly the disulfide ion (S^2- ), bind strongly to the gold surface. This is consistent with the strong association between non-metals such as carbon, with gold, as used to recover gold from cyanide solutions within the gold mining and processing industry. When disulfides are attached to gold surfaces, this may be illustrated schematically as an internal disulfide bond, although the exact nature of the bond remains obscure. However the gold-sulfur bond is quite strong and exceeding 126 kJ/mol. With such strong bond energy the reaction is generally considered to be practically irreversible. With small particulate gold, such as that released from ‘solid solution’ with pyrite minerals, the role of the gold-sulfur bond is regarded as important. Although not wishing to be bound by theory, an effective process for oxidation of refractory sulfide ores requires an understanding the nature of the gold- sulfur bond, since this chemical bond was competing with the conventional use of another non-metals, namely carbon, in the recovery of gold and silver from pregnant cyanide liquors. The inventor also noted that many polymetallic mines which include refractory sulfides with their ore body, generally were focussed on the production of a lead and zinc concentrates using froth flotation. For this reason, the inventors wished to identify an improved process which best integrated the oxidation pre- treatment process with the existing infrastructure, while not interfering with the recovery of these concentrates by conventional froth flotation. However a key commercial consideration is that some concentrates, such as lead sulfide, are sent to smelters who also recover any contained gold or silver. It is customary with these arrangements, for the shipper to be paid for these metals at a negotiated price, less the costs associated with recovery, typically 3%. The inventors reasoned that recovering 97% of the value of a component by including it in a lead concentrate was likely to be the most cost-effective way to monetise that component. By way of example, a first stage treatment (“leach”) with nitric acid was shown to initially oxidise the galena, PbS, component of many of the refractory ores and tailings ore and also commence oxidation of the sphalerite, (Zn,Fe)S, and chalcopyrite, CuFeS₂ components. The dissolution of these minerals was found to generally have a neutral reaction temperature of less than 30^oC. By way of example, a second stage leach with nitric acid was found to complete the oxidation of the remaining sulfides including the arsenopyrite and pyrite. These minerals contain the bulk of the precious metals (e.g. gold) within their crystalline structures described as being in solid solution. Testing of this stage demonstrated a generally exothermic reaction but with maximum temperature approximately 90^oC. Following these two stage leaches, the residue was shown to contain >90% of the gold and silver originally contained within the ore now amenable to subsequent recovery, such as via conventional cyanidation. The identified improved process for leaching of refractory tailings or ores proposes use of low acid concentrations, low to moderate reaction temperatures and normal atmospheric pressure. These parameters support potential for low capital costs for materials of construction and lower concerns regarding material corrosion. By way of comparison, in other processes such as the Nitrox process, nitric acid consumption was expected during modelling studies to be a major cost consideration due to escape of NO and NO₂ gases. In contrast, according to the present invention the nitric acid leach utilizes the manufactured sulfuric acid (see Eq.1) and uses nitrogen oxide (e.g nitric acid) as a catalyst and which is not significantly consumed in the reaction. Any potential loss of NO and NO₂ gases may be at least partly avoided overcome by modifying the acid leach circuit to eliminate potential for off gassing. In addition this invention overcomes the traditional challenges to achieve NO and NO2 off-gas recovery and recycling. Recycling of nitric acid was previously used at other locations in order to alleviate the high cost of nitric acid as an oxidant. The most widely used recycling technique was wet scrubbing using a water mist - however the low solubility of nitrous oxide in water required multiple scrubbing stages. Wet scrubbing was found to be particularly ineffective for NO prior to its oxidation to NO₂ due to the significantly lower solubility of the gas. This is believed to be one of the key factors determining the apparent inability to recover nitric acid consumed in previous nitric acid based leaching processes. To improve solubility of nitrogen oxide gases (NO, NO₂) previous developments have included adding an alkali to water used in the scrubbing process to accelerate dissolution of the acid fumes. As an example by incorporating a weak calcium hydroxide alkali into the scrubbing liquor whereby the nitrogen oxides were favourable dissolved to form chemically stable calcium nitrate. This solution was then used as the main scrubbing liquor, increasing the solubility of any escaping gases by several orders of magnitude. This goes some way to eliminate losses of nitrogen oxides (NOx) to the circuit but the inventors have proposed another mechanism. As hereinbefore described, treatment of mineral sources (e.g sulfide ores) under acidic conditions, such as nitric acid, at temperatures below about 100^oC and at atmospheric pressure, facilitate release of metals such as gold and silver from being in “solid solution” within the mineral source, whereby the released metals may be subjected to further processing for retrieval. As generally used herein, the term “aqueous leach liquor” , “aqueous leach solution” or “pregnant leach solution (PLS)” refers to the aqueous solution produced following acid treatment of the mineral source. In one embodiment, the invention includes removal of arsenic from the acid leach liquor. Arsenic, when present in arseno-pyrite, becomes soluble in the first stage leach and rapidly builds to the point that requires removal. As there is no commercial market for arsenic, it is preferred to react it with ferric ion and modify the pH sufficiently to enable removal of arsenic from the solution by precipitating as ferric arsenate, the mineral scorodite. In a further embodiment, the method includes the step of recovering one or more other target metals from the leach liquor. Non-limiting examples of one or more other target metals include iron, zinc, lead, cobalt and copper. Various hydrometallurgical, galvanic and electrowinning and related electrorefining processes exist to facilitate the recovery of copper, nickel and zinc, but there has been limited success with the metals aluminium, iron, lead and cobalt other than in specialised applications such as refining from purified alumina. Recovery of iron is rarely attempted in ores of less than shipping grade, typically in the range 56-59% Fe. As commercially available pyrite is typically only 33-35% metallic iron, it has been generally ignored as an ore of iron. Alternative mechanisms to recover metals include a wide range of solvent extraction (SX) and ion-exchange (IX) resins which can be used to attract metal ions for later recovery by unloading of the resins. There are also numerous hydro chemical methods with varying degrees of efficiency. Finally, there are numerous electrolytic and electrowinning processes where cathodes are used to recover metals from within solution. These various approaches are well documented and understood by persons skilled in the art. In work leading to the present invention, the occurrence of the mineral jarosite in waste emissions from electrolytic zinc smelting was noted. One form comprises potassium and has the chemical formula of KFe₃(SO₄)₂(OH)₆. Other forms also exist including ammonio-jarosite, argento-jarosite, natro-jarosite, and plumbo- jarosite, in which ammonium, silver, sodium, and lead, respectively, substitute for potassium, such as PbFe₃(SO₄)₂(OH)₆. Subsequent work has identified other jarosite- related minerals such as beudantite, such as shown in Tables 13 and 14. Accordingly, the term “jarosite” as used herein encompasses other jarosite-related minerals including but not limited to beudantite and segnitite. It is also proposed that the choice of alkali being used for neutralisation was the primary determinant. Using aqueous ammonia (NH₄+) and sodium hydroxide (NaOH) as alkalis respectively, both ammonio-jarosite NH₄Fe₃(SO₄)₂(OH)₆ and natro- jarosite NaFe3(SO₄)2(OH)6 formed. Given the possible presence of argento-jarosite and plumbo-jarosite, it is proposed that depending on the initial concentrations in the mineral source, it is preferable to recover silver and lead. Although Jarosite minerals have high water solubility, they generally form crystallites at increasing concentrations and can be separated from solution when component minerals approach, or reach saturation, in the15pproxr. This may be facilitated by ‘seeding’, where the formation of a particular double salt can be encouraged or facilitated by the addition of the corresponding metal ion(s). In this process, a semi-saturated solution containing various ion-species, can be encouraged to form ion-pairs by the addition of ions to an excess. This has the effect of exceeding the saturation level, which can be further enhanced by (for example) reducing the temperature of the solution whilst maintaining the concentration. The process of ‘seeding’, provides multiple sites or ‘nuclei’ for the soluble ions to form first ion-pairs and then in turn, crystallites, around the nuclei. These can then be separated for further refining – typically electrolysis, known to those skilled in the art, whereas without this intervention they would become entrained in the residue as insoluble forms such as lead sulfate or silver chloride at levels unfavourable for further recovery. In one embodiment, a metal “seed” is added as a water-soluble salt of the metal In another embodiment, the metal “seed” is added electrochemically by the use for example, of a sacrificial anode. Typically the use of a water-soluble salt would be more cost effective than using a sacrificial anode of a metal. However, other economic considerations include: where the cost of soluble salts is greater on a unit-component basis than the actual metal – due to manufacturing costs, freight and related matters (e.g. copper sulfate contains 25% copper but typically costs 65-75% of the pure metal when purchased as copper cathodes. Hence the unit cost of delivered copper (excluding energy) can be (75/25) or 3 times the cost of using copper as a sacrificial anode; similarly waste or by-product metal compounds can be added to the semi-saturated solution to trigger crystallite formation; zinc-containing wastes are an example. Where energy costs are subsidised or artificially low due to an abundance of either renewable energy (e.g. wind, hydro, solar), anodic oxidation is relatively inexpensive compared to addition as a soluble salt in providing both OH- and atomic H (hydrogen); and a cation has more protons than electrons, giving it a net positive charge, and using a sacrificial anode avoids the need to add an additional anion (e.g sulfate in the example of copper sulfate) but in the case of the jarosite series, avoids adding potassium, sodium or ammonium (K⁺, Na²⁺ or NH₄ ⁺) as alkalis, thereby avoiding unnecessary addition of ion species that are already in (relative) abundance (for the required stoichiometry). In embodiments where a sacrificial anode is used for seeding, reducing conditions can be generated as hydrogen gas formation occurs from water electrolysis; reduction of nitrite anions (NO²⁻ ) to firstly nitric oxide (NO) and then nitrous oxide (N₂O) allows potential nitrate losses to tailings to be reduced by returning active catalysts to the oxidation process; reduction of ferric Fe³⁺ ion to ferrous Fe²⁺ ion provides a pathway to recover magnetic iron oxides maghemite or magnetite as discussed below; and/or electrolysis of water continually generates hydroxyl ions (OH-) which in a closed system serves to raise the pH thereby reducing jarosite solubility, thereby enriching the jarosite in the sacrificial metal, e.g. lead or zinc, without the need for a balancing sodium, ammonium or potassium cation as is the case when alkali addition is required (e.g. NaOH, NH4OH or KOH). The latter can have implications where concentrates are generated for smelting since lead concentrates for example are relatively tolerant of lead as sulphate or plumbo-jarosite. With regard to the above reduction of ferric Fe³⁺ ion to ferrous Fe²⁺ ion, jarosite forms as the liquor leaving the reactor cools. This cooling process can be expedited by extracting heat via ammonia gasification from an aqueous form. This is the essence of recovering gaseous ammonia from an aqueous NH₄+ ion source. This is an alternative to the use of quicklime which generates heat exothermically. Following jarosite crystallisation, the liquor is essentially a mixed ferric sulfate mixture. (Fe₂(SO₄)₃ ). This cannot be used alone to generate magnetite, as it is in ferric Fe ³⁺ state. This requires a stoichiometric balance of ferrous Fe ²⁺ ions. Accordingly, the sacrificial anode at the jarosite stage adds both hydroxyl ions to raise the pH as well as hydrogen as a reductant to reduce Fe ³⁺ into the Fe ²⁺ state by donating an electron at the anode contributed by the formation of first H⁺ and then H and H₂ (gas) at the cathode . Other metals can be added or removed as desired using similar processes particularly the use of sacrificial anodes to encourage early seed formation in the case of jarosite minerals or recovered cathodically where a metal plates effectively using such galvanic reactions e.g. copper. After arsenic has been removed as scorodite as described above, further recovery of iron from solution is obtained by further raising the pH and precipitating ferric hydroxide. However, the present invention realizes that alkali consumption is a major cost consideration and hence it is advantageous to limit alkali consumption. This has been achieved by limiting pH increase above 2.7 and return of this liquor to the nitric acid treatment step. The recovery of ferric hydroxide sludges is problematic mainly because polymeric ferric hydroxide lattice incorporates massive volumes of water and the filtration and recovery of the material is grossly inefficient. In the past such sludges have been relegated to the tailings dam where they settle extremely slowly and contribute to the waste materials. Instead, according to the present invention, ferric hydroxide sludge is used as an intermediate in a process for the more efficient recovery of magnetic iron oxide as magnetite. A feature of the current invention is that part of the jarosite earlier recovered at low pH by crystallisation can be used in the subsequent conversion stage to prepare a magnetic iron ore mineral such as maghemite ^Fe₂O3 and/or magnetite Fe3O4. The process involves first reducing a portion of the ferric sulfate in semi saturated solution to ferrous sulfate. The mechanism to achieve this first depends on the other metals present in the leach solution. Many Volcanogenic Metal Sulfide (VMS) orebodies carry a polymetallic suite of metals, typically lead, zinc, copper, cadmium, chromium and nickel. If these metals are present in the tailings it demonstrates that : the primary flotation 18pprox.s to remove them from the ore as sulfide minerals has almost certainly failed; and they are probably present in oxidised forms generally amenable to acid dissolution and hence recovery; or lead, in particular, will be lost to tailings residue as the insoluble lead sulfate if an alternative process step cannot be developed. For example if lead is present in the initial leach but reacting with sulfate to form insoluble lead sulfate, it is preferable to enable the lead to be converted to plumbo-jarosite since this will enable it to remain in solution until further dissolution or recovery of lead is no longer possible. Another aspect of the invention includes a target metal recovery system or apparatus comprising a reactor vessel for treating the mineral source with under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to facilitate extraction of the target metal from the mineral source. In this context, reference is now made to FIGS 1 and 2 in describing a particular embodiment of the present invention. Suitable ores for treatment may include pyrite, arsenopyrite, chalcopyrite, sphalerite, galena, tetrahedrite, argentite and/or other ores typically of volcanogenic metal sulfide (VMS) origin. Non-limiting examples of catalysed electrochemical oxidation processes are: CuFeS₂ ^ CuS +·Fe²⁺ + S + 2e ZnFeS₂ ^ ZnS +·Fe²⁺ + S + 2e PbFeS₂ ^ PbS +·Fe²⁺ + S + 2e NiFeS₂ ^ NiS +·Fe²⁺ + S + 2e CoFeS₂ ^ CoS +·Fe²⁺ + S + 2e MnFeS₂ ^ MnS +·Fe²⁺ + S + 2e As shown in FIG.1, in leach apparatus 100 comprises pipe reactor circuit 10 for treatment of sulfide ores with nitric acid, referred to as the ”primary leach”, which occurs in pipe reactor 17 after mixing in slurry tank 16 with process water from source 11. The primary leach reaction occurs in at pH about 1.5-2 breaks down sulfide ores such as arsenopyrite and pyrite from mineral source 12 (such as finely ground tailings from a VMS orebody) with dilute nitric acid from supply tank 14, but as the exothermic reaction proceeds and reaches operating temperature, the generation of sulfuric acid from the pyrite requires only that nitrogen oxides (N0, N2O, N02) act in a catalytic role, being neither generated nor consumed in the reaction. It will therefore be appreciated that the method of treatment is a “continuous flow” method whereby initiation with nitric acid is followed by generation of gaseous phase nitrogen oxides such as NO, NO₂, N₂O₄ enabling the subsequent use of other sources of nitrogen oxide catalysts at higher valence states (+3, +4) including NaNO3, KNO3, NaNO2, NH4NO3, Ca(NO3)2, Mg(NO3)2 or Fe(NO3)3. These are provided by supply tank 13. An exogenous source of additional O₂ may be provided, although not shown in FIG.1. The decision to add oxygen or oxygen sources such as air, is determined by the extent to which sulphur is desired as a by-product in order to entrap gold as herein described. If this is not a requirement, then the sulphur can be conveniently further oxidised by the addition of oxygen in some gaseous, solid or liquid form. Examples include peroxide, solid persulphates and gaseous oxygen. As nitrogen oxide gas escape present potential losses of reactants, it is preferable that the pipe reactor 17 entrains any evolved gases and prevent their escape from the pipe reactor 17. Furthermore, as the various nitrous oxides vary in their water solubility, with NO₂ being the most soluble and N₂O the least, the design of the pipe reactor 17 overcomes the need to have one or more water scrubbers to recapture these gases as and when they are evolved. This design ensures the maximum possible recovery of the catalytic gaseous reactants compared to previous designs. Pipe reactor 17 typically is a metal or HDPP pipe that can operate up to 120^oC for a desired residence time. The residence time for the pipe reactor 17 is defined as the time it takes for a particle of ore to travel from the start to the end of the reaction vessel, where the particle is being subjected to aggressive leaching to solubilize and contained minerals. For arsenopyrite and pyrite, the residence time can vary from 30 minutes to 2 hrs (120 minutes) or longer depending on (1) the acid concentration (2) the slurry density; and (3) the ease of oxidation of the various contained minerals. Accordingly the pipe reactor 17 residence time is primarily a feature of the dissolution characteristics of the various ores proposed to be leached. Suitably, by appropriate selection of terrain to locate the pipe reactor 17, pumping can be minimised by utilizing gravity flow wherever possible. Depending on (a) the particle size of the ore in microns (b) concentration of the acid, in ww % and (c) the slurry density (gm/L) the reaction can become exothermic and rapidly achieve temperatures in the range 60^oC- 90^oC. Accordingly, pipe reactor 17 can be either metallic or polymer (e.g. HDDP) based, providing the material chosen is essentially resistant to acid attack and temperatures up to approx. 120^oC . A proportion of the recirculating PLS liquor is regularly bled from the circuit to enable recovery of metal components. This can be typically 25-35% of the volume circulating liquor and additional mineral feed material is added at this stage to compensate for the removal. The proportion removed is determined by the saturation of key components such as iron, sulphur and target metals. Additional aqueous nitrate may also be added to the pipe reactor at this stage to compensate for any losses from the total system. A preferred embodiment is to add aqueous calcium nitrate since this immediately reacts with circulating sulphuric acid to form NO species via the intermediate nitrous or nitric acid coupled with removal of excess sulphate as either gypsum, bassanite, anhydrite or jarosite. An example of an analysis of this PLS liquor is shown in Table 12. Referring again to FIG.1, a flotation circuit 18 facilitates recovery of evolved sulfur and other components from the primary leach as a “flotation concentrate”. The secondary flotation circuit 18 recovers sulfur formed as an oxidation product of sulfide ores such as either arsenopyrite or pyrite. With particular reference to the target metals gold and silver, the strong sulfur-gold bond acts to entrain nano and micron scale gold and silver released from solid solution within the ore. In addition the sulfur forms an efficient collector of gold and silver and can also attract and bind particulate gold. There are frequently good commercial reasons for recovering as much gold or silver as possible in this stage, particularly in VMS type deposits where lead, zinc, copper or similar concentrates are produced, typically by froth flotation. In particular, many smelters pay for contained precious metals after deducting a relatively minor processing fee, which can be as small as 2-3% of the total value of contained metals on the metal exchange. Depending on the contained metals, and concentrations, this can be a cost-effective means to recovering precious metals without having to independently develop other process options to recover them. Flotation may produce some excess liquid attached to the foam and there will be additional particulate minerals which are attached to the foam bubbles and to the sulfur. To maximise the recovery of precious metals, it is preferable to dewater said concentrates by thickening in thickener 19. This may, or may not, require added polymers or coagulants and is generally determined on a case-by-case basis. Optionally, fine grinder 20 may be used in this process. In this regard, particulate gold or silver may attach to the sulfur flotation media from above and some additional ore particles may also attach and so it may be preferable to grind floated material further to ensure maximum adhesion to the sulfur of any gold, nickel, silver and other potentially recoverable. In this process, sulfur acts similarly to carbon as used in carbon-in-pulp gold extraction with the strong gold-sulfur bond acting to bind gold nanoparticles or microparticles as they are released by further grinding of the ore. Ultimately, the decision to provide finer grinding may be determined on a case-by-case basis. The loaded sulfur following flotation (and optional fine grinding) is now suitable for potential blending with other metal concentrate materials in vessel 24 prior to despatch to the smelter 25. The smelter will have upper and lower limits established for a range of components and penalties for any that are considered undesirable. By way of example, typically the smelter will set a penalty for arsenic above 1% by weight. In the case of lead concentrates, providing the lead content exceeds 36-40% and the zinc is less than 8%, the smelter in many cases will pay up to 97% of the value for contained gold and silver. In the case of flotation concentrates comprising gold, silver and sulfur, this provides an effective mechanism to reduce the processing and recovery costs for the contained gold simply by blending with, in the example above, lead or lead zinc concentrate. The blending of concentrates to maximise the return to the processor from the smelter is known to those skilled in the art. In an alternative or additional embodiment, shown in FIG.1, the flotation concentrate is not sent to the smelter 25 but to sulfur roaster 26 to recover the metals after burning off the sulfur, or to cyanidation circuit 23, following thickening in tails thickener 21, to recovery via a cyanide leach circuit 23. This process of precious metals recovery will be well known to those skilled in the art. Optionally, the tailings from the secondary flotation circuit 18 may be subjected to thickening in tailings thickener 21 and then to dissolved metals recovery circuit (i.e. jarosite circuit) 30 as shown in FIG.2. In this regard, the tailings from the secondary flotation circuit 18 comprises the balance of gold, silver and typically any other metals that have either (1) not dissolved / transferred into solution or (2) dissolved but reprecipitated. A non-limiting example of the former is particulate gold that is insoluble in dilute nitric acid. A non-limiting example of the latter is lead from galena that dissolves rapidly as lead nitrate but then reacts with sulfate ions as they are produced by the oxidation of pyrite or arsenopyrite. Although it is advantageous to recover lead as plumbo-jarosite, any lead that reacts with sulfate directly to form dense lead sulfate will generally be found in the flotation tailings. Similarly where silver first dissolves in the dilute acid leach, the presence of chloride or other halogens or sulfate ions react with silver to form insoluble silver sulfate, chloride (or equally insoluble bromides and other halogens). These will generally be resident in the flotation tailings, but amenable to recovery in cyanide leach circuit 23. The decision whether or not to conduct the process shown in FIG.2 will depend on the economics of the individual process. Referring now to FIG.2, the transfer of dilute acid liquor “aqueous leach liquor” comprising the bulk of the dissolved pyrite and arsenopyrite minerals from pipe reactor circuit 10 provides a mechanism to recover individual metals selectively depending on their final concentration within the liquor in jarosite circuit 30. The design of the pipe reactor circuit 10 shown in FIG.1 is preferably such that the leach liquor on exit from the pipe reactor circuit 10 is close to the saturation point for at least some of the base metals that are now intended to be recovered. Suitably, the design of the process enables recovery of lead, zinc, silver and nickel as double salts with jarosite (plumbo-jarosite, argento-jarosite etc). The preferred way to achieve this is cool the leach liquor whereby the jarosite minerals will begin to form during the cooling process in jarosite crystallizer 31. As the leach liquor is cooled the latent heat associated with the crystallisation of the jarosite minerals will be released, enabling heat recovery to be transferred for other uses. As ammonia is used elsewhere in the process to recover magnetite, the ammonia can be effectively recycled by transferring aqueous ammonia solutions to a heat exchanger whereby the surplus heat can be used to recover ammonia for re-use. As previously described, during the cooling process, the jarosite minerals will begin to crystallise and settle in jarosite crystallizer 31. Depending on the initial concentrations of lead, zinc, copper, silver, nickel and other target metals 32, in one embodiment recovery of the those approaching saturation may be achieved by ‘seeding’ these other target metals in seeding tank 33. In the seeding process a crystalline form of the metal salt is added to the jarosite solution during the cooling process, to accelerate the nucleation of specific metal jarosite solutions e.g. argento jarosite. As each metal is sequentially removed, recovery of the next metal can be attempted as determined by process economics and the value of the respective metal being recovered. An alternative process to seeding is to use a sacrificial anode (not shown) to add metal ions to the jarosite circuit and thereby encourage crystal formation. The commercial economics of this can be assessed on an individual metal basis. For example, the cost of this approach is often cheaper than adding metal salt when freight and other costs are taken into consideration. For example the cost of copper supplied as a sacrificial anode can be less expensive on a unit copper basis than the equivalent quantity of copper as the hydrated metal sulfate. This applies particularly to metallic aluminium, copper and zinc salts. Following cooling and jarosite recovery, the next process step requires alkali addition from alkali source 34 to commence nucleation of the long chain polymeric ferric hydroxide minerals of general formula Fe-O-OH in iron recovery circuit 35. This commences typically at pH 2.6-2.9 with nuclei of the ferric hydroxide minerals forming. It is important not to generate too much ferric hydroxide as this is undesirable from a filtering and recovery basis. Hence choice of alkali is to encourage nucleation with the minimal quantity of alkali additive, hereby limiting the cost of this stage. Typically, sodium hydroxide or calcium hydroxide can be used preferably to increase the pH to above 2.6 and monitor the removal of undesirable arsenic from solution as the stable ferric arsenate mineral scorodite. Then as arsenic is progressively eliminated from solution, attention can be diverted to focus on recovery of saleable/marketable iron minerals, particularly hematite and magnetite. Magnetite formation: FeSO₄ +2FeO.OH + NH₄OH +NH₄SO₄ +Fe₃O₄ +H₂O Ammonia recovery: NH₄SO₄ + CaO ^ NH³ (gas) + CaSO₄.2H₂O (to tailings) A commercial decision will generally be required to be made in the negotiating the arsenic content of any Direct Shippable Iron (DSI) ore minerals contracted for shipment to an industry partner. Once the residual arsenic in solution falls below this limit, further scorodite formation and removal is not required. Final precipitation of ferrihydrite-goethite occurs by raising the pH above 2.9 in reactor vessel 36. Above the pH range 2.6-2.9, polymeric ferric hydroxide rapidly forms and coagulates. This is an intermediate product only, since it is too difficult to filter or remove from the process as a by-product. Their formation is only to enable final changes to the solution chemistry to recover directly shippable ore from the process. This is achieved in reactor vessel 38 following goethite/ferrihydrite conversion to maghemite or magnetite. This step requires simultaneously acting alkali ammonium hydroxide 37 to the nucleating ferric hydroxide sludge and maintaining the pH of the solution in the range 2.9-3.2 whilst adding spent liquor from the jarosite crystalliser circuit. With careful control of alkali addition, the red-brown ferric hydroxide sludge will immediately turn black and start to settle much more quickly as a blend of maghemite and magnetite. Optimal addition of aqueous ammonia together with spent liquor from the jarosite circuit comes together to form a thick blanket of magnetite which is of very fine particle sizes and can divert back to red-brown ferrihydrite/goethite in overly oxidative liquor. The reaction end point is monitored by control of the NH4⁺ ion concentration so as to regularly enable diversion of ammonium sulfate back to the ammonia recovery heat exchanger circuit. Excess ammonium is undesirable as it encourages formation of ammonium jarosite, an undesirable by-product. The reaction proceeds according to the following equation; FeSO₄ + 2 Fe-O-OH +NH₄OH + NH₄SO₄ + Fe₃O₄ +H₂0 As the magnetite formed during the earlier magnetite preparation process is nano or micro-particulate, in oxidative liquors it has a tendency to oxidise rapidly back to hematite or maghemite which is less desirable as an end-product. For this reason the magnetic separation of the magnetite in magnetic separator 39 is best practiced under a nitrogen or other inert atmosphere whereby the recovered magnetite can be blended with bentonite and pelleted in pelletizer 40 before any significant conversion back to hematite occurs. Choice of bentonite addition and moisture content of the magnetite is chosen to expedite pelleting into those sizes of pellets most amenable to export to the sea- going Iron Ore trade. These aspects of the trade in iron ore minerals are known to those versed in the trade. Other forms of pelletisation can be used equally as effectively including extrusion through dies as well as roller pelletising where the two surfaces of an impinging roller have indentations to enable pellet formation on the roller. Choice of pelleting process is beyond the scope of this invention and known to those expert in the field. Once pelletized, magnetite is sent to magnetite stockpile 41. Ammonium sulfate liquor from the magnetite separation phase is sent to the ammonia recovery circuit 38. Addition of calcium oxide (“quicklime”) results in a rapid increase of temperature and evolution of gaseous ammonia. This is evacuated to enable redissolving in water to reform the ammonium hydroxide / aqueous ammonia alkali required for magnetite formation. The calcium oxide rapidly converts to calcium sulfate which is discharged to tailings. Ideally separate pondage or several days residence time enables calcium levels to fall below levels which would impact the overall process. Recovered process water, preferably largely devoid of calcium, is sent back for acid makeup. So that the invention may be readily understood and put into practical effect, reference is made to the following non-limiting examples. EXAMPLES Nitric acid leaching as an alternative to sulphuric and hydrochloric acid leaching of both oxide and sulphide feedstocks has been studied since 1909 and commercial plants employing nitric acid operated successfully for many years in treating concentrates by both Electrolytic Zinc Corporation, Kennecott Copper and others. The clear advantage of a nitric acid system over other acids is its ability to achieve target extractions in a shorter time at less aggressive conditions of temperature and pressure, resulting in a reduction in plant capital cost. However, nitric acid is more costly than other acids, and therefore recovery, regeneration and recycle of nitric acid to the leach is key to achieving an economic flowsheet. The chemistry of nitric acid leaching results in the high-rate evolution of NO and/or NO2 (collectively Nox). Various formulas for the oxidation of pyrite by nitric acid have been proposed. These are shown in equations 1-4 below: 2FeS₂ +10HNO₃ → Fe₂(SO₄)₃ + H₂SO₄ + 4 H₂O + 10NO (Eq1 to4) 2FeS₂ + 8HNO3 → Fe₂(SO₄)3 + S⁰ + 8NO + 4H₂O FeS₂ + 8HNO3 → Fe(NO₃)₃ + 2H₂SO₄ + 2H₂O + 5NO FeS₂ + 18HNO₃ → Fe(NO₃)₃ + 7H₂O + 2H₂SO₄ + 15NO₂ In operating plants – NO will be oxidised back to HNO3 in the presence of oxygen and water as follows: 2NO + O2 → 2NO₂ (Eq5) Nitrogen dioxide can then be absorbed in water to regenerate nitric acid and NO: 3NO₂ +H₂O → 2HNO₃ + NO (Eq6) A simplification of the overall reaction, from those presented above is: 2FeS₂ + 8HNO3 → Fe₂(SO₄)3 + S⁰ + 8NO + 4H₂O (Eq7) However unless 100% oxygen is introduced in equation 5, the use of air, containing 21% oxygen but 78% inert nitrogen, results in large volumes of gas needing to be scrubbed. Scrubbing typically uses a water mist; however, the low solubility of nitrous oxide in water requires multiple scrubbing stages. This adds significantly to CAPEX of plants particularly when endeavouring to process low grade materials such as tailings, rather than concentrates in the Kennecott and Electrolytic Zinc examples above. For this reason the use of nitric acid has not been widely adopted, for the treatment of refractory gold in tailings, unless there are very large reserves of tailings material to be re-processed. To overcome this deficiency of prior processes, the present invention incorporates a pipe reactor to overcome the requirement for wet gas scrubbing of large volumes of essentially inert air. This is because Nox reconversion to HNO3 occurs within the head space of the pipe-reactor described herein. In a preferred embodiment, this includes the direct injection of oxygen or the addition of an oxidant such as a peroxygen component within the feed. Where additional oxygen is provided to the reactor, formation of elemental sulfur (S^o) in equation 2 above, is minimised. An unknown fraction of metals may also be solubilised as nitrates rather than sulphates, further reducing Nox gas evolution. The formation of metal nitrates and the addition of metal nitrates, such as calcium nitrate, provides a means to maintain the process once it has been instigated. As described in the equations above, the process can mostly recover the nitric acid, and when recycled efficiently, can be considered to function simply as a catalyst, with losses typically as low as 2%, (i.e., essentially unconsumed in the reaction). But at sufficiently high solution ORP, FeSO₄ will be further oxidised to Fe₂(SO₄)₃ and much of the elemental sulphur will be oxidised to sulphuric acid as has been confirmed by recent test work, thus minimising the formation of elemental sulphur, if this is desired. In the traditional recovery of gold from ores using cyanide leaching and carbon in pulp (CIP) or carbon in leach (CIL) processes, the gold is recovered onto activated carbon or activated charcoal through adsorption. The gold cyanide complex (usually in the form of Au(CN)2-) can be selectively adsorbed onto the carbon surface because of these interactions. Later, the gold-loaded carbon is typically treated with a solution such as sodium hydroxide or sodium cyanide to desorb the gold from the carbon surface, allowing for gold recovery. In these processes, the interaction between gold and the carbon surface is primarily a result of weak van der Waals forces, particularly London dispersion forces, which are relatively weak compared to covalent or ionic bonds. By contrast the bond between gold nanoparticles and colloidal sulfur is significantly stronger, which has resulted in sulfur in pregnant liquor solutions (PLS) obtaining a reputation as a ‘preg robber’ – in the sense that sulfur can bond to gold more strongly than to carbon. However the inventor believes that there can be benefits in the production of sulfur acts as a nonmetal during the leach process. If gold is being liberated from ‘solid-solution’ (i.e. essentially nano-particulate gold) then this results in the formation of the gold-sulfur bond, which has potentially stronger attraction than the weak van der Waals forces described above. Gold monosulfide, AuS, for example, shows symmetries of the ground and low-lying electronic excited states by application of laser excitation techniques . The bond length and dissociation energy of the ground X2Πi state are determined to be 2.156(2) Å. And 298 ± 2 kJ/mol, respectively. The covalent interaction at the gold–sulfur interface requires formation of gold–thiolate bond(s): that is, the sulfhydryl group is deprotonated, creating formally a thiyl radical (RS·), whereas the protonated SH group can interact with gold only by weaker coordination-type bonds through the sulfur lone-pair electrons. The thiolate– gold (RS–Au) bond has a strength close to that of the gold–gold bond, so it can significantly modify the gold–gold bonding at the gold–sulfur interface. The importance of this detail has become clear only recently when new experimental and theoretical evidence has become available about interfacial structures in thiolate-SAMs on Au(111), atomic structures of thiolate-protected gold nanoclusters and gold–thiolate–gold molecular junctions. This is relevant to the separation of gold from refractory ores according to the present invention. The practical application of this discovery is shown on the following laboratory data as it refers to the FT1 Con (i.e. a sulphur float). In the example, 48.8% of the gold and 46% of the silver in the primary ore is contained in a mass pull during the flotation using pine oil as the flotation agent, of only 15.7% of the initial mass. If then blended with concentrate, the gold and silver values can be recovered via conventional smelting at relatively low cost. Table 1 shows a typical mineral concentration in flotation feed and flotation tails. The data in Tables 2-11 and FIGS 3-11 show various quantitative analyses at various stages of the leach process. In this regard, Table 2 shows an example of initial mineral source feed before commencing leaching. Table 3 shows the leach solution composition after leach completion (10% w/w./, 1 hr residence time). Table 4 shows the composition of a flotation concentrate. The gold and silver distribution is particularly noted. Table 5 shows the composition of tailings following flotation. Table 6 shows the residue remaining after digestion. Table 7 shows the leach solution after cyanide leach (20% w/w, pH 10.5, 500 ppm NaCN, 2 hr residence time). Table 8 shows the cyanide leach residue. Table 9 shows head characterisation of Stage 1 to 3 ZST Sample while Table 10 shows QXRD Results (HGM tailings and ZST) Results from Assay by Size analysis are presented in Table 11. Table 12 shows an example of analysis of this PLS liquor, as previously described. Tables 13 and 14 show XRD analysis of a sample which indicates the presence of jarosite and the jarosite-related mineral beudantite. Tables 15 and 16 show XRD analysis of another sample comprising jarosite. Referring now to FIGS 3-10, these show various leach run oxidation reduction potential (ORP), and temperature versus residence time and target metal extraction percentages for each leach run. In particular, from FIGS.5 and 6 it is clear that gold and silver remain largely undissolved over time and remain in residue for recovery by cyanidation or similar process. Laser Size Analyses cumulative size distribution is shown in Figure 11 and indicates that the ZST feed sample tested was approximately P80 of 35 microns. Throughout the specification the aim has been to describe the preferred embodiments of the invention without limiting the invention to any one embodiment or specific collection of features. It will therefore be appreciated by those of skill in the art that, in light of the instant disclosure, various modifications and changes can be made in the particular embodiments exemplified without departing from the scope of the present invention. All computer programs, algorithms, patent and scientific literature referred to herein is incorporated herein by reference.

TABLES Table 1: Typical mineral concentration in flotation feed (left) and flotation tails (right)

Table 2: Feed composition (typical) with commencing distribution at 100% Leach 10 (LCH10) example L^{CH10 Feed} 100 Grade Dist% Au, g/t 1.78 100.0 .0 .0 .0 .0 .0 .0 .0 Table 3 :LCH10 Le

Solution Composition after leach completion Grade Dist% .0 5 .8 .7 .8 .8 .6 .9

Table 4: Flotation concentrate (FT1 Con) note gold & silver distribution 15.7 Grade Dist% Au, g/t 5.14 48.8 .1 .8 .9 .2 .2 .4 .7

Table 5: Tailings following flotation F^T1Tail 31.8 Grade Dist% Au /t 255 491 7 4 4 0 0 0 4

Table 6: LCH10 r^esidue 47.5 Grade Dist% Au, g/t 3.22 97.9 8 2 3 2 2 4 1

Table 7: LCH11 C^{N leach} Grade Dist% Au mg/ml 021 261 1 0 0 0

Table 8: LCH11 r^esidue 31.8 Grade Dist% 0 6 4 4 0 3

Table 9: Head Characterisation of Stage 1 to 3 ZST Sample Parameter Unit Sample 1 Sample 2 (16/12/2020) (04/03/2021) Gold g/t 2.16 1.78 48 91 85 20 00 72 10 50 58 00

Table 10: QXRD Results (HGM tailings and ZST) M^{ineral Phase, wt%} _{(HGM F} ^H l_o ^G t^M a_ti ^T o^a n^il Pⁱⁿ l_a ^g n^s Zinc Scav Tail t _Feed) (HGM Flotation Plant Tails) P rite 531 590

Table 11: Assay by Size Analysis Results Phase 1 sample Sample: ZST Sample P80: <38 Size Fraction Assay M M P i A A C C F M Pb S S Si Zn % 30 67 84 03 19 09 87 74 64 71 73 n % 56 43 44 86 65 1.5 07 28 3.2 0.0

Table 12: Relationship of metals to pH in PLS

Table 13: XRD phase quantification

Table 14: XRD elemental composition results

Table 15: XRD phase quantification

Table 16: XRD elemental composition results

Claims

CLAIMS 1. A method of recovering a target metal from a mineral source, said method including the step of treating the mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to thereby recover the target metal from the mineral source and produce an aqueous leach solution comprising jarosite, or a jarosite- related mineral, that comprises an extraneous other target metal added to the aqueous leach solution and the said other target metal originally present in the mineral source. 2. The method of Claim 1, wherein the nitrogen oxide is of general formula NxOy, wherein X is 1 or 2 and Y is 1,

2, 3 or 4.

3. The method of Claim 1 or Claim 2, wherein the nitrogen oxide includes NO, NO₂, NO₃, N₂O₄, HNO₃ NaNO₃, KNO₃, NaNO₂, NH₄NO₃, Ca(NO₃)₂, Mg(NO₃)₂ and/or Fe(NO₃)₃.

4. The method of any preceding claim, wherein, the treatment is initiated by addition of nitric acid (HNO3) to the mineral source.

5. The method of Claim 4, wherein further treatment occurs in the presence of one or more other nitrogen oxides of general formula NxOy, wherein X is 1 or 2 and Y is 1, 2, 3 or 4.

6. The method of any preceding claim, which is a continuous flow method.

7. The method of any preceding claim whereby initiation of the treatment facilitates generation of one or more gaseous phase nitrogen oxides.

8. The method of Claim 7, wherein the gaseous phase nitrogen oxides include one or more of NO, NO₂ and/or N₂O₄.

9. The method of any one of Claims 4-8, wherein subsequent to initiation with nitric acid one or more other nitrogen oxides at higher valence states (+3, +4) are added.

10. The method of Claim 9, wherein the one or more other nitrogen oxides include NaNO₃, KNO₃, NaNO₂, NH₄NO₃, Ca(NO₃)₂, Mg(NO₃)₂ and/or Fe(NO₃)₃.

11. The method of any preceding claim, wherein the nitrogen oxide is substantially not consumed during treatment.

12. The method of any preceding claim, wherein the mineral source is, or comprises, a metal sulfide ore.

13. The method of Claim 12, wherein the metal sulfide or is a volcanogenic metal sulfide (VMS).

14. The method of any preceding claim, wherein the target metal is gold.

15. The method of Claim 14, wherein the gold is recovered as a gold sulfur compound, such as a gold disulfide.

16. The method of Claim 14 or Claim 15, wherein the gold is recovered by adsorption or chemically combining with colloidal sulphur.

17. The method of any preceding claim, wherein: (i) the extraneous target metal is added as a water-soluble salt; and/or (ii) the extraneous target metal is added electrochemically.

18. The method of Claim 17, wherein (ii) is by the use of a sacrificial anode comprising the extraneous target metal.

19. The method of any preceding claim, wherein the one or more other target metals include silver and/or lead.

20. The method of any preceding claim, which further includes treating the jarosite or jarosite-related mineral to thereby produce an iron oxide.

21. The method of Claim 20, wherein the iron oxide includes ferrous and/or ferric oxides.

22. The method of Claim 21, wherein the iron oxides include magnetite (Fe₃O₄) and/or maghemite ( γ Fe₂O₃).

23. The method of any preceding claim, wherein treatment with the nitrogen oxide is performed in a reactor vessel of a generally tubular configuration.

24. The method of Claim 23, wherein the reactor vessel is a pipe reactor.

25. A target metal recovery system or apparatus comprising a reactor vessel of a generally tubular configuration for treating a mineral source under acidic conditions in the presence of a nitrogen oxide at a temperature no greater than about 100⁰C and at about atmospheric pressure to facilitate recovery of the target metal from the mineral source.

26. The target metal recovery system or apparatus of Claim 25, wherein the reactor vessel is a pipe reactor.

27. The target metal recovery system or apparatus of Claim 25 or Claim 26, wherein the reactor vessel operates in a closed system with loading and unloading provisions, such that there is substantially minimal or no gas escape without deliberate venting .

28. The target metal recovery system or apparatus of any one of Claims 25-27, which further comprises a seeding vessel for producing jarosite or jarosite-related mineral comprising one or more other target metals.

29. The target metal recovery system or apparatus of any one of Claims 25-28, further comprising one or more slurry tanks, grind mills, flotation tanks, magnetic separators and/or pelletizers.

30. The target metal recovery system or apparatus of any one of Claims 25-29, for use according to the method of any one of Claims 1-24.