Dalradian’s Northern Ireland property has potential to host two distinct deposit types. Licences DG1, DG3,DG4, DG5 and DG 6 which includes the Curraghinalt gold deposits, has potential to host orogenic gold deposits. Licence DG2, underlain by the Tyrone Igneous Complex, has potential to host volcanic massive sulphide mineralization (VMS), as well as porphyry copper-gold, and irongold exhalites (Hollis et al., 2014; Hollis et al., 2015; British Geological Survey, 2016).

Orogenic Gold Deposits
Rice et al. (2016) noted that the timing of gold mineralization at Curraghinalt (ca. 462.7 – 452.8 Ma) closely followed peak metamorphism associated with the Grampian event of the Caledonian Orogeny. It is temporally linked with an extensional setting following orogenic uplift and collapse. Rice et al. (2016) concluded that Curraghinalt is more likely an orogenic (rather than intrusion related) gold deposit. Thus an orogenic gold deposit model best describes the Curraghinalt vein system.

Volcanic Massive Sulphide Deposit Model
The volcanic stratigraphy of the Tyrone underlying licence DG2 is a potential host to volcanic massive sulphide deposits. VMS deposits are syngenetic, stratabound, and in part stratiform accumulations of massive to semi-massive sulphide that form seafloor hydrothermal systems at or near the seafloor (Gibson et al., 2007; Galley et al., 2007). The deposits consist of two parts: a concordant massive sulphide lens (>60 percent sulphide minerals), and discordant vein-type sulphide mineralization, commonly called the stringer or stockwork zone, located within an envelope of altered footwall volcanic and or sedimentary rocks (Gibson et al., 2007).

The Curraghinalt Gold Deposit
High-grade gold mineralization occurs as a series of west-northwest trending, moderately to steeply dipping, subparallel stacked veins and arrays of narrow extension veinlets. These veins are hosted by the Neoproterozoic Dalradian rocks in the central section of the Sperrin Mountains, and represents the largest known gold deposit in the United Kingdom.

The mineral resource model discussed herein focusses on a set of 21 prominent gold-bearing quartz veins that occur mainly within psammites, semi-pelites, and pelites of the Dalradian Argyll Group, within the Mullaghcarn Formation. Auriferous quartz veins exist between the main modelled veins, but their continuity is difficult to demonstrate at the current drill spacing. The quartz vein system was investigated by core drilling and is partly exposed in underground workings. Surface exposures of the vein system are limited to the Curraghinalt and Attagh Burns (creeks), as well as a variety of surface trenching excavations completed in 2003 and in the late 1980s. The veins range from a few centimetres to five metres wide.

The vein swarm has been modelled along strike for approximately 2,300 metres, across strike for approximately 800 metres and down dip for over 1,200 metres by prospecting, trenching, and drilling. Though the modelled veins extend 2,300 metres along strike, the vein system is traceable with similar strike aligned veins occurring over approximately four kilometres from Alwories in the east to Scotchtown in the west. The vein system remains open along strike and at depth. On average, the quartz veins dip between 50 degrees and 75 degrees to the north northeast.

The gold-bearing vein system in the Curraghinalt deposit is crosscut by several generations of structures with different characteristics and orientations. The two most prominent structures in the deposit are the Kiln and Crow shear zones. These shear zones are east-west trending, either south-dipping or north-dipping. The Kiln shear shows evidence of brittle reactivation as indicated by the presence of gouge zones along the contact between the highly strained ductile rocks within the shear zone and the Dalradian metasedimentary wall rocks. The Kiln shear disrupts and displaces the vein zones (D veins) with observed oblique dextral-normal kinematics.

Vein zones are entrained within the Kiln Shear and previous workers (Boland, 1997) have suggested that the shears have controlled vein emplacement or at least served to produce wider mineralized segments.

In addition, other shear zones and faults with gouge development crosscut the Curraghinalt deposit. These structures trend northeast and dip moderately northwest. Fault thicknesses and characteristics vary based on the country rock lithology. Faults and shear zones within the pelite, north of the No.1 vein, tend to be wider (typically 0.5 to1.0 metre wide), either crosscut or disrupt the pre-existing regional foliation in the pelite, and in some cases, have measurable offsets (e.g., four metre offset of the T17 vein).

Limited information is available for the kinematics of the crosscutting fault systems. However, in 3D the Kiln shear zone is currently modelled to offset the auriferous vein system with a dextral strikeseparation. Underground mapping by Dalradian shows sinistral strike-separation of the vein system along northeast-trending faults, particularly where faults intersect the western section of veins (e.g., T17 west, eight metre offset by FZ_107). However, the Dalradian mapping also shows the dextral strike-separation of veins where faults intersect the eastern section of veins, for example the V75 vein east is offset by three metres. This fault contains steep- plunging slickenlines, suggesting the movement direction was dominantly vertical. This is possibly related to dominantly normal movement creating sinistral or dextral separations given the dip of the faults.

Petrographic work by Clarke (2004) has documented that the gold mineralization at Curraghinalt occurs in quartz-pyrite-carbonate veins and is associated with variable abundances of carbonate, chalcopyrite, and tennantite-tetrahedrite. Gold is commonly in the form of native gold and more rarely as electrum (>20 weight percent silver), and occurs primarily along fractures in pyrite, as inclusions in pyrite, and at pyrite grain contacts with carbonate and quartz. Most native gold grains are associated paragenetically with carbonate, chalcopyrite, tennantite-tetrahedrite, and telluride minerals infilling fractures in pyrite. The seven veins studied at the time have similar mineralogy. Native gold was observed in samples from all veins and grains range in size from 2 micrometres to 150 micrometres.

The FS mine plan is based on a ramp access underground mining operation producing and average of 1,400 t/d of ore from a blend of mining methods, including lateral development, cut & fill and longhole (i.e. longhole stoping, upper retreat and pillar recovery).

Initial mine development at Curraghinalt will take place from the existing exploration adit and from a newly ramp access which will be collared near the proposed process plant site. Mining methods selected for the Curraghinalt project were chosen to maintain mining flexibility and selectivity for the various anticipated ground conditions. The majority of Mineral Reserves will be mined by longhole open stoping (66% combined longhole, uppers and pillar recovery), plus 17% cut & fill and 17% from development in ore. The mine design will use 18 m sub-levels (floor to floor) and strike lengths averaging 15 m depending on geotechnical parameters. The average diluted stope width will be 1.6 m; however, widths can vary between 1.2 to 3.5 m. The minimum mining width is 1.0 m. Paste backfill, produced in a backfill plant located underground, will be made from a mixture of tailings and cement and will be the primary backfill material, with unconsolidated waste rock from development headings used where possible. An internal ore pass system will direct ore and waste to the haulage levels where diesel powered haul trucks will transport the material to surface via the main ramp that will daylight adjacent to the process plant.

Three underground mining methods were selected for Dalradian; sub-level longhole open stoping (LH), cut & fill (C&F), and longhole upper stoping (LHUP). The bulk of the mining will be done with a longhole drill in a combination of LH and LHUP. C&F stopes will be prevalent in areas where ground is weaker and the vein is less continuous. Uppers will be taken after three lifts of C&F are mined and the vertical mining height of the upper is 7.25 m. Pillars will be mined using the LH method in all geotechnical zones, but will only be mined on retreat near the end of the mine life.

Longhole Mining
LH will be used where rock quality and ground conditions allow and where vein thickness is relatively uniform. It is the highest productivity method selected for the production phase, but given many veins are relatively narrow, considerable preparation is required to maintain stope inventory.

LH is the least selective of the mining methods when applied over long vertical distances due to drill hole deviation and vein geometry variations. To mitigate these effects, longhole drilling distances at Dalradian will be limited to 14.5 m, which is well within equipment capabilities and will give little drill hole deviation. Sub-level spacing will be kept consistent at 18 m (floor to floor) throughout the mine. However, depending on the geotechnical zone, the strike length of the LH stopes varies. In the Green geotechnical zone, a strike of 20 m is used, while 15 m and 8 m are used in the Yellow and Pink zones, respectively. A minimum mining width of 1.0 m was used for planning purposes. If the mineralized vein was less than the minimum width, internal planned dilution was included to increase the width to 1.0 m.

The longhole mining cycle will begin with blasting the slot raise to provide a free face for the first longhole round and initial empty volume for blasted swell. Production blasting will begin at the stope ends and retreat to the cross-cut. Figure 16.10 shows a typical mining sequence for LH at Dalradian.

Resue Mining
On vein ore development and C&F mining at Dalradian is primarily done using the resue or split- shot mining method. Ore and waste are blasted and mucked in two different steps. Depending on the vein width, the ore or the waste might be blasted first. If more waste than ore is present in the current production face, then waste is blasted first. If there is more ore than waste, then the ore is taken first. For design purposes, a minimum mineralized vein width of 0.7 m was used.

The entire round is drilled using a single-boom jumbo. Either the ore or waste is blasted, and then mucked using a 4 tonne class Load-Haul-Dump unit (LHD). On the next shift, the remaining holes are loaded and the rest of the round is taken and mucked with the 4 t LHD. After both blasts, the area is scaled and ground support installed. Resue mining, while it has slightly higher costs and lower productivities because of two blasting and mucking cycles, the method reduces the dilution tonnage to the mill and allows for a larger longhole drill to be used. There are enough production faces open at any given time to maintain the average 1,400 t/d mine production feed.

Cut-and-Fill
The C&F mining cycle begins with developing the first cut from the cross-cut. Once the first cut is mined to completion, it is backfilled using paste, waste, or a combination of the two. The backfill should be placed as close to the back as possible. Once backfilling is nearly complete, the back is taken down and used to form an internal 15% ramp to the next C&F lift. The process repeats until the third and final cut is complete. The last C&F lift will remain open until the LHUPs are mined. At that point the LHUP and final C&F lift are backfilled from the level above. Avoca mining and backfilling can occur if there is access to both ends of the production panel. Whenever possible, internal ramping on vein is utilized to minimize waste generation and reduce costs. The initial C&F lift is 3.5 m high with a width of 2.3 m at the back and 4.6 m at the sill with one rib being vertical and the opposite side dipping parallel to the vein. The second and third lifts are 3.625 m high and have a width of 2.0 m at the back and 4.5 m at the sill. The deposit average for C&F shanty back is 56°.

Stopes are developed on the lowest level first, and each subsequent stope or 3.625 m lift is developed above the depleted and backfilled stope. Mining direction is bottom-up. Cut-and-fill mining will be utilized in thinner areas where LH stopes are not economic. C&F will also be used in areas of poor ground conditions where larger stopes are not geotechnically possible.

Longhole Uppers Stoping
Longhole uppers stoping accounts for 26% of planned mill feed and are an important option as it offers additional production flexibility and minimizes access development of individual 3.6 m high lifts. Longhole upper stopes are designed and planned in areas where the ground is not geotechnically suitable for LH. Uppers are taken in areas where C&F mining occurs and are typically mined from the third and final lift of the C&F to the next mining level. The vertical height of the upper stope from back to sill is 7.25 m, which is half the height of LH.

Uppers stope development begins with sill cuts driven on vein at the top and bottom of the uppers stope block. Before uppers are drilled and blasted, mine geologists will inspect, channel sample, and mark the boundaries of economic mineralization. Strike lengths of 15 m or 7.5 m are used and varied based on geotechnical parameters. Blasted ore is mucked from the lower extraction drift. If access is available from both ends of the upper level, the mining sequence can be nearly continuous with Avoca backfilling. If there is not access at either end of the stope, then the uppers void will be backfilled with structural paste. Once the paste has cured, the next upper stope will be mined.

Crushing
Material from underground mining operations will feed an apron feeder - primary jaw crusher system, which produces a product size P 80 of 110 mm.

Feed material will be hauled by 30 t haul trucks to the run of mine stockpile or from underground. Material will be stockpiled near the jaw crusher or direct dumped through a static grizzly into a dump pocket. Stockpiled material will be re-handled. Oversize material from the static grizzly will be removed reduced using a rock breaker.

An apron feeder will draw material from the dump pocket at a nominal rate of 189 t/h and discharge directly into the primary jaw crusher (1,200 mm x 870 mm, 160 kW). The crusher discharge will feed the stockpile feed conveyor. The crusher will operate 24 hours per day but the production can be done in 12 hours to reduce the noise impact if needed.

The crushed ore storage facility will consist of a dome-covered stockpile with two in-line belt feeders located within a corrugated pipe reclaim tunnel. The belt feeders will transfer ore to the conveyor feeding the SAG mill. The fine ore stockpile will have a 1,500 t live capacity that can support process plant operations for 24 hours when the crushing plant is not operating. The stockpile will be approximately 35 m in diameter and 14.7 m high which will correspond to approximately three days storage. Each belt feeder will be capable of providing a total throughput of 68 t/h to the plant when required.

Grinding
The grinding circuit will consist of a SAG mill operating in closed circuit with a hydrocyclone cluster. Material from the crushed ore stockpile will be fed to the SAG mill via the SAG mill feed conveyor. The grinding circuit will operate at a nominal throughput of 68 t/h (fresh feed), and produce a final particle size P 80 of 240 µm. The SAG mill will be 6.1 m in diameter by 2.44 m effective grinding length driven by a 1.3 MW motor.

Water will be added to the SAG mill to maintain the ore charge in the mill at a constant slurry density of 70%. Slurry will overflow from the mill to a trommel screen, attached to the mill discharge end. The mill trommel screen oversize will overflow into a trash bin for removal from the system.

The SAG mill hydrocyclone cluster will classify the feed slurry into coarse and fine fractions. The coarse underflow will flow back to the SAG mill feed end for additional grinding. The overflow with a nominal P 80 of 240 µm will flow by gravity to the flotation conditioning tank.

Regrind
The rougher concentrate will be pumped to a vibrating trash screen for removal of trash material and then feed the concentrate thickener. Flocculant solution (anionic polyacrylamide) will be added to the thickener feed to promote the settling of the solids. The concentrate thickener will have a diameter of 9 m and produce a thickened product of 50% solids in the underflow. The thickener overflow will flow by gravity to the process water tank. The underflow slurry from the concentrate thickener will be pumped to a regrind mill, 244 kW. Ceramic media will be used to reduce the concentrate to a final produce size of P 80 of 50 µm. The slurry at a density of 50% solids will flow from the regrind to the CIL circuit.

The process selected for the Curraghinalt project and consists of a flotation cyanide leach and carbon adsorption process comprising crushing, grinding, cyanide leaching of flotation concentrate, carbon adsorption, cyanide destruction, carbon elution and regeneration, gold refining, dry stack tailings and paste backfill.

The process plant is designed with a nominal capacity of 1,500 t/d. The crushing circuit is designed to operate 24 hours per day at a utilization of 33% with sufficient capacity to crush the entire daily production within 12 hours if necessary. The milling and leaching circuits will operate 24 hours per day, 365 days per year at an availability of 92%.

The overall LOM recoveries based on test work are expected to be approximately 94.3% for gold and 57.6% for silver. The grinding circuit product size is targeted at 80% passing (P80) 240 microns and the flotation concentrate will undergo further grinding to P80 50 microns before the leaching stage ........