The Property captures the known extent of the Lac des Iles intrusive complex, an irregularlyshaped Neoarchean-age mafic-ultramafic intrusive body having maximum dimensions of approximately 9 km in the north-south direction and approximately 4 km in the east-west direction. The complex is currently interpreted to be made up of three discrete intrusive bodies:

• The North Lac des Iles intrusion, characterized by a series of relatively flat-lying and nested ultramafic bodies with subordinate mafic rocks;
• The Mine Block intrusion (MBI), characterized by an early series of gabbronorite and magnetite-rich gabbronorite and a later series of noritic rocks that host all of the known mineral resources on the Property; and,
• The South Lac des Iles intrusion, comprising noritic rocks having many similarities to the western part of the MBI in terms of petrology and textures.

The results of academic and internal research into the characteristics of PGE-copper-nickel mineralization in the MBI provide the basis for an updated deposit and genetic model. In summary, PGE mineralization in the MBI is found in a variety of structural and geological settings but in general is characterized by the presence of small amounts (less than 1 to 3%) of fine- to medium-grained disseminated iron-copper-nickel sulphides within broadly stratabound zones of PGE and gold enrichment associated with VGAB rocks, coarsegrained noritic rocks and local, intensive zones of amphibolitization, chloritization and shearing. An important, distinguishing characteristic of the MBI mineralization relative to other PGE deposits is the consistently high palladium: platinum ratio, commonly averaging 10:1 or more in most of the known zones. Although many ideas have been advanced to account for the high palladium: platinum ratios, it remains equivocal what specific magmatic and/or post magmatic processes have led to the extreme enrichment in palladium over platinum in the MBI. Sulphide mineralogy is dominated by pyrite with lesser pyrrhotite, chalcopyrite, pentlandite and millerite.

The MBI includes several distinctive styles of mineralization:

• Roby and Offset Zone-style mineralization, comprising:
- a relatively thick and lower-grade footwall zone comprising minor <1-2% disseminated Fe-Cu-Ni sulfide hosted by VGAB and local heterolithic gabbro breccia; and
- a thinner and generally higher-grade, tabular hangingwall zone typically hosted by a schistose and recrystallized norite to melanorite unit (PYXT) containing <1-2% disseminated Fe-Cu-Ni sulfide

• Sub-vertical to steeply-south plunging cylindrical-shaped bodies of sulfide-bearing VGAB associated with variably mineralized norite and confined by older, barren EGAB and/or magnetite gabbro and/or gabbronorite units and showing local elongation along cross-cutting faults (e.g., B2 Zone, Mystery Zone, Sheriff North Zone);
• North VT Rim Zone-style mineralization, featuring low-sulfide palladium mineralization with highly variable but locally elevated palladium grades of up to ~3 oz/t Pd, negligible base metal contents, and associated with a narrow (approximately 1 to 10m thick), sheared package of VGAB, heterolithic gabbro breccia, mafic dikes and granitic veins; and,
• Poddy palladium and disseminated base metal sulfide mineralization associated with localized sulfide-bearing magmatic breccia, VGAB, norite and north-trending and east-northeasttrending faults and shear zones (e.g., Baker Zone, Moore Zone, Creek Zone).

All of these mineralization styles can be classified as belonging to the magmatic sulfide deposit class. Magmatic sulfide deposits are defined as disseminated to massive sulfide deposits that contain economic concentrations of base metals and/or precious metals (PGE, Au, Ag) that occur in mafic or ultramafic rocks in intrusions, intrusive complexes and lava flows (Naldrett, 2004). Like lode gold deposits, the palladium mineralization at the LDI mine property is structurally controlled. However, unlike lode gold deposits, PGE, Au, Ni and Cu were deposited in the MBI by volatile-rich noritic magmas rather than lower-temperature hydrothermal fluids.

Lac des Iles Mine (LDIM) began open pit mining in 1993 and progressed underground in 2005 via the in-pit portal and ramp access into the Roby zone immediately beneath the Roby Pit. Underground production began in 2006, using a longhole mining method with an unconsolidated fill approach, with crown, rib, and sill pillars strategically established to allow top-down mining, avoiding the use of cement in the backfill.

The Roby pushback design is approximately 700m long and brings the overall pit width 600m wide with a surface area of 491,000m2 , with a pit depth of 280m to the designed pit bottom bench of 235m. The pit ramp enters the pit on the southwest side at the 515m elevation. The ramp follows the west side of the existing pit down to the 465m elevation where the first switch back occurs. The second switch back will be at the 405m elevation. At the 3rd switch back, rock fill is required to elevate the switch back to an elevation of 345 to be level with the existing east haul ramp as-built for a second access to the Roby open pit push back. The deepest part of the Roby open pit pushback will be at the 235 m elevation. It will be necessary to build a rock filled ramp to the 235 elevations from Bench 245.

Mining will be carried out by contractors using drill and blast and conventional shovel loading and truck hauling equipment, with haulage trucks in the 70 to 90.4 tonnes capacity range and appropriately sized excavators for ore and waste.

Mining bench heights of 10 m were selected to be within the safe and efficient digging envelope for an 8.1 m3 bucket capacity diesel hydraulic shovel. In some instance, the CAT 990 II or CAT 992G loader will be used in areas where excavators are not applicable to use in the mining operations. The 10 m production mining bench height takes into account swell and the maximum reach of the shovel.

For the 90.4 tonne trucks and 8.1 m3 shovel fleet, a minimum pit bottom width of 35 m has been selected, along with a minimum pushback width of 40 m. The Mining in each bench starts from the ramp and progresses either from the north to south with single face mining.

LDIM has historically used a top down longhole method with unconsolidated fill approach, strategically leaving rib pillars and sill pillars (based on geotechnical requirements) as mining progressed. The Roby Zone level intervals are relatively closely spaced, ranging in spacing from 25 to 35 m, with the 4.5 in (114 mm nominal) diameter down the hole drills used at the time. With the Offset zone, the vertical spacing of levels increased slightly, ranging from 30 to 60 m apart, still using the 4.5 in (114 mm nominal) diameter down the hole drill. Being a thicker ore body and deeper in the mine, the Offset zone left thicker rib pillars separating the stopes, with plans to mine these at a later date at a reduced recovery and higher dilution after all of the primary stopes have been mined.

In 2016, LDIM began the transition of the lower Offset zone to a sub-level shrinkage (SLS) method. This mining approach is still top down with the main difference being no pillars are left. Also in 2016, a definition drilling program better defined the B2 zone to the point that an upper mining plan using long hole open stoping and a lower mining plan of shrinkage or sublevel open stoping could be developed.

The plant is designed to operate for 365 days per year at a 92% availability, which is based on 18 days of scheduled downtime, 2.5 days of curtailment for peak power, and 97.5% operating reliability.

The run-of-mine (ROM) ore feed is crushed prior to reporting to the milling circuit. The ore is first fed to a primary 54” x 75” gyratory crusher driven by a 447 kW (600 hp) motor. Pan feeders feed the conveyor between the stockpile and the main grinding circuit, with the capability of diverting a variable proportion of the coarse ore to the secondary crusher. The proportion fed to the Secondary crushing circuit is regulated by the operator to maximize throughput in the grinding circuit. Secondary crushing is accomplished with a cone crusher (standard HP800), which discharges back on the mill feed conveyor producing a finer feed to the SAG mill.

The SAG mill , which can be operated in autogenous or semi-autogenous configurations, has dimensions of 9.14 m in diameter by 4.27 m equivalent grinding length (EGL), and is driven by a 6,400 kW (8,500 hp) motor. The mill output is passed over a vibrating screen where oversized particles are are directed by conveyor to the pebble crusher (short head cone HP800). The pebble crusher product is then returned to the grinding mill. The pebble circuit is configured to allow oversized particles to bypass the pebble crusher and return directly to the SAG mill feed.

Undersized particles from the SAG mill output are directed by the vibrating screen to the ball mill diverter box, which splits the stream to feed between two ball mills. The two ball mills each have dimensions of 6.10 m in diameter by 10.36 m EGL and are also each fitted with 6,400 kW (8,500 hp) motors. Each ball mill is in closed circuit with a hydrocyclone cluster. The underflows from the hydrocyclones report back to the ball mills and the overflows report to a collection tank. The primary hydrocyclones were upgraded to GMAX 20” and the ball mill media was reduced to 1” high chrome balls in order to produce the current K80 of 55 µm.

The current mill’s SABC configuration can also be augmented with a tertiary grinding circuit consisting of two Vertimills® in close circuit with tertiary cyclone packs to achieve a finer grind. A third Vertimill® exists but is currently not in service. The Vertimills® have 930 kW (1,250 hp) motors installed. At the rated mill capacity of 13,500 tpd volumetric constraints within the tertiary grinding circuit hydrocyclone feed lines limit the quantity that can be processed through tertiary grinding to 80% of the plant feed, while the remaining 20% is directly fed to the rougher conditioning tank prior to flotation. With the addition of the tertiary grinding stage, flotation feed is reduced to 53 µm.

The tertiary grinding circuit includes a Flash flotation cell for the hydrocyclone underflow (commissioned in 2015). The tails from the Flash flotation cell are split into the Vertimills®. The Flash flotation cell concentrate is pumped directly to final concentrate and allows an outlet for fast floating PGMs and base metals to avoid overgrinding. The tertiary grinding circuit, including the Flash flotation cell, can be bypassed with existing valving. The impact of bypassing the teritiary grinding circuit has not been fully evaluated and a trade-off study is recommended to evaluate the cost of the tertiary grind versus the increase in recovery.

The rougher conditioning tank receives feed from either the tertiary grinding cyclone overflow or\ from the of primary hydrocyclone overflow. The conditioner is configured to feed two flotation lines, each with a rougher flotation cell having a capacity of 50 m3. The concentrate from the rougher flotation cells (cells 1 from both line 1 and line 2), report to the 1.7m rougher column cleaner cell (column cell C). The tailings from rougher flotation move forward to a series of scavenger flotation cells having a capacity of 130 m3 each (cells 1B to 1G on line 1 and 2B to 2G on line 2). Tailings from scavenger flotation are final tailings and are pumped to the tailings management facility (TMF) for impoundment. Scavenger concentrates from cells 1A, 2A, 1B, 2B, 1C, and 2C are combined and pumped to the 2nd cleaner feed bank of flotation cells. The scavenger concentrates from 1D, 2D, 1E, 2E, 1F, 2F, 1G, and 2G are combined and pumped to the 1st cleaner feed bank of flotation cells.

The cleaner circuit consists of two main divisions, one for rougher cleaning (3 column cells) and the other for scavenger cleaning (4 conventional banks of cells).

Rougher concentrate reporting to the 1.7m rougher column is cleaned once to produce final concentrate. The tailings from this column are split and fed into the two 1.3m columns (columns A & B) which operate in parallel. The concentrate from these columns are sent to final concentrate while the tailings from the two columns enter the scavenger cleaner circuit at the feed to the 4th cleaner bank (Denver).

Final concentrate streams are combined and pumped to the 17m high-rate concentrate thickener. The thickener underflow is pumped into a 4.3m by 7m storage tank prior to filtration. Two pressure filters (PF-19) are operated to reduce concentrate moisture to 10%. The concentrate is shipped via trucks to to the smelter.