Overview

Deposit type

• VMS
• Stratabound

Summary:

Deposit Types
The Aripuanã polymetallic deposits are typical volcanogenic massive sulphide (VMS) deposits associated with felsic bimodal volcanism. The Aripuanã VMS deposits have been subsequently deformed and metamorphosed under greenschist facies conditions .

· Host rocks: All mineralized bodies are located on the upper levels of a felsic volcanic unit, in association with finely laminated exhalites, at or close to the contact with an overlying sedimentary unit.
· Mineralization zonality and predominant textures: Three types of mineralization are found at the Project, and are typical of VMS deposits elsewhere:
o Stringer facies: Cu-Au bearing stringers in the footwall of the stratabound mineralization, containing chalcopyrite and pyrrhotite, with stockwork and breccia textures corresponding to hydrothermal feeder zones.
o Proximal sulphide facies: mixed bodies of stratabound massive and disseminated Zn-Pb mineralization, overlying stringer mineralization.
o Geochemical zonality. The Cu/Cu+Zn ratio is higher in the proximity of the copper rich feeder zones and decreases upward from the footwall and towards the distal zinc rich stratabound mineralization.

Facies associated with the feeder zones are located in the middle of the volcanic unit and are characterized by pyrrhotite and/or chalcopyrite stockworks in a zone of intense chloritic hydrothermal alteration. The sulphide association represents a feeder zone at higher temperature. There is Cu-Au association in these zones.

Mineralization
Three main elongate mineralized zones, Arex, Link, Ambrex, and Babaçú, have been defined in the central portion of the Project. Limited exploration has identified additional, possible mineralized bodies including Massaranduba, Boroca, and Mocoto to the south and Arpa to the north.

Where outcropping, sulphide mineralization has been oxidized forming gossanous bodies which frequently mark the position of overthrust faults. These gossans are generally small and contain low levels of gold. They do not appear to be of economic interest at this time.

The individual mineralized bodies have complex shapes due to intense tectonic activity. Stratabound mineralized bodies tend to follow local folds, however, local-scale, tight isoclinal folds are frequently observed, usually with fold axes that are parallel to major reverse faults, causing rapid variations in dips. The Arex, Link, Ambrex, and Babaçú deposits are the best understood and are described below.

Hydrothermal alteration is commonly directly adjacent to the Arex, Link, Ambrex, and Babaçú deposits, and according to Leite et al. (2005, as cited in AMEC, 2007) presents a zonal and symmetrical standard:
· External zone: Sericite and muscovite in a fine-grained matrix with minor chlorite content. Where present, the low sulphide content is dominated by pyrrhotite.
· Intermediate zone: Transition of sericite to chlorite halo on stringer zones. Tremolite and chlorite alteration with minor carbonatization and silicification.
· Internal zone: Stringer zones are characterized by pervasive chlorite alteration accompanied by quartz veins. Sulphide content is dominated by chalcopyrite and pyrrhotite. Porphyroblastic magnetite and biotite locally substitutes within the sulphide matrix. The stratabound zones are dominated by tremolite, talc and carbonate alteration, accompanied by sphalerite, galena, and pyrite, with minor magnetite and fluorite. The stratabound zone may be brecciated.

Arex
Mineralization at the Arex deposit strikes at approximately 110° azimuth, extending over a 1,400 m strike length. Upper portions of the deposit tend to be near-vertical, while lower portions dip at 60° to the northeast. The Arex deposit is characterized by well-defined stringer and stratabound zones. Discrete lenses of stratabound and stringer mineralization, ranging from less than one metre to 15 m thick, interplay within a 100 m to 150 m wide zone, separated by barren, hydrothermally altered rocks. Mineralization comes close to outcropping at surface and extends to almost 500 m below surface. Discrete lenses may be continuous for up to 300 m down dip. The Arex deformation pattern is made of tight, foliation-parallel folds, and reverse faults which overthrust in the same direction. The Arex deposit presents strong dip variations that are often parallel to foliation and faults. In some areas, this may cause the stratabound and stringer mineralization to be parallel, despite its original perpendicular position.

Link
The Link deposit, first discovered in 2014, is interpreted to be the westward extension of the Ambrex deposit towards the Arex deposit. It is located approximately 300 m southeast of the Arex deposit and exhibits shape, mineralization, and alteration features similar to Ambrex, the largest deposit. Link mineralization strikes at approximately azimuth 125° and has a strike extent of approximately 450 m, based on current drilling. Mineralization thicknesses typically range between 10 m and 50 m, with a maximum of 150 m. Mineralization comes close to outcropping at surface and extends to almost 700 m below surface. The degree of folding at Link is gentler than at Arex and hosts well marked overthrust faults, which are parallel to metamorphic foliation. The orientation of the stratabound mineralization is generally parallel to the original bedding, while the stringer zone is often approximately perpendicular to the stratabound zone.

Ambrex
The Ambrex deposit represents the largest of the known mineralized zones on the Project that is included in the LOM plan. The Ambrex deposit is located approximately 1,300 m southeast of the Arex deposit. Ambrex mineralization strikes at approximately azimuth 125° and has a strike extent of approximately 1,050 m, based on current drilling. The dip varies from near vertical to 70° to the northeast. Mineralization thicknesses typically ranges between 10 m and 50 m, with a maximum of 150 m. The Ambrex deposit has an upper depth of 60 m below surface, with a lower depth of approximately 700 m. The degree of folding at Ambrex is gentler than at Arex and hosts well marked overthrust faults, which are parallel to metamorphic foliation. The orientation of the stratabound mineralization is generally parallel to the original bedding, while the stringer zone is often approximately perpendicular to the stratabound zone.

Babaçú
Located southeast of the Ambrex deposit, the Babaçú deposit is 1,300 m long and also dips to the northeast. The Babaçú deposit is interpreted to be similar in shape and style of mineralization to the Ambrex deposit.

Mining Methods

• Longitudinal retreat
• Transverse stoping
• Bench stoping

Summary:

The Aripuana Project is targeted on mining three main elongate mineralized zones, Arex, Link, and Ambrex, that have been defined in the central portion of the Project. The Arex and Ambrex deposits are separate VMS deposits with differing mineral compositions in stratabound and stringer forms and complex geometric shapes.

The deposit geometry is amenable to a number of underground mechanized mining techniques including cut and fill and bulk stoping methods. A nominal production target of 6,065 tpd has been used as the basis for the mine production schedule.

Mining will be undertaken using conventional mechanized underground mobile mining equipment via a network of declines, access drifts, and ore drives. Access to the Arex, Link, and Ambrex deposits will be from separate portals, which will access the deposits from the most favourable topographic locations.

Mine Design
The mine design has been based on using modern mobile trackless equipment with independent decline accesses into the Arex, Link, and Ambrex deposits. The three deposits will be accessed from three independent surface cut and cover portals and ramps designed at a gradient of 14% to be driven with an arched profile and cross sectional area (CSA) of 27 m2 to accommodate the selected major equipment. The main loading and hauling equipment will be 12.5 t class load haul dump units (LHD) combined with 35.5 t class haul trucks.

Main mining sublevels will be spaced 75 m apart, with stope sublevels placed at 25 m spacing. The upper sublevel in each level will contain a five metre sill pillar. The two mining methods will be longitudinal retreat longhole mining, and VRM with primary and secondary sequencing. Backfilling of stopes will be completed using paste fill, cemented rockfill, and rockfill.

Material movement at Aripuanã will be completed via ramps using haulage trucks. Primary development consists of ramps and raises. Secondary development consists of cross cuts, level access, footwall drives, ore drives, and all infrastructure development (sumps, remucks, etc.).

Mining Method
A nominal production target of 6,065 tpd (2.2 million tonne per annum (Mtpa)) has been used as the basis for the Aripuanã production schedule.

Nexa has undertaken a number of mining method option studies, which have selected a combination of longitudinal longhole retreat stoping (bench stoping) for narrow zones and VRM for thicker zones of the deposits. To increase the extraction ratio, a primary and secondary stoping sequence will be used in the VRM areas with cemented paste fill used to backfill stopes. Finished longhole retreat stopes will be backfilled with rockfill.

The primary mining method selected for the Arex deposit is longitudinal retreat mining. The majority of the Link and Ambrex deposits will be mined using VRM, with longitudinal longhole retreat mining utilized in minor areas. The tonnage split between VRM and bench stoping is approximately 60:40.

An estimate of the potentially mineable tonnage has been generated based upon the estimated Mineral Resources. The estimate includes both Measured and Indicated Mineral Resources. DSO was used to generate stope shapes to a minimum dip of 50°. A Minimum Mining Width (MMW) of four metres has been applied.

No hanging wall or footwall dilution was added in the DSO analysis, however, it was accounted for in the mine scheduling.

Backfill
The process plant will produce tailings quantities of approximately 90% of the plant feed. Tailings will be dry stacked on surface or used as backfill for underground voids. It is planned that backfill be placed as consolidated paste fill. The strengths achieved by consolidated paste fill meet the geomechanical requirements for primary and bench stopes.

In general, waste rock will be used as backfill for bench stoping areas and the remaining waste generated will be hauled to the surface and placed in waste dumps.

Production Schedule
A nominal target of 6,065 tpd was used in preparing the mining schedule, with feed to the plant consisting of campaigns of stratabound and stringer material types, managed via stockpiling.

The deposits support a production rate of 2.2 Mtpa, with average annual metal production of:
· Zinc: 72,700 t.
· Lead: 25,200 t.
· Copper: 3,600 t.
· Silver: 1.85 Moz (contained in copper and lead concentrates); and
· Gold: 14,300 oz (contained in copper and lead concentrates).

This average annual production is equivalent to 122,000 t zinc per annum (tpa), after converting other metals based on net revenue.

Comminution

Crushers and Mills

Summary:

Key elements of the process flowsheet include primary crushing, a SAG mill followed by a ball milling and pebble crushing (SABC) circuit, talc pre-flotation, and sequential flotation of copper, lead, and zinc for stratabound mineralization, and copper flotation for stringer mineralization.

ROM material will be trucked from the underground mine to the ROM stockpiles near the primary crushing area. ROM material will be directly discharged into the 80 t capacity (two truckloads) primary crusher dump hopper or held temporarily in four stockpiles based on mineralization type and grade (approx. 7,000 t each, one stringer stockpile and three mixed stockpiles of different grades) and recovered later by front end loader. A static grizzly on top of the dump hopper with 600 mm by 600 mm openings will prevent oversize material from reaching the discharge of the hopper. Discharge of ROM material from the dump hopper will be via apron feeder, which will discharge to a vibrating grizzly with an aperture of 130 mm. Oversize material from the grizzly will feed the primary (jaw) crusher while undersize will bypass the crusher.

The crusher product, with a top size of 140 mm, will be collected on a conveyor belt together with the fines from the apron feeder and grizzly undersize. A metal detector will remove scrap metal from the crushed ore that could damage downstream conveyors and equipment. The conveyor will feed the crushed ore bin with a capacity of 2,500 t. Two additional crushed ore bins may be added at a later stage if required that would bring the combined capacity to 9,100 t. Two variable speed apron feeders per bin will withdraw the crushed product from the bins and deliver it to the grinding circuit via conveyor belt. A belt scale on the conveyor will control the speed of the apron feeders, and as a result the feed rate to the grinding circuit. The grinding circuit will consist of a conventional SAG mill, ball mill, and pebble crusher (SABC configuration). Both grinding mills will have variable speed drives to allow for process optimization over a range of ore competencies and hardnesses.

A double-deck vibrating screen at the discharge of the SAG mill will separate SAG mill discharge into screen oversize (scats and pebbles) and screen undersize. Scats will be separated from the pebbles by belt magnets, leaving the pebbles to be recycled to the SAG mill feed conveyor via the pebble crusher, or directly to the SAG mill feed conveyor when the pebble crusher is undergoing maintenance. Screen undersize will discharge to the grinding circuit pump sump together with the ball mill discharge.

SAG mill discharge screen undersize material and ball mill discharge will be pumped to a set of hydrocyclones that will classify the material into oversize and undersize. The hydrocyclone underflow (oversize) material will return to the ball mill while the hydrocyclone overflow (undersize) with P80 150 µm will be transferred to the flotation feed pump box. An online particle size analyzer will provide periodic measurement of the hydrocyclone overflow stream. A trommel screen on the ball mill discharge will remove scats and the slurry will be combined with the SAG mill discharge screen undersize and recirculated as feed to the hydrocyclones.

Processing

• Crush & Screen plant
• Flotation
• Dewatering
• Filter press

Summary:

The Aripuanã process flowsheet has been developed through metallurgical test work and the use of conventional technologies for the treatment and recovery of copper, lead, and zinc as separate concentrates. Plant throughput is planned to be 2.214 Mtpa of run of mine (ROM) ore from the Arex, Link, and Ambrex underground mines. Two main ore types are present at Aripuanã, stratabound and stringer, that have different hardnesses and therefore different throughput rates. Stratabound material, however, will make up the majority of the ore to be processed (approximately 89%) and the feed blend to the plant is expected to peak at 21% stringer material during Year 5. Estimated processing rates for the two ore types individually based on hardness are approximately 5,000 tpd (dry basis) for stringer material and 6,300 tpd (dry basis) for stratabound material. Throughput for the blended ore is estimated as a weighted average of the throughputs of the two ore types. Key elements of the process flowsheet include primary crushing, a SAG mill followed by a ball milling and pebble crushing (SABC) circuit, talc pre-flotation, and sequential flotation of copper, lead, and zinc for stratabound mineralization, and copper flotation for stringer mineralization.

Flotation
Flotation will be conducted sequentially, i.e., the production from the comminution circuit will pass through four independent circuits in sequence. The first flotation circuit is for talc and light mineral flotation (mainly minerals containing magnesium) to remove naturally hydrophobic minerals and prevent them from contaminating the sulphide concentrates or interfering with sulphide flotation. Due to the high content of light minerals in stratabound mineralization, these minerals must be removed prior to sulphide flotation, however, since stringer mineralization contains only minor amounts of these minerals, processing of stringer ore only would generally by-pass the talc flotation step as the minor talc content can be depressed with carboxymethyl cellulose (CMC). The flotation circuits for the recovery of copper, lead, and zinc follow talc flotation.

Hydrocyclone overflow slurry will be conditioned prior to being fed to the talc flotation circuit, which consists of three column flotation stages: rougher, cleaner, and reverse copper flotation. Reagents added to the conditioning tank for talc flotation include MIBC as a frother and SMBS as a depressant of iron sulphides (pyrite and pyrrhotite). The talc rougher concentrate will be cleaned in the second column, with the cleaned concentrate reporting to the reverse copper flotation column where talc will be depressed with CMC while copper in the talc concentrate is recovered and reports to downstream sulphide flotation. The final talc concentrate will be combined with the sulphide flotation tailings for disposal. The talc flotation tailings containing copper, lead, and zinc minerals (including copper recovered from the talc concentrate during reverse copper flotation), will proceed to the copper flotation circuit.

Prior to conditioning and cleaner flotation, copper rougher flotation concentrate will be reground in a vertical stirred mill to P80 45 µm to increase sulphide liberation and promote cleaner stage recovery. Two stages of cleaning in column cells will produce the final copper concentrate. Copper rougher-scavenger concentrate will be recycled to the rougher flotation feed, while copper rougher-scavenger flotation tailings will feed the lead flotation circuit (stratabound or blended ore). If processing stringer ore, however, the rougher-scavenger flotation tailings can be pumped directly to tailings dewatering . Cleaner circuit tailings are recycled to the copper rougher feed.

Prior to lead rougher flotation, the feed slurry will be conditioned with the aforementioned reagents in two tanks in series. The lead circuit consists of rougher flotation, rougher concentrate regrinding to P80 75 µm, rougher-scavenger flotation, and two-stage cleaner flotation. The product from the lead cleaner flotation circuit will be the final lead concentrate. The lead rougher-scavenger flotation tailings will be pumped to feed the zinc flotation circuit.

The zinc flotation circuit is similar to the copper and lead flotation circuits, however, the quantity of flotation cells and some of the reagents used are different. Lead rougher-scavenger flotation tailings will be thickened prior to being pumped to the zinc flotation circuit. The reagents used for zinc flotation include:
-AERO 208 or A208 (dialkyl dithiophosphate collector for zinc)
-Copper sulphate (sphalerite activator)
-Lime (pH regulator)
-MIBC (frother)

Prior to zinc rougher flotation, the feed slurry will be conditioned with the aforementioned reagents in two tanks in series. The zinc circuit consists of rougher flotation, rougher concentrate regrinding to P80 75 µm, rougher-scavenger flotation, and two-stage cleaner flotation. The product from the zinc cleaner flotation circuit will be the final zinc concentrate. Zinc rougher-scavenger flotation tailings will be pumped to tailings dewatering for disposal.

Thickening and Filtration
Thickening and filtration will be performed on flotation concentrates and tailings. Three concentrates (copper, lead, and zinc) will be produced, with each thickened and filtered separately. Tailings generated from flotation will consist of talc concentrate and sulphide flotation tailings, which are also thickened separately, then combined and filtered.

The copper, lead, and zinc concentrate slurries will be pumped to storage tanks feeding pressure filters dedicated to each concentrate, which will reduce the moisture to approximately 10%. The filtered copper and zinc concentrates will fall by gravity onto belt conveyors that will deliver the material to segregated covered storage areas. These concentrates will be reclaimed by front end loader and loaded into trucks for shipping. Lead concentrate will be bagged in lined supersacs and loaded into containers for shipping. All filtrates will be recovered for re-use in their respective flotation circuits. Excess filtrate and concentrate thickener overflow will be discharged to an engineered wetland treatment system and can then be recycled to the processing plant as make-up water as required or discharged.

Sulphide flotation tailings will be thickened and filtered in three pressure filters and combined with filtered talc concentrate prior to disposal. The sulphide flotation tailings thickener underflow will be filtered to produce a filter cake with approximately 10% moisture that is suitable for dry stacking in two stockpiles (capacity of approximately 5,200 t each). This stockpiled material will be recovered by front end loader and loaded into trucks for transport to the dry stack tailings dump or the paste backfill plant. Filtrates from the pressure filters will be pumped to a recovered water pond and reclaimed for return to the process.

Water Supply

Summary:

The Aripuanã Project water balance requires a top-up fresh water supply of approximately 150m³/h. Nexa has undertaken a water supply engineering study based on the construction of a water dam and creation of a freshwater lake in a valley adjacent to the Aripuanã Project site. Nexa has obtained authorization from the regional authority to construct the dam and to draw up to 378 m³/h of fresh water from the dam to supply the Aripuanã Project.

At the Aripuanã project, Nexa has built a water dam to supply water to the plant.

The freshwater dam will be the source of make-up water for ore processing.

Recovered and Make-Up Water Systems
The water system is designed to maximize water recovery and recirculation. Water from tailings thickener overflow and tailings filtration will be pumped to a recovered water pond with a two-day retention capacity. After treatment with hydrogen peroxide, the water is pumped to a 600 m3 recovered water tank, which will receive make-up water as required. Make-up water will be collected from a storage pond close to the processing facilities and will be pumped to a 400 m3 make-up water tank. This water will be used as make-up for recovered water and for specific uses including feed for the water treatment station, pump seal water, fire suppression, vacuum pump seal water, reagent preparation, potable water, and feed to various points in the plant circuit.

Production

Operational metrics

Personnel

Job Title	Name	Profile	Ref. Date
Consultant - Mining & Costs	Jason Cox		Nov 17, 2020
Consultant - Recovery Methods	Brenna J.Y. Scholey		Nov 17, 2020
General Manager Operations	Evandro Figueiredo Reis Faria		Aug 21, 2024
Mining Manager	Celso Lima		Aug 21, 2024
Technical Services Manager	Vitor Ferraz Viana		Aug 21, 2024