hellyer tailings retreatment project, tasmania...bsc (hons), mausimm signature: coordinating author...

Competent Persons Report

Hellyer Tailings Retreatment Project, Tasmania

CSA Global Report Nº R301.2017 20 November 2017

www.csaglobal.com

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NQ MINERALS PLC HELLYER TAILINGS RETREATMENT PROJECT, TASMANIA

CSA-Report Nº: R301.2017

Report prepared for

Client Name NQ Minerals Plc

Project Name/Job Code NQMCPR01

Contact Name Mike Barden

Contact Title Chief Development Officer

Office Address Finsgate, 5-7 Cranwood St, London, UK

Report issued by

CSA Global Office

CSA Global Pty Ltd

Level 2, 201 Leichhardt Street

Spring Hill QLD 4000

AUSTRALIA

PO Box 1077

Spring Hill QLD 4004

AUSTRALIA

T +61 7 3106 1200

F +61 7 3106 1201

E [email protected]

Division Corporate

Report information

File name R301.2017 NQMCPR01 CPR - NQ Minerals Hellyer Tailings Retreatment Project FINAL_v2

Last edited 31-Dec-17 5:35:00 AM

Report Status Final

Author and Reviewer Signatures

Coordinating Author

David Williams

BSc (Hons), MAusIMM Signature:

Coordinating Author

Peter Cranfield

BSc (Hons) ARSM, MIMMM, C.Eng

Signature:

Peer Reviewer Galen White

BSc (Hons), FAusIMM, FGS Signature:

CSA Global Authorisation

Galen White

BSc (Hons), FAusIMM, FGS Signature:

© Copyright 2017

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Disclaimers

Purpose of this document

This Report was prepared exclusively for NQ Minerals Plc (“the Client”) by CSA Global Pty Ltd (“CSA Global”). The quality of information, conclusions, and estimates contained in this Report are consistent with the level of the work carried out by CSA Global to date on the assignment, in accordance with the assignment specification agreed between CSA Global and the Client.

Notice to third parties

CSA Global has prepared this Report having regard to the particular needs and interests of our client, and in accordance with their instructions. This Report is not designed for any other person’s particular needs or interests. Third party needs and interests may be distinctly different to the NQ Minerals Plc’s needs and interests, and the Report may not be sufficient nor fit or appropriate for the purpose of the third party.

CSA Global expressly disclaims any representation or warranty to third parties regarding this Report or the conclusions or opinions set out in this Report (including without limitation any representation or warranty regarding the standard of care used in preparing this Report, or that any forward-looking statements, forecasts, opinions or projections contained in the Report will be achieved, will prove to be correct or are based on reasonable assumptions). If a third party chooses to use or rely on all or part of this Report, then any loss or damage the third party may suffer in so doing is at the third party’s sole and exclusive risk.

CSA Global has created this Report using data and information provided by or on behalf of the Client [and the NQ Minerals Plc agents and contractors]. Unless specifically stated otherwise, CSA Global has not independently verified that all data and information is reliable or accurate. CSA Global accepts no liability for the accuracy or completeness of that data and information, even if that data and information has been incorporated into or relied upon in creating this Report.

Results are estimates and subject to change

The interpretations and conclusions reached in this Report are based on current scientific understanding and the best evidence available to the authors at the time of writing. It is the nature of all scientific conclusions that they are founded on an assessment of probabilities and, however high these probabilities might be, they make no claim for absolute certainty.

The ability of any person to achieve forward-looking production and economic targets is dependent on numerous factors that are beyond CSA Global’s control and that CSA Global cannot anticipate. These factors include, but are not limited to, site-specific mining and geological conditions, management and personnel capabilities, availability of funding to properly operate and capitalise the operation, variations in cost elements and market conditions, developing and operating the mine in an efficient manner, unforeseen changes in legislation and new industry developments. Any of these factors may substantially alter the performance of any mining operation.

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1 Executive Summary

CSA Global Pty Ltd (CSA Global) was engaged by NQ Mineral Plc (NQM) to compile a Competent Persons

Report (CPR) setting out, in addition to background information, the Mineral Resource estimate and Ore

Reserve estimate relating to the Hellyer Tailings Retreatment Project (Hellyer or Hellyer Project), located

in Tasmania, Australia.

1.1 Tenure

NQM holds a consolidated granted mining leases CML103M/1987 over the area including the tailings dam

and the processing plant. This lease covers all the areas required for the proposed tailings reclaim

operations. The mining lease is held by Hellyer Gold Mines Pty Limited, a wholly owned subsidiary of NQM.

The lease was granted on 24 February 1988 and was renewed and extended in June 2009 until 30 June

2020. The lease can then be further extended, for as long as the mine is operated, as advised by Mineral

Resources Tasmania. The annual permit has been paid in full until 15 February 2018.

1.2 Project Location, Access and Climate

The Hellyer Mine is located in north-western Tasmania 80 km south of Burnie, just off the main highway

between Burnie and Queenstown and four hours’ drive northwest of the capital, Hobart. The area

surrounding the project is mainly forest reserve with farmland. There has been extensive historical mining

activity in the area. The licence area has a polymetallic (Cu-Pb-Zn-Au-Ag) deposit and associated ore

processing facility at Hellyer, adjacent to the tailings dam.

This part of Tasmania lies in a high rainfall area with annual average precipitation of 2,180 mm (Bureau of

Meteorology Waratah) falling all year although higher falls and snowfall occur over winter months.

1.3 Geology and Nature of the Deposit

The Hellyer deposit is a volcanic hosted polymetallic massive sulphide deposit located within the Mount

Read volcanic arc of western Tasmania. This region also hosts similar deposits such as Hercules, Que River,

Roseberry and Mount Lyell. Mineralisation is sulphide hosted and comprised predominantly of pyrite,

with lesser sphalerite, galena and arsenopyrite. The pyrite must be carefully managed when exposed to

air to prevent its potential for oxidation and consequent and potential acid mine drainage issues. The

economic metals mined at Hellyer were lead, zinc, copper, gold and silver. The Hellyer deposit was mined

by underground methods during the period 1989 to 2000. The Fossey deposit extends down plunge from

the Hellyer deposit and was mined by Bass Metals Pty Ltd (Bass Metals) between 2010 and 2012.

Tailings from the mill were deposited in a depression approximately 1 km to the west of the Hellyer mill.

The tails were simultaneously inundated with water to prevent oxidation of the sulphide species present

in the tails. The tails were partially dredged by Polymetals Group (Polymetals) and retreated between

2006 and 2008, with the reprocessed tails deposited in the Shale Pit and Western Arm. Tails from

processed Fossey deposit ore were also discharged into the Western Arm.

1.4 Mineral Resource Estimate

The Mineral Resource estimate for the Hellyer Tailings Storage Facility (TSF) is presented in Table 1. The

Mineral Resource estimate is reported in accordance with the JORC Code1. Table 2 presents the metal

totals, as calculated from Table 1 (only gross totals are presented).

1 Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves. The JORC Code, 2012 Edition. Prepared by: The Joint Ore Reserves Committee of The Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia (JORC).

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CSA Global considers that data collection techniques are consistent with industry good practice and

suitable for use in the preparation of a Mineral Resource estimate. Vibracore drill samples were used to

interpolate grades into blocks using ordinary kriging. Several methods were used to validate the block

model, including visual review and a comparison of sampling and block model grades. A three-dimensional

block model representing the mineralisation was created using Datamine software.

The Mineral Resource is considered to have reasonable prospects for eventual economic extraction (a

required condition under JORC) due to the volume and grade of mineralisation, the tailings are readily

amenable for extraction by dredging, and the project has access to critical infrastructure, including a

processing plant. The potential economic viability has also been assessed with respect to the preparation

of Ore Reserves and economic analysis (see Sections 10 and 11 of this report).

Table 1: Hellyer Tailings Storage Facility, Mineral Resource estimate

JORC classification

Gross Net attributable

Operator Tonnage (Mt)

Zn %

Pb %

Ag g/t

Au g/t

Cu %

Tonnage (Mt)

Zn %

Pb %

Ag g/t

Au g/t

Cu %

Measured 2.05 3.31 3.35 94 2.63 0.2 2.05 3.31 3.35 94 2.63 0.2

NQM Indicated 5.99 2.29 2.95 93 2.55 0.18 5.99 2.29 2.95 93 2.55 0.18

Inferred 1.21 1.00 2.60 86 2.57 0.19 1.21 1.00 2.60 86 2.57 0.19

Total 9.25 2.35 2.99 92 2.57 0.19 9.25 2.35 2.99 92 2.57 0.19

Note: No lower cut-off reporting grade has been applied. Differences may occur due to rounding. (Datamine model: hel717md.dm).

Table 2: Hellyer Tailings Storage Facility, Mineral Resource estimate – metal tonnes and ounces, gross total only, all attributable to NQM

JORC classification Gross

Tonnage Zn (t) Pb (t) Ag (oz) Au (oz) Cu (t)

Measured 2,050,000 67,900 68,700 6,195,400 173,300 4,100

Indicated 5,990,000 137,200 176,700 17,910,200 491,100 10,800

Inferred 1,210,000 12,100 31,500 3,345,600 100,000 2,300

Total 9,250,000 217,400 276,600 27,360,300 764,300 17,600

Note: Metal tonnages and ounces rounded from calculated values.

1.5 Previous Mining Studies

During the period 2006 to 2008, Polymetals successfully dredged and processed approximately 2 million

tonnes (Mt) of tailings at Hellyer, targeting the Zn value through the sale of a bulk Zn/Pb concentrate.

Operations ceased in September 2008.

In 2010, CSA Global prepared a Resource Report for the Hellyer tailings as part of a Resource Estimate and

Mining Study (CSA Global, 2010). This report was used by Como Engineers Pty Ltd (Como) as a part of a

broader project evaluation for the then owners of Hellyer, Bass Metals.

In 2013, Ivy Resources Pty Ltd (Ivy Resources) acquired Hellyer and in November 2013 completed a

Feasibility Study of the Hellyer Tailings Retreatment Project, which presented results including marketing

studies, geology, Mineral Resource estimation, mining, mineral processing, infrastructure, permitting, and

financial analysis (Ivy Resources, 2013). The mining technology considered in the study was dredging.

As part of the Ivy Resources Feasibility Study, Como completed a “Definitive Feasibility Study” (DFS) on

the project with a specific scope related to direct cyanide leach and leaching of the Albion process product

(Como, 2013). This was updated in 2015 to include costs for various feed rates to the cyanide leach circuit

and subsequent processes (Como, 2016).

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In 2016, Ausenco produced a report on the tailings treatment restart cost estimate following a request by

NQM based on a refined processing flowsheet for sequential flotation of the tailings (Ausenco, 2016).

NQM also requested Pitt & Sherry to prepare a report in November 2016 to assess the status of

metallurgical testwork in connection with the sequential flotation of the Hellyer tailings (Pitt & Sherry,

2016).

A relevant study was also conducted by Commodity and Mining Insight Ltd (CM Insight, 2016) that

included preliminary economic modelling of the potential reprocessing of the Hellyer tailings.

1.6 Current Mining Study

In 2017, AusGEMCO Pty Ltd (AusGEMCO) were engaged by NQM to complete an Ore Reserve estimate

over the Hellyer Tailings Project. This study is documented in a report titled “Hellyer Tailings Retreatment

Project – Tasmania, Ore Reserve Estimate” dated September 2017. This report, along with supporting data

and information compiled during the study was made available to CSA Global for review and CSA Global

completed a gap analysis and peer review of this study prior to inclusion of the study contents in this

report as current material information.

A summary of this study is presented in the body of this CPR, referenced as appropriate and augmented

by CSA Global independent comment as appropriate.

1.6.1 Mine Design and Planning

The mine design for the Hellyer Tailings Retreatment Project has been developed by AusGEMCO by

modelling the profit from the exploitation of each block of the Mineral Resource block model (CSA Global,

2017) and by assessment of the Net Smelter Return from selling concentrates as final products of the mine

production. For the purpose of this study three types of concentrate are planned as final products: lead

concentrate, zinc concentrate and pyrite/gold/silver concentrate.

The Block Profit Modelling is organised for each impoundment area comprising the Hellyer TSF. The

economic inputs and mining and processing assumptions are based on those used in the financial model

described in Section 10. Two criteria are used in the block modelling: Block Unit Profit per dry metric tonne

of in-situ tailings (US$/dmt) of a block and Block Profit (US$) per block of the geological model.

The results obtained for the Main Dam consider the tails topography to the terrain bottom after the

dredging operations that were conducted by Polymetals during the period November 2006 to August

2008. A further variant covers the full Mineral Resource including the reprocessed tailings deposited in

the Shale Pit as a result of this previous dredging.

The average Block Unit Profit estimate is US$51.84/dmt of tails and the standard deviation is US$5.25/dmt.

The results of the Block Profit (US$) vary within a range from US$100,000 up to US$1,400,000 per block

of the geological model. The (X, Y) size of a block is 25 m x 25 m while block height is variable. The entire

area of Main Dam is characterised with positive estimates for both the economic criteria used.

For the purposes of mining, the sequencing of panels is based on the analysis of the Block Unit Profit

model of the dam.

Dredging is the major mining method for the extraction of Hellyer tailings. Hydraulic mining with slurry

pumps mounted on a platform and water cannon wash are the other mining methods to be used.

Production scheduling of Hellyer Tailings Dam is organised using the dredging assumptions as follows:

dredging rate = 3,234 dmt/day or 154 dmt/hour (21 hours per day); resources time of dredging = 355 days

per year (10 days for planned maintenance); number of shifts = two shifts per day (each shift = 12 hours);

and working days per week = seven days/week. These assumptions are based on analysis of the dredge

performance during the 2006 to 2008 dredging period.

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1.6.2 Environmental, Law Compliance and Permits

Caloundra Environmental Pty Ltd and Hellyer Gold Mine Pty Ltd (HGM) have conducted preliminary

assessments to identify key environmental aspects associated with the proposed development.

Emission limits have been set by the Environmental Protection Authority (EPA) to manage the point source

pollution (the TSF outflow), and the environmental authority for the tailings reprocessing operation, PCE

7386, sets emission limits from the Main Dam in its condition EF2.

Historically, zinc has been the most difficult limit to meet at the TSF outfall. Since the Aberfoyle Resources

Ltd operation ceased production in June 2000 and closed in 2003, sulphidic tailings have been left

exposed in the upper reaches of the eastern arm of the Main Dam and in Mill Creek. The acidity generated

from these sulphides caused ongoing issues with pH in the Main Dam. The pH in the Main Dam is

important due to the relationship between a pH above 8.0 and total zinc concentrations at the TSF outfall.

HGM’s improved management protocols, such as adding a lime slurry directly into the eastern arm

spillway before it overflows into the main TSF, have been responsible for most of the improvements seen

since 2016.

Environmental licence conditions for the site since 2006 have required a minimum pH of 8.0 and a

maximum Total Zn of 0.8 mg/L at the TSF outfall. A review of long-term water quality records indicates

that with good management procedures and the remediation proposed by HGM, this should be readily

achievable going forward.

HGM holds the consolidated mining lease, CML 103M/87 and the environmental licences PCE 7386

(tailings mining and reprocessing) and PCE 7759 (Fossey underground mine). HGM plans to reprocess

tailings under PCE 7386.

The mining lease encompasses the area at Hellyer where the tailings impoundment and the processing

plant are situated. The original lease (CML 103M/1987) was granted on 24 February 1988. It was renewed

and extended in June 2009. It is currently valid until 20 June 2020. An application for an extension of the

lease can only be made no more than three months before and one month after the lease ceases to be

effective.

In October 2017, following submission by HGM, the EPA of Tasmania approved the Environmental

Management Plan (EMP) for implementation of the Hellyer Project, which was prepared and submitted

by Caloundra Environmental Pty Ltd in accordance with the Condition G7 of Permit Conditions –

Environmental (PCE) No. 7386.

1.6.3 Mineral Processing

A processing plant exists at Hellyer with all the relevant equipment required to reprocess tailings. The

plant was used previously to produce a range of concentrates from mined ore and a bulk concentrate

from previous tailings reprocessing.

The plant is a relatively modern base metals processing facility that was commissioned in early 1989 and

designed to treat the complex Hellyer fine grained copper-silver-lead-zinc orebody. Its initial capacity was

for 1.0 million tonnes per annum (Mt/a) with a circuit consisting of a semi-autogenous grind (SAG) mill,

ball mill and sequential copper, lead, zinc and bulk flotation circuits. The plant was expanded to 1.25 Mt/a

in early 1990 with the addition of pebble crushing, DSM screens in the SAG mill circuit and extra lead and

bulk circuit flotation capacity.

Early processing performance was typified by low metal recoveries to concentrates, but these improved

during subsequent operation.

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This plant is in reasonable condition and currently under a maintenance program that includes some

activities focused on preparation for the proposed retreatment project. However further refurbishment

will be required before it can be brought back into production.

The processing plant is expected to operate at an average tailings feed rate of 154 dmt/hour with

availability of 92%. This performance was achieved during the previous dredging operations (2006 to

2008). The plant is already equipped with a high degree of instrumentation, automation and process

control equipment enabled by online, real-time measurement and recording.

The metallurgical processing characteristics of the Hellyer ore have been extensively tested and are well

known. Recent tests have been conducted by ALS Laboratories (ALS) in Burnie and reviewed by Pitt &

Sherry Group (2016) with the objective of defining a new flowsheet for the reprocessing of Hellyer tailings

and optimising metal recoveries.

Based on the analysis of these test results, understanding of the potential reprocessing characteristics has

been improved and a number of options to optimise the process identified.

Arsenic (As) is the only significant deleterious element in the proposed concentrate production. NQM

advised that it will target As grades below 1% in the planned concentrate products and/or use the blending

capabilities of marketing agents to enable product saleability.

Table 3 below summarises the assumptions about concentration production in the current study.

Table 3: Planned parameters of concentrate production

Parameter Unit Lead concentrate Zinc concentrate Au/Ag/pyrite concentrate

Pb grade % 37 5

Pb recovery % 47

Zn grade % 6 45

Zn recovery % 38

Au grade g/t 6.9 2 2.77

Ag grade g/t 850 160 64

Concentrate mass recovery % 51

The residuals from tailings retreatment in the Hellyer plant need additional storage capacity due to the

specific characteristics of the residuals and the proposed use of dredging for mining. The Hellyer residue

will be stored underwater to prevent potential oxidisation and acid formation.

The total quantity of residue that will be released from the processing plant after the treatment of the

tailings (9.7 Mt) is 4.2 Mt.

A new TSF (TSF2) will be built and ready for storage of the residue by April 2019 (Longey, 2017). The

embankment is proposed to be located approximately 550 m downstream of the Hellyer Main Dam,

located approximately 2 km northwest of the Mill Site. The initial phase will hold 3.0 Mt, with a planned

lift, scheduled for 2024, that will provide storage for the remaining 1.2 Mt. The new TSF2 footprint is

shown in Figure 1.

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Figure 1: New TSF2 downstream of Main Dam (after Longey, 2017)

For the period from April 2018 until March 2019, a quantity of 0.49 Mt residue will need to be temporarily

stored in accordance with the planned schedule detailed in Section 9. The Finger Pond will be used for

this purpose and to enable this the tailings that are currently located in the Finger Pond will be transferred

to the Main Dam during the pre-production period. The quantity of 0.49 Mt residue temporarily stored in

the Finger Pond will be transferred to TSF2 by using slurry pumps prior to lowering the water level in the

Main Dam and Finger Pond as described in the mining plan.

The transport of the residue from the processing plant to the storage dams will be done with a slurry

pump. Deposition of residuals into the storage areas will be sub-aqueous. This approach will minimise any

risks associated due to contact with oxygen in the air and potential oxidation.

1.6.4 Economic Evaluation

All capital cost estimates for the project were supplied to AusGEMCO by NQM. The total estimate is

US$31,253,570. All costs are projected over the LOM period of 114 months (March 2017 until September

2026 for the purposes of evaluation). Most of the capex is allocated over the first two years of the project

commencement. The highest component of the capital expenditure is the cost for building a new TSF

(US$16,136,364).

NQM supplied the operating costs for the Hellyer Tailings Project. Table 4 below reflects a summary of

some operating costs per unit product and run of mine (ROM) tailings.

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Table 4: Hellyer Tailings Project operating expenditure

Cost Unit Value

Dredging A$/dmt 2.84

Hydraulic mining A$/dmt 2.84

Processing A$/dmt 18.5

Logistics EXW to FOB A$/wmt 18.55

Logistics FOB to CIF A$/wmt 25

Sales and marketing A$/month 20,000

Shutdown maintenance A$/month 400,000

General and administration A$/LOM 12,234,527

Royalty Percent net revenue 5

Mine closure/rehabilitation A$ 2,000,000

Two variants of the economic evaluation of Hellyer Tailings Retreatment Project have been assessed.

The first variant reflects a discounted cash flow (DCF) model based on the Ore Reserve estimate of the

Hellyer Project. The basic metrics of the analysis are presented in Table 5 below. The NPV is positive

(US$113.2 million) for the case that incorporates the acquisition cost of the Project. The Discounted Cash

Flow Internal Rate of Return (DCFIRR) is 91% which is higher than the project discount rate of 10%. The

NPV calculated without the acquisition cost is US$128.5 million and the DCFIRR is 197%. This gives grounds

to accept the Hellyer Tailings Retreatment Project as economically effective, which is a precondition for

reporting its Ore Reserve in accordance with the JORC Code (2012).

Table 5: Summary of the DCF model results based on Ore Reserve estimate

Parameters Unit Value

Net revenue US$M 628

Operating expenses US$M 343

Cash flows before tax US$M 285

Cash flows after tax and acquisition US$M 183

With Acquisition Cost

NPV US$M 113.2

DCFIRR % 90.94

DCFPBP months 28

Without Acquisition Cost

NPV US$M 128.4

DCFIRR % 197.42

The second variant is an option of the evaluation of the Hellyer Project with the inclusion of the Inferred

Mineral Resource of Shale Pit which could be extracted at the end of the life of mine (LOM). This

assessment is a requirement of the 2012 JORC Code, Table 1 “Modifying Factors or Assumptions” about

the manner in which the Inferred Mineral Resources are utilised and the sensitivity of the outcome of

their inclusion. The resource of Shale Pit is 1.2 Mt and the total quantity of tailings included in the DCF

model for this variant is 9.7 Mt.

The basic metrics of the DCF model are summarised in Table 6. The NPV is positive (US$124.7 million) for

the case that incorporates the acquisition cost of the Project. The DCFIRR is 91% which is higher than the

project discount rate of 10%. This give grounds to accept the Hellyer Tailing Project as economically

effective. The Discounted Cash Flow Payback Period (DCFPBP) is 28 months. The NPV calculated without

the acquisition cost is US$139.9 million and the DCFIRR is 197%.

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Table 6: Summary of the DCF model results with inclusion of the Shale Pit resource







NPV US$M 124.7

DCFIRR % 91.26

DCFPBP months 28


NPV US$M 139.9

DCFIRR % 197.46

1.7 Ore Reserve Estimate

The Hellyer TSF Ore Reserve estimate is reported in Table 7 and was compiled for NQM as a part of the

mine planning study undertaken by AusGEMCO. This study followed a process of detailed mine

optimisation using AusGEMCO’s original method that incorporated Geological Block Profit Modelling,

mine design, sequencing, production scheduling economic evaluation and mine project risk assessment.

The resulting Ore Reserve estimate was prepared using the Guidelines of the Australasian Code for

Reporting of Ore Reserves (JORC Code, 2012).

The Ore Reserve estimate relates specifically to the conversion of Measured and Indicated Mineral

Resources of the Hellyer TSF Project and includes consideration of the modifying factors documented in

Appendix 1 – JORC Code Table 1.

No cut-off grade(s) has been used in the reserve estimation due to the specific nature of mining the tailings

by dredging. Details on the implementation of other modifying factors are presented in Section 3 of the

AusGEMCO technical report (AusGEMCO, 2017), and are summarised in Section 7 of this report.

All material classified as Measured Mineral Resources in the Hellyer TSF has translated into Proved Ore

Reserves while all material classified as Indicated Mineral Resources has translated to Probable Ore

Reserves. The reserve conversion is based on the CSA Global Mineral Resource estimate updated 24

August 2017 (CSA Global, 2017).

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Table 7: Hellyer TSF Ore Reserve estimate* (6 September 2017)

Ore Reserves

Gross Operator

Tonnage (Mt)

Grade Contained metal

NQM

Zn % Pb % Ag g/t Au g/t Cu % Zn (t) (‘00)

Pb (t) (‘00)

Ag (oz) (‘000)

Au (oz) (‘000)

Cu (t) (‘00)

Proved 2.05 3.31 3.35 94 2.63 0.21 679 687 6,212 173 43

Probable 5.99 2.29 2.95 93 2.55 0.18 1,372 1,767 17,941 491 108

Total 8.04 2.55 3.05 93 2.57 0.19 2,050 2,454 24,153 664 151

Net attributable

Tonnage (Mt)


Zn % Pb % Ag g/t Au g/t Cu %

Zn (t) (‘00)

Pb (t) (‘00)

Ag (oz) (‘000)

Au (oz) (‘000)

Cu (t) (‘00)

Proved 2.05 3.31 3.35 94 2.63 0.21 679 687 6,212 173 43

Probable 5.99 2.29 2.95 93 2.55 0.18 1,372 1,767 17,941 491 108

Total 8.04 2.55 3.05 93 2.57 0.19 2,050 2,454 24,153 664 151

* Mt tonnes and Zn, Ag, Au and Cu grades are rounded to two decimal places, Ag to the nearest whole number. Contained metal tonnes are reported to nearest hundred, and ounces to nearest thousand tonnes.

Table 8: Hellyer TSF non-JORC Reserve deposited material (6 September 2017)

Non-JORC Reserves

Gross Operator

Tonnage (Mt)


NQM


Pb (t) (‘00)

Ag (oz) (‘000)

Au (oz) (‘000)

Cu (t) (‘00)

Fossey 0.45 1.84 1.37 35 1.80 0.03 83 62 512 26 1

Shale Pit 1.21 1.00 2.60 86 2.57 0.19 121 315 3,346 100 23

Total 1.66 1.23 2.27 72 2.36 0.15 204 376 3,857 126 24

* Mt and Zn, Ag, Au and Cu grades are rounded to two decimal places, Ag to the nearest whole number. Contained metal tonnes are reported to nearest hundred, and ounces to nearest thousand tonnes.

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Table 8 above shows the estimates of additional deposited tailings material at Hellyer. This is not classified

as a JORC Code Ore Reserve category. The Fossey tailings were deposited on top of the existing tailings in

the Western Arm Dam after preparation of the 2010 CSA Global Mineral Resource estimate. At present,

the Fossey tailings are unclassified while the tailings beneath them are classified as a Probable Reserve.

However, the extraction of the Probable Reserve of the Western Arm Dam is possible only after the

extraction of Fossey tailings which lie above them. For this reason, the Fossey tailings have been taken

into account in the mine plan and financial model of Hellyer Project.

Additional deposited tailings materials are located in the Shale Pit. The recent update of the Hellyer

Resource model in 2017 has classified these tailings as an Inferred Mineral Resource, which is insufficient

for their translation into a JORC Ore Reserve category. These tailings have an economic potential and have

been analysed in the aforementioned variant to mine plan and financial model for the Hellyer Project.

1.8 Risk Assessment

Ore Reserve Risk has been assessed by using the Monte Carlo method to generate randomly distributed

sample estimates for geological parameters such as in-situ Pb grade, Zn grade, Au grade, Ag grade and

tailings density. This risk evaluation used a similar cash flow model to that developed for the economic

evaluation of the Hellyer Tailings Retreatment Project in Section 10 of the Report. However, because this

risk model applied a simplified assumption to the timing of tax payments it resulted in a slightly different

deterministic Project NPV (US$118.9 million) from the Project NPVs referenced above.

For the purpose of Ore Reserve Risk, all other inputs to the cash flow model are assumed to be constant.

Two variants were used for the assessment of the standard deviations of the geological variables:

• Variant 1 used the monthly estimates of the grades in the production schedule to determine the

standard deviation.

• Variant 2 assessed standard deviations by using the results of CSA Global variogram modelling of the

Hellyer Mineral Resource (CSA Global, 2010), which captures the short-range variability.

Estimates of variation of tailings density were obtained using a standard deviation of 10% around the

assumed constant tailings density (1.93 t/m3) as applied in the Geological Block Model.

Two criteria are used for assessing the Ore Reserve Risk:

• Risk that the stochastic NPV is less than the deterministic NPV.

• Risk that the stochastic NPV is less than zero.

1.8.1 Variant 1

The estimates of the Ore Reserve Risk are presented in Table 9. The criteria using the stochastic project

NPV (US$122.5 million) relative to the deterministic Project NPV (US$118.9 million) as a critical level of the

Ore Reserve risk shows a risk estimate of 32%, which means that the stochastic NPV is higher than the

deterministic NPV for 68% of estimates. In other words, this also means that the deterministic NPV, using

only average estimates of the tonnage and grades of Ore Reserves, is an underestimate. The results using

the second criterion of the risk analysis (Table 9) indicate a zero Ore Reserve Risk because the stochastic

Project NPV (US$122.5 million) is much higher than zero.

These results give ground to conclude that the current estimates of the variability of the tailings grades

and density will not have a significant negative impact on the successful and profitable exploitation of

Hellyer tailings.

XIII



Table 9: Ore Reserve Risk estimates – Variant 1

NPV mean (US$) NPV Std Dev (US$) Critical level (US$) Probability (%) Risk (%)

122,494,493 7,724,451 118,901,024 67.91 32.09

122,084,493 7,724,451 0 100 0

1.8.2 Variant 2

The estimates of the Ore Reserve Risk using Variant 2 are summarised in Table 10.

The risk that the stochastic Project NPV (US$120.9 million) is less than the deterministic NPV

(US$118,901,024) is 35.24%, while the risk that the stochastic Project NPV is less than zero is 0%. These

estimates confirm the validity of the conclusions made for Variant 1 that:

• The stochastic NPV is higher than the deterministic NPV, which indicates an underestimation of the

deterministic NPV.

• The stochastic NPV is significantly higher than zero.

• The estimates of ore grades and density will not have a significant negative impact on the profitable

exploitation of Hellyer tailings.


NPV mean (US$) NPV Std Dev (US$) Critical level (US$) Probability (%) Risk (%)

120,940,367 5,381,536 118,901,024 64.76 35.24

120,940,367 5,381,536 0 100 0

The Mining Project Risk assessment followed the methodology described in the references (Halatchev et

al., 2005; Halatchev, 2007; Davis et al., 2007). This methodology deals with the concept of the

development of a stochastic DCF model over LOM and treatment of all parameters of the mining project

as random quantities or functions depending on available input data.

The criterion for defining the Mining Project Risk are the DCFs of the mining project assessed with the

discounted cash flow model. This criterion is based on the risk of not achieving positive discounted cash

flows at each time step of the DCF analysis, which is a logical requirement of mining business. Such a

formulation of the Mining Project Risk model means that a strategy for achieving positive DCFs over the

LOM is set by the mining company.

To assess the Mining Project Risk, 300 Monte Carlo simulations were run, and their analysis indicates a

negative cash flow for the first period till the 14th month. After that period all cash flows are positive except

for the 23rd and 83rd months, which are affected by the increase of capex and mine closure bond costs

(23rd) and the increase in mining costs because of using water cannon washing technology and additional

capex (83rd).

The payback period of the stochastic cash flows profile is 28 months. As noted above, it is important to

note that the cash flows are assessed with income tax paid every month of the year.

The results of the Mining Project Risk assessment indicate that the risk is 100% for the period until the 14th

month. This is to be expected because the planned cash flows are negative for that period. After that

period the risk varies within the range from 0 to 10%. There are three short periods of higher risk after

the 14th month, the first two are due to the negative cash flow months mentioned above, and the third

during shutdown of the mining operations and mine closure.

Generally, the predicted profile of the Mining Project Risk is attractive for investment and useful for the

decision-making process. It can also help the management of NQM in shaping a strategy about the

mitigation of Mining Project Risk and achieving sustainable exploitation of the Hellyer tailings.

XIV



Contents

Report prepared for ............................................................................................................................................ I

Report issued by ................................................................................................................................................. I

Report information ............................................................................................................................................. I

Author and Reviewer Signatures ........................................................................................................................ I

DISCLAIMERS ......................................................................................................................................................... II

Purpose of this document ................................................................................................................................. II

Notice to third parties ....................................................................................................................................... II

Results are estimates and subject to change .................................................................................................... II

1 EXECUTIVE SUMMARY ............................................................................................................................... III

1.1 Tenure ................................................................................................................................................ III

1.2 Project Location, Access and Climate ................................................................................................. III

1.3 Geology and Nature of the Deposit .................................................................................................... III

1.4 Mineral Resource Estimate................................................................................................................. III

1.5 Previous Mining Studies ..................................................................................................................... IV

1.6 Current Mining Study .......................................................................................................................... V 1.6.1 Mine Design and Planning .......................................................................................................... V 1.6.2 Environmental, Law Compliance and Permits ........................................................................... VI 1.6.3 Mineral Processing .................................................................................................................... VI 1.6.4 Economic Evaluation ............................................................................................................... VIII

1.7 Ore Reserve Estimate .......................................................................................................................... X

1.8 Risk Assessment ................................................................................................................................ XII 1.8.1 Variant 1 ................................................................................................................................... XII 1.8.2 Variant 2 .................................................................................................................................. XIII

2 INTRODUCTION ........................................................................................................................................... 1

2.1 Context, Scope and Terms of Reference ............................................................................................. 1

2.2 Compliance with the JORC Code ......................................................................................................... 1

2.3 Principal Sources of Information ......................................................................................................... 1

2.4 Authors of the Report – Qualifications, Experience and Competence ................................................ 2

2.5 Independence ...................................................................................................................................... 3

2.6 Declarations ......................................................................................................................................... 3

2.7 Results are Estimates and Subject to Change...................................................................................... 3

2.8 About this Report ................................................................................................................................ 4

2.9 Competent Person Statement ............................................................................................................. 4

3 PROPERTY DESCRIPTION AND TENURE........................................................................................................ 5

3.1 Property Location and Climate ............................................................................................................ 5

3.2 Property Description ........................................................................................................................... 6

3.3 Tenure ................................................................................................................................................. 7

4 GEOLOGY AND PROJECT HISTORY ............................................................................................................... 8

4.1 Geology ................................................................................................................................................ 8

XV



4.2 Project History ..................................................................................................................................... 8 4.2.1 Hellyer Base Metals Mine ........................................................................................................... 8 4.2.2 Metallurgical Accounting ............................................................................................................ 8 4.2.3 Retreatment of Tails ................................................................................................................... 8 4.2.4 Fossey Base Metals Mine ........................................................................................................... 9

5 MINERAL RESOURCE ESTIMATE ................................................................................................................. 11

5.1 Exploration History ............................................................................................................................ 11

5.2 Mineral Resource Estimate................................................................................................................ 12

6 PREVIOUS MINING STUDIES ...................................................................................................................... 15

7 CURRENT MINING STUDY .......................................................................................................................... 17

7.1 Mine Design and Planning ................................................................................................................. 17 7.1.1 Mine Design based on Block Profit Modelling .......................................................................... 17 7.1.2 Production Scheduling.............................................................................................................. 22

8 ENVIRONMENTAL, LAW COMPLIANCE AND APPROVED PERMITS ............................................................. 27

8.1 Key Environmental Aspects and their Management ......................................................................... 27

8.2 Relevant Legislation, Regulations, Codes and Polices ....................................................................... 30

8.3 Environmental Licences and Permits................................................................................................. 31

8.4 Environmental Permitting History ..................................................................................................... 32

9 MINERAL PROCESSING .............................................................................................................................. 33

9.1 Prior Processing Plant Preparation .................................................................................................... 33

9.2 Current Plant Configuration .............................................................................................................. 34

9.3 Planned Plant Operation ................................................................................................................... 35

9.4 Planned Concentrate Production ...................................................................................................... 37

9.5 Tailings Residue Storage .................................................................................................................... 42

10 ECONOMIC EVALUATION .......................................................................................................................... 46

10.1 Capital Cost Estimates ....................................................................................................................... 46

10.2 Operating Cost Estimates .................................................................................................................. 47

10.3 Financial Model Assumptions ............................................................................................................ 47

10.4 Discounted Cash Flow Model ............................................................................................................ 48 10.4.1 Discounted Cash Flow Model based on the Ore Reserves Estimate ........................................ 48 10.4.2 Discounted Cash Flow Model with Inclusion of Shale Pit Inferred Mineral Resource .............. 48

11 ORE RESERVE ESTIMATE ............................................................................................................................ 53

12 RISK ASSESSMENT ..................................................................................................................................... 57

12.1 Ore Reserve Risk Assessment ............................................................................................................ 57

12.2 Mining Project Risk Assessment ........................................................................................................ 59

13 CONCLUSIONS ........................................................................................................................................... 64

14 REFERENCES .............................................................................................................................................. 66

15 GLOSSARY ................................................................................................................................................. 67

XVI



16 ABBREVIATIONS AND UNITS OF MEASUREMENT ...................................................................................... 69

Figures Figure 1: New TSF2 downstream of Main Dam (after Longey, 2017) ...................................................................... VIII Figure 2: Project location ............................................................................................................................................ 5 Figure 3: Site layout .................................................................................................................................................... 6 Figure 4: Collar plan for Hellyer TSF .......................................................................................................................... 11 Figure 5: Zn % grade variation with depth, by year of drilling .................................................................................. 12 Figure 6: Block Unit Profit spatial distribution of Main Dam .................................................................................... 18 Figure 7: Block Profit spatial distribution of Main Dam ............................................................................................ 18 Figure 8: Block Profit spatial distribution of Finger Pond Dam ................................................................................. 19 Figure 9: Block Profit spatial distribution of Western Arm Dam ............................................................................... 19 Figure 10: Block Profit spatial distribution of Shale Pit ............................................................................................... 20 Figure 11: Mine sequence of Main Dam for benches B1 and B2 ................................................................................ 22 Figure 12: Monthly tails production schedule ............................................................................................................ 23 Figure 13: Monthly Zn head grade distribution .......................................................................................................... 24 Figure 14: Monthly Pb head grade distribution .......................................................................................................... 24 Figure 15: Monthly Au head grade distribution .......................................................................................................... 25 Figure 16: Monthly Ag head grade distribution .......................................................................................................... 25 Figure 17: TSF outflow – pH versus total zinc (operations compared with closure) ................................................... 28 Figure 18: Emission improvement since 2016 ............................................................................................................ 29 Figure 19: TSF discharge pH and Total Zn since January 2016 .................................................................................... 29 Figure 20: TSF discharge pH versus Total Zn ............................................................................................................... 30 Figure 21: Hellyer mining lease ................................................................................................................................... 32 Figure 22: Hellyer Tailings lead and zinc grades by year ............................................................................................. 33 Figure 23: Inside view of the processing plant ............................................................................................................ 35 Figure 24: Flowsheet of the Hellyer processing plant ................................................................................................. 36 Figure 25: Tailings storage tanks behind the mill ........................................................................................................ 37 Figure 26: Schematic view of the proposed flotation technology .............................................................................. 38 Figure 27: Schedule of zinc production in concentrate............................................................................................... 40 Figure 28: Schedule of lead production in concentrate .............................................................................................. 40 Figure 29: Schedule of gold production in concentrate .............................................................................................. 41 Figure 30: Schedule of silver production in concentrate ............................................................................................ 41 Figure 31: Schedule of monthly tailings residues over LOM ....................................................................................... 43 Figure 32: New TSF2 downstream of Main Dam ........................................................................................................ 45 Figure 33: Histogram of NPV estimates – Variant 1 .................................................................................................... 58 Figure 34: Histogram of NPV estimates – Variant 2 .................................................................................................... 59 Figure 35: Stochastic DCFs profile ............................................................................................................................... 61 Figure 36: Mining Project Risk profile over LOM ........................................................................................................ 61

Tables Table 1: Hellyer Tailings Storage Facility, Mineral Resource estimate ..................................................................... IV Table 2: Hellyer Tailings Storage Facility, Mineral Resource estimate – metal tonnes and ounces, gross total

only, all attributable to NQM ...................................................................................................................... IV Table 3: Planned parameters of concentrate production........................................................................................ VII Table 4: Hellyer Tailings Project operating expenditure ........................................................................................... IX Table 5: Summary of the DCF model results based on Ore Reserve estimate ......................................................... IX Table 6: Summary of the DCF model results with inclusion of the Shale Pit resource .............................................. X Table 7: Hellyer TSF Ore Reserve estimate* (6 September 2017) ............................................................................ XI Table 8: Hellyer TSF non-JORC Reserve deposited material (6 September 2017) .................................................... XI

XVII



Table 9: Ore Reserve Risk estimates – Variant 1 .................................................................................................... XIII Table 10: Ore Reserve Risk estimates – Variant 2 .................................................................................................... XIII Table 11: Summary table of assets .............................................................................................................................. 7 Table 12: Metallurgical accounting, Hellyer .............................................................................................................. 10 Table 13: Hellyer Tailings Storage Facility, Mineral Resource estimate .................................................................... 14 Table 14: Hellyer Tailings Storage Facility, Mineral Resource estimate – metal tonnes and ounces (gross total

only, all attributable to NQM) .................................................................................................................... 14 Table 15: Parameters of mine design of all dams ...................................................................................................... 21 Table 16: TSF discharge emission limits ..................................................................................................................... 27 Table 17: Planned parameters of concentrate production........................................................................................ 39 Table 18: Schedule of residues storage ..................................................................................................................... 43 Table 19: Residue storage capacities ......................................................................................................................... 44 Table 20: Hellyer Tailings Project – capital expenditure ............................................................................................ 46 Table 21: Hellyer Tailings Project – operating expenditure ....................................................................................... 47 Table 22: Forecasting of Zn, Pb, Au and Ag prices ..................................................................................................... 48 Table 23: Summary of the DCF model results for Ore Reserve estimate .................................................................. 50 Table 24: Summary of the DCF model results including Shale Pit Inferred Mineral Resource .................................. 50 Table 25: DCF model of Hellyer Project based on Ore Reserve estimate .................................................................. 51 Table 26: DCF model of Hellyer Project with inclusion of Inferred Mineral Resource............................................... 52 Table 27: Hellyer TSF Ore Reserve estimate* (6 September 2017) ........................................................................... 54 Table 28: Hellyer Non-JORC Reserves (31 August 2017) of Fossey tails in the Western Arm and Shale Pit tails ...... 54 Table 29: Hellyer Ore Reserve Datamine files ........................................................................................................... 55 Table 30: Hellyer Ore Reserve Grid Datamine files .................................................................................................... 55 Table 31: Statistical input parameters of Ore Reserve Risk assessment ................................................................... 58 Table 32: Input parameters of variogram modelling ................................................................................................. 58 Table 33: Ore Reserve Risk estimates – Variant 1 ..................................................................................................... 59 Table 34: Ore Reserve Risk estimates – Variant 2 ..................................................................................................... 59 Table 35: Results of Mining Project Risk assessment ................................................................................................ 62

Appendices Appendix 1: JORC Code (2012 Edition) – Table 1

1



2 Introduction

2.1 Context, Scope and Terms of Reference

CSA Global Pty Ltd (CSA Global) was engaged by NQ Mineral Plc (NQM) to compile a Competent Persons

Report (CPR) setting out, in addition to required background information, the Mineral Resource estimate

and Ore Reserve estimate relating to the Hellyer Tailings Retreatment Project (Hellyer or Hellyer Project),

located in Tasmania, Australia.

This report was prepared exclusively for NQM by CSA Global. The quality of information, conclusions and

estimates contained in this report are consistent with the level of the work carried out by CSA Global to

date on the assignment, in accordance with the assignment specification agreed between CSA Global and

NQM.

In preparing this report, CSA Global has:

1. Adhered to the JORC Code2.

2. Relied on the accuracy and completeness of the data provided to it by NQM, and that NQM made

CSA Global aware of all material information in relation to the project asset.

3. Relied on the accuracy and completeness of the data, information and reporting provided to it by

AusGEMCO Pty Ltd (AusGEMCO) where this relates to the Ore Reserve Study, having completed

appropriate gap analysis and peer review of this work prior to inclusion of this study in this report.

4. Relied on NQM’s representation that it will hold adequate security of tenure for assessment of the

project to proceed.

5. Has independently verified the data used to prepare this report and concludes that the data provide

reasonable grounds for CSA Global’s conclusions reached in this report.

6. Required that NQM provide an indemnity to the effect that NQM would compensate CSA Global in

respect of preparing the report against any and all losses, claims, damages and liabilities to which

CSA Global or its Associates may become subject under any applicable law or otherwise arising from

the preparation of the report to the extent that such loss, claim, damage or liability is a direct result

of NQM or any of its directors or officers knowingly providing CSA Global with any false or misleading

information, or NQ Minerals, or its directors or officers knowingly withholding material information.

7. Required an indemnity that NQM would compensate CSA Global for any liability relating to any

consequential extension of workload through queries, questions or public hearings arising from the

reports.

2.2 Compliance with the JORC Code

This report has been prepared in accordance with the JORC Code (2012 Edition), which is binding upon

Members of the Australian Institute of Geoscientists (AIG) and the Australasian Institute of Mining and

Metallurgy (AusIMM) for the reporting of Mineral Resources and Ore Reserves.

2.3 Principal Sources of Information

CSA Global has based its review based on information made available to the authors by NQM, along with

technical reports prepared by consultants, previous tenement holders and other relevant published and

unpublished data.

2 Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves. The JORC Code, 2012 Edition. Prepared by: The Joint Ore Reserves Committee of The Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia (JORC).

2



CSA Global has relied upon discussions with NQM’s management, as well as previous company and

consultants reports for information contained within this assessment.

CSA Global visited the Project on 17 July 2017, reviewing the project setting, infrastructure and discussing

the Project’s history with the Resident Manager, Mr Bob Quilliam.

This report has been based upon information available up to and including 7 September 2017.

CSA Global has endeavoured, by making all reasonable enquiries, to confirm the authenticity, accuracy,

and completeness of the technical data upon which this report is based. Unless otherwise stated,

information and data contained in this technical report or used in its preparation has been provided by

NQM in the form of documentation, and from data obtained and observations made during the site visits.

NQM was provided a final draft of this report and requested to identify any material errors or omissions

prior to its lodgement.

Descriptions of the mineral tenure; tenure agreements, encumbrances and environmental liabilities were

provided to CSA Global by NQM. NQM has warranted to CSA Global that the information provided for

preparation of this report correctly represents all material information relevant to the mineral assets. Full

details on the tenement is provided in Section 3.

2.4 Authors of the Report – Qualifications, Experience and Competence

CSA Global is a privately owned, mining industry consulting company headquartered in Perth, Western

Australia, with regional offices throughout the world. CSA Global provides geological, resource, mining,

management and corporate consulting services to the international resources sector and has done so for

more than 30 years.

This report has been prepared by a team of consultants sourced from CSA Global’s Perth and Horsham

(UK) offices. The individuals who have provided input to the report have extensive experience in the

mining industry and are members in good standing of appropriate professional institutions. The

Consultants preparing this report are specialists in the fields of geology, exploration and Mineral Resource

estimation, in particular relating to tailings dam deposits.

The following individuals, by virtue of their education, experience and professional association, are

considered Competent Persons, as defined in the JORC Code (2012), for this report. The Competent

Persons’ individual areas of responsibility are presented below:

• Coordinating author – David Williams (Principal Resource Geologist) is responsible for Sections 1.1 to

1.4 and 2 to 5 of the report.

• Contributing author – Peter Cranfield (Associate Principal Mining Engineer) is responsible for Sections

1.5 to 1.8 and 6 to 12 of the report.

• Peer reviewer – Mr Galen White (Director – Europe and Africa) has peer reviewed all sections of this

report.

David Williams, CSA Global Principal Resource Geologist, is a resource geologist with 25 years’ experience

in Mineral Resource estimation and mine geology. He has worked on a variety of commodities including

gold, iron ore, uranium, nickel laterite, graphite and base metals in Australia, Indonesia and Namibia.

David is a Competent Person for JORC reporting of Mineral Resources and is also a Qualified Person for

Canadian NI 43-101 reporting for gold and base metals. David is able to provide advice on due diligence

studies, JORC and NI 43-101 reporting and Independent Geologist Reports.

Peter Cranfield, Associate Principal Mining Engineer is an engineer with over 40 years’ experience in the

mining industry throughout South-East Asia and Australia, specialising in alluvial mining and tailings

projects. He has 10 years’ experience operating and managing narrow vein underground mines and 30

3



years’ experience in alluvial and tailings mining and processing, covering exploration, development and

production for tin, wolfram, gold, mineral sands and diamonds. Alluvial resource evaluation, feasibility

studies, planning and development of new mines. Exploration and evaluation through drilling, bulk

sampling, and pilot plants to generate sufficient data for mine planning.

Galen White, CSA Global (UK) Director and Principal Geologist, is a geologist with over 20 years’

experience in the mining industry, with the last 12 years spent in consulting. He has experience in mineral

exploration, Mineral Resource estimation and mine geology in a variety of geological settings and in

relation to a variety of mineral commodities including gold, silver, base metals, uranium and iron ore

across Europe, Africa, Australia, Asia and the Americas. He has completed project reviews, Mineral

Resource estimates and audits, due diligence reviews, project management and formal stock exchange

reporting. He is a Competent Person as defined in the JORC Code (2012) and a Qualified Person in relation

to CIM compliance and NI 43-101 Technical Reporting.

2.5 Independence

Neither CSA Global, nor the authors of this report, has or has had previously, any material interest in

NQM, or the mineral properties in which NQM has an interest. CSA Global’s relationship with NQM is

solely one of professional association between client and independent consultant.

CSA Global is an independent geological consultancy. Fees are being charged to NQM at a commercial

rate for the preparation of this report, the payment of which is not contingent upon the conclusions of

the report.

No member or employee of CSA Global is, or is intended to be, a director, officer or other direct employee

of NQM. No member or employee of CSA Global has, or has had, any shareholding in NQM.

CSA Global has completed technical work previously for Bass Metals Pty Ltd (Bass Metals), the previous

owners of the Project, and as such disclosure is required including declaration of any previous reports that

the Practitioner has prepared relating to the Mineral Assets being assessed. To meet this requirement,

the reader is advised that CSA Global completed the following work for Bass Metals:

• CSA Global prepared the Mineral Resource estimate for the Hellyer Tailings Retreatment Project in

2010.

• CSA Global prepared a Mining Study for the Project in 2010.

This report was not influenced by NQM, and reflects CSA Global’s objective critical analysis and

professional judgement.

2.6 Declarations

This report has been prepared by CSA Global at the request of, and for the sole benefit of NQM. Its

purpose is to provide a technical assessment of the Hellyer Tailings Retreatment Project. It is not intended

to serve any purpose beyond that stated and should not be relied upon for any other purpose.

The statements and opinions contained in this report are given in good faith and in the belief that they

are not false or misleading. The conclusions are based on the reference date of 31 July 2017 and could

alter over time depending on exploration results, mineral prices and other relevant market factors.

2.7 Results are Estimates and Subject to Change

The interpretations and conclusions reached in this report are based on current scientific understanding

and the best evidence available to the authors at the time of writing. It is the nature of all scientific

conclusions that they are founded on an assessment of probabilities and, however high these probabilities

might be, they make no claim for absolute certainty.

4



The ability to achieve forward-looking production and economic targets is dependent on numerous

factors that are beyond CSA Global’s control and that CSA Global cannot anticipate. These factors include,

but are not limited, to changes to site-specific mining and geological conditions, management and

personnel capabilities, availability of funding to properly operate and capitalise the operation, variations

in cost elements and market conditions, developing and operating the mine in an efficient manner,

unforeseen changes in legislation and new industry developments. Any of these factors may substantially

alter the performance of any mining operation.

2.8 About this Report

This report discusses the Mineral Resource estimate of the tailings material impounded in the Hellyer

Tailings Retreatment Project. The geology and mineralisation are discussed, as well as the drilling activities

completed which support the Mineral Resource estimate. The Project has been part of the Hellyer and

Fossey polymetallic underground mine, which operated from 1989 to 2000 and 2010 to 2102 respectively.

Management of the tailings facility has continued to the present day. No discussion is provided concerning

the Hellyer or Fossey underground orebodies, apart from a discussion on the grades of ore feed into the

Hellyer mill, and tailings subsequently discharged.

2.9 Competent Person Statement

The information that relates to Mineral Resources is based on information compiled by Mr David Williams,

a Competent Person, who is a Member of the Australasian Institute of Mining and Metallurgy. Mr Williams

is employed by CSA Global Pty Ltd, an independent consulting company. Mr Williams has sufficient

experience, which is relevant to the style of mineralisation and type of deposit under consideration, and

to the activity he is undertaking, to qualify as a Competent Person as defined in the 2012 Edition of the

“Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves”.

Mr Williams consents to the inclusion of the matters relating to Mineral Resources, based on his

information in the form and context in which it appears.

The information that relates to Ore Reserves is based on information received and reviewed by Mr Peter

Cranfield, a Competent Person, who is a Member of the Institute of Materials, Minerals and Mining.

Mr Cranfield is an Associate of CSA Global Pty Ltd, an independent consulting company. Mr Cranfield has

sufficient experience, which is relevant to the style of mineralisation and type of deposit under

consideration, and to the activity he is undertaking, to qualify as a Competent Person as defined in the

2012 Edition of the “Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore

Reserves”. Mr Cranfield consents to the inclusion of the matters relating to Ore Reserves, based on his

information in the form and context in which it appears.

5



3 Property Description and Tenure

3.1 Property Location and Climate

Hellyer is located in north-western Tasmania 80 km south of Burnie, just off the main highway between

Burnie and Queenstown and four hours’ drive northwest of the capital, Hobart. The area surrounding the

project is mainly forest reserve with farmland. There has been extensive historical and recent mining

activity in the area, including the Hellyer base metals mine located adjacent to the Hellyer. The project

location is presented in Figure 2.

This part of Tasmania lies in a high rainfall area with annual average precipitation of 2,180 mm (Bureau of

Meteorology, Waratah) falling all year although higher falls, and snowfall, occur over winter months.

Average daytime temperatures range from 7°C (winter) to 18°C (summer).

Figure 2: Project location

6



3.2 Property Description

The property under study is a tailings dam with a capacity of approximately 10 Mt at an in-situ density of

approximately 1.9 t/m3. The tailings are the waste product from the Hellyer mill and comprise fine sands

containing significant values of zinc, lead, copper, gold and silver, a portion of which can possibly be

recovered by further milling and retreatment by flotation and other means.

The depth of the tailings varies from 1 m below water level to approximately 20 m below water level at

the deepest point beside the dam wall. There is an environmental constraint to maintain a minimum water

cover of 1 m over the surface of the tailings. The tailings contain sulphides susceptible to oxidation and

the formation of acids if exposed to the atmosphere. This imposes a constraint on the mining method,

resulting in the choice of a dredge as the most suitable means of mining. A Seabird III 300 kW Electric

Drive Cutter Suction Dredge is currently moored on the tails dam and has been used in the past to mine

approximately 2 Mt of the tails material for retreatment in the Hellyer mill. Figure 3 presents a map

showing the site layout.

Figure 3: Site layout

7



3.3 Tenure

NQM holds a consolidated granted mining leases CML103M/1987 over the area including the tailings dam

and the processing plant. This lease covers all the areas required for the proposed tailings reclaim

operations. The mining lease is held by Hellyer Gold Mines Pty Ltd (HGM), a wholly owned subsidiary of

NQM. The lease was granted on 24 February 1988 and was renewed and extended in June 2009 until

30 June 2020. The lease can then be further extended, for as long as the mine is operated, as advised by

Mineral Resources Tasmania. The annual permit has been paid in full until 15 February 2018.

Tenure details are presented in Table 11, and the tenement outline is shown in Figure 2 and Figure 3.

In relation to further extension of the lease beyond June 2020, NQM has been in direct communication

with the Inspection Team at Mineral Resources Tasmania. The Inspection Team undertakes mining lease

renewal assessments whose manager reports directly to the Director of Mines (Tasmania), who then

makes recommendations to the Minister for Mines. NQM was provided with the following email advice

on 30 August 2016 from the Manager, Inspection Team, as follows:

• Under Section 97(2) of the MRDA, the Minister must grant the application if satisfied that:

o The lessee has submitted a mining plan for the renewal period;

o The lessee has complied with the conditions of the lease and provisions of this Act;

o A failure to comply with any conditions of a lease was exempted under section 86; and

o The lessee has provided a security deposit.

• The legislation is worded in favour of ongoing security of tenure as long as a lessee provides the

information, security deposit and has demonstrated compliance.

The proposed Hellyer Project requires government approvals to operate and, as such, a permit currently

exists for the reprocessing of tailings (2 Mt/a) at Hellyer (PCE No. 7386). In October 2017, following a

request by HGM the Environmental Protection Authority (EPA) of Tasmania approved the 2017

Environmental Management Plan (EMP) for implementation of the Hellyer Project, in accordance with the

Condition G7 of Permit Conditions – Environmental (PCE) No. 7386, as contained in Permit No. DA

138/2006. As a result, all permits and approvals required for the tailings reclaim and retreatment

operation are either in place or can be obtained by NQM within the necessary timeframe.

Table 11: Summary table of assets

Asset Holder Interest

(%) Status

Lease expiry date

Lease area

Comments

Australia, Hellyer Gold Mine

Hellyer Gold Mines Pty Ltd

100 Production 30 June

2020 16.95 km2

Final permitting for retreatment complete.

8



4 Geology and Project History

4.1 Geology

The Hellyer deposit is a volcanic hosted polymetallic massive sulphide deposit located within the Mount

Read volcanic arc of western Tasmania. This region also hosts similar deposits such as Hercules, Que River,

Roseberry and Mount Lyell. Mineralisation is sulphide hosted and comprised predominantly of pyrite,

with lesser sphalerite, galena and arsenopyrite. The pyrite must be carefully managed when exposed to

air to prevent its oxidation and potential to form acid mine drainage issues. The economic metals mined

at Hellyer were lead, zinc, copper, gold and silver. The Hellyer deposit was mined by underground

methods during the period 1989 to 2000. The Fossey deposit extends down plunge from the Hellyer

deposit and was mined by Bass Metals between 2010 and 2012.

Tailings from the mill were deposited in a depression approximately 1 km to the west of the Hellyer mill.

The tails were inundated with water to prevent oxidation of the sulphide species present in the tails. The

tails are typically stratified, having been deposited from discharge pipes set around the dam. The

sedimentary stratification is also reflected in a mild gradient in grades of Zn and Pb, with grades typically

higher in the deeper parts of the tails and decreasing as the depth of the tails sediments shallow. The tails

sediments are unconsolidated.

4.2 Project History

4.2.1 Hellyer Base Metals Mine

The Hellyer base metals mine was operated by Aberfoyle Resources Ltd (Aberfoyle), then Western Metals

Ltd (Western Metals) between 1989 and 2000. Tails were deposited in a storage facility constructed to

the west of the mine site, making use of a broad valley. The dam wall is of earth construction.

4.2.2 Metallurgical Accounting

Aberfoyle and Western Metals produced a metallurgical account of the materials that were treated

through their operation during the life of the operation. Table 12 presents the inventory of tails deposited

in the tailings dam between 1989 and 2000. Table 12 shows the tails deposited in the earlier years

contained higher grades of Cu, Zn, Pb, Ag and Au than in later years. These higher-grade tails are located

in the deeper parts of the tailings dam, as discussed in Section 5.2.

4.2.3 Retreatment of Tails

The tailings dam was partially dredged between November 2006 and August 2008, for a total of 2.026 Mt

of tails. These tails were treated at the Hellyer mill. The tails were sourced from the following areas of the

tails dam, which are depicted in Figure 3:

• Western Arm

• Eastern Arm

• Main Dam.

The retreated tails were redeposited into the following locations:

• Shale Pit (1.31 Mt)

• Western Arm Dam (0.605 Mt).

No tails were deposited into the Eastern Arm (Mill Creek) because it was set aside as an environmental

remediation dam to submerge legacy beached tails mainly emanating from operations during the 1990s

(Bolger, 2007).

9



4.2.4 Fossey Base Metals Mine

Bass Metals mined the Fossey underground deposit between 2010 and 2012. The tails were discharged

into the Western Arm of the tailings dam. Table 12 presents the inventory of tails deposited in the tailings

dam, from ore sourced from Fossey, between 2010 and 2012. No mining production statistics for mill ore

feed were provided to CSA Global.

10



Table 12: Metallurgical accounting, Hellyer

Year Mill Ore Feed Tails

Tonnes % Cu % Zn % Pb g/t Ag g/t Au Tonnes % Cu % Zn % Pb g/t Ag g/t Au

1989 582,022 0.38 13.05 7.08 150.00 2.67 440,503 0.24 5.90 3.66 92.18 2.89

1990 995,087 0.36 13.12 7.24 166.00 2.63 744,267 0.21 5.18 4.18 105.62 2.33

1991 1,248,881 0.37 13.78 7.39 178.00 2.77 878,240 0.16 3.39 3.83 101.70 3.13

1992 1,355,658 0.32 12.55 7.08 170.00 2.80 1,001,451 0.19 3.16 3.93 119.69 3.13

1993 779,256 0.31 12.33 6.70 157.00 2.41 576,366 0.15 3.23 3.50 99.56 2.71

1994 1,305,946 0.32 13.07 6.55 165.00 2.43 942,371 0.16 2.97 3.67 102.76 2.76

1995 1,322,518 0.36 13.14 6.38 155.00 2.30 945,779 0.16 2.45 3.48 94.07 2.69

1996 1,335,749 0.29 12.57 5.66 148.00 2.23 978,643 0.12 2.54 2.86 85.56 2.54

1997 1,392,528 0.32 11.94 6.01 155.00 2.47 1,017,928 0.15 2.17 2.57 85.68 2.80

1998 1,436,210 0.29 10.60 4.91 130.00 2.03 1,105,009 0.14 2.08 2.01 71.00 2.23

1999 1,491,888 0.32 10.46 5.14 114.00 1.96 1,150,603 0.16 1.99 2.06 66.92 2.17

2000 1,368,980 0.27 8.63 4.57 106.00 1.89 1,101,027 0.13 1.48 1.71 58.39 1.98

Total 14,615,210 0.32 11.97 6.11 148.03 2.35 10,882,187 0.16 2.80 3.00 88.42 2.58

2010-2012 447,798 0.03 1.84 1.37 35.38 1.80

Note: Between the years 1989 and 1992, the tonnages were calculated to the second week of November; from 1993 to mine closure the tonnages were calculated over the financial year. No production figures available for Fossey ore delivered to mill (2010 to 2012).

11



5 Mineral Resource Estimate

5.1 Exploration History

Drill sampling of the tails dam was undertaken during 1998 and 2000. The drilling method used was

Vibracoring, which uses a core tube that is inserted into soft sediments. A vibrating mechanism

(Vibrahead) assists this process. When the insertion is completed, the Vibracorer is switched off and the

core tube is recovered. No information is available for the 1998 or 2000 drilling campaigns, with respect

to detailing the type of Vibracorer, core recoveries or geological logging. Figure 4 shows the distribution

by year of drilling, of drillholes across the tails dam. Figure 4 also shows that drilling was concentrated

through the centre of the dam and into the Eastern Arm, with relatively few holes testing the western

quarter of the dam. None of the drill collars were preserved following completion of sample coring, due

to the water covering the tails sediments.

Tails deposition from the Hellyer processing plant continued after completion of the 1998 drilling

campaign, therefore the 1998 collars are generally located below the 1998 depositional surface. As a

result, there are samples from the 2000 series located at a higher elevation than the collars of the 1998

data.

Figure 4: Collar plan for Hellyer TSF

Note: Also shown are tails limits excluding the Eastern Arm, and dam wall. Collars colour coded by year of drilling.

Drill collars were surveyed using a handheld global positioning system (GPS) device at the drill collar during

drilling. The dam water level was surveyed by a licensed surveyor and the elevations were used as a proxy

for the collar elevations.

Drill samples were assayed at AMMTEC Research Laboratory (Burnie) using X-ray fluorescence (XRF) to

analyse most elements, while Au was analysed by fire assay. No records of quality assurance/quality

control (QAQC) testing or results are known to exist from the drilling programs.

12



5.2 Mineral Resource Estimate

The Hellyer Mineral Resource is based upon a 3D block model encapsulating key features defining the

dam construction and a top of tails surface. Blocks were interpolated with assayed sample grades

including Cu (%), Zn (%), Pb (%), Ag (g/t) and Au (g/t). An average density value of 1.93 t/m3 was applied

to each block so that a tonnage could be calculated from the known block volumes. The block model was

classified in accordance with the JORC Code and the tonnes and grades in the Mineral Resource model

were reported.

A wireframe solid of the tailings dam wall was initially prepared for a Mineral Resource estimate prepared

in 1998 and was used for to model the Mineral Resource discussed here. A topographic digital terrain

model (DTM) of the original surface of the tailings dam prior to discharge of tails was constructed from

surveyed contours. The surface of the unconsolidated tails sediments was estimated from drill depths

recorded during Vibracore drilling.

A block model with parent cell sizes 25 mE x 25 mN x 5 mRL was constructed, with sub-celling to 5 mE x

5 mN x 2.5 mRL, sufficient to provide resolution at the wireframe domain boundaries. The block model

was split in the vertical direction with 5 m parent cell resolution, in cognisance of the observed vertical

stratification of the assay grades for the key metal elements being modelled.

Drill samples were composited to 8 m lengths and statistically analysed. Metal grades were noted to

increase with depth, an example of which is presented in Figure 5, showing the Zn (%) grades. This vertical

zonation is due to the depositional history of the tails, as presented in Table 12. Similar vertical zonation

was observed for Cu and Pb, with less discrete zonation observed for Ag and Au.

Figure 5: Zn % grade variation with depth, by year of drilling

Variograms were modelled for Zn, Pb, Cu, Au and Ag from the composited drill samples. Low nugget

effects were modelled with short ranges of between 60 m and 80 m. Long ranges were modelled to

between 200 m and 300 m.

626

628

630

632

634

636

638

640

642

644

646

648

0 1 2 3 4 5 6 7

RL

(m)

Metal (%)

Zinc Composited (8m) data by depth

ZN 1998

ZN 2000

13



Grade was interpolated into all the blocks using ordinary kriging. A primary search ellipse of 100 m along

strike by 100 m across strike by 5 m vertical was used for the first pass estimation run. The search radii

were increased to 150 m for the second estimation run, then 1,000 m for a third run if necessary.

The minimum number of samples used within the search was two, with the maximum being six. These

parameters reflect low nugget effect, and the broad compositing interval (8 m) that was used, wherein

many samples were combined into composited samples prior to the grade estimation. A maximum of two

samples per drillhole per block estimate was applied, with discretisation of 3 x 3 x 1 used. No octant based

search was used. The samples were not top-cut prior to grade interpolation.

Samples from the 2000 Vibracore drilling campaign were retained at the AMMTEC Laboratory (Burnie)

where they performed 39 bulk density determinations on these samples, in preparation for an historical

Mineral Resource estimate in 2006 (Tyrell, 2006). The results ranged from 1.30 to 2.63 t/m3 and averaged

1.93 t/m3. Tyrell (2006) notes a discrepancy between reported tonnages in the 2006 Mineral Resource

estimate and the mill production figures, as presented in Table 12. CSA Global used the same model files

for the Mineral Resource reported in this report as were used by Tyrell (2006), and the discrepancy is

noted. Tyrell (2006) suggests the reason for the discrepancy may be due to the manner in which the

density of tailings originally discharged was calculated. CSA Global also note the variability of sample

grades in the vertical profile, with higher grades of base metals (particularly Pb and Zn) expected to impart

a higher density on the local volume of tailings.

The interpolated block grades were validated by:

• A visual comparison of block grades with sample grades, in cross section.

• Swath plots which plot the trends of the grade variable (block grade and sample grade) along

northing, easting and vertical directions.

• Comparing the mean grade of the blocks with the mean grade of the samples.

The Mineral Resource is classified in accordance with the JORC Code (2012) as a combination of Measured,

Indicated and Inferred Mineral Resources, and is presented in Table 13. Table 14 presents the metal totals,

as calculated from Table 13 (gross totals only are presented). Classification was based upon drill spacing

and considerations regarding reasonable prospects for the eventual economic extraction of the tails. The

area where drillholes are spaced at 50 m x 50 m intervals was classified as Measured Mineral Resources.

The remainder of the tails in the Main Dam were classified as Indicated Mineral Resources. The tails

contained within the Shale Pit are classified as Inferred Mineral Resources. These block grades are

estimated based upon a back calculation of assays from reprocessed tails using calculated recoveries. No

sampling has been carried out to date of the Shale Pit tails.

It is the Competent Person’s opinion that there are reasonable prospects for eventual economic extraction

of the Hellyer tails due to the following reasons:

• The Mineral Resource is at a shallow depth and amenable to dredging. An operational dredge is

moored on the tailings dam surface.

• The Project is located within 1 km of an ore processing facility, which was used to process the ore

from the Hellyer underground mine during the period 1989 to 2000. The processed tails were

discharged into the tailings dam.

• The Project is located near the north coast of Tasmania, with deep water ports which can transport

the concentrates to domestic or international customers. A rail line was constructed at the time of

construction of the Hellyer mine, to allow rail transportation of concentrate to Burnie Port.

• The Project is currently under care and maintenance. Power and water are currently supplied to the

Project.

14



• The metal sulphide species contained within the dam are amenable to further processing and

recoveries of 45% to 68% for Zn have been demonstrated through recent metallurgical testwork.

• A mining lease is current over the property.

Table 13: Hellyer Tailings Storage Facility, Mineral Resource estimate

JORC classification

Gross Net attributable

Operator Tonnage (Mt)

Zn %

Pb %

Ag g/t

Au g/t

Cu %

Tonnage (Mt)

Zn %

Pb %

Ag g/t

Au g/t

Cu %

Measured 2.05 3.31 3.35 94 2.63 0.2 2.05 3.31 3.35 94 2.63 0.2

NQM Indicated 5.99 2.29 2.95 93 2.55 0.18 5.99 2.29 2.95 93 2.55 0.18

Inferred 1.21 1.00 2.60 86 2.57 0.19 1.21 1.00 2.60 86 2.57 0.19

Total 9.25 2.35 2.99 92 2.57 0.19 9.25 2.35 2.99 92 2.57 0.19

Note: No lower cut-off reporting grade has been applied. Differences may occur due to rounding. (Datamine model: hel717md.dm).

Table 14: Hellyer Tailings Storage Facility, Mineral Resource estimate – metal tonnes and ounces (gross total only, all attributable to NQM)

JORC classification Gross

Tonnage Zn (t) Pb (t) Ag (oz) Au (oz) Cu (t)

Measured 2,050,000 67,900 68,700 6,195,400 173,300 4,100

Indicated 5,990,000 137,200 176,700 17,910,200 491,100 10,800

Inferred 1,210,000 12,100 31,500 3,345,600 100,000 2,300

Total 9,250,000 217,400 276,600 27,360,300 764,300 17,600

Note: Metal tonnages and ounces rounded from calculated values.

15



6 Previous Mining Studies

During the period 2006 to 2008, Polymetals Group (Polymetals) successfully dredged and processed

approximately 2 Mt of tailings at Hellyer, targeting the Zn value through the sale of a bulk Zn/Pb

concentrate. Operations ceased in September 2008 following a severe drop in zinc prices associated with

the global financial crisis. Polymetals did not focus on precious metal extraction.

In 2010, CSA Global prepared a Resource Report for the Hellyer tailings as part of a Resource Estimate and

Mining Study (CSA Global, 2010). This report was used by Como Engineers Pty Ltd (Como) as a part of a

broader project evaluation for the then owners of Hellyer, Bass Metals. This study included consideration

of dredging technology, mine sequencing, production scheduling, mineral processing and economic

evaluation. The mining study focused on the extraction of the tailings contained in the Main Dam.

In 2013, Ivy Resources Pty Ltd (Ivy Resources) acquired Hellyer and in November 2013 completed a

Feasibility Study of the Hellyer Tailings Retreatment Project, which presented results including marketing

studies, geology, Mineral Resource estimation, mining, mineral processing, infrastructure, permitting, and

financial analysis (Ivy Resources, 2013). The mining technology considered was dredging. The mine

sequence and schedule were adopted from the CSA Global Resource Estimate and Mining Study.

This Feasibility Study concluded that tailings recovery and treatment using sequential bulk flotation and

pyrite flotation with pyrite concentrate treated using the Albion process (partial and complete oxidation)

and subsequent cyanide leaching would result in a 26% internal rate of return (IRR). The capital cost was

estimated to be A$232 million and the operating cost approximately A$100/dmt. Ivy Resources did not

proceed with the project.

As part of the Ivy Resources Feasibility Study, Como completed a “Definitive Feasibility Study” (DFS) on

the Hellyer TSF Project in 2013 with a specific scope related to direct cyanide leach and leaching of the

Albion process product (Como, 2013). The study addressed capital and operating expenditure, project

management and risk analysis.

In 2015, Como updated the prior DFS (Como, 2016) and included costs for various feed rates to the cyanide

leach circuit and subsequent processes.

In 2016, Ausenco produced a report on the tailings treatment restart cost estimate following a request by

NQM based on a refined processing flowsheet for sequential flotation of the tailings. The report reviewed

cost and capital estimates and made recommendations for improvement in the confidence of the cost

estimate for refurbishment of the existing dredge and processing plant and the restart of the tailings

retreatment. The Ausenco report noted that “potentially saleable Pb/Ag, Zn and pyrite concentrates can

be produced in the laboratory. The recovery of the Au and Ag to dore is technically achievable, but

economically challenged due to the ore’s refractory nature. Oxidation of at least 50% of the pyrite is

required to achieve high Au recoveries with subsequent cyanide leaching” (Ausenco, 2016).

In November 2016, Pitt & Sherry was requested by NQM to prepare a report assessing the status of

metallurgical testwork in connection with the sequential flotation of the Hellyer tailings (Pitt & Sherry,

2016). In the report, Pitt & Sherry analysed the results of testwork conducted by the ALS Laboratories

(ALS) in Burnie, Tasmania (approximately 40 batch tests and three locked cycle). The results indicated the

opportunity to produce the following three concentrates: (1) a combined Pb/Ag concentrate grading 37%

Pb and maximising silver content; (2) a Zn concentrate grading 45% Zn; and (3) a pyrite concentrate

containing 46% S and potentially payable Au and Ag. The testwork conducted identified that a combined

Pb/Ag concentrate could be selectively floated separately from a subsequent Zn concentrate. In both

cases, the rougher concentrate was reground prior to cleaning. Both these base metal concentrates were

of a saleable quality.

16



A relevant study was also conducted by Commodity and Mining Insight Ltd (CM Insight, 2016) that

included preliminary economic modelling of the potential reprocessing of the Hellyer tailings. CM Insight

performed this modelling based on assumptions drawn from previous reports, the results of the ALS

laboratory testwork and information provided by Hellyer management.

The above cited dredging operations, studies and reports are evidence about the progressive trend of

searching for an efficient technological solution of the exploitation of the Hellyer tailings with a

commercial effect. They are a sufficient base for undertaking an Ore Reserve estimation of the Hellyer

Tailings Retreatment Project in accordance with the requirements of the JORC Code (2012).

17



7 Current Mining Study

In 2017, AusGEMCO was engaged by NQM to complete an Ore Reserve estimate over the Hellyer Tailings

Retreatment Project. This study is documented in a report titled “Hellyer Tailings Retreatment Project –

Tasmania, Ore Reserve Estimate” dated September 2017. This report, along with supporting data and

information compiled during the study was made available to CSA Global for review and CSA Global

completed a gap analysis and peer review of this study prior to inclusion of the study in this report as

current material information.

A summary of this study is present in the body of this CPR, referenced as appropriate and augmented by

CSA Global independent comment as appropriate.

7.1 Mine Design and Planning

The following section has been summarised from the AusGEMCO technical report (AusGEMCO, 2017),

augmented by CSA Global comments in bold, where appropriate.

7.1.1 Mine Design based on Block Profit Modelling

The mine design for Hellyer Tailings Dam has been developed using the approach of modelling the profit

from the exploitation of each block of the Mineral Resource block model (CSA Global, 2017) by assessment

of the Net Smelter Return of selling concentrates as final products of the mine production. For the purpose

of this study three types of concentrate are planned as final products: lead concentrate, zinc concentrate

and pyrite/gold/silver concentrate.

The Block Profit Modelling is organised for each impoundment area comprising the Hellyer TSF. The

economic inputs and mining and processing assumptions are based on those used in the financial model

described in Section 10. Two criteria are used in the block modelling: Block Unit Profit per dry metric tonne

of in-situ tailings (US$/dmt) of a block and Block Profit (US$) per block of the geological model.

The results obtained for the Main Dam which covers the full Mineral Resource quantity from the tails

topography (after the dredging operations had been conducted by Polymetals in the period from

November 2006 until August 2008) to the terrain bottom, are graphically illustrated in Figure 6 and

Figure 7. The average Block Unit Profit estimate is US$51.84/dmt of tails and the standard deviation is

US$5.25/dmt. The results of the Block Profit (US$) vary within a range from US$100,000 up to

US$1,400,000 per block of the geological model. The (X, Y) size of a block is 25 m x 25 m while block height

is variable. The entire area of Main Dam is characterised with positive estimates of both criteria used.

The comparison of both graphs indicates that the prior dredging zone has the maximum estimates of Block

Unit Profit, which is logically correct because the tails with higher grades of useful components are located

in depth and the prior dredging uncovered tailings with higher grades. As to the graph of Block Profit, its

analysis indicates that the dredged area has minimum estimates while the maximum estimates are

inherent for the zone close to the Main Dam wall.

The method of kriging is used in the modelling for interpolation of the Block Unit Profit/Profit results in

Figure 6 and Figure 7.

The results for Block Profit models of Finger Pond Dam, Western Arm Dam and Shale Pit are graphically

illustrated in Figure 8, Figure 9 and Figure 10. Block Unit Profit for the Shale Pit estimates are constant

across the entire dam area due to the utilisation of average estimates for the tails qualities from the

Mineral Resource block model.

18



Figure 6: Block Unit Profit spatial distribution of Main Dam

Figure 7: Block Profit spatial distribution of Main Dam

19



Figure 8: Block Profit spatial distribution of Finger Pond Dam

Figure 9: Block Profit spatial distribution of Western Arm Dam

20



Figure 10: Block Profit spatial distribution of Shale Pit

The proposed dredge path (strip) direction has been optimised for each impoundment area because it has

a strong impact on the NPV of the Project. The results of the Block Profit modelling show a well-expressed

heterogeneity regarding Block Unit Profit which has a strong correlation with the geological heterogeneity

of Hellyer TSF. This evidence makes the problem for optimising the path direction complex. The

AusGEMCO method has been applied in the study for optimising the dredge path direction in each dam

at Hellyer. The method uses economic, geological and technological inputs and is based on a well-

grounded mathematical procedure (Halatchev, 2009).

Figure 11 shows the optimal orientation of the grid of the dredge paths (strips) of the Main Dam, which

has an azimuth of 282° (or an angle of 102° in a Cartesian coordinate system with regards to X-axis).

The parameters of the mine design for each dam are summarised in Table 15. The mining design for the

Main Dam is based upon splitting the tailings body into four vertical benches from the tails top to the tails

bottom. Bench B1 provides a dredging depth of 10.40 m from the current water level in the dam. Bench

B2 has a dredging depth of 5 m tails. In combination these benches provide an overall depth of 15.40 m,

which is achievable by dredging using the removable ladder extension piece.

21



Table 15: Parameters of mine design of all dams

Mining area Bench From RL (m) To RL (m) Depth (m)

Main Dam

B1 649.4* 639 10.4

B2 639 634 5

B3 634 629 5

B4 629 627.5 1.5

Western Arm Dam B1 653.0* 639 14

Finger Pond Dam B1 650.0* 639 11

B2 639 635 4

Shale Pit Dam

B1 681.0* 665 16

B2 665 650 15

B3 650 645 5

* Water level

The width of each dredge path is 30 m, which was used in the mine plan of the prior dredging operations

(CSA Global, 2010). The blocks are designed with a length of 50 m each. The working area in each dam is

designed based on two of constraints: a safety buffer zone from each dam wall and a minimum dredging

depth of 1.5 m. With respect to the Main Dam this working area is refined into two separate areas based

on distance from the Main Dam wall. The first of these areas is based on a 300 m buffer zone (a temporary

restriction) while the second of these areas is based on a 50 m buffer zone.

Mine sequencing has been organised following economic and technological requirements. There is a

pre-dredging period that is used for moving tails from the Finger Pond Dam into the previously dredged

area of the Main Dam. Extraction for production purposes commences with dredging in the Main Dam.

Initially the dredge will take the top bench (B1) to a depth of 10.4 m across the entire working area of the

main dam (excluding buffers zones near existing dam walls). The next bench (B2) will then be dredged

down to a depth of 15.4 m below water level. The mining sequencing has been organised around panels

each of which includes a number of dredge paths and blocks. The panel layout and mining sequence for

benches B1 and B2 of the Main Dam is shown in Figure 11 (the mining panels located in the 300 m and

50 m buffer zones are shown in different colours). The sequencing of panels is based on the analysis of

the Block Unit Profit model of the dam.

After completion of dredging of B1 and B2 of the Main Dam, mining operations move to the Western Arm

Dam, which is designed with a single bench. Upon completion of Western Arm Dam, the dredging will

continue in the Main Dam for taking the last two benches, B3 and B4, after releasing water from the dam.

The last dredging area is the Shale Pit.

22



Figure 11: Mine sequence of Main Dam for benches B1 and B2

CSA Global considers the approach adopted here to be reasonable and physically attainable using the

proposed dredging, pipeline and pumping equipment. The dredge unit has previously demonstrated its

suitability and production capacity in very similar tailings material. The mining block sequence selection

appears logical, and the face widths and block depths to be mined are suited to the equipment. The

mining sequence is not constrained by a closely connected floating processing plant with tailings

deposition immediately behind. The dredge has freedom of movement around the dam and can be

moved from block to block with a minimum of downtime.

7.1.2 Production Scheduling

Dredging is the major mining method for the extraction of Hellyer tailings. Hydraulic mining with slurry

pumps mounted on a platform and water cannon wash are the other mining methods to be used.

Hydraulic mining will be used only for transferring the tailings from the bottom of Western Arm Dam and

Finger Pond Dam to the Main Dam after the release of water from the top two benches in the Main Dam.

Water cannon washing will be used for washing down the tails in the buffer zones adjacent to each dam;

assisting with hydraulic mining; and clearance of tails from the base of the Main Dam after the release of

water.

Production scheduling for the Hellyer Project is organised using the following dredging assumptions:

dredging rate = 3,234 dmt/day or 154 dmt/hour (21 hours per day); resources time of dredging = 355 days

per year (10 days for planned maintenance); number of shifts = two shifts per day (each shift = 12 hours);

23



and working days per week = seven days/week. These assumptions are based on the analysis of the dredge

performance during the 2006 to 2008 dredging period.

The production schedule is organised on a monthly basis and the schedule of run of mine (ROM) quantities

of tailings is shown in Figure 12. The schedule horizon is 102 months (commencing from April 2018 for

scheduling purposes). The associated schedules of the head grades are shown Figure 13 for Zn, Figure 14

for Pb, Figure 15 for Au, and Figure 16 for Ag.

Figure 12: Monthly tails production schedule

24



Figure 13: Monthly Zn head grade distribution

Figure 14: Monthly Pb head grade distribution

25



Figure 15: Monthly Au head grade distribution

Figure 16: Monthly Ag head grade distribution

26



CSA Global considers that the dredge production rates quoted and tabulated above demonstrate a

reasonable approach that can be achieved. Hourly throughput rates have been demonstrated as

regularly achievable and the subsequent calculation of annual and monthly effective working hours are

at industry standards. The grades of material to be mined are predicated on drill results and shown to

be statistically valid, but one cannot expect that the actual monthly rates achieved will follow the

predictions precisely. There will be variations between months, but annual results can be expected to

follow predictions closely.

As noted in the section above, the dredge unit has previously demonstrated its suitability and

production capacity in very similar tailings material. To continue to achieve these rates over the longer

term requires that good maintenance standards be upheld.

27



8 Environmental, Law Compliance and Approved Permits

The following section has been summarised from the AusGEMCO technical report (AusGEMCO, 2017).

8.1 Key Environmental Aspects and their Management

Caloundra Environmental Pty Ltd and HGM have conducted preliminary assessments to identify key

environmental aspects associated with the proposed development. The key environmental issues for the

construction and operation of new TSF (TSF2) are expected to be:

• Surface water quality

• Groundwater quality

• Management of AMD

• Tailings management

• Potential impacts on fauna and flora

• Geotechnical Issues and dam safety

• Rehabilitation and closure issues

• Consultation and communication.

Emission limits have been set by the EPA to manage the point source pollution (the TSF outflow) to

safeguard the protected environmental values (PEVs) for the receiving waters (in this case the Que River).

The environmental authority for the tailings reprocessing operation, PCE 7386, sets emission limits from

the Main Dam in its condition EF2. Table 16 provides these emission limits.

Table 16: TSF discharge emission limits

Parameter Minimum emission limit (mg/L)

pH 8 (pH units)

Sulphate 300

Total lead 0.6

Total zinc 0.8

Total copper 0.2

Total aluminum 0.5

Total arsenic 0.02

Total suspended solids 30

Historically, zinc has been the most difficult limit to meet at the TSF outfall. Since the Aberfoyle operation

ceased production in June 2000 and closed in 2003, sulfidic tailings have been left exposed in the upper

reaches of the eastern arm of the Main Dam and in Mill Creek. The acidity generated from these sulphides

caused ongoing issues with pH in the Main Dam. The pH in the Main Dam is important due to the

relationship between a pH above 8.0 and total zinc concentrations at the TSF outfall. In Figure 17, it can

be seen that zinc concentrations increased after the closure of the Aberfoyle operation. The main reason

for this is that during operations, the average pH in the tailings dam was much higher, this would have led

to high zinc precipitation rates within the TSF.

28



Figure 17: TSF outflow – pH versus total zinc (operations compared with closure)

Figure 18 shows emission limits compared to compliance limits for the parameters shown in Table 16

except pH and Total Zn since 2015. Significant improvements and stability have been noticeable since

2016. HGM’s improved management protocols, such as adding a lime slurry directly into the eastern arm

spillway before it overflows into the main TSF, have been responsible for most of the improvements seen

in Figure 18. Figure 19 shows the pH and Total Zn in the TSF discharge since the beginning of 2016. There

has been a steady decline in Zn concentrations since an autumn flush in May 2016.

As noted in Table 16, environmental licence conditions for the site since 2006 have required a minimum

pH of 8.0 and a maximum Total Zn of 0.8 mg/L at the TSF outfall. Figure 20 shows the relationship

between the TSF outfall discharge pH and the Total Zn from mid-2006 until mid-2012. When the pH is

above 8.0, the discharge is usually compliant. A review of long-term water quality records (Figure 20)

indicates that with good management procedures and the remediation proposed by HGM, this should be

readily achievable going forward.

CSA Global notes that on 17 October 2017, HGM received confirmation from the EPA that Condition G7 of

PCE 7386 have been met through the submission of the final version of the EMP. As such, the EPA has

approved the 2017 EMP for implementation.

29



Figure 18: Emission improvement since 2016

Figure 19: TSF discharge pH and Total Zn since January 2016

30



Figure 20: TSF discharge pH versus Total Zn

8.2 Relevant Legislation, Regulations, Codes and Polices

HGM operates under and complies with a variety of acts, regulations, policies and guidelines that are likely

to be significant for this development, including:

• Aboriginal Relics Act 1975

• Crown Land Act 1976

• Environmental Management and Pollution Control Act 1994 (associated policies and regulations)

• Environment Protection Policy (Noise) 2009

• Environment Protection and Biodiversity Conservation Act 1999

• Forest Practices Act 1985

• Historic Cultural Heritage Act 1995

• Land Use Planning and Approvals Act 1993

• Mineral Resources Development Act 1995

• National Parks and Reserves Management Act 2002

• National Environment Protection Council (Tasmania) Act 1995

• Native Forestry Agreement Act 1980

• Native Title (Tasmania) Act 1994

• Forestry (Rebuilding the Forest Industry) Act 2014

• Resource Management and Planning Appeal Tribunal Act 1993 and associated amendments

• State Policy on Water Quality Management 1997

• Threatened Species Protection Act 1995

• Water Management Act 1999 and associated regulations

• Weed Management Act 1999

• Workplace Health and Safety Act 1995.

31



8.3 Environmental Licences and Permits

HGM holds the consolidated mining lease, CML 103M/87 (Figure 21) and the environmental licences PCE

7386 (tailings mining and reprocessing) and PCE 7759 (Fossey underground mine). HGM plans to reprocess

tailings under PCE 7386.

The mining lease encompasses the area at Hellyer where the tailings impoundment and the processing

plant are situated. The original lease (CML 103M/1987) was granted on 24 February 1988. It was renewed

and extended in June 2009. It is currently valid until 20 June 2020. An application for an extension of the

lease can only be made no more than three months before and one month after the lease ceases to be

effective.

Section 97(2) of the Mineral Resources Development Act 1995 (MRDA Act) places requirements which

must be met before an application for a lease renewal can be approved. These are:

a. The lessee has submitted a mining plan for the renewal period;

b. The lessee has complied with the conditions of the lease and provisions of this Act;

c. A failure to comply with any conditions of a lease was exempted under section 86; and

d. The lessee has provided a security deposit.

This means that the lease renewal will be granted if the Mine Plan is updated, the security deposit is

reviewed, and there have not been any significant compliance issues relevant to the site.

Condition G7 of PCE 7386 required that a comprehensive EMP review must be submitted to the Director

(of the EPA) for approval by 31 January 2008 and for every three years thereafter, by the third yearly

anniversary. The Hellyer Zinc Concentrate Project Joint Venture (HZCJV) obtained PCE 7386 in October

2006 and operated under that approval until 31 August 2008. As a result, the requirement for a triennial

EMP review was not met. HGM has submitted an updated EMP to the Director (of the EPA) on

6 September 2017 to fulfil this permit requirement.

Currently, there are a number of licences, permits and plans which are effective with the Tasmanian

authorities. This includes the current water licence, approval for use of radiation gauges as well as various

environmental management plans contained within the EMP.

On 17 October 2017, the EPA of Tasmania notified HGM of its acceptance of the abovementioned EMP for

implementation of the Hellyer Project in accordance with the Condition G7 of Permit Conditions –

Environmental (PCE) 7386, as contained in Permit No. DA 138/2006 (EPA, 2017).

For the purposes of this report, it is assumed that all permits and approvals required for the tailings

reclaim and retreatment operation are either in place, or can be obtained by NQM within the necessary

timeframe.

32



Figure 21: Hellyer mining lease

8.4 Environmental Permitting History

Intec and Polymetals formed the HZCJV in 2006. Under this agreement, Polymetals obtained PCE 7386,

refurbished the Hellyer Processing Plant and acted as the operators of the HZCJV until the HZCJV ceased

on 31 August 2008. The Hellyer operation continued under Intec control until 9 September 2008 when it

was placed into care and maintenance by Intec.

Intec sold the Hellyer site to Bass Metals during early 2009. Bass Metals operated the site, obtained PCE

7785 and 7759 and developed the Fossey underground mine, which operated until March 2012 when the

site was placed into care and maintenance due to low metal prices, lower-than-planned metal recoveries

and higher-than-expected operating costs.

33



9 Mineral Processing



9.1 Prior Processing Plant Preparation

The Hellyer plant is a relatively modern base metals processing facility that was commissioned in early

1989 and designed to treat the complex Hellyer fine grained copper-silver-lead-zinc orebody. It was

constructed by Aberfoyle as a 1.0 Mt/a concentrator with then “state of the art” processing equipment

and process control with a large flotation capacity due to the long slurry residence times required to treat

the complex ore.

The Hellyer plant initially treated 1.0 Mt/a with a circuit consisting of a semi-autogenous grind (SAG) mill,

ball mill and sequential copper, lead, zinc and bulk flotation circuits. Testwork and a 30 t/h Luina pilot

plant determined that fine grinding, long circuit residence times (due to slow flotation rates) would be

needed to ensure good separation of the valuable sulphide minerals from each other and from the

dominant pyrite matrix. The plant design was heavily influenced by Cominco who had a major

shareholding in Aberfoyle at the time. The plant was expanded to 1.25 Mt/a in early 1990 with the addition

of pebble crushing, DSM screens in the SAG mill circuit and extra lead and bulk circuit flotation capacity.

Early performance was typified by low overall metal recoveries and low metal recoveries to primary

concentrates. There were problems with understanding and dealing with the metallurgy of the fine-

grained minerals. Gradual improvements in the plant performance took place, which led to improved

selectivity and recovery with an associated reduction in lead and zinc levels in the tailings, as shown in

Figure 22.

Figure 22: Hellyer Tailings lead and zinc grades by year

Source: after Ivy Resources, 2013

At present there is a significant number of reports available for the Hellyer plant, which has a rich history

of metallurgical studies and developments covering nearly 40 years. The original flowsheet and reagent

regime of the Hellyer polymetallic ores were documented in 1988. The Hellyer processing plant was

commissioned in early 1989 with a nominal capacity of 1.0 Mt/a (Ivy Resources, 2013).

34



A summary of important stages and events in processing Hellyer ore is as follows:

• 1986 to 1988: A 250 kt parcel of Hellyer ore was treated through the Cleveland concentrator at Luina

at a throughput of approximately 30 t/h.

• 1989: The Hellyer concentrator with a nominal capacity was commissioned making a copper-silver

concentrate, a lead concentrate, a zinc concentrate and a zinc-lead bulk concentrate.

• 1990: Optimisation increased plant capacity to 1.25 Mt/a.

• 1993: Addition of pebble crushing, reallocating duties of tower mills, use of high intensity

conditioning, and changes in reagent regime and better operating and maintenance practices

continually improved plant performance and operating cost.

• 1999: Pilot-scale testing of tailing retreatment options.

• 2000: Last Hellyer ore processed in June.

• 2000 to 2006: Processing facilities on care and maintenance.

• 2006: In November, reclamation of tailings by dredging commenced with a bulk zinc-lead

concentrate produced by flotation by a joint venture between Intec and Polymetals.

• 2008: In September, bulk concentrate production ceased due to worsening economic conditions.

• 2011 to 2012: Processing of ore mined from the Fossey underground mine.

9.2 Current Plant Configuration

The existing processing plant is in reasonable condition requiring some refurbishment before being

brought back into production. The major equipment items currently available include (Ivy Resources,

2013):

• 1 x 60x48 Kemco Primary Jaw Crushing Plant

• 1 x 25KT Stockpile – Conveyor and Reclaim System

• 1 x SAG Mill - Allis Chalmers 6.7 m x 2.4 m, 2100 kW

• 1 x Ball Mill – Allis Chalmers 4.6 m x 6.1 m 2100 kW

• 3 x vertical Tower Mills – Kubota 335 kW each

• 3 x High Intensity Conditioners – 32 m³, 810 kW each

• 25 x 54 m³ Maxwell Flotation Tank Cells

• 15 x 38 m³ Maxwell Flotation Tank Cells

• 8 x Dorr Oliver Trough Cells – 17 m³

• 3 x Dorr Oliver Trough Cells – 8.5 m³

• 10 x Agitair Trough Cells – 2.5 m³

• 4 x Agitair Trough Cells – 1 m³

• 2 x Ingersoll Rand, Lasta 60 Plate Pressure Filter Presses

• 2 x 20 m diameter Thickeners

• 1 x Lime Slaking Plant

• Bailey 90 Control System

• Courier 5 On Stream XRF Analysers.

The processing plant is contained within a fully enclosed building which is well serviced by several high

capacity overhead traveling gantry cranes throughout each major section (Figure 23). A large covered

concentrate storage shed is also available for concentrate storage and loadout.

35



The processing facilities are currently under a maintenance program, which includes a preparation of the

plant for the new project of retreatment of the Hellyer tailings.

Preventative and mandatory testing is also undertaken such as testing of transformers, pressure vessels,

cranes and electrical systems.

Figure 23: Inside view of the processing plant

9.3 Planned Plant Operation

The planned reprocessing of Hellyer Tailings is designed to produce three saleable products:

• Lead concentrate

• Zinc concentrate

• Gold/silver/pyrite concentrate.

The flowsheet of the Hellyer processing plant is shown in Figure 24.

36



Figure 24: Flowsheet of the Hellyer processing plant

37



The processing plant is expected to operate at an average tailings feed rate of 154 dmt/hr with availability

of 92%, which was achieved during the previous dredging operations in 2006 to 2008. The plant is already

equipped with a high degree of instrumentation and process control equipment with a high level of

automation and on line, real-time measurement and recording.

The rate of 154 dmt/hr has also to be achieved by the feed pumps of the surge tanks near the Main Dam

via the 2,000 m pipeline to the mill.

There are two mill feed tanks outside the mill as shown in Figure 25. Each of them has 12,000–13,000 m3

capacity or 26,000 m3 in total. Both filled can allow a 12 hours work of the processing plant when the

dredge moves to another location.

Figure 25: Tailings storage tanks behind the mill

9.4 Planned Concentrate Production

The metallurgical processing characteristics of the Hellyer ore have been extensively tested and are well

known. Recent tests have been conducted by ALS in Burnie and reviewed by Pitt & Sherry Group (2016)

with the objective of defining a new flowsheet for the reprocessing of Hellyer tailings and optimising metal

recoveries. The testwork was relatively conventional and replicated the original Hellyer flowsheet albeit

at a finer primary and regrind size target, as well as with the inclusion of a pyrite float with payable

precious metals (gold and silver). Forty batch tests were also conducted by ALS as well as three locked

cycle tests to simulate the recirculating loads under balanced circuit conditions. The results obtained

confirmed that a combined lead/silver concentrate can be floated separately from a subsequent zinc

concentrate and thus the opportunity to produce the following three saleable concentrates:

• Combined lead/silver/copper concentrate grading 37% lead and maximising silver content.

• Zinc concentrate grading 45% zinc.

• Gold/silver/pyrite concentrate with a 46% sulphide sulphur content.

To achieve this, in both cases the base metal rougher concentrates were reground prior to cleaning and

following base metal flotation, a high grade (S) pyrite was floated from the residual out of the zinc

38



scavenger circuit tailings. The final tails from the pyrite flotation circuit comprised a low sulphur grade

and other gangue material. Figure 26 below lays out an overall schematic view of the proposed flotation

technology.

Figure 26: Schematic view of the proposed flotation technology

Source: after NQM, 2017

Based on the analysis of these test results, the understanding of the potential reprocessing characteristics

has been improved and a number of options to optimise the process identified.

The metallurgical testwork was undertaken at the ALS in Burnie, who are well known for their capability

with respect to relevant metallurgical testing. The results of this work were reviewed in a report prepared

by Pitt & Sherry Group (2016).

Arsenic (As) is the only deleterious element in the concentrate production. The arsenic is predominantly

presented as arsenopyrite with other arsenic minerals such as tennantite and tetrahedrite. NQM advised

that the concentrates to be produced will contain less than 1% arsenic. The testwork conducted show a

range of 0.5% to 1.5% variation of arsenic in the lead concentrate and averaged under 1%. The results

obtained for the zinc concentrate around 45% zinc grade, indicated that the arsenic values were 0.6% to

1.3% and averaged around 1%. An option may be to blend each of the base metal concentrates to maintain

the 1% limit. The gold/silver/pyrite concentrates contained 0.7% to 1.3% arsenic and averaged around 1%

arsenic.

The assumptions used for the prediction of the concentration production in the current study are

summarised in Table 16. It is worth noting that the grades of the elements in the planned concentrates

differ from the actual average grades from prior reprocessing operations due to the development of the

new flowsheet which includes the production of separate base metal concentrates in comparison with

the production a single bulk concentrate. For example, the lead (Pb) grade of 37% is higher than the actual

average grade of 11% which was achieved during the previous tailings retreatment.

39



Table 17: Planned parameters of concentrate production

Parameter Unit Lead concentrate Zinc concentrate Au/Ag/pyrite concentrate

Pb grade % 37 5

Pb recovery % 47

Zn grade % 6 45

Zn recovery % 38

Au grade g/t 6.9 2 2.77

Ag grade g/t 850 160 64

Mass recovery to concentrate % 51

The schedule for the zinc concentrate production over the life of mine (LOM) is shown in Figure 27. The

schedule reflects the performance of the zinc circuit of the processing plant. Analysis of the figure

indicates that zinc concentrate production is almost stationary with the exception of the period after the

80th month where the production gets a significant increase because of the high-grade tailings of bench

B3 and bench B4 of the Main Dam. After the 100th month, the zinc production drops which is due to the

tailings of the Shale Pit.

The schedule for the lead concentrate production over LOM are shown in Figure 28. The schedule reflects

the performance of the lead circuit. The analysis of the figure indicates that the lead production is almost

stationary with a wide range of fluctuations between 600 and 1,500 Mt/month.

Scheduled production of gold is predicted for all three flotation circuits and the results are shown in

Figure 29. Analysis indicates that the highest production of Au will be achieved from the Au/Ag/pyrite

circuit. The Pb circuit also will produce a high Au concentrate, while the Zn circuit will produce a small

quantity of Au.

The scheduled production of silver for all three concentrates is graphically presented in Figure 30. Analysis

indicates that the highest production of Ag will be achieved from the Au/Ag/pyrite concentrate. The

production of Ag in the Pb concentrate is also high, but it is characterised with obvious variability over the

LOM. The lowest production of Ag is expected in the Zn concentrate, which is pretty stable over the LOM.

A general conclusion can be made that the production of all metals in the concentrates increases in the

period from the 80th until the 100th month of the scheduling horizon due to the extraction of tailings from

the bench B3 and bench B4 of the Main Dam, which cover the lowest zone of the dam having higher

grades.

40



Figure 27: Schedule of zinc production in concentrate

Figure 28: Schedule of lead production in concentrate

41



Figure 29: Schedule of gold production in concentrate

Figure 30: Schedule of silver production in concentrate

CSA Global considers the approach adopted here to be reasonable. As noted, metallurgical testwork

and analyses, equipment selection and performance predictions, have been conducted over a number

of years by reputable organisations such as Como, ASL and the Pit & Sherry Group. The current study

and predictions are now based on this previous work.

Processing plant throughput rates are expected to be achieved with a high degree of confidence.

42



The achievement of improved metallurgical recovery in practice, as predicted by the comprehensive

testwork, is very important for this project. The NPV calculations of this project are based on this.

However, CSA Global notes that to achieve the improved metallurgical recoveries predicted, it is

essential that the plant is appropriately refurbished and necessary improvements/modifications are

completed to align with the proposed flowsheet.

9.5 Tailings Residue Storage

The residue from tailings retreatment in the Hellyer plant needs additional storage capacity due to the

specific characteristics of the residue and the proposed use of dredging for mining. In contrast with the

conventional technology of mining which uses waste dumps, the Hellyer residue needs to be stored under

water to prevent potential oxidisation and acid formation.

The total quantity of residue that will be released from the processing plant after the treatment of the

tailings (9.7 Mt) is 4.2 Mt. The schedule of the residue is graphically illustrated in Figure 31 and presented

in Table 18. The commencement of the schedule is April 2018 when the production process in Hellyer is

planned to start. The schedule covers the whole production period until September 2026 (102 months).

The planned residuals storage capacities are shown in Table 19.

A new TSF (TSF2) will be built and ready for storage of residue by April 2019 (Longey, 2017). The

embankment is proposed to be located approximately 550 m downstream of the Hellyer Main Dam,

located approximately 2 km northwest of the Mill Site. The starter dam will be constructed to RL 638 m,

then be raised progressively by downstream construction method to the ultimate RL 646 m, which is just

below the current crest of the Main Dam at RL 650 m. The new TSF2 footprint is shown in Figure 32. The

TSF2 project engineering will be done by GHD and assumes a settled density for the residue of 1.3 t/m3

versus 1.93 t/m3 density of the tailings due to the removal of the dense sulphides as part of the process.

Column testing is proposed by GHD to commence in November 2017, once reprocessed tailings become

available.

The residue management plan requires 2 m water cover to be maintained due to the acid forming nature

of the tailings, albeit reduced from the current high sulphide risk level post reprocessing. The Hellyer

retreatment process will result in two separate residue streams, one of which would have a lower sulphide

content. The lower sulphide stream could be used in beaching pending its geochemical characterization

and lag time until formation of acid and metalliferous drainage (AMD). The higher sulphide tailings stream

could be deposited sub-aqueously at the pond end of the storage to reduce the risk of oxidation. This plan

could be utilised to reduce the surface area of the higher risk sulphide stream (possible to increase water

depth over the higher risk tailings or lime/cover material applied to reduce oxidisation risk on closure).

The details of the residue management plan will be refined once geochemical characterisation testing and

proportions/settled density of each stream are known. GHD may consider alternatives for the high

sulphide stream, such as deposition in the underground void.

The water released by the residue deposition in the new TSF2 would be pumped back to the existing dams

for reuse in the dredging operations. Makeup water could also be retained in the new TSF2 should it be

required for reprocessing in drier months.

The initial lift of TSF2 will have the capacity of 3,000,000 t and be capable of accepting process residue

until April 2024. From the same month, an extension of the TSF2 will continue the accommodation of the

remaining quantity of residue.

For the period from April 2018 until March 2019, a quantity of 0.49 Mt of residue will need to be

accommodated in accordance with the schedule in Table 18. A planned place is the Finger Pond, the

tailings of which will be transferred to the Main Dam during the pre-production period. The quantity of

493,184 dmt of residue will be subsequently transferred to the new TSF2 by using slurry pump prior to

43



dropping the water level in the Main Dam and Finger Pond just before the dredging of the lower benches

B3 and B4 of the Main Dam.

The transport of the residue from the processing plant to the storage dams will be done with a slurry

pump and pipeline and deposited sub-aqueously. This means that the slurry discharge into the dams will

be submerged and minimise contact with the air and reduce any oxidation potential.

Figure 31: Schedule of monthly tailings residue over LOM

Table 18: Schedule of residue storage

No. Year Month Residuals (dmt) No. Year Month Residuals (dmt)

1 2018 4 41,430 52 2022 7 40,308

2 2018 5 38,549 53 2022 8 42,816

3 2018 6 42,537 54 2022 9 41,601

4 2018 7 40,157 55 2022 10 34,771

5 2018 8 42,992 56 2022 11 43,247

6 2018 9 41,495 57 2022 12 39,737

7 2018 10 34,727 58 2023 1 43,549

8 2018 11 43,260 59 2023 2 42,957

9 2018 12 40,496 60 2023 3 43,247

10 2019 1 43,465 61 2023 4 43,037

11 2019 2 42,059 62 2023 5 39,459

12 2019 3 42,017 63 2023 6 44,179

13 2019 4 41,834 64 2023 7 41,776

14 2019 5 38,460 65 2023 8 44,521

15 2019 6 42,199 66 2023 9 43,120

16 2019 7 39,469 67 2023 10 35,894

17 2019 8 42,128 68 2023 11 44,699

44



No. Year Month Residuals (dmt) No. Year Month Residuals (dmt)

18 2019 9 40,806 69 2023 12 42,866

19 2019 10 34,142 70 2024 1 45,822

20 2019 11 43,155 71 2024 2 41,201

21 2019 12 40,305 72 2024 3 41,956

22 2020 1 43,082 73 2024 4 43,247

23 2020 2 42,014 74 2024 5 40,364

24 2020 3 41,124 75 2024 6 44,689

25 2020 4 41,754 76 2024 7 41,805

26 2020 5 39,405 77 2024 8 42,475

27 2020 6 44,040 78 2024 9 42,278

28 2020 7 41,477 79 2024 10 33,755

29 2020 8 44,618 80 2024 11 40,315

30 2020 9 43,088 81 2024 12 38,161

31 2020 10 35,314 82 2025 1 41,712

32 2020 11 43,945 83 2025 2 40,364

33 2020 12 41,385 84 2025 3 40,665

34 2021 1 44,575 85 2025 4 41,028

35 2021 2 43,297 86 2025 5 39,670

36 2021 3 43,371 87 2025 6 42,256

37 2021 4 42,815 88 2025 7 39,045

38 2021 5 40,683 89 2025 8 42,274

39 2021 6 44,188 90 2025 9 43,516

40 2021 7 41,573 91 2025 10 36,264

41 2021 8 44,520 92 2025 11 44,967

42 2021 9 43,274 93 2025 12 42,066

43 2021 10 35,974 94 2026 1 44,967

44 2021 11 43,190 95 2026 2 43,516

45 2021 12 39,057 96 2026 3 43,516

46 2022 1 42,266 97 2026 4 43,516

47 2022 2 40,441 98 2026 5 40,615

48 2022 3 39,760 99 2026 6 44,967

49 2022 4 40,484 100 2026 7 42,066

50 2022 5 37,318 101 2026 8 44,967

51 2022 6 42,641 102 2026 9 12,320

Table 19: Residue storage capacities

Storage Capacity (dmt) Availability

TSF2 – Initial 3,000,000 From April 2019

TSF2 – Extension 1,208,482 From April 2024

Total 4,208,482

45



Figure 32: New TSF2 downstream of Main Dam

Source: after Longey, 2017

CSA Global notes that the Project has planned well forward and that a Notice of Intent (NOI) has been

submitted, with respect to the construction of TS2 and that the DA approval is to be submitted in

January 2018. It is further noted that the storage capacity of TSF2 is based on the processing plant

handling a total of 9,700,000 dmt which includes material currently considered as an Inferred Mineral

Resource. This is considered to be sound practice.

46



10 Economic Evaluation



10.1 Capital Cost Estimates

All capital and operating cost estimates for the Project were supplied to AusGEMCO by NQM and have

been reviewed by CSA Global. The total capital estimate is US$31.2 million and is detailed in Table 20. All

costs are projected over the LOM period of 114 months from March 2017 till September 2026. Most of the

capex is allocated over the first two years of the project commencement. The highest individual

component of the capital expenditure is the cost of building a new TSF (US$16.1 million).

Table 20: Hellyer Tailings Project – capital expenditure

CSA Global considers the data presented here to be reasonable and the contingency figures stated

should cover any variations that may occur. In terms of the NPV calculation, the timing of any capital

expenditure has a large impact in its effect.

47



10.2 Operating Cost Estimates

NQM supplied the operating costs of the Hellyer Tailings Project to AusGEMCO and these have been

discussed and reviewed by CSA Global.

Table 21 below reflects a summary of some operating costs per unit product and ROM tailings.

Table 21: Hellyer Tailings Project – operating expenditure

Cost Unit Value

Dredging* AUD/dmt 2.84

Hydraulic mining* A$/dmt 2.84

Processing* A$/dmt 18.5

Logistics EXW to FOB A$/wmt 18.55

Logistics FOB to CIF A$/wmt 25

Sales and marketing A$/month 20,000

Shutdown maintenance A$/year 400,000

General and administration A$/LOM 12,234,527

Royalty % net revenue 5

Increase in Mine Closure Bond A$ 2,000,000

* Variable costs

CSA Global considers the figures adopted here, based on details provided by NQM, to be reasonable

and representative. Electrical power consumption by the Project is high but, in favour of the Project,

costs per kWhr provided by hydro-generation are low.

10.3 Financial Model Assumptions

The financial model of Hellyer TSF Project is based on assumptions regarding prices, project discount rate,

penalties for deleterious components in the concentrates, and processing recoveries (NQM, 2017). These

assumptions are reflected in the discounted cash flow model (AusGEMCO, 2017 Appendix D) using a

monthly time step. Table 22 presents the forecasts of Pb, Zn, Au and Ag prices (real 2017) that are used in

the current study. These prices are based on input assumptions from Consensus Economics

(www.consensuseconomics.com).

The Project discount rate is 10%, while the income tax is 30%. The assumption about the recoveries of Zn,

Pb, and Pyrite concentrates are presented in Table 17 of Section 9 “Mineral Processing”.

Based on the analysis of these test results understanding of the potential reprocessing characteristics has

been improved and a number of options to optimise the process identified. The metallurgical testwork

was undertaken at ALS in Burnie, who are well known for their capability with respect to relevant

metallurgical testing. The results of this work were reviewed in a report prepared by Pitt & Sherry Group

(2016).

Arsenic (As) is the only deleterious element in the concentrate production. The arsenic is predominantly

presented as arsenopyrite with other arsenic minerals such as tennantite and tetrahedrite. NQM advised

that the concentrates to be produced will contain less than 1% arsenic. The testwork conducted show a

range of 0.5% to 1.5% variation of arsenic in the lead concentrate and averaged under 1%. The results

obtained for the zinc concentrate around 45% zinc grade, indicated that the arsenic values were 0.6% to

1.3% and averaged around 1%. An option may be to blend each of the base metal concentrates to maintain

the 1% limit. The gold/silver/pyrite concentrates contained 0.7% to 1.3% arsenic and averaged around 1%

arsenic.

The assumptions used for the prediction of the concentration production in the current study are

summarised in Table 17. It is worth noting that the grades of the elements in the planned concentrates

48



differ from the actual average grades from prior reprocessing operations due to the development of the

new flowsheet which includes the production of separate base metal concentrates in comparison with

the production a single bulk concentrate. For example, the Pb grade of 37% is higher than the actual

average grade of 11% which was achieved during the previous tailings retreatment.

Table 22: Forecasting of Zn, Pb, Au and Ag prices

Year Zn (US$/mt) Pb (US$/mt) Au (US$/oz) Ag (US$/oz)

2018 2,800 2,300 1,250 17.7

2019 2,666 2,086 1,249 17.7

2020 2,507 2,028 1,254 17.99

2021 2,433 1,944 1,251 18.24

2022 2,406 1,915 1,215 18

2023 2,228 1,833 1,198 17.87

2024 2,228 1,833 1,198 17.87

2025 2,228 1,833 1,198 17.87

2026 2,228 1,833 1,198 17.87

CSA Global considers the approach adopted here to be reasonable. Beyond physical mining and

processing plant production rates, which have engineering predictability, the data presented must use

forecasting for future metal sales prices. Professional advice has been taken in this regard.

10.4 Discounted Cash Flow Model

Two variants of the economic evaluation of Hellyer Tailings Retreatment Project are summarised (more

detail can be found in AusGEMCO, 2017 Appendix D).

10.4.1 Discounted Cash Flow Model based on the Ore Reserves Estimate

The discounted cash flow (DCF) model of Hellyer Project is organised with the JORC Code (2012) Ore

Reserve estimate of Hellyer Project presented in Table 27. The Mineral Reserve of Main Dam includes the

tailings located in the Finger Pond (0.465 Mt), which is achieved with the development of a combined

Reserve block model with Datamine software. The Ore Reserve model of Western Arm Dam is also

developed also as a combined block model, which includes the Probable Reserve of 0.265 Mt deposited

initially and the unclassified resource of Fossey mine of 0.448 Mt, which was deposited later in the

Western Arm Dam as a top layer. Technologically it is impossible to mine the Reserve layer of tailings

before mining the top layer of Fossey mine tailings. This is the reason to account for the Fossey tailings in

the financial model of Hellyer Project in order prove the economic potential of the Reserve estimates.

The basic metrics of the DCF model are summarised in Table 23. They are calculated using real dollars and

the financial assumptions described above. The NPV is positive (US$113.2 million) for the case which

includes the acquisition cost of the project. The Discounted Cash Flow Internal Rate of Return (DCFIRR) is

90.94% which is much higher than the project discount rate of 10%. This give grounds to accept the Hellyer

Tailing Project as economically effective, which is a precondition for reporting its Ore Reserves in

accordance with the JORC Code (2012). The Discounted Cash Flow Payback Period (DCFPBP) is 28 months.

The NPV calculated without the acquisition cost is US$128.5 million and the DCFIRR is 197.42%. The DCF

model of Hellyer is also presented with an annual time step in Table 25.

10.4.2 Discounted Cash Flow Model with Inclusion of Shale Pit Inferred Mineral Resource

An option of the evaluation of the Hellyer Project is the inclusion of the Inferred Mineral Resource of Shale

Pit. This is a requirement of the JORC Code (2012) Table 1 “Modifying Factors or Assumptions” about the

49



manner in which the Inferred Mineral Resources are utilised and the sensitivity of the outcome of their

inclusion. The resource of Shale Pit is 1.205 Mt and the total quantity of tailings of the DCF model is 9.7 Mt.

The basic metrics of the DCF model are summarised in Table 24. They are calculated using real dollars.

The NPV is positive (US$124.7 million) for the case that includes the acquisition cost of the Project. The

DCFIRR is 91.26% which is much higher than the project discount rate of 10%. This give grounds to accept

the Hellyer Tailing Project as economically effective. The DCFPBP is 28 months. The NPV calculated

without the acquisition cost is US$139.9 million and the DCFIRR is 197.46%. The DCF model of Hellyer is

presented with an annual time step in Table 26.

CSA Global has reviewed the detailed cash flow models and considers the approach to be reasonable

under the particular circumstances of the Hellyer tailings dam which relate to the classifications of

Resources and Reserves compliant with the JORC Code (2012). The Reserves are discussed in greater

detail in the following Section 11 (Ore Reserve Estimate).

Regarding the economic evaluation and calculation of the NPV for the Project, a particular discussion

relates to the inclusion or exclusion of certain unclassified amounts of tailings in this evaluation,

namely:

• 448,000 t of tailings from the Fossey mine deposited on top of partially reworked material in the

Western Arm of the Main Dam where the Reserve was already classified as Probable. To access

this Probable Reserve requires removal of the upper layer of tailings. This material has an

estimated tonnage and contains metal grades based on processing plant data that is considered

acceptably accurate by engineering standards, but does not meet JORC (2012) standards and has

been included in the primary evaluation (Variant 1); and

• 1,210,000 t of tailings held in the Shale Pit which have been assessed as an Inferred Mineral

Resource by CSA Global (2010). It is reasonably expected that this resource will be upgraded by

further testing during the life of the Project and this material is then mined and treated in the final

years of the Project. This extra material has been included in the calculation of NPV (Variant 2).

Particular attention should be paid to Table 23 and Table 24 below and the details in Table 25 and Table

26. Table 23 and Table 25 relate to Variant 1, including Fossey tails (with and without “Acquisition

Costs”) and Table 24 and Table 26 relate to Variant 2 which includes the Inferred Mineral Resource of

the Shale Pit (with and without “Acquisition Costs”).

CSA Global considers that the inclusion of the Fossey Tails in Variant 1 of the NPV calculation is a

reasonable decision necessitated by the physical nature of the upper and lower layers of tailings to be

handled. The calculation of tonnage and grade is based on data from the processing plant when this

material was fed into the plant from production from the Fossey mine with recorded tonnages and

grade. The tonnage and grade of tailings is then back calculated by deduction of the known recovered

tonnage and metal content of concentrates produced for sale.

The inclusion of the Shale Pit Resource in Variant 2 is essentially an observation that the Project life is

very likely to be extended to include this material. Importantly, this extra material has been considered

in the planning of the construction of TSF2 and the permits related to it.

CSA Global notes that overall the Project appears to be economically robust. As such, the inclusion or

exclusion of certain portions of the Resource which are not compliant with JORC (2012), does not appear

to affect the economic viability to any large extent.

50



Table 23: Summary of the DCF model results for Ore Reserve estimate







NPV US$M 113.2

DCFIRR % 90.94

DCFPBP months 28


NPV US$M 128.4

DCFIRR % 197.42

Table 24: Summary of the DCF model results including Shale Pit Inferred Mineral Resource


Net Revenue US$M 708



Cash flows after tax & acquisition US$M 209


NPV US$M 124.7

DCFIRR % 91.26

DCFPBP months 28


NPV US$M 139.9

DCFIRR % 197.46

51



Table 25: DCF model of Hellyer Project based on Ore Reserve estimate

Item Unit Year 2017 2018 2019 2020 2021 2022 2023 2024 2025

Totals 0 1 2 3 4 5 6 7 8

Net Revenue

Lead Silver Concentrate Sales US$ (‘000) 353,121 0 41,417 53,591 46,192 41,589 54,459 36,428 43,961 35,484

Zinc Concentrate Sales US$ (‘000) 112,305 0 12,435 17,915 13,882 13,528 16,748 12,187 16,048 9,562

Gold Silver Concentrate Sales US$ (‘000) 162,747 0 16,493 22,179 22,179 22,241 22,179 22,179 22,179 13,120

Total Net Revenue US$ (‘000) 628,174 0 70,345 93,685 82,253 77,358 93,386 70,794 82,188 58,166

Direct Operating Expenses

Dredging US$ (‘000) -23,065 0 -1,973 -3,630 -2,981 -2,987 -2,980 -3,392 -3,481 -1,639

Processing US$ (‘000) -136,794 0 -13,881 -18,643 -18,643 -18,689 -18,643 -18,643 -18,643 -11,008

Logistics US$ (‘000) -90,318 0 -9,293 -12,568 -12,071 -11,934 -12,775 -11,713 -12,301 -7,663

Sales and marketing US$ (‘000) -1,333 0 -136 -182 -182 -182 -182 -182 -182 -106

General and administration US$ (‘000) -12,124 0 -1,240 -1,653 -1,654 -1,653 -1,653 -1,653 -1,654 -964

Annual and major R&M US$ (‘000) -2,424 0 -303 -303 -303 -303 -303 -303 -303 -303

Operating Contingency US$ (‘000) -13,303 0 -1,341 -1,849 -1,792 -1,787 -1,827 -1,794 -1,828 -1,084

Other Operating Expenses

Royalty US$ (‘000) -35,405 0 -2,420 -4,668 -4,277 -3,773 -4,651 -3,841 -3,629 -3,736

Closure/rehabilitation costs US$ (‘000) -1,515 0 -1,515 0 0 0 0 0 0 0

CAPEX US$ (‘000) -31,254 -15,534 -5,985 -3,295 0 0 0 -2,809 -3,630 0

Total Operating Expenses US$ (‘000) -388,609 -15,534 -38,087 -46,792 -41903 -41308 -43015 -44,332 -45,651 -26,503

Operating Cashflow b/f Tax US$ (‘000) 319,499 -15,534 32,258 46,893 40,350 36,050 50,371 26,462 36,536 31,663

Tax US$ (‘000) -96,168 0 -7,791 -13,820 -12,268 -9,497 -14,232 -9,819 -8,081 -11,681

Net Cashflow after Tax US$ (‘000) 223,332 -15,534 24,467 33,072 28,082 26,553 36,139 16,643 28,456 19,982

Less Acquisition Cost US$ (‘000) -15,152 -15,152 0 0 0 0 0 0 0 0

Cashflow after Tax and Acquisition US$ (‘000) 208,180 -30,685 24,467 33,072 28,082 26,553 36,139 16,643 28,456 19,982

Discount factor Unit 1 1 1 1 1 1 1 1 0

Discounted cashflows @ 10% US$ (‘000) -30,685 22,243 27,332 21,099 18,136 22,440 9,395 14,602 9,322

Project NPV @ 10% US$ (‘000) 113,263

DCFIRR percent 90.94

DCFPBP months 28

52



Table 26: DCF model of Hellyer Project with inclusion of Inferred Mineral Resource

Item Unit Year 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026

LOM 0 1 2 3 4 5 6 7 8 9

Net Revenue

Lead Silver Concentrate Sales US$ (‘000) 401,136 0 41,417 53,591 46,192 41,589 54,459 36,428 43,961 54,224 29,275

Zinc Concentrate Sales US$ (‘000) 119,579 0 12,435 17,915 13,882 13,528 16,748 12,187 16,048 12,746 4,090

Gold Silver Concentrate Sales US$ (‘000) 187,393 0 16,493 22,179 22,179 22,241 22,179 22,179 22,179 22,241 15,525

Total Net Revenue US$ (‘000) 708,109 0 70,345 93,685 82,253 77,358 93,386 70,794 82,188 89,211 48,889

Direct Operating Expenses

Dredging US$ (‘000) -26,022 0 -1,973 -3,630 -2,981 -2,987 -2,980 -3,392 -3,481 -2,731 -1,865

Processing US$ (‘000) -157,653 0 -13,881 -18,643 -18,643 -18,689 -18,643 -18,643 -18,643 -18,689 -13,178

Logistics US$ (‘000) -103,176 0 -9,293 -12,568 -12,071 -11,934 -12,775 -11,713 -12,301 -12,497 -8,024

Sales and marketing US$ (‘000) -1,545 0 -136 -182 -182 -182 -182 -182 -182 -182 -136

General and administration US$ (‘000) -14,053 0 -1,240 -1,653 -1,654 -1,653 -1,653 -1,653 -1,654 -1,653 -1,240

Annual and major R&M US$ (‘000) -2,727 0 -303 -303 -303 -303 -303 -303 -303 -303 -303

Operating Contingency US$ (‘000) -15,259 0 -1,341 -1,849 -1,792 -1,787 -1,827 -1,794 -1,828 -1,803 -1,237

Other Operating Expenses

Royalty US$ (‘000) -35,405 0 -2,420 -4,668 -4,277 -3,773 -4,651 -3,841 -3,629 -4,866 -3,281

Closure/rehabilitation costs US$ (‘000) -1,515 0 -1,515 0 0 0 0 0 0 0 0

CAPEX US$ (‘000) -31,254 -15,534 -5,985 -3,295 0 0 0 -2,809 -3,630 0 0

Total Operating Expenses US$ (‘000) -388,609 -15,534 -38,087 -46,792 -41,903 -41,308 -43,015 -44,332 -45,651 -42,724 -29,264

Operating Cashflow b/f Tax US$ (‘000) 319,499 -15,534 32,258 46,893 40,350 36,050 50,371 26,462 36,536 46,488 19,625

Tax US$ (‘000) -96,168 0 -7,791 -13,820 -12,268 -9,497 -14,232 -9,819 -8,081 -14,970 -5,690

Net Cashflow after Tax US$ (‘000) 223,332 -15,534 24,467 33,072 28,082 26,553 36,139 16,643 28,456 31,517 13,935

Less Acquisition Cost US$ (‘000) -15,152 -15,152 0 0 0 0 0 0 0 0 0

Cashflow after Tax and Acquisition US$ (‘000) 208,180 -30,685 24,467 33,072 28,082 26,553 36,139 16,643 28,456 31,517 13,935

Discount factor unit 1 0.91 0.83 0.75 0.68 0.62 0.56 0.51 0.47 0.42

Discounted cashflows @ 10% US$ (‘000) -30,685 22,243 27,332 21,099 18,136 22,440 9,395 14,602 14,703 5,910

Project NPV @ 10% US$ (‘000) 124,749

DCFIRR percent 91.26

DCFPBP months 28

53



11 Ore Reserve Estimate



The Hellyer TSF Ore Reserve estimate is reported in Table 27 and was compiled for NQM as a part of the

mine planning study undertaken by AusGEMCO, which has been reviewed by CSA Global. Following a

process of detailed mine optimisation using the author’s original method that incorporated Geological

Block Profit Modelling, mine design, sequencing, production scheduling economic evaluation and mine

project risk assessment, the Ore Reserve estimate was prepared using the Guidelines of the Australasian

Code for Reporting of Ore Reserves (JORC Code, 2012).

The Ore Reserve estimate relates specifically to the conversion of Measured and Indicated Mineral

Resources of the Hellyer Tailings Project and includes consideration of the modifying factors documented

in Appendix 1 – JORC Code Table 1.

After the completion of the Mineral Resource estimate in 2010 by CSA Global, tailings were deposited in

the Western Arm Dam from the processing of ore mined in the Fossey Mine. The quantity and quality of

the tailings were recorded by the management of the Hellyer processing plant and they are excluded in

the current Ore Reserve estimate. At present the Fossey tailings are unclassified while the tailings beneath

them in the Western Arm Dam are classified as a Probable Reserve. The extraction of the Probable Reserve

of the Western Arm Dam is possible only with the extraction of the Fossey tailings. This is the reason for

considering the Fossey tailings in the mine plan and financial model of Hellyer Project.

The tailings of Shale Pit, which have been re-classified by CSA Global in August 2017 as “Inferred” Mineral

Resource, are also excluded in the Ore Reserve estimate because of the requirements of the JORC Code

(2012). These tailings have an economic potential and are also analysed in the mine plan and financial

model of the Hellyer Project as an option.

A summary of the quantities and qualities of the tailings of Fossey Mine and Shale Pit is presented in

Table 28.

No cut-off grade(s) has been used in the reserve estimation due to the specific nature of mining the tailings

by dredging. Details on the implementation of other modifying factors are presented in Section 7 “Mine

Design, Sequencing and Scheduling”. Environmental, Law Compliance and Permitting details outlined in

Section 8 are considered, along with Mineral Processing details in Section 9.

All material classified as Measured Mineral Resources in Hellyer TSF has translated into Proved Ore

Reserves while all material classified as Indicated Mineral Resources has translated to Probable Ore

Reserves. The reserve conversion is based on the CSA Global Mineral Resources estimate updated 24

August 2017 (CSA Global, 2017).

54



Table 27: Hellyer TSF Ore Reserve estimate* (6 September 2017)

Ore Reserves per asset

Gross Operator

Tonnage (Mt)


NQM


Pb (t) (‘00)

Ag (oz) (‘000)

Au (oz) (‘000)

Cu (t) (‘00)

Proven 2.05 3.31 3.35 94 2.63 0.21 679 687 6,212 173 43

Probable 5.99 2.29 2.95 93 2.55 0.18 1,372 1,767 17,941 491 108

Total 8.04 2.55 3.05 93 2.57 0.19 2,050 2,454 24,153 664 151

Net attributable

Tonnage (Mt)


Zn % Pb % Ag g/t Au g/t Cu %

Zn (t) (‘00)

Pb (t) (‘00)

Ag (oz) (‘000)

Au (oz) (‘000)

Cu (t) (‘00)

Proven 2.05 3.31 3.35 94 2.63 0.21 679 687 6,212 173 43

Probable 5.99 2.29 2.95 93 2.55 0.18 1,372 1,767 17,941 491 108

Total 8.04 2.55 3.05 93 2.57 0.19 2,050 2,454 24,153 664 151

Note: Mt and Zn, Ag, Au and Cu grades are rounded to two decimal places, Ag to the nearest whole number. Contained metal tonnes are reported to nearest hundred, and ounces to nearest thousand tonnes.

Table 28: Hellyer Non-JORC Reserves (31 August 2017) of Fossey tails in the Western Arm and Shale Pit tails

Non-JORC Reserves per asset

Gross Operator

Tonnage (Mt)


NQM


Pb (t) (‘00)

Ag (oz) (‘000)

Au (oz) (‘000)

Cu (t) (‘00)

Fossey 0.45 1.84 1.37 35 1.80 0.03 83 62 512 26 1

Shale Pit 1.21 1.00 2.60 86 2.57 0.19 121 315 3,346 100 23

Total 1.66 1.23 2.27 72 2.36 0.15 204 376 3,857 126 24

Note: Mt and Zn, Ag, Au and Cu grades are rounded to two decimal places, Ag to the nearest whole number. Contained metal tonnes are reported to nearest hundred, and ounces to nearest thousand tonnes.

CSA Global notes that the total tonnages now amount to Proven 2.05 Mt plus Probable 5.99 Mt plus non-JORC 1.66 Mt = 9.7 Mt as used in Variant 2 NPV

calculation.

55



The Hellyer Ore Reserve Datamine files are described in Table 29. The grid files of each dam with the

BLOCKID data and (X, Y) coordinates of each block of the mine design are described in Table 30. The grid

files have to be used for location of each block during the dredging operations. The Z coordinates of the

blocks correspond to the RLs of dam benches, which are summarised in Table 15 of Section 7.

Table 29: Hellyer Ore Reserve Datamine files

Dam File Comment

Main

Reserve_mainbench1.dm

Reserve_mainbench1co.dm With Finger Pond tailings


Reserve_mainbench2co.dm With Finger Pond tailings



Western Arm Reserve_westbench1.dm

Reserve_wadfos1.dm With Fossey tailings

Finger Pond Reserve_eastb1.dm

Reserve_eastb2.dm

Shale Pit

Reserve_shaleb1.dm

Reserve_shaleb2.dm

Reserve_shaleb3.dm

Table 30: Hellyer Ore Reserve Grid Datamine files

Dam Datamine File Excel file

Main MainDam12_grid.dm MainDam12_grid.csv

Western Arm WesternArm_grid.dm WesternArm_grid.csv

Finger Pond FingerPond_grid.dm FingerPond_grid.csv

Shale Pit ShalePit_grid.dm ShalePit_grid.csv

CSA Global has reviewed the Ore Reserve estimates shown in Table 27 Hellyer TSF Ore Reserve Estimate

(6 September 2017) and these are fully compliant with JORC Code (2012).

Table 28 shows additional material marked Fossey Tails (0.45 Mt) and Shale Pit tails (1.21 Mt).

The Shale Pit tails have been clearly classified as an Inferred Mineral Resource by CSA Global (2010).

This resource requires further testing to be upgraded to an Indicated Mineral Resource and thence a

Probable Ore Reserve if demonstrated economic. This is most likely to occur, and this extra material

has been included in the calculation of the NPV (Variant 2), but the Resource is excluded from the JORC

(2012) Ore Reserve estimates.

Fossey tails, on the other hand, pose a dilemma which has been discussed in detail by CSA Global

competent personnel together with the author of the AusGEMCO Report which includes this material

in Variant 1 of the NPV calculation. These tails overlie a Probable Resource, known as Western Arm. A

portion of Western Arm Resource was dredged and treated in an earlier phase of operations by Bass

Metals, leaving a void. This void was then conveniently used for deposition of tailings during the

working of Fossey mine. This material has not been drilled or tested in any way. However, it is

considered that the tonnage and grades estimates based on feed data to the processing plant, less

metals recovered as concentrates, provide sufficiently accurate information for this material to be

incorporated into the Block Models of the AusGEMCO study. The AusGEMCO study then includes this

material in the Mining Schedule which is then utilised in the economic evaluation to generate a project

NPV. This material must be removed in any case prior to the mining of the underlying Western Arm Ore

56



Reserve. This Fossey material carries valuable metal grades, albeit at a lower grade than the average

grades of the project, but nevertheless economic under similar mining and treatment methods. It would

be detrimental to the Project to discard this material as valueless.

As stated above, AusGEMCO includes this material in the NPV calculation named Variant 1.

CSA Global has noted that project economics are robust. The inclusion of Fossey tails material, which is

not compliant with JORC (2012), does not affect the conclusion in regard to the economic viability of

the Measured and Indicated Resources to Ore Reserve status. The inclusion of this material is done to

pursue a logical mining schedule, maximising mining recovery and delivery to the processing plant.

CSA Global considers this to be a reasonable and valid method to handle the valuation of this material.

57



12 Risk assessment



12.1 Ore Reserve Risk Assessment

In this Report the Ore Reserve Risk has been assessed by using the Monte Carlo method to generate

randomly distributed sample estimates for geological parameters such as in-situ Pb grade, Zn grade, Au

grade, Ag grade and tailings density. However, it is important to note that, for the purposes of calculation

efficiency, the cash flow model used for this risk analysis was simplified in terms of tax and similar

quarterly based calculations. This simplified model resulted in a slightly lower NPV (US$118.9 million)

when compared to the NPV (US$124.7 million) calculated from the more detailed monthly cash flow for

the project model referred in the economic evaluation in Section 10. Consequently, for the purposes of

the risk analysis below the author has used the value of US$118.9 million as the deterministic project value

in the context of the approach applied for assessing the Ore Reserve Risk. All other inputs to the cash flow

model are assumed to be constant.

Two variants were used for the assessment of the standard deviations of the geological variables:

• Variant 1 used the monthly estimates of the grades in the production schedule to determine the

standard deviation. The estimates are summarised in Table 31.

• Variant 2 assessed standard deviations by using the results of CSA Global variogram modelling of the

Hellyer Mineral Resource (CSA Global, 2010), which captures the short-range variability. The

estimates are summarised in Table 32.

Estimates of variation of tailings density were obtained using standard deviation of 10% around the

assumed constant tailings density (1.93 t/m3) as applied in the Geological Block Model.

Two criteria are used for assessing the Ore Reserve Risk:

• Risk that the stochastic NPV is less than the deterministic NPV.

• Risk that the stochastic NPV is less than zero.

The results obtained from 300 Monte Carlo simulations of the NPV for Variant 1 by treating the Ore

Reserve parameters as random quantities are graphically illustrated as a histogram in Figure 33. The

estimates of the Ore Reserve Risk are presented in Table 33. The variant of using the NPV (US$119 million)

of the deterministic DCF analysis as a critical level of the Ore Reserve Risk shows the risk estimate of 32%,

which means that the stochastic NPV is higher than the deterministic NPV. In other words, this also means

an underestimation of the deterministic NPV using only average estimates of the tonnage and grades of

Ore Reserves.

The results of using the second criterion of the risk analysis presented in Table 33 show a zero Ore Reserve

Risk because the stochastic NPV of US$122 million is much higher than zero. These results give ground to

conclude that the current estimates of the variability of the tailings grades and density will not have a

significant negative impact on the successful and profitable exploitation of Hellyer tailings.

The results obtained from the simulation of the NPV by using the input data of Variant 2 are shown as a

histogram in Figure 34. The estimates of the Ore Reserve Risk are summarised in Table 34. The risk that

the stochastic NPV is less than the deterministic NPV is 35%, while the risk that the stochastic NPV is less

than zero is 0%. These estimates confirm the validity of the conclusions made for Variant 1 that:

• Stochastic NPV is higher than the deterministic NPV, which indicates an underestimation of the

deterministic NPV.

58



• Stochastic NPV is significantly higher than zero.

• The estimates of ore grades and density will not have a significant negative impact on the profitable

exploitation of Hellyer tailings.

Table 31: Statistical input parameters of Ore Reserve Risk assessment

Statistics Pb (%) Zn (%) Au (g/t) Ag (g/t) Density (t/m3)

Mean 2.89 2.28 2.52 89.5 1.93

Standard deviation 0.54 0.77 0.19 13.67 0.19

Table 32: Input parameters of variogram modelling

Element Nugget Sill Variance Standard deviation

Pb 0.1 0.21 0.11 0.33

Zn 0.2 0.57 0.37 0.61

Au 0.2 0.8 0.6 0.78

Ag 0.2 0.29 0.09 0.3

Source: after CSA Global, 2010

Figure 33: Histogram of NPV estimates – Variant 1

59



Figure 34: Histogram of NPV estimates – Variant 2


NPV mean (US$) NPV Std. Dev. (US$) Critical level (US$) Probability (%) Risk (%)

122,494,493 7,724,451 118,901,024 67.91 32.09

122,084,493 7,724,451 0 100 0


NPV mean (US$) NPV Std. Dev. (US$) Critical level (US$) Probability (%) Risk (%)

120,940,367 5,381,536 118,901,024 64.76 35.24

120,940,367 5,381,536 0 100 0

12.2 Mining Project Risk Assessment

The methodology used for Mining Project Risk assessment is described in the references (Halatchev et al.,

2005; Halatchev, 2007; Davis et al., 2007). It deals with the concept of the development of a stochastic

DCF model over the LOM and treatment of all parameters of the mining project as random quantities or

functions depending on the available input data. The criterion for defining the Mining Project Risk are the

DCFs of the mining project assessed with the DCF model. This criterion defines the risk of not achieving

positive DCFs at each time step of the DCF analysis, which is a logical requirement of mining business.

Such a formulation of the Mining Project Risk model means that a strategy for achieving positive

discounted cash flows over the LOM is set by the mining company.

The Mining Project Risk model can be presented with the formula:

Rp (ti ) PrDCF(ti ) 0, i

Where: DCF(ti ) are the discounted cash flows at time ti

The Mining Project Risk assessment in this study has been performed with regards to the following

parameters:

• Geological parameters – grades of Pb, Zn, Au, Ag and density of the tailings.

60



• Mineral processing parameters – recoveries of Pb in the Pb concentrate, Zn in the Zn concentrate,

and pyrite in the pyrite concentrates; grades of Pb, Zn, Au and Ag in the Pb, Zn and Pyrite

concentrates.

• Economic parameters – prices of Pb, Zn, Au, Ag and Pyrite concentrate; operating costs for mining

(dredging and water cannon washing), mineral processing, logistics, sales and marketing, general and

administration, maintenance, contingency, royalty, and mine closure as well as capex.

The standard deviations of the geological parameters are assessed using the data of the production

scheduling model developed on a monthly basis. The standard deviations of the mineral processing and

economic parameters are assessed as 10% of the mean estimates, which is considered a reasonable

estimate at this stage of the Project’s development.

To assess the Mining Project Risk, 300 Monte Carlo simulations were run; Table 35 summarises the results

of this analysis (the monthly mean and standard deviation of DCFs). These stochastic DCFs are illustrated

in Figure 35 and analysis of them indicates negative cash flows for the first period till the 14th month. After

that period, the cash flows have a non-stationary character of variation. There are a few stages which

start with some peak estimate and then it declines to a certain level. All cash flows are positive except the

ones predicted for 23rd and 83rd months, which are affected by the increase of capex and mine closure

costs (step 23) and increase of mining costs because of using water cannon washing technology and

additional capex (step 83). The payback period of the stochastic cash flows profile is 28 months. It is

important to note that the cash flows are assessed with an income tax paid every month of the year.

The results of Mining Project Risk assessment are shown in Figure 36, which indicates that the risk is 100%

for the period till the 14th month. This is to be expected because the planned cash flows are negative for

that period. After that period, the risk varies within the range from 0 to 10.00%. There are three short

periods of volatile risk assessments. The first two of these periods have a risk increase up to 60.00% and

the reason for that increase is explained above regarding the profile of the stochastic cash flows. The third

period is during shutdown of the mining operations and mine closure.

Generally, the predicted profile of the Mining Project Risk is attractive for investment and useful for the

decision-making process. It can help the management of NQM for shaping a strategy about the future

mitigation of the Mining Project Risk and achieving a sustainable exploitation of the Hellyer tailings.

61



Figure 35: Stochastic DCFs profile

Figure 36: Mining Project Risk profile over LOM

62



Table 35: Results of Mining Project Risk assessment

Time (month)

DCF mean (US$)

DCF Std. Dev. (US$)

Risk (%)

Time (month)

DCF mean (US$)

DCF Std. Dev. (US$)

Risk (%)

0 0 0 0 59 2,156,163 863,961 0.63

1 0 0 0 60 2,067,160 819,251 0.58

2 -15,343,557 0 100 61 2,083,450 788,281 0.41

3 -1,727,944 0 100 62 2,274,467 858,015 0.4

4 -430,200 0 100 63 1,938,114 738,917 0.44

5 -821,584 0 100 64 1,923,831 716,660 0.36

6 -2,078,675 0 100 65 1,680,292 741,559 1.17

7 -2,155,284 0 100 66 1,644,968 728,831 1.2

8 -879,689 0 100 67 1,750,163 666,327 0.43

9 -4,027,231 0 100 68 1,674,539 724,785 1.04

10 -269,194 0 100 69 1,273,666 548,627 1.01

11 -2,070,144 0 100 70 1,642,420 717,899 1.11

12 -11,570 0 100 71 1,663,927 632,739 0.43

13 -11,478 0 100 72 1,503,483 711,801 1.73

14 -11,388 0 100 73 1,104,077 568,653 2.61

15 3,222,566 1,098,168 0.17 74 1,000,785 550,287 3.45

16 2,743,346 1,161,728 0.91 75 1,108,925 614,944 3.57

17 3,015,227 1,164,503 0.48 76 1,214,809 624,338 2.58

18 2,613,430 1,184,736 1.37 77 1083,307 598,756 3.52

19 2,781,726 1,240,824 1.25 78 901,063 561,926 5.44

20 2,707,118 1,359,857 2.33 79 1,045,314 576,317 3.49

21 548,740 833,170 25.51 80 962,459 636,287 6.52

22 1,284,957 1,211,830 14.45 81 742,374 430,074 4.22

23 -375,631 1,132,449 62.99 82 88,434 611,388 44.25

24 1,567,468 1,047,957 6.74 83 -140,387 500,002 61.06

25 1,515,260 943,281 5.41 84 163,418 524,223 37.76

26 1,793,814 875,714 2.03 85 934,100 572,262 5.13

27 1,812,371 856,139 1.71 86 883,125 612,660 7.47

28 2,150,726 1,041,484 1.95 87 639,952 626,283 15.34

29 2,352,329 1,075,374 1.44 88 873,666 508,280 4.28

30 2,610,722 1,126,853 1.03 89 876,565 496,279 3.87

31 2,927,492 1,068,187 0.31 90 906,174 524,741 4.21

32 2,779,520 1,238,564 1.24 91 1,556,278 573,350 0.33

33 2,067,231 716,396 0.2 92 1,230,474 572,027 1.57

34 2,352,805 1,016,481 1.03 93 1,306,998 560,526 0.99

35 2,216,302 903,096 0.71 94 1,991,042 663,038 0.13

36 2,386,726 1,135,469 1.78 95 1,748,127 613,932 0.22

37 1,991,821 974,277 2.05 96 1,744,858 650,112 0.36

38 2,471,540 946,519 0.45 97 1,631,364 548,622 0.15

39 2,146,210 893,051 0.81 98 1,544,452 563,035 0.3

40 1,979,834 940,677 1.77 99 1,466,855 633,427 1.03

41 1,673,789 844,394 2.37 100 1,473,742 588,225 0.61

42 1,555,004 882,848 3.91 101 1,503,839 590,226 0.54

43 1,528,943 888,962 4.27 102 1,541,311 556,752 0.28

63



Time (month)

DCF mean (US$)

DCF Std. Dev. (US$)

Risk (%)

Time (month)

DCF mean (US$)

DCF Std. Dev. (US$)

Risk (%)

44 1,467,885 890,371 4.96 103 1,552,018 559,514 0.28

45 1,462,540 666,622 1.41 104 801,455 534,871 6.7

46 1,844,794 818,417 1.21 105 623,899 336,128 3.17

47 1,411,815 719,725 2.49 106 860,219 513,017 4.68

48 1,431,476 883,082 5.25 107 780,484 374,388 1.85

49 1,292,134 711,470 3.47 108 891,370 461,414 2.67

50 1,259,286 864,546 7.26 109 984,438 452,841 1.49

51 1,458,474 774,706 2.99 110 795,499 430,308 3.23

52 1,726,871 802,822 1.57 111 780,583 418,814 3.12

53 1,367,052 792,240 4.22 112 721,791 428,185 4.59

54 1,315,993 749,012 3.95 113 786,606 448,559 3.97

55 1,396,630 772,860 3.54 114 749,072 440,629 4.46

56 1,221,762 762,370 5.45 115 868,508 505,369 4.28

57 1,018,433 636,821 5.49 116 57,715 144,785 34.51

58 1,901,282 748790 0.56 117 138,316 106,443 9.69

CSA Global considers the approach taken to assess economic and mining risk by detailed statistical

means to be entirely appropriate.

CSA Global cautions that, notwithstanding these analyses, the possibility of a catastrophic natural event

severely affecting production time, and hence forecasts of economics, still exists.

64



13 Conclusions

CSA Global concludes the following;

• CSA Global was engaged by NQM to compile and report a Mineral Resource estimate and to complete

a gap analysis and review the Ore Reserve Study completed by AusGEMCO of the Hellyer Tailings

Retreatment Project, located in Tasmania, Australia.

• NQM holds a consolidated granted mining lease CML103M/1987 over the area including the tailings

dam and the processing plant at Hellyer which cover all areas required for the proposed tailings

reclaim operations. The mining lease is held by Hellyer Gold Mines Pty Limited, a wholly owned

subsidiary of NQM. The lease was granted for a period of 32 years from 1 February 1988 to 30 June

2020. The annual permit has been paid in full until 15 February 2018.

• The lease area has a polymetallic (Cu-Pb-Zn-Au-Ag) deposit and associated ore processing facility at

Hellyer, adjacent to the Tailings Dam, and is currently under care and maintenance. The economic

metals mined at Hellyer were lead, zinc, copper, gold and silver. The Hellyer deposit was mined by

underground methods during the period 1989 to 2000. The Fossey deposit extends down plunge

from the Hellyer deposit and was mined by Bass between 2010 and 2012.

• Tailings from the mill were deposited in a depression approximately 1 km to the west of the Hellyer

mill. The tails were subsequently inundated with water to prevent oxidation of the sulphide species

present in the tails. The tails were partially dredged by Polymetals and retreated between 2006 and

2008, with the reprocessed tails deposited in the Shale Pit and Western Arm. Tails from processed

Fossey deposit ore were also discharged into the Western Arm.

• CSA Global estimated a Mineral Resource for the Hellyer Tailings Retreatment Project, which is

reported in accordance with the JORC Code. Samples from Vibracore drilling carried out in 1998 and

2000 were used to interpolate grades for Zn, Pb, Cu, Ag and Au into a block model, constructed using

digital terrain models representing original topography and top of tailings surfaces. The block model

is depleted to capture the volumes of tailings dredged in the years 2006 to 2008. A density of

1.93 t/m3 was applied to all blocks in the Mineral Resource model. The block model was classified as

a combination of Measured, Indicated and Inferred Mineral Resources. The Measured and Indicated

Mineral Resource categories were assigned to volumes supported by varying density of Vibracore

drilling. The Shale Pit, located 200 m to the southeast of the Main Dam, was filled with reprocessed

tails following dredging activities, and is classified as Inferred Mineral Resources.

• Testing of the resource using Vibracore drilling methods, sample collection and subsequent assay

methods are considered valid and reliable and stated as such in reports by CSA Global (2010 and

2017). This data has been used to generate Block Profit Models, and organise a mining path and

mining schedule, which in turn generates product values.

• The choice of the mining and transport methods for delivery of these tailing materials to the

processing plant for further processing and recovery of valuable metals is considered entirely

appropriate. There are clear historical records of production rates using this equipment to back up

these estimates. Operating costs have been calculated in detail and also capital costs to refurbish old

equipment to full operating status. This data is considered entirely appropriate.

• Detailed metallurgical sampling and testing has been carried out to provide recommendations for

processing plant modifications, together with engineering studies to estimate the capital costs and

operating costs using these modifications, with the aim of improving recovery of metal values. Based

on this work the assumptions used in regard to metallurgical recovery are considered appropriate

and valid. Capital costs and operating costs have been researched and estimated in detail. These

figures are considered appropriate and valid.

65



• Forecasts of future metal value to estimate sales value of the concentrates produced are reliant on

professional third-party consultants and are appropriate.

• The economic viability of the Project has been estimated using traditional and standard means of

DCF analysis which then provide an estimate of NPV and project rates of return.

• In conclusion, with respect to Ore Reserves, CSA Global considers the amounts stated in Table 27 of

Section 11 (Ore Reserve Estimate) to be compliant with JORC Code (2012) and is prepared to sign-off

on this basis.

66



14 References

33rd APCOM Symposium, April 24-27, 2007, Santiago, Chile, pp.729-739.

Ausenco, 2016. NQ Minerals Hellyer Project. Tailings Treatment Restart Cost Estimate. Report, September 2016.

Bolger, C. 2006, “Resource Estimate, Hellyer Tailings Project”. PolyMetals Group.

Bolger, C. 2007 “Resource Estimate Update January 2007, Hellyer Tailings Project”. PolyMetals Group.

CM Insight, 2016. Project Hellyer. Evaluation of Tailings Reprocessing Options, Report, October 2016

Como Engineers, 2013. Hellyer Gold Project. Definitive Feasibility Study. Report, August 2013.

Como Engineers, 2016. Hellyer Gold Project. Pre-Feasibility Study. Report, June 2016.

CSA Global, 2010. Resource estimate and mining study. Hellyer Tailings Retreatment Project-Tasmania.

CSA Global, 2017. Hellyer Tailings Retreatment Project-Tasmania. Competent Persons Report, August 2017. Report No. Nº R301a 2017.

Davis, G.A., Halatchev, R.A., and Potvin, Y., 2007. Risk analysis and economic valuation of mining projects. Manual. Mercure Hotel, Brisbane, 3-5 April 2007. Course No. 07AA, Australian Centre for Geomechanics, Perth, WA.

Department of Infrastructure, Energy and Resources – Tasmania, 2012. Lease Search as at 17 October 2012.

EPA, 2017. Hellyer Gold Mines 2017 Environmental Management Plan, Letter, 17 October 2017.

Halatchev, R.A., 2007. An approach to variable discount rate modelling of open pit gold mine projects. Proc.

Halatchev, R.A., 2009. Profit Margin Ranking of the Elimatta Coal Project. Sponsor: Minserve Group Pty Ltd.

Halatchev, R.A., Lever, P.J.A., 2005. Risk model of long-term production scheduling in open pit gold mining.

Halatchev, R.A. & Gabeva, D., 2017. Hellyer Tailings Retreatment Project, Tasmania – Ore Reserve Estimate, September 2017.

Ivy Resources, 2013. Hellyer Gold Mines Project. Project Feasibility Study, Report, November 2013.

Joint Ore Reserves Committee 2012. “Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves. The JORC Code, 2012 Edition”. Prepared by: The Joint Ore Reserves Committee of The Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia (JORC).

Longey, R., 2017. Memorandum. New TSF options study. GHD, 10 September 2017.

NQ Minerals, 2017. Hellyer financial model 2017_03_28.xlsx.

Pitt & Sherry, 2016. Hellyer Tailings Reprocessing Metallurgical Status Report, November 2016.

Proc. 2005 Australian Mining Technology Conference, pp.285-296, Fremantle, WA.

Tyrell, J.P. 2006 “Hellyer Tailings Resource Estimate”. AMC Report 205082, for Intec Hellyer Metals Pty Ltd.

Tyrell, J.P. 2009 “Hellyer Tailings Resource Estimate, 2009 Update”. AMC Report 209038, for Bass Metals Ltd.

67



15 Glossary

Below are brief descriptions of some terms used in this report.

Sections 1 to 5

Arsenopyrite: (Fe, AsS), a sulphide mineral, the chief ore mineral of arsenic and can be associated with Au.

Base Metals: In mining and economics, the term “base metals” refers to industrial non-ferrous metals excluding precious metals. These include copper, lead, nickel and zinc.

Block Model: A three-dimensional model of blocks, into which grade data is interpolated.

Fire Assay: A quantitative analytical technique for the determination of metal or metals from impurities via a fusion process.

Galena: PbS, lead sulphide, the chief ore mineral of lead

Gangue: The waste, or non-economic component, of ore; for example, quartz, and non-economic sulphides such as pyrite.

Indicated Mineral Resource: An “Indicated Mineral Resource” is that part of a Mineral Resource for which quantity, grade (or quality), densities, shape and physical characteristics are estimated with sufficient confidence to allow the application of Modifying Factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes, and is sufficient to assume geological and grade (or quality) continuity between points of observation where data and samples are gathered.

Inferred Mineral Resource: An “Inferred Mineral Resource” is that part of a Mineral Resource for which quantity and grade (or quality) are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade (or quality) continuity. It is based on exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes.

JORC: The Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves, as published by the Joint Ore Reserves Committee of The Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia.

Mill: In mining, a processing facility.

Mineral Resource: A “Mineral Resource” is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade (or quality), and quantity that there are reasonable prospects for eventual economic extraction. The location, quantity, grade (or quality), continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling. Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured Mineral Resource categories.

Measured Mineral Resource: A “Measured Mineral Resource” is that part of a Mineral Resource for which quantity, grade (or quality), densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit. Geological evidence is derived from detailed and reliable

68



exploration, sampling and testing gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes, and is sufficient to confirm geological and grade (or quality) continuity between points of observation where data and samples are gathered.

MRDA: Mineral Resources Development Act (1995).

Ordinary Kriging: A geostatistical method of grade interpolation.

Plunge: An inclination of a surface or axis to the horizontal.

Pyrite: FeS2, iron sulphide, can be associated with Au.

Sphalerite: (Zn, Fe)S, zinc sulphide, the chief ore mineral of zinc.

Tailings: By-products left over from mining and ore processing plants, usually discharged as a wet slurry into a tailings dam.

Tailings Dam: An earth-fill embankment dam used to store by-products of mining operations after separating the ore from the gangue.

Variogram: A statistical function describing the degree of spatial dependence.

XRF: X-ray fluorescence, an analytical technique to obtain values for the concentration of elements of interest.

Sections 6 to 12

AMD: Acid and Metalliferous Drainage

A$: Australian dollars

CRM: Certified Reference Material

CSV: Comma Separated Values

DCF: Discounted cash flow

DTM: Digital Terrain Model

EMP: Environmental Management Plan

EPA: Environmental Protection Authority

FY: Financial year

IP: Intellectual property

LOM: Life of mine

LOR: Life of Reserve

ML: Mining Lease

NPV: Net present value

QAQC: Quality assurance/quality control

PCE: Permit Environmental Condition

ROM: Run of mine

SG: Specific gravity

USGS: United States Geological Survey

YTD: Year to Date

69



16 Abbreviations and Units of Measurement

Sections 1 to 5

QAQC quality assurance/quality control (for sampling and assaying)

g/t grams per tonne

ha hectares

km kilometres

km2 square kilometres

kt/a thousands of tonnes a year, kt/yr

Mt millions of tonnes

m metre(s)

% percent

ppm parts per million; 1 ppm = 1 g/t

Sections 6 to 12

° degrees of angle

°C degrees Celsius

Au gold

dmt dry metric tonne(s)

g gram(s)

koz kilo-ounce(s)

kt kilo-tonne(s)

M million(s)

m3 cubic metre(s)

Ma million years ago

mE metres east

mm millimetre(s)

mN metres north

Mt/a million metric tonnes per annum

oz ounces

RL Relative level

mt metric tonne(s)

toz Troy ounces

t/a metric tonnes per annum

US$ United States dollars

UTM Universal Transverse Mercator coordinate system

wmt wet metric tonne(s)

Appendix 1: JORC Code (2012 Edition) – Table 1

Section 1 Sampling Techniques and Data (Criteria in this section apply to all succeeding sections)

Criteria JORC Code explanation Commentary

Sampling techniques

• Nature and quality of sampling (e.g. cut channels, random chips, or specific specialised industry standard measurement tools appropriate to the minerals under investigation, such as downhole gamma sondes, or handheld XRF instruments, etc). These examples should not be taken as limiting the broad meaning of sampling.

• Include reference to measures taken to ensure sample representivity and the appropriate calibration of any measurement tools or systems used.

• Aspects of the determination of mineralisation that are Material to the Public Report.

• In cases where ‘industry standard’ work has been done this would be relatively simple (e.g. ‘reverse circulation drilling was used to obtain 1 m samples from which 3 kg was pulverised to produce a 30-g charge for fire assay’). In other cases, more explanation may be required, such as where there is coarse gold that has inherent sampling problems. Unusual commodities or mineralisation types (e.g. submarine nodules) may warrant disclosure of detailed information.

• Sampling was carried out using Vibracore drilling, with full core dispatched for assaying.

• Assaying was carried out by AMMTEC Burnie Research Laboratory and Becquerel Laboratories (Sydney).

• AMMTEC used XRF technique to analyse for the suite of elements, with Au assayed by fire assay. Becquerel used Neutron Activation Analysis (NAA) to assay for Au as a check against the fire assay results.

Drilling techniques

• Drill type (e.g. core, reverse circulation, open-hole hammer, rotary air blast, auger, Bangka, sonic, etc) and details (e.g. core diameter, triple or standard tube, depth of diamond tails, face-sampling bit or other type, whether core is oriented and if so, by what method, etc).

• Sampling was carried out by Nick Poltock Field Exploration, using Vibracore drilling, which involved lowering a vibrating core tube into soft sediments. The vibrations provide energy for rearranging the particles within the sediment in such a way that the core tube penetrates under the static weight of the vibracoring apparatus. The Vibracore drill rig was mounted on a barge, and moored above each drillhole.

• No core diameters were documented in the provided data, however clear PVC piping is typically used to capture the core samples.

Drill sample recovery

• Method of recording and assessing core and chip sample recoveries and results assessed.

• Measures taken to maximise sample recovery and ensure representative nature of the samples.

• Whether a relationship exists between sample recovery and grade and whether sample bias may have occurred due to preferential loss/gain of fine/coarse material.

• No sample recovery data was recorded during sampling. An early analysis of sample data revealed that equal drill sample lengths do not necessarily equate to equal sample volumes. Records indicate that the drilling contractor added varying quantities of water to the samples during drilling in order to get sample return, especially from the deeper, more consolidated (and drier) tails sediments.


Logging • Whether core and chip samples have been geologically and geotechnically logged to a level of detail to support appropriate Mineral Resource estimation, mining studies and metallurgical studies.

• Whether logging is qualitative or quantitative in nature. Core (or costean, channel, etc) photography.

• The total length and percentage of the relevant intersections logged.

• There are no records of the samples having been geologically logged. All samples would have contained a mixture of gangue sands and sulphide minerals (based upon the sample assays), with a variability of concentrations not discernible to the human eye.

• All drill samples were assayed.

Sub-sampling techniques and sample preparation

• If core, whether cut or sawn and whether quarter, half or all core taken.

• If non-core, whether riffled, tube sampled, rotary split, etc and whether sampled wet or dry.

• For all sample types, the nature, quality and appropriateness of the sample preparation technique.

• Quality control procedures adopted for all sub-sampling stages to maximise representivity of samples.

• Measures taken to ensure that the sampling is representative of the in-situ material collected, including for instance results for field duplicate/second-half sampling.

• Whether sample sizes are appropriate to the grain size of the material being sampled.

• The complete core sample was removed from the PVC core and placed into plastic buckets, then sealed and individually numbered.

• The sample preparation technique is considered appropriate for the style of mineralisation.

• All samples were wet and unconsolidated sediments.

• Drilling in 1998 collected samples generally of 1 m, 1.5 m and 2 m lengths. Drilling in 2000 collected samples ranging from 2.5 m to 10 m lengths. Reasons for the varying sample lengths were not documented at time of drilling, however it is believed sample compositing took place at the drill rig during the 2000 drilling campaign.

• There are no records of field duplicates or field repeat samples having been taken. Assay results show low variability and it is the Competent Person’s opinion that sample duplicates would not have provided additional quality assurance to the data. Sample assays are also of similar tenor to the tails assays as recorded by the project operators at time of ore processing.

• Sample sizes are considered appropriate to the grain size of the material being sampled.

Quality of assay data and laboratory tests

• The nature, quality and appropriateness of the assaying and laboratory procedures used and whether the technique is considered partial or total.

• For geophysical tools, spectrometers, handheld XRF instruments, etc, the parameters used in determining the analysis including instrument make and model, reading times, calibrations factors applied and their derivation, etc.

• Nature of quality control procedures adopted (e.g. standards, blanks, duplicates, external laboratory checks) and whether acceptable levels of accuracy (i.e. lack of bias) and precision have been established.

• Assaying was done by XRF method which is considered appropriate for the style of mineralisation. The technique is considered total.

• There are no records of certified standards or blanks used during the sampling. There are no records for laboratory repeats.

• Other QAQC information is limited, including comparisons of Au assays by fire assay and NAA; and a comparison of Ag assays by XRF and NAA.

• All sample assays are consistent in grade, within expected limits, of the grade of tailings deposited as monitored during deposition of tails.


Verification of sampling and assaying

• The verification of significant intersections by either independent or alternative company personnel.

• The use of twinned holes.

• Documentation of primary data, data entry procedures, data verification, data storage (physical and electronic) protocols.

• Discuss any adjustment to assay data.

• All drill intersections were reviewed by alternative Western Metals personnel at time of data receipt.

• Holes TA12 and TA51 were drilled 10 m apart and demonstrate similar Zn assay grades from similar depths downhole.

• Drill data was stored in un-secure spreadsheets. Collar and assay data are stored. All drillholes are vertical and of shallow depth therefore no downhole surveys were taken. Geological logs were not taken.

• No adjustments have been recorded as having taken place.

Location of data points

• Accuracy and quality of surveys used to locate drill holes (collar and down-hole surveys), trenches, mine workings and other locations used in Mineral Resource estimation.

• Specification of the grid system used.

• Quality and adequacy of topographic control.

• Eastings and northings of drill collars were surveyed using Differential Global Positioning System (DGPS), recorded during drilling activities. Collar elevation was recorded as the level of the surface of the tails at the time of drilling.

• A local grid was used for all survey records.

• All dam structures and features were surveyed using Differential GPS; this included the original topography pre-filling with tails and the dam wall.

Data spacing and distribution

• Data spacing for reporting of Exploration Results.

• Whether the data spacing and distribution is sufficient to establish the degree of geological and grade continuity appropriate for the Mineral Resource and Ore Reserve estimation procedure(s) and classifications applied.

• Whether sample compositing has been applied.

• The holes drilled in 1998 were drilled on 100 m spacing (easting and northing), and infill drilling was to 50 m x 50 m. The holes drilled in 2000 were drilled on 100 m x 100 m spacing, infilled to 70 m x 70 m spacing. Both sets of drill data (by year of drilling) overlap.

• The data spacing is sufficient to support geological and grade continuity for the Mineral Resource estimate (MRE).

• Sample compositing is believed to have taken place at the drill rig during the 2000 drilling campaign.

Orientation of data in relation to geological structure

• Whether the orientation of sampling achieves unbiased sampling of possible structures and the extent to which this is known, considering the deposit type.

• If the relationship between the drilling orientation and the orientation of key mineralised structures is considered to have introduced a sampling bias, this should be assessed and reported if material.

• All holes were drilled orthogonal to the interpreted directions of geological and grade continuity.

• No sampling bias is perceived based upon the orientation of key mineralisation structures and the drilling orientation.

Sample security • The measures taken to ensure sample security. • Chain of custody for sample security is believed to have been maintained by Western Metals Resources at the time of drilling. All samples were ticketed and then transported by truck to the analytical laboratory in Burnie, where they came under the security of the lab.

Audits or reviews • The results of any audits or reviews of sampling techniques and data. • Mineral Resources were previously estimated by AMC (2006, 2009) and Polymetals Mining Services (2006), who independently reviewed the data. All the Mineral Resource estimates were classified partly as Measured, providing a high level of confidence in the sampling techniques and data.

Section 2 Reporting of Exploration Results (Criteria listed in the preceding section also apply to this section)


Mineral tenement and land tenure status

• Type, reference name/number, location and ownership including agreements or material issues with third parties such as joint ventures, partnerships, overriding royalties, native title interests, historical sites, wilderness or national park and environmental settings.

• The security of the tenure held at the time of reporting along with any known impediments to obtaining a licence to operate in the area.

• The Hellyer Tails Facility is located on consolidated granted mining lease CML 103M/1987 which is held by Hellyer Gold Mines Pty Ltd, a wholly owned subsidiary of NQ Minerals (NQM). NQM has contracted to purchase Ivy Resources, the shareholders of which will retain a 29.9% equity in NQM. The lease was granted for a period of 32 years from 1 February 1988 to 30 June 2020. An annual permit fee of $44,645.40 has been paid in full until 15 February 2018.

Exploration done by other parties

• Acknowledgment and appraisal of exploration by other parties. • The Hellyer polymetallic orebody was discovered in 1983 by Aberfoyle Ltd. Underground mining commenced in 1989 under the ownership of Aberfoyle, who later sold the operation to Western Metals Resources Ltd. Mining ceased in 2000 when economically available ore became exhausted. The mine produced 15 million tonnes (Mt) of ore in the 11 years it was in operation that yielded 601,000 t of bulk concentrate, 2.7 Mt of zinc concentrate and 728,000 t of lead concentrate.

• A total of 11 Mt of mine tailings were deposited grading 2.7 g/t Au and 3% Zn during the life of mining. Feasibility studies were undertaken and a demonstration plant constructed on site to assess the viability of mining tailings left over from the earlier mining activity.

• The Tailings dam was drilled in 1998 and 2000, during which tails were continuing to be deposited.

• The refurbishment of the Hellyer ore processing plant, following the closure of the mine, was completed in 2007 by the new owner of the project, Intec Limited, and this allowed the retreatment of the tailings to go ahead as planned. An amount of 2.026 Mt of tailings were put through the upgraded plant from 2006 to 2008, with 1.92 Mt of tails redeposited. When the tailings treatment program ceased in 2008 the processing plant was sold to Bass Metals Pty Ltd.

Geology • Deposit type, geological setting and style of mineralisation. • The deposit is a tailings dam, containing tailings from the processing of ore mined from the Hellyer polymetallic mine. The Hellyer deposit is a volcanic-hosted polymetallic massive sulphide deposit located with the Mount Read Volcanic Arc of western Tasmania, which also hosts deposits such as Mount Lyell, Que River, Henty and Roseberry. Mineralisation is comprised predominantly of pyrite and sphalerite, with lesser galena and arsenopyrite.

• The tails are predominantly crushed and ground waste products from the processed ore, with the bulk of the volume being sands, with lesser amounts of sulphides and some free metals.

• The tailings were deposited into a natural depression approximately 1 km to the west of the mill. The tails material is covered by over 1 m of water to control oxidation of the sulphide species.


Drillhole information

• A summary of all information material to the understanding of the exploration results including a tabulation of the following information for all Material drillholes:

o easting and northing of the drillhole collar

o elevation or RL (Reduced Level – elevation above sea level in metres) of the drillhole collar

o dip and azimuth of the hole

o down hole length and interception depth

o hole length.

• If the exclusion of this information is justified on the basis that the information is not Material and this exclusion does not detract from the understanding of the report, the Competent Person should clearly explain why this is the case.

• Exploration results are not being reported in this document. All drillhole information was incorporated in the MRE.

Data aggregation methods

• In reporting Exploration Results, weighting averaging techniques, maximum and/or minimum grade truncations (e.g. cutting of high grades) and cut-off grades are usually Material and should be stated.

• Where aggregate intercepts incorporate short lengths of high grade results and longer lengths of low grade results, the procedure used for such aggregation should be stated and some typical examples of such aggregations should be shown in detail.

• The assumptions used for any reporting of metal equivalent values should be clearly stated.

• Exploration results are not being reported in this document.

Relationship between mineralisation widths and intercept lengths

• These relationships are particularly important in the reporting of Exploration Results.

• If the geometry of the mineralisation with respect to the drill hole angle is known, its nature should be reported.

• If it is not known and only the down hole lengths are reported, there should be a clear statement to this effect (e.g. ‘down hole length, true width not known’).

• Exploration results are not being reported in this document. Holes were drilled vertically and are orthogonal to the stratification of the tails mineralisation.

Diagrams • Appropriate maps and sections (with scales) and tabulations of intercepts should be included for any significant discovery being reported These should include, but not be limited to a plan view of drill hole collar locations and appropriate sectional views.

• Maps and sections are presented in the body of the report.

Balanced reporting

• Where comprehensive reporting of all Exploration Results is not practicable, representative reporting of both low and high grades and/or widths should be practiced to avoid misleading reporting of Exploration Results.

• Exploration results are not being reported in this document. All drillhole information was incorporated in the MRE.


Other substantive exploration data

• Other exploration data, if meaningful and material, should be reported including (but not limited to): geological observations; geophysical survey results; geochemical survey results; bulk samples – size and method of treatment; metallurgical test results; bulk density, groundwater, geotechnical and rock characteristics; potential deleterious or contaminating substances.

• Vibracore drilling samples were collected in 2016 as part of a metallurgical testwork program. Samples of 1 m in length were assayed at ALS in Burnie, with assay results corroborating the sample grades from the 1998/2000 vibracore drilling and estimated block grades.

• Ultrasound survey in 2017 has provided contours for the top of tails surface. No elevation was provided with the data and is unable to be used to refine the Mineral Resource model. However, contour maps corroborate the depth to tails as built into the Mineral Resource model.

• No other substantive data related to the Mineral Resource estimate has been collected by NQM or the previous owners, since the MRE was prepared in 2010.

Further work • The nature and scale of planned further work (e.g. tests for lateral extensions or depth extensions or large-scale step-out drilling).

• Diagrams clearly highlighting the areas of possible extensions, including the main geological interpretations and future drilling areas, provided this information is not commercially sensitive.

• No further work is planned to add data to the MRE.

Section 3 Estimation and Reporting of Mineral Resources (Criteria listed in section 1, and where relevant in section 2, also apply to this section)


Database integrity

• Measures taken to ensure that data has not been corrupted by, for example, transcription or keying errors, between its initial collection and its use for Mineral Resource estimation purposes.

• Data validation procedures used.

• No drill samples were geologically logged. The original collar survey records and laboratory assay certificates have not been cited and there are no records from the previous MREs of this data being verified. Laboratory assay certificates would have been delivered to the client following approval for release by the lab manager using data transmission protocols available at the time, and imported into geological modelling software.

• Data validation procedures included checks for absent collar data, multiple collar entries, questionable downhole survey results, absent survey data, overlapping intervals, negative sample lengths, and sample intervals which extended beyond the hole depth defined in the collar table. Three holes from the 2000 drilling program have collars but no assays.

Site visits • Comment on any site visits undertaken by the Competent Person and the outcome of those visits.

• If no site visits have been undertaken indicate why this is the case.

• The Competent Person (Mineral Resources) visited site on 17 July 2017, and noted the general layout of the Tailings Dam facility, the current water level with respect to the tails deposit, and held discussions with the resident caretaker on the history of, and current plans for the project.


Geological interpretation

• Confidence in (or conversely, the uncertainty of) the geological interpretation of the mineral deposit.

• Nature of the data used and of any assumptions made.

• The effect, if any, of alternative interpretations on Mineral Resource estimation.

• The use of geology in guiding and controlling Mineral Resource estimation.

• The factors affecting continuity both of grade and geology.

• The deposit is a tailings dam, comprised of sediments pumped into a natural depression with an earthenware dam. The tails sediments were sourced from the adjacent Hellyer Mine processing facility which processed polymetallic ore sourced from the Hellyer underground mine. The tails are sands with a minor amount of sulphide species, but sufficient in quantity to define a Mineral Resource. The sands were deposited laterally across the tails facility over the course of the Hellyer mine life (1989 to 2003). The tails sediments are covered by a 1.5 m depth of water to prevent the oxidation of the sulphides, thus preventing the formation of acidic groundwater and consequent potential seepage into the local river systems.

• There is a high confidence in the geological interpretation used to constrain the Mineral Resource.

• The tails are constrained within a natural depression and behind an earthen dam, which were surveyed prior to tails deposition. Digital terrain models (DTMs) were constructed for the topography. Drillhole data was used to locate the positions of the sample data.

• It is assumed the drill collar surveys are as originally recorded. CSA Global do not have reason to suspect any adjustments have been made to the collar surveys.

• No alternative geological interpretations have been considered.

• Some stratification of the tails sediments was noted, based upon the grades of the main elements under consideration, and the grade interpolation attempted to honour this stratification.

• Tails sere deposited according to the location of the discharge points, which would have varied over the life of the tailings facility during tails discharge. This will have allowed deposition of tails with varying grades of metals, dependent upon the performance of the processing facility at the time (recoveries of ore).

Dimensions • The extent and variability of the Mineral Resource expressed as length (along strike or otherwise), plan width, and depth below surface to the upper and lower limits of the Mineral Resource.

• The tailings dam is of irregular dimensions, with the tails having been deposited into a natural depression. The dimensions of the tails deposit are approximately 750 m x 750 m. The depth of tails varies between 1 m and 21 m.

Estimation and modelling techniques

• The nature and appropriateness of the estimation technique(s) applied and key assumptions, including treatment of extreme grade values, domaining, interpolation parameters and maximum distance of extrapolation from data points. If a computer assisted estimation method was chosen include a description of computer software and parameters used.

• The availability of check estimates, previous estimates and/or mine production records and whether the Mineral Resource estimate takes appropriate account of such data.

• The assumptions made regarding recovery of by-products.

• A DTM of the pre-depositional surface was constructed, which acts as a base for the tails MRE. Wireframe solids for the dam walls (Main Dam, Western Arm and Eastern Arm dams) were also constructed. The geological models were constructed using Datamine, and geostatistics carried out using GeoAccess Professional and Snowden Supervisor software. No grade domaining was carried out; all the tails are contained within the one geological domain for the purposes of grade interpolation and Mineral Resource reporting.

• Top cuts were not used to cap high grade data due to the absence of very high-grade data. The high-grade metal was recovered during processing of the primary ore prior to tails deposition.


• Estimation of deleterious elements or other non-grade variables of economic significance (e.g. sulphur for acid mine drainage characterisation).

• In the case of block model interpolation, the block size in relation to the average sample spacing and the search employed.

• Any assumptions behind modelling of selective mining units.

• Any assumptions about correlation between variables.

• Description of how the geological interpretation was used to control the resource estimates.

• Discussion of basis for using or not using grade cutting or capping.

• The process of validation, the checking process used, the comparison of model data to drill hole data, and use of reconciliation data if available.

• A statistical analysis of the assay data suggested an 8 m composite length be used to composite the sample data into similar sample lengths, with raw sample lengths varying between 1.5 m and 8 m. A vertical zonation of grade was noted, with Zn, Pb and Cu increasing in grade with depth. Au and Ag grades did not present a noticeable increase in grade with depth. Variograms for Zn, Pb, Cu, Au, Ag and As were modelled, all showing low relative nugget effects (between 10% and 20%), but with variable ranges.

• A block model with parent cell sizes 25 mE x 25 mN x 5 mRL was employed, with sub-celling to 5 mE x 5 mN x 2.5 mRL.

• Grade estimation was by ordinary kriging (OK). A search ellipse with radii 100 m (along strike) by 5 m (down dip) by 100 m (across strike) was used, with minimum and maximum samples per block estimate of 2 and 6 respectively. The low number of samples used is due to the larger number of samples contained within a single 8 m composite. A maximum of 2 samples per drillhole per block estimate was used. Cell discretisation of 3 x 3 x 1 was utilised. The primary search ellipse radii were increased by 50% and then by 10 times if a parent cell was not able to be interpolated. Search ellipse orientations were flat.

• This MRE is of similar tonnage and grade to previous MREs.

• No assumptions have been made regarding the recovery of by-products.

• Arsenic was modelled as a deleterious product.

• The block size is approximately half the typical drill spacing of the well drilled areas.

• No assumptions were made regarding selective mining units.

• Some correlation is observed between some of the metals, although each element was individually interpolated into the model.

• The flat repose of the tails strata required a flat search ellipse to be used to interpolate the block grades.

• The grade model was validated by: (1) creating slices of the model and comparing to drillholes on the same slice; (2) swath plots comparing average block grades with average sample grades on nominated easting, northing and elevation slices; and (3) mean grades per domain for estimated blocks and flagged drillhole samples. Each validation step complemented the others.

• The tonnage weighted and accumulated block grades were compared to mill tailings records and compare favourably.

Moisture • Whether the tonnages are estimated on a dry basis or with natural moisture, and the method of determination of the moisture content.

• Tonnages are based upon a dry basis.


Cut-off parameters

• The basis of the adopted cut-off grade(s) or quality parameters applied. • No cut-off grade is used to report the MRE. All blocks within the block model are reported.

Mining factors or assumptions

• Assumptions made regarding possible mining methods, minimum mining dimensions and internal (or, if applicable, external) mining dilution. It is always necessary as part of the process of determining reasonable prospects for eventual economic extraction to consider potential mining methods, but the assumptions made regarding mining methods and parameters when estimating Mineral Resources may not always be rigorous. Where this is the case, this should be reported with an explanation of the basis of the mining assumptions made.

• The tails deposit will likely be mined using a dredging operation, due to the deposit covered by at least 1.5 m of water.

Metallurgical factors or assumptions

• The basis for assumptions or predictions regarding metallurgical amenability. It is always necessary as part of the process of determining reasonable prospects for eventual economic extraction to consider potential metallurgical methods, but the assumptions regarding metallurgical treatment processes and parameters made when reporting Mineral Resources may not always be rigorous. Where this is the case, this should be reported with an explanation of the basis of the metallurgical assumptions made.

• Recent metallurgical testwork was completed on 10 composited samples obtained from vibracore drilling of the Tails dam. The following initial results were obtained:

• A silver/copper/lead flotation concentrate at 36-38% lead grade can be attained at variable recoveries from 26-54% Pb.

• A zinc concentrate of 40-42% Zn grade was readily attained at recoveries from 45-68% for the samples tested.

• A gold/silver/pyrite concentrate containing 46% Sulphur is readily achievable at high recoveries (50-70%).

• Arsenic grade reporting to the concentrates is variable and exceeded the 1% target limit for some of the lead tests, half of the pyrite tests and for the zinc flotation tests where the grade was below 48% Zn grade.

Environmental factors or assumptions

• Assumptions made regarding possible waste and process residue disposal options. It is always necessary as part of the process of determining reasonable prospects for eventual economic extraction to consider the potential environmental impacts of the mining and processing operation. While at this stage the determination of potential environmental impacts, particularly for a greenfields project, may not always be well advanced, the status of early consideration of these potential environmental impacts should be reported. Where these aspects have not been considered this should be reported with an explanation of the environmental assumptions made.

• All tails are required to be impounded beneath water to prevent oxidation of the sulphide species, which will potentially create acid mine drainage.

• NQM have put in place plans for environmental surveys of the proposed tails storage facility to be constructed downstream from the current dam wall.

Bulk density • Whether assumed or determined. If assumed, the basis for the assumptions. If determined, the method used, whether wet or dry, the frequency of the measurements, the nature, size and representativeness of the samples.

• The bulk density for bulk material must have been measured by methods that adequately account for void spaces (vugs, porosity, etc), moisture and differences between rock and alteration zones within the deposit.

• Discuss assumptions for bulk density estimates used in the evaluation process of the different materials.

• Buckets containing samples from the 2000 drilling campaign were retained at the AMMTEC Laboratory (Burnie) where they performed 39 bulk density determinations on these samples. The results ranged from 1.30 to 2.63 t/m3 and averaged 1.93 t/m3.


Classification • The basis for the classification of the Mineral Resources into varying confidence categories.

• Whether appropriate account has been taken of all relevant factors (i.e. relative confidence in tonnage/grade estimations, reliability of input data, confidence in continuity of geology and metal values, quality, quantity and distribution of the data).

• Whether the result appropriately reflects the Competent Person’s view of the deposit.

• The Mineral Resource is classified as Measured, Indicated and Inferred, in accordance with the JORC Code (2012 Edition).

• Classification was based upon drill spacing, and considerations regarding reasonable prospects for the eventual economic extraction of the tails. The area where drillholes are spaced at 50 m x 50 m intervals was classified as Measured Mineral Resources. The remainder of the tails within the Main Tails Dam, Western and Eastern Arms were classified as Indicated Mineral Resources.

• The Tails contained within the Shale Pit are classified as Inferred Mineral Resources. The block grades are estimated based upon a back calculation of assays from reprocessed tails using calculated recoveries. No sampling has been carried out to date of the Shale Pit tails.

• The volume of tails which were dredged in the period 2006 to 2008 are retained in the block model but are not classified.

• The classification of the Mineral Resource appropriately reflects the Competent Person’s view of the deposit.

Audits or reviews • The results of any audits or reviews of Mineral Resource estimates. • There have been at least six MREs prepared for the Hellyer Tails project, including the current one. Previous MREs were either upgrades to earlier models, or a revised approach to the grade interpolation. Resource reports supporting the latter included brief discussions on previous MREs and the basis for a change in strategy for estimating the Mineral Resources.

• Earlier block models did not include vertical grade zonation, whilst latter resource models did include the perceived zonation.

• All models since 2006 have been either wholly or partially classified as Measured Mineral Resources.

Discussion of relative accuracy/ confidence

• Where appropriate a statement of the relative accuracy and confidence level in the Mineral Resource estimate using an approach or procedure deemed appropriate by the Competent Person. For example, the application of statistical or geostatistical procedures to quantify the relative accuracy of the resource within stated confidence limits, or, if such an approach is not deemed appropriate, a qualitative discussion of the factors that could affect the relative accuracy and confidence of the estimate.

• The statement should specify whether it relates to global or local estimates, and, if local, state the relevant tonnages, which should be relevant to technical and economic evaluation. Documentation should include assumptions made and the procedures used.

• These statements of relative accuracy and confidence of the estimate should be compared with production data, where available.

• The Mineral Resource accuracy is communicated through the classification assigned to this Mineral Resource. The MRE has been classified in accordance with the JORC Code, 2012 Edition using a qualitative approach. All factors that have been considered have been adequately communicated in Section 1 and Section 3 of this Table.

• The Mineral Resource statement is a global estimate. No domaining of grade has taken place and all classified blocks in the tails model are reported.

• The tails were deposited as a discharge from the Hellyer ore processing plant, with annual records for tonnes and grade of tails as deposited. The block model reconciles well against this production data.

• Some of the tails were dredged and treated in the period 2006-2008, and production records demonstrate good reconciliation.

Section 4 Estimation and Reporting of Ore Reserves (Criteria listed in section 1, and where relevant in sections 2 and 3, also apply to this section)


Mineral Resource estimate for conversion to Ore Reserves

• Description of the Mineral Resource estimate used as a basis for the conversion to an Ore Reserve.

• Clear statement as to whether the Mineral Resources are reported additional to, or inclusive of, the Ore Reserves.

• The Mineral Resources of Hellyer TSF were estimated by a team of CSA Global Pty Ltd, which includes Peter Davies, Peter Cranfield and David Williams in 2010. David Williams did an update of the Mineral Resource estimates in August 2017. The Mineral Resource estimate with full recovery of tailings and using the CSA Global resource model as a Datamine file is as follows:

Classification Tonnes (Mt) Zn % Pb % Au g/t

Measured 2.05 3.31 3.35 2.63

Indicated 5.99 2.29 2.95 2.55

Inferred 1.21 1.00 2.60 2.57

Total 9.25 2.35 2.99 2.57

Classification Cu % Ag g/t As %

Measured 0.20 94 1.44

Indicated 0.18 93 1.42

Inferred 0.19 86 1.26

Total 0.19 92 1.40

Notes:

• Figures in the tables may not sum due to rounding

• The Mineral Resource is reported as wholly inclusive of the Ore Reserve.

Site visits • Comment on any site visits undertaken by the Competent Person and the outcome of those visits.

• If no site visits have been undertaken indicate why this is the case.

• A site visit was made by Dr Rossen Halatchev, Project Manager for this Ore Reserve work and Principal Consultant for the Ore Reserve estimation.

• His observations confirm the physical presence of the tailings dam, mining equipment (dredge), pipelines, treatment plant and all facilities as outlined in the accompanying report by AusGEMCO and other studies detailed below. In addition, CSA Global Principal Geologist David Williams visited the Project on July 2017.

Study status • The type and level of study undertaken to enable Mineral Resources to be converted to Ore Reserves.

• The Code requires that a study to at least Prefeasibility Study level has been undertaken to convert Mineral Resources to Ore Reserves. Such studies will have been carried out and will have determined a mine plan that is technically achievable and economically viable, and that material Modifying Factors have been considered.

• A prior Prefeasibility Study was undertaken by COMO Engineers Pty Ltd in 2013, which was updated in 2015.

• A Feasibility Study was conducted by Ivy Resources Pty Ltd in 2013.

• As part of the Ivy Resources Feasibility Study, Como Engineers Pty Ltd (Como) completed a Definitive Feasibility Study (DFS) in 2013 with a focus only on a closely defined scope related to direct cyanide leach and leaching of the Albion process product. Como updated the prior DFS in 2015 and included costs for various feed rates to the cyanide leach circuit and subsequent processes.


• A mining study by CSA Global in 2010 was conducted dealing with mining and metallurgical assumptions. The study determined a mine plan which is technically achievable and economically viable based on some mining assumptions.

• Polymetals Pty Ltd, as a former owner of Hellyer TSF, conducted profitable dredging operations in the Main Dam for the period 2006-2008 by mining about 2.0 Mt of tailings and processing them in the currently existed processing plant.

Cut-off parameters

• The basis of the cut-off grade(s) or quality parameters applied. • No cut-off grades have been applied because the tailings extraction does not allow a physical separation of the material into grade categories. Tailings are dredged and the whole lot is pumped to the processing plant for reprocessing.

• Quality parameters are not applied due to the specificity of the tailings extraction by dredging. These parameters are normally used for the development of Geological Block Profit Model and preliminary assessment of the economic potential of the tailings retreatment of Hellyer TSF.

Mining factors or assumptions

• The method and assumptions used as reported in the Pre-Feasibility or Feasibility Study to convert the Mineral Resource to an Ore Reserve (i.e. either by application of appropriate factors by optimisation or by preliminary or detailed design).

• The choice, nature and appropriateness of the selected mining method(s) and other mining parameters including associated design issues such as pre-strip, access, etc.

• The assumptions made regarding geotechnical parameters (eg pit slopes, stope sizes, etc), grade control and pre-production drilling.

• The major assumptions made and Mineral Resource model used for pit and stope optimisation (if appropriate).

• The mining dilution factors used.

• The mining recovery factors used.

• Any minimum mining widths used.

• The manner in which Inferred Mineral Resources are utilised in mining studies and the sensitivity of the outcome to their inclusion.

• The infrastructure requirements of the selected mining methods.

• An assumption was used for the mining factors in the Prefeasibility and Feasibility studies that tailings are calculated as dry metric tonnes (dmt) without loss and dilution. The mining method used was dredging.

• The current Ore Reserve work includes a detailed pit design of all dams (Main Dam, Finger Pond Dam, Western Arm Dam, and Shale Pit Dam). The pit design was optimized with the utilisation of the results of Geological Block Profit models. The design is based on the optimum dredge path (strip) direction which was assessed with the Block Unit Profit criterion measured in US$/dmt. Each pit is designed with benches allowing dredging to a depth of 10.5 m and to 16 m with the utilisation of a special dredge frame extension. The design also includes discretization of each bench into paths (strips) and blocks of 30 m width and 50 m length. The mine sequence is organised with panels where each panel covers a few strips and blocks. A working dredging zone of each pit was designed outsides of the buffer zones from the dams’ walls, which is a requirement for ensuring safe and normal conditions of dredging. A minimum dredging depth of 1.5 m was used for contouring the working zones. Two scenarios of the mine sequence of Main Dam were developed – Base Case Scenario and Scenario 300. The first scenario deals with the design of buffer zones of 50 m from all walls and full extraction of the tailings, while the second scenario uses a buffer zone of 300 m only from the Main Dam wall and partial extraction of the tailings within a depth of 16 m from water level.

• Dredging is adopted as the major mining method, which proved its advantages during the previous period 2006–2008 of dredging the tailings by Polymetals Pty Ltd. The dredge is currently located on the mine site and it is in very good condition allowing its further utilisation after a scheduled maintenance. Another method is hydraulic mining with slurry pump for the tailings at the bottom of


Western Arm Dam and Shale Pit Dam. The purpose of utilising the method of hydraulic mining is the transfer of tailings from Western Arm Dam and Finger Pond Dam to the Main Dam during the operations at the dam bottom. Main Dam favours the method of dredging because of the configuration of the dam bottom terrain. Another method is water cannon wash of the tailings left on the bottom of upper benches with water cannons. These tailings will be directed to the working dredge zone of each dam. The extraction of the lower benches will be done by lowering the water level in each dam within the depth of the upper benches allowing further dredging operations.

• There are four dam walls in Hellyer TSF. They were designed and constructed in the past. Each of them is periodically inspected and there are no symptoms so far indicating processes of deformation. A risk plan is proposed in this Ore Reserve work for maintaining the wall slope stability by implementing the concept of equalising the water levels in the dams during the periods of water release from the dredged benches. This will ensure a constant hydrostatic pressure on both sides of each wall, which is a major pre-condition for maintaining the wall slope stability during operations.

• There is no a major assumption used in the Mineral Resource Model of 2017. A tailings density of 1.93 t/m3 was used in the development of the Mineral Resource model.

• Mining dilution is zero with all material consisting of tailings from previous treatment plant operations. The dredge is equipped with a device for a precise control of the dredging depth and its x-y axis location.

• Mining recovery is accepted as 100% because of using water cannon washing technology, which provides maximum mining recovery.

• A minimum mining width of 30 m was used for the design of dredging strips (paths). This width was also used in the previous design and dredging operations during the 2006-2008 period.

• Inferred Mineral Resources belong to Shale Pit Dam and they are of a good value in accordance with the tailings records of the processing plant made during the exploitation of Hellyer gold mines. This minimises the risk of non-confirmation of the resource quantity and grades. They are included in the mine production schedules for extraction as there is sufficient time for an update of this resource category because the Shale Pit Dam is scheduled as the last dam to be exploited.

• Mining infrastructure is well-established following the previous dredging operations. The pipeline for transport of tailings from the dredge to the processing plant already exists on the mine site, but needs some scheduled maintenance prior to the new dredging operations.


Metallurgical factors or assumptions

• The metallurgical process proposed and the appropriateness of that process to the style of mineralisation.

• Whether the metallurgical process is well-tested technology or novel in nature.

• The nature, amount and representativeness of metallurgical test work undertaken, the nature of the metallurgical domaining applied and the corresponding metallurgical recovery factors applied.

• Any assumptions or allowances made for deleterious elements.

• The existence of any bulk sample or pilot scale test work and the degree to which such samples are considered representative of the orebody as a whole.

• For minerals that are defined by a specification, has the ore reserve estimation been based on the appropriate mineralogy to meet the specifications?

• The Hellyer Processing Plant exists at present with all technological equipment required for reprocessing of the tailings. The plant was used previously to produce concentrates from the ore supplied from the gold mines and a bulk concentrate from the tailings retreatment of the old dredging period. The original flowsheet includes three roughers for production of zinc concentrate, combined lead/silver/copper concentrate and gold/silver/pyrite concentrate. The plant is in very good condition and only requires a scheduled maintenance prior to its use in the production process.

• The metallurgical process of the Hellyer tailings retreatment uses well-tested technology. Recent tests have been conducted by ALS laboratories (ALS) in Burnie (Tasmania) and reported on by Pitt & Sherry Group (2016) for developing a new flowsheet for the processing of Hellyer tailings with the objective of optimising the recovery of all valuable components. The testwork was relatively conventional and replicates the original Hellyer flowsheet but at a finer primary and regrind size, as well as the inclusion of a pyrite float for precious metals gold and silver. 40 Batch tests were also conducted by ALS as well as 3 locked cycle tests which simulate the recirculating loads to project a balanced circuit performance. The results obtained indicate the opportunity for producing the following three concentrates:

• Combined lead/silver/copper concentrate grading 37% lead (Pb) and maximising silver content.

• Zinc concentrate grading 45% zinc (Zn).

• Gold/silver/pyrite concentrate with a 46% sulphur content.

• Based on the analysis of the test results there are a number of opportunities identified to improve understanding of the current primary metal and precious metal characteristics and assess options to optimise the process to further increase metal recovery rates.

• The metallurgical testwork undertaken is highly representative and conducted at the ASL laboratories who are known for their benchmarking approach of metallurgical testing. The results of this work were presented in an official report prepared by the Pitt & Sherry Group.

• Arsenic is the only deleterious element in the concentrate production. NQM advised that the concentrates to be produced will contain less than 1% arsenic. The arsenic is predominantly present as arsenopyrite with other arsenical minerals such as tennantite and tetrahedrite. The testwork conducted by ALS shows a range of 0.5-1.5% variation of arsenic in the lead concentrate and averaged less than 1%. An option may be to blend the lead concentrates to maintain the 1% limit. The results obtained for the zinc concentrates of around 45% zinc, indicated that the arsenic values were 0.6-1.3% and averaged around


1%. The gold/silver/pyrite concentrates contained 0.7-1.3% arsenic and averaged around 1% arsenic.

• The pilot scale test can be taken as the actual performance of the existing processing plant, which was successfully used for processing the tailings from the old dredging operations in 2006-2008. The tailings processed were dredged from a large and representative area of the Main Dam.

• The numerous laboratory tests conducted in the past and present give grounds to accept that the Ore Reserve estimation is based on an appropriate mineralogy and treatment process that meets the specifications required.

Environmental • The status of studies of potential environmental impacts of the mining and processing operation. Details of waste rock characterisation and the consideration of potential sites, status of design options considered and, where applicable, the status of approvals for process residue storage and waste dumps should be reported.

• Hellyer TSF has been the object of various environmental studies since its construction and commencement of exploitation. The focus of these studies involved many factors, such as minimum water cover on the tailings to avoid oxidation process, slope stability of the Main Dam wall and other walls, and various chemical component contents in the water in the TSF and surrounds. The previous and current management of Hellyer paid serious attention to these factors and achieved satisfactory results. There is an adequate number of reports with results from the investigation of potential impacts as well as with estimates of the components of the tailings to satisfy regulatory requirements. The environmental status of the mine site is under permanent control of the relevant Tasmanian authorities. NQM has already developed a plan for building a new tailings storage facility (TSF2) near the Main Dam wall, which will be used for storing the residuals from the reprocessing of the tailings in the processing plant and ensuring sufficient storage capacity for the new mining operations. The TSF2 is planned to be in use in May 2019. In accordance with the mine plan developed in this Ore Reserve study, there will be a pre-production period for transferring the tailings of the Finger Pond Dam into the old void in the Main Dam, which was created during the 2006-2008 dredging period. This production period will cover only Main Dam, Western Arm Dam and Shale Pit Dam. An approval for building the TSF2 has already been granted.

• The Environmental Protection Authority (EPA) of Tasmania approved the 2017 Environmental Management Plan (EMP) of Hellyer Project for implementation, which was prepared and submitted by Caloundra Environmental Pty Ltd in accordance with Condition G7 of Permit Conditions – Environmental (PCE) No. 7386, as contained in Permit No. DA 138/2006.


Infrastructure • The existence of appropriate infrastructure: availability of land for plant development, power, water, transportation (particularly for bulk commodities), labour, accommodation; or the ease with which the infrastructure can be provided, or accessed.

• The Hellyer site has extensive existing infrastructure that has been adequately maintained during the various care and maintenance periods that followed Hellyer operation’s original closure (June 2000), the cessation of tailings re-treatment (September 2008), and the more recent cessation of processing activity by Bass Metals Ltd (Bass) (July 2012).

• The Hellyer mining lease intersects Tas Networks’s 110kV and 220kV high tension power lines that supply the northern part of Tasmania. A newly upgraded 22kV substation located adjacent to the Que River Mine was recently installed and approximately 8 km of 22kV line connects the Hellyer Site where it is transformed down to 3.3kV and 415 volts for use at the plant.

• Water is drawn from the Southwell River via electric pumps located below the plateau and within the north-eastern boundary of the lease. The water is of excellent quality and the license conditions permit a maximum rate of extraction of 12.96 megalitres per day. This is enough water to meet site processing and other needs.

• Access to the site from government maintained roads is via two privately owned and maintained unsealed access roads with the preferred route being the gravel Hellyer mine site access road from the Cradle Mountain Link Road.

• There is an existing 3’6” gauge railway spur into the Hellyer plant. This 11 km section of railway line onto the mine site is owned by the Tasmanian Government and links the operation to the Port of Burnie line. This track will be used to transport the zinc, lead/copper/silver and gold/silver/pyrite concentrates to Burnie.

• There is a 100-pair cable delivering phone and fax services to site. There is also a fibre-optic server connection between the mill and the current main office/administration building that is currently connected and capable of sharing files and data. Internet access is provided by two satellite connections.

• A large, enclosed store/warehouse exists at the Hellyer Processing Plant.

Costs • The derivation of, or assumptions made, regarding projected capital costs in the study.

• The methodology used to estimate operating costs.

• Allowances made for the content of deleterious elements.

• The source of exchange rates used in the study.

• Derivation of transportation charges.

• The basis for forecasting or source of treatment and refining charges, penalties for failure to meet specification, etc.

• The allowances made for royalties payable, both Government and private.

• Capital and operating costs are provided by NQM and were discussed in detail about their validity.

• Derivation of costs by NQM was based on historical data of unit operating costs, updated for current fuel, power and labour costs and professional engineering estimates and quotations for capital expenditures.

• Historical data and current Tasmanian regulations were used in estimating the costs of transport and royalties payable.

• Operating costs are estimated in A$, whilst estimates for capital expenditure for the duration of the project are in US$.

• The source of data for these costs estimates and the associated exchange rates to


US$ are contained within a financial model generated by NQM. The Excel file is dated 28 May 2017.

• Allowance was made for the content of arsenic in concentrates.

• Allowances were made for Government royalties and refining charges

Revenue factors • The derivation of, or assumptions made regarding revenue factors including head grade, metal or commodity price(s) exchange rates, transportation and treatment charges, penalties, net smelter returns, etc.

• The derivation of assumptions made of metal or commodity price(s), for the principal metals, minerals and co-products.

• The Treatment Plant head grades of Zn, Pb, Au, Ag, and Cu were assessed in accordance with the optimization of mine sequence and production schedule and utilization of the Mineral Resource model.

• In accordance with the life-of-mine plan developed, the average annual processing throughput of tailings is 1.14 Mt/a.

• A$:US$ exchange rate of 1.32 was used as advised by NQM.

• Recovered metal values and any penalties were based on information from NQM following their market studies and price forecasts and applied to each individual element of all concentrates quoted in the financial model.

Market assessment

• The demand, supply and stock situation for the particular commodity, consumption trends and factors likely to affect supply and demand into the future.

• A customer and competitor analysis along with the identification of likely market windows for the product.

• Price and volume forecasts and the basis for these forecasts.

• For industrial minerals the customer specification, testing and acceptance requirements prior to a supply contract.

• AusGEMCO is reliant on the metal price projections advised by NQM. Prices are presented by NQM using the forecast of Consensus Economics (August 2017) (www.consensuseconomics.com):

• Zn – 2,028 USD/dmt

• Pb – 1,918 UAD/dmt

• Au – 1,232 USD/oz

• Ag – 17.70 USD/oz

• The pyrite concentrate FOB price is A$50.00/dmt.

• The total production of metals over the life-of-mine is as follows:

• Au – 680,332 oz

• Ag – 24,293,563 oz

• Pb – 246,433 t

• Zn – 208,488 t

• Cu – 14,747 t

• NQM market is oriented to customers in Asia and predominantly in China.

Economic • The inputs to the economic analysis to produce the net present value (NPV) in the study, the source and confidence of these economic inputs including estimated inflation, discount rate, etc.

• NPV ranges and sensitivity to variations in the significant assumptions and inputs.

• The technological input of the NPV assessment in terms of total production of metals over the life-of-mine is as follows:

• Au – 680,332 oz

• Ag – 24,293,563 oz

• Pb – 246,433 t

• Zn – 208,488 t

• Cu – 14,747 t

• Grade deduction of the basic metal component (Zn, Pb, Au) is used for each concentrate type in order to minimize the risk of non-confirmation of the metal

http://www.consensuseconomics.com/


grades in the Mineral Resource Model.

• Models of Ore Reserve Risk and Mining Project Risk assessment were developed to account for the project uncertainty regarding geological, mining, processing and economic factors having impact on the project NPV.

• Discount rate of 10% was used for the NPV assessment.

• The cost of tailings processing may be considered as the most sensitive in the discounted cash flow analysis.

• Grades of Zn, Pb, Au, Ag and Cu vary in a small range in the Mineral Resource model, which make Hellyer TSF a pretty homogeneous deposit of tailings. Higher grades are related to zones close to the dam bottom because of using old processing technology in the treatment plant at the commencement of the tails storage in Hellyer TSF.

• Inputs to the economic analysis included Modifying Factors as described above.

Social • The status of agreements with key stakeholders and matters leading to social license to operate.

• All key stakeholder agreements are in place. The Company has close relationships with the communities surrounding the Project.

Other • To the extent relevant, the impact of the following on the project and/or on the estimation and classification of the Ore Reserves:

• Any identified material naturally occurring risks.

• The status of material legal agreements and marketing arrangements.

• The status of governmental agreements and approvals critical to the viability of the project, such as mineral tenement status, and government and statutory approvals. There must be reasonable grounds to expect that all necessary Government approvals will be received within the timeframes anticipated in the Pre-Feasibility or Feasibility study. Highlight and discuss the materiality of any unresolved matter that is dependent on a third party on which extraction of the reserve is contingent.

• There are no known significant naturally occurring risks to the Project. A similar operation was conducted for the period 2006-2008 without any negative consequences for the environment and surrounding communities.

Classification • The basis for the classification of the Ore Reserves into varying confidence categories.

• Whether the result appropriately reflects the Competent Person’s view of the deposit.

• The proportion of Probable Ore Reserves that have been derived from Measured Mineral Resources (if any).

• Measured Mineral Resources have been converted to Proven Reserves. Indicated Mineral Resources have been converted to Probable Reserves.

• No Measured Mineral Resources were downgraded to Probable Reserves.

• The estimated Ore Reserves are, in the opinion of the Competent Person, appropriate for this style of deposit.

Audits or reviews

• The results of any audits or reviews of Ore Reserve estimates. • AusGEMCO Pty Ltd has completed a review of the Ore Reserve estimate resulting from this study.


Discussion of relative accuracy/ confidence

• Where appropriate a statement of the relative accuracy and confidence level in the Ore Reserve estimate using an approach or procedure deemed appropriate by the Competent Person. For example, the application of statistical or geostatistical procedures to quantify the relative accuracy of the reserve within stated confidence limits, or, if such an approach is not deemed appropriate, a qualitative discussion of the factors which could affect the relative accuracy and confidence of the estimate.

• The statement should specify whether it relates to global or local estimates, and, if local, state the relevant tonnages, which should be relevant to technical and economic evaluation. Documentation should include assumptions made and the procedures used.

• Accuracy and confidence discussions should extend to specific discussions of any applied Modifying Factors that may have a material impact on Ore Reserve viability, or for which there are remaining areas of uncertainty at the current study stage.

• It is recognized that this may not be possible or appropriate in all circumstances. These statements of relative accuracy and confidence of the estimate should be compared with production data, where available.

• The level of study carried out as part of this Ore Reserve is to a Feasibility Study. The relative accuracy of the estimates is reflected in the reporting of the Ore Reserves as per guidelines re: modifying factors, study levels and Competent Persons contained in the JORC 2012 Code.

• The statement relates to global estimates of tonnes and grade.

• Ore Reserve Risk and Mining Project Risk analyses were carried out. Standard deviations of all project variables were observed. The Ore Reserve Risk is assessed with standard deviations of the geological variables using the Variogram modelling results of the Mineral Resource study. The standard deviations of metal prices were also observed. Globally, the Project is susceptible to fluctuations on metal prices and processing costs.

• The relative accuracy and confidence of the Ore Reserve estimate was compared with the available production data obtained during the previous dredging period 2006 and 2008 and there is strong evidence to confirm its reliability.

hellyer tailings retreatment project, tasmania...bsc (hons), mausimm signature: coordinating author...

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