Exhibit 99.1

AMC Mining Consultants (Canada) Ltd. BC0767129   200 Granville Street, Suite 202 Vancouver BC V6C 1S4 Canada   T +1 604 669 0044 E vancouver@amcconsultants.com W amcconsultants.com  

Technical Report

Mineral Resource and Reserve Update for the Tucano Gold Mine

Great Panther Silver Limited

Amapá (Brazil)

In accordance with the requirements of National Instrument 43-101 “Standards of Disclosure for Mineral Projects” of the Canadian Securities Administrators

Qualified Persons:

B Wolfe, MAIG

N Spicer, C.Eng.

M Batelochi, MAusIMM (CP)

R Lacourt, SME - RM

J Moreno, MIEAust, CPEng.

P O’Bryan, MAusIMM (CP)

R. Walton, P.Eng.

G Methven, P.Eng.

AMC Project 718060

Effective date 11 May 2018

Report date 31 October 2018

Unearth a smarter way

Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Quality control

The signing of this statement confirms this report has been prepared and checked in accordance with the AMC Peer Review Process.

Project Manager	The signatory has given permission to use their signature in this AMC document	31 October 2018
	Gary Methven	Date
Peer Reviewer	The signatory has given permission to use their signature in this AMC document	31 October 2018
	Gary Methven	Date
Author	The signatory has given permission to use their signature in this AMC document	31 October 2018
	Carl Kottmeier	Date

Important information about this report

Confidentiality

This document and its contents are confidential and may not be disclosed, copied, quoted or published unless AMC Mining Consultants (Canada) Ltd. (AMC) has given its prior written consent.

No liability

AMC accepts no liability for any loss or damage arising as a result of any person other than the named client acting in reliance on any information, opinion or advice contained in this document.

Reliance

This document may not be relied upon by any person other than the client, its officers and employees.

Information

AMC accepts no liability and gives no warranty as to the accuracy or completeness of information provided to it by or on behalf of the client or its representatives and takes no account of matters that existed when the document was transmitted to the client but which were not known to AMC until subsequently.

Precedence

This document supersedes any prior documents (whether interim or otherwise) dealing with any matter that is the subject of this document.

Recommendations

AMC accepts no liability for any matters arising if any recommendations contained in this document are not carried out, or are partially carried out, without further advice being obtained from AMC.

Outstanding fees

No person (including the client) is entitled to use or rely on this document and its contents at any time if any fees (or reimbursement of expenses) due to AMC by its client are outstanding. In those circumstances, AMC may require the return of all copies of this document.

Public reporting requirements

If a Client wishes to publish a Mineral Resource or Ore / Mineral Reserve estimate prepared by AMC, it must first obtain the Competent / Qualified Person’s written consent, not only to the estimate being published but also to the form and context of the published statement. The published statement must include a statement that the Competent / Qualified Person’s written consent has been obtained.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

1 Summary

This report has been prepared for Great Panther Silver Limited (GPR) in connection with the proposed acquisition (the Acquisition) by GPR of Beadell Resources Limited (BDR) pursuant to a Scheme Implementation Deed entered into between GPR and BDR on September 23, 2018. AMC Mining Consultants (Canada) Ltd. (AMC) has been retained by GPR to provide this Technical Report in accordance with National Instrument 43-101 (NI 43-101) “Standards of Disclosure for Mineral Projects”, of the Canadian Securities Administrators, to support the disclosure to be made by GPR regarding BDR’s Tucano Property (Property) including the Tucano Gold Mine in the information circular for the meeting of the shareholders of GPR to be held to approve the Acquisition.

1.1 Project summary

The Tucano Gold Project is currently the 2nd largest gold producer in Brazil, producing around 145,000 oz per annum from several open pit mines in the northern state of Amapá, Brazil. Formerly a heap leaching operation, it ran from late 2005 until January 2009, producing approximately 292,000 troy ounces (oz) of gold from 8.8 million tonnes (Mt) of ore. The property was put under care and maintenance due to problems treating the increasingly transitional material. BDR acquired the property in 2010 and the process plant was upgraded with the addition of a SAG mill and carbon-in-leach (CIL) infrastructure, capable of treating sulphide ore. From November 2012 to 30 June 2017, 17.0 Mt at 1.34 g/t Au containing 735,294 oz Au were processed, feed was made up of a blend of pit oxide ore, sulphide ore, spent ore, and historic stockpiles.

GPR entered into a Scheme Implementation Deed dated 23 September 2018, that will result in the acquisition of all of the issued ordinary shares of BDR. Upon completion of the Acquisition, existing BDR and GPR shareholders are expected to own approximately 38% and 62%, respectively, of the combined company.

1.2 Location

The Tucano Property is located at latitude 0.85°N and longitude 52.90°W in Amapá, the most north-eastern state of Brazil, approximately 15 km from the town of Serra do Navio. The mine is approximately 200 km from Macapá, the state capital, and is accessible by the Brazilian federal highway BR-210, or by chartered aircraft.

1.3 Ownership

On completion of the Scheme Implementation Deed BDR will be wholly owned by GPR and through the wholly owned subsidiary, Beadell Brasil Limitada (Beadell Brasil), GPR will hold a 100% interest in the Tucano Property.

1.4 Geology

The Tucano mineral deposits are located within the Guyana Craton, described by several authors as the Maroni - Itacaiunas mobile belt. This belt runs from Pará and Amapá States of northern Brazil through the Guyanas and into Venezuela. The regional structure is marked by a north-western trending foliation parallel to the main lithologic contacts. The western part of the Tucano property (about 25% of the area) is underlain by basement gneiss. The balance of the property area is underlain by ortho-amphibolite and meta-sedimentary rocks of the Vila Nova Group (1.75 to 1.9 billion years old), composed essentially of gneisses, granites, amphibolites, carbonates / calc-silicates, banded iron formations (BIF), schists, and quartzite. These units are intruded by granitic pegmatites, diabase dykes, and gabbro bodies.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

1.5 Mineralization

Mineralization at Tucano is controlled by a major north-south shear zone. The texture and mineralogy along the shear zone indicates high-temperature hydrothermal alteration. This zone exhibits intense hydrothermal alteration, particularly silicification and sulfidation, bearing auriferous pyrrhotite, and pyrite. The alteration is most intense in the proximity of reactive meta-sediments such as BIF and carbonate and, to a lesser extent in calc-silicates, schists, and amphibolites. Late cross cutting intrusions are typically barren.

1.6 History

Anglo American discovered a mineralized shear zone in 1994 and undertook extensive exploration between 1995 and 2002. AngloGold completed the initial feasibility study of the oxide resources in October 2002 and from this time until the acquisition of the project by BDR in 2010, regional exploration outside of the Tucano mining concession involved ground IP surveys, mapping, stream, soil, and auger sampling. In 2008 - 2009 a campaign of diamond drilling was undertaken in some of the regional exploration targets using a man-portable modular diamond drill rig.

1.7 Drilling and exploration

After acquisition of the Property, BDR undertook an extensive resource delineation diamond drilling within the mining concession to improve and expand the resource base as part of the Feasibility Study (FS) for the CIL plant expansion. In 2012 BDR acquired a new T60 Schramm Reverse Circulation (RC) rig from Australia which was later sold to McKay drilling and continued operating on site under contract. The T60 Schramm is the main tool used for grade control drilling and resource extension holes, up to 400 m deep, and until now, has only been used to a limited extent to test regional exploration targets. Shallow testing of regional targets is currently done using motor driven auger drills and a track mounted blasthole rig that is very effective in penetrating the laterite profiles. BDR has undertaken an extensive review of geophysical, geochemical, and mapping datasets to generate a list of ranked targets which are the focus of the 2018 near mine and regional exploration program.

GPR is expected to benefit from near-term resource growth opportunities with multiple in-mine lease discoveries at Tucano and longer-term exploration optionality from BDR’s 2,500 square kilometre highly prospective land package, which is the property reported in this report (Property).

1.8 Sample preparation and analysis

All sample preparation is done on site using separated sample preparation areas (one for mine grade control samples and one for resource infill and exploration samples). Grade control samples are analysed at the mine laboratory by 30 g fire assay with an atomic absorption spectrometer (AAS) finish. Exploration and resource holes were analysed at SGS Geosol Laborator’s in Belo Horizonte using 30 g fire assay with an AAS finish. An auditable chain of custody was established for the sample handling, data reporting, and data capturing using a Datashed SQL database. The reliability of the gold assay results was based on a well designed and implemented quality assurance and quality control protocol that includes the analysis of blanks, duplicates, and certified reference materials. Pulps from the mine laboratory are routinely sent to SGS Geosol to monitor its performance. Overall the laboratories returned acceptable results for the certified reference materials and blanks.

Based on this review work, the Qualified Person (QP) is of the opinion that the dataset provided by BDR is of an appropriate standard to use for resource estimation work. In the opinion of the QP the sample preparation, security and analytical procedures are adequate, and the data is fit for purpose.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

1.9 Mineral Resources

The Property contains an estimated Measured and Indicated Resource of 40.7 Mt at 1.71 g/t Au containing 2.24 Moz Au and an Inferred Resource of 16.4 Mt at 2.19 g/t Au containing 1.15 Moz Au (Table 1.1).

The Mineral Resources for the Property have been estimated by the following Qualified Persons (QPs) Mr Brian Wolfe, Principal Resource Geologist of International Resource Solutions Pty Ltd, takes responsibility for the Tap AB and Urucum estimates of the open pit MIK models and Mr Marcelo Batelochi, Principal Resource Geologist of Soluções em Geologia e Mineração Ltda, takes responsibility for the remaining estimates.

The Mineral Resources were estimated by a combination of Multiple Indicator Kriging (MIK) and Ordinary Kriging (OK).

For the purpose of reporting open pit Mineral Resources, the QPs tabulated the Mineral Resources within pit shells optimized using a gold price of US$1,500 per ounce of gold for the Tap AB and Urucum deposits.

The Mineral Resources have been reported in accordance with NI 43-101, and Definition Standards for Mineral Resources and Mineral Reserves adopted by the CIM Council on May 2014. Measured and Indicated Mineral Resources are inclusive of those Mineral Resources modified to produce the Mineral Reserves. A summary of the Mineral Resources as of 30 June 2017 are reported in Table 1.1.

It should be noted that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

Table 1.1 Summary of Mineral Resources as of 30 June 2017

Classification	Tonnes (kt)	Au grade (g/t)	Contained Au metal (koz)
Measured	18,448	1.41	835
Indicated	22,293	1.96	1,405
Measured and Indicated	40,742	1.71	2,240
Inferred	16,351	2.19	1,150

Notes:

• The June 2017 Tap AB UG Resource includes 173 kt at 4.68 g/t Au of Inferred Oxide within the Inferred Primary Category.

• CIM Definition Standards (2014) were used for reporting the Mineral Resources.

• Mr B. Wolfe, MAIG is the Qualified Person under NI 43-101 and takes responsibility for the Mineral Resource estimate for the Tap AB Open Pit and Urucum Open Pit Resource. Mr M. Batelochi, MAusImm (CP) is the Qualified Person under NI 43-101 and takes responsibility for all remaining resources.

• Tap AB and Urucum open pit Mineral Resources are constrained by an optimized pit shell at a gold price of US$1,500 oz and using Measured, Indicated, and Inferred categories. The cut-off grade (COG) applied to the open pit resources is 0.5 g/t Au. The COG applied to ore stockpiles is 0.5 g.t with the exception of Marginal ore which has a COG of 0.3 g/t Au.

• The underground COG is 1.2 g/t Au.

• Drilling results are up to the 30 June 2017 for the Tap AB Open Pit resource. For all other deposits, drilling results up to the 31 December 2016, excluding the Urucum North Underground resource which has drilling results up to September 2015.

• Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

• Mineral Resources are inclusive of their derived Mineral Reserves.

• The numbers may not compute exactly due to rounding.

The QPs are not aware of any known environmental, permitting, legal, title, taxation, socioeconomic, marketing, political, or other similar factors that could materially affect the stated Mineral Resource estimates.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

1.10 Mineral Reserves

Tucano Mineral Reserve estimates as of 30 June 2017 stated in this report are based on the Mineral Resources reported above and updated by Nigel Spicer of Minesure Pty Ltd. The Urucum underground Mineral Reserve was restated by AMC Consultants based on the PFS study completed in 2016 and in line with NI 43-101 guidelines. The cut-off grade (COG) used for the underground Mineral Reserves at Urucum was 1.6 g/t Au. The key modifying parameters for the optimized pit shells constraining the open pit reserves as at 30 June 2017 open pit Mineral Reserve estimates were made are summarized in Table 1.2.

It is noted that following reconciliation studies conducted in 2017, the modelling parameters for both Mineral Resource and Mineral Reserves are likely to be updated, which will be incorporated in the next Mineral Resource and Mineral Reserve estimate.

Table 1.2 Mineral Reserve key modifying parameters

Applied modifying parameters	Urucum and Tap C	Tap AB
Optimized	March 2017	June 2017
Gold price	US$1,100/oz	US$1,150/oz
Royalty	2.00%	2.00%
Total process cost US$/t milled	10.71 - 14.34	9.98 - 14.1
Process Recovery	93 - 95%	93 - 95%
Mining + grade control cost US$/t milled	9.26	6.48
Mining dilution	15 - 20%	15 - 20%
Mining loss	0%	10%
Pit slopes	40 - 55 degrees	30 - 56 degrees
G&A cost US$/t milled	1.88	2.08
Mining in-situ Au COG	0.74 - 0.94 g/t	0.66 - 0.82 g/t

Based on the updated Measured and Indicated Mineral Resources for the various mineral deposits at Tucano, the total Proved and Probable Mineral Reserves using a gold price of US$1,200/oz for the open pit reserves and US$1,120 for the Urucum underground reserves, inclusive of stockpiles and as of 30 June 2017 is estimated to be 24.5 Mt at 1.79 g/t Au for 1.41 Moz of gold (Table 1.3). Open pit reserves for Tap AB and Urucum have been reported within pit designs optimized at a gold price of US$1,150/oz and US$1,100/oz respectively. The Mineral Reserves have also been reported in accordance with NI 43-101 and Definition Standards for Mineral Resources and Mineral Reserves adopted by the CIM Council on May 2014.

Exchange rates of US$ to the Brazilian Real (R$) used for the estimation of Mineral Reserves are discussed below Table 1.3.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Table 1.3 Tucano gold operation - Mineral Reserves as of 30 June 2017

Brazil	COG	Proved Reserve			Probable Reserve			Total Mineral Reserve
	Au (g/t)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)
Urucum Open Pit oxide	0.7	374	1.21	15	180	1.18	7	554	1.20	21
Urucum East Open Pit oxide	0.7	0	0.00	0	151	1.71	8	151	1.71	8
Tap AB Open Pit oxide	0.6	2,384	2.16	166	2,024	1.76	115	4,408	1.98	280
Tap C Open Pit oxide	0.6	362	1.08	13	191	0.99	6	553	1.05	19
Total oxide open pit		3,120	1.92	193	2,545	1.66	136	5,666	1.80	329
Urucum Open Pit primary	0.8	4,446	1.69	241	5,246	1.81	306	9,692	1.76	547
Urucum East Open Pit fresh	0.7	0	0.00	0	16	1.50	1	16	1.50	1
Tap AB Open Pit primary	0.8	1,274	2.00	82	1,188	1.72	66	2,462	1.86	147
Tap C Open Pit primary	0.8	229	1.31	10	387	1.44	18	615	1.39	28
Total primary open pit		5,949	1.74	333	6,836	1.77	390	12,785	1.76	723
Urucum Open Pit total		4,820	1.65	256	5,425	1.79	312	10,245	1.72	568
Urucum East Open Pit total		0	0	0	167	1.69	9	167	1.69	9
Tap AB Open Pit total		3,658	2.10	247	3,212	1.75	181	6,870	1.94	428
Tap C Open Pit total		591	1.17	22	577	1.29	24	1,168	1.23	46
Total oxide and primary open pit		9,069	1.80	525	9,381	1.74	526	18,450	1.77	1,051
Open pit stockpile	0.5	1,985	0.66	42	0	0.00	0	1,985	0.66	42
Spent ore stockpile	0.5	1,311	0.61	26	0	0.00	0	1,311	0.61	26
ROM expansion stockpile	0.5	345	0.80	9	0	0.00	0	345	0.80	9
Total stockpiles	0.5	3,642	0.66	77	0	0.00	0	3,642	0.66	77
Total Tucano Open Pit & stockpiles		12,711	1.47	603	9,381	1.74	526	22,092	1.59	1,128
Urucum Underground primary		0	0.00	0	2,378	3.64	278	2,378	3.64	278
Total Underground primary		0	0.00	0	2,378	3.64	278	2,378	3.64	278
Grand total Tucano		12,711	1.47	603	11,759	2.12	804	24,470	1.79	1,406

Notes:

• 2018 Open Pit Mineral Reserves are stated as of 30 June 2017 based on the COGs as detailed in Table 1.3 above. Mineral Reserves are reported within open pit designs based on pit optimized shells using the following optimization COGs; Tap AB Oxide: 0.66 g/t Au, Tap AB Fresh: 0.82 g/t, Urucum Oxide 0.74 g/t Au, Urucum Fresh 0.94 g/t Au, Tap C Oxide 0.66 g/t Au, Tap C Fresh 0.86 g/t Au, Urucum East Oxide 0.74 g/t Au, and Urucum East Primary 0.94 g/t Au. The COG is based on total costs per pit (inclusive of processing, G&A, tech services and haulage, exclusive of mining costs) and averages US$12.8 for oxide and US$16.8 for fresh. Metallurgical recoveries are 95% for oxide and 93% for sulphides. The gold price used for the open pit optimizations was US$1,150/oz for Tap AB and US$1,100 for Tap C and Urucum. The gold price used to calculate the COGs for the reserves reported within the optimized pit designs was US$1,200 per ounce. An exchange rate of US$:R$ of 1:3.2 for Tap AB and 1:3.4 for Urucum was used. Open Pit Mining Recoveries of 100% for oxide and fresh were used at Urucum, 90% for oxide and fresh at Tap C and Tap AB and 90% oxide, 100% fresh at Urucum East. Dilution used for all pits was 15% for oxide and 20% for fresh.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

• 2018 Underground Mineral Reserves are stated as of 30 June 2017 and based on a COG of 1.6 g/t Au. The COG is based on an incremental mining and processing cost of US$50/t, metallurgical recovery of 90%, a gold price of US$1,120/oz and an exchange rate of US$:R$ of 1:3.8.

• A 0.5 m dilution skin was assumed for the HW and FW side of the underground production stopes making the total dilution 1.0 m in width. Ore losses were assumed to be 8% for the Benching method, 14% for the Up-Hole retreat method, and ore losses for the recovery of the sill pillars were assumed to be 35%.

• The JORC reserve released on 19 December 2017 to the ASX ("Interim Mineral Resource and Ore Reserve Update as at 30 June 2017") stated the total Urucum Underground Reserve was 2,972 kt at 3.61 g/t Au for 345 koz Au. This reserve was reported by AMC inclusive of Inferred resource blocks captured inside the stope designs that were optimized on Measured and Indicated Resources in the 2016 PFS. The Inferred Resources falling within the stope designs was removed from the calculations and reporting of the NI 43-101 compliant Urucum Underground Reserves by AMC Vancouver in 2018 and has been quoted in the Table 1.3 above.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

1.11 Mining

The Tucano mine commenced operations in 2005 and continued as a heap leach operation until 2009 when it was placed on care and maintenance following processing problems caused by increasingly transitional ore. In 2010 BDR acquired the project and commenced construction of a CIL plant to add onto the existing process infrastructure. Mining and stockpiling of ore commenced in 2011 and the CIL plant was commissioned in November 2012. To date, 21.6 Mt at 1.31 g/t Au containing 912,690 oz Au has been processed made up of a blend of pit oxide ore, sulphide ore, spent ore, and historic low-grade stockpiles. Tucano uses a conventional open pit mining method, employing a new mine contractor Unienge & Modulo (U&M).

Underground mine

Access to site is well established as open pit mining operations commenced in 2012. The journey from the port town of Macapa takes approximately four hours by car during the dry season and up to six hours during the wet season. Access to the Urucum North underground mine would be via a portal located at the north end of the Urucum North pit.

BDR undertook a program of resource infill drilling below the Urucum North Pit. Most of these holes have been geotechnically logged for structural data. The geotechnical data logged from these holes has been analysed to provide preliminary estimates of stable stope development and pillar dimensions, using industry standard empirical techniques.

The mine layout is based on the following criteria:

• Twin declines (north and south based on orebody strike length of >800 m.

• Single access portal located at the northern end of the mine.

• North and south exhaust rises.

• Level interval of 20 m (based on 15 m production blastholes and the narrow width of some lodes).

• Crosscuts between the decline and the mineralization, every 20 vertical meters (vm).

• Access for decline development available in the Urucum North pit in January 2019.

• Targeting Measured and Indicated Mineral Resources down to 500 mRL (750 mbs). The basis for the scheduling includes:

— Focus on mining higher grade areas in 2021 and 2022.

— Assume a top-down stope extraction sequence in upper levels to allow early production.

— Assumes early development to the higher-grade areas in the lower levels where down-hole benching commences in a bottom-up orientation.

The Project sits below the Urucum North open pit and power is already available for the Urucum North underground mine. The underground mine will make use of the open pit mine’s existing explosives magazines and warehouse.

1.11.1 Underground mine economics

Project capital is estimated at $44M; pre-production capital (Year 1) is estimated to be $10M. Sustaining capital ($34M) is defined as all capital following pre-production and includes ongoing mine development after production commences, and fixed equipment replacements and rebuilds over the mine life.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Total site operating costs have been estimated at $88.67/t milled. The unit costs are broken down as follows:

• Mining: $71.67/t milled

• Milling and General and Administration: $17.00/t milled

The cost estimate and projected revenue were prepared with a base date of Year 1 and use constant Year 1 dollars (no inflation). For net present value (NPV) estimation, all costs and revenues are discounted at 5% from the base date. An exchange rate of US$:R$ of 1:3.8 and a royalty of 2% has been assumed, changes to the exchange rate affect labour costs only.

Revenues from gold production have been based on:

• Metallurgical recovery of 90%

• Gold royalty of 2%

AMC has conducted a pre-tax economic assessment of the Urucum North underground. Over an 8-year operating life, the mine is projected to generate approximately $25.5M pre-tax NPV at 5% discount rate, with a pre-tax internal rate of return (IRR) of 19.9%.

1.12 Metallurgy and process plant

The Tucano processing plant uses a conventional carbon in leach (CIL) gold recovery process. The comminution circuit of the process plant is comprised of a primary jaw crusher, followed by a SAG mill. The milled product is processed in the CIL circuit where gold contained in the ore is leached and adsorbed onto activated carbon. The CIL tails slurry undergoes cyanide destruction prior to disposal in the tailings dam. Loaded carbon is rinsed prior to elution. A Pressure Zadra elution circuit is used to remove the gold from the loaded carbon and after which the rich solution flows through a bank of two electrowinning cells where the gold in the solution is deposited on the cathodes within the cells.

Following elution, the carbon is re-generated in the carbon re-generation kiln prior to being returned to the CIL circuit. The sludge containing the gold (which is weakly adhered to the cathodes) is dislodged with a high-pressure water spray and the sludge is dried in a drying oven before being direct smelted with fluxes in a furnace. The doré metal and slag separate in the furnace, the slag is poured off to slag pots and the doré metal is poured into bars. Doré bars are cleaned, weighed, stamped, sampled, and then stored in a safe, contained within a strong room, while awaiting dispatch.

With the current configuration of the grinding circuit, the plant can process up to 3.6 Mt per annum and a blend ratio of 80:20, oxide:sulphide. Increasing sulphide feed in future years meant that this blend ratio was not sustainable and in 2017 work commenced to expand the CIL plant to include a 6 MW ball mill, extra CIL tank, oxygen plant, and pre-leach thickener. The aim of this was to accommodate the planned 20:80, oxide:sulphide ratio planned for future years as well as maintain a consistent 75 um grind size. With the ball mill in conjunction with the extra CIL tank and oxygen plant, recoveries are expected to increase from 88% to 93%. As of September 2018, commissioning of the ball mill was complete and the additional CIL tank and oxygen plant remains on track for late-October 2018 and early-November 2018, respectively.

1.13 Infrastructure

The site is connected to the national electrical grid by a 69 kilovolt (kV), 20 megavolt-ampere (MVA) power line via Companhia de Eletricidade do Amapá (CEA), the local supply authority. Power is also provided by an independent 11 MW continuous rated Aggreko diesel-powered generation system. Water is provided from recycled process water and water storage dams. The mine currently employs a total of 1,116 persons, of whom 416 are Beadell Brasil employees and 700 contractors. Approximately 99% are Brazilian nationals with the majority of the workforce from local towns in proximity to the mine. Accommodations for senior and junior staff members are provided by a camp located 2 km from the plant which is capable of housing 122 persons. The plant contains the mechanical treatment (crusher, mill and thickener) and chemical process (CIL, detox, oxygen plant, elution) areas. The other major infrastructure buildings include the reagents storage warehouse, reagents preparation plant, chemical laboratory and process laboratory, tailings dam and power distribution system. The main administration area includes maintenance (electrical and mechanical) workshops, mine administration facilities, technical services offices, stores warehouse, core yard, mess facility, and medical centre.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

1.14 Market studies and contracts

Gold output from Tucano is in the form of doré bars which are shipped to Sao Paulo for refining by Umicore, the contract refiner. A number of operational duties have been contracted out to suppliers, most notably U&M is the principal mining contractors. SGS Geosol operates the offsite certified laboratory in Belo Horizonte, Minas Gerais and McKay operates a Schramm RC rig for grade control and resource drilling. The various contracts were awarded following a competitive bidding process, prices are within the industry range and comparable to other operations in Brazil.

1.15 Environmental and social

In 2011 SRK Consultores do Brasil Ltda produced an Environmental Control Plan (PCA) to present changes implemented to the project’s environmental controls as a condition for the renewal of the Operating Licence issued by the State Secretariat of Environment, Amapá (SEMA). This followed a period of care and maintenance from 2009 until BDR acquired the project in 2010. The PCA reviewed changes to the proposed process plant configuration (from heap leach to CIL) and ways of minimizing the environmental impact of the plant up-grade. The PCA also covered the environmental aspects of mining and plans to monitor and manage these as well as implementing various programs for the local community.

1.16 Capital and operating costs

Major capital expenditures for 2018 will include US$16.4M for the process plant upgrade, US$2.7M for the tails storage facility, US$1.5M for the powerline upgrade, and US$0.7M for clearing and hydro seeding. Capitalized stripping or miscellaneous minor capital requirements have not been reported. In 2017 cash operating costs for Tucano was US$1,100/oz of gold sold, which includes all mining, treatment, inventory adjustments and site general, and administrative costs. It excludes depreciation, amortization, sustaining capital, and corporate general and administration costs.

1.17 Conclusions

Tucano is a successfully operating gold mine that started commercial production by BDR in January 2011 following a period of care and maintenance by the former owners, New Gold (operating as Mineração Pedra Branca do Amapari (MPBA)). BDR added a CIL circuit to the existing heap leach plant and commissioned it in November 2012.

The operation is projected to continue until 2023 based on currently available open pit Mineral Reserves. The exploration database for Tucano is reliable for the purpose of resource estimation. The Mineral Resources and Mineral Reserves have been updated to 30 June 2018.

As from 2018, a total of 24.7 Mt of ore is planned to be mined at an overall strip ratio of 5.6 to 1. Grade control reconciliation has confirmed the Mineral Resources and Mineral Reserves as previously stated for Tucano. The results of this update to the Mineral Resource and Mineral Reserve evaluation confirm the continued economic viability of exploiting the Tucano Gold Deposit. The current open pit LOM production schedule 2018 to 2023 considers 23.3 Mt of ore at an average grade of 1.43 g/t Au containing a total of 1.07 Moz of Au. In 2017, Tucano produced 129,764 oz Au at an all-in sustaining cost (AISC) of US$1,180/oz. which includes all mining, treatment, general and administrative costs, sustaining capital, and royalties which are incurred at the mine site. The mine level AISC costs exclude depreciation, amortization, and corporate general and administrative costs.

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1.18 Recommendations

Near mine exploration and resource definition drilling on key targets within and near the Urucum pit is recommended. This could have the potential to add low strip, near surface ounces to the reserves for 2018. Continued regional exploration on the established ranked list of targets is recommended to discover and define significant new gold deposits outside the mine corridor. Regional exploration work will involve stream and soil geochemistry with follow up diamond and RC drilling for the advanced projects. Total budget for Near Mine Exploration, Resource Confirmation, and Regional Exploration in 2018 is US$6.0M. The proposed work includes 25,200 m of RC and 5,500 m of diamond core, 5,700 m of Rotary Air Blast (RAB) and Auger, and 9,290 soil samples.

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Contents

1	Summary			iii
	1.1	Project summary		iii
	1.2	Location		iii
	1.3	Ownership		iii
	1.4	Geology		iii
	1.5	Mineralization		iv
	1.6	History		iv
	1.7	Drilling and exploration		iv
	1.8	Sample preparation and analysis		iv
	1.9	Mineral Resources		v
	1.10	Mineral Reserves		vi
	1.11	Mining		ix
		1.11.1	Underground mine economics	ix
	1.12	Metallurgy and process plant		x
	1.13	Infrastructure		x
	1.14	Market studies and contracts		xi
	1.15	Environmental and social		xi
	1.16	Capital and operating costs		xi
	1.17	Conclusions		xi
	1.18	Recommendations		xii
2	Introduction			29
	2.1	Terms of reference		30
	2.2	Purpose of report		30
	2.3	Sources of information		30
3	Reliance on other experts			32
4	Property description and location			33
	4.1	Location		33
	4.2	Property ownership		33
	4.3	Mineral tenure		34
	4.4	Royalties and agreements		38
	4.5	Environmental liabilities		39
	4.6	Permits and other		39
5	Accessibility, climate, local resources, etc.			40
	5.1	Accessibility		40
	5.2	Physiography		40
	5.3	Climate		40
	5.4	Infrastructure		41
		5.4.1	Water	42
		5.4.2	Power	42
		5.4.3	Workforce	42
		5.4.4	Commercial resources and services	42
		5.4.5	Transportation	42
6	History			43
7	Geological setting and mineralization			44
	7.1	Regional geology		44
	7.2	Property geology		45
	7.3	Local geology and mineralization		48
		7.3.1	Tap AB and Urucum	49

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		7.3.2	Urucum East		51
		7.3.3	Duckhead		51
	7.4	Mineralization weathering types			51
		7.4.1	Primary or sulphide mineralization		51
			7.4.1.1	Saprolite mineralization	51
		7.4.2	Colluvial mineralization		52
8	Deposit types				53
9	Exploration				54
	9.1	Airborne magnetics, radiometrics, and DEM			54
	9.2	Airborne electromagnetic survey			55
	9.3	Ground geophysics			56
	9.4	Petrophysics			57
	9.5	Geochemistry			57
	9.6	Geological mapping			59
10	Drilling				60
	10.1	Drilling methods			60
	10.2	Surveys and survey coordinate system			60
	10.3	Diamond drill core			60
	10.4	Recovery			61
	10.5	Drilling orientation			61
	10.6	Drillhole spacing			61
	10.7	Sampling			61
11	Sample preparation, analyses, and security				62
	11.1	Sample preparation			62
	11.2	Sample security			62
	11.3	Sample analysis			62
	11.4	Sample QAQC			62
	11.5	QAQC data analysis			62
	11.6	Statistical evaluation			63
	11.7	Onsite laboratory			63
		11.7.1	Tap AB grade control and resource dataset		63
		11.7.2	Tap C grade control and resource dataset		64
		11.7.3	Urucum grade control and resource dataset		65
	11.8	SGS Belo Horizonte laboratory			65
		11.8.1	Tap AB grade control and resource dataset		65
		11.8.2	Tap C grade control and resource dataset		66
		11.8.3	Urucum grade control and resource dataset		66
	11.9	QP conclusions			67
12	Data verification				69
13	Mineral processing and metallurgical testing				70
	13.1	Summary			70
	13.2	Test work review			71
	13.3	Ausenco FS metallurgical program			71
	13.4	BDR FS metallurgical program			71
	13.5	Testwork Process Development Ltd.			72
	13.6	SGS Geosol			72
	13.7	Conclusions and design criteria			73
		13.7.1	Ausenco and Ammtec FS program		73
		13.7.2	Beadell FS program		74
	13.8	Predicted recoveries			76

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14	Mineral Resource estimates				77
	14.1	Data used			78
		14.1.1	Drillhole database		78
		14.1.2	Bulk density		80
	14.2	Interpretation and modelling			82
		14.2.1	Geological interpretation		82
			14.2.1.1	Tap AB and Urucum	82
			14.2.1.2	Tap C	84
			14.2.1.3	Duckhead	85
		14.2.2	Mineralization interpretation		86
			14.2.2.1	Tap AB and Urucum Open pit model	86
			14.2.2.2	Tap AB and Urucum Underground model	88
			14.2.2.3	Urucum North and South Underground model	90
			14.2.2.4	Urucum East model	91
			14.2.2.5	Tap C model	91
			14.2.2.6	Duckhead model	92
	14.3	Compositing			93
	14.4	Statistical analysis			94
		14.4.1	Tap AB Open Pit model		94
			14.4.1.1	Descriptive statistics	94
			14.4.1.2	High grade outlier analysis	95
			14.4.1.3	MIK cut-offs and indicator class statistics	95
		14.4.2	Urucum Open Pit model		96
			14.4.2.1	Descriptive statistics	96
			14.4.2.2	High grade outlier analysis	96
			14.4.2.3	MIK cut-offs and indicator class statistics	97
		14.4.3	Tap AB Underground model		98
			14.4.3.1	Descriptive statistics	98
			14.4.3.2	High grade outlier analysis	100
		14.4.4	Urucum North Underground model		100
			14.4.4.1	Descriptive statistics and top cuts	100
		14.4.5	Urucum South Underground model		102
			14.4.5.1	Descriptive statistics	102
		14.4.6	Urucum East model		105
			14.4.6.1	Urucum East summary statistics	105
		14.4.7	Tap C (Other) model		105
			14.4.7.1	Raw sample data	105
			14.4.7.2	2 m composite data	106
			14.4.7.3	Treatment of high grade gold values (g/t)	106
		14.4.8	Tap C (C3W and C3N) model		107
		14.4.9	Duckhead model		107
			14.4.9.1	Raw sample data	107
			14.4.9.2	2 m composite data	108
			14.4.9.3	Top cuts	109
	14.5	Variography			110
		14.5.1	Introduction		110
		14.5.2	Tap AB Open Pit model variography		110
		14.5.3	Urucum Open Pit model variography		111
		14.5.4	Tap AB Underground model variography		111
		14.5.5	Urucum North Underground model variography		111
		14.5.6	Urucum South Underground model variography		111
		14.5.7	Urucum East model variography		111
		14.5.8	Tap C (other) model variography		111

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		14.5.9	Tap C (C3N and C3W) model variography		112
		14.5.10	Duckhead model variography		112
	14.6	Block modelling			113
		14.6.1	Tap AB and Urucum Open Pit block models		113
		14.6.2	Tap AB and Urucum Underground block models		113
		14.6.3	Urucum East, Tap C, and Duckhead block models		114
	14.7	Grade estimation			115
		14.7.1	Tap AB and Urucum Open Pit block models		115
			14.7.1.1	Introduction	115
			14.7.1.2	The MIK method	115
			14.7.1.3	Post MIK processing - E-type estimates	115
			14.7.1.4	Support correction (selective mining unit estimation)	117
			14.7.1.5	Indirect lognormal correction	117
			14.7.1.6	MIK and OK parameters	118
			14.7.1.7	Change of support	120
		14.7.2	Tap AB Underground - OK estimation		121
			14.7.2.1	Estimate validation	124
		14.7.3	Urucum North Underground OK estimation		124
			14.7.3.1	Urucum North Underground model - estimate validation	125
		14.7.4	Urucum South Underground model - OK estimation		125
			14.7.4.1	Urucum South Underground model - estimate validation	128
		14.7.5	Urucum East model-OK estimation		128
			14.7.5.1	Estimate validation	128
		14.7.6	Tap C (other) estimation		128
			14.7.6.1	Estimate validation	129
		14.7.7	Tap C Model (Tap C3N and Tap C3W areas)		131
		14.7.8	Stockpile estimation		131
		14.7.9	Duckhead estimation		132
	14.8	Depletion for mining activity			134
	14.9	Resource classification			134
	14.10	Resource reporting			136
	14.11	Comparison with previous estimates			138
15	Mineral Reserve estimates				142
	15.1	Summary			142
	15.2	Open pit			144
		15.2.1	Reserve block model		144
		15.2.2	Open pit optimization		145
		15.2.3	Dilution and extraction		145
		15.2.4	Gold grade data		145
		15.2.5	Cost parameters for pit optimization		145
		15.2.6	Pit optimization mill recovery		146
		15.2.7	Tap AB optimization results		146
		15.2.8	Urucum optimization results		147
		15.2.9	Open pit COG - Pit optimization and Reserves		148
	15.3	Underground mining			151
		15.3.1	Underground mine Mineral Reserve estimate		151
		15.3.2	Orebody geometry		151
		15.3.3	Cut-off grade (COG)		152
		15.3.4	Dilution and mining recovery estimates		152
		15.3.5	Mining method criteria and modifying factors		153
	15.4	Combined Mineral Reserve estimate			153
	15.5	Factors that may impact Mineral Reserve estimates			156

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16	Mining methods				157
	16.1	Open pit mining			157
	16.2	Geotechnical assumptions			157
		16.2.1	Overall pit slopes		157
			16.2.1.1	Urucum	158
			16.2.1.2	Urucum East	158
			16.2.1.3	Tap AB	158
			16.2.1.4	Pegmatite	159
			16.2.1.5	Proposed overall slope angles	159
		16.2.2	Engineered pit designs		160
			16.2.2.1	Tap AB	160
			16.2.2.2	Urucum	162
			16.2.2.3	Urucum East	163
			16.2.2.4	Tap C	164
	16.3	Contract mining			165
		16.3.1	Grade control and dig plans		166
	16.4	Underground mining			167
		16.4.1	Geotechnical and hydrogeological considerations		167
		16.4.2	Data used		167
		16.4.3	Major structures		168
		16.4.4	In-situ stress environment		169
		16.4.5	Weathering		169
		16.4.6	Ground water		170
		16.4.7	Rock mass classification		170
		16.4.8	Stable stope dimensions		171
		16.4.9	Pillar dimensions		172
			16.4.9.1	Crown pillar	172
			16.4.9.2	Sill pillars	173
			16.4.9.3	Rib pillars	173
		16.4.10	Excavation and support requirement		173
			16.4.10.1	Lateral development	173
			16.4.10.2	Vertical development (ventilation shafts)	174
			16.4.10.3	Portal locations	174
		16.4.11	Underground mine design basis		174
			16.4.11.1	Mining limits	174
			16.4.11.2	Mineral Resource estimate	174
			16.4.11.3	Stope optimization	174
			16.4.11.4	MSO estimates	175
		16.4.12	Access development		177
			16.4.12.1	Decline stand-off distance	177
			16.4.12.2	Capital development estimates	178
			16.4.12.3	Operating development estimates	178
		16.4.13	Mining methods		178
		16.4.14	Up-hole retreat mining method		179
		16.4.15	Down-hole benching mining method		180
		16.4.16	Sill pillar recovery		181
		16.4.17	Ventilation strategy		182
			16.4.17.1	Level ventilation	183
			16.4.17.2	Development ventilation	184
		16.4.18	Ventsim modelling		185
			16.4.18.1	Modelling results	185
		16.4.19	Egress		185
		16.4.20	Underground mine production and development schedules		185

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			16.4.20.1	Development profiles and scheduling rates	186
		16.4.21	LOM underground production schedule		186
		16.4.22	Underground mobile equipment		189
			16.4.22.1	Development drills / jumbos	190
			16.4.22.2	Production drills	190
			16.4.22.3	Underground trucks	191
			16.4.22.4	LHDs	192
			16.4.22.5	Ground support rigs	193
			16.4.22.6	Utility vehicles	193
		16.4.23	Underground mine personnel		193
			16.4.23.1	Working shifts	193
			16.4.23.2	Rosters	194
	16.5	Underground infrastructure			195
		16.5.1	Underground access		195
		16.5.2	Surface ventilation fans		196
		16.5.3	Change rooms and lamp room		196
		16.5.4	Underground mine offices		196
		16.5.5	Surface workshop		196
		16.5.6	Mines rescue centre		197
		16.5.7	Mine portal		197
		16.5.8	Water supply		197
		16.5.9	Existing facilities		197
		16.5.10	Power supply and distribution		197
		16.5.11	Water supply and distribution		198
		16.5.12	Dewatering		198
			16.5.12.1	Pump selection	198
			16.5.12.2	Sump design	198
			16.5.12.3	Dewatering system	199
		16.5.13	Compressed air		199
		16.5.14	Emergency egress		200
		16.5.15	Refuge chambers		200
		16.5.16	Underground communications		201
		16.5.17	Future data management systems		201
		16.5.18	Underground service bay		202
17	Recovery methods				203
	17.1	Summary			203
	17.2	Crushing and ore stockpiling			203
	17.3	Milling and classification			203
	17.4	CIL			203
	17.5	Tailings			204
	17.6	Elution			204
	17.7	Carbon regeneration			204
	17.8	Electrowinning and gold recovery			205
18	Project infrastructure				206
	18.1	Context and regional infrastructure			206
	18.2	Site access and logistics			207
	18.3	Mining and support facilities			207
	18.4	Sub-contractor facilities			208
	18.5	Process plant			209
	18.6	Water			209
		18.6.1	Raw water		209
		18.6.2	Potable water		209

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		18.6.3	Process water		209
	18.7	Power			210
	18.8	Fire protection			210
	18.9	Communications			210
	18.10	Security			211
	18.11	Personnel and personnel logistics			211
	18.12	Accommodation and medical care			211
	18.13	Tailings Storage Facility (TSF)			211
		18.13.1	Tailings operation		211
		18.13.2	Description of the constructed TSFs		212
			18.13.2.1	Tap D TSF	212
			18.13.2.2	North Mill Pond TSF	214
			18.13.2.3	WPP1 TSF	216
			18.13.2.4	Integrated WPP1 and NMP TSFs	217
		18.13.3	Extended tailings storage capacity		218
			18.13.3.1	Barragem Leste TSF	220
			18.13.3.2	WPP2 TSF	222
		18.13.4	Permitting of the TSFs		223
		18.13.5	Conclusions		223
	18.14	Underground infrastructure			223
19	Market studies and contracts				224
	19.1	Markets			224
	19.2	Contracts			224
20	Environmental studies, permitting, and social or community impact				225
	20.1	Site visit			225
		20.1.1	Groundwater and surface water monitoring field check		225
		20.1.2	Socio-economics aspects field check		226
	20.2	Known environmental liabilities			226
	20.3	Biological resources (Fauna and Flora)			226
		20.3.1	Flora		227
		20.3.2	Fauna		227
		20.3.3	Threatened and special status species		228
	20.4	Air quality			228
	20.5	Noise quality and vibrations			229
	20.6	Geotechnical			229
	20.7	Hydrogeology (groundwater), hydrology (surface water)			229
	20.8	Cultural and archaeological resources			230
		20.8.1	Regional archaeological context		230
		20.8.2	Archaeological studies at Tucano Project		230
	20.9	Socio-economics			231
21	Capital and operating costs				233
	21.1	Capital expenditures			233
	21.2	Operating costs			233
		21.2.1	Underground capital costs - basis of estimate		233
		21.2.2	Contractor mining versus owner mining		233
		21.2.3	Total underground capital cost estimate		234
		21.2.4	Underground lateral development capital costs		236
		21.2.5	Underground vertical development capital costs		236
		21.2.6	Underground mobile equipment capital costs		236
		21.2.7	Underground infrastructure capital costs		236
		21.2.8	Surface infrastructure capital costs		237
		21.2.9	Exclusions from the underground cost estimate		237

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		21.2.10	Indirect capital costs	237
		21.2.11	Capital contingency	237
		21.2.12	Underground operating costs - basis of estimate	238
		21.2.13	Underground mine total operating costs	238
			21.2.13.1        Underground mining cost	239
		21.2.14	Underground mobile equipment assumptions and costs	240
		21.2.15	Employee costs / labour costs	240
22	Economic analysis			242
	22.1	Underground mine economics		242
		22.1.1	Assumptions	242
		22.1.2	Economic analysis	242
	22.2	Taxes and Royalties		245
	22.3	Economic sensitivity analysis		246
23	Adjacent properties			249
24	Other relevant data and information			250
25	Interpretation and conclusions			251
		25.1	Risk assessment	251
		25.2	Mineral Resources and Mineral Reserves	251
		25.3	Mining	251
		25.4	Processing	251
		25.5	Production and all-in sustaining costs	252
		25.6	Health and safety	252
		25.7	Underground mining	252
26	Recommendations			254
	26.1	Operations		254
	26.2	Exploration and Resource Delineation		254
		26.2.1	Near mine and resource definition	254
		26.2.2	Regional exploration	254
	26.3	Underground mining		255
27	References			257
28	QP Certificates			258

Tables
Table 1.1	Summary of Mineral Resources as of 30 June 2017	v
Table 1.2	Mineral Reserve key modifying parameters	vi
Table 1.3	Tucano gold operation - Mineral Reserves as of 30 June 2017	vii
Table 2.1	Persons who prepared or contributed to this Technical Report	30
Table 4.1	GPR Mineral Tenure	35
Table 5.1	Monthly rainfall at Tucano from 2008 to 2016 inclusive	41
Table 7.1	Stratigraphic column (from AMC 2009)	46
Table 7.2	Description of mineralized zones in each deposit	49
Table 9.1	Conductivity test results of a sample of pyrrhotite rich core from Urucum	57
Table 9.2	Number of exploration samples collected by sample type and company	58
Table 13.1	SGS Geosol samples	72
Table 13.2	SGS Geosol bottle roll results	73
Table 13.3	Selected design criteria (Ausenco)	74

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Table 14.1	Summary of estimated Mineral Resources (as of 30 June 2017)	78
Table 14.2	Drillhole data used in the Open Pit Mineral Resource estimates	79
Table 14.3	Summary of bulk density measurements by weathering type	81
Table 14.4	Urucum: Comparative statistics of bulk density	81
Table 14.5	Tap AB: Comparative statistics of bulk density	81
Table 14.6	Tap C: Comparative statistics of bulk density	81
Table 14.7	Duckhead: Comparative statistics of bulk density	82
Table 14.8	Tap AB Gold Deposit - Open Pit model summary statistics gold (g/t)	94
Table 14.9	Tap AB Gold Deposit - Open Pit Resource indicator class statistics	96
Table 14.10	Urucum Gold Deposit - Open Pit model; summary statistics gold (g/t)	97
Table 14.11	Urucum Gold Deposit - Open Pit model indicator class statistics (selective)	98
Table 14.12	Tap AB Underground model; statistics by domain	99
Table 14.13	Tap AB Underground Resource; 3 m composite top cut capping by domain	100
Table 14.14	Urucum North Underground Resource statistics; raw, composited, and capped	101
Table 14.15	Urucum South Underground Resource statistics; Au (g/t)	102
Table 14.16	Urucum South Underground Resource statistics; raw gold data (g/t)	104
Table 14.17	Urucum East 1 m composites by domain - summary statistics	105
Table 14.18	Raw data statistic summary - grouped ore domains gold (ppm)	105
Table 14.19	Tap C and gap, 2 m composite statistical summary of variogram sector VERT	106
Table 14.20	Tap C (other) model; top grade for 2 m composite - Au (g/t)	106
Table 14.21	Tap C3W and C3N 2 m composites by domain - summary statistics	107
Table 14.22	Tap C3W and C3N 2 m applied top cuts by domain	107
Table 14.23	Duckhead; raw data Au (g/t), statistical summary by domain	108
Table 14.24	Duckhead; 2 m composite statistical summary - grouped by ore lode	109
Table 14.25	Duckhead determined top cuts for each lode	110
Table 14.26	Tap AB Gold Deposit - Open Pit Resource block model parameters	113
Table 14.27	Urucum Gold Deposit - Open Pit Resource block model parameters	113
Table 14.28	Tap AB Underground block model parameters	114
Table 14.29	Urucum North Underground block model parameters	114
Table 14.30	Urucum South Underground block model parameters	114
Table 14.31	Urucum East Gold Deposit block model parameters	114
Table 14.32	Tap C Gold Deposit block model parameters	115
Table 14.33	Duckhead Gold Deposit block model parameters	115
Table 14.34	Tap AB and Urucum open pit block model Indicator cut-off and probability	116
Table 14.35	Tap AB Gold Deposit - Open Pit Resource MIK sample search criteria	119
Table 14.36	Urucum Gold Deposit - Open Pit Resource MIK sample search criteria	119
Table 14.37	Tap AB Gold Deposit - Open Pit Resource OK sample search criteria	120
Table 14.38	Urucum Gold Deposit - Open Pit Resource OK sample search criteria	120
Table 14.39	Variance adjustment ratios (3 mE x 5 mN x 4 mRL SMU)	121
Table 14.40	Tap AB Gold Deposit - Underground Resource; orientation and radius of search radius of individual lodes / oxide domains	122

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Table 14.41	Urucum South Underground model: orientation of search radius of individual lodes	126
Table 14.42	Urucum East summary estimation parameters	128
Table 14.43	Statistical validation of Au (g/t) of 1st neighbourhood	130
Table 14.44	Estimation parameters for Tap C (Tap C3N and Tap C3W areas)	131
Table 14.45	Duckhead search ellipses by lode / oxidation domain	133
Table 14.46	Resource classification criteria	135
Table 14.47	Tucano Gold Resources 30 June 2017	137
Table 14.48	Mineral Resource changes by deposit and physical reporting limits	139
Table 14.49	Comparison of Mineral Resource estimates	141
Table 15.1	Proportion of reserves by deposit	144
Table 15.2	Optimization cost parameters	146
Table 15.3	COG estimation parameters using pit shell optimization gold price	149
Table 15.4	COG estimation parameters using the US$1,200 gold price	150
Table 15.5	Urucum North Underground Mineral Reserve estimate	151
Table 15.6	Total Mineral Reserves - 30 June 2017	154
Table 16.1	Recommended wall design parameters	157
Table 16.2	Measured overall slope angles (degrees)	158
Table 16.3	Proposed overall slope angles (degrees)	160
Table 16.4	Urucum pits design data	163
Table 16.5	Tucano mining fleet	166
Table 16.6	Grade control grade ranges by material type	167
Table 16.7	Diamond drillholes used in the geotechnical analysis	168
Table 16.8	Range of Q’ and RMR values	171
Table 16.9	Stope design factors and hydraulic radii	171
Table 16.10	Urucum North Underground Mineral Resource estimate - based on COG of 1.6 g/t Au	174
Table 16.11	MSO parameters	176
Table 16.12	Lateral capital development estimates	178
Table 16.13	Vertical capital development estimates	178
Table 16.14	Operating lateral development estimates	178
Table 16.15	Surface fan details	185
Table 16.16	Lateral development details	186
Table 16.17	Vertical development details	186
Table 16.18	LOM underground production schedule	188
Table 16.19	Major equipment required over LOM	189
Table 16.20	Underground workforce time allocation during shift	194
Table 16.21	Mining department employees - excluding UG equipment operators and maintenance personnel	195
Table 18.1	Summary of TSF storage capacity	212
Table 21.1	Cash operating costs 2017	233
Table 21.2	Pre-production capital and total capital cost estimate	234
Table 21.3	Capital cost estimate by year	235

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Table 21.4	Annual capital expenditure	235
Table 21.5	Underground infrastructure capital costs	236
Table 21.6	Surface infrastructure capital costs	237
Table 21.7	Capital contingency	238
Table 21.8	LOM underground total operating costs by major area	238
Table 21.9	LOM underground total operating costs by year	239
Table 21.10	LOM average underground total operating costs	239
Table 21.11	Underground mining costs	239
Table 21.12	Underground mobile equipment productivity assumptions	240
Table 21.13	Mobile equipment operating cost assumptions	240
Table 21.14	Mobile equipment maintenance cost assumptions	240
Table 21.15	Employee wages	241
Table 21.16	Employee benefits	241
Table 22.1	Key economic assumptions and results for the underground mine	243
Table 22.2	Urucum North Underground LOM production and cash flow forecast	244
Table 22.3	Economic sensitivity analysis (pre-tax)	247
Table 26.1	2018, Resource Definition, Near mine, and regional exploration budget	255
Figures
Figure 4.1	Tucano mine location map	33
Figure 4.2	GPR tenement ownership map with tenement status	37
Figure 5.1	Monthly rainfall at Tucano from 2008 to 2016 inclusive	41
Figure 7.1	Geology map of basement geology overlain by the Property tenements	44
Figure 7.2	Property geology map	47
Figure 7.3	Tucano deposit location map	48
Figure 7.4	Tap AB Gold Deposit section 94,550 mN	50
Figure 8.1	Reconstruction of the Guiana Craton and West Africa Shield	53
Figure 9.1	Merged Airborne Magnetic Total Magnetic Image (AS)	55
Figure 9.2	SPETREM electromagnetic image	56
Figure 9.3	Photo of the conductivity test sample	57
Figure 9.4	Tenement status map showing the coverage of soil sampling	58
Figure 9.5	Serra da Canga East stream sediment anomaly	59
Figure 11.1	Control chart for standard GE076	64
Figure 11.2	Control chart for Standard G311-5	67
Figure 13.1	Proposed Tucano expansion flow sheet	75
Figure 14.1	Plan view of drilling Urucum (left) and Tap AB (right)	80
Figure 14.2	Urucum Gold Deposit cross section displaying geology and coded block model	83
Figure 14.3	Tap AB Gold Deposit cross section displaying geology and coded block model	84
Figure 14.4	Isometric view of Tap C lithology solids	85
Figure 14.5	Plan view of Duckhead showing the modelled lithological units	86
Figure 14.6	Tap AB Gold Deposit isometric SE view	87
Figure 14.7	Urucum Gold Deposit isometric SE view	88
Figure 14.8	Oblique view of the Tap AB constructed mineralization wireframes	89

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Figure 14.9	Cross section of all Urucum North Underground domains	90
Figure 14.10	Long section view of Urucum showing the underground mineralization domains	91
Figure 14.11	Tap C gold mineralization domains - oblique view	92
Figure 14.12	Mineralized wireframes of Duckhead and surrounding deposits	93
Figure 14.13	Urucum East Model - resource classification - HG Lode	136
Figure 15.1	Deposit location plan	143
Figure 15.2	Tap AB optimization results	147
Figure 15.3	Urucum optimization results	148
Figure 16.1	Tap AB1/2 final design	161
Figure 16.2	Tap AB3 final design	162
Figure 16.3	Urucum final design	163
Figure 16.4	Urucum East final design	164
Figure 16.5	Tap C final design	165
Figure 16.6	Long projection of Urucum North Underground looking west, showing geotechnical drillhole locations	168
Figure 16.7	Cross-section showing weathering profile (Section 99,865 mN)	170
Figure 16.8	Modified Mathews short-term stability estimate	172
Figure 16.9	Tonnage / grade curves based on MSO stope shapes generated above 1.6 g/t Au COG	175
Figure 16.10	3D view of mine design and MSO stopes - by lode number	177
Figure 16.11	Long projection indicating up-hole retreat mining sequence	179
Figure 16.12	Long projection indicating down-hole benching mining method and sequence	180
Figure 16.13	Sill pillars	181
Figure 16.14	Options for sill pillar recovery	182
Figure 16.15	Long projection of the primary ventilation system	183
Figure 16.16	Forced ventilation for a typical level	184
Figure 16.17	Projected production profile	187
Figure 16.18	Major equipment required over LOM	189
Figure 16.19	Simba H1354 production drill	191
Figure 16.20	Scania P420 truck being loaded by a Volvo front-end loader	192
Figure 16.21	Section of proposed sump layout	199
Figure 16.22	MineArc refuge chambers	201
Figure 18.1	U&M workshop facilities	208
Figure 18.2	Tap D embankments	213
Figure 18.3	Typical wall cross section	214
Figure 18.4	NMP embankments	215
Figure 18.5	WPP1 embankments	216
Figure 18.6	Wall 11 cross section	217
Figure 18.7	Integrated WPP1 and NMP TSFs	218
Figure 18.8	Proposed TSF expansion	219
Figure 18.9	Typical cross section of Walls 21 and 22a	221
Figure 18.10	WP TSF layout	222
Figure 20.1	Location map of field visit monitoring points	226

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Figure 22.1	Production costs, revenues, and cumulative pre-tax cash flow by year	245
Figure 22.2	Sensitivity analysis - pre-tax NPV at 5% discount rate	247
Figure 22.3	Pre-tax NPV and IRR sensitivities to changes in the gold price	248
Figure 23.1	Tenement ownership map	249

Distribution list

1 e-copy to Great Panther Silver Limited

1 e-copy to AMC Vancouver office

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Acronyms and abbreviations

	Description		Description
$	United States of America dollars	M	million
µm	micrometre	m	metres
2D	two dimensional	m3	cubic metre
3D	three dimensional	m3/h	cubic metre per hour
AAS	atomic absorption spectrometer	m3/s	cubic metre per second
AEM	airborne electromagnetic	Ma	Million years
AISC	all-in sustaining cost	MACA	Mining and Civil Australia
ANM	Agência Nacional de Mineração (National Mining Agency, Brazil)	mbs	metres below surface
ANQUIM	Análise Química Mineral, Ambiental e Industrial Mineral	mE	metres east
ASX	Australian Securities Exchange	MII	Measured + Indicated + Inferred
Au	gold	MIK	Multiple Indicator Kriging
AWD	all-wheel drive	min	minute
bcm	bank cubic metres	ml	millilitre
BDR	Beadell Resources Ltd	mm	millimetres
BIF	Banded Iron Formations	Mm3	Million cubic metres
BVI	Blast Vector Indicators	MMI	mobile metal ion
ºC	degrees centigrade	mN	metres north
CC	correlation coefficient	mRL	metres reduced level
ccdf	conditional cumulative distribution function	Moz	million ounces
CEA	Centrais Elétricas do Amapá	MPBA	Mineração Pedra Branca do Amapari
CFEM	Compensation for Exploitation of Mineral Resources	MPC	Mine Power Centres
cfm	cubic feet per minute	MPH	MPH Minerals Consultancy Ltd.
CIC	carbon in column	MSO	Mineable Shape Optimizer
CIL	carbon-in-leach	Mtpa	million tonnes per annum
CIP	carbon-in-pulp	Mt	Million tonnes
cm	centimetre	MVA	megavolt-ampere
CO	carbon monoxide	MW	megawatt
CO2	carbon dioxide	N	northing
COG	cut-off grade	NaCN	sodium cyanide
CSLL	Social Contribution Tax on Profits	NATA	National Association of Testing Authorities
CSS	closed side setting	NN	Nearest Neighbour
cusum	cumulative sum of the deviations	NMP	North Mill Pond
CV	coefficient of variation	NO	nitric oxide
dBA	decibels	NO2	nitrogen dioxide
DDH	diamond drillhole	NPV	net present value
FS	Feasibility Study	NQ2	size of diamond drill rod / bit / core
DEM	Digital Elevation Model	O2	oxygen

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	Description		Description
DNPM	Departamento Nacional de Produção Mineral (National Department of Mineral Production, Brazil)	OK	Ordinary Kriging
DTM	digital terrain model	oz	troy ounce
E (X)	easting	P80 -75 µm	80% passing 75 microns
EBX	EBX Gold Ltd	PAL	pulverize and leach
EDM	electronic distance measuring	PBX	private branch exchange
EGL	effective grinding length	PCA	Environmental Control Plan
EPCM	engineering, procurement, and construction management	PDC	Power Distribution Centre
EPS	Earth Works Production Scheduling	PETN	Penta Erythritol Tetra Nitrate
EV	expected value	PFS	Pre-Feasibility Study
FIFO	fly in fly out	ppb	parts per billion
g	gram	ppm	parts per million
G&A	General and Administration	psi	pounds per square inch
g/m³	grams per cubic metre	PVC	poly vinyl chloride
GPR	Great Panther Silver Limited	QC	quality control
g/t	grams per tonne	QP	Qualified Persons
ha	hectare	Q-Q	quantile-quantile
HARD	half the absolute relative difference	R$	Brazilian Real
HCl	hydrochloric acid	RAB	rotary air blast
HDPE	high density poly ethylene	RC	reverse circulation
HQ2	size of diamond drill rod / bit / core	RL (Z)	reduced level
HR	hydraulic radii	ROM	run-of-mine
hr	hours	RQD	rock quality designation
HRD	half relative difference	RWC	Ray Walton Consulting Inc.
ICP-MS	inductivity coupled plasma mass spectroscopy	SCS	Southern Geoscience Consultants
ID	Inverse Distance weighting	SD	standard deviation
ID²	Inverse Distance Squared	SEMA	State Secretariat of Environment, Amapá
IP	Induced Polarization	SGS	Société Générale de Surveillance
IPD	Inverse Power Distance	SMBS	sodium metabisulfite
IPS	integrated pressure stripping	SMU	selective mining unit
IRPJ	Corporate Income Tax	SO2	sulphur dioxide
IRR	internal rate of return	SSAG	single-stage, semi-autogenous grinding
IRS	Inter Ramp Slope	SUDAM	Superintendence of Amazonas Development
ISO	International Standards Organization	t	tonnes
IT	Integrated Tool carrier	t/hr	tonnes per hour
ITS	Inchcape Testing Services	t/m³	tonnes per cubic metre
kg	kilogram	TFRM	Control, Monitoring and Supervision of Research Activities, Mining, Exploration and Exploitation of Mineral Resources Fee
kg/m3	kilogram per cubic metre	TMI	Total Magnetic Intensity
kg/t	kilogram per tonne	tpd	tonnes per day

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	Description		Description
km	kilometres	TSF	Tailings Storage Facility
km²	square kilometres	U&M	Unienge & Modulo
kt	kilotonne	US$	United States of America dollars
ktpa	thousand tonnes per annum	UTE	Thermoelectric Plant
kV	kilovolt	VLF	Very Low Frequency
kW	kilowatt	vm	vertical metre
kWh	kilowatt hour	VPN	Virtual Private Network
LHD	Load Haul Dump	WPP1	West Pond Phase 1
LOM	life of mine	WPP2	West Pond Phase 2
l/hr/m²	litres per hour per square metre

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2 Introduction

This report has been prepared for Great Panther Silver Limited (GPR) in connection with the proposed acquisition (the Acquisition) by GPR of Beadell Resources Limited (BDR) pursuant to a Scheme Implementation Deed entered into between GPR and BDR on 23 September 2018. GPR has announced the friendly acquisition of BDR. The all-shares transaction will result in the consolidation of the two companies in order to create an emerging intermediate gold producer. Subsequently, GPR is required to file a technical report on the Tucano Property (Property) in order to support its disclosure within the Information Circular (expected late November) for the meeting of the shareholders of GPR to be held to approve the Acquisition.

AMC Mining Consultants (Canada) Ltd. (AMC) was retained by GPR to prepare a Technical Report in accordance with National Instrument 43-101 “Standards of Disclosure for Mineral Projects” of the Canadian Securities Administrators (NI 43-101) for the Tucano Gold Mine and surrounding concessions (Property). The Property is located in Amapá State in northern Brazil. The nearest major populated centre to the Tucano site is Macapá, situated on the northern bank of the Amazon River.

Tucano Gold Mine is currently wholly owned by BDR and its Brazilian subsidiary, Beadell Brasil Ltda (Beadell Brasil). The Tucano Gold Mine is an operating mine that produced 129,764 oz’s of gold in 2017. Current operations are from Tap AB, Tap C, and Urucum deposits.

AMC is responsible for managing and preparing the Technical Report with inputs from the following Qualified Persons (QP): Mr G. Methven from AMC Mining Consultants (Canada) Ltd., Mr B. Wolfe from International Resource Solutions Pty Ltd., Mr N. Spicer from Minesure Pty Ltd., Mr R. Lacourt from LOM Consultants, Mr J. Moreno from SRK Consultants, Mr P. O’Bryan from Peter O’Bryan and Associates, Mr R. Walton from Ray Walton Consulting Inc., and Mr M. Batelochi from Soluções em Geologia e Mineração Ltda.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Table 2.1 Persons who prepared or contributed to this Technical Report

Qualified Persons responsible for the preparation of this Technical Report
Qualified Person	Position	Employer	Independent of GPR	Date of last site visit	Professional designation	Sections of report
Mr G. Methven	Principal Mining Engineer	AMC Mining Consultants (Canada) Ltd.	Yes	None	P.Eng. (BC)	1 (part), 2, 3 (part), 15 (part), 16 (part), 19, 21, 22 (part), 24, 25 (part), 26 (part).
Mr B. Wolfe	Principal Resource Geologist	International Resource Solutions Pty Ltd	Yes	3 Aug 2015	MAIG	1 (part), 4 (part), 5 (part) 6 - 12, 14 (part), 23, 25 (part), 26 (part), 27 (part).
Mr N. Spicer	Principal Mining Engineer	Minesure Pty Ltd	Yes	8 Feb 2017	C.Eng. MIMMM	1 (part), 3 (part), 15 (part), 16 (part), 22 (part), 25 (part.
Mr R. Lacourt	Principal Mining Engineer	LOM Consultants	Yes	4 April 2018	SME - RM	Section 5 (part), 18 (part).
Mr J. Moreno	Principal Geotechnical Engineer	SRK Consultants	Yes	21 Oct 2016	MIEAust CPEng.	Section 18 (part).
Mr P. O'Bryan	Principal Mining Engineer	Peter O'Bryan and Associates	Yes	2 June 2015	MAusIMM (CP)	Section 16 (part).
Mr R. Walton	Principal Engineer	Ray Walton Consulting Inc	Yes	15 Jan 2016	P.Eng. (Ontario)	Section 13 (part), 17 (part), 25 (part).
Mr M. Batelochi	Principal Resource Geologist	Soluções em Geologia e Mineração Ltda	Yes	20 March 2018	MAusIMM (CP)	1 (part), 4 (part), 14 (part), 20, 26 (part), 27 (part).
Other Experts who assisted the Qualified Persons
Expert	Position	Employer	Independent of GPR	Visited site		Sections of report
Mr R. Rocha	Civil Engineer	SRK Consultants	Yes	18 April 2018		Section 18.
Mr P. O'Bryan	Principal Mining Engineer	Peter O'Bryan and Associates	Yes	2 June 2015		Section 15.
Mr L. Azevedo	Partner	FFA Legal	Yes	None		Section 4.
Mr R. Lessa	Attorney	FFA Legal	Yes	None		Section 4 and 22

2.1 Terms of reference

Great Panther Silver Limited (GPR) entered into a Scheme Implementation Deed dated 23 September 2018, that will result in the acquisition of all of the issued ordinary shares of BDR. Upon completion of the transaction, the Tucano Project will be 100% owned and operated by GPR. GPR is a primary silver mining and exploration company listed on the Toronto Stock Exchange trading under the symbol GPR, and on the NYSE American under the symbol GPL. AMC has prepared this Technical Report for the Tucano Gold Property on behalf of GPR according to NI 43-101.

2.2 Purpose of report

The purpose of this report is to restate GPR’s Mineral Resource and Mineral Reserve statement for the Tucano Gold Mine as of 30 June 2017. This Technical Report conforms to the standards set out in NI 43-101, pursuant to Form 43-101F.

2.3 Sources of information

This report relies on historic and recent data generated by BDR and GPR. GPR has re-engaged a number of specialist consultants and information from these reports prepared by previous independent consultants has been utilized in the compilation of this report. A full listing of the principle sources of information are included in Section 27 of this report.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

This report has been prepared by AMC for GPR. The information, conclusions, opinions, and estimates contained herein, for which the named Qualified Persons (QPs) take responsibility, are based on:

• Information available to AMC, at the time of preparation of this report.

• Assumptions, conditions, and qualifications as set forth in this report.

• Data, reports, and other information supplied by GPR and from other sources.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

3 Reliance on other experts

The QPs have relied, in respect of legal and taxation aspects, upon the work of the Experts listed below. To the extent permitted under NI 43-101, the QPs disclaim responsibility for the relevant section of the Technical Report.

The following disclosure is made in respect of this expert:

• Luis Mauricio Ferraiuoli de Azevedo, Partner FFA Legal.

Report, opinion, or statement relied upon:

• Title Opinion on the Mineral Rights held by Beadell’s group dated 17 September 2018

Extent of reliance:

• Full reliance following a review by the Qualified Person(s).

Portion of Technical Report to which disclaimer applies:

• Section 4.3

The following disclosure is made in respect of this expert:

• Rodrigo Simões Lessa, Attorney, FFA Legal.

Report, opinion, or statement relied upon:

• Taxation and royalties

Extent of reliance:

• Full reliance following a review by the Qualified Person(s).

Portion of Technical Report to which disclaimer applies:

• Sections 4.4 and 22.1 for royalties and Section 22.2 for taxation

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

4 Property description and location

4.1 Location

The Property is located at latitude 0.85°N and longitude 52.90°W in Amapá, the most north-eastern state of Brazil, approximately 15 km from the town of Serra do Navio and was formerly an open pit mining and heap leaching operation. It ran from late 2005 until January 2009, producing approximately 292,000 troy ounces (oz) of gold from 8.8 million tonnes (Mt) of ore. The property was acquired by BDR in 2010 and the plant upgraded with the addition of SAG mill and carbon-in-leach (CIL) infrastructure capable of treating sulphide ore. From November 2012 to 30 June 2017, 17.0 Mt at 1.34 g/t Au containing 735,294 oz Au was processed made up of a blend of pit oxide ore, sulphide ore, spent ore, and historic stockpiles. The mine is approximately 200 km from Macapá, the state capital, and is accessible by the Brazilian federal highway BR-210, or by chartered aircraft (Figure 4.1).

Figure 4.1 Tucano mine location map

Source: BDR

4.2 Property ownership

The Property and the Tucano Gold Mine is currently wholly owned by BDR and its Brazilian subsidiary, Beadell Brasil Ltda (Beadell Brasil), which is to be acquired by GPR. The Tucano Gold Mine is an operating mine that produced 129,764 oz’s of gold in 2017. Current operations are from Tap AB, Tap C, and Urucum deposits. Upon completion of the Acquisition, existing BDR and GPR shareholders are expected to own approximately 38% and 62%, respectively, of the combined company.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

4.3 Mineral tenure

GPR will hold tenements totalling approximately 2,500 km². Details of each tenement including area and status is provided in Table 4.1.

Outside of the Mining Concession 851.676 / 1992, GPR holds a 2,500 km² area of prospective exploration tenements at a grassroots level of exploration (Mining Concession is centred in the circle; Figure 23.1). This ground spans a length of 100 km of Paleoproterozoic greenstones and is 100% GPR owned with the exception of 4 tenements with Iron Ore concessions currently held by Zamin Ferrous (Zamin). Within Zamin’s leases, GPR has 100% gold rights whereas Zamin retains the iron ore rights - refer to Figure 7.2. In 2017 BDR reached an agreement with joint venture partner Mineração Vale dos. Reis Ltda (MVR) to convert its interest from 70% to 100%.

Table 4.1 includes all GPR's mineral rights (including tenement status), clarifying that all mineral rights are owned by GPR, but there are some tenements still held by Mineração Tanagra, Mineração Dorica and Mineração Vale doc Reis, which are current titleholders registered before the Brazilian National Mining Agency (Agência Nacional de Mineração (ANM)) that will replace current regulator National Department of Mineral Production (Departamento Nacional de Produção Mineral (DNPM)) as Table 4.1 shows. This means that the assignment of the tenements to BDR's Brazilian subsidiaries (Marina Norte and Serra da Canga) have already been requested before the ANM (replacing the DNPM) but have not been approved yet. In most of those cases, since the tenements are in exploration application phase, they can only be assigned after the ANM grants the respective exploration permits which has not happened yet.

In cases where there is already an exploration permit granted, GPR must pay TAH, which means Annual Fee per Hectare and shall be paid until January or July each year, depending on the date in which the exploration permit was published and is a fixed value per hectare, being R$3.21 during the original period of the exploration permit and R$4.86 during its extension period.

Figure 4.2 shows a map of these tenements as they relate to the mining concession shown in red. The peripheral tenements are treated as part of the property as any discovery would be treated in a central processing facility.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Table 4.1 GPR Mineral Tenure

Process

Titleholder

Status

Area (ha)

Project

Phase

License number

Publication date

Renewal date

Expiration date

Tah / June 2018

Substance

District

858.046/09

Beadell Brasil Ltda

In progress

1,445.09

Tucano

Exploration license extended

14,660

29 Apr 2015

27 Apr 2018

Gold

Pedra Branca do Amapari e Porto Grande

858.017/04

Beadell Brasil Ltda

In progress

2,527.46

Tucano

Exploration license extended

9,436

29 Apr 2015

27 Apr 2018

Manganese

Porto Grande

850.852/87

Mineração Vale dos Reis Ltda

Total assignment submitted

3,453.00

Tucano

Exploration license extended

14,658

29 Apr 2015

27 Apr 2018

Molybdenum

Amapá e Macapá

858.044/09

Mneração Serra da Canga Ltda

Total assignment submitted

2,249.40

Tucano

Exploration license extended

14,557

29 Apr 2015

27 Apr 2018

Gold

Pedra Branca do Amapari e Porto Grande

854.262/93

Mineração Tanagra Ltda

Total assignment submitted

9,793.10

Tucano

Exploration license extended

4,120

15 Jan 2016

15 Jan 2019

R$47,594.47

Gold

Macapá

858.080/14

Beadell Brasil Ltda

In progress

4,593.14

Tucano

Exploration license

11056

6 Oct 2016

5 Jul 2019

4 Sep 2019

Gold

Serra do Navio e Pedra Bca Amapari

858.002/16

Beadell Resources Mineração Ltda

In progress

9,408.23

Tartarugalzinho

Exploration license

11,060

6 Oct 2016

5 Jul 2019

4 Sep 2019

Gold

Tartarugalzinho

858.001/16

Mneração Serra Da Canga Ltda

In progress

543.03

Tucano

Exploration license

11,059

6 Oct 2016

5 Jul 2019

4 Sep 2019

Gold

Serra do Navio e Pedra Bca Amapari

858.091/13

Beadell Brasil Ltda

In progress

4,469.33

Tucano

Exploration license

11,279

17 Oct 2016

16 Jul 2019

17 Sep 2019

Gold

Serra do navio e pedra bca amapari

858.098/14

Beadell Resources Mineração Ltda

Pending appeal

8,006.31

Tucano

Exploration license

11,281

17 Oct 2016

16 Jul 2019

17 Sep 2019

Gold

Pedra Branca do Amapari e Porto Grande

858.078/09

Mneração Serra da Canga Ltda

In progress

9,823.41

Tucano

Exploration license extended

663

18 Apr 2017

17 Mar 2020

R$47,741.77

Gold

Serra do Navio e Pedra Branca do Amapari

858.124/13

Beadell Resources Mineração Ltda

In progress

5,813.46

Tucano

Exploration license extended

3,140

14 Aug 2017

14 Jul 2020

Gold

Pedra Branca do Amapari E Porto Grande

858.010/10

Mineração Vale DOS Reis Ltda

In progress

289.94

Tucano

Exploration license extended

290

30 Aug 2017

28 Jul 2020

Gold

Pedra Branca do Amapari e Porto Grande

858.076/09

Mneração Serra da Canga Ltda

In progress

8,807.03

Tucano

Exploration license extended

14,559

17 Jan 2018

17 Dec 2020

R$42,802.17

Gold

Serra do Navio

858.022/14

Beadell Resources Mineração Ltda

In progress

9,348.46

Tucano

Exploration license

363

18 Jan 2018

18 Oct 2020

18 Dec 2020

R$30,008.56

Gold

Porto Grande

858.007/18

Beadell Resources Mineração Ltda

In progress

9,992.91

Tucano

Application for exploration

N/A

Gold

Ferreira Gomes e Porto Grande

858.008/18

Beadell Resources Mineração Ltda

In progress

6,002.44

Tucano

Application for exploration

N/A

Gold

Serra do Navio / Pedra Branca do Amarapari

858.100/14

Beadell Brasil Ltda

In progress

1,376.59

Tucano

Application for exploration

N/A

Gold

Serra do Navio e Pedra Bca Amapari

858.103/11

Beadell Brasil Ltda

Pending appeal

5,530.25

Tucano

Application for exploration

N/A

Iron

Pedra Branca do Amapari

858.099/14

Beadell Resources Mineração Ltda

Pending appeal

9,999.78

Tucano

Application for exploration

N/A

Gold

Pedra Branca do Amapari e Porto Grande

858.081/14

Beadell Resources Mineração Ltda

In progress

4,874,53

Tucano

Application for exploration

N/A

Gold

Porto Grande

855.033/94

Marina Norte Empreendimentos de Mineraação S.A

In progress

3,262.30

Tucano

Application for exploration

N/A

Gold

Ferreira Gomes

851.766/94

Mineração Dorica Ltda

Total assignment submitted / pending appeal

8,982.86

Tucano

Application for exploration

N/A

Gold

Serra do Navio

851.770/94

Mineração Dorica Ltda

4,746.61

Tucano

Application for exploration

N/A

Gold

Serra do Navio

851.771/94

MINERAÇÃO DORICA LTDA

Pending appeal

8,750.06

Tucano

Application for exploration

N/A

Gold

Serra do Navio e Pedra Bca Amapari

858.062/95

Mineração Tanagra Ltda

In progress

8,982.41

Tucano

Application for exploration

N/A

Gold

Macapá

858.082/17

Beadell Resources Mineração Ltda

In progress

1,038.31

Tucano

Application for exploration

N/A

Gold

Pedra Branca do Amapari

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Process

Titleholder

Status

Area (ha)

Project

Phase

License number

Publication date

Renewal date

Expiration date

Tah / June 2018

Substance

District

858.079/17

Beadell Resources Mineração Ltda

In progress

268.94

Tucano

Application for exploration

N/A

Gold

Porto Grande

858.095/17

Beadell Resources Mineração Ltda

In progress

643.77

Application for exploration

N/A

Gold

Porto Grande

858.099/17

Beadell Resources Mineração Ltda

In progress

996.82

Application for exploration

N/A

Gold

Porto Grande

858.050/95

Mineração Tanagra Ltda

Total assignment submitted / Faixa de Fronteira (Brazilian border zone)

9,300.09

Tucano

Application for exploration

N/A

Gold

Macapá

858.051/95

Mineração Tanagra Ltda

Total assignment submitted / Faixa de Fronteira (Brazilian border zone)

6,371.28

Tucano

Application for exploration

N/A

Gold

Macapá

858.079/14

Beadell Brasil Ltda

Leasehold partial

150.08

Duckhead

Partial lease

N/A

Gold

Pedra Branca do Amapari

858.010/99

Beadell Brasil Ltda

Pending appeal

8,634.00

Tucano

Bid

N/A

Pedra Branca do Amapari

850.865/87

Mneração Serra da Canga Ltda

In progress

6,111.83

Tucano

Exploitation application

6,389

24 Mar 2017

N/A

Gold & iron

Pedra Branca do Amapari e Porto Grande

851.439/80

Beadell Resources Mineração Ltda

Final report submitted

9,601.89

Tartarugalzinho

Exploitation application

3,162

10 May 2016

N/A

Gold

Tartarugalzinho

858.000/98

Beadell Brasil Ltda

Positive final report submitted

7,884.08

Tucano

Final report presented

2,949

5 May 2006

N/A

Gold & aluminum

Ferreira Gomes e Porto Grande

858.263/96

Beadell Brasil Ltda

Positive final report submitted

9,690.63

Tucano

Final report presented

1,578

5 May 2006

N/A

Gold

Porto Grande

858.052/02

Mneração Serra da Canga Ltda

Pending appeal

8,064.79

Tucano

Final report presented

1,362

19 Jun 2006

N/A

Gold

Serra do Navio e Ferreira Gomes

858.077/09

Beadell Brasil Ltda

Partial report submitted

123.92

Tucano

Final report presented

14,560

10 Dec 2009

N/A

Gold

Pedra Branca do Amapari e Porto Grande

858.054/04

Beadell Brasil Ltda

Partial report submitted

971.40

Tucano

License extension requested

859

24 Jan 2005

N/A

Gold

Pedra Branca do Amapari

858.029/13

Beadell Resources Mineração Ltda

Partial report submitted

919.03

Tucano

License extension requested

835

29 Jan 2014

N/A

Gold & iron

Pedra Branca do Amapari e Porto Grande

858.015/12

Beadell Resources Mineração Ltda

Partial report submitted

8,280.13

Tartarugalzinho

License extension requested

1,153

9 Apr 2012

N/A

Gold

Tartarugalzinho

858.032/12

Beadell Resources Mineração Ltda

In progress

8,476.34

Tartarugalzinho

License extension requested

3,643

19 Apr 2013

N/A

Gold

Tartarugalzinho

850.858/87

Mineração Vale dos Reis Ltda

Partial report submitted / total assignment submitted

5,643.84

Tucano

License extension requested

325

6 Jan 2014

N/A

Titanium

Serra do Navio e Pedra Bca Amapari

858.002/11

Mneração Serra da Canga Ltda

Partial report submitted / pending appeal

76.90

Tucano

License extension requested

9,522

4 Jul 2011

N/A

Iron

Porto Grande

851.676/92

Beadell Brasil Ltda

In progress

3,971.41

Tucano

Mining concession

73

13 Apr 2004

N/A

Gold

Pedra Branca do Amapari

852.730/93

Zamin Amapá Mineração Ltda

In progress / leasehold

5,529.63

Duckhead

Mining concession

91

6 Apr 2009

N/A

Gold & iron

Pedra Branca do Amapari

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Figure 4.2 GPR tenement ownership map with tenement status

Source: BDR

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Tenements that are under one of the phases following do not have an expiration date (in some of these cases it depends on an analysis / decision by the ANM. For such reason the wording “N/A” has been used in the schedule with respect to expiration date: Application for Exploration, Partial Leasehold (Duckhead lease), Bid, Exploitation Application, Final Report Presented, License Extension Requested, or Mining Concession. “Bid” means the processes which are currently under a public tender process, in which interested companies have presented their proposals but the ANM has not analysed yet.

The following tenements are discussed further:

• 858.008/2018; Awaiting confirmation from ANM regarding the area requested by BDR (6,002.44 ha) being accepted by ANM.

• 858.010/1999 - An area of 8,634.00 ha was the size of the area that BDR was qualified for in the availability notice. Availability to date is on appeal. The area is exclusive of 930.73 ha of the Zamin mining concession area within the tenement 858.010/1999.

• 852.730/1993 -An area of 5,529.63 ha was the size of the area that BDR was qualified for in the availability notice. Availability to date is on appeal.

4.4 Royalties and agreements

GPR is subject to both federal and state royalties and other royalty agreements in relation to mineral product sales from its Tucano Gold Mine.

In summary these are; (a) the Compensation for Exploitation of Mineral Resources (CFEM), (b) the Control, Monitoring and Supervision of Research Activities, Mining, Exploration and Exploitation of Mineral Resources Fee (TFRM), (c) the Social and Community Development Funds, and (d) a Commodities Royalty.

CFEM is a federal royalty and is calculated over the amount of gross revenue obtained in the sale of mineral products. In respect of gold sales, the company is liable to pay a royalty of 1.5% on gross revenue from production at its Tucano Gold Mine.

TFRM is a state royalty levied by the state of Amapá, in which GPR’s Tucano Gold Mine is located. TFRM is currently calculated based on the grams of gold produced multiplied by the state index rate of BRL2.25 which multiplied by a factor of 0.4. The Company has recently finalized negotiations with the state of Amapá to reduce the factor applied; 0.1 for the 2018 and 2019 years and 0.25 for the years 2020 and beyond. The reduced factors have not yet been enacted into law.

The Social and Community Development funds have resulted from various agreements with the state of Amapá and the municipalities of Pedra Branca do Amapari and Serra do Navio. The municipalities of Pedra Branca do Amapari and Serra do Navio are located nearby the Tucano Gold Mine.

Under the terms of the agreements the Company will pay a maximum 1% royalty over the gross proceeds from gold sales from the Tucano Gold Mine to support the socio-economic and community development of the municipalities of Pedra Branca do Amapari and Serra do Navio.

The Commodities Royalty is payable to the previous holders of 13 tenements acquired by GPR. The Commodities Royalty is levied at 0.75% of commodity sales revenue arising from those 13 tenements minus taxes, transport and insurance expenses and royalties payable.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

4.5 Environmental liabilities

The Project area is an active mining operation centred in dense tropical rain forest located in Amapá State, Brazil, consisting of open pit operations, crushing, milling and CIL processing that has a direct impact on the environment and the ecosystem of the region. The main impacts are on biological resources, air quality, noise and vibration, archaeology, hydrogeology and hydrology (ground and surface water), and community stakeholders.

4.6 Permits and other

All permits are in place for the Project. A recent field trip by Marcelo Batelochi the QP for the environmental - Section 20, has confirmed that all relevant monitoring and reporting to government environmental agencies is being undertaken as per legislation and that the mine site is fully compliant in the areas of:

• Characterization of Flora and Fauna.

• Documentation and retrieval of archaeological artefacts.

• Fauna monitoring programs.

• Vegetation clearing and rehabilitation processes.

• Dust, emissions and noise control.

• Control of drainage and erosion.

• Monitoring of surface and ground water.

• Classification and management of industrial solid and liquid wastes.

• Environmental emergency plans

• Details of funding for community development programs.

• Positive impact on the state economy through employment.

• Development of educational programs for the community.

• Support in maintaining local community infrastructure and services.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

5 Accessibility, climate, local resources, etc.

5.1 Accessibility

The Property is located at latitude 0.85°N and longitude 52.90°W, approximately 200 km northwest of Macapá, the capital of Amapá state in northeast Brazil. Access to the site is made via road or by chartered aircraft.

Road access to Tucano is made using Brazilian federal highway BR-210 from Macapá. The first 100 km of the highway from Macapá is sealed - the balance of the road is unsealed. Total driving time from Macapá to the site is approximately four to six hours, depending the season and on road conditions.

The Tucano mine site is also serviced by a 1,100 m airstrip located approximately 800 m from the main gate. Charter flights from Macapá to the mine take about 50 minutes.

The nearest accessible community to the Tucano operation is the town of Serra do Navio, with a population of approximately 3,300. It is located adjacent to the Amapari River and is approximately 200 km from Macapá. Macapá has a population of approximately 474,000 and lies on the equator on the north side of the Amazon River. Most of the workforce are transported by bus from Serra do Navio, Pedra Branca and other small surrounding towns. Professional staff commute from either Macapá or other cities in Brazil on a fly-in-fly-out (FIFO) basis.

5.2 Physiography

The Tucano project is situated in low hills to the north of the Rio Amapari valley. The original heap leach pad and the plant site was cut from the top of existing hills. The ground drops away quickly to the east of the process plant, first into the valley of Igarapé Taperebá (Taperebá Creek) and then into the valley of the Igarapé William (William Creek). The Taperebá AB pit complex is on the other side of Igarapé William and is bounded on the east by some slightly higher hills. The process plant is at an altitude of approximately 170 m above sea level and the adjacent Igarapé William valley about 110 m above sea level. The hills behind the Taperebá AB pits rise to a height of about 320 m above sea level and form part of a ridge running along the eastern side of Igarapé William. The area is located within tropical forest, with a high vegetation density.

5.3 Climate

The Project site is located in an area considered to have an equatorial climate, typified by year round high humidity and rainfall. The wet season occurs from January to June and is associated with frequent torrential rain events of long duration. Less rainfall is received during the remaining months however rainfall events are still frequent. Measurements of site rainfall between 2008 and 2016 inclusive show an average annual rainfall of 2,259 mm. Table 5.1 and Figure 5.1 below show the monthly rainfall measurements for 2008 - 2016. According to the Brazilian Water Agency (ANA), average monthly evaporation rates range from 800 mm (February to April) and 1,900 mm (August to October), with an annual evaporation of 1,424 mm.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Table 5.1 Monthly rainfall at Tucano from 2008 to 2016 inclusive

Year	Jan	Feb	Mar	April	May	June	July	Aug	Sept	Oct	Nov	Dec	Total
2008	262	225	358	292	442	258	319	30	90	77	5	150	2,507
2009	208	404	218	247	413	263	178	51	11	34	5	150	2,181
2010	304	225	262	334	268	160	196	184	57	33	91	186	2,300
2011	249	244	306	320	317	160	238	147	43	95	69	71	2,257
2012	222	281	257	240	182	166	171	109	41	104	52	218	2,043
2013	218	268	196	284	236	103	248	209	104	72	24	151	2,115
2014	196	361	297	303	474	233	145	88	58	133	65	64	2,417
2015	233	318	278	362	367	191	164	45	38	34	29	145	2,204
2016	255	213	447	362	165	267	189	73	64	54	9	208	2,306
2017	331	259	283	276	192	302	88	99	123	130	36	224	2,343
Average	248	280	290	302	306	210	194	104	63	77	38	157	2,259

Figure 5.1 Monthly rainfall at Tucano from 2008 to 2016 inclusive

Source: BDR

5.4 Infrastructure

Infrastructure at the Tucano Mine site includes a central administration building and medical facility, offices (mining, geology, survey, mining engineering, supply, environment, and safety), core shed, nursery, light and heavy equipment workshops, warehouse storage facilities, tire shop, refuelling and washing bays, fire-fighting systems, and explosive magazines. The project also encompasses a CIL gold plant with administration buildings, workshop, a power generation and distribution facility, security office, sample preparation and assay laboratory, storage area, and a reagent preparation facility. Mine subcontractors have separate facilities.

Site communication comprises private branch exchange (PBX) telephone systems and internet links.

All installations are fenced and monitored by armed guards and cameras. The main entrance to the site is located at the main site access road. A 122 person accommodation camp is located near the access road to site. This camp is well equipped and includes a gymnasium, games room, laundry, and restaurant area.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Doctors and nurses are stationed on the site permanently. An established system is in place for emergency removal of personnel if required.

5.4.1 Water

Water is provided from water storage dams and recycled process water. Potable water is processed onsite through a potable water treatment plant, where it undergoes filtration and chlorination to produce water of drinkable quality. The water is transferred from the potable water treatment plant into two potable water tanks.

Process water consists of recycled and other waste streams used in the process, including return water from the Tailings Storage Facility (TSF).

5.4.2 Power

The site is connected to the national electrical grid by a 69 kilovolt (kV), 20 megavolt-ampere (MVA) power line via the state-owned Centrais Elétricas do Amapá (CEA - the local supply authority), and an independent 12 MW continuous rated Aggreko diesel-powered generation system.

5.4.3 Workforce

The mine currently employs a total of 1,116 persons, comprising 416 employees and 700 contractors. With the exception of a few senior management staff and the expats working for contract companies MACA and McKay, all staff (approximately 99%) are Brazilian nationals. The majority of the workforce live in towns in close proximity to the mine site. Managers and professional staff either commute from the state capital Macapá or from other cities in Brazil on a FIFO basis.

Shift workers work a rotating 8 hr shift over a 24-hr period with a roster in days of 6 on 1 off, 6 on 2 off, 6 on 3 off basis. Managers and professional staff either work a roster in days of 19 on, 9 off or 5 on, 2 off. Expats work on a FIFO basis of 6 weeks on, 3 weeks off.

5.4.4 Commercial resources and services

Contracted services and the procurement of goods is managed by the site supply department. Many of the ancillary services required by the site are located in Macapá (e.g. cleaning, transport, and catering). Larger construction projects such as the process plant and tails dam expansion are contracted using companies based in Para and Mina Gerais. GPR will establish long term contracts with companies in Belo Horizonte, which is the hub for many mining industry services in Brazil. The supply of materials is either directly out of Macapá or shipped in from other cities via the port of Belem to the port of Santana located 10 km from Macapá.

5.4.5 Transportation

The mine site is located approximately 200 km from Macapá. This access comprises:

• Sealed road for approximately 100 km, from the city of Macapá to Porto Grande.

• Unsealed road from Porto Grande to Pedra Branca do Amapari, approximately 70 km.

• Unsealed road from Pedra Branca do Amapari to the site, approximately 30 km.

The journey from Macapá to site takes approximately four hours by car during the dry season and up to six hours during the wet season, depending on the condition of the unsealed roads. The unsealed roads are passable all year around and no blockages are experienced except during road maintenance. The transport of personnel, bulk materials, fuel, and other supplies passes over various bridges along the route to the mine. These are mostly wood construction material and are limited to a 45 t of total load (carrier plus load). Most of the bridges have been upgraded recently, allowing for a reliable means of transport to the site.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

6 History

Anglo American discovered a mineralized shear zone in 1994 and undertook extensive exploration between 1995 and 2002. AngloGold completed a feasibility study of the oxide resources in October 2002.

The mine was subsequently acquired by EBX Gold Ltd (EBX) in May 2003. EBX carried out a feasibility study for the oxide mineral resources and a prefeasibility study for mining the sulphide ore.

The property was acquired by Wheaton River Minerals Ltd (which later merged with Goldcorp Inc) in January 2004. Mine construction began in July 2004, with the first gold production in late 2005. The mine then operated as Brazilian company Mineração Pedra Branca do Amapari (MPBA) as a heap leach operation until 2 January 2009, when it was placed on care and maintenance. A significant factor in the decision to close the operation was the increasing occurrence of transition material in the pits and an inability to treat this material economically with the existing process route.

Goldcorp Inc sold the mine to Peak Gold Ltd in April 2007. Peak Gold Limited later merged with New Gold Inc.

The operation suffered from problems with handling the clayey nature of the saprolitic oxide ore and did not reach the predicted recoveries. In response to the operational problems, some test work was conducted on site during 2007 to investigate the potential improvements from installing a wash-plant (integrated with downstream CIL facilities and oxide and sulphide milling circuits) to remove fines and clays and improve heap percolation characteristics. However, it was not possible to reach firm conclusions regarding the potential benefits of the proposed wash plant. As a consequence of the test work ambiguities, continued heap leach improvements and the relative shortage of oxide resources, the plans for the integrated wash-plant and oxide and sulphide milling circuit were put aside in favour of a milling and CIL circuit to treat predominantly sulphide plus remnant oxide material.

Until closure in 2009, mining operations extracted 8.8 Mt of gold ore from four areas, Taperebá (Tap) AB (pits 1, 2, and 3), Tap C, Tap D, and Urucum. Total gold production from the heap leach operations was approximately 292,000 oz.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

7 Geological setting and mineralization

7.1 Regional geology

The South American Precambrian Shield comprises some 50% of the bedrock in Brazil and consists of major Proterozoic deformation zones surrounding cratonic nuclei of Archean age. The three principal cratons are the Guyana Craton in the north, the Amazon (or Guapore) craton immediately south of the Amazon River, and the São Francisco Craton situated between the Amazon Craton and the coast. The cratons are mostly a granite gneiss complex including some highly metamorphosed supracrustal belts, of which greenstone belts represent a small portion. Remoteness and lack of outcrops due to deep weathering prevent detailed stratigraphic and structural mapping across most of the greenstone belts. However, stratigraphic and structural elements typical for greenstone belts worldwide are well recognized in most South American examples.

The consistent north-easterly trend of inferred extensions within granite-greenstone belts across north-eastern South America suggest they have formed during a single major event. The Tucano mineral deposits are located within the Guyana Craton, described by several authors as the Maroni - Itacaiunas mobile belt. This belt runs from Pará and Amapá States of northern Brazil through the Guyanas and into Venezuela. The regional structure is marked by a north-western trending foliation parallel to the main lithologic contacts. Figure 7.1 below shows the basement geology overlain with the Property outline.

Figure 7.1 Geology map of basement geology overlain by the Property tenements

Source: BDR

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

7.2 Property geology

The western part of the Property (about 25% of the area) is underlain by basement gneiss. The balance of the Property area is underlain by ortho-amphibolite and meta-sedimentary rocks of the Vila Nova Group (1.75 to 1.9 billion years old), composed essentially of gneisses, granites, amphibolites, banded iron formations (BIF), massive iron formations, schists, and quartzites (Table 7.1). These units are intruded by granitic pegmatites, diabase dykes, and gabbro bodies. The metasediments are similar to what has been named the Serra do Navio Formation in the nearby area that was mined by Indústria e Comércio de Minerios S/A for manganese. The metasediments hosting the gold mineralization within the Tucano property lie along the limb of a syncline (Figure 7.2).

On the south end of the syncline, the strike is north-south, with dips approximately 65 degrees to the west. To the north the dip becomes vertical to dipping slightly northeast, with the strike trending north-northwest. The gold mineralization is associated with iron and carbonate-rich units of the chemical sedimentary sequence known as the William Formation (Table 7.1). Gold mineralization is predominantly associated with the BIF facies, but carbonate and calc-silicate rocks also host economic mineralization. Sub-economic mineralization may be hosted by any of the lithological units shown in the stratigraphic column (Table 7.1) except the late intrusions.

Laterite and lateritic colluvium are common throughout the area and tend to be thicker over topographic highs, with thick laterites developed in proximity to BIF. Colluvium is characterized by deeply weathered rock that no longer displays any trace of original structure and commonly has angular fragments that indicate limited transport. The colluvium hosts secondary mineralization. A north-south trending shear zone appears to control gold mineralization by carrying gold-bearing hydrothermal fluids or remobilizing existing mineralization; this can be justified by the presence of mineralization found in BIF that was not deformed by shearing. Mineralization is both conformable and not conformable with meta-sedimentary contacts in the William Formation.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Table 7.1 Stratigraphic column (from AMC 2009)

Alluvium / laterites
	Intrusives	Basic	Db	Gabbro / diabase	Dykes of diabase and gabbro
		Acid	Gran. / Peg	Pegmatites
Vila Nova Group	Clasto-pelitic Sedimentary Unit	Quartz Domain	QMX	Muscovite quartzites	Muscovite-quartz schist, muscovite quartzite, locally with fuchsite and / or sillimanite
		Pelitic Domain	QBX	Quartz-mica schists	Quartz-biotite-muscovite schist with garnets, interspaced with lenses of calc-silicates, iron formation and muscovite
	Clasto-chemical Sedimentary Unit	Transitional Unit	RANF	Quartz-grünerite cummingtonite schist w/ garnets, chlorite and biotite	Quartz-amphibole schists and amphibole schists with lenses of silicate facies iron formation and calc-silicates
	Chemical Sedimentary Unit (William Fm.)	Ferruginous Domain	FFER	Iron formation, silicate facies	Garnet-hornblende-grünerite- diopside; magnetite-grünerite- hornblende schists
			BIF-O	Iron formation oxide facies	(Grünerite)-quartz- magnetite / hematite, sometimes with garnets and diopside
			BIF-OS	Iron formation oxide-silicate facies	(Hornblende)-diopside-grünerite- quartz with magnetite; (diopside)- hornblende-garnet, quartz- magnetite-grünerite schist
		Calc- magnesian Domain	CALC	Schists with diopside porphyroblasts	Actinolite-tremolite-diopsides, hornblende-diopsides, amphibole- diopsides, with epidote, biotite and garnets
			RCB	Marble and carbonate schists	Calcic marble, serpentine marble with tremolite, forsterite, fayalite, hastingsite, chlorite and magnetite; actinolite-tremolite- carbonate schists
	Volcanic Unit		MTB or ANF	Metabasics ortho- amphibolites	Plagioclase amphibolite’s, biotite- amphibole schists, plagioclase- cummingtonite-hornblende schists

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Figure 7.2 Property geology map

 

Source: BDR

In Figure 7.3 the Tucano deposit is shown on a lithology background along with some infrastructure. The circled areas show key near mine exploration targets.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Figure 7.3 Tucano deposit location map

 

Source: BDR

7.3 Local geology and mineralization

The mineralization at Tucano occurs in a series of deposits over a 7 km strike length associated with a north-south trending shear zone occurring coincident with a north-south line of topographic ridges. From south to north, these deposits have been named Tap A, B, C, and Urucum. Tap D is a separate structure in the west. The locations of the deposits are shown in Figure 7.3. Higher grades are associated with the more intensely hydrothermally-altered rocks.

Continuous high grade shallowly plunging ore lodes are developed along mineralized shear zone hosts. The geometry and plunge of the ore shoots is interpreted to be controlled by gently plunging F2 fold hinges and more steeply dipping fault intersections.

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Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

The texture and mineralogy along the shear zone indicates a high-temperature hydrothermal system. The interaction of favourable rock type, structure, heat, and mineralized solutions has resulted in the deposition of gold bearing, non-refractory sulphides concentrated near or on major lithological contacts. This mineralized zone exhibits intense hydrothermal alteration, particularly silicification and sulfidation, bearing auriferous pyrrhotite, and pyrite. Alteration is most intense in the proximity of reactive meta-sediments such as BIF, followed by amphibolite, carbonate, schist, and to a lesser extent, calc-silicate rocks. Table 7.2 below shows a description of length, width, depth, and continuity for each deposit.

Table 7.2 Description of mineralized zones in each deposit

Deposit	Pit / area	Total horizontal length of mineralization (m)	Average width (m)	Depth defined by drilling (m)	Continuity of mineralization down plunge (m)
Tap AB	Tap AB1	580	9	300	450
	Tap AB2	600	5.5	350	440
	Tap AB3	420	5.5	460	480
Tap C	Tap C1	800	6	150	150
	Tap C3	450	7	300	150
	Tap C3N	350	5.5	90	150
	Gap	170	11	130	100
Urucum	Uucum North	1,000	4	780	550
	Urucum South	1,300	10	360	750
Urucum East	Urucum East	230	7	75	180
Duckhead	Main Lode	80	12	210	220
	HW Lode	90	15	165	180

7.3.1 Tap AB and Urucum

The Tap AB and Urucum Gold Deposits of the Tucano Gold Project are emplaced within a north-south trending, multiply deformed volcano-sedimentary sequence of rocks that is bounded to the west by the Amapari Granite. The deformation history has resulted in steeply dipping N-S shear zones and the gold mineralization is associated with alteration and sulphide mineralization of various lithologies within the deformed sequence. Mineralization occurs over at least 7 km of strike length and has been interpreted to extend to more than 700 m below the surface. A suite of late post-gold pegmatites has been emplaced throughout the sequence in various orientations and is largely barren of gold.

Mineralization also occurs in colluvium overlying the sequence and this mineralization is spatially related to the bedrock mineralization and is interpreted to be derived thereof. Typically for many deposits of its type, the mineralization often presents as generally somewhat discontinuous and irregularly distributed on the scale of approximately 50 m to 100 m. Figure 7.4 presents a sectional view from Tap AB 30 m apart demonstrating variability in grade, thickness, and orientation of gold mineralization.

Gold mineralization at Urucum is predominantly strata bound to specific sheared lithological units within the BIF and is characterized by strong disseminated and shear fabric pyrrhotite sulphide. The strong association between gold and pyrrhotite results in a highly visual ore in fresh rock that is easily discernible from non-mineralized BIF and other rock.

The Urucum underground resource covers a strike length of approximately 800 m down to a depth of approximately 500 m below the open pit. The lodes form continuous subparallel ore shoots hosted within an approximately 100 m wide Banded Iron Formation (BIF). Three main ore lode horizons have been defined by the drilling and are named Lode 1, Lode 2, and Lode 300 with each lode dipping sub vertically and generally separated by 20 m to 30 m. The average true width of each lode is 4 m.

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Figure 7.4 Tap AB Gold Deposit section 94,550 mN

Source: BDR

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7.3.2 Urucum East

The Urucum East deposit is located east of the Urucum Mine in the northern anticlinal fold hinge of the Tucano mine stratigraphy. The deposit consists of a single, east-west striking and flat (-30°) north dipping sulphide lode which has an average thickness of 7 m and a strike length of 230 m. Mineralization at Urucum East is hosted in a wedge of carbonate and altered banded iron formation located inside a swarm of parallel, east-west striking and -30° north dipping pegmatite dykes which have intruded the host schist unit. The deposit is covered by a blanket of poorly mineralized colluvium up to 10 m thick and weathered to a depth of approximately 50 m below the surface.

7.3.3 Duckhead

Mineralization at Duckhead is controlled by the recently interpreted intersection of steep east-west striking shear zones with a banded iron formation lithological contact to form steeply west plunging high grade shoots. The texture and mineralogy along the shear zone indicates high-temperature hydrothermal alteration, particularly silicification and sulfidation, bearing auriferous pyrite.

7.4 Mineralization weathering types

Deep weathering is present in a majority of the deposits with high grade mineralization extending right to the surface through a layer of colluvium several metres thick. Gold mineralization can be found in the fresh rock at depth, in the saprolite zone created by insitu weathering of the underlying rocks and also in colluvial deposits which overlie the saprolite mineralization as a blanket, spreading out over the hill slopes. The saprolite and the colluvial mineralization are collectively grouped together as ‘oxide mineralization’.

Sulphide zones follow shear plane foliation, often crosscutting the BIF and other host meta-sediments and as bedding parallel lenses dipping either east or west along the limbs of the folded BIF units. Host rocks are poor in sulphide and gold, outside the shears and faulted zones. The accumulation of auriferous massive and / or disseminated sulphides in zones of fractures and folds, and forming plunging ore shoots, often crossing lithological contacts, suggests an epigenetic or remobilization event.

7.4.1 Primary or sulphide mineralization

Sulphide mineralization within fresh rock is only found in drill core and does not outcrop at Amapari. Pyrrhotite and pyrite are the most abundant sulphides. Chalcopyrite, arsenopyrite, sphalerite, and galena occur in trace amounts (less than 1%). At Urucum, the mineralization occurs with intense silicification and pyrrhotite is the dominant sulphide. At Tap AB the gold is associated with masses containing 5% to 10% pyrite, and pyrrhotite only occurs in trace amounts.

The individual sulphide masses are several metres thick and can be elongated on strike along north-northwest or north-south orientations. The sulphides extend in depth along a plunge dipping from 10º to 40º at about N10ºW. The dip ranges from almost vertical at Urucum, to 30°W at the south of the Tap AB deposit. Studies show that the gold occurs as free gold and not tied into the crystal lattice of the sulphide minerals. Mineralization is not confined to any particular lithology, nor is it concordant with lithological contacts.

7.4.1.1 Saprolite mineralization

Intense tropical weathering reaching down 30 m to 300 m has caused the formation of saprolite, that is, an in-situ oxidation of the primary sulphide mineralization and host rocks. The saprolite consists mainly of iron oxides and hydroxides, clay, and silica. Ore zones within saprolite follow the strike, dip, and plunge of the primary sulphide mineralization. Semi-decomposed remnants of the primary mineralization become more frequent with depth.

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Weathering has left much of the saprolite material amenable to free-digging but some of the saprolite has required blasting prior to excavation. The mineralization in these more competent zones may range from entirely oxidized to partially or even completely primary sulphides.

7.4.2 Colluvial mineralization

The colluvial deposits occur along north-south trending ridges cut by William Creek. The creek is at about 115 m elevation and the ridges reach 300 m. The top and slopes of the ridges are covered by alluvial and colluvial sediments. It is difficult to separate alluvial and colluvial sediments in the field and therefore they are collectively named colluvium.

All primary ore shoots are covered by mineralized colluvium, which vary according to the subsurface lithology. The grade of mineralization in colluvium tends to reflect the grade of underlying ore shoots, with patches of low grade or barren colluvium usually reflecting low grade or barren underlying saprolite zones. However, zones of mineralization in colluvium tend to be wider than in underlying saprolite due to mechanical transport and development of some secondary mineralization due to variations in surface soil chemistry.

Deep weathering and intense fixation of iron in the upper portions of the soil often create a laterite horizon. The top of the colluvium is usually a layer rarely more than 1 m thick composed of silty, clayey, and sandy materials, poor in fragments of limonite. Immediately below there is a variable layer up to 10 m thick containing lateritic fragments rich in iron oxide dispersed in a ferruginous clay-sand matrix which becomes rich in manganese at the base.

The colluvium deposits are heterogeneous, reflecting the varied subsurface lithologies. Occasionally there are semi-decomposed, angular fragments of these lithologies within the colluvium.

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8 Deposit types

The gold deposits at the property are hosted within shear zones within a number of different Paleoproterozoic, metasedimentary host rocks. This mineralization style is considered to represent an orogenic, structurally controlled gold mineralizing system. The strong association of gold mineralization within and proximal to a major BIF unit makes the setting of the Tucano deposit unique in the district however the age and shear structural controls show similarities to Birimian Orogenic Gold District deposits in West Africa. Figure 8.1 shows a reconstruction of the Guiana Craton with the West Africa Shield showing the location of the major gold deposits in the area. Deep and highly variable thickness weathering is a feature shared by Tucano with other gold deposits in the Guiana Craton.

Figure 8.1 Reconstruction of the Guiana Craton and West Africa Shield

 

Source: BDR

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9 Exploration

9.1 Airborne magnetics, radiometrics, and DEM

The Property has been covered by an airborne magnetic, radiometric, and Digital Elevation Model (DEM) survey conducted for previous operator, AngloGold in 1998. The survey was flown by LASA Engenharia e Prospecções S.A covered a total area of 10,377 km² on 250 m spaced lines at an elevation of 100 m. Two areas; "AP1" (Eastern Area) and "AP2" (Western Area) were flown. In 2001 Spectrem Air Ltd conducted an airborne geophysics program for AngloGold involving magnetic, electromagnetic, radiometric, and DEM on 200 m spaced lines. In 2004, CPRM - Serviço Geológico do Brasil (Geological Survey of Brazil) flew a government funded airborne magnetic, radiometric, and DEM survey covering the Araguari River Project area in Amapa state. The survey covered 10,872 km² at an elevation of 110 m and a line spacing of 500 m. The justification for the 2004 program was on the basis that geological characteristics indicated potential for gold, base metals (copper and others), and chromium.

During 2015, BDR contracted Southern Geoscience Consultants (SCS) in Australia to merge and re-process both airborne and ground geophysical datasets. For the airborne magnetic datasets, the "AP1" survey was excluded as it was superseded by the higher resolution SPECTREM survey. The processing of the airborne magnetic datasets involved re-projection to a common SAD69 22N datum, re-gridding, merging, and filtering. The merged Total Magnetic Intensity (TMI) filters used included:

• RTP; Reduction to Pole

• 1VD; First Vertical Derivative

• AS; Analytical Signal of the TMI

• ASVI; Analytical Signal of the Vertical Integral of the TMI.

A Total Magnetic Image (AS) from the SCS merged data along with the Property outline is shown in Figure 9.1. BIFs and late cross cutting diorite dykes are shown as magnetic highs in red.

In 2015 SCS also merged all four datasets for Radiometric elements; K, U, Th, and TC (total count) using a grid cell of 100 m for the Araguari dataset and 50 m for the others. K and U were not used from the SPECTREM dataset due to problem with the processing of these channels. Prior to merging the data was scaled and levelled using a histogram approach.

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Figure 9.1 Merged Airborne Magnetic Total Magnetic Image (AS)

 

9.2 Airborne electromagnetic survey

In January 2001 SPETREM Air Ltd flew an airborne electromagnetic (AEM) survey over the Amapa area for previous operator, AngloGold. A total of 4,601 acceptable line kilometres of data were collected on 200 m spaced lines before the bird (sensor) was lost and thus an estimated 750 line kilometres were never captured. In addition to AEM, aeromagnetic, radiometric, and DEM data were also collected. A large number of good bedrock conductors were identified in the largely resistive environment. Most of these conductors seem to be due to carbonaceous units although many of those could also be ascribed to sulphide-rich horizons. A number of conductors were identified as BIF with associated sulphides are of interest as gold exploration targets. High conductivity-thickness product AEM anomalies with a magnetic association are probably indicative of pyrrhotite rich sulphide conductors.

The SPECTREM data did not require levelling or filtering by SCS in 2015 as it was deemed to be of good quality. The data was re-gridded to 40 m to improve image detail.

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Figure 9.2 SPETREM electromagnetic image

 

9.3 Ground geophysics

In 2004 previous operator, MPBA contracted EQUIPE AFC GEOFÍSICA LTDA to carry out ground geophysics using the following methods; Magnetometer, Very Low Frequency (VLF), and Induced Polarization, Time Domain (IP) over the projects of Tap AB, Tap D, Urucum, Vila do Meio, and Urucum East. The first stage of the geophysical survey was to base-line against known mineralization at Urucum, Tap D, and Tap AB already defined by 4 lines drill ranging from 300 m to 800 m in length. This involved magnetometer, VLF, and IP time domain surveys.

At Vila do Meio and Urucum East only, IP and magnetometer were used with 8 lines completed and a total of 6,100 and 8,350 line metres respectively. At the first stage, spacing of the magnetometer readings was 10 m and IP dipole-dipole arranged on a 25 m spacing on short lines up to 400 m with a combination of di-pole and gradient array on the longer lines up to 800 m. At Urucum East and Vila do Meio magnetometer reading spacing was at 25 m and IP dipole was at 50 m with some lines employing a combination of dipole and gradient array. Mineralization from all lines in stage 1 was detected with IP. The magnetometer was excellent in defining the Iron Formation. Conductors in areas without drill information were identified and flagged for follow-up work.

In 2015 SCS processed the ground magnetic data for Vila do Meio for BDR with the aim to provide information in the SPECTREM data gap area however the resolution of the image was considered poor. It was recommended to close the reading spacing to 10 m to improve resolution. A review of the ground IP dataset was not undertaken by SCS.

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9.4 Petrophysics

In 2016, BDR contracted Western Geoscience Pty Ltd to undertake physical property test work on selected drill cores for the Tucano project. Results confirmed that pyrrhotite mineralization was insufficient concentration to be conductive. IP surveys have been proposed to test targets in the Tucano trend for possible sulphide conductors at depth. Table 9.1 below shows the test results of a sample of pyrrhotite rich ore from Urucum along with a photo of the core (Figure 9.3).

Table 9.1 Conductivity test results of a sample of pyrrhotite rich core from Urucum

Id	No. reading	Grade (ppm Au)	CondS/m	MSus10-3
1	1		1,305.00	2.61
1	2		1,545.00	2.83
1	3		832.00	2.85
1	4		815.00	2.06
1	5		517.00	2.43
1	6		855.00	2.45
Average		48.1	978.17	2.54

Figure 9.3 Photo of the conductivity test sample

 

9.5 Geochemistry

Previous operator, AngloGold had conducted extensive soil sampling across the Property. Results show soil anomalism over many areas, some of which require base line levelling due to a several methods of analysis and various detection limits being used. BDR has assessed this data along with an interpretation of the geophysics and geological mapping as part of its exploration target generation and ranking review. Figure 9.4 below shows the coverage of existing soil sampling and tenement status. It illustrates that for many tenements, first pass soil sampling is still to be completed. BDR has undertaken soil sampling in its exploration tenements using 400 m spaced lines and 40 m sample intervals with infill sampling closing down to 200 m spaced lines. Soil samples around 1 kg are collected from the B soil horizon using a post hole digger to a depth of 30 cm to 50 cm. Table 9.2 below shows the breakdown of sampling work undertaken by tenement.

BDR carried out an extensive program of stream sediment sampling in 2017 with a method of fine fraction (200 mesh) BLEG analysis. Samples were collected at approximately 700 m spaced centres, typically 1 sample per tributary stream. The sampling involved collection of one composite of 35 sample points along a 100 m stretch of the stream collection point. The method has so far proved effective in identifying a large anomaly in the Serra da Canga East area and is now the subject of a follow-up soil grid sampling program. Figure 9.5 shows the extent of the Serra da Canga stream sediment anomaly.

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Figure 9.4 Tenement status map showing the coverage of soil sampling

 

Table 9.2 Number of exploration samples collected by sample type and company

Company	Sample type	No. samples
Anglo Gold	Soil	78,134
Anglo Gold	Stream sediment	1,349
Anglo Gold	Rock chip	3,674
BDR	Soil	3,907
BDR	Stream sediment	263
BDR	Rock chip	1,073

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Figure 9.5 Serra da Canga East stream sediment anomaly

 

9.6 Geological mapping

The tenement package has been extensively mapped by the previous company MPBA and work continued under the direction of BRD's regional exploration team. Outcrop is generally scarce however the composition of the soil and colluvium cover gives indicators to the underlying geology and this in combination with the geophysics aids in the geological interpretation. Mapping usually occurs in conjunction with soil sampling along defined sampling lines.

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10 Drilling

10.1 Drilling methods

Auger, RAB, RC, and Diamond Drilling methods are employed at Tucano. Auger drilling is undertaken using a three-man crew operating a man-portable motor driven auger drill with a barrel cutting bit to collect a core sample of loose friable material. While unable to penetrate hard laterite or large rock fragments this can penetrate up to 21 m in softer material with average depths of around 4 m commonly achieved. This is also a useful tool aid for mapping as Tucano and the surrounding district is covered by an extensive colluvium blanket and outcrop is often rare.

RAB drilling is undertaken by a mine owned track mounted PW blasthole rig trailing an air compressor, it is used exclusively by geology for both grade control and exploration. The rig is capable of penetrating to 30 m but is generally used to test to depths of around 4 m to 8 m. It is used where access is difficult to undertake grade control in the mine with RC drilling. Also provides a rapid means to drill test exploration targets with a minimal amount of clearing required, especially where there is extensive laterite development in the regolith.

RC Drilling uses a track mounted Schramm T60 with a track mounted auxiliary booster operated by McKay Sondagens (a Brazilian company established by the Australian owned parent company McKay Drilling). The rig has a mounted cyclone and cone splitter which are both accessible for cleaning during drilling operations. The rig has previously drilled to depths of 400 m and is the main tool used for in-pit grade control as well as resource extension and definition work.

Diamond drilling utilizes two skid mounted Brazilian Maquesonda Mach 1200 diamond rigs operated by Brazilian drilling contractor Geosol. The rigs have previously drilled up to 878 m on site and are mainly employed to drill deeper underground resource extensions.

10.2 Surveys and survey coordinate system

Drillhole collar locations and the collar azimuth and dip of the drill rods string are surveyed using a Total Station Leica 407 in the SAD69 22N coordinate system. Downhole surveys for the RC drilling are undertaken using a Reflex Gyro tool. Downhole surveys are collected in diamond holes utilizing a Reflex Maxibore II. These surveys are generally undertaken every 100 m as the hole progresses (i.e.: 0 - 100 m, 0 - 200 m, 0 - 300 m, etc.) to monitor drilling deviation. Maxibore survey data is collected on 3 m intervals during each survey. Geology staff check survey data for repeatability.

10.3 Diamond drill core

Diamond core drilling utilizes both HQ2 and NQ2 diameter core. Normally holes are started utilizing HQ2 size in oxide and are reduced to NQ2 once fresh rock is encountered. Drill core is orientated with a Reflex ACT III orientation tool in fresh rock. Core orientations and depths are checked by site geologists and core yard personnel for accuracy before the core is photographed, geologically and geotechnically logged.

Geological logging is done using "Logchief" software linked to the Datashed SQL database that captures lithology, structure, mineralization, alteration, weathering, and material resistance / hardness. Core is routinely measured for bulk density (SG) with the oxide tested immediately after drilling to determine wet bulk density before being dried to determine dry bulk density. The method of density measurement used is the Jolly Method (see Section 10.7). Once the core has been bulk density tested and photographed, it is cut in half for assay. The remaining half is stored in purpose built, undercover racks at the site core yard.

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10.4 Recovery

Core recovery is measured along with rock quality designation (RQD) in core run intervals. Average recoveries of core from all Tucano deposits is 94%. Both core recovery and RQD are recorded together in the drillhole database.

10.5 Drilling orientation

Owing to the N-S strike of the Tucano system, drillholes are typically orientated to either Azimuth 090° or 270°, dependent on the dip of the lodes and / or availability of accessible sites. Dips of -50° to -90° can be achieved by the T60 Schramm RC rig whereas dips of -55° to -90° are possible with the diamond rigs. The vast majority of holes are orientated with dips of around -60°.

10.6 Drillhole spacing

Grade control drilling is carried out on a nominal 10 mN x 12 mE spacing with infill to 10 mN x 6 mE in zones of lode mineralization. In-reserve pit drilling has been carried out on a spacing of between 20 m and 40 m within the plane of the mineralized structure. Deeper underground intercepts have been completed on a long-section spacing of between 40 m and 80 m.

10.7 Sampling

1 m RC samples were obtained by an adjustable cone splitter attached to the base of the cyclone (1.5 kg to 6.0 kg) and were utilized for both lithology logging and assaying.

Diamond core is predominantly sampled at 1 m intervals, with some sampling on geological intervals (0.6 m to 1.4 m). Diamond core was used for structural, geotechnical, and density measurements as well as lithology logging and assaying. Density measurements were completed for both oxide and fresh whole core with the oxide densities calculated before and after drying to determine wet bulk density, dry bulk density, and moisture content.

The method of density measurement used was the Jolly Method. The weight of an intact 15 cm length of whole core is measured in air and again in water. The density is calculated by the formula SG = weight in air / (weight in air - weight in water). As oxide and transitional rock normally contains moisture, the sample is wrapped in plastic film and weighed while the core is still wet / moist, soon after being received from the drill rig. This enables a wet density to be calculated. The core is then oven dried, and the procedure repeated to attain the dry density.

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11 Sample preparation, analyses, and security

11.1 Sample preparation

Sample preparation is completed at the Tucano Mine sample preparation facility.

Core samples are dried at 105°C, crushed to -8 mm then to -2 mm and split to 0.9 kg to 1 kg before being pulverized to 1 mm. The 1 mm pulverized sample is quarter cut to between 200 g to 400 g before being pulverized to 95% passing 105 microns (μm). The final pulp is quartered again to achieve a sample of 100 g to 200 g and is sent to SGS laboratories in Belo Horizonte for fire assay. RC samples are dried at 140°C, crushed to -2 mm (if aggregated) and riffle split to 1 kg. The 1 kg sample is then pulverized to 1 mm and quarter cut to between 200 g and 400 g. This sample is then pulverized to 95% passing 105 μm and quarter cut to a 100 g to 200 g sample to send to SGS Geosol in Belo Horizonte.

11.2 Sample security

Samples are securely sealed and stored onsite, until delivery to Macapá via the company contracted driver, who then also delivers the samples directly to airlines cargo dispatch facility for delivery to Belo Horizonte. Sample submission forms are sent with the samples to the laboratory and the laboratory emails a confirmation that the samples have been received along with a job number for tracking purpose.

11.3 Sample analysis

Sample analysis is completed by both the site laboratory and an external laboratory (SGS Geosol Laboratorios Ltda in Belo Horizonte, Brazil). SGS Geosol Laboratorios has been accredited by ABS Quality Evaluations Inc and complies with the requirements for ISO 9001-2015. Further details of certifications for SGS Geosol Laboratorios can be found at their website: www.sgsgeosol.com.br.

Sample analysis is undertaken at both laboratories using a 30 g charge, fire assay with an AAS finish. For holes analysed at the Site laboratory, intervals that yield positive gold results have the corresponding pulverized sample sent to the external laboratory for analysis using the same methodology. Received results from the external laboratory then supersede the results of the Site Laboratory in the drillhole database and are used as an additional means to monitor site laboratory performance.

11.4 Sample QAQC

Samples were sent to SGS Geosol in Belo Horizonte for analysis. Certified standards were inserted every 20^th sample to assess the accuracy and methodology of the laboratory. Field duplicates were inserted every 20^th sample of diamond core only to assess the repeatability and variability of the gold mineralization. The field duplicate sample used is the remaining half of the core from the same sample interval. No field duplicates are produced for the RC drilling which represents the majority for drilling completed. A blank standard was inserted at the start of every batch of approximately 150 samples. In addition, the contract laboratory’s SGS Geosol also carried out their own internal standards and laboratory duplicates for each lot. Results of the QAQC sampling were assessed on a batch by batch basis and were considered acceptable.

11.5 QAQC data analysis

Assay QAQC data analysis for Beadell Brasil was undertaken by Maxwell Geoservices (Maxwell) of Fremantle, WA. Available QAQC data for gold assays from BDR’s various drilling programs within the date range of May 2012 to September 2017 is presented in the following sections.

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In early 2016 the grade control data, which was stored in Minesight Torque database, was integrated to the Maxwell Datashed (Datashed) database. The Minesight Torque database did not store the QC or Standards data for the grade control data and only some of the early batches were fully loaded to Datashed. The Datashed database is therefore incomplete for QC for the period prior to the start of 2016.

Notwithstanding the above, note that previous reports by Maxwell prior to 2015 have been reviewed with only minor issues identified.

The following Tucano Gold Project data has been assessed:

• Standards: The standards assays provide a measure of relative analytical precision. Sample batches which fall outside specified tolerance ranges are flagged and investigated.

• Laboratory Repeats (i.e. represents a second 50 g pulp sample taken from the first 200 g pulp).

• Field Duplicates (i.e. represents a second 3 kg sample split taken from the original RC field sample interval. This sample is taken at the same time as the original sample).

A total of 8 datasets are available (4 deposit areas) for two separate laboratories, with gold assaying completed by both the site laboratory and external laboratories (SGS Geosol Laboratorios Ltda in Belo Horizonte). A total of 6 datasets are relevant to this report and are discussed here. A large number of standards are available for review and a total of 47 have been deleted from the dataset due to low usage (<10 times each).

11.6 Statistical evaluation

The quality control data has been statistically evaluated, and summary plots have been produced for interpretation using the functions available in Maxwells QAQC Report software. The various available reports are summarized below.

11.7 Onsite laboratory

11.7.1 Tap AB grade control and resource dataset

The dataset contains a total of 245,690 drillhole samples from 4,124 batches. A total of 597 QC samples and 7,712 standards have been reviewed. Laboratory and field duplicate ratios are extremely low at 1:435 and 1:7,678 respectively. The ratio of client and laboratory inserted standards is also relatively low at 1:54 and 1:78 respectively.

A total of 85 separate client inserted standards have been recorded as included in the sample stream. Many appear to have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Overall, despite some individual standards performing poorly, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 11 separate laboratory inserted standards have been recorded as included in the sample stream. A total of four have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Consistently with the client included standards, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 36 field duplicate checks are available results demonstrate equivalency between the first and second assays.

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A total of 563 drill sample laboratory duplicates are available and demonstrate equivalency between the first and second assays.

No drill sample pulp checks are available. No drill sample laboratory repeats are available. No drill sample umpire checks have been undertaken.

Figure 11.1 shows a control chart for standard GE076 (outliers included) for the site laboratory for Tap AB.

Figure 11.1 Control chart for standard GE076

Source: BDR

11.7.2 Tap C grade control and resource dataset

The dataset contains a total of 98,631 drillhole samples from 1,542 batches. A total of 26 QC samples and 1,554 standards have been reviewed. Laboratory duplicate ratios are extremely low at 1:3,794. There are no recorded field duplicate samples. The ratio of client and laboratory inserted standards is also relatively low at 1:107 and 1:157 respectively.

A total of 51 separate client inserted standards have been recorded as included in the sample stream. Many (more than eight in total) appear to have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Overall, despite some individual standards performing poorly, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 5 separate laboratory inserted standards have been recorded as included in the sample stream. A total of two have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Consistently with the client included standards, there does not appear to be a consistent pattern of bias in the dataset as a whole.

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A total of 26 drill sample laboratory duplicates are available and demonstrate equivalency between the first and second assays.

11.7.3 Urucum grade control and resource dataset

The dataset contains a total of 133,831 drillhole samples from 2,164 batches. A total of 30 QC samples and 4,423 standards have been reviewed. Laboratory duplicate ratios are extremely low at 1:4,461. There are no field duplicate samples available. The ratio of client and laboratory inserted standards is also relatively low at 1:52 and 1:73 respectively.

A total of 58 separate client inserted standards have been recorded as included in the sample stream. Many appear to have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Overall, despite some individual standards performing poorly, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of five separate laboratory inserted standards have been recorded as included in the sample stream. A total of four have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Consistently with the client included standards, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 29 drill sample laboratory duplicates are available and demonstrate equivalency between the first and second assays.

No drill sample pulp checks are available. No drill sample laboratory repeats are available. No drill sample umpire checks have been undertaken.

11.8 SGS Belo Horizonte laboratory

11.8.1 Tap AB grade control and resource dataset

The dataset contains a total of 48,732 drillhole samples from 812 batches. A total of 13,302 QC samples and 5,488 standards have been reviewed. Laboratory and field duplicate and laboratory repeat ratios are low at 1:4, 1:761, and 1:42 respectively. The field duplicate ratio is low. The ratio of client and laboratory inserted standards is acceptable at 1:23 and 1:15 respectively.

A total of 59 separate client inserted standards have been recorded as included in the sample stream. Many appear to have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Overall, despite some individual standards performing poorly, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 41 separate laboratory inserted standards have been recorded as included in the sample stream. A number have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Consistently with the client included standards, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 24 field duplicate checks are available. Results do not demonstrate equivalency between the first and second assays. A total of 4 paired results are considered poor with differences of up to 167%. The dataset is not sufficiently large to make a meaningful conclusion.

A total of 121 drill sample laboratory duplicates are available and demonstrate equivalency between the first and second assays.

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A total of 822 drill sample laboratory repeats are available and demonstrate equivalency between the first and second assays.

No drill sample pulp checks are available. No drill sample umpire checks have been undertaken.

11.8.2 Tap C grade control and resource dataset

The dataset contains a total of 4,291 drillhole samples from 180 batches. A total of 328 QC samples and 1,256 standards have been reviewed. Laboratory duplicate ratios of 1:1 is unlikely and a database error is suspected. The laboratory repeat ratio is 1:13 which is acceptable. There are only 3 field duplicate samples. The ratio of client and laboratory inserted standards is acceptable at 1:12 and 1:5 respectively.

A total of 39 separate client inserted standards have been recorded as included in the sample stream. Many appear to have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Overall, despite some individual standards performing poorly, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 36 separate laboratory inserted standards have been recorded as included in the sample stream. Many have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Consistently with the client included standards, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 53 drill sample laboratory duplicates are available and demonstrate equivalency between the first and second assays.

Only two field duplicate checks are available therefore no meaningful conclusion can be drawn. 237 drill sample laboratory repeats are available and demonstrate equivalency between the first and second assays.

No drill sample umpire checks have been undertaken. No drill sample pulp checks are available.

11.8.3 Urucum grade control and resource dataset

The dataset contains a total of 19,174 drillhole samples from 326 batches. A total of 7,450 QC samples and 2,576 standards have been reviewed. Laboratory duplicate ratios (6,697) are acceptable at 1:3. There are 149 field duplicate samples available for a relatively low ratio of 1:129. The ratio of laboratory repeats (604) is a relatively low 1:32. The ratio of client and laboratory inserted standards is also relatively low at 1:23 and 1:11 respectively.

A total of 53 separate client inserted standards have been recorded as included in the sample stream. Many appear to have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Overall, despite some individual standards performing poorly, there does not appear to be a consistent pattern of bias in the dataset as a whole.

A total of 32 separate laboratory inserted standards have been recorded as included in the sample stream. Many have been used in small numbers and through limited date ranges and this can be considered less useful as reliable indicators of overall laboratory performance. Consistently with the client included standards, there does not appear to be a consistent pattern of bias in the dataset as a whole.

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A total of 441 drill sample laboratory repeats are available and demonstrate equivalency between the first and second assays.

A total of 42 drill sample laboratory duplicates are available and demonstrate approximate equivalency between the first and second assays albeit this may be affected by a single high-grade outlier with poor repeatability.

A total of 149 drill sample field duplicates are available and demonstrate approximate equivalency between the first and second assays and again this appears be affected by several high-grade outlier with poor repeatability.

No drill sample pulp checks are available. No drill sample umpire checks have been undertaken.

Figure 11.2 shows a control chart for Standard G311-5 with outliers included from SGS Geosol date from Urucum.

Figure 11.2 Control chart for Standard G311-5

Source: BDR

11.9 QP conclusions

The QP has reviewed the summary reports and Plots produced by Maxwell for the 2012 - 2017 dataset. Previous reports by Maxwell up to 2015 have also been reviewed with only minor issues noted (although not discussed here). The analysis has not been undertaken by the QP. The following are the conclusions and recommendations of the QP.

• Standards, field duplicates and laboratory repeats have been reported for 8 datasets and two laboratories within the date range of 16 May 2012 to 8 September 2017.

• Standards have been reported with charts displaying assays plotted by date and by instance and plotted according to whether they are client standards submitted by site, or laboratory standards included as part of laboratory QC processes.

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• Several standards in the different reporting groups appear to have used only during relatively brief periods of time of 6 months or less and may not be reliable indicators of laboratory performance.

• Overall, a number of standards appear to be been mislabelled having assays which are significantly different to the expected results. Some standards charts in each group show a negative or positive bias in the displayed results.

• Field duplicates and laboratory repeats are included in the reports where available. Overall Field duplicates and laboratory repeats appear to be performing within expected limits.

• Insufficient field duplicate assays have been undertaken over the period and it is recommended to increase the ratio of field duplicates to approximately 1:20. It is recommended to commence a program of field duplicate samples for RC drilling.

• Despite the poor performance of certain standards, it is the QP’s opinion that there is no systematic bias present in the assay databases sufficient to make the grade estimates inaccurate. Overall, Certified Reference Materials (CRMs) have been inserted at a sufficient rate to monitor ongoing laboratory performance.

• The inclusion of standards and various duplicates in the batches does not appear to follow a consistent pattern with strongly varying ratios in the proportion of the various QC samples to drill samples. It is recommended to undertake a thorough review and implementation of a more systematic QC program.

• A very large number of different CRMs have been used by BDR since 2012. A substantial number of individual CRMs have been used only a few times or over a very limited date range and this provides only a very limited indication of the laboratories performance. It is recommended to rationalize the number of CRMs in use over any one period and ensure that an adequate grade range is maintained.

Based on this review work, the Qualified Person is of the opinion that the dataset provided is of an appropriate standard to use for resource estimation work.

In the opinion of the QP the sample preparation, security and analytical procedures are adequate, and the data is fit for purpose.

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12 Data verification

The Qualified Person, Mr Brian Wolfe, visited the Tucano Gold Project in July 2015. Steps undertaken to verify the integrity of data used in this report include:

• Field visits to the areas outlined in this report including the Tucano Gold Project permit and various gold deposits as discussed in the text.

• Inspection of drill core and RC drill chips.

• Inspection of AC, RC, and DC drilling activities, sampling, and logging procedures.

• Review of BDR’s data collection, database management, and data validation procedures.

• Review of the previous QA / QC reports for the project up to 2015.

• Validation of drilling, geology, and assay database (including checks overlapping intervals, samples beyond hole depth, and other data irregularities).

• Review of Maxwell Geoscience QA / QC reports and charts for standards, blanks, and duplicates.

Based on this review work, the Qualified Person is of the opinion that the dataset provided by BDR is of an appropriate standard to use for resource estimation work.

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13 Mineral processing and metallurgical testing

13.1 Summary

The existing Tucano gold processing plant was designed by Ausenco. After approximately four years of operation, an expansion was planned to install a 3 MW ball mill alongside the 7 MW single stage SAG mill, as well as a tailings thickener to accommodate the increasing proportion of harder sulphide ore being mined. The Expansion Project, which is in construction, is defined as follows:

• An additional 6 MW ball mill to maintain throughput and obtain a finer product size of 80% < 75 μm at a higher proportion of sulphide ore.

• A pre-leach thickener to allow both effective operation of the cyclones and adequate retention time in the CIL recovery plant.

• Additional leach residence time.

• Oxygen addition to CIL.

• Lead nitrate addition to CIL.

The mill sizing calculations determined that a 4 MW mill should be adequate to grind the tonnages in the proposed mine plant to 80% < 75 μm; however, a larger 6 MW ball mill is recommended for the following reasons:

• The selection of the smaller 4 MW ball mill assumes that the SAG and ball mills can operate continuously at 6.5 MW and close to 4 MW, respectively. This is difficult in practice. Selection of the larger 6 MW ball mill will allow optimization, and some “drift” of the respective SAG and ball mill power draws. It is likely that the actual optimum power draws for the SAG and ball mill will be closer to approximately 4 MW and 6 MW respectively.

• A 6 MW ball mill is a better fit to work alongside the 6.5 MW SAG. It is very unusual in SAG / ball mill grinding installations that the ball mill has a smaller motor, than the SAG mill.

• Future operations may decide to remove some of the current processing bottle-necks which limit production to 450 t/hr.

• There is also some doubt as to the exact specific energy of all the sulphides. Selection of a 6 MW mill will ensure the treatment rate is maintained if harder sulphide ore is mined.

• As gold prices increase, and power costs drop, the optimum grind calculation can decrease from 75 μm towards 53 μm, requiring more power.

• Selection of a 6 MW mill allows some standardization of mill components with the existing Outotec SAG mill.

The cyclones are currently not operating effectively, which is affecting grind and recovery. The slurry has a high viscosity which is a function of the high percent solids that is required in the cyclone overflow to maintain adequate residence times in CIL. It is therefore recommended that a pre-leach thickener be installed.

At a throughput of 450 t/hr and a slurry density of 45% solids, a total volume of 17,200 m³ is required to provide a leach residence time of 24 hours. The existing CIL provides 15,400 m³; therefore, one additional leach tank, 15 m × 15 m is required. The new tank will be located ahead of the existing CIL tanks.

As of September 2018, commissioning of the ball mill was complete and the additional CIL tank and oxygen plant remains on track for late-October 2018 and early-November 2018, respectively.

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13.2 Test work review

Three separate metallurgical test work programs have been carried out to date:

• Amapari work program by NewGold Inc. (original heap leach).

• Ausenco Feasibility Study program by Ausenco and Ammtec Limited (original mill).

• Beadell FS program by BDR and their consultants, Ray Walton Consulting Inc. (RWC) and MPH Minerals Consultancy Ltd. (MPH) for the mill expansion.

The Amapari program by NewGold Inc. was focused on a heap leach operation and is not relevant to the grind and cyanidation process later adopted by BDR. The bulk of the test work to support the design of the original mill was carried out by Ausenco at Ammtec Limited (Ammtec). This was later supplemented by an operations review and other test work at SGS Geosol - Laboratórios Ltda (SGS Geosol) to support the design of the expansion. This was supervised by BDR. The latter two programs are summarized below.

13.3 Ausenco FS metallurgical program

A detailed metallurgical test work program was conducted to support the initial feasibility study for the Tucano project. The program was supervised by Ausenco and was carried out by Ammtec in Perth, Western Australia, during the second half of 2010. It included test work to establish:

• The comminution characteristics of four main ore type composite samples of Tap AB oxide, Tap AB sulphide, Urucum oxide and Urucum sulphide, and 15 variability samples.

• The optimum grind size and cyanidation conditions for the four main ore type composite samples.

• The gravity recovery response of the four main ore type composite samples.

• The leach performance of 29 variability samples, including five Tap C oxide, two Tap C sulphide, and two Tap D oxide ores.

• The leach performance of eight samples of ‘spent ore’ (heap leach tailings) and two samples of stockpiled low-grade ore.

• Engineering data that included testing of oxygen uptake, carbon kinetics, slurry viscosity, and lime demand.

• The performance of air / SO₂ process and hydrogen peroxide (H₂O₂) for cyanide destruction.

• The mineralogy of the four main ore type composites.

In addition, Ammtec prepared samples for dispatch to:

• Coffey International Limited to establish physical characteristics of plant tailings.

• ALS Environmental to establish geochemical characteristics of plant tailings.

• FLSmidth Pty Ltd for thickening test work.

13.4 BDR FS metallurgical program

In mid-2016 BDR, together with their consultants, RWC and MPH, reviewed the operation of the plant with specific emphasis on the low gold recovery and coarser grind that was being experienced at the time. Additional test work was performed, and the plant operating data was analysed and compared with the original mill design. It was concluded that certain critical areas required attention. An Expansion Project was defined as follows:

• Additional grinding power provided by an additional ball mill.

• A pre-leach thickener to allow both effective operation of the cyclones and adequate retention time in CIL.

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• Additional leach residence time.

• Oxygen addition to CIL.

• Lead nitrate addition to CIL.

13.5 Testwork Process Development Ltd.

In mid-2016, samples were submitted to Testwork Process Development Ltd. for bottle roll testing as listed below:

• Urucum sulphide

• Urucum

• Tap AB2 oxide

• Tap AB3 sulphide

• Tap AB2 transition

• Spent ore

Leaching curves of gold recovery versus time were produced, with and without oxygen addition. The source of the samples was not recorded. Generally, the results were highly variable and not as expected, therefore an additional program at SGS Geosol was planned.

13.6 SGS Geosol

In January 2017, a more definitive metallurgical test program was initiated. Ten composites were prepared, nine sulphides and one oxide, and were submitted to SGS Geosol for bottle roll testing as listed below:

Table 13.1 SGS Geosol samples

Metallurgical test composite	Lode description	Mine	Pit
BDRMET1	AB1 Flat Sulphide	Tap AB	Tap AB1
BDRMET2	AB1 Trough Oxide	Tap AB	Tap AB1
BDRMET3	AB2 Sulphide	Tap AB	Tap AB2
BDRMET4	AB3 Sulphide	Tap AB	Tap AB3
BDRMET5	URS Sulphide	Urucum	Urucum South (URS)
BDRMET6	URCS Sulphide	Urucum	Urucum Central South (URCS)
BDRMET7	URCN SIB Sulphide	Urucum	Urucum Central North (URCN)
BDRMET8	URCN SCS Sulphide	Urucum	Urucum Central North (URCN)
BDRMET9	URN Sulphide	Urucum	Urucum North (URN)
BDRMET10	URN UG Sulphide	Urucum	Urucum North Underground (URN UG)

A summary of the selected drillhole intervals used to select the samples and sections showing the drillhole locations, within the proposed open pit, was provided. The metallurgical results are summarized in the following table.

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Table 13.2 SGS Geosol bottle roll results

Metallurgical test composite	Gold grade (g/t)	53 μm	75 μm	106 μm
BDRMET1	2.17	98.3	94.9	96.7
BDRMET2	2.41	96.6	96.1	97.6
BDRMET3	2.91	94.6	94.2	91.6
BDRMET4	2.81	97.3	95.7	97.6
BDRMET5	3.44	91.5	91.3	90.3
BDRMET6	2.48	90.4	93.0	91.0
BDRMET7	2.19	96.1	93.6	91.1
BDRMET8	7.14	97.6	95.7	91.9
BDRMET9	1.52	91.2	88.9	88.5
BDRMET10	5.71	96.5	95.7	93.2
Average	3.28	95.0	93.9	93.0

Leaching curves of gold recovery versus time were produced at various grind sizes. The data was analysed by BDR and their consultants and it was concluded that a residence time of 24 hours and a grind size of 80% < 75 μm was optimum. Although it can be seen that recoveries were higher at 53 μm than at 75 μm, the calculations carried out to determine the optimum grind, showed that net operating revenues at both grinds were similar. According to plant experience, recoveries fall significantly with grinds coarser than 80% < 106 μm.

13.7 Conclusions and design criteria

The conclusions and design criteria from the two programs, i.e. the Ausenco and Ammtec FS and the Beadell FS, are described separately as follows.

13.7.1 Ausenco and Ammtec FS program

Conclusions drawn from the Ausenco and Ammtec FS test work program are given below.

• The oxide and sulphide ores have very dissimilar physical and comminution characteristics. The oxide ore has very low competency and low-to-average hardness, whereas the sulphide ores are competent to very competent and average to above average hardness. Co-processing of these ores is best suited to a SAG mill. A single-stage SAG mill is recommended for the first few years, when a high-oxide-low-sulphide blend is planned. A ball mill is to be added around Year 4 when the sulphide content of the blend is expected to rise.

• Gravity test work in the laboratory showed a high recovery (35% to 56%) of gold by centrifugal gravity concentration and intensive cyanide of gravity concentrate. However, the overall gold recovery, including leaching of gravity tails, showed no recovery improvement in recovery over a whole ore leach. Therefore, the gravity circuit was not included in the processing plant design in order to reduce capital costs.

• The Tucano ores are free-milling and are not “preg-robbing” and are amenable to gold extraction by conventional cyanidation.

• As a group, the oxide ores respond in a similar manner to each other, giving very high recoveries (typically >95%) at a grind size of 80% < 75 µm, with moderate reagent consumptions of 0.55 kg/t NaCN and 5.5 kg/t lime (60% CaO).

• The sulphide ores also have high recoveries, but typically 3% lower than oxide ores, at a grind size of 80% < 75 µm, and moderate reagent consumptions of 0.49 kg/t NaCN and 0.57 kg/t lime (60% CaO).

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• The ‘spent ore’ samples stockpiled from the previous heap leach operation had a recovery of 87% at a grind size of 80% < 75 µm, the low-grade ore stockpiled during the previous heap leach operation exhibited a recovery of 93.5% at a grind size of 80% < 75 µm.

• Oxygen addition increases leach kinetics, primarily for the sulphide ores. The use of oxygen is recommended for the high sulphide ore blends planned for the fourth year of operation.

• Pb(NO₃)₂ addition (at 100 g/t) appears to increase the gold leach kinetics for Tap AB oxide, such that extractions previously achieved at 48 hours are achieved at between 12 and 24 hours.

• The planned Tucano ore blends both have thickener specific settling rates of 1.2 t/m². The high sulphide blend requires 5 g/t to 7.5 g/t of flocculant and should achieve underflow solids concentration of 60% to 62% solids. By comparison, the high oxide blend would require twice as much flocculant and be expected to achieve an underflow solids concentration of 50% to 52% solids.

• The air / SO₂ process can be used to reduce cyanide in CIL tailings slurry to <3 ppm CN_WAD and <25 ppm CN_TOTAL.

The process design criteria recommended from this test work program by Ausenco are summarized in Table 13.3.

Table 13.3 Selected design criteria (Ausenco)

Parameter	Unit	Oxide	Sulphide
Comminution
Bond Rod Mill Work Index	kWh/t	7.9	20.0
Bond Ball Mill Work Index	kWh/t	11.3	16.8
Bond Abrasion Index	kWh/t	0.03	0.43
JK Parameter A x b		192	34.9
Ore SG		3.57	3.27
Grind Size P80	µm	75	75
CIL
CIL residence time	h	24	24
Residual CNFREE Level	mg/L	100	100
Tails thickening
Specific Settling Rate	t/m2.h	1.2	1.2
Underflow Density	% w/w	50-52	60-62
Cyanide destruction
Operating pH		8.5-9.0	8.5-9.0
Suggested residence time	h	1	1
CIL Residual CNWAD Level	mg/L	110	110
Tailings CNWAD	mg/L	<50	<50
Tailings CNWAD (target)	mg/L	<10	<10
Solution discharge CNTOTAL limit	mg/L	<0.2	<0.2

13.7.2 Beadell FS program

The relevant studies and evaluations carried out by BDR and their consultants are summarized in detail in the Beadell 2016 FS. An alternative expansion strategy to the original Ausenco plan was recommended to maintain adequate throughput and recovery.

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Conclusions drawn from the Plant Operation Review and the more recent SGS Geosol test work program, carried out as part of the Beadell FS, are given below.

• The nominal throughput would be 3.6 Mtpa, 10,000 tpd, or 450 t/hr of predominantly sulphide ore, at a grind of 80% < 75 μm.

• The new equipment would comprise a 6 MW ball mill, a 28 m diameter pre-leach thickener and an additional leach tank, located ahead of the existing CIL. An oxygen plant would also be included (see Figure 13.1).

Figure 13.1 Proposed Tucano expansion flow sheet

 

The design, sizing, and location of each new item of equipment are described below.

Ball mill

The SRK report, included in the Beadell FS, contains a description of the plant trials, mill sizing, and power requirement calculations. The conclusion was that to achieve the optimum grind of 80% < 75 μm, 7.1 kWh/t for oxide and 24.6 kWh/t for sulphide, would be required. The mill sizing calculations by SRK have determined that a 4 MW mill should be adequate to grind the tonnages in the proposed mine plant to 80% < 75 μm, however, a larger 6 MW ball mill is recommended, the reasons for which are outlined earlier.

Pre-leach thickener

In the current operation, the cyclones are operating inefficiently due to the high viscosity of the feed slurry. Operating experience is that the actual grind size currently being achieved is between 80% < 110 μm and 80% < 150 μm. The high viscosity is due to the high percent solids in the cyclone overflow which is required to maintain adequate residence times in CIL. It was therefore recommended by BDR and their consultants to install a pre-leach thickener. A 28 m diameter thickener was selected.

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Thickening tests were done by Outotec during the Beadell FS. This thickener will provide two functions:

• Allow the cyclones to be operated at a lower percent solid, and hence classify more effectively at the optimum grind of 75 μm.

• Provide a slurry at a higher percent solid to the new leach tank and CIL, and hence maintain adequate residence times.

Additional leach tank

At a throughput of 450 t/hr, and a slurry density of 45% solids, a total volume of 17,200 m³ is required to provide a leach residence time of 24 hours. The existing CIL provides 15,400 m³, therefore, one additional tank, 15 m × 15 m, is required. The additional leach tank will be located ahead of the existing CIL tanks.

Reagents

The benefits of lead nitrate addition are shorter leach time and reduced cyanide consumption, particularly on the oxides ores, explained in Mike Hallewell’s report. Additions fall in the range of 100 g/t to 250 g/t as lead nitrate (PbNO₃). A design figure of 150 g/t has been selected and a lead nitrate mixing, and dosing plant has already been installed.

RWC estimated oxygen consumption using benchmarks from other operations, and discussions with companies specializing in oxygenation. These figures were checked against the previous experience of personnel at Tucano. These benchmarks were also compared against typical specific aeration rates, adjusted for the improved mass transfer efficiency experienced with oxygen. A final check was made against an “oxygen uptake rate‟ measured using dissolved oxygen meters on slurries during a previous test work program. A plant capacity of 7.5 tpd of oxygen is recommended, i.e. approximately 0.75 kg of oxygen per tonne of slurry.

13.8 Predicted recoveries

Using the recent data from the SGS Geosol testing and plant operating experience for sulphides, and previous data on oxide leaching from the Ausenco BFS, it is recommended that gold recoveries of 93.9% for sulphide and 97.0% for oxide are used in the economic evaluation. The numbers should be reduced by 0.6% to account for the soluble gold loss that occurs in CIL, i.e. 93.3% for sulphide and 96.4% for oxide.

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14 Mineral Resource estimates

The Mineral Resources for the Property have been estimated by the following QPs Mr Brian Wolfe, Principal Resource Geologist of International Resource Solutions Pty Ltd, takes responsibility for the Tap AB and Urucum estimates of the open pit MIK models and Mr Marcelo Batelochi, Principal Resource Geologist of Soluções em Geologia e Mineração Ltda, takes responsibility for the remaining estimates.

The Mineral Resource estimates for the Tucano operation were reported in accordance with NI43-101 and the Canadian Institute of Mine Metallurgy and Petroleum Definition Standards for Mineral Resources and Mineral Reserves 2014 (CIM Standards).

The QPs are not aware of any known environmental, permitting, legal, title, taxation, socioeconomic, marketing, political, or other similar factors that could materially affect the stated Mineral Resource estimates.

The Tap AB Open Pit Mineral Resource has been revised as of 30 June 2017 based on new drilling completed in 2017, an updated US$1,500 optimized pit shell using Measured, Indicated, and Inferred resources and the 30 June 2017 topographic surface.

The Tap AB Underground Mineral Resource was revised as of 30 June 2017 by reporting the 31 December 2016 model below a US$1,500 optimized pit shell that was generated using the 30 June 2017 Tap AB Open Pit Mineral Resource model.

Models for the Tucano Gold Project that have not been revised since 31 December 2016 include the resources listed below. The reporting of Mineral Resources in relation to these models was released by BDR as a JORC compliant resource on the ASX dated 11 April 2017, "Ore Reserve and Mineral Resource Update". The reporting used reserve pit designs and optimized resource pit shells applicable at the time of reporting.

• Urucum Open Pit

• Urucum South Underground

• Urucum North Underground

• Tap AB Underground

• Tap C

• Duckhead

• Urucum East

The reporting of Mineral Resources in relation to these models was released by BDR on the ASX dated 11 April 2017, "Ore Reserve and Mineral Resource Update". The details of this modelling are discussed below. These have been reviewed and are now reported compliant with the CIM standards.

Gold grade estimation for Tap AB and Urucum Open Pit resources was done using a combination of Multiple Indicator Kriging (MIK) and Ordinary Kriging (OK) and were constrained by grade shells derived from indicator kriging of drillhole gold grades.

Grade estimation was completed for Tap AB Underground, Urucum Underground, Tap C, Urucum East, and Duckhead using OK. OK estimates were constrained within manually interpreted mineralization wireframes and in the case of Duckhead, nested grade wireframes were used to tightly constrain the higher grades in the core of the Main Lode.

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Table 14.1 Summary of estimated Mineral Resources (as of 30 June 2017)

Classification	Tonnes (kt)	Au grade (g/t)	Contained Au metal (koz)
Measured	18,448	1.41	835
Indicated	22,293	1.96	1,405
Measured and Indicated	40,742	1.71	2,240
Inferred	16,351	2.19	1,150

Notes:

• The June 2017 Tap AB UG Resource includes 173 kt at 4.68 g/t Au of Inferred Oxide within the Inferred Primary Category.

• CIM Definition Standards (2014) were used for reporting the Mineral Resources.

• Mr B. Wolfe, MAIG is the Qualified Person under NI 43-101 and takes responsibility for the Mineral Resource estimate for the Tap AB Open Pit and Urucum Open Pit Resource. Mr M. Batelochi, MAusImm (CP) is the Qualified Person under NI 43-101 and takes responsibility for all remaining resources.

• Open pit Mineral Resources are constrained by an optimized pit shell at a gold price of US$1,500 oz. The cut-off grade (COG) applied to the open pit resources is 0.5 g/t Au. The COG applied to ore stockpiles is 0.5 g/t Au; with the exception of Marginal ore which has a grade COG 0.3 g/t Au.

• The underground COG is 1.2 g/t Au.

• Drilling results are up to the 30 June 2017 for the Tap AB Open Pit resource. For all other deposits, drilling results up to the 31 December 2016, excluding the Urucum North Underground resource which has drilling results up to September 2015.

• Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

• Mineral Resources are inclusive of their derived Mineral Reserves.

• The numbers may not compute exactly due to rounding.

14.1 Data used

14.1.1 Drillhole database

The Mineral Resource estimation was based on the available exploration drillhole database which was compiled in-house by BDR.

The database includes samples from trenching, costean, auger, channel sampling, RAB, RC, and diamond drilling. RAB, RC, Auger, and diamond drilling have been included for the purpose of modelling and grade estimation at Tap AB and Urucum with all other data excluded from the estimates Tap AB and Urucum samples were composited to 3 m for the purpose of the June 2017 MIK Open Pit Resource estimate. Holes were sampled at 1 m intervals. The below table shows the breakdown of composites by hole type and total sample length.

A total of 5,836 drillholes were selected at Urucum, 179 drillholes at Urucum East, 5,357 at Tap AB, 8,995 at Tap C and 4,815 at Duckhead. The data type and number of drillholes in the resource area are shown in Table 14.2.

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Table 14.2 Drillhole data used in the Open Pit Mineral Resource estimates

Deposit name	Drillhole type	# holes	# composite assays	Metres
Urucum MIK Open Pit.	Auger	163	163	304
	RAB	2,017	5,042	12,842
	RC	3,356	22,168	59,506
	Diamond	300	4,238	11,258
Tap AB MIK Open Pit	RAB	0	0	0
	RC	4,742	28,770	77,581
	Diamond	615	6,142	16,200
Tap C OK Open Pit	RAB	2,098	4,628	8,204
	RC	2,346	19,928	36,701
	Auger	689	1,250	2,156
	Trench	3,691	37,600	71,602
	Channel	38	292	550
	Diamond	133	1,930	3,478
Duckhead	RAB	391	983	1,805
	RC	1,010	6,661	12,362
	Auger	108	174	285
	Channel	14	93	171
	Trench	33	310	591
	Diamond	66	415	770
Urucum East	Auger	71	176	174
	RC	8	51	51
	Diamond	100	669	665

Note: Urucum and Tap AB Open Pit Resources were composited to 3 m, Tap C and Duckhead composited to 2 m, and Urucum East to 1 m.

Plan view of the drilling for Urucum and Tap AB are shown in Figure 14.1 below.

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Figure 14.1 Plan view of drilling Urucum (left) and Tap AB (right)

Source: B. Wolfe (International Resource Solutions Pty Ltd)

14.1.2 Bulk density

A total of 113,629 density measurements of diamond core were collected from the property. The collection of bulk density measurements is described in Section 10. The spacing and distribution of the measurements allows for density to be estimated throughout the block models.

For the Urucum deposit bulk density was estimated using OK and Nearest Neighbour (NN). For the Tap AB deposit bulk density was estimated using OK. For the Tap C deposit bulk density was estimated using Inverse Power Distance (IPD). For the Duckhead deposit bulk density was estimated using NN.

Table 14.3 shows the number of density samples collected at each deposit by weathering type.

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Table 14.3 Summary of bulk density measurements by weathering type

	# samples oxide	# samples transitional	# samples fresh	# samples total
Urucum	414		37,583	37,998
Tap AB	2,755		30,347	33,102
Tap C	7,167	132	30,714	38,013
Duckhead	739		3,777	4,516

Table 14.4 to Table 14.7 compare the original density measurements with the results from the block model estimate by lithology for each deposit. Blocks are in general agreement with the original measurements.

Table 14.4 Urucum: Comparative statistics of bulk density

	Measured density		Estimated density
Lithology	Oxide t/m3	Fresh t/m3	Oxide t/m3	Fresh t/m3
Colluvium	1.90-2.00		1.60
BIF	2.08	3.30	2.17	3.34
Carbonate	2.27	3.10	2.22	3.10
Schist	2.03	2.80	2.17	2.90
Pegmatite	2.27	2.77	1.79	2.79
Amphibolite			1.79	3.20

Table 14.5 Tap AB: Comparative statistics of bulk density

	Measured density		Estimated density
Lithology	Oxide t/m3	Fresh t/m3	Oxide t/m3	Fresh t/m3
Colluvium	1.89	-	1.80	-
BIF	2.76	3.35	2.23	3.34
Carbonate	2.27	3.11	1.90	3.10
Schist	1.99	2.89	1.92	2.90
Pegmatite	2.27	2.79	2.06	2.79
Amphibolite	1.76	3.39	1.93	3.20

Table 14.6 Tap C: Comparative statistics of bulk density

	Measured density			Estimated density
Lithology	Oxide t/m3	Trans t/m3	Fresh t/m3	Oxide t/m3	Trans t/m3	Fresh t/m3
Colluvium and soil	1.75			1.92
Amphibolites / diabases	1.85	2.89	3.17	2.22	2.79	2.98
Banded iron formation	2.32	2.17	3.39	2.50	3.23	3.36
Calcsilicates	1.95	2.20	3.13	1.76	2.37	3.19
Cabonates / hidrot. alter.	2.02	3.13	3.06	2.01	2.83	3.11
Schists and quartzites	2.40	2.70	2.79	1.79	2.61	2.76
Iron formation	1.77	3.13	3.34	2.47	3.08	3.32
Granites and pegmatites			2.78	2.03	0	2.87

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Table 14.7 Duckhead: Comparative statistics of bulk density

Lithology	Measured density		Estimated density
Description	Oxide g/cm3	Fresh g/cm3	Oxide g/cm3	Fresh g/cm3
Schist	2.74	2.85	2.59	2.84
Carbonate	2.37	3.09	2.77	3.07
BIF	2.91	3.19	2.95	3.05
Amphibolites	2.99	3.04	3.00	3.04
Outer schists amphibolite	3.00	3.01	2.98	3.03

14.2 Interpretation and modelling

There are four deposits, Urucum, Tap C, Tap AB, and Duckhead.

For modelling purposes Urucum is further subdivided into Urucum North open pit and underground, Urucum South open pit and underground, and East Urucum models. Tap AB is also split into open pit and underground models. Tap C is also divided into two areas: Tap C (other) and Tap C (C3N and C3W).

14.2.1 Geological interpretation

14.2.1.1 Tap AB and Urucum

Based on grade information and geological observations, colluvium and oxidation boundaries have been interpreted and wireframes modelled to constrain resource estimation. Interpretation and digitizing of all constraining boundaries has been undertaken on cross sections orientated at 90º (drill line orientation). The resultant digitized boundaries have been used to construct wireframe surfaces or solids defining the three-dimensional geometry of each interpreted feature.

In a similar fashion, geological logging of drill core and RC chips has been input into the construction of lithological wireframes for the project. At Urucum and Urucum East, modelled lithologies comprise pegmatites and granites, BIF, schist, and carbonate rocks. At Tap AB an additional lithology amphibolite has also been modelled. Schematic cross sections are presented in Figure 14.2 and Figure 14.3. Note topography is pre-mining topography.

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Figure 14.2 Urucum Gold Deposit cross section displaying geology and coded block model

Note: Section 99,050 mN

Source: B. Wolfe (International Resource Solutions Pty Ltd)

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Figure 14.3 Tap AB Gold Deposit cross section displaying geology and coded block model

Note: Section 93,930 mN

Source: B. Wolfe (International Resource Solutions Pty Ltd)

14.2.1.2 Tap C

Although reported as one Mineral Resource, the geological model was developed differently in different zones. For the purpose of geological domaining the area is divided into the C3N and C3W zones and “other”.

For C3N and C3W zones, the lithological domains were manually modelled based on 10 m space sections. Outside of these zones, only the pegmatite that cross cuts the mineralization was modelled using a wireframe solid. The remaining lithologies were estimated using a NN approach.

Figure 14.5 shows the lithological solids as used for the C3N and C3W areas.

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Figure 14.4 Isometric view of Tap C lithology solids

Note: Tap C Model Area; Green = Schist, Brown = BIF, Blue = Carbonates, Red = Pegmatites. As used for areas C3N and C3w.

Source: BDR.

14.2.1.3    Duckhead

The Duckhead geological model was rebuilt on 10 m spaced oblique NE-SW sections cut oblique to the strike of the stratigraphy. The modelling approach used was the same as Tap AB and Urucum. Figure 14.5 below shows the main banded iron unit with internal carbonates, schists, and amphibolites.

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Figure 14.5 Plan view of Duckhead showing the modelled lithological units

Note: Large Green Area= Schist, Minor Green Areas = Amphibolites, Brown = BIF, Blue = Carbonates.

Source: BDR.

14.2.2 Mineralization interpretation

The models were estimated by either OK or a combination of MIK and OK. As such mineralization domains were built based on wireframing on a sectional and plan basis as well as being generated by grade shell via indicator kriging at a selective cut-off grade (COG).

14.2.2.1 Tap AB and Urucum Open pit model

The style of mineralization is discussed in Section 7.3. Typically, for many deposits of its type, the mineralization often presents as generally somewhat discontinuous and irregularly distributed on the scale of approximately 50 m to 100 m. Sectional views, even 30 m apart, demonstrate variability in grade, thickness, and orientation of gold mineralization.

This commonly makes the traditional approach of wireframing on a sectional and plan basis somewhat difficult with multiple plausible geometrical solutions often existing.

To establish appropriate grade continuity, the mineralization models for the Tap AB and Urucum Open Pit were therefore based upon a nominal 0.3 ppm Au indicator mineralization shell estimated using 3 m unconstrained downhole composites. This interpretation is designed to capture the broad mineralization halo that encompasses the geological vein system and is not intended to constrain individual veins or vein clusters. As the main grade estimation technique is MIK with change of support technique, this type of mineralization constraint is deemed appropriate.

The mineralization grade shell was generated by grade shell via indicator kriging at a single COG, 0.3 g/t Au. Grade estimation was into block models with cell dimensions of 2 mE × 2 mN × 2 mRL. Grade shell triangulations were then generated by constraining the block model at a 35% probability cut-off for Tap AB and a 31% probability cut-off for Urucum. The probability cut-offs may be considered somewhat subjective and may seem arbitrary, however were selected based on extensive review of a range of probability cut-off levels. The selected probability shells are considered optimal to capture the observed continuity and tenor of mineralization while excluding obvious low-grade material. Grade shells were reviewed in multiple orientations and in plan and section view prior to being accepted for grade estimation and block modelling purposes.

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Mineralization estimation domains were thus defined with further sub-division being differentiated on the basis of orientation, flexures in the shear and tenor of gold grade. In the case of Tap AB, a total of seven estimation domains have been defined.

An isometric SE view of the estimation domains is presented in Figure 14.6.

Figure 14.6 Tap AB Gold Deposit isometric SE view

Source: B. Wolfe (International Resource Solutions Pty Ltd).

At Urucum a total of ten estimation domains have been defined with six minor flat lying domains relating to bodies of colluvium having been generated by manual wireframing and four major steeply orientated domains generated via grade shell estimate as previously described. Generalized relationships between the mineralized domains are presented in Figure 14.7.

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Figure 14.7 Urucum Gold Deposit isometric SE view

Source: B. Wolfe (International Resource Solutions Pty Ltd).

14.2.2.2 Tap AB and Urucum Underground model

Both the Tap AB and Urucum Underground models had manually constructed wireframes built on 10 m vertical E-W sections using both drillhole geology and assay grades to drive the interpretation and digitization.

For the Tap AB underground, 21 mineralization domains were built for the purpose of estimating the underground Mineral Resource (Figure 14.8). The digitizing was based on a nominal 1.4 g/t Au COG for drillholes below a 2016 US$1,500 optimized pit shell.

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Figure 14.8 Oblique view of the Tap AB constructed mineralization wireframes

Source: BDR.

For the Urucum North Underground model eight principal mineralization domains were constructed north of 99,400 mN for the purpose of the Mineral Resource estimate. In addition to interpreting these mineralization domains as major planar structures, the internal high-grade shoots within these lodes were also modelled based on detailed long section analysis and field observation of structures in the pit (Figure 14.9). These shoots were modelled as separate domains in the OK estimation process and in total 24 domains were used in the OK estimation. All wireframes were manually built by digitizing on 10 m sections and at a nominal COG of 1.4 g/t Au.

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Figure 14.9 Cross section of all Urucum North Underground domains

 

Notes: Section 99,800 mN.

Source: BDR.

14.2.2.3 Urucum North and South Underground model

For the Urucum South Underground block model, 70 mineralization domains were constructed and updated with new drillhole results as of March 2016. Lodes south of 99,400 mN were modelled manually on 10 m section using a nominal COG of 1.4 g/t Au below a 2016 US$1,500 optimized pit shell.

North of 99,400 mN, for the Urucum North underground block model (reference: urucum_ug_ resource2015_v17.mdl). Figure 14.10 below shows the underground resource Urucum North and Urucum South lode wireframes and the divisor at 99,400 mN.

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Figure 14.10 Long section view of Urucum showing the underground mineralization domains

 

Note: red line shows the divisor between Urucum South and Urucum North at 99,400 mN.

Source: BDR.

14.2.2.4 Urucum East model

The Urucum East deposit consists of two mineralization domains. The first domain is the bedrock mineralization that is cross cut by pegmatite dykes. The second domain is the mineralized colluvium.

Where the pegmatite partially crosses the mineralization wireframe, this is deemed to have introduced zero Au grade. No separate colluvium mineralized envelope was created for gold mineralization within the colluvium but rather the entire lithological unit for the colluvium was used in the gold estimation.

14.2.2.5 Tap C model

For the Tap C deposit 25 mineralization domains were manually digitized and wireframed on 10 m sections using a nominal 0.3 g/t Au COG (Figure 14.11). A low-grade envelope was also constructed using a 0.1 g/t Au envelope using Implicit Modeller in Minesight.

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Figure 14.11 Tap C gold mineralization domains - oblique view

Source: BDR.

14.2.2.6 Duckhead model

For the Duckhead deposit, 24 mineralization domains were modelled and were also split into oxide and sulphide sub domains for gold grade estimation where sufficient data existed. The modelling for Duckhead covers the main Duckhead pit and also the surrounding smaller deposits as shown in Figure 14.12. Due to the extremely high grades in Duckhead it was necessary to create the model using nested grade shells, one inside the other, to restrict the influence of very high grades in the core of the deposit.

The following nominal COGs were used to drive the digitization of the nested grade shells:

• Duckhead Main Lode Extreme Grade Core: >50 g/t Au = Lode 1.

• Duckhead Main Lode High Grade Envelope: >2 g/t Au = Lode 2.

• Duckhead Main Lode Low Grade Envelope: >0.3 g/t Au = Lode 3.

• Duckhead Hangingwall Lode High Grade Envelope: >2 g/t Au = Lode 4.

• Duckhead Hangingwall Lode High Grade Envelope: >0.3 g/t Au = Lode 4.

The rest of the lodes were modelled using a 0.3 g/t Au COG.

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Figure 14.12 Mineralized wireframes of Duckhead and surrounding deposits

Source: BDR

14.3 Compositing

At Tap AB and Urucum, drillhole samples were flagged with the relevant mineralization wireframes, indicator grade shells, topographical surfaces, and the oxidation wireframes described in previous sections. Coding was undertaken on the basis that if the individual sample centroid fell within the grade shell boundary it was coded as within the grade shell. Each domain has been assigned a unique numerical code to allow the application of hard boundary domains if required during grade estimation.

The drillhole database coded within each grade shell or mineralization wireframe was then composited as a means of achieving a uniform sample support. It should be noted, however, that equalizing sample length is not the only criteria for standardizing sample support. Factors such as angle of intersection of the sampling to mineralization, sample type and diameters, drilling conditions, recovery, sampling / sub-sampling practices and laboratory practices all affect the ‘support’ of a sample. Exploration / mining databases which contain multiple sample types and / or sources of data provide challenges in generating composite data with equalized sample support, and uniform support is frequently difficult to achieve.

After consideration of relevant factors relating to geological setting and mining, including likely mining selectivity and bench / flitch height, a regular 3 m run length (down hole) composite was selected as the most appropriate composite interval to equalize the sample support for the Tap AB and Urucum Deposits. Compositing was broken when the routine encountered a change in flagging (grade shell boundary) and composites with residual intervals of less than 1.2 m were retained by addition to the previous composite resulting in a composite file containing composites between 1.2 m and 4.2 m in length.

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At Tap C and Duckhead, downhole composites of 2 m were created inside the mineralization solids and flagged by lode code. Composites with a majority length inside the cross-cutting pegmatite solid were removed to minimize internal dilution. A minimum composite length was set at 0.95 m. Comparison between total sample length and the mean gold grade of flagged and composited data showed negligible bias and reduction in metal content.

The Urucum North Underground estimate used a 1 m composite length. The 31 December 2016 Underground Resources for both Tap AB and Urucum South both used a 3 m composite length keeping consistent with the composite length used in the Open Pit Resource estimates for these deposits. Investigation into composite lengths showed no expressive bias going from a 1 m to 3 m composite length.

14.4 Statistical analysis

14.4.1 Tap AB Open Pit model

14.4.1.1 Descriptive statistics

The composites flagged as described in the previous section were used for subsequent statistical, geostatistical, and grade estimation investigations.

Summary descriptive statistics were generated for all domains (Table 14.8). The grade distributions are typical for gold deposits of this style and show a positive skew or near lognormal behaviour. The coefficient of variation (CV) is moderately high, consistent with the presence of high outlier grades that potentially require cutting (capping) for grade estimation. Table 14.8 shows statistics for composited data and for data following capping outlier grades.

Table 14.8 Tap AB Gold Deposit - Open Pit model summary statistics gold (g/t)

Composited
Domain	10	101	102	110	200	1,000	2,000	3,000
Count	4,096	9,449	189	379	5,833	2,194	4,845	7,927
Minimum	0.001	0.001	0.004	0.019	0.001	0.001	0.001	0.001
Maximum	95.1	652.45	9.873	68.837	327.48	551.957	106.94	79.283
Mean	0.541	1.217	0.762	1.341	1.426	4.992	1.302	1.559
SD	2.257	7.177	1.049	4.199	6.290	20.509	3.839	3.561
CV	4.171	5.897	1.376	3.132	4.412	4.108	2.948	2.284
Capped
Domain	10	101	102	110	200	1,000	2,000	3,000
Count	4,096	9,449	189	379	5,833	2,194	4,845	7,927
Minimum	0.001	0.001	0.004	0.019	0.001	0.001	0.001	0.001
Maximum	20	24	9.873	8	30	100	30	36
Mean	0.506	1.121	0.762	1.038	1.311	4.421	1.230	1.536
SD	1.427	2.010	1.049	1.511	2.895	13.619	2.759	3.236
CV	2.819	1.793	1.376	1.456	2.207	3.081	2.242	2.107

SD = Standard Deviation; CV = Coefficient of Variation

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14.4.1.2 High grade outlier analysis

MIK has been selected as an appropriate method to estimate the gold grades for the Tap AB Gold Deposit. However, the grade datasets for the various estimation domains are characterized by moderately high CV values, indicating that high grade values may contribute significantly to the mean grades reported for the various datasets.

It should be noted that while gold grades are not cut or capped for the purposes of MIK estimation the use of cut grades is often employed for variography and the change of support process. As MIK estimates are essentially a series of OK estimates applied to the binary transformation of a series of indicator cut-offs, high grade cutting will have no effect on the resultant MIK estimate unless the high-grade cut is lower than the chosen upper indicator cut-off and this scenario would be considered highly sub-optimal in the context of MIK estimation.

The effects of the highest grade composites on the mean grade and standard deviation (SD) of the gold dataset for each of the estimation domains have been investigated by compiling and reviewing statistical plots (histograms and probability plots). The resultant plots were reviewed together with probability plots of the sample populations and an upper cut for each dataset was chosen coinciding with a pronounced inflection or increase in the variance of the data. A list of the determined upper cuts applied, and their impact on the mean grades of the datasets is provided in Table 14.8. Composite data was viewed in 3D to determine the clustering or otherwise of these highest grades observed in each domain to assess the appropriateness of the high-grade cut. Clustering of the highest grades in one or more areas may indicate that the grades do not require to be cut and need to be dealt with in a different way.

14.4.1.3 MIK cut-offs and indicator class statistics

Indicator Kriging cut-offs or indicator bins were selected for each domain to be estimated by MIK. Cut-offs were based upon population distributions and metal proportions above and below the mean composite value of the proposed cut-off bins. Conditional statistics for data for the two largest domains to be estimated by MIK are listed in Table 14.9. A total of 17 cut-offs were applied to each domain for estimation via MIK. Top cuts have not been applied for the purposes of conditional statistics calculation.

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Table 14.9 Tap AB Gold Deposit - Open Pit Resource indicator class statistics

Domain
Zone 101			Zone 3000
Probability threshold	Grade threshold Au (g/t)	Class mean Au (g/t)	Probability threshold	Grade threshold Au (g/t)	Class mean Au (g/t)
0.225	0.3	0.1667	0.217	0.3	0.1573
0.361	0.4	0.3461	0.355	0.4	0.345
0.469	0.5	0.4442	0.452	0.5	0.4442
0.544	0.6	0.5463	0.522	0.6	0.5453
0.626	0.75	0.6672	0.578	0.7	0.6462
0.688	0.9	0.8224	0.637	0.85	0.768
0.746	1.1	0.9892	0.683	1	0.9199
0.788	1.3	1.191	0.733	1.2	1.0816
0.827	1.55	1.4231	0.783	1.5	1.337
0.854	1.8	1.6698	0.817	1.8	1.6326
0.881	2.1	1.9481	0.843	2.2	1.9846
0.906	2.5	2.2952	0.875	2.7	2.4381
0.924	2.9	2.6887	0.903	3.5	3.1036
0.941	3.5	3.1618	0.927	4.5	4.0061
0.960	4.5	3.9364	0.949	6	5.2296
0.976	6	5.1924	0.968	8.5	7.1726
0.992	11	7.7969	0.987	15	10.9904
Max	Max	17.6222	Max	Max	23.6843

Note: statistics are shown for the two largest domains only.

14.4.2 Urucum Open Pit model

14.4.2.1 Descriptive statistics

Summary descriptive statistics were generated for all domains and the statistical results for all domains are shown in Table 14.10. The grade distributions are typical for gold deposits of this style and show a positive skew or near lognormal behaviour. The CV is moderate to high, consistent with the presence of high outlier grades that potentially require cutting (capping) for grade estimation. This is evident in the domains selected for estimation via MIK.

14.4.2.2 High grade outlier analysis

MIK has been selected as an appropriate method to estimate the gold grades for the main estimation domains of the Urucum Gold Deposit. However, the grade datasets for some of the estimation domains are characterized by moderately high CV values, indicating that high grade values may contribute significantly to the mean grades reported for the various datasets.

The effects of the highest grade composites on the mean grade and standard deviation of the gold dataset for each of the estimation domains have been investigated by compiling and reviewing statistical plots (histograms and probability plots). The resultant plots were reviewed together with probability plots of the sample populations and an upper cut for each dataset was chosen coinciding with a pronounced inflection or increase in the variance of the data. A list of the determined upper cuts applied and their impact on the mean grades of the domains is provided in Table 14.10.

Composite data was viewed in 3D to determine the clustering or otherwise of these highest grades observed in each domain to assess the appropriateness of the high-grade cut. Clustering of the highest grades in one or more areas may indicate that the grades do not require to be cut and need to be dealt with in a different way.

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14.4.2.3 MIK cut-offs and indicator class statistics

Indicator Kriging cut-offs or indicator bins were selected for each domain to be estimated by MIK. Cut-offs were based upon population distributions and metal proportions above and below the mean composite value of the proposed cut-off bins. Conditional statistics for data within each domain to be estimated by MIK are listed in Table 14.11. A total of 17 cut-offs were applied to each Domain for estimation via MIK. Top cuts have not been applied for the purposes of conditional statistics calculation. Note that due to a paucity of samples in Domain 300, the conditional statistics for Domain 400 have been adopted for the purposes of change of support process.

Table 14.10 Urucum Gold Deposit - Open Pit model; summary statistics gold (g/t)

Composite
Domain	100	200	300	400	510	520	530	540	550	700
Count	2,195	14,676	179	7,389	2,667	458	203	596	87	3,161
Minimum	0.001	0.001	0.001	0.001	0.001	0.004	0.01	0.04	0.013	0.001
Maximum	21.907	82.867	32.666	59.99	15.48	7.667	2.91	5.89	5.567	20.584
Mean	1.073	1.225	1.684	1.053	1.094	0.719	0.611	0.905	1.020	1.181
SD	1.656	2.408	3.468	2.221	1.129	0.698	0.442	0.597	0.981	1.403
CV	1.543	1.966	2.059	2.110	1.032	0.970	0.725	0.660	0.961	1.188
Capped
Domain	100	200	300	400	510	520	530	540	550	700
Count	2,195	14,676	179	7,389	2,667	458	203	596	87	3,161
Minimum	0.001	0.001	0.001	0.001	0.001	0.004	0.01	0.04	0.013	0.001
Maximum	15	20	20	20	12	7.667	2.91	5.89	5.567	13
Mean	1.069	1.201	1.613	1.025	1.093	0.719	0.611	0.905	1.020	1.173
SD	1.611	2.031	2.921	1.794	1.114	0.698	0.442	0.597	0.981	1.317
CV	1.507	1.690	1.810	1.751	1.019	0.970	0.725	0.660	0.961	1.123

SD = Standard Deviation; CV = Coefficient of Variation

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Table 14.11 Urucum Gold Deposit - Open Pit model indicator class statistics (selective)

Domain
Zone 400			Zone 200
Probability threshold	Grade threshold Au (g/t)	Class mean Au (g/t)	Probability threshold	Grade threshold Au (g/t)	Class mean Au (g/t)
0.242	0.3	0.1726	0.245	0.3	0.1494
0.371	0.4	0.3480	0.373	0.4	0.3450
0.480	0.5	0.4459	0.468	0.5	0.4466
0.560	0.6	0.5482	0.541	0.6	0.5403
0.623	0.7	0.6470	0.600	0.7	0.6456
0.667	0.8	0.7447	0.642	0.8	0.7511
0.709	0.9	0.8447	0.678	0.9	0.8497
0.742	1	0.9472	0.706	1.0	0.9474
0.781	1.15	1.0655	0.741	1.15	1.0740
0.811	1.3	1.2204	0.771	1.3	1.2279
0.840	1.5	1.3927	0.785	1.5	1.3957
0.863	1.7	1.5959	0.813	1.7	1.5889
0.893	2	1.8436	0.842	2.0	1.8474
0.919	2.4	2.1820	0.871	2.4	2.1520
0.939	2.9	2.6449	0.901	2.9	2.6419
0.959	3.8	3.2882	0.931	3.8	3.3125
0.978	6	4.7374	0.968	6	4.8105
Max	Max	12.1008	Max	Max	9.5410

Note: statistics are shown for the two largest domains only.

14.4.3 Tap AB Underground model

14.4.3.1 Descriptive statistics

Raw statistics of each lode split by oxide and fresh surfaces were generated. The results of this statistical study have shown high variability, some characteristics common in gold deposits with coefficients of variation often high due to dilution caused by internal waste within the lodes. As a consequence of the high variability shown in the gold distribution, a large part of the gold metal is contained in the tail of the distribution and required analysis and treatment of the outliers in the composite dataset. Table 14.12 below shows the raw and composited data for the domains.

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Table 14.12 Tap AB Underground model; statistics by domain

Raw
Domain	Count	Minimum	Maximum	Mean	SE mean	SD	CV
LODE10	15,530	0.005	1877.3	1.18	0.12	15.27	12.89
LODE11A	634	0.005	24.94	1.36	0.11	2.71	1.99
LODE11B	1,650	0.01	48.36	1.43	0.07	2.72	1.91
LODE12A	520	0.005	40.63	0.93	0.11	2.48	2.68
LODE12B	286	0.005	8.75	0.58	0.06	0.94	1.63
LODE12C	635	0.005	16.34	0.95	0.06	1.53	1.60
LODE12D	1,210	0.005	33.34	1.44	0.09	3.08	2.14
LODE14A	1,805	0.005	81.38	3.19	0.16	6.65	2.08
LODE14B	2,751	0.005	394.44	5.35	0.45	23.76	4.44
LODE1A	19,588	0.005	498.78	1.58	0.04	6.21	3.93
LODE1B	239	0.010	12.87	1.11	0.12	1.88	1.69
LODE1C	596	0.005	36.64	1.45	0.12	2.94	2.03
LODE3	1,560	0.005	109.53	2.47	0.16	6.43	2.61
LODE4	3,491	0.005	498.78	1.80	0.18	10.78	5.98
LODE5	2,552	0.005	145.60	2.95	0.15	7.82	2.65
LODE6	508	0.005	41.64	0.83	0.11	2.52	3.04
LODE9A	6,165	0.005	143.27	1.30	0.04	3.47	2.66
LODE9B	308	0.010	8.84	0.62	0.04	0.78	1.27
MLODE2A	2,629	0.005	109.07	0.74	0.06	3.22	4.32
MLODE2B	1,166	0.005	33.15	1.03	0.07	2.47	2.39
MLODE2C	258	0.005	20.20	0.91	0.13	2.06	2.25
Composite
Domain	Count	Minimum	Maximum	Mean	SE mean	SD	CV
LODE10	9,693	0.005	527.39	1.16	0.06	6.02	5.18
LODE11A	417	0.005	21.02	1.31	0.12	2.36	1.80
LODE11B	804	0.005	20.43	1.30	0.07	1.94	1.49
LODE12A	593	0.005	15.97	1.01	0.07	1.72	1.71
LODE12B	98	0.044	4.80	0.66	0.08	0.80	1.20
LODE12C	272	0.039	8.37	0.98	0.07	1.18	1.20
LODE12D	549	0.005	28.16	1.62	0.14	3.30	2.04
LODE14A	827	0.005	37.91	2.79	0.16	4.62	1.66
LODE14B	2,317	0.005	547.96	5.75	0.43	20.45	3.55
LODE1A	43	0.018	11.35	1.23	0.27	1.77	1.44
LODE1B	97	0.005	7.05	1.15	0.13	1.31	1.14
LODE1C	223	0.01	15.48	1.41	0.12	1.83	1.29
LODE3	684	0.005	67.89	2.39	0.17	4.49	1.88
LODE4	2,909	0.005	327.53	1.95	0.16	8.67	4.45
LODE5	1,010	0.005	53.80	2.91	0.18	5.75	1.98
LODE6	538	0.005	16.66	1.04	0.08	1.80	1.73
LODE9A	3,182	0.005	74.20	1.28	0.05	2.70	2.11
LODE9B	171	0.041	7.71	0.76	0.07	0.86	1.13
MLODE2A	1,820	0.005	37.86	0.86	0.04	1.89	2.20
MLODE2B	627	0.005	28.64	1.09	0.09	2.18	2.00
MLODE2C	145	0.005	13.14	1.16	0.17	2.01	1.73

SD = Standard Deviation; CV = Coefficient of Variation

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14.4.3.2 High grade outlier analysis

Three metre composite samples have had a top cut applied for each lode based on the outlier shown in logarithmic histogram and log probability plots. Grades were cut to within the 98th to 100th percentile and in the case of lodes with less than 100 samples there was no top cut applied however the range of estimation for the outliers was reduced to 20 m.

Table 14.3 shows the capping by domains.

Table 14.13 Tap AB Underground Resource; 3 m composite top cut capping by domain

Domain	Domain is split into separate Oxide and Sulphide Domains	Au capping (g/t)
lode10	Yes	9
lode11a	No	9
lode11b	No	9
lode6	Yes	9
lode12a	Yes	7
lode12b	No	No
lode12c	No	5
lode12d	Yes	15
lode14a	No	20
lode14b	Yes	80
lode1a	No	3
lode1b	Yes	No
lode3	Yes	20
lode5	Yes	25
lode4	Yes	18
lode9a	Yes	10
lode9b	Yes	3
lode1c	Yes	7
minor_lodes_v2a	Yes	8
minor_lodes_v2b	Yes	8
minor_lodes_v2c	Yes	8

14.4.4 Urucum North Underground model

14.4.4.1 Descriptive statistics and top cuts

The statistics analysis was carried out using raw data, 1 m composites, and also 1 m composites with selected top cuts applied by domains. The results of this statistical study demonstrate high variability with characteristics common in gold deposits with CVs often high due to dilution caused by internal waste within the domains. Top cuts for the 1 m composites were selected following a review of outliers in the tail of the log histogram and log probability plots. Table 14.14 below shows the raw, 1 m composite and top cut 1 m composite statistics. Values in the Maximum column indicate the top cut values where the grades have been rounded.

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Table 14.14 Urucum North Underground Resource statistics; raw, composited, and capped

SHT_type	Assay					Composited					Capped
	Count	Mean	SD	Median	Maximum	Count	Mean	SD	Median	Maximum	Count	Mean	SD	Median	Maximum
Lode 1
LODE0001	283	1.264	2.495	0.570	30.320	277	1.245	2.135	0.611	24.188	277	1.185	1.664	0.611	10.000
LODE1001	311	3.769	6.649	1.565	50.040	308	3.742	6.353	1.650	49.617	308	3.564	5.308	1.650	30.000
LODE1002	209	6.822	14.046	2.120	121.070	200	7.118	13.599	2.279	87.975	200	6.312	9.985	2.279	40.000
LODE1003	79	1.271	2.363	0.850	20.420	79	1.264	2.342	0.888	20.418	79	1.069	0.961	0.888	5.000
LODE1004	67	5.530	15.360	1.270	99.150	66	5.610	13.800	1.520	78.090	66	3.677	5.150	1.519	20.000
Lode 2
LODE0002	276	1.197	2.227	0.560	23.800	269	1.163	1.986	0.578	20.236	269	1.097	1.474	0.578	10.000
LODE2001	294	3.294	5.272	1.650	54.280	277	3.131	4.100	1.830	31.718	277	3.050	3.631	1.830	20.000
LODE2002	73	3.337	3.916	1.810	16.400	71	3.413	3.675	1.940	16.337	71	3.187	3.073	1.940	10.000
LODE2003	64	3.813	5.112	1.605	28.170	62	3.925	5.194	1.656	28.136	62	3.475	3.675	1.656	12.000
LODE2005	16	4.680	7.590	1.440	29.970	16	4.740	6.460	1.240	21.460	16	3.514	3.649	1.243	10.000
Lode 300
LODE0300	473	1.128	2.387	0.600	38.850	470	1.113	2.304	0.612	37.015	470	1.071	1.754	0.612	20.000
LODE3001	116	2.127	2.781	1.335	16.940	108	2.047	2.459	1.372	16.782	108	1.892	1.759	1.372	8.000
LODE3002	34	3.164	4.099	1.685	18.210	34	3.166	4.088	1.721	18.207	34	2.620	2.383	1.721	8.000
LODE3003	15	4.570	6.130	2.270	24.790	14	4.254	2.954	4.550	10.446	14	4.254	2.954	4.550	10.446
LODE3004	13	1.364	1.714	1.030	6.400	13	1.443	1.668	0.685	6.398	13	1.182	0.929	0.685	3.000
LODE3005	30	6.000	8.750	2.170	37.310	30	6.000	7.790	2.210	27.800	30	5.420	6.320	2.210	20.000
Lode 400
LODE0400	171	0.647	0.827	0.400	5.230	168	0.646	0.796	0.429	5.153	168	0.646	0.796	0.429	5.153
LODE4001	12	7.660	18.900	2.000	67.420	12	7.690	18.620	2.090	66.570	12	2.807	2.322	2.089	8.000
LODE4002	65	2.184	2.486	1.580	12.820	65	2.137	2.346	1.580	12.559	65	2.137	2.346	1.580	12.559
LODE4005	5	3.870	3.180	3.230	8.550	5	3.860	3.180	3.230	8.550	5	3.860	3.180	3.230	8.550
LODE0100	231	1.645	7.213	0.190	78.890	234	1.630	6.900	0.210	77.682	234	1.151	3.255	0.210	20.000
LODE0299	263	0.768	1.924	0.310	26.300	264	0.747	1.862	0.315	26.265	264	0.682	1.116	0.315	9.000
LODE0301	294	0.624	1.038	0.310	12.370	290	0.617	1.015	0.320	12.370	290	0.582	0.735	0.320	4.000
LODE0500	28	3.350	3.191	1.965	12.190	28	3.350	2.845	2.223	11.911	28	3.350	2.845	2.223	11.911
CO	79	0.689	0.698	0.550	5.700	80	0.689	0.683	0.556	5.699	80	0.689	0.683	0.556	5.699

SD = Standard Deviation; CV = Coefficient of Variation.

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14.4.5 Urucum South Underground model

14.4.5.1 Descriptive statistics

Raw statistics were reviewed for each domain and split by oxide and fresh where sufficient data existed to do so. Compositing was selected on 3 m intervals to be consistent with the Urucum MIK Open Pit resource estimate. A review of 2 m vs 3 m composites was undertaken grouped by variogram sector and showed no significant bias in the 3 m composites.

Composite statistics were generated for the 78 domains. Table 14.15 below shows 3 m composite statistics for domains with counts over 100. The results of this statistical study have shown high variability, some characteristics common in gold deposits with CVs often high due to the dilution caused by internal waste within the lodes. As a consequence of the high variability shown in the gold distribution, a large part of the gold metal is contained in the tail of the distribution and required analysis and treatment of the outliers.

Table 14.15 Urucum South Underground Resource statistics; Au (g/t)

LodeType	Count	Minimum	Maximum	Mean	SE mean	SD	CV
LOCN702	3,920	0.003	51.50	1.76	0.04	2.62	1.49
LOCN700	3,831	0.003	24.19	1.16	0.02	1.39	1.20
LOUN510	2,978	0.011	36.90	1.10	0.02	1.29	1.18
LOCN704	1,686	0.005	56.07	1.48	0.07	3.06	2.07
LOCN703OX	1,518	0.002	34.90	1.03	0.04	1.69	1.64
LOCN703FR	1,361	0.005	69.63	1.60	0.08	2.97	1.85
LOLN001	1,198	0.005	48.10	1.78	0.11	3.71	2.08
LOLN300	703	0.005	19.76	1.14	0.07	1.76	1.55
LOUN540	606	0.038	5.55	0.90	0.02	0.60	0.66
LOLN002	553	0.005	19.20	1.71	0.10	2.33	1.36
LOUS902	533	0.005	14.35	2.05	0.11	2.42	1.18
LOUN602	515	0.021	14.70	1.16	0.06	1.43	1.24
LOLN299	385	0.005	44.49	0.78	0.13	2.63	3.40
LOUN520OX	363	0.013	6.14	0.66	0.03	0.60	0.90
LOLN301	341	0.005	19.00	0.78	0.07	1.32	1.69
LOUS900OX	307	0.004	10.63	0.85	0.07	1.22	1.43
LOLN400	295	0.005	34.01	1.00	0.13	2.28	2.27
LOUN601	260	0.001	22.49	1.20	0.11	1.71	1.42
LOUS903	260	0.005	20.92	1.11	0.13	2.01	1.82
LOUN530	210	0.01	3.39	0.62	0.03	0.45	0.72
LOUN606	184	0.07	5.72	1.21	0.07	0.90	0.74
LOUN603	179	0.02	9.60	1.13	0.08	1.05	0.93
LOUS901OX	175	0.005	3.96	0.70	0.05	0.62	0.89
LOCS800	166	0.005	17.03	0.88	0.12	1.53	1.73
LOCN709	153	0.011	16.07	1.20	0.14	1.68	1.40
LOCN705FR	147	0.04	5.68	0.92	0.09	1.04	1.13
LOUS900FR	125	0.005	4.31	0.50	0.06	0.66	1.31
LOUN626	114	0.01	2.13	0.56	0.03	0.37	0.66
LOUN520FR	112	0.036	3.34	0.76	0.06	0.63	0.83
LOCN707OX	107	0.02	9.97	0.84	0.14	1.42	1.70

Notes: Table shows 3 m composite statistical summary high grade outlier analysis. Only domains with > 100 counts shown.

SD = Standard Deviation; CV = Coefficient of Variation

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Three metre composite samples have had a top cut applied to each lode based on the outliers shown in the logarithmic histograms and probabilities plots. Top cuts were within the 98th to 100th percentile. In case of lodes with less than 100 samples, no top cuts were applied to the outliers however the estimation search for these samples was reduced to 20 m. Table 14.16 shows the top cuts applied to the 3 m composites by lode.

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Table 14.16 Urucum South Underground Resource statistics; raw gold data (g/t)

Lode No	Lode type	No. 3 m composite	Top cut-off	High grade value - search ellipsoid reduction	Lode No	Lode type	No. 3 m composite	Top cut-off	High grade value - search ellipsoid reduction	Lode No	Lode type	No. 3 m composite	Top cut-off	High grade value - search ellipsoid reduction
800	LOCS800	166	4.8		722	LOCN722	18	9,999		617	LOUN617	13	1.25
801	LOCS801	70	5.2		001	LOLN001	1,198	18		618	LOUN618	20	9,999
802	LOCS802	59	9,999	2.7	002	LOLN002	553	12		620	LOUN620	6	9,999
803	LOCS803	16	9,999	1.7	299	LOLN299	385	5		621	LOUN621	2	9,999
804	LOCS804	7	9,999		300	LOLN300	703	9		623	LOUN623	11	9,999	2
902	LOUS902	533	9,999		301	LOLN301	341	6		624	LOUN624	3	9,999
903	LOUS903	260	6		400	LOLN400	295	5		625	LOUN625	7	9,999
904	LOUS904	31	9,999		500	LOLN500	10	9,999		626	LOUN626	114	2
905	LOUS905	44	9,999		510	LOUN510	2,978	6		627	LOUN627	25	9,999	10
700	LOCN700	3,831	6.3		530	LOUN530	210	2.5		628	LOUN628	9	9,999
701	LOCN701	4	9,999		540	LOUN540	606	3		900	LOUS900OX	307	5.5
702	LOCN702	3,920	11		550	LOUN550	92	9,999	3	900	LOUS900FR	125	5.5
704	LOCN704	1,686	13		600	LOUN600	38	9,999	5.3	901	LOUS901OX	175	2.56
706	LOCN706	57	1.5		601	LOUN601	260	6.3		901	LOUS901FR	92	2.56
708	LOCN708	35	9,999		602	LOUN602	515	7.3		703	LOCN703OX	1,518	11
709	LOCN709	153	9		603	LOUN603	179	3.6		703	LOCN703FR	1,361	11
710	LOCN710	29	9,999		605	LOUN605	61	9,999	3	705	LOCN705OX	57	7
711	LOCN711	65	9,999	1.6	606	LOUN606	184	4.3		705	LOCN705FR	147	7
712	LOCN712	14	9,999		607	LOUN607	7	9,999		707	LOCN707OX	107	9,999
713	LOCN713	47	9,999	1.56	608	LOUN608	89	10		707	LOCN707FR	56	9,999
715	LOCN715	40	9,999	2.51	609	LOUN609	30	9,999		721	LOCN721OX	75	9,999
716	LOCN716	31	9,999	4.5	610	LOUN610	23	9,999	1.8	721	LOCN721FR	38	9,999
717	LOCN717	45	9,999	2.62	612	LOUN612	2	9,999		520	LOUN520OX	363	3
718	LOCN718	3	9,999		614	LOUN614	59	9,999	2.5	520	LOUN520FR	112	3
719	LOCN719	8	9,999		615	LOUN615	53	9,999	5	613	LOUN613OX	38	20
720	LOCN720	31	0.95		616	LOUN616	12	9,999		613	LOUN613FR	65	20

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14.4.6 Urucum East model

14.4.6.1 Urucum East summary statistics

Drillholes within the database were assigned a unique 'zonecode' based on the numbering system. The unique database coding was used to control the compositing process. A downhole composite length of 1 m was used for all zones with a minimum composite length of 0.75 m set at the domain boundaries.

As expected, statistical analysis confirms that the HG Lode Zone is preferentially mineralized having a raw mean gold grade of 1.77 g/t and exhibiting moderate variability with a CV of around 1.29. The 99th percentile was 12.1 g/t Au and the maximum was 14.9 g/t Au. Based on the lack of any extreme outliers in the Lode or the colluvium, a top cut was not deemed necessary. Table 14.7 summarizes statistics for the high-grade zone and colluvium at Urucum East.

Table 14.17 Urucum East 1 m composites by domain - summary statistics

	High grade zone	Colluvium
	Raw Au ppm	Raw Au ppm
Number	284	612
Minimum	0.005	0.005
Maximum	14.9	1.0
Mean	1.77	0.023
SD	2.27	0.07
CV	1.29	3.20

SD = Standard Deviation; CV = Coefficient of Variation

14.4.7 Tap C (Other) model

14.4.7.1 Raw sample data

Statistics were calculated for raw data separated by azimuth / dip groups, in each group the information was consolidated into sub-domains for variography. The main objective of analysing the statistics of the raw data was to understand the behaviour of gold grade for domains by analysing mean, standard deviation (Variance) to support the variography study.

The results of this shown high variability and some characteristics common in gold deposits with CVs often high due to the inclusion of internal waste dilution. As a consequence of the high variability shown in the gold distribution, a large part of the gold is situated at the tail of the distribution and in some cases will require that a top cut be applied.

Table 14.18 Raw data statistic summary - grouped ore domains gold (ppm)

Au (PPM)	LODE111	LOGR020	LODE103
Count	3,406	1,271	161
Minimum	0.01	0	0.03
Maximum	24.6	30.2	7.92
Mean	1.119	0.721	0.735
SE of mean	0.031	0.042	0.066
Standard deviation	1.79	1.5	0.84
CV	1.602	2.081	1.143

SD = Standard Deviation; CV = Coefficient of Variation

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14.4.7.2 2 m composite data

As with the raw data, the 2 m composites from Tap C were analysed using the same approach. In addition to this, the composites were used to undertake top cut analysis.

The 2 m composites were separated by azimuth / dip groups for geostatistical analysis. These groups were further consolidated in sub-domains for variography. The VERT sector is shown in the following tables and figures.

Table 14.19 Tap C and gap, 2 m composite statistical summary of variogram sector VERT

Au (PPM)	LODE3_4_5101	LODE102	LODE111	LOGR020	LODE103
Count	22,752	2,878	1,821	796	86
Minimum	0.00	0.00	0.01	0.01	0.07
Maximum	90.59	28.05	18.17	17.42	4.06
Mean	1.003	1.594	1.095	0.715	0.710
SD	2.44	2.46	1.61	1.22	0.62
CV	2.433	1.540	1.475	1.706	0.879

SD = Standard Deviation; CV = Coefficient of Variation

14.4.7.3      Treatment of high grade gold values (g/t)

The two metre composite samples have had a top grade for each lode based on the outlier shown in their respective logarithmic histograms and probability plots. The top cut grades are within the 98th to 100th percentile of the grade distribution. In some cases where there was a low quantity of samples present, the search ellipsoid was reduced to 20 m.

Table 14.20 details the high-grade Au treatment for each estimated ore lode.

Table 14.20 Tap C (other) model; top grade for 2 m composite - Au (g/t)

Lode code		Dip group	Variography group	Top cut-off		Red ellipsoid
				Au (g/t)	%	Au (g/t)
LODE3_4_5101	3101	VERT	VERT_101_102	10.0	98.94%	9,999.0
LODE102	102	VERT	VERT_101_102	12.0	98.71%	9,999.0
LODE111	111	VERT	VERT	8.0	98.68%	9,999.0
LOGR020	20	VERT	VERT	6.0	98.99%	9,999.0
LODE103	103	80E	VERT	9,999.0	100.00%	9,999.0
LODE201	201	VERT	VERT	15.0	97.90%	9,999.0
LODE104	104	40E	40E	9,999.0	100.00%	9,999.0
LODE105	105	40E	40E	9,999.0	100.00%	9,999.0
LODE108	108	40E	40E	9,999.0	100.00%	9,999.0
LODE107	107	40E	40E_106_107	5.0	98.67%	9,999.0
LODE106	106	60E	40E_106_107	10.0	98.68%	9,999.0
LODE109	109	N20/60E	N20/60E	9,999.0	100.00%	2.0
LODE110	110	N20/60E	N20/60E	5.0	98.96%	9,999.0
LGEV3_4_5010	3,010	60W	60W	1.5	99.01%	9,999.0
LODE2101	2,101	OMNI	LODE2101	10.0	98.84%	9,999.0
LGEV2010	2,010	OMNI	LGEV2010	4.0	98.90%	9,999.0

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14.4.8 Tap C (C3W and C3N) model

A statistical analysis based on 2 m composites for gold was undertaken within each mineralized domain. Table 14.21 shows summary statistics of gold for the main mineralized domains at Tap C3W and Tap C3N.

Top cuts have been applied to the 97.5% percentile of each estimation group to exclude outliers from the estimation (Table 14.21).

Table 14.21 Tap C3W and C3N 2 m composites by domain - summary statistics

Zone	1	2	7
Number	2162	461	154
Minimum	0.00	0.00	0.06
Maximum	27.73	4.06	5.03
Mean	1.06	0.81	0.87
SD	1.43	0.56	0.62
CV	1.35	0.69	0.71

SD = Standard Deviation; CV = Coefficient of Variation

Table 14.22 Tap C3W and C3N 2 m applied top cuts by domain

Zone	Top-cut Au (g/t)
1	4.72
2	2.29
7	2.51

14.4.9 Duckhead model

14.4.9.1 Raw sample data

Statistics were calculated for raw data and reviewed in their respective azimuth / dip groups (groupings determined for variography). A statistical summary of the domains is shown in Table 14.23.

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Table 14.23 Duckhead; raw data Au (g/t), statistical summary by domain

Domain	Count	Minimum	Maximum	Mean	SD	CV
LODE1	313	0	6844.6	165.3	441.9	2.68
LODE100	12	0.04	2.87	0.655	0.836	1.28
LODE101	54	0.08	4.09	0.869	0.828	0.95
LODE102	67	0.07	95.37	4.4	12.91	2.94
LODE103	9	0.4	2.06	0.857	0.535	0.62
LODE104	2	0.44	0.53	0.485	0.064	0.13
LODE105	6	0.42	0.7	0.535	0.095	0.18
LODE106	110	0.005	6.32	0.708	0.929	1.31
LODE107	232	0.005	95.34	2.225	8.159	3.67
LODE108	3	0.66	4.62	2.58	1.98	0.77
LODE109	35	0.15	13.47	1.335	2.685	2.01
LODE110FR	41	0.005	21.91	1.15	3.471	3.02
LODE110OX	37	0.02	13.64	1.193	2.278	1.91
LODE2	2,781	0.01	14018.6	11.4	267.3	23.40
LODE3	2,145	0.005	42.24	0.556	1.175	2.11
LODE31	857	0.005	36.94	0.639	1.806	2.83
LODE4FR	231	0.005	30.56	2.986	4.118	1.38
LODE4OX	1,991	0.005	93.09	5.17	8.056	1.56
LODE51FR	88	0.005	20.28	1.808	3.254	1.80
LODE51OX	92	0.02	14.34	1.264	1.938	1.53
LODE5FR	313	0.005	18.77	0.626	1.511	2.41
LODE5OX	1,493	0.005	17.62	0.789	1.359	1.72
LODE6FR	15	0.22	10.03	4.108	2.264	0.55
LODE6OX	22	0.3	31.86	6.33	8.45	1.33
LODE71	706	0.005	47.44	1.127	2.652	2.35
LODE7FR	123	0.005	9.68	0.832	1.257	1.51
LODE7OX	866	0.005	32.5	0.611	1.289	2.11
LODE8	594	0.005	53.63	0.811	2.441	3.01
LODE9	2789	0.005	24.81	0.73	1.163	1.60
LODE90	123	0.005	12.94	0.766	1.638	2.14

SD = Standard Deviation; CV = Coefficient of Variation

14.4.9.2 2 m composite data

Samples regularized on 2 m composite and then reviewed for appropriate top cuts for the grade outliers. Statistics for the 2 m composites by domain are shown in Table 14.24.

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Table 14.24 Duckhead; 2 m composite statistical summary - grouped by ore lode

LODETYPE	Count	Minimum	Maximum	Mean	SD	CV
LODE1	199	0	3,557.5	166.2	300.4	1.81
LODE100	6	0.042	2.87	0.889	1.036	1.17
LODE101	30	0.244	2.385	0.869	0.561	0.65
LODE102	38	0.08	53.78	3.96	9.67	2.44
LODE103	6	0.314	1.6	0.794	0.478	0.60
LODE104	2	0.44	0.5297	0.4848	0.0634	0.13
LODE105	4	0.48	0.6999	0.555	0.0988	0.18
LODE106	56	0.01	3.951	0.701	0.755	1.08
LODE107	128	0.005	72.588	2.105	6.834	3.25
LODE108	2	2.437	2.994	2.716	0.394	0.15
LODE109	21	0.253	13.436	1.536	2.923	1.90
LODE110	46	0.007	11.434	1.147	2.074	1.81
LODE2	1,505	0.01	13,849.6	15.5	357.4	2.30
LODE3	1,129	0.005	21.1534	0.5527	0.829	1.50
LODE31	460	0.005	18.926	0.6329	1.3549	2.14
LODE4	1,213	0.008	68.86	4.894	6.789	1.39
LODE5	978	0.005	13.94	0.7875	1.2491	1.59
LODE51	94	0.025	14.31	1.612	2.294	1.42
LODE6	19	1.63	31.86	6.16	7.31	1.19
LODE7	519	0.005	16.4353	0.6389	0.9504	1.49
LODE71	376	0.005	35.078	1.099	2.22	2.02
LODE8	306	0.005	23.2124	0.8143	1.6019	1.97
LODE9	1,425	0.005	14.6913	0.7269	0.9674	1.33
LODE90	73	0.044	7.05	0.721	1.215	1.68

SD = Standard Deviation; CV = Coefficient of Variation

14.4.9.3 Top cuts

The 2 m composite samples have had a top cut applied to each domain based on an analysis of outliers in logarithmic histogram and probability plots. The top cut grades are within 98th to 100th percentile of the grade distribution. In case of domains with less than 100 composites, no grade capping was applied however the range of the estimation was restricted to 20 m.

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Table 14.25 Duckhead determined top cuts for each lode

Domain codes		Au capping (g/t)	Lode codes		Au capping (g/t)
LODE1	1	900	LODE2	2	45
LODE100	100	No	LODE3	3	3
LODE101	101	No	LODE31	31	4
LODE102	102	4.5	LODE4	4	33
LODE103	103	No	LODE5	5	6.5
LODE104	104	No	LODE51	51	8
LODE105	105	No	LODE6	6	No
LODE106	106	No	LODE7	7	4
LODE107	107	13.5	LODE71	71	6.5
LODE108	108	No	LODE8	8	6
LODE109	109	No	LODE9	9	8
LODE110	110	No	LODE90	90	No

14.5 Variography

14.5.1 Introduction

Variography is used to describe the spatial variability or correlation of an attribute (gold, silver, etc.). The spatial variability is traditionally measured by means of a variogram, which is generated by determining the averaged squared difference of data points at a nominated distance (h), or lag (Srivastava and Isaacs, 1989). The averaged squared difference (variogram or γ (h)) for each lag distance is plotted on a bivariate plot, where the X-axis is the lag distance and the Y-axis represents the average squared differences (γ(h)) for the nominated lag distance.

Several types of variogram calculations are employed to determine the directions of the continuity of the mineralization:

• Traditional variograms are calculated from the raw assay values.

• Log-transformed variography involves a logarithmic transformation of the assay data.

• Gaussian variograms are based on the results after declustering and a transformation to a Normal distribution.

• Pair wise-relative variograms attempt to ‘normalize’ the variogram by dividing the variogram value for each pair by their squared mean value.

• Correlograms are ‘standardized’ by the variance calculated from the sample values that contribute to each lag.

Fan variography involves the graphical representation of spatial trends by calculating a range of variograms in a selected plane and contouring the variogram values. The result is a contour map of the grade continuity within the domain.

14.5.2 Tap AB Open Pit model variography

Grade and indicator variography was generated to enable grade estimation via MIK and change of support analysis to be completed. In addition, Gaussian variograms were also examined as part of the change of support process. Indicator thresholds for Domains to be estimated via MIK had variograms modelled with every third variogram typically modelled. Variograms not modelled have had their parameters interpolated based on the bounding modelled variograms. Interpreted anisotropy directions correspond well with the modelled geology and overall geometry of the interpreted domains. Grade variography was modelled for OK grade estimation.

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14.5.3 Urucum Open Pit model variography

The Urucum open pit model variography was done using the same procedures outline above for the Tap open pit model.

14.5.4 Tap AB Underground model variography

The analysis of spatial variability of the TAP-AB Deposit was performed by analysis of experimental and fitting pair wise semi-variograms using 3 m composites combined into 8 variography groups.

14.5.5 Urucum North Underground model variography

The analysis of spatial variability of the Urucum Underground Project was carried out by analysis of experimental and modelled semi-variograms. 1 m lode composites were top cut and grouped into a single domain for variogram analysis. Despite the high variability of the Au grades, the raw semi-variograms were made using the traditional variogram function. A variogram was modelled down-hole to obtain the nugget effect.

14.5.6 Urucum South Underground model variography

The analysis of spatial variability for the Urucum South Underground model was carried out by analysis of experimental and fitted semi-variograms. 3 m composites were used with a top cut applied and the data grouped within variography groups as defined by differences in lode morphology, azimuth, and dip.

14.5.7 Urucum East model variography

Variography has been used to analyse the spatial continuity within the mineralized zones and to determine appropriate estimation inputs to the interpolation process. The variogram modelling process generally followed these steps:

• Calculate and model the omni-directional or down hole variogram to characterize the Nugget effect.

• Systematically calculate orientated variograms in 3 dimensions to identify the plane of greatest continuity.

• Calculate a fan of variograms within the plane of greatest continuity to identify the direction of maximum continuity within the plane. Model the variogram in the direction of maximum continuity and the orthogonal directions.

Experimental normal variography with lags of 25 m was performed on 1 m downhole composites. Directional variography for gold within the HG zone revealed generally well-structured behaviour with maximum ranges of 76 m (major range) and 52 m (semi-major range) within the plane of mineralization. The mineralized plane has a gentle plunge towards the North (-30° to > 0°). An omni-directional variogram model with maximum ranges of 125 m in the horizontal plane was modelled for the colluvium domain.

14.5.8 Tap C (other) model variography

The analysis of spatial variability for the Tap C model was carried out by analysis of experimental and fitted semi-variogram models. 2 m composites with an applied top cut grade were used and the data within the domains grouped into their variography groups as defined by statistical analysis. This procedure of grouping was applied to also improve the variography of discontinuous lodes with a small number of samples.

The criteria used to select the groups were based on statistical analysis and cumulative probability plots that showed no significant differences between the grades inside the domains. Domain groups however had key differences in morphology, azimuth, and dip, and this was reviewed with site geologists.

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Due to the high variability of the Au grades, the raw variograms were generally impossible to interpret.

The variogram methodology is pair wise relative semi-variogram. Journell 2008 - GSLIB Second Edition “With this variogram, each pair is normalized by the squared average of the tail and head values.

 

Practical experience has shown that the general relative and pair wise relative sample semi-variograms are resistant to data sparsity and outliers when applied to positively skewed sample distribution”. It should be made clear that the transformation of the original grades into Gaussian distributed values (Gaussian anamorphosis) before calculating the variograms was a solution considered, however, it would have generated effectively the same results.

In general, the variograms show a relatively low nugget effect but the ranges of the first geostatistical structure are short and often over 80% of the total variability is reached within 20 m to 40 m.

Fitting the semi-variogram was achieved by first fitting the omnidirectional semi-variogram with a short lag to get nugget effect and then fitting the directional semi-variogram.

14.5.9 Tap C (C3N and C3W) model variography

Variography has been used to analyse the spatial continuity within the mineralized zones and to determine appropriate estimation inputs to the interpolation process. The variogram modelling process generally followed these steps:

• Review the omni-directional or down-hole variogram to characterize the Nugget Effect. Due to the poor resultant downhole variogram, the nugget was modelled from the orientated variograms.

• Systematically calculate orientated variograms in 3 dimensions to identify the plane of greatest continuity.

• Calculate a fan of variograms within the plane of greatest continuity to identify the direction of maximum continuity within the plane. Model the variogram in the direction of maximum continuity and the orthogonal directions.

Experimental pair wise variography with lags of 10 m was performed on 2 m downhole composites.

Directional variography for gold within the 3 zones defining the C3W and C3N areas show a high nugget and maximum ranges up to 161 m with the first structure range between 16 m and 46 m within the plane of the dipping domains. A lag spacing of 10 m was selected to approximate the drillhole spacing within these shallow domains.

14.5.10 Duckhead model variography

The 2 m composites grouped into 4 variogram sectors based on domain orientation. Lode 1 (the core of Duckhead Main mineralization) has been separated out with its own variogram sector.

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14.6 Block modelling

14.6.1 Tap AB and Urucum Open Pit block models

For the Tap AB and Urucum Open Pit Resources 3D block models were created in the National grid (UTM SAD69 Zone 22N) using Vulcan mining software. The parent block size was selected on the basis of the average drill spacing together with consideration of mining parameters currently in use at the Tucano Gold Project. A parent cell size of 9 mE by 20 mN by 8 mRL which was sub-blocked down to 3 mE by 5 mN by 4 mRL (to ensure adequate volume representation). The models covered all the interpreted mineralization zones and included suitable additional waste material to allow later pit optimization studies. Block coding was completed on the basis of the block centroid, wherein a centroid falling within any wireframe was coded with the wireframe solid attribute. The block models are non-rotated.

The main block model parameters are summarized below in Table 14.26 and Table 14.27. Variables were coded into the block models to enable MIK and OK estimation and subsequent MIK change of support and grade tonnage reporting. A visual review of the wireframe solids and the block model indicated correct flagging of the block model. Additionally, a check was made of coded volume versus wireframe volume which confirmed the above.

Table 14.26 Tap AB Gold Deposit - Open Pit Resource block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. coordinates	401,640	92,740	-627
Max coordinates	402,855	95,640	325
Block size (m)	9.0	20.0	8.0
Sub block size (m)	3.0	5.0	4.0
Rotation (° around axis)	0°	0°	0°

Table 14.27 Urucum Gold Deposit - Open Pit Resource block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. coordinates	401,500	97,940	-683
Max coordinates	402,319	100,440	317
Block size (m)	9.0	20.0	8.0
Sub block size (m)	3.0	5.0	4.0
Rotation (° around axis)	0°	0°	0°

14.6.2 Tap AB and Urucum Underground block models

For the Tap AB and Urucum Underground Resources 3D block models were created in the National grid (UTM SAD69 Zone 22N) using Surpac mining software. Estimation was completed in Isatis using parent cells and then script imported into an attributed and sub blocked Surpac model on a domain by domain basis such that the domains were completely filled to sub cell resolution with the estimated blocks. The block models for the Urucum North Underground and Urucum South Underground both extend into one another however the resource reporting for these uses the 99,400 mN northing as the official divisor between the 2 models.

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Table 14.28 Tap AB Underground block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. Coordinates	401,640	92,740	-627
Max Coordinates	402,855	95,640	325
Block size (m)	9.0	20.0	8.0
Sub Block size (m)	1.125	5	2.0
Rotation (° around axis)	0°	0°	0°

Table 14.29 Urucum North Underground block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. coordinates	401,500	97,940	-683
Max coordinates	402,316	100,440	317
Block size (m)	8.0	20.0	20.0
Sub block size (m)	1.0	5.0	2.5
Rotation (° around axis)	0°	0°	0°

Table 14.30 Urucum South Underground block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. coordinates	401,500	97,940	-683
Max coordinates	402,319	100,440	317
Block size (m)	9.0	20.0	8.0
Sub block size (m)	1.125	5.0	2.0
Rotation (° around axis)	0°	0°	0°

14.6.3 Urucum East, Tap C, and Duckhead block models

The Urucum East, Tap C and Duckhead block models were created in the National grid (UTM SAD69 Zone 22N) using Surpac software. For Duckhead and the Tap C Other Areas, Isatis estimated centroids were imported using the same methodology as the Tap AB Underground and Urucum Underground models. Urucum East and the Tap C (C3N and C3W areas) were estimated directly inside Surpac.

Table 14.31 Urucum East Gold Deposit block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. coordinates	402,600	100,000	0
Max coordinates	403,250	100,490	320
Block size (m)	5.0	10.0	2.0
Sub block size (m)	0.625	2.5	0.5
Rotation (° around axis)	0°	0°	0°

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Table 14.32 Tap C Gold Deposit block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. coordinates	401,600	95,640	-355
Max coordinates	402,704	97,940	285
Block size (m)	8.0	20.0	4.0
Sub block size (m)	2.0	5.0	2.0
Rotation (° around axis)	0°	0°	0°

Table 14.33 Duckhead Gold Deposit block model parameters

	Easting (X)	Northing (Y)	RL (Z)
Min. coordinates	407,750	87,750	-250
Max coordinates	409,750	90,250	300
Block size (m)	5.0	10.0	2.0
Sub block size (m)	0.625	2.5	1.0
Rotation (° around axis)	0°	-45°	0°

14.7 Grade estimation

14.7.1 Tap AB and Urucum Open Pit block models

14.7.1.1  Introduction

MIK was applied to grade estimation at the Tap AB and Urucum Gold Deposits within the defined indicator mineralization shells. A number of minor domains were estimated via OK. Estimation was completed in the mining package Vulcan using the GSLib geostatistical software while geostatistical change of support parameters was developed in Isatis geostatistical software. MIK is considered a robust estimation methodology for grade estimates for gold deposits such as Tap AB and Urucum where high levels of short scale variability are present. MIK grade estimation with change of support has been applied to produce ‘recoverable’ gold estimates targeting a selective mining unit (SMU) of 3 mE x 5 mN x 4 mRL.

14.7.1.2 The MIK method

The MIK technique is implemented by completing a series of OK estimates of binary transformed data. A composite sample, which is equal to or above a nominated cut-off or threshold, is assigned a value of 1, with those below the nominated indicator threshold being assigned a value of 0. The indicator estimates, with a range between 0 and 1, represent the probability the point will exceed the indicator COG. The probability of the points exceeding a cut-off can also be considered broadly equivalent to the proportion of a nominated block that will exceed the nominated COG.

The estimation of a complete series of indicator cut-offs allows the reconstitution of the local histogram or conditional cumulative distribution function (ccdf) for the estimated point. Based on the ccdf, local or block properties, such as the block mean and proportion (tonnes) above or below a nominated cut-off grade can be investigated.

14.7.1.3 Post MIK processing - E-type estimates

The E-type estimate provides an estimate for the grade of the total block or bulk-mining scenario. This is achieved by discretizing the calculated ccdf for each block into a nominated number of intervals and interpolating between the given points with a selected function (e.g. the linear, power or hyperbolic model) or, by applying intra-class mean grades. The sum of all these weighted interpolated points or mean grades enables an average whole block grade to be determined.

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The following example shows the determination of an E-type estimate for a block containing three indicator cut-offs.

The indicator cut-offs and associated probabilities calculated are shown in Table 14.34.

Table 14.34 Tap AB and Urucum open pit block model Indicator cut-off and probability

Indicator	COG Au (g/t)	Indicator probability (cumulative)
Minimum grade *	0	0.00 **
Indicator 1	1	0.40
Indicator 2	2	0.65
Indicator 3	3	0.85
Maximum grade *	4	1.00 **

Notes:

* COGs determined by the user.

** Indicator probability is assumed at the minimum and maximum cut-off.

The whole block grade can now be determined in this block with the following parameters used for the purposes of the interpolation:

• Number of discretization intervals: 4.

• Linear extrapolation between all points (median grade between nominated cut-offs).

The worked example is then calculated with the following steps:

• Interval 1 (0-1 g/t Au)      median grade x probability / proportion attributed to the interval (0.5 g/t Au x 0.40 = 0.200).

• Interval 2 (1 - 2 g/t Au)     median grade x proportion (1.5 g/t Au x 0.25 = 0.375).

• Interval 3 (2 - 3 g/t Au)     median grade x proportion (2.5 g/t Au x 0.20 = 0.500).

• Interval 4 (3 - 4 g/t Au)     median grade x proportion (3.5 g/t Au x 0.15 = 0.525).

• Calculate total grade         average all calculated intervals ((0.2 + 0.375 + 0.500 + 0.525) / 1) = 1.60 g/t Au.

It is also possible from this example to calculate the proportion and grade above a nominated cut-off (e.g. 2 g/t - at sample support or complete selectivity). The following steps would be undertaken to calculate the tonnes and grade at sample selectivity using a 2 g/t Au COG:

• Interval 3 (2 - 3 g/t Au)      median grade x proportion (2.5 g/t Au x 0.20 = 0.500).

• Interval 4 (3 - 4 g/t Au)      median grade x proportion (3.5 g/t Au x 0.15 = 0.525).

• Calculate total grade average all calculated intervals ((0.500 + 0.525) / 0.35) = 2.93 g/t Au with 0.35% of the block above the COG.

The effect of using a non-linear model to interpolate between COGs is to shift the grade weighting associated with that cut-off away from the median. For Tap AB and Urucum, the intra-class means based on the cut composite data have been used to reconstitute the ccdf and produce block statistics.

It is noted, however, that the calculation of the E-type estimate and complete selectivity often does not allow mine planning to the level of selectivity which is proposed for production. To achieve an estimate which reflects the levels of mining selectivity envisaged, a selective mining unit (SMU) correction is often applied to the calculated ccdf.

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14.7.1.4 Support correction (selective mining unit estimation)

A range of techniques are known to produce a support correction and therefore allow for selective mining unit emulation. The common features of the support correction are:

• Maintenance of the mean grade of the histogram (E-type mean).

• Adjustment of the histogram variance by a variance adjustment factor (the ‘f’ factor).

The variance adjustment factor, used to reduce the histogram or ccdf variance, can be calculated using the variogram model. The variance adjustment factor is often modified to account for the likely grade control approach or ‘information effect’.

In simplest terms, the variance adjustment factor takes into account the known relationship derived from the dispersion variance.

Total variance = variance of samples within blocks + variance between blocks.

The variance adjustment factor is calculated as the ratio of the variance between the blocks and the variance of the samples within the blocks, with a small ratio (e.g. 0.10) indicating a large adjustment of the ccdf variance and large ratio (e.g. 0.80) representing a small shift in the ccdf.

Two simple support corrections that are available include the Affine and Indirect Lognormal correction, which are both based on the permanence of distribution. The discrete Gaussian model is often applied to global change of support studies and has been generated on the composite dataset as a comparison. The indirect lognormal correction was applied to the MIK grade estimates.

14.7.1.5 Indirect lognormal correction

The indirect lognormal correction can be implemented by adjusting the quantiles (indicator cut-offs) of the ccdf with the variance adjustment factor so that the adjusted ccdf represents the statistical characteristics of the block volume of interest.

This is implemented with the following formula:

 

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At the completion of the quantile adjustments, grades and tonnages (probabilities are then considered a pseudo tonnage proportion of the blocks) at a nominated COG can be calculated using the methodology described above (E-type). The indirect lognormal correction, as applied to the Tucano Gold Project, is the best suited of the common adjustments applied to MIK to produce selective mining estimates for positively skewed distributions.

14.7.1.6 MIK and OK parameters

MIK estimates were completed using the indicator variogram models and a set of ancillary parameters controlling the source and selection of composite data. The sample search parameters were defined based on the variography and the data spacing, and a series of sample search tests performed in Isatis geostatistical software. A total of 17 indicator thresholds were estimated for all estimation domains.

OK estimates were completed on the minor estimation domains and the colluvium using the grade variogram models (Section 14.6), and a set of ancillary parameters controlling the source and selection of composite data. The sample search parameters were defined based on the variography and the data spacing, and a series of sample search tests performed in Isatis geostatistical software.

The sample search parameters for the MIK estimations are provided in Table 14.35 and Table 14.36.

A combination of soft and hard domain boundaries was used for the MIK portion of the estimate to reflect continuity between domains or otherwise. A three-pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations (9 mE by 20 mN by 8 mRL) were applied throughout and discretization was applied on the basis of 3 X by 3 Y by 3 RL for 27 discretization points per block.

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Table 14.35 Tap AB Gold Deposit - Open Pit Resource MIK sample search criteria

Domain	Pass	Sample search orientation (dip / dip direction˚)			Sample search distance (m)			Numbers of samples			Boundary condition
		Major	Semi- major	Minor	Major	Semi-major	Minor	Min.	Max.	Max per drillhole
101	Pass 1	0/10	-20/280	-70/100	50	50	15	32	32	5	Hard
	Pass 2	0/10	-20/280	-70/100	100	100	30	24	32	-	Soft
	Pass 3	0/10	-20/280	-70/100	300	300	50	12	32	-	Soft
200	Pass 1	-90/0	0/0	0/270	50	50	15	24	24	5	Hard
	Pass 2	-90/0	0/0	0/270	100	100	30	24	24	-	Soft
	Pass 3	-90/0	0/0	0/270	300	300	50	12	24	-	Soft
1,000	Pass 1	-80/290	0/20	-10/110`	50	50	15	32	32	5	Hard
	Pass 2	-80/290	0/20	-10/110`	100	100	30	24	32	-	Hard
	Pass 3	-80/290	0/20	-10/110`	500	500	50	12	32	-	Hard
2,000	Pass 1	0/20	-80/290	-10/121	50	50	15	32	32	5	Soft
	Pass 2	0/20	-80/290	-10/121	100	100	30	24	32	-	Soft
	Pass 3	0/20	-80/290	-10/121	500	500	50	12	32		Soft
3,000	Pass 1	-40/0	-50/180	0/90	50	50	15	32	32	5	Hard
	Pass 2	-40/0	-50/180	0/90	100	100	30	24	32	-	Soft
	Pass 3	-40/0	-50/180	0/90	300	300	50	12	32		Soft

Table 14.36 Urucum Gold Deposit - Open Pit Resource MIK sample search criteria

Domain	Pass	Sample search orientation (dip / dip direction˚)			Sample search distance (m)			Numbers of samples			Boundary condition
		Major	Semi- major	Minor	Major	Semi- major	Minor	Min.	Max.	Max per drillhole
100	Pass 1	0/350	-70/80	-20/260	50	50	15	24	32	5	Hard
	Pass 2	0/350	-70/80	-20/260	100	100	30	24	32	5	Soft
	Pass 3	0/350	-70/80	-20/260	300	300	50	12	32	-	Soft
200 north1	Pass 1	0/0	-70/90	-20/270	50	50	15	24	32	5	Hard
	Pass 2	0/0	-70/90	-20/270	100	100	30	24	32	5	Soft
	Pass 3	0/0	-70/90	-20/270	300	300	50	12	32	-	Soft
200 south2	Pass 1	0/350	-70/80	-20/260	50	50	15	24	32	5	Hard
	Pass 2	0/350	-70/80	-20/260	100	100	30	24	32	5	Soft
	Pass 3	0/350	-70/80	-20/260	300	300	50	12	32	-	Soft
300, 4003	Pass 1	0/0	-75/90	-15/270	50	50	15	24	32	5	Hard
	Pass 2	0/0	-75/90	-15/270	100	100	30	24	32	-	Hard
	Pass 3	0/0	-75/90	-15/270	300	300	50	12	32	-	Soft

Notes:

(1) for block centroids north of 98,915 mN.

(2) for block centroids south of 98,915 mN.

(3) Domains 300 and 400 estimated simultaneously.

The sample search parameters for the OK estimations are provided in Table 14.37 and Table 14.38.

Hard boundaries were applied throughout in the case of the domains estimated by OK. Only one estimation pass was generally considered with a search neighbourhood of sufficient parameters to enable estimation of all required blocks. If required a second neighbourhood with expanded dimensions and relaxed sample requirements was applied to complete the estimation. Estimations were on the basis of SMU block dimensions throughout (3 mE by 5 mN by 4 mRL) and discretization was applied on the basis of 2 X by 2 Y by 2 RL for 8 discretization points per block.

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Table 14.37 Tap AB Gold Deposit - Open Pit Resource OK sample search criteria

Domain	Pass	Sample search orientation (dip / dip direction˚)			Sample search distance (m)			Numbers of samples
		Major	Semi- major	Minor	Major	Semi- major	Minor	Min.	Max.	Max per drillhole	Boundary condition
102	Pass 1	0/0	-20/270	-70/90	200	200	20	6	6	3	H
110	Pass 1	-13/38	16/312	69/90	200	200	20	6	12	3	H

Table 14.38 Urucum Gold Deposit - Open Pit Resource OK sample search criteria

Domain	Pass	Sample search orientation (dip / dip direction˚)			Sample search distance (m)			Numbers of samples
		Major	Semi- major	Minor	Major	Semi- major	Minor	Min.	Max.	Max per drillhole	Boundary condition
510 to 550	Pass 1	0/0	-15/270	-75/90	50	50	15	6	8	3	H
	Pass 2	0/0	-15/270	-75/90	200	200	30	4	8	-	H
700	Pass 1	0/0	-15/270	-75/90	50	50	15	6	8	3	H
	Pass 2	0/0	-15/270	-75/90	100	100	30	4	8	-	H

All relevant statistical information was recorded to enable validation and review of the MIK estimates. The recorded information included:

• Number of samples used per block estimate.

• Number of drillholes from which samples selected.

• Average distance to samples per block estimate and distance to nearest sample.

• Estimation flag to determine in which estimation pass a block was estimated.

• Number of drillholes from which composite data were used to complete the block estimate.

The estimates were reviewed visually and statistically prior to being accepted. The review included the following activities:

• Comparison of the E-type estimate versus the mean of the composite dataset, including weighting where appropriate to account for data clustering.

• Comparison of the reconstituted cumulative conditional distribution functions of the estimated blocks (indicator kriging) versus the input composite data.

• Visual checks of cross sections, long sections, and plans.

Alternative estimates were also completed to test the sensitivity of the reported model to the selected MIK interpolation parameters. An insignificant amount of variation in overall grade was noted in the alternate estimations.

14.7.1.7 Change of support

Applying the modelled variography, variance adjustment factors were calculated to emulate a 3 mE x 5 mN x 4 mRL selective mining unit (“SMU”) via the indirect lognormal change of support. The intra-class composite mean grades were used in calculating the whole block and SMU grades. The change of support study also included the calculation of the theoretical global change of support via the discrete Gaussian change of support model. An ‘information effect’ factor is commonly applied to the originally derived panel-to-block variance ratios to determine the final variance adjustment ratio. The goal of incorporating information effect is to calculate results taking into account that mining takes place based on grade control information. There will still be a quantifiable error associated with this data and it is this error we want to incorporate. This is achieved in practice by running a test kriging estimation of an SMU using grade control data (the results required to incorporate this option in the change of support do not depend on the assay data, so the grade control data can be hypothetical). The incorporation of the information effect is commonly found to be negligible, however it can have a significant effect in some cases. In this case, the information effect factor was found to have a minor effect and has been incorporated in the calculation. The variance adjustment ratios are provided in Table 14.39.

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Table 14.39 Variance adjustment ratios (3 mE x 5 mN x 4 mRL SMU)

Tap AB
Zone	101	200	1000	2000	3000
Variance adjustment factor (f)	0.20	0.25	0.09	0.30	0.15
Urucum
Zone	100	200	300	400	-
Variance adjustment factor (f)	0.15	0.30	0.15	0.25	-

14.7.2 Tap AB Underground - OK estimation

Gold grades were estimated by three consecutive OK runs performed using 3 m composites, parent cells of 9 m x 20 m x 8 m (x, y, z), hard wireframe boundaries and variograms determined by variogram sector. The key parameters of the estimation are chosen after several tests, a critical parameter being the number of composites to be used for the estimates. Table 14.40 details the first and second neighbourhood searches for each of the lode / oxide domains.

The search radii are related to the semi-variogram ranges for first and second runs and for the third run a large neighbourhood was used to fill the lode wireframe for blocks not estimated.

• 1st neighbourhood: search radius value varying to fit 85 - 90% of the variogram sill. Minimum number of samples = 4, Maximum = 32. 8 sectors and minimum of 3 drillholes in a grid of 120° spaced, 3D continuity.

• 2nd neighbourhood: search radius 2x of neighbourhood 1. Minimum number of Samples = 4, Maximum = 32, 8 sectors and minimum of 2 drillholes, 2D continuity.

• 3rd neighbourhood represents the search ellipsoid to estimate remain blocks inside the lode wireframe = 1,500 x 1,500 x 100 m. Minimum of 1 Sample and a maximum of 4.

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Table 14.40 Tap AB Gold Deposit - Underground Resource; orientation and radius of search radius of individual lodes / oxide domains

Ore shoots		VarSector	No blocks	Az	Dip	Plunge	Range neigh 1			Range neigh 2
Sample index	Block index						Dist X	Dist Y	Dist Z	Dist_U	Dist_V	Dist_W
LODE10OX	3,010	LODE10	2,905	10	-50	0	65	65	25	130	130	50
LODE11A	111	LODE6_11	4,027	0	85	0	35	55	10	70	110	20
LODE11B	112	LODE6_11	1,713	0	90	0	35	55	10	70	110	20
LODE12AOX	3,121	LODE12	393	0	-85	0	55	55	20	110	110	40
LODE12B	122	LODE12	102	10	-75	0	55	55	20	110	110	40
LODE12C	123	LODE12	218	0	-85	0	55	55	20	110	110	40
LODE12DOX	3,124	LODE12	153	0	85	0	55	55	20	110	110	40
LODE14A	141	LODE14	1,654	0	90	0	35	55	10	70	110	20
LODE14BOX	3,142	LODE14	1,655	15	-85	0	35	55	10	70	110	20
LODE1A	11	LODE1_3_5	99	-10	-85	0	35	55	10	70	110	20
LODE1BOX	3,012	LODE1_3_5	60	10	90	0	35	55	10	70	110	20
LODE1COX	3,013	DIP45W	98	10	-60	0	35	55	10	70	110	20
LODE3OX	3,003	LODE1_3_5	296	5	90	0	35	55	10	70	110	20
LODE4OX	3,004	LODE4	2,513	0	90	0	25	45	30	50	90	60
LODE5OX	3,005	LODE1_3_5	418	0	90	0	35	55	10	70	110	20
LODE6OX	3,006	LODE6_11	820	-10	90	0	35	55	10	70	110	20
LODE9AOX	3,091	DIP45W	1,514	0	-45	0	35	55	10	70	110	20
LODE9BOX	3,092	DIP45W	121	15	-45	0	35	55	10	70	110	20
MLODE2AOX	3,211	MLODES	2,359	0	-80	0	35	55	10	70	110	20
MLODE2BOX	3,212	MLODES	466	5	90	0	35	55	10	70	110	20
MLODE2COX	3,213	MLODES	141	0	90	0	35	55	10	70	110	20
LODE10FR	5,010	LODE10	1,762	10	-50	0	65	65	25	130	130	50
LODE12AFR	5,121	LODE12	434	0	-85	0	55	55	20	110	110	40
LODE12DFR	5,124	LODE12	505	0	85	0	55	55	20	110	110	40
LODE14BFR	5,142	LODE14	2,236	15	-85	0	35	55	10	70	110	20
LODE1BFR	5,012	LODE1_3_5	54	10	90	0	35	55	10	70	110	20
LODE1CFR	5,013	DIP45W	62	10	-60	0	35	55	10	70	110	20
LODE3FR	5,003	LODE1_3_5	118	5	90	0	35	55	10	70	110	20
LODE4FR	5,004	LODE4	1,324	0	90	0	25	45	30	50	90	60
LODE5FR	5,005	LODE1_3_5	249	0	90	0	35	55	10	70	110	20

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Ore shoots		VarSector	No blocks	Az	Dip	Plunge	Range neigh 1			Range neigh 2
Sample index	Block index						Dist X	Dist Y	Dist Z	Dist_U	Dist_V	Dist_W
LODE6FR	5,006	LODE6_11	585	-10	90	0	35	55	10	70	110	20
LODE9AFR	5,091	DIP45W	960	0	-45	0	35	55	10	70	110	20
LODE9BFR	5,092	DIP45W	312	15	-45	0	35	55	10	70	110	20
MLODE2AFR	5,211	MLODES	3,380	0	-80	0	35	55	10	70	110	20
MLODE2BFR	5,212	MLODES	203	5	90	0	35	55	10	70	110	20
MLODE2CFR	5,213	MLODES	163	0	90	0	35	55	10	70	110	20

VarSector = Variogram Sector, Dist = Distance

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14.7.2.1 Estimate validation

The statistical validation of the OK estimated grades involved a comparison of these grades by lode in each neighbourhood with the top cut 3 m composite mean grade (3M CMP), NN estimated grade and the slope of regression. This comparison was carried out to have a better understanding of the performance of kriging and allow parameter adjustments for each neighbourhood.

Swath Plot analysis was applied to validate the grades along the block model in 3 main directions: X, Y, and Z. All cases observe a reasonable correlation of samples and block models, especially for the Measured and Indicated categories. Some differences were observed, however, related to Inferred blocks and small number of blocks and samples.

14.7.3 Urucum North Underground OK estimation

Au block grades (g/t) were estimated by three consecutive OK runs performed using 1 m composites, parent cells of 8 m x 20 m x 20 m (x, y, z), hard wireframe boundaries and a variogram determined from all lodes. The OK run was carried out in three successive runs with its parameters based on QKNA analysis.

The search ellipse is divided in eight octants and an optimum number of four data points per octant was sought (i.e. 32 data points in total). In the first run, a minimum of four data points (in total) was required for the block to be estimated. This condition is relaxed to one in subsequent runs as shown below.

Estimation Pass 1:

1 Minimum number of samples: 12.

2 Search Ellipsoid - 130 x 120 x 25 m - Semi-variogram range considering all structures.

3 Optimum number of samples for each octant: 3.

4 Maximum number of consecutive empty octants: 3.

Estimation Pass 2:

5 Minimum number of Samples: 12.

6 Search Ellipsoid - 130 x 120 x 25 m - Semi-variogram range considering all structures.

7 Optimum number of samples for each octant: 3.

8 Maximum number of consecutive empty octants: 7.

Estimation Pass 3:

9 Minimum number of Samples: 4.

10 Search Ellipsoid - Large enough to estimate all blocks.

11 Optimum number samples per octant: No restriction.

12 Maximum number of consecutive empty octants: No Restriction.

The search ellipses were generally orientated in the same directions as the variogram anisotropy axes.

Search ellipsoids were reduced to 20 m for composites greater than a defined cut-off. The aim of the reduced search was to avoid smoothing of very high grades in the estimate. The quality of the OK is variable based on the criteria of the slope of regression, however many blocks were poorly estimated due to the paucity of data and relatively poor grade continuity in some areas, consequently the grades could be smoothed, and it is suggested to implement a closer spaced drilling to reduce this effect.

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14.7.3.1 Urucum North Underground model - estimate validation

The statistical validation of gold estimated grades involved a comparison of grades estimated in each neighbourhood with the sample mean grade, NN estimated grades, and the slope of regression separated by lodes. This comparison was undertaken to gain a better understanding of the performance of kriging, allow adjustments of parameters for each neighbourhood, and to get information on the number of estimated blocks.

A Swath Plot analysis was also applied to validate the grades along the block model in 3 main directions: X, Y, and Z. A comparison was made against sample mean grade, OK estimated grades, and NN estimated grades. All cases observe a reasonable correlation of samples and block models, especially at Measured and Indicated regions. It also showed some differences but related to Inferred blocks and a small number of blocks and samples.

14.7.4 Urucum South Underground model - OK estimation

Gold grades are estimated by OK using 3 m composites within parent cells of dimension 9 m x 20 m x 8 m (x, y, z) using hard wireframe boundaries and applying variography based on domains grouped by variogram sector. The key parameters of the estimation are chosen after several tests, being the number of composites to be used for the estimates and the best number for the 8 estimation sectors. A treatment of outliers was applied for samples for domains with less than 100 composites. These had no top cut applied however had the search radius reduced to 20 m for outliers above the determined limits.

The search radii are related to the semi-variogram ranges for first and second runs with the third using a larger neighbourhood to fill blocks not estimated within the lodes. The orientation of the search ellipsoid was based on the orientation of the individual lodes. A summary of the estimation parameters by neighbourhood are follows:

• 1st neighbourhood: search radius 35 m x 30 m x 15 m. Minimum number of samples = 4, maximum number of samples = 32, 8 sectors and a minimum 3 drillholes around a 120° spaced, 3D continuity.

• 2nd neighbourhood: search radius 70 m x 60 m x 30 m. Minimum number of Samples = 4, maximum number of samples = 32, octant search and a minimum of 2 drillholes around 180° spaced, 2D continuity.

3rd neighbourhood: The search ellipsoid to estimate remain blocks within the lode wireframe. Search radius 500 m x 500 m x 100 m. Minimum of 1 Sample and a maximum of 4 samples. Details of the search orientation for each lode domain are detailed in Table 14.41.

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Table 14.41 Urucum South Underground model: orientation of search radius of individual lodes

Lode code	VarSector	No blocks	Az	Dip	Plunge	Lode code	VarSector	No blocks	Az	Dip	Plunge
LOCS800	350to360_55to70E	80	-10	70	10	LOUN600	350to360_35to50E	37	-5	50	0
LOCS801	350to360_55to70E	110	-10	70	10	LOUN601	350to360_55to70E	157	-5	55	10
LOCS802	335to340_55to70E	67	-20	60	10	LOUN602	350to360_0to30E	281	-5	30	0
LOCS803	350to360_55to70E	34	0	70	10	LOUN603	350to360_35to50E	135	-5	45	0
LOCS804	350to360_55to70E	218	-10	70	10	LOUN605	350to360_35to50E	67	-5	40	0
LOUS902	350to360_55to70E	1,133	0	55	10	LOUN606	350to360_35to50E	119	-5	45	0
LOUS903	350to360_55to70E	1,441	-10	60	10	LOUN607	350to360_35to50E	12	0	50	0
LOUS904	350to360_55to70E	319	-5	65	10	LOUN608	350to360_75to90E	62	-5	80	10
LOUS905	350to360_0to25W	68	-20	-10	0	LOUN609	350to360_55to70E	29	-5	55	10
LOCN700	350to360_0to25W	2,330	0	-15	0	LOUN610	335to340_55to70E	23	-20	60	10
LOCN701	350to360_55to70E	245	0	70	10	LOUN612	350to360_35to50E	3	0	40	0
LOCN702	350to360_55to70E	5,606	-10	70	10	LOUN614	350to360_75to90E	74	-10	75	10
LOCN704	350to360_55to70E	2,888	-10	70	10	LOUN615	350to360_75to90E	61	0	85	10
LOCN706	350to360_75to90E	74	-10	75	10	LOUN616	335to340_55to70E	16	-20	60	10
LOCN708	350to360_55to70E	48	0	60	10	LOUN617	350to360_35to50E	25	0	-40	0
LOCN709	350to360_35to50E	79	0	35	0	LOUN618	335to340_55to70E	22	-20	60	10
LOCN710	350to360_75to90E	29	-5	80	10	LOUN620	350to360_0to30E	3	0	30	0
LOCN711	350to360_55to70E	31	0	60	10	LOUN621	350to360_55to70E	6	-5	60	10
LOCN712	350to360_75to90E	38	0	90	10	LOUN623	350to360_0to25W	8	0	-25	0
LOCN713	350to360_0to30E	16	0	25	0	LOUN624	350to360_55to70E	5	0	55	10
LOCN715	350to360_0to25W	26	0	-25	0	LOUN625	350to360_35to50E	5	0	45	0
LOCN716	350to360_55to70E	39	0	60	10	LOUN626	350to360_0to25W	129	0	-25	0
LOCN717	335to340_55to70E	44	-20	70	10	LOUN627	350to360_75to90E	82	0	85	10
LOCN718	350to360_55to70E	57	0	70	10	LOUN628	350to360_75to90E	45	10	85	10
LOCN719	350to360_75to90E	42	0	80	10	LOUS900OX	350to360_55to70E	180	-10	60	10
LOCN720	350to360_55to70E	23	-20	-65	10	LOUS900FR	350to360_55to70E	441	-10	60	10
LOCN722	350to360_0to25W	28	0	-5	0	LOUS901OX	335to340_55to70E	103	-25	60	10
LOLN001	350to360_75to90E	7,171	-5	80	10	LOUS901FR	335to340_55to70E	176	-25	60	10
LOLN002	350to360_75to90E	5,480	-5	80	10	LOCN703OX	350to360_55to70E	481	-5	70	10
LOLN299	350to360_75to90E	2,685	-5	80	10	LOCN703FR	350to360_55to70E	1498	-5	70	10
LOLN300	350to360_75to90E	5,827	-5	80	10	LOCN705OX	350to360_55to70E	48	-5	70	10

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Lode code	VarSector	No blocks	Az	Dip	Plunge	Lode code	VarSector	No blocks	Az	Dip	Plunge
LOLN301	350to360_75to90E	3,109	0	80	10	LOCN705FR	350to360_55to70E	97	-5	70	10
LOLN400	350to360_75to90E	2,950	-5	80	10	LOCN707OX	335to340_55to70E	72	-20	55	10
LOLN500	350to360_75to90E	149	0	80	10	LOCN707FR	335to340_55to70E	44	-20	55	10
LOUN510	350to360_0to25W	2,229	25	-15	0	LOCN721OX	350to360_0to30E	80	0	0	0
LOUN530	350to360_0to25W	201	0	-15	0	LOCN721FR	350to360_0to30E	51	0	0	0
LOUN540	350to360_0to25W	322	0	-15	0	LOUN520OX	350to360_0to25W	404	0	-15	0
LOUN550	350to360_0to25W	85	5	-30	0	LOUN520FR	350to360_0to25W	228	0	-15	0
						LOUN613OX	350to360_75to90E	49	-10	80	10
						LOUN613FR	350to360_75to90E	57	-10	80	10

Notes:

Azimuth Negative = Northwest quadrant

Dip Negative = Southwest quadrant

Plunge Negative = South quadrant

VarSector = Variogram sector

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14.7.4.1 Urucum South Underground model - estimate validation

The statistical validation of estimated grades consisted in a comparison of these grades by neighbourhood (“NN”) and lode with the sample mean grades, NN estimated grades and the slope of regression.

The Swath Plot analysis was applied to validate the grades along the block model in 3 main directions: X, Y, and Z. The procedure involved a comparison between the sample mean grade, OK estimated grade, and NN estimated grade along slices. A high difference with a large number of samples indicates a problem with estimation process and requires further investigation. In all cases there was a reasonable correlation of samples and the block model grades especially in the Measured and Indicated regions. There were some differences however these related to Inferred blocks and small number of blocks and samples.

14.7.5 Urucum East model-OK estimation

Traditional 3D OK in Surpac software was used to interpolate gold into Y = 10 m x X = 5 m x Z = 2 m parent cells for all Urucum East mineralized domains. All domains were estimated individually using the uniquely coded 1 m downhole composite data. Urucum East estimation parameters are summarized in Table 14.42.

Table 14.42 Urucum East summary estimation parameters

Parameter	HG zone	Colluvium
Minimum number of composites	3	4
Maximum number of composites	20	24
Search major distance	76	155
Search orientation	0	0
Plunge of major axis	-30	0
Dip of major axis	0	0
Anisotropy major / semi-major	1.44	1
Anisotropy major / minor	10	1
Block discretization	3x2x2	3x2x2

14.7.5.1 Estimate validation

A variety of validation checks on the final estimated block model were undertaken. The validation checks included swath plots and visual inspection.

Swath plots were used to detect any obvious spatial biases between the OK estimates and the informing composites. As expected, in the swath plots the OK estimates are somewhat more smoothed than the raw composite grades. Some minor local discrepancies exist (e.g. 403,030 mE), however excluding these there is little evidence of obvious spatial bias.

A visual validation was undertaken of composite grades and estimated model for all estimated domains. It was concluded that the grade estimation process adequately reflects the input data.

14.7.6 Tap C (other) estimation

Gold block grades are estimated by OK performed using parent cells, sample composites within hard boundaries, and estimation based on the variography of domains grouped by variogram sectors.

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The key parameters of the estimation were chosen after several tests. A critical parameter was the number of composites to be used for the estimates and the best number for each sector. The choice was guided by optimizing the slope of regression.

The kriging took place in three successive runs with the search radii related to the semi-variogram ranges and the third run using an increasingly larger neighbourhood to fill in the blocks not estimated within the mineralization wireframes. The search ellipses are generally orientated in the same direction as the variogram’s anisotropic axes.

Typically, the search ellipse is divided in eight octants, and an optimum number of four data points per octant is sought (i.e. 32 data points in total). In the first run, a minimum of four data points in total was required for the block to be estimated. This condition is relaxed to one in subsequent runs.

• 1st neighbourhood represents the distance of 70% to 80% of Sill, the distance of the main inflection of the semi-variogram, minimum number of samples 4, Maximum 32, octants search and a minimum of 2 drillholes to estimate a block. Blocks estimated in this search strategy were classified as Measured.

• 2nd neighbourhood represents the distance of 100% of Sill, minimum number of samples 2, maximum 16, octant search and a minimum of 2 drillholes to estimate a block. Restrictions were applied to the search radius for high grade values. Blocks estimated in this search strategy were classified as Indicated

• 3rd neighbourhood represents the search ellipsoid to estimate remain blocks, sectors North and South: U = 720; V = 240; W = 80; minimum of 1 Sample, maximum of 4. Restrictions were applied to the search radius for high grade values. Blocks estimated in this search strategy were classified as Inferred.

The quality of the OK is variable based on the criteria of the slope of regression. Many blocks were poorly estimated, mostly due to the paucity of data in some areas and the relatively poor grade continuity, and consequently the grades could be smoothed. It is suggested to implement a closer spaced drilling in grade-smoothed areas to reduce this effect.

14.7.6.1 Estimate validation

The statistical validation of Au estimated grades (g/t) was undertaken by a comparison of these grades estimated in each neighbourhood against the sample mean grade, nearest neighbourhood estimated grades and the slope of regression, separated by lode.

Table 14.43 gives an example for the first neighbourhood.

This comparison was carried out to gain an understanding of the performance of the OK estimation allowing for some parameter adjustments of each neighbourhood and to obtain information of the number of estimated blocks. This work assisted in defining cut-offs for high grade transition (e.g. reduced search ellipsoid to 20 m for Au (g/t) greater than the defined COG). The overall aim was to avoid excessive smoothing of grades and control the differences so as to not exceed 10%.

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Table 14.43 Statistical validation of Au (g/t) of 1st neighbourhood

Lode_type		No blocks			Statistics of Au estimated blocks (g/t)				Au grades to control estimation (g/t)			Percentile
Samp index	Block index	Total	Estimated	Percent	Count	Mean grade	Mean slope	Mean slope	2M CMP No capping	2M CMP capping	Dif Krig / CMP
LODE3_4_5101	EX_LODE3_4_5101	8,121	3,276	40.34%	3,276	0.870	0.713	0.713	0.919	1.003	-13.2%	9.2%
LODE102	EX_LODE102	3,770	659	17.48%	659	1.345	0.659	0.659	1.543	1.594	-15.6%	3.3%
LODE111	EX_LODE111	1,681	499	29.68%	499	0.915	0.644	0.644	1.054	1.095	-16.4%	3.9%
LOGR020	EX_LOGR020	1,262	153	12.12%	153	0.687	0.523	0.523	0.679	0.715	-4.0%	5.3%
LODE103	EX_LODE103	130	45	34.62%	45	0.651	0.545	0.545	0.710	0.710	-8.3%	0.0%
LODE104	EX_LODE104	207	32	15.46%	32	0.708	0.714	0.714	0.709	0.709	-0.2%	0.0%
LODE105	EX_LODE105	80	28	35.00%	28	0.543	0.721	0.721	0.537	0.537	1.1%	0.0%
LODE108	EX_LODE108	343	71	20.70%	71	0.526	0.572	0.572	0.549	0.549	-4.3%	0.0%
LODE107	EX_LODE107	126	21	16.67%	21	1.052	0.787	0.787	1.089	1.106	-4.9%	1.5%
LODE106	EX_LODE106	443	130	29.35%	130	1.574	0.783	0.783	1.715	1.807	-12.9%	5.4%
LODE109	EX_LODE109	604	89	14.74%	89	0.969	0.615	0.615	0.996	0.996	-2.7%	0.0%
LODE110	EX_LODE110	514	126	24.51%	126	0.879	0.716	0.716	0.870	0.908	-3.2%	4.4%
LGEV3_4_5010	EX_LGEV3_4_5010	11,309	4,254	37.62%	4,254	0.193	0.381	0.381	0.196	0.214	-10.0%	9.2%
LODE2101	EX_LODE2101	2,895	2,110	72.88%	2,110	1.450	0.768	0.768	1.425	1.486	-2.5%	4.3%
LGEV2010	EX_LGEV2010	4,840	2,047	42.29%	2,047	0.502	0.543	0.543	0.522	0.559	-10.2%	7.1%
Total		36,325	13,540	37.27%	13,540	0.719	0.583	0.583	0.747	0.794	-9.5%	6.3%

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Swath plot analysis was undertaken to validate the block model grades in 3 main directions: X, Y, and Z. The procedure involved a comparison of sample mean grade, OK estimated grades, nearest neighbourhood estimated grades in each direction. If the difference compared to the block mean is high with a large number of samples then the estimation process has problems and needs to be revised. In all cases there was observed a reasonable correlation of samples and block model grades, especially in the Measured and Indicated regions. There were noted some differences but these related to Inferred blocks and a small number of blocks and samples.

14.7.7 Tap C Model (Tap C3N and Tap C3W areas)

Estimation of the Tap C 31 (Tap C3N and C3W) update was principally carried out on 3 grouped domains. The groupings were determined by domain orientation and location. Estimation was carried out by OK in Surpac software using the mineralization wireframes as hard boundaries. The estimation was carried out in 2 passes, the first to the maximum range of the variogram and the second extended to fill the lode solid. An elliptical search was used with a minimum of 3 and maximum of 15 samples used to inform the estimation in both passes. OK parameters used for each domain are detailed in Table 14.44 below.

Table 14.44 Estimation parameters for Tap C (Tap C3N and Tap C3W areas)

			Sill (relative variance)	Range	Azimuth	Plunge	Dip	Major / semi major ratio	Major / minor ratio
Zone 1	Moderate East Dipping Lodes	Nugget	0.38
		S 1	0.02	18.9	0	0	-50	1.4	0.9
		S 2	0.07	112.7	0	0	-50	1.9	1.9
Zone 2	Flat W Dipping Lodes	Nugget	0.26
		S 1	0.08	15.8	0	0	10	0.6	0.6
		S 2	0.04	79.5	0	0	10	1.2	2.7
Zone 7	Flat Dipping Lodes	Nugget	0.22
		S 1	0.02	46.7	0	0	0	1.4	5.5
		S 2	0.10	94.3	0	0	0	1.1	3.3

14.7.8 Stockpile estimation

Stockpiles are calculated on a monthly basis by balancing mass and metal of the survey reconciled mine production, mill throughput, and previous end of month stockpiles. The mine production is estimated using the grade control model for gold and the resource model for density. As-mined, adjusted dig polygons, and end of month survey pickups are used to obtain mined ore volumes which are then interrogated with the grade control model to determine tonnes and grade.

Ore categories are divided on the basis of grade range (Very High Grade, High Grade, Low Grade, and Marginal; see Table 16.6) and material type (Oxide or Sulphide). The survey reconciled mine production is distributed to the various intermediary "live" stockpiles located proximal to active pit areas and the ROM stockpiles on the basis of these categories. Tallied truck tonnes are compared against the survey reconciled production to check for any discrepancies between actual material movement and the polygons mined. Ore movement from live stockpiles to the ROM during the month is also determined as part of the distribution of monthly mine production, to its final destination.

Daily stockpile blending on the ROM is adjusted to plant reconciled total dry tonnes and grade for the month. The monthly stockpile balance for both mass and Au metal is then undertaken as follows; Opening Stockpile + Survey Reconciled Production = Mill Throughput + Closing Stockpile. Each stockpile is surveyed at the end of month and the balance of the mass and metal checked against the stockpile volume to ensure that swelled density is within the acceptable range for the material type.

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Spent Ore from the previous heap leach operation is periodically added to the blend. The mass and grade of the Spent Ore has been determined from a grade control model based initially on a RC drilling pattern of 40 m x 40 m grid and later a 20 m x 20 m grid. Density was determined with 52 density test pits completed across the deposit. Both the pre-mining total tonnage and grade of Spent Ore from the model was a close match with the final estimate of Spent Ore tonnes and grade determined by MPBA plant metallurgists at the closure of the heap leach. The tonnes and grade of Spent Ore remaining in the stockpile is calculated from the grade control model and both the pre-mining and end of month topographic surfaces.

14.7.9 Duckhead estimation

Gold grades are estimated by three consecutive OK runs performed using parent cells of dimensions 5 m x 10 m x 2 m (x, y, z), hard boundaries and 2 m composites. The key parameters for the estimation were selected, with the critical parameter being the number of composites to be used and the optimal number of composites per sector in the octant search. The search radii are related to the semi-variogram ranges for first and second estimation runs, for the third a large neighbourhood was used to fill the not estimated blocks with the orientation, based on that individual domains. The summary of the search neighbourhood parameters is following:

• 1st neighbourhood with search radius value varying according to fit 85% to 90% of the sill of the variograms, minimum number of samples 4, Maximum 32, 8 sectors and Minimum 3 drillholes in a grid of 120° spaced - 3D continuity.

• 2nd neighbourhood search radius 2x Neighbourhood 1, minimum number of Samples 4, Maximum 32, octants and minimum 2 drillholes - 2D continuity

• 3rd neighbourhood represents the search ellipsoid to estimate remain blocks, 1,500 m x 1,500 m x 100 m; minimum of 1 Sample, maximum of 4.

Table 14.45 below shows the Duckhead search ellipses by oxidation domain.

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Table 14.45 Duckhead search ellipses by lode / oxidation domain

Rock codes		VarSector	No blocks	High grade transition		Az	Dip	1st neigh			2nd neigh			3rd neigh
								Dist X	Dist Y	Dist Z	Dist X	Dist Y	Dist Z	Dist X	Dist Y	Dist Z
LODE1	1	LODE1	595	999.0	Yes	95	60	20	45	20	40	100	40	1,000	1,000	100
LODE100	100	LOWDIP	521	999.0	Yes	125	15	30	30	10	60	100	20	1,000	1,000	100
LODE101	101	LOWDIP	1,263	1.5	Yes	105	10	30	30	10	60	100	20	1,000	1,000	100
LODE102	102	LOWDIP	895	999.0	Yes	115	15	30	30	10	60	100	20	1,000	1,000	100
LODE103	103	LOWDIP	101	999.0	Yes	125	15	30	30	10	60	100	20	1,000	1,000	100
LODE104	104	LOWDIP	36	999.0	Yes	125	15	30	30	10	60	100	20	1,000	1,000	100
LODE105	105	LOWDIP	34	999.0	Yes	125	15	30	30	10	60	100	20	1,000	1,000	100
LODE106	106	LOWDIP	1,015	2.0	Yes	125	15	30	30	10	60	100	20	1,000	1,000	100
LODE107	107	LOWDIP	2,122	999.0	Yes	110	15	30	30	10	60	100	20	1,000	1,000	100
LODE108	108	HIGHDIP	3,295	999.0	Yes	135	70	30	50	15	60	100	30	1,000	1,000	100
LODE109	109	HIGHDIP	219	1.5	Yes	140	60	30	50	15	60	100	30	1,000	1,000	100
LODE2	2	LODES2_4_6	2,478	999.0	Yes	100	70	30	30	10	60	100	20	1,000	1,000	100
LODE3	3	HIGHDIP	3,716	999.0	Yes	110	70	30	50	15	60	100	30	1,000	1,000	100
LODE31	31	LOWDIP	2,381	999.0	Yes	125	45	30	30	10	60	100	20	1,000	1,000	100
LODE71	71	HIGHDIP	1,298	999.0	Yes	120	75	30	50	15	60	100	30	1,000	1,000	100
LODE8	8	HIGHDIP	1,062	999.0	Yes	115	85	30	50	15	60	100	30	1,000	1,000	100
LODE9	9	LOWDIP	3,122	999.0	Yes	110	10	30	30	10	60	100	20	1,000	1,000	100
LODE90	90	LOWDIP	2,675	3.5	Yes	135	25	30	30	10	60	100	20	1,000	1,000	100
LODE110OX	3110	HIGHDIP	1,538	2.5	Yes	115	70	30	50	15	60	100	30	1,000	1,000	100
LODE110FR	5110	HIGHDIP	1,583	2.5	Yes	115	70	30	50	15	60	100	30	1,000	1,000	100
LODE4OX	3004	LODES2_4_6	1,348	999.0	Yes	120	70	30	30	10	60	100	20	1,000	1,000	100
LODE4FR	5004	LODES2_4_6	1,383	999.0	Yes	120	70	30	30	10	60	100	20	1,000	1,000	100
LODE5OX	3005	HIGHDIP	1,806	999.0	Yes	120	70	30	50	15	60	100	30	1,000	1,000	100
LODE5FR	5005	HIGHDIP	1,976	999.0	Yes	120	70	30	50	15	60	100	30	1,000	1,000	100
LODE51OX	3051	HIGHDIP	239	999.0	Yes	125	70	30	50	15	60	100	30	1,000	1,000	100
LODE51FR	5051	HIGHDIP	768	999.0	Yes	125	70	30	50	15	60	100	30	1,000	1,000	100
LODE6OX	3006	LODES2_4_6	66	8.0	Yes	130	70	30	30	10	60	100	20	1,000	1,000	100
LODE6FR	5006	LODES2_4_6	147	8.0	Yes	130	70	30	30	10	60	100	20	1,000	1,000	100
LODE7OX	3007	HIGHDIP	749	999.0	Yes	115	70	30	50	15	60	100	30	1,000	1,000	100
LODE7FR	5007	HIGHDIP	3,980	999.0	Yes	115	70	30	50	15	60	100	30	1,000	1,000	100

VarSector = Variogram Sector, Dist = Distance

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14.8 Depletion for mining activity

Depletion to account for mining activity has been applied to the models. Depletion has been applied as of 30 June 2017 via the use of surveyed topographic surfaces as at that date. Depletion has been applied by block model flag to identify the mined and in-situ portions of the models.

14.9 Resource classification

Resource classification was completed based on an assessment of the following criteria:

• Geological knowledge and interpretation.

• Data distribution including drilling density and orientation.

• Variogram models and the ranges of the first structure in multi-structure models.

• Estimation statistics.

The model was inspected visually, and manual adjustments made where appropriate.

Table 14.46 describes the criteria used to classify each deposit.

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Table 14.46 Resource classification criteria

Deposit		Measured	Indicated	Inferred	Notes
Urucum Open Pit		- Blocks insitu (not mined) - Blocks estimated in pass 1 - Distance to nearest data ≤ 20 m - Average distance to samples ≤ 35 m - ≥ 24 data used from ≥ 9 drillholes	- Blocks insitu (not mined) - Blocks estimated in pass 1 or 2 - Average distance to samples ≤ 75 m - ≥ 24 data used from ≥ 9 drillholes	- All other insitu blocks (not mined)
Tap AB Open Pit		- Blocks insitu (not mined) - Blocks estimated in pass 1 - Distance to nearest data ≤ 15 m - 24 data used from ≥ 7 drillholes	- Blocks insitu (not mined) - Blocks estimated in pass 1 or 2 - Average distance to samples ≤ 60 m - ≥ 24 data used from ≥ 9 drillholes	- All other insitu blocks (not mined)
Tap AB Underground		- Blocks estimated by ≥ 3 drillholes in pass 1 - minimum number of samples required	- Blocks estimated with ≥ 2 and ≤ 16 samples in octant search - minimum 2 drillholes	- All other blocks	Moving average post processing applied to individual blocks. Grouped on panels lode thickness x 60 m in North direction15m in Z direction to minimize spottiness of classification
Urucum North Underground	Lode 1 (Ore shoots 1001-1004)	- Slope of regression > 0.7	- Slope of regression 0.4 to 0.7	- Slope of regression < 0.4	Classified based on geological knowledge of lode/ore shoot, drill grid and slope of regression, subsequent visual inspection and manual adjustment
	Lode 2 (Ore shoots 2001-2003, 2005)	- Slope of regression > 0.7	- Slope of regression 0.4 to 0.7	- Slope of regression < 0.4
	Lode 300 (Ore shoot 3001-3005)	- Slope of regression > 0.7	- Slope of regression 0.4 to 0.7	- Slope of regression < 0.4
	Lode 400 (Ore shoot 4001-4002, 4005)	- Slope of regression > 0.7	- Slope of regression 0.4 to 0.7	- Slope of regression < 0.5
	Lode 100, 299, 301	N / A	N / A	N / A	Adjusted to be unclassified
	Lode 500	N / A	N / A		Manually classified as inferred
Urucum East 2013		N / A	- Drill spacing sufficient to demonstrate confidence in geometry and continuity of grade - Blocks within 30 m of sample	- Infill drilling required - Blocks between 30 and 76 m from sample	Blocks <76 m from sample not classified. Applied to both mineralized lode and colluvium domains.
Tap C (other)		- At least 3 drillholes in first pass - minimum number of samples used in kriging	- Blocks estimated with ≥ 2 and ≤ 16 samples in octant search - minimum 2 drillholes	- All other blocks	Distribution of data, number of samples used to estimate the block and range of the semi-variogram. Slope of regression was used to validate.
Tap C (Tap C3N and C3W areas)		- Average distance to sample <20 m	Average distance to sample <5 0m	- All other blocks within domain wireframe.

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Figure 14.13 shows an example of the resource classification for the Urucum East model.

Figure 14.13 Urucum East Model - resource classification - HG Lode

Note: Green = Indicated, Purple = Inferred.

Source: BDR

14.10 Resource reporting

The summary total Mineral Resource for the Tucano Gold Project is provided as of 30 June 2017 in Table 14.47 below. The COG for reporting is 0.5 g/t Au for open pit resources and most stockpiles and 1.2 g/t Au for underground resources. In view of the nature and style of the mineralization and potential mining approach and method, this is considered an appropriate COG. It should be noted that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Mineral Resources have been estimated in accordance with Canadian National Instrument 43-101 “Standards of Disclosure for Mineral Projects” ("NI 43-101") of August 2011 and “Definition Standards for Mineral Resources and Mineral Reserves” adopted by the CIM Council on May 2014. Measured and Indicated Mineral Resources are inclusive of those Mineral Resources modified to produce the Mineral Reserves.

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Table 14.47 Tucano Gold Resources 30 June 2017

Brazil	Cut-off	Measured			Indicated			Total Measured & Indicated			Inferred
	Au (g/t)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)
Urucum Open Pit oxide	0.5	678	1.15	25	344	1.08	12	1,021	1.13	37	70	0.97	2.16
Urucum East Open Pit oxide	0.5	0	0.00	0	200	1.88	12	200	1.88	12	9	1.58	0.44
Tap AB Open Pit oxide	0.5	2,870	2.18	201	3,889	1.76	220	6,759	1.94	421	796	1.11	28.47
Tap C Open Pit oxide	0.5	566	0.97	18	312	0.90	9	878	0.95	27	88	0.70	1.99
Duckhead Open Pit oxide	1	89	4.24	12	140	1.74	8	229	2.71	20	60	1.56	3.01
Total oxide		4,202	1.90	256	4,885	1.66	261	9,087	1.77	517	1,022	1.10	36.07
Urucum Open Pit primary	0.5	5,783	1.54	286	7,120	1.71	393	12,903	1.63	678	393	1.57	20.00
Urucum East Open Pit fresh	0.5	0	0.00	0	211	1.45	10	211	1.45	10	84	0.94	3.00
Tap AB Open Pit primary	0.5	2,356	1.74	132	4,176	1.57	210	6,531	1.63	342	1,053	1.34	45.00
Tap C Open Pit primary	0.5	482	1.22	19	1,980	1.22	78	2,462	1.22	97	1,044	1.35	45.00
Duckhead surface primary	1	115	2.28	8	264	2.26	19	379	2.26	28	262	1.81	15.00
Open Pit primary		8,736	1.58	445	13,750	1.60	709	22,486	1.60	1,154	2,837	1.41	128.00
Urucum Open Pit total	0.5	6,460	1.50	311	7,464	1.69	404	13,924	1.60	715	463	1.48	21.97
Urucum East Open Pit total	0.5	0	0.00	0	411	1.66	22	411	1.66	22	92	1.00	2.97
Tap AB Open Pit total	0.5	5,225	1.98	333	8,065	1.66	430	13,290	1.79	764	1,849	1.24	73.86
Tap C Open Pit total	0.5	1,048	1.09	37	2,292	1.18	87	3,340	1.15	123	1,133	1.30	47.32
Duckhead Open Pit total	1	205	3.14	21	404	2.08	27	608	2.43	48	322	1.76	18.25
Total oxide & primary open pit		12,938	1.69	701	18,635	1.62	970	31,573	1.65	1,671	3,859	1.32	164.37
Open pit stockpile	0.5	1,985	0.66	42	0	0.00	0	1,985	0.66	42	0	0.00	0.00
Spent ore stockpile	0.5	1,311	0.61	26	0	0.00	0	1,311	0.61	26	0	0.00	0.00
Rom expansion stockpile	0.5	345	0.80	9	0	0.00	0	345	0.80	9	0	0.00	0.00
Marginal ore stockpiles	0.3	1,586	0.44	22	0	0.00	0	1,586	0.44	22	0	0.00	0.00
Total stockpiles		5,228	0.59	99	0	0.00	0	5,228	0.59	99	0	0.00	0.00
Total Tucano Open Pit & stockpiles		18,166	1.37	800	18,635	1.62	970	36,801	1.50	1,770	3,859	1.32	164.37
Tap AB Underground primary*	1.2	22	1.49	1	1,025	2.3	76	1,047	2.28	77	3,653	3.19	375
Urucum Underground primary - pfs	1.6	260	4.06	34	2,634	4.24	359	2,894	4.22	393	5,106	2.71	445
Urucum Underground primary - non pfs	1.2	0	0.00	0	0	0	0	0	0.00	0	3,733	1.38	166
Total Tucano Underground		282	3.86	35	3,659	3.70	435	3,941	3.71	470	12,492	2.45	986
Grand total Tucano		18,448	1.41	835	22,293	1.96	1,405	40,742	1.71	2,240	16,351	2.19	1,150

Notes:

* The June 2017 Tap AB UG Resource includes 173 kt at 4.68 g/t Au of Inferred Oxide within the Inferred Primary Category.

• CIM Definition Standards (2014) were used for reporting the Mineral Resources.

• Mr B. Wolfe, MAIG is the Qualified Person under NI 43-101 and takes responsibility for the Mineral Resource estimate for the Tap AB Open Pit and Urucum Open Pit Resource. Mr M. Batelochi, MAusImm (CP) is the Qualified Person under NI 43-101 and takes responsibility for all remaining resources.

• Tap AB and Urucum Open Pit Mineral Resources are constrained by an optimized pit shell at a gold price of US$1,500 oz. The COG applied to the open pit resources is 0.5 g/t Au. The COG applied to ore stockpiles is 0.5 g/t Au with the exception of Marginal Ore which has a COG of 0.3 g/t Au.

• The underground COG is 1.2 g/t Au.

• Drilling results up to 30 June 2017 for the Tap AB Open Pit resource and 31 December 2016 for all other deposits.

• Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

• The numbers may not compute exactly due to rounding.

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14.11 Comparison with previous estimates.

The previous Mineral Resource estimate on the Property was published in a BDR ASX announcement “Ore Reserve and Mineral Resource Update” dated 11 April 2017. The previous Mineral Resources had an effective date of 31 December 2016.

Table 14.48 shows the Mineral Resource changes by deposit and physical reporting limits. Note that many deposits are unchanged since the 2016 estimate.

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Table 14.48 Mineral Resource changes by deposit and physical reporting limits

Deposit	Pit area	Mine method	Estimation method	Previous effective resource date	Current effective resource date	Change from previous to current effective date	Physical limits for resource reporting
Urucum	Urucum North	Open Pit (Surface)	MIK and OK	31 Dec 2016	30 June 2017	Model Re-reported Only.	North of 99,300 mN and above the US$1,100, 31 Dec 2016 reserve pit design. Also included is near surface resource located outside of the reserve pit and inside of the US$1,500 Measured + Indicated + Inferred (MII) optimized pit shell which does not form part of the Urucum North Underground Resources.
Urucum	Urucum South	Open Pit (Surface)	MIK and OK	31 Dec 2016	30 June 2017	No Change	South of 99,300 mN and above the 31 Dec 2016 US$1,500 MII optimized pit shell.
Urucum	Urucum North	Underground	OK	31 Dec 2016	30 June 2017	No Change	North of 99,300 mN and below 31 Dec 2016 reserve pit design (based on US$1,100 MII optimized pit shell).
Urucum	Urucum South	Underground	OK	31 Dec 2016	30 June 2017	No Change	South of 99,300 mN and below the 31 Dec 2016 US$1,500 MII optimized pit shell.
Urucum	Urucum East	Open Pit (Surface)	OK	31 Dec 2016	30 June 2017	No Change	No Pit Shell Reporting Limit Applied.
Tap C	Tap C	Open Pit (Surface)	OK	31 Dec 2016	30 June 2017	No Change	No Pit Shell Reporting Limit Applied.
Tap AB	Tap AB	Open Pit (Surface)	MIK and OK	31 Dec 2016	30 June 2017	Model Re-estimated and Reported	Above the 30 June 2016 US$1,500 MII optimized pit shell.
Tap AB	Tap AB	Underground	OK	31 Dec 2016	30 June 2017	Model Re-reported Only	Below the 30 June 2016 US$1,500 MII optimized pit shell.
Duckhead	Duckhead	Open Pit (Surface)	OK	31 Dec 2016	30 June 2017	No Change	No Pit Shell Reporting Limit Applied.

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Changes to the Mineral Resource estimated in this report are due predominately to:

• Additional drilling completed within the Tap AB optimized resource pit shell completed between the 31 December 2016 and 30 June 2017 resource estimates.

• Re-optimization of the Tap AB US$1,500 MII shell used to constrain the Tap AB Open Pit Resource and Tap AB Underground Resource as at 30 June 2017.

• The 31 December 2016 Mineral Resource estimate for Urucum North Open Pit resource included shallow, near surface Mineral Resources occurring outside of the US$1,500 MII Optimized Pit Shell and north of 99,300 mN. These resources were removed in the 30 June 2017 reported resources.

• 31 December 2016 Tap AB Underground Mineral Resources in the Measured and Indicated category were classified to Inferred by reason that a PFS had not been undertaken to determine viability of the underground resource. The 30 June 2017 resource has reported from the same model using the original geostatistical resource classification and not converting these Measured and Indicated Resources to Inferred Resource.

• Mining depletion of the Tap AB Open Pit Resource and Open Pit Stockpiles between 31 December 2016 and 30 June 2017.

• Conversion of part of the Open Pit Stockpile to the new ROM Expansion Stockpile.

• Increase in Marginal Stocks not processed by the plant.

Table 14.49 shows the comparison to the previous estimate.

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Table 14.49 Comparison of Mineral Resource estimates

Brazil	Cut-off	Measured - 31 Dec 2016			Measured - 30 June 2017				Indicated - 31 Dec 2016			Indicated - 30 June 2017			Total Measured & Indicated - 31 Dec 2016			Total Measured & Indicated - 30 June 2017				Inferred - 31 Dec 2016			Inferred - 30 June 2017
	Au (g/t)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)		Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)		Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)
Urucum Open Pit oxide	0.5	713	1.13	26	678	1.15	25	409		1	13	344	1.08	12	1,122	1.08	39	1,021	1.13	37	87		0.9	3	70	0.97	2
Urucum East Open Pit oxide	0.5	0	0	0	0	0	0	200		1.88	12	200	1.88	12	200	1.88	12	200	1.88	12	9		1.58	0	9	1.58	0
Tap AB Open Pit oxide	0.5	2,980	2.05	197	2,870	2.18	201	4,559		1.94	285	3,889	1.76	220	7,540	1.99	481	6,759	1.94	421	1,228		1.58	62	796	1.11	28
Tap C Open Pit oxide	0.5	566	0.97	18	566	0.97	18	312		0.9	9	312	0.9	9	878	0.95	27	878	0.95	27	88		0.7	2	88	0.7	2
Duckhead Open Pit oxide	1	89	4.24	12	89	4.24	12	140		1.74	8	140	1.74	8	229	2.71	20	229	2.71	20	60		1.56	3	60	1.56	3
Total oxide		4,348	1.81	252	4,202	1.9	256	5,621		1.81	327	4,885	1.66	261	9,969	1.81	579	9,087	1.77	517	1,472		1.49	70	1,022	1.1	36
Urucum Open Pit primary	0.5	5,822	1.53	287	5,783	1.54	286	7,184		1.71	395	7,120	1.71	393	13,007	1.63	682	12,903	1.63	678	410		1.53	20	393	1.57	20
Urucum East Open Pit primary	0.5	0	0	0	0	0	0	211		1.45	10	211	1.45	10	211	1.45	10	211	1.45	10	84		0.94	3	84	0.94	3
Tap AB Open Pit primary	0.5	2,155	1.78	123	2,356	1.74	132	4,233		1.81	247	4,176	1.57	210	6,388	1.8	370	6,531	1.63	342	2,804		1.62	146	1,053	1.34	45
Tap C Open Pit primary	0.5	482	1.22	19	482	1.22	19	1,980		1.22	78	1,980	1.22	78	2,462	1.22	97	2,462	1.22	97	1,044		1.35	45	1,044	1.35	45
Duckhead surface primary	1	115	2.28	8	115	2.28	8	264		2.26	19	264	2.26	19	379	2.26	28	379	2.26	28	262		1.81	15	262	1.81	15
Open Pit primary		8,575	1.59	437	8,736	1.58	445	13,872		1.68	749	13,750	1.6	709	22,447	1.64	1,186	22,486	1.6	1,154	4,604		1.55	229	2,837	1.41	128
Urucum Open Pit total	0.5	6,535	1.49	313	6,460	1.5	311	7,594		1.67	408	7,464	1.69	404	14,129	1.59	721	13,924	1.6	715	497		1.42	23	463	1.48	22
Urucum East Open Pit total	0.5	0	0	0	0	0	0	411		1.66	22	411	1.66	22	411	1.66	22	411	1.66	22	92		1	3	92	1	3
Tap AB Open Pit total	0.5	5,135	1.94	320	5,225	1.98	333	8,792		1.88	532	8,065	1.66	430	13,928	1.9	852	13,290	1.79	764	4,032		1.61	208	1,849	1.24	74
Tap C Open Pit total	0.5	1,048	1.09	37	1,048	1.09	37	2,292		1.18	87	2,292	1.18	87	3,340	1.15	123	3,340	1.15	123	1,133		1.3	47	1,133	1.3	47
Duckhead Open Pit total	1	205	3.14	21	205	3.14	21	404		2.08	27	404	2.08	27	608	2.43	48	608	2.43	48	322		1.76	18	322	1.76	18
Total oxide & primary open pit		12,923	1.66	690	12,938	1.69	701	19,492		1.72	1,075	18,635	1.62	970	32,415	1.69	1,765	31,573	1.65	1,671	6,077		1.53	299	3,859	1.32	164
Open pit stockpile	0.5	2,581	0.73	61	1,985	0.66	42	0		0	0	0	0	0	2,581	0.73	61	1,985	0.66	42	0		0	0	0	0	0
Spent ore stockpile	0.5	1,721	0.65	36	1,311	0.61	26	0		0	0	0	0	0	1,721	0.65	36	1,311	0.61	26	0		0	0	0	0	0
Rom expansion stockpile	0.5	0	0	0	345	0.8	9	0		0	0	0	0	0	0	0	0	345	0.8	9	0		0	0	0	0	0
Marginal ore stockpiles	0.3	1,517	0.43	21	1,586	0.44	22	0		0	0	0	0	0	1,517	0.43	21	1,586	0.44	22	0		0	0	0	0	0
Total stockpiles		5,819	0.63	118	5,228	0.59	99	0		0	0	0	0	0	5,819	0.63	118	5,228	0.59	99	0		0	0	0	0	0
Total Tucano Open Pit & stockpiles		18,742	1.34	807	18,166	1.37	800	19,492		1.72	1,075	18,635	1.62	970	38,235	1.53	1,883	36,801	1.5	1,770	6,077		1.53	299	3,859	1.32	164
Tap AB Underground primary*	1.2	0	0	0	22	1.49	1	0		0	0	1,025	2.3	76	0	0	0	1,047	2.28	77	4,216		2.85	386	3,653	3.19	375
Urucum Underground primary (Cut-off >1.6 g/t Au)	1.6	260	4.06	34	260	4.06	34	2,634		4.24	359	2,634	4.24	359	2,894	4.22	393	2,894	4.22	393	5,106		2.71	445	5,106	2.71	445
Urucum Underground primary (Cut-off 1.2 g/t -1.6 g/t Au)	1.2 to 1.6	3	1.4	0	0	0	0	26		1.44	1	0	0	0	29	1.43	1	0	0	0	3,704		1.38	165	3,733	1.38	166
Total Tucano Underground		263	4.03	34	282	3.86	35	2,660		4.21	360	3,659	3.7	435	2,923	4.2	394	3,941	3.71	470	13,026		2.38	996	12,492	2.45	986
Grand total Tucano		19,005	1.38	841	18,448	1.41	835	22,152		2.02	1,436	22,293	1.96	1,405	41,157	1.72	2,277	40,742	1.71	2,240	19,102		2.11	1,295	16,351	2.19	1,150

Please see footnotes under Table 14.47 for the 30 June 2017 report.

Footnotes for the 31 Dec 2016 Mineral Resource estimate follow:

* The Dec 2016 Tap AB UG Resource includes 169 kt at 4.51 g/t Au for 25 koz Au of Inferred Oxide in the Inferred primary Category.

* The Dec 2016 Tap AB UG Resource was all classed as Inferred by BDR. The June 2017 Tap AB UG Resource used the geostatistical resource classification of the original estimated model.

Source: BDR.

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15 Mineral Reserve estimates

15.1 Summary

The Mineral Reserve estimates for the Tucano operation were reported in accordance with the guidelines of NI43-101 and the Canadian Institute of Mine Metallurgy and Petroleum Definition Standards for Mineral Resources and Mineral Reserves 2014 (CIM Standards).

“A Mineral Reserve is the economically mineable part of a Measured and/or Indicated Mineral Resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at Preliminary Feasibility or Feasibility level as appropriate that includes application of Modifying Factors. Such studies demonstrate that, at the time of reporting, extraction could be reasonably justified.”

Open Pit Mineral Reserves were estimated through an open pit optimization exercise using the Measured and Indicated Resources in the block models provided by B Wolfe for the Tap AB and Urucum deposits and Measured and Indicated Resources in the block models provided by Marcello Batelochi and Paul Tan for the Urucum East and Tap C deposits.

Mineral Reserves were reported within detailed engineered pit designs and the LOM plans based on these pit shells.

The Mineral Reserves inside the engineered pit designs for the Tap AB open pits were estimated using COGs estimated for each deposit and ore type based on the following gold prices; Open Pit Optimization Shell used for Pit Design:US$1,150/oz and a US$:R$ exchange rate of 1:3.2. Reserves reported within the Pit Design: US$1,200/oz and a US$:R$ exchange rate of 1:3.25.

The Mineral Reserves inside the engineered pit designs for the remaining open pits were estimated using COGs estimated for each deposit and ore type based on the following gold prices; Open Pit Optimization Shell used for Pit Design: US$1,100/oz and a US$:R$ exchange rate of 1:3.4. Reserves reported within the Pit Design: US$1,200/oz and a US$:R$ exchange rate of 1:3.25.

The Mineral Reserves are contained within four open pit deposits, one underground zone, and surface stockpiles. Proved and Probable Mineral Reserves for the mine as of 30 June 2017 are estimated to be 24.5 Mt at 1.79 g/t Au containing approximately 1.41 Moz of gold. The Mineral Reserve estimate for Tucano as of 30 June 2017 is summarized in Table 15.6.

Open Pit Mineral Reserves represent approximately 74% of the total tonnes and 71% of the total ounces of contained gold. Current open pit mine life is seven years (2018 to 2024). Underground Mineral Reserves represent approximately 12% of the total tonnes and 23% of the total ounces of contained gold. Current underground mine life is projected to be seven years (2018 to 2024).

Minesure is of the opinion that the 30 June 2017 Mineral Reserves relating to the open pits have been estimated in a manner consistent with industry practices and NI 43-101 guidelines.

AMC is of the opinion that the 30 June 2017 Mineral Reserves relating to the underground have been estimated in a manner consistent with industry practices and NI 43-101 guidelines.

There are two principal locations of the Tucano Deposit namely, Tap AB and Urucum. There are two smaller locations being Urucum East and Tap C. Tap AB is located 1.2 km east of the processing plant and infrastructure and comprises three conjoined deposits being AB1, AB2, and AB3. Urucum centroid is located 5 km north-north east of the processing plant and comprises four conjoined deposits, Urucum South, Urucum Central South, Urucum Central North, and Urucum North. Urucum East is located 1.4 km north east of Urucum North pit. Tap C is located 2.8 km north-north east of the processing plant.

The locations of the deposits are shown in Figure 15.1. 

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Figure 15.1 Deposit location plan

Source: N. Spicer (Minesure Pty Ltd).

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Open cut mining has previously been undertaken on all deposits except Urucum East.

Mineral Reserves for the Underground were estimated by AMC in a pre-feasibility study completed in 2016. The study estimated the cash flow of establishing and operating an underground operation at Urucum North using mechanized long-hole stoping techniques. The recovery of Mineral Reserves would sustain the underground mining for up to eight years and realize 2.4 Mt of ore grading 3.6 g/t Au and containing 278,000 oz of gold.

To date there has been no underground mining production.

The proportion of Reserves by deposit is shown in Table 15.1.

Table 15.1 Proportion of reserves by deposit

Deposit	% of total
	Tonnes	Au (oz)
Tap AB	26%	29%
Urucum	42%	39%
Tap C	5%	3%
Urucum East	1%	1%
Urucum Underground	12%	23%
Stockpiles	14%	5%

Test work and processing results indicate that the Mineral Reserves are all amenable to processing using Carbon in Leach (CIL) technology.

15.2 Open pit

15.2.1 Reserve block model

The mining engineering work related to the pit optimizations and engineered pit designs was carried out using the block models prepared by B Wolfe in January 2017 for the Urucum deposit and in June 2017 for the Tap AB deposit. The block models prepared by Marcello Batelochi and Paul Tan in February 2017 were used for the Tap C and Urucum East deposits. A parent block size of 9 m x 20 m x 8 m (Y, X, Z) metres was used for the Tap AB and Urucum deposits, and 20 m x 8 m x 4 m (Y, X, Z) metres was used for the Tap C deposit, and 10 x 5 x 2 (Y, X, Z) metres was used for the Urucum East deposit. The models contain blocks coded with the following information:

• Gold grade

• Block proportion of ore

• Resource category (Measured, Indicated, and Inferred)

• Density

• Oxidation factor

• Lithology and domain

• Material volumes.

Life of mine (LOM) final designs have been compiled for the open pits and these were the basis of estimating the Mineral Reserves for the Tucano project. Intermediate or annual pit designs have not been compiled.

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15.2.2 Open pit optimization

The open pit optimizations were carried out for all deposits using the Lerchs-Grossman 3D algorithm in the Whittle 4D Software. Using mining costs, processing costs, gold recovery values and an overall pit slope, the pit optimizer determines an ultimate pit shell that delineates the volume of extraction that maximizes value.

A series of pit optimizations were produced using a range of gold selling prices in order to produce an industry standard pit-by-pit graph. This process is used to evaluate the sensitivity of the pit optimizations to changes in mineral selling prices, as well as to evaluate the effect of the pit size and stripping ratios on the project Net Present Value (NPV). The optimization process produces a series of nested pit shells that prioritize the mining of the most economic material. Less profitable material (lower grade and / or high strip ratio) is by definition only mined in later pit shells as the input mineral selling price is increased.

From these results, appropriate pit shells for the deposits were selected as a basis for the engineered pit designs and Mineral Reserve estimates. All pit optimizations were run using reasonable and relevant economic, cost, recovery, and pit slope assumptions, and were run on diluted gold grades. In accordance with the guidelines of NI 43-101 and CIM Standards, only resource blocks classified as either Measured or Indicated were allowed to drive the pit optimizer for Mineral Reserve reporting purposes.

15.2.3 Dilution and extraction

Resources for Tap AB and Urucum were estimated using MIK which include a proportion of dilution when estimating the resource block grades. MIK estimation techniques have only been in use at Tucano for a limited amount of time so no real trends can be discerned from the reconciliation data. Based on the limited reconciliation data available it was decided to increase the proportion of included dilution. Accordingly, Resources were diluted by an additional 15% in the oxide and 20% in the fresh.

Mining recovery (extraction) varied by deposit and ore type. It was found that the quantity of ore mined from Urucum was consistent with model predictions for both oxide and fresh ores. Reconciliation data indicated that the quantity of ore mined from Tap AB was generally lower than that indicated by the model for both oxide and fresh ores.

No mining has yet taken place at Urucum East so it was not possible to derive dilution and recovery factors from reconciliation data. The factors used for this deposit were derived from using the Tap AB oxide ore factors and Urucum fresh ore factors.

15.2.4 Gold grade data

The gold grade data (undiluted) was provided in the mineral resource block model and is the main item used for pit optimization and Mineral Reserve estimates. The grades were modified to account for mining dilution and the block volumes were modified to allow for dilution and mining recovery.

15.2.5 Cost parameters for pit optimization

The key pit optimization parameters used to derive the economic pit shells for the deposits are summarized in Table 15.2. The optimizations were based on parameters and cost data projected for the 2017 Budget.

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Table 15.2 Optimization cost parameters

Item	Unit	Tap AB		Urucum
		Oxide	Fresh	Oxide	Fresh
		Value	Value	Value	Value
Gold price	$/oz	1,150	1,150	1,100	1,100
	$/gm	36.97	36.97	35.37	35.37
Optimization selling price	$/gm	36.23	36.23	34.66	34.66
Metallurgical recovery for Au	%	93	93	95	93
Gold value	$/gm	34.39	34.39	33.60	32.89
State royalty	%	2.0	2.0	2.0	2.0
Realized value	$/gm	33.70	33.70	32.93	32.23
Treatment cost	$/t ore	10.33	13.96	9.62	13.74
Crusher feed	$/t ore	0.38	0.38	0.36	0.36
Admin cost	$/t ore	2.08	2.08	1.88	1.88
Corp overheads	$/t ore	0.00	0.00	0.00	0.00
Technical services	$/t ore	0.80	0.80	0.74	0.74
Dewatering	$/t ore	0.98	0.98	0.56	0.56
Grade control	$/t ore	1.12	1.12	1.24	1.24
Mining contractor fixed	$/t ore	3.58	3.58	4.38	4.38
Ore haulage	$/t ore	0.00	0.00	2.34	2.34
Total processing cost	$/t ore	19.27	22.90	21.11	25.23
Mining dilution	%	15	20	15	20
Mining recovery	%	90	90	100	100

Mining costs were based on the Mining Contract rates escalated in accordance with the Rise and Fall clause of the mining contract.

15.2.6 Pit optimization mill recovery

Test work and operations has indicated that a consistent metallurgical recovery of 93% is projected for the oxide and fresh ore for Tap AB. Metallurgical recovery for Urucum is estimated to be 95% for the oxide and 93% for the fresh ore following the planned plant upgrade.

15.2.7 Tap AB optimization results

A series of pit shells for Tap AB were run using gold selling prices ranging from US$550/oz to US$1,650/oz at US$ / R$ exchange rate of 1:3.2 and using the other parameters listed in the sections above. The results of the pit optimization are presented in Figure 15.2.

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Figure 15.2 Tap AB optimization results

Source: N. Spicer (Minesure Pty Ltd).

Shell 21 at the gold price of US$1,150/oz generated the maximum average discounted cash flow and was chosen as the basis for the Tap AB final design.

15.2.8 Urucum optimization results

A series of pit shells for Urucum were run using gold selling prices ranging from US$550/oz to US$1,650/oz at US$:R$ exchange rate of 1:3.4 and using the other parameters listed in the sections above. The results of the pit optimization are presented in Figure 15.3.

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Figure 15.3 Urucum optimization results

Source: N. Spicer (Minesure Pty Ltd).

Shell 21 at the gold price of US$1,100/oz generated the maximum average discounted cash flow and was chosen as the basis for the Urucum final design.

15.2.9 Open pit COG - Pit optimization and Reserves

COGs used in the open pit optimization were estimated for each deposit based on the following criteria:

• The gold price - US$1,150/oz (Tap AB) and US$1,100 (Tap C and Urucum).

• The applicable mine royalty payments.

• Metallurgical Recoveries.

• Specific ore mining and hauling costs.

• The process operating costs and on-site (and off-site) metal recoveries by material type, applicable or selected process method, and orebody.

Gold price assumptions are dictated by the corporate office. Royalties are a percentage due to the local and federal governing bodies.

A set of process and overhead costs for processing oxide and fresh ore were estimated by the process manager based on current operating costs and projected operating costs. Costs used to estimate COGs do not include mining costs, unless there is a marginal cost for ore compared to waste in a given pit. COGs consider the general and administration (G&A) costs as part of the process cost. Estimation of COGs for the various pits is shown in Table 15.3.

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Table 15.3 COG estimation parameters using pit shell optimization gold price

Deposit	Ore type	Processing cost ($/t)	G&A cost ($/t)	Tech services $/t	Haulage ($/t)	Royalty (%)	Met Rec (%)	Dilution (%)	Mining Rec (%)	COG Au (g/t)
Tap AB	Oxide	10.71	2.08	0.80	0	2.0	93	15	90%	0.66
	Fresh	14.34	2.08	0.80	0	2.0	93	20	90	0.82
Urucum	Oxide	9.98	1.88	0.74	2.34	2.0	95	15	100	0.74
	Fresh	14.10	1.88	0.74	2.34	2.0	93	20	100	0.94
Tap C	Oxide	9.98	1.88	0.74	2.34	2.0	95	15	90	0.66
	Fresh	14.10	1.88	0.74	2.34	2.0	93	20	90	0.86
Urucum East	Oxide	9.98	1.88	0.74	2.34	2.0	95	15	90	0.74
	Fresh	14.10	1.88	0.74	2.34	2.0	93	20	100	0.94

G&A = General and Administration, Tech = Technical, Met Rec = Metallurgical recovery, Mining Rec = Mining recovery

In November 2017, the gold price for the reporting of Mineral Reserves was increased to US$1,200/oz and the COG for each pit was recalculated based on this price, the optimization process costs and the revised haulage costs. A metallurgical recovery of 93% was applied for both oxide and sulphide. These COGs were used to define the Mineral Reserve COGs for reporting within the reserve pits designed with the lower gold price optimized shells (Table 15.2). Table 15.4 below shows the COG estimation parameters using a gold price of US$1,200/oz.

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Table 15.4 COG estimation parameters using the US$1,200 gold price

Deposit	Ore type	Processing cost ($/t)	G&A cost ($/t)	Tech services ($/t)	Haulage ($/t)	Royalty (%)	Met Rec (%)	Dilution (%)	Mining Rec (%)	Calculated COG Au (g/t)	COG Au Reserve (g/t)
Tap AB	Oxide	10.57	2.05	0.79	0.00	2	93	15	90	0.62	0.6
	Fresh	14.11	2.05	0.79	0.00	2	93	20	90	0.77	0.7
Urucum	Oxide	10.57	2.05	0.79	2.45	2	93	15	100	0.70	0.7
	Fresh	14.11	2.05	0.79	2.45	2	93	20	100	0.85	0.8
Tap C	Oxide	10.57	2.05	0.79	1.40	2	93	15	90	0.67	0.6
	Fresh	14.11	2.05	0.79	1.40	2	93	20	90	0.82	0.8
Urucum East	Oxide	10.57	2.05	0.79	2.69	2	93	15	90	0.69	0.7
	Fresh	14.11	2.05	0.79	2.69	2	93	20	90	0.86	0.7

G&A = General and Administration, Tech = Technical, Met Rec = Metallurgical recovery, Mining Rec = Mining recovery

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15.3 Underground mining

15.3.1 Underground mine Mineral Reserve estimate

The underground Mineral Reserves were developed from the Mineral Resource block model (reference: urucum_ug_resource2015_v17.mdl) which was provided to AMC by BDR.

The estimation of Mineral Resources that form the basis of the Mineral Reserve estimates is described in Section 14 of this report. Mr Gary Methven, P.Eng., takes responsibility for the estimation of Urucum North Underground Mineral Reserves.

Underground Mineral Reserves have been estimated based on:

• COG criteria indicated in Section 15.3.

• Dilution and mining recovery estimates as discussed in Section 15.4 and Section 15.5.

• In accordance with NI 43-101 standards, no Inferred material has been included in this Mineral Reserve estimate.

The estimated Mineral Reserve is based only on Measured and Indicated Mineral Resource and is effective as of 30 June 2017 and is provided in Table 15.5. AMC notes that although there is approximately 10% of the Mineral Resource estimated as a Measured resource, given the accuracy of the PFS it is reasonable to convert Measured Resources to Probable Reserves due to uncertainties with the modifying factors.

Table 15.5 Urucum North Underground Mineral Reserve estimate

Category	Ore tonnes (x 1,000)	Grade of contained metal Au (g/t)
Proven
Probable	2,378	3.64
Total	2,378	3.64

Notes:

• 2018 Mineral Reserves are as of 30 June 2017 and based on a COG of 1.6 g/t Au. The COG is based on an incremental mining and processing cost of US$50/t, metallurgical recovery of 90%, a gold price of US$1,120/oz and an exchange rate of US$:R$ of 1:3.8.

• A 0.5 m dilution skin was assumed for the HW and FW side of the production stopes making the total dilution 1.0 m in width. Ore losses were assumed to be 8% for the Benching method, 14% for the Up-Hole retreat method, and ore losses for the recovery of the sill pillars were assumed to be 35%.

15.3.2 Orebody geometry

The Urucum North Underground mineralization covers a strike length of approximately 800 m down to a depth of approximately 500 m below the current open pit design (BDR pit design of 11 March 2016). Mineralization is hosted within the banded iron formation (BIF) unit which is approximately 100 m wide.

Continuous high grade, shallow plunging ore lodes exist along the mineralized BIF. BDR have modelled 3 main ore lodes, namely, Lode 1, Lode 2, and Lode 300, each lode dipping sub-vertically and generally separated by 20 m to 30 m.

Since the earliest days of exploration and mining at Tucano, it has been recognized that a number of ore shoots exist within the mineralized lodes. These shoots plunge in two main orientations, namely:

• Shallow dipping ore shoots plunging 15˚ to 25˚ towards the north.

• Steeply dipping ore shoots plunging 50˚ to 60˚ towards the north.

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Ore shoots are defined as contiguous regions where drill intersections meet specific grade and metal content criteria.

15.3.3 Cut-off grade (COG)

An incremental COG of 1.6 g/t Au has been chosen as the basis for estimating Mineral Reserves. The basis for this COG is:

• Optimization studies undertaken as part of August 2014 Scoping study based on evaluation of COGs from 1.0 g/t Au through to 3.0 g/t Au, in 0.5 g/t Au increments.

• An incremental mining and processing cost of US$50/t.

• Metallurgical recovery of gold of 90%.

• A gold price of US$1,120/oz equating to US$36/g.

The incremental costs do not include any development costs or contractor equipment leasing costs. AMC has undertaken sufficient design, scheduling and costing work to test the effect of increasing the COG above 1.6 g/t Au. This work indicates that the low-grade stopes (between 1.6 g/t Au and 2.0 g/t Au added value to the overall project economics, partially because of already expensed “fixed costs” and already costed development that is required to access higher grade stopes.

15.3.4 Dilution and mining recovery estimates

There are two main sources of dilution in narrow vein stopes:

1 Planned dilution: this is the dilution required to achieve the designed stope shape. Designed dilution can result from waste included:

— To achieve minimum mining width.

— To achieve a viable mining shape.

2 Unplanned dilution: this is dilution that is outside of the designed stope shape. Depending on the mining method, it may include both overbreak and floor dilution.

Overbreak is typically a result of blasting practices and geotechnical conditions. Floor dilution is the result of mucking waste rock from the rock fill floor.

AMC has applied a planned dilution of 0.50 m to each of the hangingwall and the footwall with a minimum stope width of 2 m. The stope shape is then evaluated against the Mineral Resource model. After the Mineable Shape Optimizer (MSO) wireframes were generated, the following ore losses have been applied to the stope tonnes:

• Up-hole Retreat mining areas (mining in a down-dip direction)

— 6 m rib pillar every 50 m equivalent to 12% ore loss.

— 2% ore loss for difficulty to muck ore at the bottom of stopes.

— Total ore loss = 14%.

• Bench stoping areas (mining in an up-dip direction)

— 3 m rib pillar every 50 m equivalent to 6% ore loss.

— 2% ore loss for difficulty to muck ore at the bottom of stopes.

— Total ore loss = 8%.

• Sill pillar recovery

— Ore loss = 35% due to mining below unconsolidated rockfill.

• Ore development recovery

— Recovery of development ore has been assumed to be 100% (Note: development dilution is captured in the development wireframes when assessed by Mine24D software).

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15.3.5 Mining method criteria and modifying factors

Up-hole Retreat mining parameters include:

• 20 m level interval (15 m - 20 m production holes).

• Minimum mining width of 2.0 m (plus dilution).

• 6 m rib pillars every 50 m along strike.

• Generally, 3 x 20 m lifts between sill pillars, equating to an ore block height of 60 m.

• Mining proceeding in a top down sequence.

• Ore dilution based on 0.5 m of FW and 0.5 m of HW dilution.

• Mining recovery of 86% assumed.

Down-Hole Benching mining parameters include:

• 20 m level interval (15 m - 20 m production holes).

• Minimum mining width of 2.0 m (plus dilution).

• 3 m rib pillar every 50 m along strike.

• Generally, 3 x 20 m lifts between sill pillars, equating to an ore block height of 60 m.

• Mining proceeding in a bottom up sequence.

• Ore dilution based on 0.5 m of FW and 0.5 m of HW dilution.

• Mining recovery of 92% assumed.

Sill Pillar mining parameters include:

• 20 m level interval (15 m - 20 m production holes).

• Minimum mining width of 2.0 m (plus dilution).

• Continuous along strike.

• Recovered using up-holes drilled from below the sill.

• Dilution based on 0.5 m of FW and 0.5 m of HW dilution.

• Mining recovery of 65% assumed.

15.4 Combined Mineral Reserve estimate

Open Pit Proved and Probable Mineral Reserves were estimated using the engineered pit designs, the Mineral Resource block models, and modifying factors. The Urucum underground Mineral Reserve was restated by AMC based on the PFS study completed in 2016 and in line with NI 43-101 guidelines with an effective date of 30 June 2017. The Mineral Reserves have been estimated in accordance with Canadian NI 43-101 Standards of Disclosure for Mineral Projects of August 2011 and Definition Standards for Mineral Resources and Mineral Reserves adopted by the CIM Council on May 2014. The estimated Mineral Reserves for the Tucano project as of 30 June 2017 are shown in Table 15.6.

Exchange rates of US$ to the Brazilian Real (R$) used for the estimation of Mineral Reserves are discussed below Table 1.3.

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Table 15.6 Total Mineral Reserves - 30 June 2017

Brazil	Cut-off	Proved Reserve			Probable Reserve			Total Mineral Reserve
	Au (g/t)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)	Tonnes (‘000)	Grade Au (g/t)	Ounces (‘000)
Urucum Open Pit oxide	0.7	374	1.21	15	180	1.18	7	554	1.20	21
Urucum East Open Pit oxide	0.7	0	0.00	0	151	1.71	8	151	1.71	8
Tap AB Open Pit oxide	0.6	2,384	2.16	166	2,024	1.76	115	4,408	1.98	280
Tap C Open Pit oxide	0.6	362	1.08	13	191	0.99	6	553	1.05	19
Total oxide open pit		3,120	1.92	193	2,545	1.66	136	5,666	1.80	329
Urucum Open Pit primary	0.8	4,446	1.69	241	5,246	1.81	306	9,692	1.76	547
Urucum East Open Pit primary	0.7	0	0.00	0	16	1.50	1	16	1.50	1
Tap AB Open Pit primary	0.8	1,274	2.00	82	1,188	1.72	66	2,462	1.86	147
Tap C Open Pit primary	0.8	229	1.31	10	387	1.44	18	615	1.39	28
Total primary open pit		5,949	1.74	333	6,836	1.77	390	12,785	1.76	723
Urucum Open Pit total		4,820	1.65	256	5,425	1.79	312	10,245	1.72	568
Urucum East Open Pit total		0	0	0	167	1.69	9	167	1.69	9
Tap AB Open Pit total		3,658	2.10	247	3,212	1.75	181	6,870	1.94	428
Tap C Open Pit total		591	1.17	22	577	1.29	24	1,168	1.23	46
Total oxide and primary open pit		9,069	1.80	525	9,381	1.74	526	18,450	1.77	1,051
Open pit stockpile	0.5	1,985	0.66	42	0	0.00	0	1,985	0.66	42
Spent ore stockpile	0.5	1,311	0.61	26	0	0.00	0	1,311	0.61	26
ROM expansion stockpile	0.5	345	0.80	9	0	0.00	0	345	0.80	9
Total stockpiles	0.5	3,642	0.66	77	0	0.00	0	3,642	0.66	77
Total Tucano Open Pit & stockpiles		12,711	1.47	603	9,381	1.74	526	22,092	1.59	1,128
Urucum Underground primary		0	0.00	0	2,378	3.64	278	2,378	3.64	278
Total Underground primary		0	0.00	0	2,378	3.64	278	2,378	3.64	278
Grand total Tucano		12,711	1.47	603	11,759	2.12	804	24,470	1.79	1,406

Notes:

• 2018 Open Pit Mineral Reserves are stated as of 30 June 2017 based on the COGs as detailed in Table 1.3 above. Reserves are reported within open pit designs based on pit optimized shells using the following optimization COGs; Tap AB Oxide: 0.66g/t Au, Tap AB Fresh: 0.82 g/t Au, Urucum Oxide 0.74 g/t Au, Urucum Fresh 0.94 g/t Au, Tap C Oxide 0.66 g/t Au, Tap C primary 0.86 g/t Au, Urucum East Oxide 0.74 g/t Au, Urucum East Fresh 0.94 g/t Au. The COG is based on total costs per pit (inclusive of processing, G&A, tech services and haulage, exclusive of mining costs) and averages US$12.8 for oxide and US$16.8 for primary. Metallurgical recoveries are 95% for oxide and 93% for sulphides. The gold price used for the open pit optimizations was $1,150/oz for Tap AB and $1,100 for Tap C and Urucum. The gold price used to calculate the COGs for the reserves reported within the optimized pit designs was US$1,200 per ounce. An exchange rate of US$:R$ of 1:3.2 for Tap AB and 1:3.4 for Urucum was used. Open Pit Mining Recoveries of 100% for oxide and primary were used at Urucum, 90% for oxide and primary at Tap C and Tap AB and 90% oxide, 100% primary at Urucum East. Dilution used for all pits was 15% for oxide and 20% for primary.

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• 2018 Underground Mineral Reserves are stated as of 30 June 2017 and based on a COG of 1.6 g/t Au. The COG is based on an incremental mining and processing cost of US$50/t, metallurgical recovery of 90%, a gold price of US$1,120/oz and an exchange rate of US$:R$ of 1:3.8.

• A 0.5 m dilution skin was assumed for the HW and FW side of the underground production stopes making the total dilution 1.0 m in width. Ore losses were assumed to be 8% for the Benching method, 14% for the Up-Hole retreat method, and ore losses for the recovery of the sill pillars were assumed to be 35%.

• The JORC reserve released on 19 December 2017 to the ASX ("Interim Mineral Resource and Ore Reserve Update as at 30 June 2017") stated the total Urucum Underground Reserve was 2,972 kt at 3.61 g/t Au for 345 koz Au. This reserve was reported by AMC inclusive of Inferred resource blocks captured inside the stope designs that were optimized on Measured and Indicated Resources in the 2016 PFS. The Inferred Resources falling within the stope designs was removed from the calculations and reporting of the NI 43-101 compliant Urucum Underground Reserves by AMC Vancouver in 2018 and has been quoted in the Table 1.3 above.

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15.5 Factors that may impact Mineral Reserve estimates

The main factors that may impact the mineral reserve estimates are as follows:

• Metal prices and exchange rate assumptions.

• Mining, process, and operating cost or recovery assumptions.

• Ability to obtain permits in a timely manner.

• Social license to operate being maintained.

The QP is not aware of any other modifying factors that could materially impact the mineral reserves of the Tucano operation.

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16 Mining methods

16.1 Open pit mining

Open pit mining is currently being undertaken by new mine contractor Unienge & Modulo (U&M), the largest mining and locally experienced contractor in Brazil. The mining operations are based on the use of hydraulic excavators and a haul truck fleet engaged in conventional open pit mining techniques. Details of the pit slope design parameters based on independent review work is provided in Section 16.2.1.

Excavated material is loaded to trucks and hauled to either the ROM, long term low grade or marginal stockpiles, the iron ore stockpile, or the waste dump. Ore excavation and haulage is monitored by quality control personnel employed by the Geology department and details of material movement are recorded by a radio dispatch system “Tecmine”. Oxide material is free dig with transitional material lightly blasted to loosen it for digging. Fresh rock is typically blasted on 4 m benches however 8 m benches are also fired. For wide sulphide ore lodes, the ore can sometimes be blasted separately if the shot is sufficiently large to minimize dilution. Normally ore and waste are fired together requiring the use of 8 m long poly pipes or Blast Vector Indicators (BVI) at each node of the ore polygon to determine post-blast movement. Ore sent to the ROM is stockpiled on fingers of oxide or sulphide within various grade cut-off categories to facilitate blending to achieve the desired oxide / sulphide proportions and grade.

16.2 Geotechnical assumptions

16.2.1 Overall pit slopes

SRK and Peter O’Bryan and Associates made recommendations regarding the wall design parameters for the Tucano Project Pits. These are shown in Table 16.1.

Table 16.1 Recommended wall design parameters

Pit	Domain (above interface)	Bench height (m)	Berm width (m)	Batter face angle	Bench stack height (m)	Inter-ramp angle	Geotechnical Berm width (m)
Tap C	Friable (W3)	4	3	70°	80	41.9°	10
	Unweathered	20	8	75°	120	56.3°	n/a
Tap AB	Friable (R2)	4	3	70°	60	41.9°	40
	Unweathered	20	8	75°	120	56.3°	n/a
Urucum	Friable (R2)	4	3	70°	60	41.9°	40
	Unweathered	20	8	75°	120	56.3°	20

Additionally, Geotechnical Consultant Luis Navarro has recommended (Site Visit Report October 2015) that walls in weathered pegmatite have batter angles reduced to 60°. This particularly applies to Urucum Central South pit and potentially Urucum Central North Pit which has massive pegmatite running parallel to the strike of the orebody in the final west wall position.

The current Urucum Central South pit is designed to 158 m below the surface and should require a 20 m wide geotechnical berm at approximately 150 m below surface however the benches at the bottom 20 m of the pit are essentially narrow and short, so a geotechnical berm is not required. Following a review of the AB pits in March 2017 Peter O'Bryan advised that the following regimes for geotechnical berms could be adopted:

• Slopes up to 120 m (= two 60 m stacks): One geotechnical berm at 20 m width.

• Slopes up to 180 m (= three 60 m stacks): Two geotechnical berms at 30 m width.

• Slopes 240 m (three or more 60 m stacks): Three geotechnical berms at 40 m width.

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• The heights are measured above the oxide / fresh interface.

Overall slope angles for the existing pits were measured from the 2016 final designs and these are shown in Table 16.2. The angles included berms and access roads where intersected.

Table 16.2 Measured overall slope angles (degrees)

Pit	Total depth (m)	Quadrant
		North	East	South	West
URN	160	28	47	35	45
URCN	150	21	49	20	49
URCS	150	33	48	44	49
URS	115	45	52	48	55
UR_East	96	30	38	28	17
AB3	160	45	37	41	46
AB1/2	270	22	26	27	38

URN = Urucum north, URCN = Urucum Central North, URCS = Urucum Central South, UR_East = Urucum East.

The majority of the measured slope angles are lower than the recommended FS Inter ramp slope (IRS) angles. This indicates that access requirements are greater than estimated in the FS.

For most of the deposits the North and South slope angles are dictated by the access ramp and in practice tend to be lower than the recommended slope angles. For the optimization the recommended IRS angles were used for North and South walls of all pits.

Different slope characteristics were estimated for oxide and fresh rocks because of the difference in competence. Fresh rocks are by nature stronger and able to remain stable at steeper slope angles whereas the oxidized rocks tend to crumble and fail at too steep an angle.

16.2.1.1 Urucum

At Urucum the east and west walls are of similar heights. The depth of weathering is between 20 m and 30 m below surface. The fresh rock is very competent except where there are extensive pegmatite intrusions. Pegmatite affects all of the Urucum pit walls. It was proposed that the oxide zone in the Urucum Pits use IRS angles of 40°. The exception is Urucum South which is basically a circular pit with a spiralling ramp. The current design has a 14 m wide ramp which is only suitable for two-way operation for Volvo trucks. It was proposed to use 42° on all oxide walls.

IRS angles for the fresh zones were measured from the 2016 designs and are listed in Table 16.3.

16.2.1.2 Urucum East

Urucum East is a small stand-alone deposit located 700 m east of the Urucum north zone. The deposit is wholly oxide. Due to the small size of the deposit the overall slopes are flattened by the access road and also by the shallow footwall on the north and west side. An existing design has overall slope angles of 30° in the north quadrant which has a 96 m high wall incorporating a 40 m wide geotechnical berm at the 253 mRL. Slope angles for the remaining quadrants are shown in Table 16.3.

16.2.1.3 Tap AB

Tap AB3 has been mined down to the 81 mRL. The east wall and west walls are mostly fresh from the 129 mRL down however the middle of the zone is oxidized down to the 17 mRL. The north wall is all oxide and the south wall has fresh rock occurring at the 140 mRL. The south wall has an overall slope angle of 41° in the oxide and 40° in the fresh. The north wall final design has an overall slope angle of 45° which includes an eight metres wide berm on the 117 mRL and a narrow access ramp. The east wall has a 40 m wide geotechnical berm at the 117 mRL and a 20 m wide access ramp which flattens to wall to 37°. It is proposed that the IRS in oxide be 33° in the North, 31° in the East, 30° in the South quadrant and 38° in the West quadrant. The overall slope angle for the fresh rock is estimated to be 47° in the East and South quadrants and 56° for the remainder.

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Tap AB1 and Tap AB2 join up as one pit in the current final design. The eastern wall is oxide down to the 33 mRL. The measured overall slope angle is 26°. The eastern wall design had four stack berms of 40 m width at 60 m vertical intervals, hence the low overall slope angle. The 2016 design was based on an outdated topography which has subsequently been re-compiled. The new topography showed that one of the stack berms in the east wall could be eliminated leading to a steeper overall slope angle of 33° in the oxide rock. The west wall is much shallower and does not require stack berms consequently the overall slope angle in oxide is 38°.

It is proposed to use 33° slope angle in oxide rock for the north, 30° for the east, 31° for the south walls, and 38° in the west wall. The overall slope angle for the fresh rock is estimated to be 56° in the north, 47° in the east, 47° in the south, and 56°for the west wall.

16.2.1.4 Pegmatite

The recommended design parameters in oxidized pegmatite are:

• Batter height 4 m

• Batter angle 60°

• Berm width 3 m

The average slope angle becomes 35° and this will be used throughout. Where possible ramps will not be mined in the oxidized pegmatite because of its inherent instability.

The current south-east wall of Tap AB1 has a highly oxidized pegmatite abutting carbonate and iron stone lithologies which are very friable. This material is unstable and readily slumps. It is particularly important that this area be mined at a shallow angle.

16.2.1.5 Proposed overall slope angles

The proposed overall slope angles are summarized in Table 16.3.

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Table 16.3 Proposed overall slope angles (degrees)

Pit	Total depth (m)	Rock type	Quadrant
			North	East	South	West
URN	160	Oxide	40	40	40	40
		Fresh	45	45	45	51
URCN	150	Oxide	40	40	40	40
		Fresh	55	55	55	55
URCS	150	Oxide	40	40	40	40
		Fresh	50	49	50	52
URS	115	Oxide	42	42	42	42
		Fresh	50	51	50	49
UR_East	96	Oxide	30	38	30	30
AB1/2/3	270	Oxide	33	30	31	38
		Fresh	56	47	47	56

URN = Urucum north, URCN = Urucum Central North, URCS = Urucum Central South, UR_East = Urucum East.

16.2.2 Engineered pit designs

The engineered pit designs were completed using the pit optimization shells as a guide in order to maximize the value and gold recovered inside the ultimate pits. The resulting pit designs include practical geometry that is required in an operational mine, such as the haul road to access all the benches, recommended pit slopes with geotechnical berms, proper benching configuration, and smoothed pit walls.

The resulting engineered pit designs were used to estimate the Mineral Reserves in this report.

16.2.2.1 Tap AB

At the end of June 2017 mining was active only in Tap AB1 and Tap AB2. Tap AB pits will be the main source of plant feed for up to 18 months because the processing plant in its current configuration can only process a proportion of the Urucum ore which is mainly primary or fresh. Modifications to the processing plant are scheduled to be completed by November 2018 to enable the primary ores to be treated effectively.

Due to overlapping, AB1 and AB2 pit designs have been combined into a single pit. The long axis length of the AB deposit is approximately 1,700 m, and the minor axis width ranges from 400 m in the AB3 pit to 950 m in the AB1/2 pit. The designed pit bottom elevation of AB1/2 is -47 mRL and the designed pit bottom for AB3 is 5 mRL.

The primary pit exit for AB1/2 is at 115 mRL and for AB3 at 160 mRL.

Access ramps are designed at eight (8) metre width for the bottom 20 m vertical, then 14 m width for the next 60 m Vertical and then 23 m width for the remainder. Ramp gradients for the 8 m wide and 14 m wide sections will be 12.5% (1 in 8) which is suitable for 40 t capacity all-wheel drive (AWD) trucks. The 23 m wide ramp will be mined at 10% (1 in 10) gradient which is suitable for Caterpillar 777 rigid frame trucks.

The final design for Tap AB1/2 is shown in Figure 16.1 and the final design for Tap AB3 is shown in Figure 16.2.

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The oxide rock is mostly free-dig whilst all fresh rock mined material is drilled and blasted. All ore is hauled to the Run of Mine stockpile where it is re-handled into the jaw crusher by a front-end loader.

Figure 16.1 Tap AB1/2 final design

Source: N. Spicer (Minesure Pty Ltd).

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Figure 16.2 Tap AB3 final design

Source: N. Spicer (Minesure Pty Ltd).

16.2.2.2 Urucum

Past mining of the Urucum pits has recovered most of the oxide ore which has been trucked to the processing plant and processed. Mining of the Urucum sub-pits has been suspended until the processing plant has been modified to process higher proportions of primary ore. The primary ore is amenable to CIL processing however recovery is grind size sensitive so additional grinding capacity will be installed to enable throughput and recovery to be maintained at current levels.

The proposed pit final design is shown in Figure 16.3.

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Figure 16.3 Urucum final design

Source: N. Spicer (Minesure Pty Ltd).

The total strike length (N-S) is approximately 2,200 m and the average cross-strike width is approximately 300 m. Relevant design data for four sub-pits is shown in Table 16.4.

Table 16.4 Urucum pits design data

Pit	Length (m)	Width (m)	Pit base (mRL)	Pit exit (mRL)	Pit depth (m)
South	330	230	15	125	110
Central South	580	350	-29	130	159
Central North	500	290	59	179	120
North	880	300	31	151	120

The access ramp is designed to be eight metres wide with a gradient of 12.5% (1 in 8) for the bottom 36 m. It is then designed to be 14 m wide with a gradient of 12.5% for the next 24 m vertical. Thereafter it is designed to be 23 m wide with a gradient of 10% (1 in 10).

Ore is stockpiled close to the pits and then re-handled to the ROM pad for re-handling into the crusher.

16.2.2.3 Urucum East

At the end of June 2017, no mining had taken place at the Urucum East deposit. The proposed final pit design is 300 m along strike and 180 m across strike. The design mines out the side of a hill so that the design has a depth of 40 m to the south and 190 m to the north. The access ramp is designed to be 14 m wide at a gradient of 12.5%. Ore is planned to be stockpiled close to the pits and then re-handled to the ROM pad for re-handling into the crusher.

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The proposed final pit design is shown in Figure 16.4.

Figure 16.4 Urucum East final design

Source: N. Spicer (Minesure Pty Ltd).

16.2.2.4 Tap C

The deposit has been mined previously and the proposed final pit design is shown in Figure 16.5.

The pit design extends over a strike length of 2 km with an approximate average width of 170 m and comprises a series of inter-connected pits. The access ramp varies in width but is generally 10 m wide at a gradient of 10%. Ore is usually hauled directly to the ROM pad for re-handling into the crusher.

Mining for all deposits is accomplished with a typical drill, blast, load, and haul cycle, using conventional truck-and hydraulic excavator equipment. Mining is undertaken by a mining contractor under the direction of BDR staff. The truck fleet comprises a mixture of AWD and rigid frame trucks. The rigid frame trucks are efficient at hauling large volumes of material and are mostly used on waste stripping. The AWD trucks are narrower than the rigid frame trucks so are mostly used to haul from the bottom of the pits up steeper ramps. This reduces the quantity of waste needed to be mined. Hydraulic excavators are used in backhoe configuration and range in size from 60 t up to 240 t. All equipment is diesel powered.

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Figure 16.5 Tap C final design

Source: N. Spicer (Minesure Pty Ltd).

16.3 Contract mining

The main mine production equipment that U&M have selected for Tucano is outlined in Table 16.5 along with equipment which is now owned by the mine. Additional mining equipment support is provided by JS and Vale do Amazonas.

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Table 16.5 Tucano mining fleet

Excavators	Number	Contactor / Beadell
Liebherr 9250	1	Beadell
Liebherr 9150	1	Beadell
VOLVO 700	1	JS Máquinas
VOLVO 700	1	JS Máquinas
HYUNDAI 220 BL Hyundai 220 Hitachi (2500) Hitachi (1200)	1 4 3 4	Vale do Amazonas U&M U&M U&M
Total excavators	16
Trucks
CAT 777	3	Beadell
CAT 777	20	U&M
Komatsu 730	6	U&M
VOLVO (A35)	4	JS Máquinas
VOLVO (A35)	3	Beadell
Total trucks	36
Dozers
CAT D9	5	U&M
CAT D6	1	U&M
CAT D6	2	Vale do Amazonas
Total dozers	8
Support Equipment
Graders (Cat 16H, 14M 140)	5	U&M
Drills (Panther 1500, DX 680)	8	U&M
Loaders  (Cat 980)	4	U&M
Water trucks (Mercedes, Volvo)	4	U&M
Fuel truck	5	U&M

16.3.1 Grade control and dig plans

The principal means of grade control is by RC drilling. Drilling is normally conducted on a 12 m x 6 m (y, x) pattern in the ore zone and 12 m x 12 m pattern outside of the ore zone. Holes are normally drilled to a depth of at least 24 m to enable testing of 20 vertical metres (5 working benches). Where access is difficult, RAB drilling performed by the geology department air track PW blasthole rig can infill drill angled holes to a depth of 8 m.

Channel samples of the pit wall are taken on every bench where the wall crosses the lode. At times face samples are taken along the face of an active bench to gain a better visualization and understanding of controls on mineralized structures. Areas of potential ore that do not have adequate sample coverage and cannot be tested with RC or RAB prior to mining due to access or timing issues have truck loads sent to the rehandle pad for further sampling. All RC, RAB, and channel sampling is undertaken on 1 m intervals, assayed and logged for lithology and weathering.

Grade control data is entered into an SQL database (Datashed) which is linked to an SQL (Torque) database in Minesight. The assays are composited to 2 m intervals coded by pre-defined estimation domains using a scripted routine. Top cuts are applied as part of the compositing process. Estimation is undertaken on a domain basis by OK on a bench basis using all data within range of the search ellipse for each domain.

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The grade estimation is a scripted routine in Minesight which is fully editable for revised estimation parameters using a simple front end. Before each dig plan is produced, the composite file is updated and the estimation re-run. Dig plans are generated by first interpreting the drillhole assay grade and geology on 10 m sections. The cross-section of the ore polygon solid is designed based on the interpreted ore contacts and where dipping, either positioned as a vertical or dipping dig contact.

Following the review in cross-section, the ore polygons are adjusted in plan view and polygon solids created. These are then imported into a Minesight (MSIP) graphical database which calculates the volume, grade and tonnage from the geological model. MSIP also allows for the inclusion of dilution and mining recovery at the calculation stage with dilution based on lode width and the proportion of the ore polygon perimeter exposed to waste. Once the grade is calculated, the dig polygon is assigned to a grade category as defined in Table 16.6.

Table 16.6 Grade control grade ranges by material type

	Grade range Au (g/t)		Destination
Grade category	Pit: Tap AB / Tap C	Pit: Urucum / Urucum East	Oxide	Sulphide
Very High Grade	>1.60	>1.60	ROM Pad	ROM Pad
High Grade	0.85 - 1.59	0.90 - 1.59	ROM Pad	ROM Pad
Low Grade	0.65 - 0.84	0.75 - 0.89	ROM Pad	Long Term LG Stockpiles
Marginal	0.50 - 0.64	0.50 - 0.74	Long Term Marginal Stockpiles	Long Term Marginal Stockpiles

The pit topography is regularly coded to the model to generate accurate dig volumes and tonnages. In addition to gold ore, Iron Ore and Waste polygons are also generated. The iron ore block model has been integrated into the gold grade control block model and although iron ore does not currently hold value for GPR, it is stockpiled separately as a future resource.

16.4 Underground mining

16.4.1 Geotechnical and hydrogeological considerations

BDR completed a program of resource infill drilling below the Urucum North Pit. Most of these holes have been geotechnically logged for structural data, with the addition of one purely geotechnical hole to investigate the nature of the large structure located between the Urucum North and South pits.

The geotechnical data logged from these holes has been analysed to provide preliminary estimates of stable stope, development, and pillar dimensions using industry standard empirical techniques.

16.4.2 Data used

Stability ratings have been calculated from data logged from the following holes (refer Table 16.7). The holes have been grouped into North, Central, and South areas to establish any trends in ground conditions. The positions of these holes, relative to the areas considered for underground mining, are indicated in Figure 16.6.

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Table 16.7 Diamond drillholes used in the geotechnical analysis

Area			Borehole ID
North	FD1393	FD1409	FD1411
Central	FD1398	FD1401	FD1402	FD1410
South	FD1399	FD1400	FD1404	FD1405	FD1418

Figure 16.6 Long projection of Urucum North Underground looking west, showing geotechnical drillhole locations

Source: AMC 2016 PFS

16.4.3 Major structures

A small number of significant faults have been logged. These appear to have two main orientations, sub-parallel to the lodes, but not directly adjacent to the mineralized areas, and striking north-northeast across the orebody.

The structures that are most likely to affect the mining and infrastructure development are the extensive pegmatite intrusions. Although the pegmatites are generally massive and competent, intense laminated structures can be present adjacent to the contacts, which may be destabilized by blasting. These intrusions take several forms:

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• Sub-vertical dykes: where these intersect the surface, deep (>200 m) weathering can be expected on the contacts, with the possibility of some water inflows.

• Cross-linking structures between sub-vertical dykes: these can be heavily structured, both within and at the contacts of the structures with the host rock.

• Sub-horizontal sills: these can be heavily structured at the contacts of the structures with the host rock.

The mine design should avoid placement of development in these areas where possible.

A north-northeast striking pegmatite intrusion has been identified between Urucum North and Central pits, which would be crossed by one of the proposed ramp routes. Logging of diamond drilled core through this feature indicates that below the base of oxidation, the ground is competent and should not present any difficulties to development. Cover drilling will be required to identify potential water sources or weathering on the contacts of the pegmatite.

16.4.4 In-situ stress environment

No in-situ stress measurements have been completed at Urucum North underground mine.

It is not expected that high stress regimes will be present in the region. However, it would be prudent to carry out stress measurements at an early stage of the underground mine life to confirm this assumption. The results from the stress measurements should be analysed to review the proposed LOM sequence, particularly for deeper areas, as the planned mine depth is beyond 700 m from surface.

Although the strength of the rock mass means that major seismic activity is unlikely, as the mine develops deeper, there could be a significant potential for strain-bursting at the contacts between the different rock types, due to their different elastic properties.

16.4.5 Weathering

Figure 16.7 presents the profiles both for the base of colluvium and for the base of weathering. The base of weathering is relatively shallow (less than 50 m over the orebody and less than 30 m over the country rock) and thus will have little effect on the underground mine. The main considerations will be the location of the portal, upper sections of vertical development (such as ventilation shafts), and the potential for much deeper weathering on the contacts of the pegmatite dykes and major faults.

The portal will be positioned within the open pit, below the level of the weathering, and vertical development will be positioned in areas of minimal thicknesses of weathered material, which would then either be dozed off or a pre-sink developed from surface to the top of the fresh rock.

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Figure 16.7 Cross-section showing weathering profile (Section 99,865 mN)

Source: AMC 2016 PFS.

16.4.6 Ground water

Little ground water is expected to be encountered underground, with the exception of contact areas between the pegmatites and the country rock and the major cross-cutting faults. It would be prudent to carry out cover drilling when approaching any of these features.

16.4.7 Rock mass classification

All holes recently drilled for the resource infill have been geotechnically logged for a distance of at least 30 m on either side of the mineralization. All holes were oriented, and logging was carried out before geological samples were removed.

The geotechnical information collected has been used to populate the mine design parameters for excavation dimensions and ground support requirements. Overall, the geotechnical data collected is of reasonable quality. Some problems have been experienced due to poor drilling practices in the earlier holes, but this has been addressed thus improving the reliability of orientation and run length in later holes.

The logging has provided sufficient data to be able to estimate both Barton’s rock quality index, Q and Bieniawski’s rock mass rating, RMR. These classifications allow the estimation of stable spans for stoping and development, as well as ground control requirements.

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Safe development spans and support requirements can be estimated from Q, and it is possible to estimate a stable hydraulic radius to determine safe stope spans from a modified Q value (Q’), using the modified Mathews stability method (Potvin, 1988).

The results for the analyses are summarized in Table 16.8.

Table 16.8 Range of Q’ and RMR values

Zone / location		Q’			RMR
		Min.	Max.	Avg.	Min.	Max.	Avg.
Hangingwall	North	3.58	16.67	12.4	63	82	74
	Central	2.78	96.71	33.04	63	84	74
	South	3.13	200.00	41.79	60	87	73
Backs	North	1.39	33.27	10.39	63	78	73
	Central	3.12	266.67	55.71	63	82	75
	South	1.39	100.00	25.79	72	87	80

The overall rock mass quality (either Q’ and RMR) is competent and does not vary much between hangingwalls and backs at the same locations.

AMC used the Q’ values to estimate the modified stability number (N’) and to determine the stable hydraulic radii (HR) for each of the stope zones. The same conservative rock mass condition (the lower bound of the average Q’ for the stoping zones) is used as Q’ for stope design. Table 16.9 presents the stope design parameters and the design limit of HR for an unsupported stable stope.

Table 16.9 Stope design factors and hydraulic radii

Q’	N'	Maximum HR if unsupported (m)
12.5	28.1	8.2

16.4.8 Stable stope dimensions

Figure 16.8 shows a summary of the maximum spans of stable stopes. In the example shown, if the spacing of the sublevel is 20 m (24 m floor to roof of the levels), an unsupported strike length of up to 50 m would remain stable, assuming no significant defects (such as major structures) are present. The 50 m will define either the rib pillar spacing, or the maximum strike length of the stope that can be left open at that height before backfilling the stope.

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Figure 16.8 Modified Mathews short-term stability estimate

  

Source: AMC 2016 PFS

16.4.9 Pillar dimensions

16.4.9.1 Crown pillar

The crown pillar between the base of the open pit and the upper underground excavation is critical to safe operations. This pillar must remain stable for the life of the excavation, either by design of sufficient size or ensuring sufficient confinement is provided to minimize pillar collapse. This may include additional ground control measures in the bottom of the pillar. Mining of blind up-hole stopes into the crown pillar should be avoided.

The height of the crown pillar will depend on the width of the stope beneath it. The minimum height to width ratio for the crown pillar should be 3:1. The actual height can be optimized by controlling the profile of the base of the pillar and with the use of ground control systems (cable bolts and shotcrete, for example).

It is essential that the crown pillar be formed with the minimum height in sufficiently strong (fresh) rock, and not within the weathered profile or colluvium horizons. The presence of deep weathering adjacent to any pegmatite contacts would need to be investigated.

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16.4.9.2 Sill pillars

Sill pillars should be a minimum of 10 m high, with a minimum height to width ratio of 2:1. However, where stacked lenses, or wider orebody zones exist, the sill pillar height may need to be increased to maintain the 2:1 height to width ratio.

16.4.9.3 Rib pillars

Rib pillars between open stopes along the strike of the same lens should be a minimum of 5 m high, with a minimum height to width ratio of 1:1. Pillars between lenses, where stacked lenses exist, should aim to form a minimum 10 m pillar. Where stacked lenses exist, the sequence of extraction should be from the hangingwall to the footwall. Stopes planned in the footwall of an existing stope should be limited in size, or small ore-rib pillars should be introduced to reduce the chances of inter-lens pillar failure.

16.4.10 Excavation and support requirement

16.4.10.1 Lateral development

Various development drive sizes have been proposed for use in the mine. The existing Q values indicate that the ground conditions will be suitable for the proposed development dimensions. Ground conditions are expected to be mostly good, which will result in low ground support requirements. Additional support might be required when crossing contacts between the pegmatites and country rock and crossing the major fault zones. Where poor ground conditions are encountered, additional support will be required, which should be designed by a competent person after inspection of the conditions.

Long term development

A standard pattern of five 2.4 m long rebar bolts per ring (nominally 1.5 m x 1.5 m spacing) is recommended, consisting of three bolts in the back and one in each shoulder. Taking into account the height of the ramp and cross-cut development, it will be impractical to either inspect or scale the backs of the drives. For this reason, it is highly recommended that either weld wire mesh (100 mm x 100 mm) or 50 mm thick fibrecrete is installed in the backs and shoulders of the ramp and cross-cut development, which can prevent the loose fall and save time in ground control maintenance (such as inspection, sound and check-scaling loose), thus to improve productivity.

Spot bolts may be required to secure wedges and smaller blocks formed by adverse joints in the backs or sidewalls.

In addition to primary support, cable bolts may be required for intersections with large span. It is recommended that all intersections be inspected by a qualified person, and the requirement for cable bolts be evaluated based on actual ground conditions.

Short term development

A similar pattern of bolts should be used in the ore drives, but these can be friction bolts and will not require mesh or shotcrete, except where ground conditions are poor.

Cable bolting might be required in sections of the orebody where brows of stopes are formed, particularly where open stopes remain. When double lift stopes are planned, the sub-level ore-drive may require hangingwall cables to reduce the effective unsupported span. This should be evaluated based on the stope dimension and actual ground condition.

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16.4.10.2 Vertical development (ventilation shafts)

Ventilation shafts will be required in various locations throughout the operation. Specific investigation holes will be required to assess the local conditions and stability of these shaft positions. However, given the current information, it is expected that stable shafts could be excavated throughout the fresh rock zones.

Instability will be present in areas of the weathered and colluvium zones. Ideally, these layers should be avoided by placing the shaft collars in areas where this material has been removed. Where shafts are required to pass through these zones, pre-conditioning or shaft sinking methods may be required to reach the fresh rock horizon.

16.4.10.3 Portal locations

Location and design of portals should be considered in fresh rock positions only. Specific investigation work may be required to ensure the locations will be suitable for the expected LOM, and that the layout of the decline does not pass into areas of significant deep weathering, or poor ground conditions related to currently undefined faults. Portal locations should aim to have a minimum of 5 m of sound, fresh rock above the completed portal excavation, and avoid being at the bottom of the open pit excavation where they might be vulnerable to flooding.

16.4.11 Underground mine design basis

16.4.11.1 Mining limits

The Urucum North Underground mine is based on the following battery limits:

• Mining below Urucum North pit design of 11 March 2016 (des_uru_nth_march_a2016.dxf) and north of 99,300 mN. Mineral Resources based on the geological block model Reference: urucum_ug_resources2015_v17.dm (8 October 2015) provided by BDR.

• LOM planning based on Measured and Indicated Mineral Resources.

• Mineral Reserve estimates based on Measured and Indicated Mineral Resources only.

• West of the mining license boundary at approximately 402,154 mE.

16.4.11.2 Mineral Resource estimate

The Urucum North Underground Mineral Resource estimate is discussed in Section 14 of this report and is the basis for the Underground Mineral Reserve estimate. This resource is based on unweathered “fresh” material reported above a COG of 1.6 g/t Au and is shown in Table 16.10.

Table 16.10 Urucum North Underground Mineral Resource estimate - based on COG of 1.6 g/t Au

Mineral Resource category	Tonnes	Au grade (g/t)	oz
Measured	258,125	4.09	33,943
Indicated	2,577,852	4.28	354,726
Total	2,835,977	4.26	388,669

16.4.11.3 Stope optimization

The MSO module from Datamine Studio 3^TM mine planning software has been used to produce the stope shapes and to generate stope inventories. The MSO process automates what a planning engineer would do manually, that is, generate stope string shapes on sections, link the string sections to create a stope wireframe shape, and then evaluate the wireframe against a block model.

MSO provides a stope shape that maximizes recovered resource value above a COG while also catering for practical mining parameters such as:

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• Minimum and maximum mining width.

• Anticipated unintentional wall dilutions.

• Minimum and maximum wall angles.

• Minimum separation distances between parallel and sub-parallel stopes.

• Minimum and maximum stope heights and lengths.

Figure 16.9 below summarizes the tonnage-grade curves for the MSO inventories based on various COGs. These MSO inventories include some out-lying (isolated) stopes that fall some distance from the main mineralized areas.

MSO parameters, including mining dilution parameters, are discussed in Section 16.2.4.

Figure 16.9 Tonnage / grade curves based on MSO stope shapes generated above 1.6 g/t Au COG

Source: AMC 2016 PFS

After the MSO process is complete, Mine24D has been used to estimate the development ore and stoping ore inventories. Mining recovery factors have only been applied to the stoping ore inventories.

16.4.11.4 MSO estimates

Mining inventories are based on the MSO stope shapes generated using the following criteria:

• Stope height generally 20 m (equivalent to level interval) although some stopes near the upper resource boundaries can be either 10 m or 15 m in height.

• Stope length of 50 m along strike (divided into 2 x 25 m scheduling increments).

• COG of 1.6 g/t Au.

• Planned dilution of 0.5 m footwall plus 0.5 m hangingwall.

• Minimum mining width of 2.0 m (plus 1.0 m of dilution).

• Minimum pillar distance between parallel lodes 10.0 m.

• Removal of isolated stopes requiring excessive development.

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Note that the MSO estimates include outlying stopes (outliers) that, for mine planning purposes, have been excluded based on incremental access development costs. It should also be noted that this estimate has not allowed for modifying factors such as, ore losses to pillars (rib and sill), mucking recovery, and other external dilutions. These modifying factors have been applied to the detailed ore production schedules, generated using Earth Works Production Scheduling (EPS) software.

Table 16.11 summarizes the MSO parameters used (for reference).

Table 16.11 MSO parameters

Item	Unit	Value
COG for MSO stope shapes (before dilution added)	Au (g/t)	1.6
Minimum mining width	m	2.0
Dilution skin width - total	m	1.0
Stope height 1	m	20
Stope height 2 from base	m	15
Stope height 3 from base	m	10
Stope increment - strike length	m	25
Minimum transverse pillar width	m	10
Include inventory within pit “reserve” pit design *	Flag (1/0)	yes
Conduct stope-adjacency smoothing	Flag (y/n)	no
Reference X start for stope shape grid & extent	Easting | m	401,880 | 200
Reference Y start for stope shape grid & extent	Northing | m	99,310 | 1,075
Reference Z start for stope shape grid & extent	Elevation | m	-590 | 800
Maximum weathered rock inside stope shape	WEATH=5,000%	10
Block model: "urucum_ug_resources2015_v17"	Version | date	ugv17d #| 8 Oct 2015

Figure 16.10 shows the MSO stope shapes and the development designs in 3D, after removal of isolated stopes.

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Figure 16.10 3D view of mine design and MSO stopes - by lode number

Source: AMC 2016 PFS

16.4.12 Access development

The key underground access features are:

• A trucking portal centreline located at 402,053 mE, 100,062 mN, 173 mRL (floor level), at the north end of the Urucum North pit.

• A ventilation portal centreline located at 402,037 mE, 99,744 mN, 136 mRL (floor level), at the south end of the Urucum North pit.

• A decline system developed to the east of the lodes in the geological hangingwall that extends from approximately 173 mRL to -490 mRL. This splits into a twin decline system between 10 mRL and -390 mRL. The lateral coverage is approximately between 99,400 mN to 100,200 mN.

• Two vertical ventilation rises exhausting to surface located at the north end of the Urucum North pit at 402,030 mN, 100,200 mE and 210 mRL.

16.4.12.1 Decline stand-off distance

The decline has been designed based on the following minimum stand-off distances:

• 5 m east of BIF contact.

• Minimum 15 m standoff between stope void and decline, down to 200 mbs.

• Minimum 20 m standoff between stope void and decline, from 200 mbs to 400 mbs.

• Minimum 25 m standoff between stope void and decline, below 400 mbs.

By keeping the decline east of the BIF unit, crosscut lengths are significant on some levels (>70 m). Once underground trials commence, and the location of the payable lodes is better understood as well as the ground conditions within the BIF, it may be preferable to develop the ramp closer to the stopes within the BIF unit itself.

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16.4.12.2 Capital development estimates

The LOM capital development estimates are summarized in Table 16.12 and Table 16.13.

Table 16.12 Lateral capital development estimates

Lateral development - capital	Dimension: W x H (m)	Total length (m)
Decline	5.0 x 5.0	10,350
Decline stockpile	5.0 x 6.0	521
Egress access	4.0 x 4.0	1,155
Exploration drive	5.0 x 5.0	1,119
Fresh airway	5.0 x 5.0	60
Infrastructure	6.0 x 6.0	125
Level crosscut access	5.0 x 5.0	2,688
Level sump	4.0 x 4.0	384
Pump station	5.0 x 6.0	293
Return airway	5.0 x 5.0	2,386
Return airway oversized	6.0 x 5.0	565
Total		19,646

Table 16.13 Vertical capital development estimates

Vertical development - capital	Dimension: Ø or W x H (m)	Total length (m)
Egress raise	0.9	1,112
Fresh air raise	3.6	133
Return air raise	3.6	282
Longhole raise	4.0 x 4.0	742
Total		2,270

16.4.12.3 Operating development estimates

The LOM operating development estimates are summarized in Table 16.14.

Table 16.14 Operating lateral development estimates

Lateral development - Operating	Dimension: W x H (m)	Total length (m)
Ore drive	4.0 x 5.0	15,252
Level stockpile	5.0 x 6.0	2,216
Total		17,468

16.4.13 Mining methods

Based on the favourable ground conditions and relatively narrow ore body widths, Longitudinal Long Hole Open Stoping methods are preferred. Variations of Longitudinal Hole Open Stoping include:

• Up-hole retreat

• Down-hole benching

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16.4.14 Up-hole retreat mining method

The Up-hole retreat method, as indicated in Figure 16.11, allows mining to proceed in a top down direction with stopes being blasted using up-holes and retreating back towards a centralized crosscut. Backfill of stopes can be delayed with rib pillars providing support to the void sidewall. Normally the large voids are eventually filled with un-cemented waste rock, to reduce the possibility of regional instability.

Figure 16.11 Long projection indicating up-hole retreat mining sequence

Source: AMC 2016 PFS

This method is recommended in the initial production build up-phase of the project, for the following reasons:

• It does not require backfill until sometime after stopes are mined and even then, only if remaining pillars are stressed and begin to affect operations on lower levels.

• Minimizes the amount of development required to be completed before production commences, unlike overhand benching methods where production commences on the lower level and proceeds in an up-dip direction.

• The fact that backfill is not required as part of the mining cycle allows higher production rates to be realized.

• The method is considered relatively simple to plan and execute.

The disadvantages of the method include:

• Ore loss resulting from leaving behind ribs, resulting in an average ore loss of around 14%. (As the resource is developed and the grade distribution of the lodes becomes better understood it may be possible to position pillars in low grade areas, resulting in lower gold loss).

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• A significant amount of ore (approximately 70%) will need to be mucked using remote loading.

• Stope fall-off can damage LHDs.

• Often it is difficult to maintain a horizontal stope back at the top of an ore block, resulting in additional ore loss.

16.4.15 Down-hole benching mining method

The Down-hole benching mining method indicated in Figure 16.12, allows mining to proceed in a bottom-up direction once access to the bottom levels is established. Generally, the method relies on down-hole drilling of blastholes. As each stope or level is mined it is back-filled with waste rock. Once a stope is filled, mining of the adjacent stope or stope on the next level above, can commence.

Pillar ore loss can be reduced for this method as pillars only need to stand until mining of the subsequent stope commences. Furthermore, some mines have removed the pillars altogether, by choke blasting up to the waste rock interface. The effect of blasting ore against the waste rock tends to compact the waste, allowing reasonable extraction of the abutting blasted ore with minimal waste dilution.

Figure 16.12 Long projection indicating down-hole benching mining method and sequence

Source: AMC 2016 PFS

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This method is recommended at Urucum North Underground in the lower parts of the mine for the following reasons:

• Higher ore recoveries (estimated at 92%) can be achieved for the method as the backfill is placed much sooner compared to that for the Up-Hole retreat method, allowing smaller ribs (or even no ribs) to be left behind.

• Allows down-hole drilling and charging of blastholes.

• The method is considered relatively simple to plan and execute.

The disadvantages of the method include:

• Ore production commences only when access to the bottom of an ore block is completed.

• A significant amount of ore (approximately 70%) will need to be mucked using remote loading.

16.4.16 Sill pillar recovery

In order to allow multiple mining fronts to be developed, allowing higher ore production rates to be achieved, sill pillars have been allowed in the design as indicated in Figure 16.13.

Figure 16.13 Sill pillars

Source: AMC 2016 PFS

The sill pillars are extracted after the stopes above or below have been mined and backfilled. Figure 16.14 indicates two options for sill pillar recovery. Both options are considered viable and should be trialled. The sill pillars are based on a standard level interval of 20 m equating to a sill pillar height of 15.5 m after allowing for the ore development.

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Alternatively, it may be preferable to consider lower height sill pillars, e.g. 6 m to 10 m, that are left in-situ (unrecovered). It is AMC’s opinion that the preferred option can only be determined from mining trials and detailed stress modelling.

Figure 16.14 Options for sill pillar recovery

Source: AMC 2016 PFS

16.4.17 Ventilation strategy

The mine will be ventilated by a Pull or exhausting type ventilation system. That is, the primary mine ventilation fans will be located at the primary exhaust airways of the mine and develop sufficient pressure to ensure all work places are supplied with the required fresh air and such that contaminants are removed to the exhaust air system and ultimately to the surface. Fresh air will be sourced from the surface through the main portal entrance and subsequently into the twin declines. Fresh air is also provided through an access drive into the open pit to provide an additional intake and ensure air velocities in the declines are not excessive.

Two primary return air raises are planned to breakthrough to the surface at the northern end of the mine, each with a single exhaust fan installation. As they are drawing from a common system, there is some redundancy provided should one be down for maintenance or breakdown. Internal return air raises are developed in tandem with each of the twin declines. A return air transfer drift is required to convey exhaust air from the southern internal return air raises across to the northern located primary raises.

The ventilation strategy is schematically represented in Figure 16.15.

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Figure 16.15 Long projection of the primary ventilation system

Source: AMC 2016 PFS

A total of 350 m³/s is planned for ventilation of the underground operation including the diesel equipment fleet, infrastructure, and personnel. This quantity reflects the requirements outlined in the Brazilian Regulatory Standards for Mining (Normas Reguladoras de Mineração) and a production rate of up to 700,000 tonnes per annum.

The main fan installations are planned to be located on surface and will be controlled with variable frequency drives to allow fluctuation in air volumes during the LOM. Ventilation network software modelling of the circuit has been undertaken to reflect the peak primary fan duties that could be reasonably expected.

Subsequent detailed scheduling confirmed maximum production rates of around 580,000 tpa, hence there is scope to optimize the ventilation requirements further.

16.4.17.1 Level ventilation

Level distribution is designed such that fresh air will be sourced from the declines and travel along each level access drive where it is picked up by an auxiliary fan and duct installation for ventilation of work activities. Contaminated air from development and productions level activities exhaust back to the return air raise in the access drive of each level (refer Figure 16.16).

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Figure 16.16 Forced ventilation for a typical level

Source: AMC 2016 PFS

16.4.17.2 Development ventilation

Ventilation will be required for all phases of the development drill and blast cycle, (e.g. drill, blast, muck, support, etc.). The ventilation requirement for development will need to:

• Clear any residual dust generated (primarily by mucking).

• Remove heat generated from diesel engines and, to a lesser extent, heat from electric / hydraulic equipment such as drill jumbos and bolters.

• Dilute and remove diesel engine exhaust to acceptable levels.

• Remove noxious gases after blasting.

The ventilation strategy for development uses a forced air fan (push) and duct system. To define the required ventilation flow for the excavation heading, a minimum flow of 0.078 m³/s per kW diesel engine power has been used (Brazilian Regulatory Standards for Mining).

For the decline development, it is assumed that there will be a maximum of one Atlas Copco ST 1030 loader (186 kW) and one Scania P20 truck (294 kW) in the decline heading. The total required airflow to be delivered to the end of the heading was estimated at 37.4 m³/s. This will be delivered through twin ventilation ducts with one duct extending to the face and the other to the truck loading bay. Each duct will require an auxiliary fan with a 90 kW motor.

For level development and also during the production cycle, it is assumed that trucks will not enter the ore drives and instead be loaded at the level accesses. As such, the mining levels will need to accommodate an Atlas Copco ST 1030 as the largest piece of equipment. At 186 kW, the ore drives will need to be ventilated with 14.5 m³/s airflow at the face. The auxiliary ventilation requirements for this is a single duct with 55 kW fan.

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16.4.18 Ventsim modelling

The ventilation system for the Urucum North Underground Mine was modelled using Ventsim Visual™ Advanced.

The ventilation network for the Urucum North Underground Mine was analysed by importing the mine design from mine design software and then applying attributes for each of the airways relative to their dimensions, frictional resistance, length, etc. The airflow previously determined was applied to the network.

16.4.18.1 Modelling results

AMC developed a Ventsim ventilation model for the Project for three primary purposes:

• To validate the operability of the ventilation circuit ensuring airflow can be provided to all the required areas and during all phases of the mine.

• To ensure compliance with design criteria and legislated ventilation requirements.

• To determine fan duties and energy requirements.

Over the life of the mine, there will be a multitude of settings for the ventilation circuit, depending on the type of activities and their location throughout the mine. AMC modelled the circuit to reflect the peak primary fan duties that could be reasonably expected.

Table 16.15 Surface fan details

Description	Specification
Type	Horizontal mount axial mine fan
Configuration	One exhausting fan per raise (two total fans)
Requirements	175 m3/s at 3,911 Pa raise collar total pressure (fan static pressure) on each raise
Each fan duty	175 m³/s at 4,464 Pa fan total pressure (air density 1.2 kg/m³ at 21.1°C)
Each fan motor	1,500 hp at 1,200 rpm complete with variable frequency drive

16.4.19 Egress

In development of the ventilation strategy for the Urucum North Underground Mine, consideration has been given to the potential for mine emergencies. As such, the following criteria have been established:

• Ramps will be in fresh air following their initial development.

• On all levels, escape can be to either the ramp or to the escape way.

• In each ramp, escape may either be up the ramp or down the ramp to a safe area.

• Refuge Stations will be portable for flexibility of location at the most appropriate points in the mine. The number of portable Refuge Stations will be confirmed in later stages of design.

• Whilst the primary means of communication will be by radio, a back-up stench system will be in place for introduction of ethyl mercaptan into both the portal and FAR.

16.4.20 Underground mine production and development schedules

The basis for the scheduling includes:

• Commence decline development on 1 January 2019 with the portal to be constructed at 172 mRL (floor level).

• Focus on mining higher grade areas in 2021 and 2022.

• Assume a top-down stope extraction sequence in upper levels to allow early production.

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• Assumes early development to the higher grade areas in the lower levels of Lode 1002 where down-hole benching commences in a bottom-up orientation.

16.4.20.1 Development profiles and scheduling rates

The lateral development profiles (jumbo development), quantities and scheduling rates estimated by AMC are indicated in Table 16.16.

Table 16.16 Lateral development details

Type	Development profile m x m	Advance rate m/month	Quantity (m)
Decline (including stockpiles)	5.0 W x 5.0 H	140	10,871
Egress access	4.0 W x 4.0 H	140	1,155
Exploration drives*	5.0 W x 5.0 H	NA	1,119
Fresh airway access	5.0 W x 5.0 H	140	60
Service bays, magazine, pump stations, infrastructure	6.0 W x 6.0 H	60	418
Crosscuts	5.0 W x 5.0 H	140	2,688
Ore drives	4.0 W x 5.0 H	60	15,252
Return airway access	5.0 W x 5.0 H	140	2,386
Return airway access - oversize	6.0 W x 5.0 H	140	565
Level sump	4.0 W x 4.0 H	140	384
Level stockpile	4.0 W x 4.0 H	60	2,216
Total			37,116

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

* Exploration drives are indicated in some of the figures and tables, but they have not been included in the cost estimate.

The vertical development profiles, quantities, and scheduling rates estimated by AMC are indicated in Table 16.17.

Table 16.17 Vertical development details

Type	Development profile	Advance rate m/month	Quantity (m)	Comments
Egress - long hole	1.1 m dia.	70	1,112	Long hole
Fresh air intake - raisebore	3.6 m dia.	70	133	Development rate inclusive of setup and tear-down time
Return air raise - raisebore	3.6 m dia.	70	282	Development rate inclusive of setup and tear-down time
Return air raise - long hole	4.0 m x 4.0 m	60	742	Long hole
Total			2,270

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

16.4.21 LOM underground production schedule

The total projected waste development tonnes, and the annual production from the underground mine as a whole is provided in Figure 16.17. The production and development schedule is summarized in Table 16.18.

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Figure 16.17 Projected production profile

Source: AMC 2016 PFS

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Table 16.18 LOM underground production schedule

Item	Unit	Year	Y1	Y2	Y3	Y4	Y5	Y6	Y7	Y8
Ore and waste		Total	2019	2020	2021	2022	2023	2024	2025	2026
Waste mined	kt	2,000	160	560	614	432	90	96	33	16
Production - stopes	kt	1,894	3	155	229	343	373	281	271	239
Production - development	kt	485	2	63	112	122	46	72	32	35
Total ore production	kt	2,378	5	218	341	465	419	353	303	274
Production tonnage gold grade	g/t	3.6	2.8	3.0	3.4	4.1	4.0	3.8	3.7	3.1
Mine development
Decline	m	10,350	1,454	3,250	3,316	2,330
Decline - stockpile	m	521	81	160	160	120
Egress access	m	1,155	8	367	348	267	60	105
Fresh airway	m	60		60
Infrastructure	m	125	125
Level crosscut access	m	2,689	100	879	1,068	524	118
Level sump	m	384	16	128	144	96
Pump station	m	293	60	63	80	90
Return airway	m	2,386	130	601	810	505	170	170
Return airway oversized	m	565	204	361
Ore drive	m	15,252	65	1,660	3,150	3,679	1,811	3,002	1,084	801
Level stockpile	m	2,217	42	755	811	509	100
Total lateral development	m	35,997	2,285	8,284	9,887	8,120	2,259	3,277	1,084	801
Major airways - raisebore	m	416	219	197
Major airways - long hole	m	742	31	163	299	156	42	51
Egress - long hole	m	1,112		338	430	218	43	82
Total vertical development	m	2,270	250	698	729	374	85	133
Backfill
Backfill placed	kt	1,307	1.6	53.8	120.9	261.8	263.2	214.5	197.4	171.5
Total trucking	M t-km	15.8	0.1	1.5	2.8	3.6	2.2	2.5	1.7	1.1

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16.4.22 Underground mobile equipment

Equipment selection has been based on using mechanized underground mobile mining equipment that matches the proposed production rates and stope / ore drive dimensions. The main fleet will be supported by an appropriate ancillary fleet. Note that equipment models listed below are examples of the class of equipment, not recommendations for specific makes and models.

Diesel equipment such as service vehicles and trucks will be serviced on surface. Electric-hydraulic units (jumbos and production drills) and LHDs travel at slower speeds and therefore will be serviced underground for daily and weekly services and on surface for major services and overhauls.

Table 16.19 and Figure 16.18 show the units of equipment by year.

Table 16.19 Major equipment required over LOM

Summary	2019	2020	2021	2022	2023	2024	2025	2026
Development jumbos	1	5	6	5	2	2	1	1
Ground support rigs (bolters, shotcreter, etc.)	3	3	4	3	3	3	3	3
Production drills	1	1	1	1	1	1	1	1
LHDs	2	4	5	5	4	4	3	3
Underground trucks	1	2	4	5	3	3	3	2
Support equipment (ITs, Grader, etc.)	8	9	10	9	8	8	7	7
4WDs	4	4	4	4	4	4	4	4
Total	20	28	34	32	25	25	22	21

Figure 16.18 Major equipment required over LOM

Source: AMC 2016 PFS

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16.4.22.1 Development drills / jumbos

Two-boom development drills (jumbos) such as Atlas Boomer 282 will be used to bore all development headings. As they will only be used for development drillhole boring, a longer feed rail can be used.

It may also be an option to use single boom jumbos and smaller LHDs in narrow ore lodes to try to minimize waste dilution. This advantage would need to weigh up against the lower productivities of both the single boom jumbos and smaller LHDs.

AMC recommends commencing operations with the larger equipment until the width of the ore lodes is better understood and the actual dilution in mining operations accurately determined.

16.4.22.2 Production drills

Production drill selection has been based on the following criteria:

• Preference for a basic drill which can be more easily maintained in Brazil.

• No requirement for an operating cabin.

• Average hole length of 15 m (with some holes to 20 m).

• Ability to drill up-holes and down-holes for near vertical and narrow orebodies.

• Hole diameter of 64 mm with ability to ream some holes to 115 mm diameter.

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The Simba H1354 (refer Figure 16.19) would be suitable and has the following features:

• Feed can be selected to suit 1.2 m, 1.5 m, or 1.8 m drill steels.

• Electric hydraulic rig controlled using remote umbilical panel.

• RHS rod handler (17+1 rods).

• Total feed length is 3.97 m (assuming 1.8 m drill steels plus back hammer facility).

• Ability to drill up-holes and down-holes.

• All drilling is undertaken remotely using 20 m stand-off umbilical cord panel.

• Slide table gives up to 3 m parallel hole capacity.

Figure 16.19 Simba H1354 production drill

Source: AMC 2016 PFS

16.4.22.3 Underground trucks

Based on an average production rate of approximately 500 ktpa, AMC has considered the following trucks for haulage of ore and waste:

• Volvo A35 articulated hauler. This 6WD truck is common in underground mines in Brazil. Its overall height is approximately 3.3 m with a tray height of approximately 2.9 m. As the truck is capable of withstanding poor roadway conditions, it is better suited to initial main ramp development and extensions of the ramps before a permanent well graded roadway is established.

• Scania P420 rigid body truck. This 38 t truck is effectively used at Ferbasa’s Ipueira underground mine. The truck is inexpensive to purchase or maintain when roadways are well maintained and graded.

• Scania 50 t version of the P420. Ferbasa have reported that the truck has similar dimensions to the P420 with the exception of the length, which will be extended by approximately 1.5 m.

• Other articulated underground trucks such as the Caterpillar AD40.

From the trucks noted above, AMC has selected the Volvo A35 articulated hauler for the following reasons:

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• The Volvo A35 is commonly used in Brazil because of its proven performance, especially where poor ramp conditions exist.

• AMC understands that the Volvo trucks can be purchased in Brazil, without attracting large import duties.

Once mining trial operations are underway, and decline roadway conditions are better understood, it may be beneficial to reconsider the truck fleet as any of the 4 trucks indicated above could be accommodated in the nominated decline profiles. The benefits of using the Scania P420 truck in particular include:

• Lower cost - the Scania trucks cost approximately US$140,000/unit which is considerably less than the cost of articulated underground trucks.

• Scania truck spares are common in Brazil, thus resulting in more available spare parts and lower maintenance costs.

Figure 16.20 shows a Scania P420 truck being loaded by a Volvo 120E front-end loader.

Figure 16.20 Scania P420 truck being loaded by a Volvo front-end loader

Source: Ferbasa

16.4.22.4 LHDs

A fleet of 10 t LHDs is proposed at this stage. These LHDs are 2.6 m wide and approximately 2.4 m in height and can easily be accommodated within the 5.0 m H x 4.0 m W ore drives.

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16.4.22.5 Ground support rigs

A dedicated ground support bolter such as the Atlas Copco Boltec M is recommended for ground support. This unit will be used for hole drilling, bolt installation, and mesh installation.

16.4.22.6 Utility vehicles

Utility vehicles include the following:

• Scaler (e.g. Scaletec UV 2 or Normet Scaler 2000).

• Mobile rock breaker.

• Integrated Tool carriers (ITs) for installation of services and transportation of materials.

• Explosives charge-up units (e.g. Charmec).

• Grader (e.g. Caterpillar 12H Grader) to maintain permanent underground roadways. If it is difficult to establish good roadways, it may also be necessary to deploy a vibratory roller periodically (e.g. every 6 months) to help compact the roads.

• 4WDs for transportation of technical staff and supervisors.

16.4.23 Underground mine personnel

AMC estimates the following workforce for the underground project once it is in full production (not including 4 Expat trainers and additional resources to cover for absenteeism):

• Underground mining equipment operators 91

• Underground maintenance personnel 60

• Management, supervision, and Technical Services 57

• Surface maintenance 15

• Underground service crews 13

• Total 236

16.4.23.1 Working shifts

Mine schedules have been based on 8 hr shifts (1 hr on surface and 7 hrs underground) as shown in Table 16.20.

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Table 16.20 Underground workforce time allocation during shift

Activity	Duration	Notes
Average travel time - Home to site	30 min	Unpaid time
Check in at mine gate	0 min	Unpaid time
Travel gate to UG office	20 min	Paid surface time
Clock in at lamp room	0 min	Paid underground time
Shift change discussions and instructions	15 min	Paid underground time
Travel to underground work place	15 min	Paid underground time
Underground duties	3 hrs	Paid underground time
Shift break	15 min	Paid underground time
Underground duties	2 hrs 45 min	Paid underground time
Travel to surface	15 min	Paid underground time
Reporting of completed duties, consumables etc.	15 min	Paid underground time
Clock out at lamp room	0 min	Paid underground time
Change / shower	20 min	Paid surface time
Travel gate to UG office	20 min	Paid surface time
Check out at mine gate	0 min	Unpaid time
Average travel time to home	30 min	Unpaid time
Total paid underground time	7 hrs
Total paid surface time	1 hr
Total underground time working on job	5 hrs 45 min

16.4.23.2 Rosters

The underground operations personnel are to work a six panel, three shift roster. Features of this roster are:

• Continuous operations - 24 hrs per day, 365 days per year, three shifts per day.

• 7 hrs per shift plus 1 hr on surface, with a maximum of 5 hrs 45 min working at the face.

• 19 days on, 9 days off basis.

• 28 days annual leave.

Expats will work a four panel, two shift roster. Features of this roster are:

• Continuous operations - 24 hrs per day, 365 days per year, two shifts per day.

• 12 hrs per shift.

• 14 days on, 14 days off basis.

• 14 days annual leave.

It is expected that senior staff will work a non-continuous roster of 5 x 12 hr shifts followed by 2 days off, with 20 annual leave days per annum.

Table 16.21 indicates the break-down of the management and supervision, technical services, surface maintenance, and underground service crew employees (excludes underground equipment operators and maintenance personnel) for the underground project once in full production.

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Table 16.21 Mining department employees - excluding UG equipment operators and maintenance personnel

Position	Class	Responsibility / duties	Max number
Mining
UG Manager	A	Overall responsibility of underground operations	1
Senior Engineer	B	Responsible for all planning / scheduling, reserve estimates, and annual budgets	1
Mining Engineers	C	Shorter term planning and scheduling, drill and blast, vent etc.	3
Sen Geotechnical Engineer	B	Responsible for GCMP and compliance	1
Technicians	D	Prod reporting, vent monitoring etc.	8
UG Foreman	B	UG supervision / management - for each shift	6
UG Workers	UG6	Building of vent barricades, clean drains etc.	15
Geology
Senior Geologist	B	Responsible for grade control, mineral resources and annual budgets	1
Geologists	C	Grade control, resource models	2
UG Geology Technicians	F	Sampling, core cutting / logging	9
Survey
Surveyor	C		3
UG Survey Technicians	E	Assist surveyors	7
Surface workshop / stores
Manager	A	Overall responsibility of underground equipment and annual maintenance budgets	1
Senior Engineer	B	Maintenance planning, spares ordering	1
Engineers / planners	C	Planning and supervision of fitters / electricians	3
Artisans	E	Fitters / electricians	10
Expert Trainers	Expat	On site trainers	4

16.5 Underground infrastructure

16.5.1 Underground access

The key underground access features are:

• A trucking portal centreline located at 402,053 mE, 100,062 mN, 173 mRL (floor level), at the north end of the Urucum North pit.

• A ventilation portal centreline located at 402,037 mE, 99,744 mN, 136 mRL (floor level), at the south end of the Urucum North pit.

• A decline system developed to the east of the lodes in the geological hangingwall that extends from approximately 173 mRL to -490 mRL. This splits into a twin decline system between 10 mRL and -390 mRL. The lateral coverage is approximately between 99,400 mN to 100,200 mN.

• Two vertical ventilation rises exhausting to surface located at the north end of the Urucum North pit at 402,030 mN, 100,200 mE, 210 mRL.

At this stage AMC has assumed a larger dedicated Personnel Carrier (or similar) will transport the underground workforce from surface to underground, together with smaller 4WDs. Materials will also be transported via the decline predominantly using Integrated Tool carrier (IT) type underground vehicles.

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16.5.2 Surface ventilation fans

Surface fans will comprise two 560 kW radial parallel fans. Fans have been based on airflow modelling.

16.5.3 Change rooms and lamp room

Three change rooms will be established for the following persons:

• Underground workforce (40 per shift).

• Managers and supervisors (10 per shift).

• Visitors (10 per shift).

Each change room will have the following facilities:

• Clean changing area including lockers.

• Dirty clothes changing area including lockers and clothes drying racks.

• Showers, toilets, and wash basins (the underground workforce will shower before proceeding home at the end of each shift.

Adjacent to the change room will be a lamp room capable of accommodating up to 200 lamps.

16.5.4 Underground mine offices

Offices will be established for the following functions and persons:

• Geology and survey

• Engineering and mine planning

• Operations control

• Electrical and mechanical engineering

• Management and supervision

The operations control will comprise of a mine clerk who will maintain all production reporting data, shift bosses (present during shift change), Health and Safety reporting, and mine communications.

This study assumes that the change rooms, lamp rooms, and offices will be rented for the duration of the project and will be delivered to site as trailer mountable / dismountable units. A lower cost alternative might be to find a local contractor to build these structures on site.

16.5.5 Surface workshop

Surface workshop will be used for regular servicing of all trucks and service vehicles (scissor lift, ITs, and light vehicles). In addition, the major maintenance of underground jumbos, production drills, and LHDs will also be undertaken at this workshop. The workshop will be a basic structure consisting of two stacked containers on either side of an arched undercover area. Any major overhauls requiring additional facilities can be undertaken at the existing open pit maintenance facilities.

Surface workshop will be constructed with the following facilities:

• 10 t overhead / gantry crane.

• Offices for supervisors.

• Toilets and wash basins.

• Workshop area to accommodate up to two trucks, LHDs, or jumbos.

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• Refuelling station and storage.

• Wash-down pad and facilities.

16.5.6 Mines rescue centre

A mines rescue centre will be constructed to accommodate:

• First aid room

• Mines rescue and training room

16.5.7 Mine portal

A portal will be established at the base of the Urucum North pit. The portal will be established in fresh rock with basic support (no steel sets or concrete anticipated).

16.5.8 Water supply

Water will be supplied from the open pit mining facilities and stored above the portal area from where it will be gravity fed through the decline to underground. AMC has allowed for four 48 kL poly tanks as water supply header tanks in its cost estimates.

16.5.9 Existing facilities

Both the existing open pit explosives magazine and warehouse facility will be used for the underground project.

16.5.10 Power supply and distribution

Medium voltage electrical power (e.g. 4,160 volt) will be supplied to the underground via both the declines using separate feeder circuits connected to the site main substation. Each of the two feeders will continue down the underground access declines.

As development progresses, the first of the feeders will be connected to a fused medium voltage switchgear unit located on the level where mining activity will occur. This medium voltage switchgear unit will supply power to a Power Distribution Centre (PDC) unit located on that same level. This PDC will step down the voltage to 660 volts and distribute power to any mining equipment, pumps, fans, small motors, and lighting demands necessary to permanently service that level and nearby sublevels.

The medium voltage switchgear unit will also supply up to six portable Mine Power Centres (MPC) located throughout the level. Each of the portable MPCs will be fed from the medium voltage switchgear via trailing cables and will step down the voltage to 660 volts to supply power to any mining equipment, pumps, fans, small motors, and lighting demands required on a temporary basis to develop and mine the level.

As development continues downward, an additional feeder cable will be installed with a permanently placed medium voltage switchgear unit and a PDC unit installed on main levels. With both ramps being developed at the same time, this equipment could be powered from either ramp.

When the lower interconnected level is reached, a medium voltage switchgear unit and PDC unit will be placed along with two tie switches that connect the two independent power feeds coming down each ramp. This will allow power to be supplied via either feed.

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16.5.11 Water supply and distribution

The mine has an existing fresh water system set up for open pit mining and processing operations.

Underground requirements are estimated to peak at around 7.6 l/s assuming a production rate of around 580 ktpa and would be met by 4” steel service piping down the declines with occasional pressure reducing valves.

16.5.12 Dewatering

The Urucum North underground dewatering system is based on a series of sumps and submersible pumps along the decline. Each of the sumps is 80 m apart vertically. Each sump discharges into the next in series until discharging out the portal into the open pit dewatering system.

Intermediate sumps on 20 m levels drain to each major sump below via drain-holes.

16.5.12.1 Pump selection

Given the low ground water inflows (expected), pumps have been selected that can generate a carrying velocity in the 6” dewatering lines. This provides a capacity that is in excess of any peak drilling water consumption. Sumps are sized to provide a 30 min on cycle for the design flow rate.

All pumps would be the same model and would be readily interchangeable. Space would be provided in the sumps for inline spare capacity. As well, the system can be quickly expanded if additional dewatering requirements are encountered during development, by installing additional pumps.

The pump selected is a submersible Weir Hazleton SHW, more commonly known as a “Warman” style centrifugal submersible slurry pump, with a nominally 50 hp motor.

16.5.12.2 Sump design

Sumps are split into a clean and dirty side. The dirty side is accessible from the ramp with mobile equipment in order to allow for cleaning if required. The clean side and dirty side are separated with a low weir.

To ensure mud build up on the dirty water side does not become excessive, AMC recommends LHDs muck one to two buckets on each shift (depending on rate of mud build up) and the small quantities of mud mixed in with development waste (or ore if required) for haulage to surface. The dirty side must never be allowed to build up to the extent where a major mud clean-up is required (major clean ups can result in LHD damage when the water / mud levels are above the LHD axles).

A submersible pump is hung from chain hoist and monorail to allow it to be removed and serviced. Bypass valves would be installed at the base of the weir to allow the two sides to be connected for maintenance work.

Figure 16.21 shows a section of a proposed sump layout.

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Figure 16.21 Section of proposed sump layout

Source: AMC 2016 PFS

The sump is intended to have a clean water capacity of 50 m³.

16.5.12.3 Dewatering system

Pumping estimates have been based on a pumping rate of 10 l/s, for both the north and south declines. This rate is based on both ground water inflows and manmade water (drilling / flushing water, washing down, etc.).

Eight sumps would be required along the north decline and six stations would be required along the south. Depending upon the mine schedule, some sumps could have drain holes that allow sumps to be bypassed as the mine development proceeds, although this would need to be confirmed during more detailed future work.

Mine dewatering will take place along the decline through six-inch standard steel lines using quick-couplings. Pumps would be connected to the dewatering line via flexible connectors. Backflow checks are required downstream of each pump.

As the decline is advanced the dewatering lines would be hung and used during development. As each sump is excavated and constructed, the dewatering lines remain in place. Discharge would take place out the portal into the open pit dewatering system.

16.5.13 Compressed air

Compressed air will be supplied by local area portable compressors.

The portable compressors will be electrically powered using the same jumbo plugs and jumbo boxes as other mobile equipment, such as the jumbos and bolters. AMC recommends that two compressors be available on site to accommodate the following potential demands:

• Meet the elevated pressure requirements for any ITH drill activities.

• Powering mechanized raise climbers (if required).

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• Miscellaneous activities such as spot bolting with airlegs, powering air tools.

All mobile drilling equipment, including jumbos, long-hole drills, and bolters will be equipped with on board compressors.

16.5.14 Emergency egress

The emergency egress will be developed in 40 m sections, connecting to every second level. AMC has assumed this will be undertaken using contract airleg miners; however, raise climbers such as alimaks may also be used as an alternative. Lateral development between the crosscut and the emergency egress has been assumed to be developed using development jumbos; however, short handheld development should also be considered if the raises are mined using airlegs.

Once the airleg rise is completed, a modular escapeway system will be installed between the two levels. The system will be extended as each second level crosscut is developed. The modular escapeway ladders are fully caged, with rest landing spaced at 6 m intervals, while inclined at 70º or steeper.

16.5.15 Refuge chambers

AMC has made allowance for four 12-person refuge chambers for use during the development phase of mine as the escapeway system is being established or where a second means of egress cannot be provided. Figure 16.22 shows a typical example of the proposed chamber.

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Figure 16.22 MineArc refuge chambers

  

Source: AMC 2016 PFS

16.5.16 Underground communications

In order to keep capital costs to a minimum, AMC has assumed a simple and cost effective leaky feeder communication system will be installed underground. This system is based on 2-way radios mounted on every piece of mobile equipment and in strategic locations.

16.5.17 Future data management systems

More expensive fiber-optic data systems today are able to control the following:

• Secondary ventilation fans (controlled from surface).

• Dewatering pumps (controlled from surface).

• Underground management reporting (production and development statistics, etc.).

• Underground communications.

• Surface fans.

• Production and development blasting.

Such a system should be considered once initial mining trials prove to be affective and would be based on a network of fiber-optic cables between stations. Information collected from equipment sensors, tele-communication terminals, and voice would be transferred to the fiber-optic facility and routed to surface, as well as to other points within the mine. The benefits of such a system are reduced ventilation costs by running ventilation based on demand and better project management using up to date underground data such as shift reports, etc.

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16.5.18 Underground service bay

Daily maintenance of Jumbos, LHDs, and production drills will be completed either at the face, or in an underground service bay located close to mining activities. Trucks and service vehicles capable of travelling at higher speeds will be serviced on surface.

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17 Recovery methods

17.1 Summary

The existing Tucano Mill consists of a jaw crusher feeding a single-stage, semi-autogenous grinding mill (SSAG), with the ore then being treated in a CIL plant. GPR is currently planning on mining and treating approximately 3.6 Mtpa of ore through the process plant, elution, and electrowinning facility to produce around 140,000 to 150,000 ounces of gold per year. The plant treats an ore blend of 80% oxide and 20% sulphide. The Tucano Expansion Project is currently in construction and is discussed in Section 13. Plant throughput will be maintained at 3.6 Mtpa. The expanded plant is described as follows:

17.2 Crushing and ore stockpiling

Ore is delivered to a run-of-mine (ROM) pad where ore of oxide and sulphide are stockpiled separately on fingers of specified grade ranges. Ore is fed to the crusher hopper by means of a front-end loader to achieve the planned sulphide / oxide blend and grade for the day. A vibrating grizzly allows the finer, sticky oxide ore to bypass the crusher. Crushing is achieved using a 1,220 mm x 1,070 mm single toggle jaw crusher with a closed side setting (CSS) of 125 mm. In addition to the jaw crusher, a separate Spent Ore Feeder is located in the emergency stockpile area to facilitate the feeding of Spent Ore direct to the SAG mill. A surge bin with 30 min residence time and a diversion chute to an emergency stockpile feeder is located halfway between the crusher and the SAG mill.

17.3 Milling and classification

Grinding is achieved with a SSAG mill with a 7 MW single-pinion drive, a diameter of 7.32 m and an effective grinding length (EGL) of 7.95 m. Cyclones operate in closed circuit with the mill to produce a product size which is currently coarser than 80% < 105 μm. After construction of the 6 MW ball mill, the grinding circuit will achieve a finer grind of 80% < 75 μm.

17.4 CIL

Cyclone overflow from the grinding circuit is screened in the trash screen and is then pumped to the CIL circuit. The CIL circuit consists of six 2,650 m³ (live volume) carbon contactors, located south of the reagent storage area on the area formerly occupied by the rich solution pond from the heap leach operation. A seventh tank will be installed as part of the current mill expansion. It will not be fitted with a screen and will operate as leach tank only, ahead of the existing CIL.

The six existing CIL tanks operate in series. Launders connect the tanks to allow slurry to flow by gravity through the tank train. All tanks are fitted with bypass launders to allow any tank to be removed from service for agitator or screen maintenance.

Each tank is fitted with a 110 kW mechanical agitator to ensure uniform mixing. Air is currently sparged into each leach tank to provide adequate oxygen for gold dissolution. An oxygen plant will be constructed during the expansion to provide increased recoveries at the higher proportion of sulphide material in the ore blend.

Each CIL tank is fitted with a single 12 m² mechanically-swept, wedge-wire inter-tank screen to retain carbon. The style of inter-tank screen selected is known as a “pump screen”, which allows each of the CIL tanks to be installed at the same elevation, in contrast to the stepped arrangement used for conventional cascade carbon-in-pulp (CIP) plants. This simplified the civil engineering requirements of the CIL area. Barren carbon enters the adsorption circuit at the last tank, CIL Tank 6, and is advanced counter-currently to the slurry flow by moving slurry and carbon from CIL Tank 6 to CIL Tank 5 using recessed impeller carbon-transfer pumps. The inter-stage screen in CIL Tank 5 retains the carbon and the slurry flows by gravity back to CIL Tank 6. This counter-current process is repeated until the carbon reaches CIL Tank 1. A single recessed-impeller pump is used to transfer slurry to a loaded-carbon recovery screen. The carbon reporting as screen oversize flows to the carbon transfer column and the screen undersize returns to CIL Tank 1.

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The loaded carbon is washed with process water in the carbon transfer column to remove slimes prior to the elution circuit. An eductor, using raw water, is used to transfer the washed carbon to the existing desorption circuit. Barren carbon returning to the adsorption circuit from the carbon regeneration kiln is screened over a vibrating screen to remove fine carbon which is then collected and stockpiled for processing off site. The sized and regenerated carbon reports directly to CIL Tank 6.

Sodium cyanide (NaCN) solution is metered into Leach Tank No. 1 and CIL Tanks 1 and 2 via a ring main system. Final pH adjustments are made by the addition of lime to the CIL feed box from a lime slurry ring main system. Tailings slurry from the last CIL tank flows to the vibrating carbon safety screen to recover any carbon that may be present due to damage, wear, or incorrect installation of the final stage inter-stage screen. Carbon recovered on the screen is delivered to a bulk bag for re-use. Tailings discharging from the carbon safety screen underflow are transferred via gravity to the cyanide detoxification circuit. The leach and adsorption area is provided with three spillage pumps, which deliver any spillage in the area back to the process. Maintenance of the CIL tanks and agitators is achieved using an overhead crane.

The new leach tank and the six CIL tanks, give the combined CIL tanks a total residence time of just over 24 hrs at the expected slurry density.

17.5 Tailings

Tailings are deposited in the West Pond area to the north of the plant. The use of the West Pond involves pumping the tailings about 3.6 km and the return water initially about 2.7 km and later up to 5.5 km. An air / SO₂ cyanide destruction circuit, using sodium metabisulfite (SMBS), lime slurry, and copper sulphate (Cu₂SO4), and consisting of two tanks in series with a total residence time of 60 min, lowers the cyanide concentration before discharging to the West Pond TSF. The detoxified tailings will leave the circuit at approximately 38% solids. Water is recovered from the TSF and pumped back to the process plant for consumption or neutralization and discharge. Natural degradation and dilution in the dam reduces the cyanide level further in the return water to approximately 10 ppm total cyanide. The majority of this water is re-used in the process plant. However, as the site water balance is positive, a portion of this water needs to be discharged to Igarapé William (William Creek).

17.6 Elution

The existing Pressure Zadra elution circuit and gold room were already in place when BDR acquired the plant from the previous heap leach operation and the CIL plant was configured to use these facilities. The elution columns are primed with a solution of 0.1% cyanide and 1% caustic soda that is prepared in the electrowinning solution tank. This solution is recirculated through the elution columns at 135°C to promote the elution of gold ions that are adsorbed in the micro pores of the loaded carbon. The heating of the recirculating solution is provided by a 720 kW electrical elution heater. The current elution circuit is sized to treat 4.8 tonne of carbon per day.

17.7 Carbon regeneration

Once the acid wash cycle is complete, the carbon is educted to the kiln feed hopper, where the carbon is screw-fed to the kiln dewatering screen, to dewater the carbon to 50% w/w solids. The undersize of the kiln dewatering screen is brought to grade, to be collected by the acid wash area sump pump. The kiln dewatering screen oversize reports to the kiln feed hopper. Carbon is withdrawn from the kiln feed hopper by a screw feeder and fed to the carbon regeneration kiln. Carbon regeneration is achieved at a temperature of 700°C. After regeneration, the barren carbon is discharged into the carbon quench tank to reduce the temperature of the carbon to room temperature. The barren carbon is then transferred back to the CIL circuit. The carbon regeneration kiln acquired from the previous heap leach operation has been extensively refurbished by BDR.

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17.8 Electrowinning and gold recovery

The rich solution leaving the elution columns passes through two heat exchangers, arranged in series, to reduce the solution temperature, then through a cyclone separator to the electrowinning feed tank. The solution flows from the electrowinning feed tank through a bank of two electrowinning cells, where the gold in the solution is deposited on the cathodes within the cells. The weak solution leaves the electrowinning cells and returns to the electrowinning solution tank. This solution is then pumped back through the elution heater to the elution columns. The solution is recirculated through this closed circuit for 12 hrs, until the disposition of gold onto the cathodes is complete. At the end of the elution cycle, the carbon is transferred to the acid wash column. The barren eluate solution is pumped from the electrowinning solution tank back to the CIL circuit.

The acid wash cycle consists of a six-hour acid wash, in which 15% hydrochloric acid (HCl) is recirculated through the column, followed by a two-hour water wash. The HCl is delivered to the acid wash column by a dedicated acid wash HCl tank and pump set, located in the desorption area. The 15% solution is prepared by the addition of 33% HCl from the main acid storage tank, with dilution provided by raw water. Wash water is added to the acid wash tank from the main raw water supply. The original design of the elution circuit included a four-hour neutralization cycle using sodium hydroxide (NaOH). However, this step was deemed to be unnecessary during the operation of the heap leach plant and was decommissioned.

The gold room was retained from the previous heap leach operation existing operation and consists of a bank of two electrowinning cells. During the elution process, the stainless-steel wool cathodes within the cells become loaded with gold. At the completion of the elution cycle, the sludge containing the gold, which is weakly adhered to the cathodes, is dislodged with a high-pressure water spray. The sludge is dried in a drying oven before being direct smelted with fluxes in a furnace. There are two operating furnaces that were retained from the previous heap leach operation, a gas-fired furnace and an electric furnace. The doré metal and slag separate in the furnace, the slag is poured off to slag pots and the doré metal is poured into bars. Doré bars are cleaned, weighed, stamped, sampled, and then stored in a safe contained within a strong room, while awaiting dispatch.

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18 Project infrastructure

18.1 Context and regional infrastructure

The Tucano site is located approximately 120 km Northwest from Macapá, the capital of Amapá state, in the Northern coast of Brazil.

The site is relatively remote and, for the development of the Tucano property, GPR and its predecessors took advantage of the existing infrastructure built in the late 1950s for the Serra do Navio manganese operations, 15 km Northwest from the Tucano site, and its upgrades, mainly established during the 1970s.

The main existing infrastructure in the vicinity of the Project consists of the towns at Serra do Navio and Pedra Branca do Amapari, the road from Macapá to Serra do Navio, the hydropower facilities and transmission line from the Coaracy Nunes complex in the Araguari river, approximately 70 km to the West from the Tucano site, the railway from Macapá do Serra do Navio, now inactive, and the port at Santana.

Macapá and the Santana port, approximately 10 km to the Southwest of Macapá, are located in the North canal of the Amazon river. The Santana port can accommodate relatively large ships, as it was built for bulk materials handling and has been in use until recently to handle iron ore from the Zamin property, adjacent to the Tucano site, to the Southeast.

Macapá is 330 km Northwest from Belém, the capital of the state of Pará, in the South canal of the Amazon river. Belém has approximately 1.5 million inhabitants and has a fairly developed industry. It is well connected to the industrial regions in Brazil Southeast and the main region’s hub.

Most of the industrial supplies and bulk materials required for Macapá and the Project are sourced in Southeast Brazil, transported via roads to Belém, where trucks are ferried to Macapá, disembarking in the Santana port.

Macapá has a population estimated as of approximately 474,000 inhabitants. The state of Amapá, in the North margin of the Amazon river is considerably less developed than its southern neighbour, Pará. Most of the industrial development of the state was first driven by a number of mining and agro-industrial projects established since late 1950s.

There are two small towns near the Tucano site that house most of the Project employees and their families: Serra do Navio, established as a mine camp to accommodate the manganese operations personnel and their families in the late 1950s, with approximately 5,100 inhabitants and Pedra Branca do Amapari, 30 km via road from the Project to the Southeast, with a population of approximately 13,000 people.

The Coaracy Nunes hydropower facilities at the Araguari river have a capacity up to 78 MW and are operated by the state-owned CEA, controlled by Eletronorte. The facilities were developed in the mid-1970s to support the Serra do Navio and Jari projects providing a reliable and affordable alternative to diesel generated energy for the state of Amapá, which has no links to the remainder of the national grid in Brazil. The 69 kV transmission line from the Coaracy Nunes complex to Serra do Navio is approximately 70 km long and was installed in the South margin of the Araguari river to its junction to the Amapari river and then follows the railway path to Serra do Navio in the South margin of the Amapari river.

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18.2 Site access and logistics

The mine site is located approximately 200 km from Macapá. Access to Serra do Navio via road is well established since the implementation of the manganese operations in the late 1950s and, later, the upgrade for the construction of BR-210, Rodovia Perimetral Norte in the mid-1970s. Open pit mining operations at Tucano commenced in 2012, when a dedicated road was upgraded from Pedra Branca do Amapari to the Mine. This access comprises:

• Sealed road for approximately 100 km, from the port town of Macapá towards Porto Grande.

• Unsealed road from Porto Grande to Pedra Branca do Amapari, approximately 70 km.

• Unsealed road from Pedra Branca do Amapari to the site, developed specifically for the Project for approximately 30 km.

The journey from Macapá takes approximately four hours by car during the dry season and up to six hours during the wet season, depending on the condition of the unsealed roads. The unsealed roads are passable all year around and no blockages are experimented, except for the road maintenance.

The main concern for the transport of personnel, bulk materials, parts, and other supplies is the existence of various bridges along the route, mostly built using wood as primary construction material and whose capacity is limited to 45 t of total load (carriers plus the deadweight). Most of the bridges have been upgraded recently, allowing for a reliable means of transport to the site. The load limitations have not been a problem for the mine operation or the plant upgrade, as most of the equipment parts to be erected are lighter than 20 t to 25 t, average deadweight transported per trip.

Diesel for the operation of the mining fleet and power generation is the main bulk material transported to the Mine, approximately 3 ML are consumed per month (equivalent to approximately 80 tpd) as of the first quarter of 2018. This demand would decrease by half once the power source is converted from diesel to hydropower. Lime for neutralization follows, with an average consumption of approximately 30 t per day, sourced from São José da Lapa, Minas Gerais. Cyanide, sourced from Bahia, is transported in briquettes and hydrated at the mine site at a rate as of 250 t per month. Balls for milling, representing approximately 150 t per month, and activated coal for the CIL process, at an approximate rate as of 24 t per month, complete the major bulk consumables. The average time spent by general cargo via road from the Southeast of Brazil to the mine site takes 12 days to be delivered, once consolidation takes place, totalling less than 20 days, including consolidation, while dedicated cargo takes 12 days to 15 days to be delivered at the Project.

A helipad was constructed near the CIL plant and is used to transport product and emergency removal of personnel. The Tucano mine site is also serviced by a 1,100 m airstrip located approximately 800 m from the main gate that serves as a backup. Charter flights from Macapá to the mine take about 50 min.

18.3 Mining and support facilities

The following facilities are located in proximity to the CIL Gold Plant:

• Central administration building and medical facility.

• Geology, survey, mining engineering, supply, environment, and safety offices.

• Core shed and exploration sample preparation facility.

• Nursery.

The mine is primarily operated with 250 t class excavators (Liebherr 9250s, Hitachi EX2500s), 100 t class trucks (Caterpillar 777s) and 40t articulated trucks, supported by 400 HP and 320 HP class bulldozers (Caterpillar D9s and D8s), 4” drill rigs and 16 feet class graders (Caterpillar 160), plus assorted support equipment.

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The heavy machinery and light vehicle workshops with adjacent warehouse storage facilities, tire shop, refuelling, and washing bays were constructed to high standards and were upgraded in 2012.

The mine planning, administrative, and messing facilities are also located in the workshop area together with the mine control and dispatch room.

An explosives magazine and accessories storage constructed to standards are located at convenient distance from the other industrial installations. More recently, a factory for producing emulsion explosives was built and is operated by the explosives supplier. The installations are fenced and surveyed by armed guards and cameras permanently.

18.4 Sub-contractor facilities

U&M will coordinate and undertake all heavy machinery maintenance at the main workshop alongside of the main supply warehouse. McKay Drilling, the geological drilling contractor, has a separate workshop (and storage sea container near the MACA workshop).

Figure 18.1 U&M workshop facilities

 

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18.5 Process plant

The processing plant and administration facilities consist of:

• Plant administration buildings.

• Plant workshop area.

• Power supply and distribution.

• Sample preparation and assay laboratory’s.

• Storage area for plant supplies and consumables.

• Reagents preparation facilities.

• Security.

18.6 Water

18.6.1 Raw water

Raw water consists of fresh water that has not been used in the process and does not contain process reagents or salts. Raw water is consumed at an average rate as of 180 m³/h, while existing permits allow for up to 450 m³/h of consumption.

Raw water is drawn from the water collection dam and pumped to the lined raw water dam (also known as the ETA dam). The raw water flows from the ETA dam to the existing raw water treatment plant. Lime and aluminium sulphate is added to the raw water before it enters two primary water treatment filters. The raw water then flows to one of two locations, to either the plant raw water tanks for consumption in the plant, or for further treatment for use as potable water.

From the total 180 m³/h flow extracted at Igarapé William, which is primary filtrated and conditioned, 100 m³/h are directly used to refrigerate the milling equipment, part of this is evaporated and the balance is diverted back to the river. The balancing 80 m³/h flow, after primary treatment, is used at a rate as of 72 m³/h for washing and industrial purposes at the various installations (including the mine workshops) and as sealing water for the pumps at the plant, the remainder 8 m³/h are further treated to obtain potable water.

18.6.2 Potable water

Raw water that is used for human consumption is transferred from the primary treatment filters to three filtered water tanks at an average rate as of 8 m³/h. This water is pumped to the potable water treatment plant, where is further filtrated and chlorinated to produce water of drinkable quality. The water is transferred from the potable water treatment plant into two potable water tanks. Two pumps (one duty, one standby) are used for distribution of potable water to end users.

18.6.3 Process water

Process water consists of recycled and other waste streams used in the process, including return water from the TSF.

The plant consumes 550 m³/h of process water, being 100 m³/h used in the mill feed, 450 m³/h at the mill discharge and further 100 m³/h for the preparation of reagents.

Process water is stored and sourced from a dedicated process water dam. Water is pumped from the process water dam by four pumps (three duty, one standby) to the following users:

• Grinding area water - mill trommel screens.

• CIL area water - screen spray water and carbon transfer column eductor water.

• Crushing area water - dust suppression sprays.

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Overflow from the process water dam flows by gravity to the detoxification pond for cyanide destruction prior to discharge to the environment.

18.7 Power

The Tucano mine power requirements are provided by two sources: 1) a 69 kilovolt (kV), 20 megavolt-ampere (MVA) power line via CEA and 2) a local supply authority, and a 12 MW continuous rated Aggreko diesel-powered generation system. The current demand is approximately 12 MW. At the completion of the additional ball mill and plant upgrade, now in progress, it is expected an increase of the demand to 15 MW.

The Aggreko generator facilities have a nominal capacity as of 12 MW, meeting the current demand, while the CEA line is being expanded to meet the total power requirements.

The CEA line capacity is currently limited to 20 MVA due to CEA distribution transformer limitations from 138 kV to 69 kV. A progressive strategy for the upgrade of the power line and transformers of the CEA line from 1.8 MW (prior to 2017) to 12 MW (already contracted) and, possibly, 15 MW (the final demand after the plant upgrade) has been implemented since 2017. The 70 km long 69 kV transmission line from the Coaracy Nunes facilities to the Mine has been continually upgraded and maintained and the capacity as of the first quarter of 2018 is 5 MW, resulting in significant increasing reliability of the system and substantial savings on diesel costs. A 30 MVA transformer from 138 kV to 69 kV to support the expansion of the capacity of the CEA line is in the late stages of procurement.

18.8 Fire protection

There is a fire-fighting system next to the fuel storage and dispensing area and another one on the outdoor premises of the warehouse area. In both cases, the systems consisted of a tank and three fire pumps. A provision for additional local fire-fighting equipment near the oxygen plant location has been made.

18.9 Communications

There are three rented interconnected private branch exchanges (PBXs). There are two Enterprise PCX services at the Tucano mine site with 300 extensions; one in the plant / administration area and the other at the residential camp. The third Enterprise PCX is at the Rio de Janeiro office and consists of one Embratel Link with 30 lines and up to 100 extensions.

• There is a satellite communication internet link and two internet radio connections to Macapá.

• At the Macapá office there is a fiber-optic link to Compuservice Webflash.

• At Rio de Janeiro there are two redundant fiber-optic links.

• There are 6 VPN site to site connections between the Rio de Janeiro office and the Tucano Mine, using the two Rio links and the three Tucano links.

• There are 3 VPN site to site connections with the Macapá Office to Rio de Janeiro and the Tucano Mine.

• Fiber-optic is available, with radio redundancy, between the mine areas, accommodation and industrial areas of the mine.

• Network equipment based on Fast Ethernet and Giga Ethernet LAN and WIFI.

Light vehicles and mine equipment are equipped with mobile radios, and there are portable radios for operations personnel. There currently are 165 vehicle and equipment mounted radios, 206 hand portable radios, and 8 repeaters on the site.

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18.10 Security

The main entrance security office is located at the Main Site access road. All visitors to the mine complex report to this security gate for authorization prior to entry. Additional fencing is provided for further safety and security within process plant areas, such as the gold room area, transformers and substations. A closed-circuit television system is installed in the plant and gold room to monitor activities.

18.11 Personnel and personnel logistics

The mine currently employs a total of 680 staff on its books and 522 contractors.

The personnel living in Amapá accounts for approximately 65% of the workforce. Most of the workforce live in towns near the mine site with transport to Serra do Navio and Pedra Branca do Amapari provided by GPR in a daily basis. Personnel from Macapá also has transport provided by buses in a weekly basis. Most staff are Brazilian nationals, commuting to the mine. A few senior management staff and some expatriates working for contract companies MACA and McKay commute from overseas.

Managers and professional staff either commute from the state capital Macapá or from other cities in Brazil on a FIFO basis. Shift workers work a rotating 8 hr shift over a 24 hr period with a roster in days of 6 on 1 off, 6 on 2 off, 6 on 3 off basis. Managers and professional staff either work a roster in days of 19 on, 9 off or 5 on, 2 off. Expats work on a FIFO basis of 6 weeks on, 3 weeks off.

18.12 Accommodation and medical care

There is an 85 person accommodation camp located to the east of the old leach pads, near the access road to the site. Senior staff uses the cabins in blocks of four in the upper part of the camp. Technical staff is accommodated in a longer barracks-style building downhill from the cabins. The rooms in the accommodation camp are air-conditioned and equipped with a refrigerator, writing desk, television, bed, bedside table and lamp, microwave oven, chair, en suite shower, toilet, and cupboards for clothing.

Communications for personnel purposes are provided at the lodging via telephone and wi-fi connections.

The camp also has a well-equipped gymnasium, games room, laundry, and restaurant area for ordering meals.

Medical facilities are close to the administrative offices and a medical doctor and nurses are permanently at the mine site to provide first care and treatment to GPR employees and contractors. There is provision for emergency removal of personnel either by ambulance, a mobile intensive care unit, and helicopter or plane, if needed.

18.13 Tailings Storage Facility (TSF)

18.13.1 Tailings operation

The Tucano Mine plant produces an average of 3.9 Mtpa of tailings (dry basis) from the gold beneficiation process; the tailings slurry is delivered to the TSFs via a fixed pipeline and is subaerially discharged into the various TSFs as a slurry at around 30% solids concentration (w/w). Tailings settle out in the TSFs to an average dry density of 1.2 t/m³, and supernatant water is removed from the pond via floating vertical pumps for recycle back into the plant circuit.

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A water balance of the LOM facility prepared by SRK (project BDR020 Water Balance Update) was used to determine spillway capacity, water reclaim availability, and storm management controls. The conclusions of this report indicated that on average, the plant water demand is met (no deficits in water availability) throughout the LOM. It is expected that the facilities will spill for post-closure conditions.

The LOM tailings production remaining, from May 2018 up to September 2025, is expected to be approximately 34.0 Mt, or a volume of 28.3 Mm³.

SRK has audited these facilities since commissioning of the TSFs, with a focus on safety and operational capacity of the embankments, pond size and ongoing construction of the facilities. The following audit site visits have been undertaken by SRK:

• 6 January 2014 - 7 January 2014: Mr Juan Moreno

• 16 November 2014 - 18 November 2014: Mr Juan Moreno and Mr. Nicholas Thompson

• 11 November 2015 - 16 November 2015: Mr Juan Moreno

• 19 October 2016 - 21 October 2016: Mr Juan Moreno

• 12 March 2017 - 15 March 2017: Mr Matheus Rocha

• 18 April 2018 - 21 April 2018: Mr Rubens Rocha.

18.13.2 Description of the constructed TSFs

The Tucano TSF system consists of three facilities currently in operating conditions: Tap D, North Mill Pond (NMP) and West Pond Phase 1 (WPP1). A summary of the storage capacity of these facilities is provided in Table 18.1.

The remaining capacity of the current facilities is approximately 3.5 Mm³, which covers the storage for 12.0 months, or from October 2018 to October 2019.

GPR plans to extend the tailings storage capacity by commissioning two new TSFs, Barragem Leste and West Pond Phase 2 (WPP2).

Table 18.1 Summary of TSF storage capacity

Facility	Tailings volume stored (m3)	Remaining storage capacity (m3)
Tap D	6,200,000	Nil
NMP	3,328,000	2,200,000
WPP1	3,900,000	1,250,000
Total	13,428,00	3,450,000

18.13.2.1 Tap D TSF

The Tap D TSF integrates three embankments and a small saddle wall that close a small valley just east of the process plant. The TSF includes two pits within its basin; namely Tap D1 and D2. SRK designed and supervised the construction of this facility.

The walls were named Wall 1, Wall 2, Wall 3, and Wall 4; these are shown in Figure 18.2. The design of these walls was provided in the Design Report (SRK projects BDR001 and BDR008). All the embankments are built with homogenous fills of low permeable material. The borrow material for the walls is a silty sand to sandy silt residual soil. A typical schematic cross section of these walls is shown in Figure 18.3.

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Figure 18.2 Tap D embankments

 

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Figure 18.3 Typical wall cross section

Active deposition in the Tap D TSF ceased in 2014 and it is since then dormant. The facility has not yet reached capacity and GPR has strategically opted to retain it as a contingency for emergency deposition, as it is located downstream of the mill. The Tap D TSF has received intermittent tailings discharge on a number of occasions, but it has been reported as minimal. The remaining capacity is not significant; however, some contingency is available since the current freeboard is significantly larger than the required (+3 m) and floating pumps are still operating in the basin.

18.13.2.2 North Mill Pond TSF

The NMP TSF is a valley-type impoundment bounded by the West Pond Phase 1 (WPP1) TSF to the north and Tap D TSF to the south. The NMP facility was intended to act as a temporary measure while alternative facilities are commissioned in order to service the remaining LOM. SRK designed and provided partial supervision of the construction of this facility.

NMP TSF is composed of Wall 17 to RL 135 m with three additional minor compacted fill embankments (Walls 16, 18, and 19) required to be built to RL 135 m in saddles to enclose the site. A layout of the NMP site is presented in Figure 18.4.

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Figure 18.4 NMP embankments

 

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18.13.2.3 WPP1 TSF

The WPP1 TSF integrates five walls that close a small valley approximately 1.5 km north east of the process plant. SRK designed and provided partial supervision of the construction of this facility.

The walls were named Wall 11, Wall 12, Wall 13, Wall 14, and Wall 15; these are shown in Figure 18.5, while the height of each wall is reported in Table 18.1. The design of these walls was provided in the Design Report for SRK project BDR005.

Figure 18.5 WPP1 embankments

 

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The highest embankment is Wall 1, which is 33 m high, with the remaining walls being less than 15 m high.

All the embankments are designed as homogenous fills of low permeability material, with slope gradients at 2.5H:1V and 5 m wide crests. A downstream drainage and seepage collection system for long-term seepage control has been considered for Walls 11 and 12. A typical cross section for Wall 11 is presented in Figure 18.6 for reference.

Figure 18.6 Wall 11 cross section

18.13.2.4 Integrated WPP1 and NMP TSFs

The WPP1 and NMP facilities were raised to integrate them into a single facility, the corresponding embankments crests were initially at RL 135 m and RL 130 m, respectively. These embankments were downstream, raised using a combination of rockfill and lateritic borrow material to RL 137 m, which allowed to raise the tailings level to a maximum elevation of RL 136 m. The spillway for the facility was relocated to Wall 16, as shown in Figure 18.7. The integrated facility was designed by SRK.

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Figure 18.7 Integrated WPP1 and NMP TSFs

  

18.13.3 Extended tailings storage capacity

The current built TSF capacity is enough to store tailings up to October 2019. Additional tailings storage capacity is planned through commissioning two new facilities, Barragem Leste and WPP2, which would cover the tailings storage demand up to September 2025.

GPR is planning to first commission Barragem Leste, while working on the development of the larger WPP2 facility. Figure 18.8 presents the proposed TSFs.

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Figure 18.8 Proposed TSF expansion

 

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18.13.3.1 Barragem Leste TSF

The Barragem Leste TSF is a valley-type impoundment bounded by the NMP TSF to the south; this facility will be constructed during the 2018 dry season (July to December) and will be commissioned in July 2019. The capacity of this facility is 4.9 Mm³, which would add 18 months of storage, up to December 2020. This facility has been designed by GeoHydroTech Engenharia to a conceptual level, and it is expected that the construction design will be ready in 2018.

The Barragem Leste TSF is composed of Walls 21 and 22 to a crest elevation of RL 137 m, these walls will be constructed as compacted low permeability homogeneous embankments. A typical cross section of the planned Walls is presented in Figure 18.9.

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Figure 18.9 Typical cross section of Walls 21 and 22a

 

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18.13.3.2 WPP2 TSF

The West Pond (WP) TSF is a valley-type impoundment that was designed to be developed in two phases, WPP1 (approximately 35 ha) and WPP2 (approximately 167 ha), located to the east and west of the valley respectively. WPP1 has been constructed and is currently active. SRK has designed this facility to a detailed level.

In total, nine earth embankments (Wall 1 to Wall 9) were designed to isolate the WPP2 catchment and limit the spread of the tailings. A downstream drainage and seepage collection system for long-term seepage control was considered for Wall 1 to Wall 8. Wall 9 is a temporary structure that allows construction of WPP2 to be staged, if required. Figure 18.10 presents a layout of WP TSFs.

Figure 18.10 WP TSF layout

 

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The storage capacity obtained after construction of the WPP2 is 23.7 Mm³, or 28.0 Mt. This TSF adds approximately 7 years of operational TSF capacity.

18.13.4 Permitting of the TSFs

Operation of TSFs in Brazil requires a license to operate issued by the Instituto de Meio Ambiente and Regulacion Territorial, and a deforestation authorization (ASV) to implement the construction of the facilities.

The current operating TSFs have been granted license to operate (Liçença de Operaçao 0223/2015) from November 2015 to November 2021 (six years).

The integrated NMP and WPP1, has been permitted via Oficio 868/2017 by the Insitituto de Meio Ambiente and Regulacion Territorial, within the same operational license.

IMAP license and an ASV have been granted to permit construction of East pond (Barragem Leste). Construction of the East pond has begun.

Construction and operation of the WPP2 TSF requires a renewal of the current ASV. GPR has advised that the application for a new IMAP for this TSF is underway.

18.13.5 Conclusions

There are three operating TSFs in the Tucano Mine, these are Tap D, NMP, and WPP1. The constructed storage capacity for tailings is 3.5 Mm³, which covers the storage for 15.7 months, or from October 2018 to October 2019. These facilities have been designed and audited by SRK since 2014 to date.

Further storage capacity is to be implemented by constructing two new TSFs, Barragem Leste and WPP2. These two facilities would provide 4.9 Mm³ and 23.7 Mm³ of tailings storage respectively, sufficient to cover the Tucano Mine LOM production to up to September 2025 at an average tailings production of 3.9 Mtpa.

18.14 Underground infrastructure

Underground infrastructure is discussed in Section 16.5.

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19 Market studies and contracts

19.1 Markets

Gold output from the Tucano Mine operation is in the form of doré bars containing approximately 85% gold and 15% silver, the balance being copper and other minor metals. Silver credits are received from the refiner. The doré is shipped to Sao Paulo for refining. Umicore is the contract refiner and they retain Brinks to transport the bullion from Tucano to their facilities. Responsibility for the gold changes hands at the gold room gate. Based on actual contract agreements, the breakdown of costs is set out below:

• Refinery charge US$3.86 oz gold (US$3.86/oz doré).

• Transport, handling, and insurance US$ 17.50/oz gold.

• Total US$ 21.4/oz gold.

The costs above are subject to some variation depending of the amount of ounces per shipment, however, they are representative of the average cost.

Penalties are applied in the refining contract for any impurities, but these are rare occurrences and the penalties are not excessive, the main impact being on the delay in final payment due to extended refining periods. The current refining contract is valid until the end of August 2019.

The market for gold doré is well established. Market predictions and discussions for gold are beyond the scope of this document. The impacts of gold price volatility on the mine plan and process operation are well understood.

19.2 Contracts

As of 30 June 2017, the main contractors involved with the mine are:

• Mining load and haul: U&M.

— Mining Additional Equipment Support; JS Máquinas (excavators), Vale do Amazonas (excavators, haul trucks, dozers, ancilliary fleet), Masters (drilling), Norte Service (rock breaker).

— RC Drilling: Geosedna.

• Fuel supply: Ipiranga.

— Power Generators: Aggreko.

• Contract security staff: Prosegur.

• Catering and camp management: JS Catering.

• Offsite Laboratory: SGS GEOSOL,

• Contract personnel transport: Vale do Amazonas & Palmerim.

• Refining: Umicore.

The various contracts were awarded following a competitive bidding process, prices are within the industry range and comparable to other operations in Brazil.

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20 Environmental studies, permitting, and social or community impact

In 2011 SRK Consultores do Brasil Ltda produced an Environmental Control Plan (PCA) to present changes implemented to the project’s environmental controls as a condition for the renewal of the Operating Licence issued by the State Secretariat of Environment, Amapá (SEMA). This followed a period of care and maintenance from 2009 until BDR acquired the project. The PCA reviewed changes to the process plant configuration (from heap leach to CIL) and ways of minimizing the environmental impact of the plant up-grade. The PCA also covered the environmental aspects of mining and plans to monitor and manage these as well as implementing various programs for the local community.

20.1 Site visit

A site visit was conducted in March 2018 by Marcelo Batelochi the QP for Section 20. The site visit concentrated on the daily routine activities of the various operational sectors, concentrating on the monitoring of water quality as there is a creek crossing through the Project area (William Creek). In addition, the visit verified the monitoring requirements of Flora and Fauna, and the development of the native seedling nursery by BDR.

Air and noise monitoring comprises of mobile equipment that collects data in compliance with legal requirements and is performed by an independent company Análise Química Mineral, Ambiental e Industrial Mineral (ANQUIM). During the visit data collection was being carried out in the areas of the administrative offices, geology laboratory and carpentry workshop.

The monitoring of Fauna and Flora was not checked, but the activities of the field team, an independent company (MRW Ambiental), were observed.

With regard to community issues, a field trip with the community agents for BDR was performed from the land settlement lines southeast of the city of Pedra Branca do Amapari to the mechanized projects in the Nova Divisão community.

20.1.1 Groundwater and surface water monitoring field check

Groundwater and Surface Water Monitoring is performed through field checking of daily monitoring activities and collection of samples for chemical analysis at the site facilities. There are also monthly, quarterly and half-yearly sample collections as required by environmental agencies, which are required to be sent to a certified external laboratory.

During the field visit, eight monitoring points were visited as shown the Google Earth map in Figure 20.1.

The first point was a solids containment dam at Urucum North pit. It was visually verified that the dam efficiently traps the solids upstream and clean and clarified water flows downstream. The remaining points visited represented piezometer monitoring at the pit, tailings dam, and industrial area in compliance with Environmental Plans approved by the governmental agencies.

The points visited for Surface Water Monitoring were selected downstream of the Taperebá project operations to check the quality of water that is returning to the environment after the interaction with the mining project. Visually, the water appeared to be of very good quality, with low turbidity and normal temperature, about 25°C.

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Figure 20.1 Location map of field visit monitoring points

20.1.2 Socio-economics aspects field check

A field check of BDR’s socio-economic aspects was carried out by completing a visit to the town of Pedra Branca do Amapari municipality, where the community investments include popular Shopping infrastructure, and a Municipal Market. In addition, community investments in the form of the community roasted milled flour, and the result of the mechanization of the banana plantations were observed.

20.2 Known environmental liabilities

The Project area is an active mining operation centred in dense tropical rain forest located in Amapá State, Brazil, consisting of open pit operations, crushing, milling and CIL processing that has a direct impact on the environment and the ecosystem of the region. The main impacts are on biological resources, air quality, noise and vibration, archaeology, hydrology (ground and surface water), and community stakeholders.

20.3 Biological resources (Fauna and Flora)

The Tucano Project Flora is typical of an Amazon biome, identified as an area of Dense Ombrophylous Forest. The enormous biodiversity existing in this Amazonian ecosystem is favoured by the high structural complexity of the environment, with great vertical stratification, diversity of niches and microhabitats, providing a great number of species of Flora and Fauna.

The project area is not recognized as a priority for conservation, however, it should be noted that the mine is surrounded by areas of conservational interest. To the east and west is the State Forest of Amapá, at the northern limit is the National Forest of Amapá and to the southern extent is the Priority Area for Conservation of Biodiversity denominated Pedra Branca, that has a biological importance rating of Very High, due to the presence of certain sensitive areas in the ecosystem that include:

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• Preservation of species of economic importance.

• Maintenance of animal species (spider monkey).

• Wildlife Maintenance (MMA).

20.3.1 Flora

The vegetation is very well conserved, with reduced stretches in regeneration that correspond to the areas of old gold panning sites. In the regional analysis of the project, a secondary vegetation strip is observed that corresponds to the headquarters of the municipalities of Serra do Navio and Pedra Branca do Amapari, which creates a geographic barrier to the flow of Fauna throughout the western flank of the mine.

The Flora Conservation Program is basically the case of germplasm collection in which the collector operates in areas under anthropic impact (Walter & Cavalcanti, 2005). In Legislative Decree No. 02 of 08/02/94, genetic resources are defined as ‘genetic material of real or potential value for human beings’, and ‘genetic material’ means all material of plant, animal or microbial origin, or other, containing functional units of heredity. Therefore, genetic resources represent a small share of biological resources, which in turn are components of biodiversity.

Based on these definitions, the Tucano Project has been involved with biodiversity conservation working primarily at the level of ecosystems and communities, while in the conservation of genetic resources, research is carried out at the level of species, that are populations and pools of a numerically very large universe of species.

The suppression of primary vegetation of Dense Ombrophylous Forest contributes to the reduction of the genetic variability of populations of the species occurring in the locality. This program uses the premise that the populations and or species affected by the anthropization contain genotypes not found in the remaining populations or elsewhere in their distribution areas, so that there will be loss of genetic variability (Walter & Cavalcanti, 2005). Even though this loss is small, it can be minimized by a detailed rehabilitation plan.

The salvage of the germplasm is based on information constructed with the data of the floristic and phytosociological inventory already carried out, and the evaluation of the biological value of the species, germplasm recovery and conservation of plant genetic resources. Rehabilitation is through the implementation of the re-vegetation sub-program on degraded areas or of future rehabilitation with planting or re-seeding, as well as the provision of genetic material for rehabilitation of the mine area.

20.3.2 Fauna

The Fauna associated with the surroundings of the Tucano Project is rich and presents typical and endemic elements of Dense Ombrophilous Forest, as well as species of conservation interest such as Harpia, Myrmecophaga tridaetyla and Panthera onca.

The Fauna conservation program is based on the Normative Instruction (NI) 146/2007 of IBAMA (2007), that it is a guide instrument composed of procedures and regulations of field and laboratory activities, that undertake studies on the Fauna of regions under the influence of anthropization, providing accurate data to detect and quickly action to protect the environment.

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The BDR Fauna monitoring program has provided to the stakeholders, governmental agencies and scientific community the composition and structure of Faunal communities occurring in the region of the Tucano Gold Project, and the means to evaluate these populations in relation to the future activities of the company.

In summary, the monitoring program consists of:

• Logging the occurrence of species of mammals, amphibians, reptiles, birds, and fish on extended areas of the project influence.

• Obtaining data on the composition, wealth, and abundance of communities.

• Detection and evaluation of the interferences on faunistic communities.

• Development and continuous improvement of controls and management to mitigate impacts on the Faunal communities.

This continuous Faunal monitoring has been carried out by an independent consultancy company (MRW Ambiental), which has been reporting the behaviour of the Fauna in the seasons of intense rains and drier periods, periodically, to the legislated governmental agency.

20.3.3 Threatened and special status species

In order to reassess the conservation status of the Fauna described for the Tucano Gold Project, the list of endangered species contained in Article L. 411-1 of the Brazil Environmental Code (FR197/2010) strictly protects wild species of plants and animals listed by Ministerial Order. These species cannot be captured, transported, intentionally disturbed or commercially exploited. These prohibitions can extend to the destruction, degradation and alteration of the habitats of these protected species. A total of 110 nationally protected species were recorded on the Project site. Of these protected species, 100 were bird species, including three species with protected habitat, seven were mammalian species, and three plant species.

The site also hosts five new plant species and seven other plants of interest (rare or endemic), as well as two fish species rare and endemic, present in 1st / 2nd order creeks.

20.4 Air quality

Control of air quality in the Brazilian territory is managed by Brazilian laws (CONAMA No. 08/90) establishing national emission limits for external combustion power up to 70 MW (seventy megawatts), Ordinance MINTER-100/80, which provides for the emission of smoke by diesel-powered vehicles, CONAMA No 01/90 and 03/90, ABNT / NBR 10151, CONAMA No. 436/2011 - NBR 9547, 12019, and NBR 12827. This set of regulations manage the emission control of gaseous effluents, airborne particulates, noise and vibration, effectively promoting the systematic monitoring of concentrations of pollutants to be discharged into the environment (atmosphere) and minimizing the degree of risk caused in the environment, keeping their concentrations below the legal limits. BDR periodically submit the monitoring reports to the governmental agency, as well as being regularly audited.

The air quality in the Tucano Project is measured on the gaseous effluents from the exhaust discharge of the generator sets of the Thermoelectric Plant (UTE), smoke indices from the diesel powered heavy equipment and through control of total suspended particles. These effluents have come from the UTE, processing plant, administrative area and explosives magazine.

The main air pollutants emitted by the UTE are nitrogen oxides composed of nitric oxide (NO), nitrogen dioxide (NO₂) and carbon dioxide (CO₂). In smaller amounts, carbon monoxide (CO), hydrocarbons, smoke, and sulphur dioxide (SO₂) are also expelled.

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In a sample analysis of reports submitted to government agencies, the concentrations of NO, NO₂, CO, CO₂, O₂, and SO₂ show satisfactory indices below the reference limits of the equipment. Suspended total particle concentrations, have higher concentration indices for periods of low rainfall but are well below the acceptable maximum limit.

Smoke ratings on off-road mobile equipment and light vehicles have satisfactory rates of 20% or less (RINGELMANN comparative scale No.1).

20.5 Noise quality and vibrations

The main noise and vibration are related to the haulage vehicles, blasting and heavy equipment - fixed and mobile:

• Vibrations and noise caused by traffic and use of heavy equipment are relatively minor and can be considered negligible.

• The crushing system and beneficiation plant generates high noise / vibration levels that have been monitored in order to ensure that maximum levels acceptable within the legal limits are not exceeded. Several items of equipment have enclosed, or acoustic barriers installed in order to guarantee the noise levels remain below the limits of the legislation. In addition, periodic and corrective maintenance is done to keep noise within the limits set by manufacturers and specific legislation.

• Blasting has been limited to a scheduled specific time and these blasting schedules must be pre-defined and marked on visible notices at the entrance to the mine areas. The Project does not have communities in its immediate surroundings and the noise and vibration of blasting is negligible to neighbourhood communities.

Noise routinely monitored in these areas shows levels below 70 dBA. On site, the compulsory use of individual hearing protection is enforced.

20.6 Geotechnical

The geotechnical control and monitoring implemented at the Project consists of designing a set of sediment retention structures and slopes in the pit and the surrounding hills inside the Project area. The slopes are controlled by the Mine Planning and Mine Operation departments with specific controls developed by specialists in geotechnical engineering and monitoring of soil / rock potential displacement points. The sediments are controlled by collecting water run-off and directing it through surface drainage systems and internal drainage systems (in the case of waste and pits), thus allowing reduction or even elimination, of the solid material suspended in the water that is carried to existing waterways, primarily William Creek.

The main objective of these geotechnical controls is to ensure efficient prevention of erosion and prevent solids in the form of sediments flowing into the waterways and the environment and, consequently to minimize the impact on the surrounding non-anthropized areas.

20.7 Hydrogeology (groundwater), hydrology (surface water)

The hydrogeology and hydrology are critically related to the equatorial regions, where rainfall levels exceed 2,000 mm per year, and it is the most relevant element of the physical environment that impacts on the ecosystem. For this reason, the characterization and constant monitoring of groundwater, surface water and rainwater is essential to minimize the impact of the Tucano operations.

The Project implemented monitoring of the hydro system, since 1999, where surface sampling points and piezometers were geo-referenced and flagged. The routine sampling has been collecting quality base data (CONAMA 20/86 standard), with which current water quality sampling is compared (CONAMA 357/05). Sampling includes physical parameters of acidity, conductivity, temperature, dissolved oxygen, and turbidity. Samples are sent for laboratory analysis for metal content (anions and solids levels).

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In 2005, before starting operations, this monitoring system was extended to cover the area considered under the feasibility study including the waste dump, tailings, industrial plant, and buildings.

Currently, the underground and surface water-monitoring system is composed of 57 points. Due to adjustments on the Project infrastructure due to the dynamics of the operation, several piezometers and water surface points were unused, and others were installed in the area of coverage.

The results of monitoring and evaluation have been systematically submitted to the environmental agencies (SEMA / IMAP) as official Brazilian governmental reports that become public information for mitigation and minimization of the impacts of the dynamic local hydro system, as a result of the mine operation.

20.8 Cultural and archaeological resources

The archaeological site studies in the Brazilian territory are conditioned by the articles 20, 23, 215, and 216 of the Federal Constitution; Laws No 3.924, 26 July 1961; Ordinance SPHAN No 07, 10 December 1988; Ordinance IPHAN No 230, 17 December 2002; Normative Instruction IPHAN, No 001, 25 March 2015. In which the archaeological studies carried out by BDR at the Tucano Project fully complied with the Brazilian regulations.

20.8.1 Regional archaeological context

The State of Amapá has known archaeological evidences dating back to the sixteenth century, from scientific expeditions by naturalists, travellers, and ethnologists, among which is the stand out contribution from Crevaux, Coudreau, P Curt Nimuendajú, and Emilio Goeldi on the archaeological ceramics.

In the Tucano Project area, municipalities of Pedra Branca do Amapari, Serra do Navio, and Porto Grande there are 68 identified archaeological sites, most of them associated with historical and pre-colonial sites.

20.8.2 Archaeological studies at Tucano Project

Archaeological research was initiated in 2005 by Nunes Filho (2007 & 2009), when the project belonged to Wheaton River Minerals, and carried out in accordance with the Archaeological Program defined in the Environmental Impact Study, approved by SEMA (state environmental agency).

2006 and 2007, carried out several surveying, prospecting (with sample collection of artefacts), elaboration of surface diagnostics (for preliminary cultural characterization of each archaeological site), archaeological excavation and laboratory study, with the main objective of training a local team in archaeology and Patrimonial Education.

All the archaeological material collected during the activities was duly packed and transported to the Archaeological Laboratory of the Archaeological Studies and Research Center of Amapá - Cepap (Unifap), where they went through the processes of cleaning, cataloguing, analysis, and restoration.

The result of the archaeological studies during 2006 and 2007 identified 11 archaeological sites, all of them prehistoric, among which three were excavated and studied. In 2008, the archaeological survey identified seven archaeological sites, all of them prehistoric, these were excavated and studied, and the material, as in previous years, was duly sent to Unifap. Also, in 2008, five carbon samples were sent to the Beta laboratory in the United States for carbon dating. In 2009, in the municipality of Pedra Branca do Amaparí, four other archaeological sites were rescued, with the excavation of three sites and surface collection of one site. Currently, this area belongs to Zamin Resources Ltd. The results of the archaeological studies from 2005 to 2009 subsidized the BDR Operation Licensing for current operation at Tucano Project.

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After 2012, when operations commenced, archaeological studies were part of the routine management of the company, working together with near mining mineral exploration, monitoring, mapping / identifying sites and restoring artefacts.

20.9 Socio-economics

The Tucano Project covers an area of approximately 4,000 km² socially affected by the mining project, partially including the municipalities of Serra do Navio and Pedra Branca do Amapari, in which the population is distributed at the town and in small communities.

The municipality of Pedra Branca do Amapari, has a population of 10,773 inhabitants with a population density of 0.50 habitants/km². The entire population is considered a stakeholder of the project within the scope activities, totalling 21 communities dispersed along the unpaved highway BR-210.

The municipality of Serra do Navio, has a population of 5,111 inhabitants with a population density of 0.66 habitants/km², where practically the entire population is included within the Property, acting both in the town and the 10 small communities surrounding the western part of the Tucano project.

BDR has two professionals working full time closely with the communities. Among the actions already developed since the beginning of its activities are:

• Community Development Funds, with the distribution of resources through the approval of specific projects submitted in partnership with the Community Councils of Pedra Branca do Amaparí and Serra do Navio.

• Job creation and increasing earnings, with the prioritization of labour and local suppliers, as well as several projects with direct participation of stakeholders, which include:

— Projects of fisheries (fish farms).

— Co-operative of seamstresses "Rosa Branca do Amaparí.

— Donation of truck to farmers.

— Libraries.

— Training courses for environmental and community agents.

— Juniors Entrepreneurship Education Project.

— Cooking project “Cozinha Brasil”.

— Implementation of computer laboratory.

— Community factories of roasted cassava flour.

— Paper recycling workshops in Pedra Branca do Amaparí and Serra do Navio.

— Project “Guarda Mirim” with support of uniforms and donation of musical instruments.

— Educational games and “Gol de Placa” Project.

— Elaboration of educational folders.

— Donation of school supplies.

• Social assistance with:

— Donation of fuel and support for the Tutelary Council of municipalities and school busses.

— Basic food baskets distribution.

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— Circle of lectures and solidarity aid related to commemorative dates, such as Solidarity Christmas, New Year’s Day (Ano Novo da Paz), International Women's Day, and Easter, among others.

— Constant visits to communities working solidarity within the community routines.

• Health, where the company donated two ambulances, one for each Municipality Health Department.

• Public Security, where the company donated vehicles to the Environmental Battalion of the Military Police of Amapá State.

• Infrastructure

— Constant support for repairs and improvement of access roads.

— Support to the improvement of sanitation infrastructure, public buildings, and streets.

— Donation of diesel fuel oil to the municipality solid waste collection service.

— Infrastructure for popular shopping.

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21 Capital and operating costs

21.1 Capital expenditures

Total capital expenditures for 2018 have been estimated at US$26.9M, inclusive of US$18.5M for the completion of the Tucano Plant Expansion Project. Other significant estimated capital expenditures calculated at an exchange rate of 3.3 in 2018 are:

• Tails storage facility - 2018; US$2.68M.

• Powerline upgrade - 2018; US$1.52M.

• Urucum site buildings - 2018; US$1.10M.

This does not include capitalized stripping or miscellaneous minor capital requirements.

21.2 Operating costs

The 2017 operating costs for the Tucano mine are presented in Table 21.1 and include all mining, treatment, and general and administrative costs, which were incurred at the mine site. The cash operating costs exclude depreciation, amortization, sustaining capital, and corporate general and administration costs.

Table 21.1 Cash operating costs 2017

Year 2017	Cost (US$M)	Unit cost
Item
Operational mining costs	96.68	4.10	US$/tonne mined
Operational processing costs	54.62	14.68	US$/tonne milled
On-site general & administrative costs	7.65	2.06	US$/tonne milled
Inventory adjustments and other costs	-17.85
Total	141.09	1,100	US$/ounce sold

Note: The operating costs for 2017 were US$1,100/oz of gold sold using a US$/R$ exchange rate of 3.193. In 2017, the mine produced 129,764 oz at an all-in sustaining cost (AISC) of US$1,180/oz.

21.2.1 Underground capital costs - basis of estimate

Capital costs have been estimated based on the following:

• Capital infrastructure - is based on preliminary design assumptions and costing undertaken for other projects during the past 5 years. Where possible labour rates in these estimates have been adjusted to reflect Brazilian labour rates. A contingency of 20% has been added to these costs to cover the fact that detailed engineering has not been undertaken for these items and for minor items not yet allowed for in the PFS.

• Capital lateral and vertical development - based on mining schedules and first principles estimates based on equipment operating costs and assumptions and employee cost assumptions. A raiseboring cost of US$5,000 per metre has been assumed for 3.6 m diameter raises connecting to surface.

All costs, unless otherwise stated, are in US dollars ($).

21.2.2 Contractor mining versus owner mining

By deploying a mining contractor and not having to purchase the mobile equipment fleet upfront, AMC estimates that some capital cost savings could be realized. Project capital costs could reduce further depending on the mobile equipment and infrastructure the contractor may already have, and how it has been depreciated in the past.

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In view of the above, AMC has based this capital cost estimate on deploying an experienced mining contractor to develop and operate the project and assumes:

• The contractor would lease the underground mobile equipment fleet and pass on the leasing costs to GPR, together with all mining costs plus a contractor margin of 10%.

• GPR would finance all underground and surface infrastructure for the project.

• GPR would be responsible for all power and diesel costs.

21.2.3 Total underground capital cost estimate

The total underground capital cost estimated for the project over the LOM is summarized in Table 21.2. The estimated capital cost by year is shown in Table 21.3.

Table 21.2 Pre-production capital considers all capital expenditure in Year 1 (2019).

Table 21.2 Pre-production capital and total capital cost estimate

		Pre-production capital ($)
Item	Total ($)	Y01
Jumbo development	19,556,460	1,992,446
Underground infrastructure	11,027,907	1,700,966
Surface infrastructure	7,259,353	3,665,152
Raiseboring	2,079,585	1,097,021
Infrastructure contingency	4,073,369	1,292,628
Total capital	43,996,674	9,748,213
Pre-production capital	9,748,213
Sustaining capital	34,248,461
Total LOM Capital	43,996,674

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

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Table 21.3 Capital cost estimate by year

Capex item	Total	2019	2020	2021	2022	2023	2024	2025	2026
Jumbo development	19.56	1.99	6.33	6.27	4.27	0.35	0.34
Underground infrastructure
Egress ladderway	0.56	-	0.17	0.22	0.11	0.02	0.04
High / medium voltage switchgear	0.47	0.47
Power supply and distribution	4.47	0.39	1.46	1.48	0.98	0.09	0.07
Mine dewatering and water supply	4.10	0.36	1.34	1.36	0.90	0.08	0.06
Compressed air - portable compressors	0.20	0.10	0.10
Refuge chambers	0.22	0.11	0.11
Fire detection and suppression	0.16	0.08	0.08
Leaky feeder communications system	0.17	0.01	0.05	0.05	0.04	0.00	0.00
Mine rescue equipment	0.10	0.05	0.05
Auxiliary ventilation fans and regulators	0.40	0.03	0.13	0.13	0.09	0.01	0.01
UG workshops - underground fitout	0.20	0.10	0.10
Surface infrastructure
Portal preparation	0.20	0.20
Surface exhaust fan (2 x 560 kW parallel fans)	2.63	1.31	1.31
Change rooms / lamp room / lamps + SRs	1.47	0.18	0.18	0.18	0.18	0.18	0.18	0.18	0.18
Underground mining offices	1.13	0.14	0.14	0.14	0.14	0.14	0.14	0.14	0.14
Surface workshop	0.30	0.30
Water supply (PRV, tanks)	0.24	0.24
Explosives magazine	-	-
Fuel storage and bowser	0.15	0.15
Stores / warehouse facility	-	-
Surveying and engineering equipment	0.20	0.20
First fills	0.94	0.94
Raiseboring	2.08	1.10	0.98
Infrastructure contingency	4.07	1.29	1.24	0.71	0.49	0.10	0.10	0.07	0.07
Total Capex	44.00	9.75	13.80	10.55	7.20	0.98	0.95	0.39	0.39

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

The projected annual capital expenditure is summarized in Table 21.4.

Table 21.4 Annual capital expenditure

Year	Date	Capital $M
Year 1	2019	9.75
Year 2	2020	13.80
Year 3	2021	10.55
Year 4	2022	7.20
Year 5	2023	0.98
Year 6	2024	0.95
Year 7	2025	0.39
Year 8	2026	0.39
Total		44.00

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

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21.2.4 Underground lateral development capital costs

Capital lateral development is expected to cost approximately US$20M over the life of the project. This cost excludes material handling costs, power, and diesel costs which have been allocated to operating costs.

21.2.5 Underground vertical development capital costs

Capital raiseboring is expected to cost around US$2M during the first 18 months of the project. Additional long-hole raising for both the airway extensions and emergency egress has been costed as an operating cost (apart from purchasing of the emergency ladderway at a cost of $0.5M).

21.2.6 Underground mobile equipment capital costs

The underground mine is assumed to use a contractor for all development and production throughout the life of the mine. The equipment costs are included in the contractor’s development and production mining rates.

The estimate for the contractor provided equipment includes the following (based on maximum requirements):

• Longhole production drill (1)

• Jumbos - double-boom (6)

• 10-tonne LHDs (5)

• Articulated trucks - 32.5-tonne (5)

• Cable bolters and shotcrete machines (4)

• Ancillary equipment

21.2.7 Underground infrastructure capital costs

The underground infrastructure capital cost estimate is $11M (Table 21.5). Costs are based upon supplier quotations and unit rates from past experience with similar projects. The underground infrastructure costs largely consist of electrical reticulation, ventilation and dewatering system costs.

Table 21.5 Underground infrastructure capital costs

Description	$M
Egress ladderway	0.56
High / medium voltage switchgear	0.47
Power supply and distribution	4.47
Mine dewatering and water supply	4.10
Compressed air - portable compressors	0.20
Refuge chambers	0.22
Fire detection and suppression	0.16
Leaky feeder communications system	0.17
Mine rescue equipment	0.10
Auxiliary ventilation fans and regulators	0.40
UG workshops - underground fitout	0.20
Total	11.03

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments

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21.2.8 Surface infrastructure capital costs

Table 21.6 shows a breakdown of the direct capital costs estimated for the surface infrastructure required to support underground mining.

Table 21.6 Surface infrastructure capital costs

Description	Total $M
Portal preparation	0.20
Surface exhaust fan (2 x 560 kW parallel fans)	2.63
Change rooms / lamp room / lamps + SRs1	1.47
Underground mining offices	1.13
Surface workshop	0.30
Water supply (PRV2, tanks)	0.24
Explosives magazine*	-
Fuel storage and bowser	0.15
Stores / warehouse facility*	-
Surveying and engineering equipment	0.20
First fills	0.94
Total	7.26

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

(1) Self-rescuers.

(2) Pressure reducing valve.

* The underground mine will use existing facilities that support the open pit mine.

21.2.9 Exclusions from the underground cost estimate

The following costs have not been included in the underground capital cost estimate:

• Mine closure costs (to be captured as part of open pit costs).

• Mill modification costs.

• Surface road costs to the Urucum North Underground Mine site.

• Additional explosives magazine facilities.

• Additional stores / warehousing facilities.

AMC understands that existing facilities can be used in terms of the surface road, explosives magazine and stores / warehouse.

21.2.10 Indirect capital costs

Indirect costs are those relating to engineering, procurement, and construction management (EPCM), the costs associated with servicing and maintaining the mine during the construction period, and operating costs incurred prior to first concentrate production. No owners’ costs or EPCM have been considered in the cost estimate.

21.2.11 Capital contingency

Capital contingency total amounts are shown in Table 21.6 and average approximately 10% of the direct capital cost.

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Table 21.7 Capital contingency

Contingency	$M
Mine infrastructure	2.2
Surface infrastructure	1.5
Raiseboring	0.4
Total	4.1

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

21.2.12 Underground operating costs - basis of estimate

Underground operating cost estimates have been built up from first principles based on the following:

• AMC benchmarked mining equipment operating costs and productivities.

• Budget quotations for the mining fleet.

• The mining schedule - output from Mine24D and EPS schedules.

• Underground mining productivities for the various equipment types.

• Employee costs based on operating and maintenance assumptions and the employee costs.

• Effective working hours per shift for the underground workforce are 5 hrs 45 mins per shift.

• Power costs for pumping and ventilation are based on calculated power consumption (by year) and a power cost of US$0.08 per kWh (based on connecting to grid power in the future).

• Diesel cost of US$0.80/l.

• Diamond drilling based on a contractor rate of US$120/m (includes assaying, etc.).

• A mining contractor margin of 10%.

• An operating cost contingency of 10%.

• R$/US$ exchange rate of 3.8:1.

• Infrastructure costs (power water etc.) based on recently completed projects outside of Brazil but with the labour component adjusted to reflect Brazilian labour costs.

• Surface ore transport cost of US$1.35/t (based on scoping study estimate).

• A cost of US$17/t has been included to cover ore processing and G&A. (based on scoping study estimate).

It is assumed that the mining contractor will lease the mobile equipment fleet and charge the leasing costs back to GPR as an operating cost. The equipment leasing costs are estimated to total US$29M over the life of the project.

21.2.13 Underground mine total operating costs

LOM underground total operating costs, by major area, are shown in Table 21.8; annual operation totals are shown in Table 21.9.

Table 21.8 LOM underground total operating costs by major area

Department	$M
Mine	170.4
Process plant and G&A	40.4
Total	210.9

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

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Table 21.9 LOM underground total operating costs by year

Year	Date	Operating cost $M
Year 1	2019	8.3
Year 2	2020	26.9
Year 3	2021	36.0
Year 4	2022	39.4
Year 5	2023	28.2
Year 6	2024	29.6
Year 7	2025	22.7
Year 8	2026	19.7
Total		210.9

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

LOM average underground total operating costs have been estimated at $88.67/t milled. The cost breakdown by major area is shown in Table 21.10.

Table 21.10 LOM average underground total operating costs

Area	$/t milled
Mine	71.67
Process plant and G&A	17.00
Total	88.67

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

21.2.13.1 Underground mining cost

Underground mining cost estimates are detailed in Table 21.11.

Table 21.11 Underground mining costs

Operating costs	$M	$/t milled
Lateral development	17.7	7.45
Stoping	5.3	2.24
Materials handling	31.3	13.17
Power	12.4	5.20
Diesel	7.6	3.20
Ancillary equipment	3.3	1.38
Management and supervision	33.5	14.10
Miscellaneous	3.5	1.46
Surface haulage HG Ore	3.9	1.64
Mining opex cost contingency	11.9	5.00
Mining contractor equipment leasing costs	29.0	12.21
Mining contractor margin	11.0	4.61
Total	170.4	71.67

Note: Totals may not necessarily equal the sum of the components due to rounding adjustments.

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21.2.14 Underground mobile equipment assumptions and costs

Underground mobile equipment requirements are based on the productivities indicated in Table 21.12.

Table 21.12 Underground mobile equipment productivity assumptions

Equipment	Instantaneous productivity		Availability %	Utilization%
Jumbo - double boom	84.3	drill m/hr	70	30
Scaler	6.0	dev m/hr	70	30
Rock bolter	14.5	bolts/hr	70	40
Production drill	34.8	drill m/hr	80	25
LHD (10 tonne)	91.0	t/hr	80	50
Front-end loader (UG truck loading)	300.0	t/hr	80	50
Articulated truck	200.0	t-km/hr	80	60

Underground mobile equipment operating costs are based on the operating labour, consumables and energy costs indicated in Table 21.13.

Table 21.13 Mobile equipment operating cost assumptions

Equipment type	Unit	Operating labour (hr/unit)	Consumables ($/unit)	Energy ($/unit)
Jumbo - double boom	drill m	0.04	1.89	0.14
Scaler	dev m	0.56	40.00	40.00
Rock bolter	bolts	0.35	48.00	8.00
Production drill	drill m	0.11	7.12	0.21
LHD (10 tonne)	t	0.02	0.20	0.34
Front-end loader (UG truck loading)	t	0.01	0.20	0.34
Articulated truck	t-km	0.01	0.08	0.27

Underground mobile equipment maintenance costs are based on the maintenance labour, consumables and energy costs indicated in Table 21.14.

Table 21.14 Mobile equipment maintenance cost assumptions

Equipment type	Unit	Maintenance labour (hr/unit)	Consumables ($/unit)	Energy ($/unit)
Jumbo - double boom	drill m	0.03	3.49	0.08
Scaler	dev m	0.30	1.20	40.00
Rock bolter	bolts	0.20	1.20	8.00
Production drill	drill m	0.04	10.63	0.09
LHD (10 tonne)	t	0.01	1.30	0.08
Front-end loader (UG truck loading)	t	0.01	0.70	0.08
Articulated truck	t-km	0.01	0.41	0.01

21.2.15 Employee costs / labour costs

Employee costs have been based on annual wages as indicated in Table 21.15. These wages have been sourced from Brazilian 2014 wages quoted in R$/month adding 10% for inflation then converted to US$ rates using an exchange rate of US$:R$ of 1:3.8. These wages are inclusive of Brazilian labour benefits indicated in Table 21.16.

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Table 21.15 Employee wages

Type	2014 wage (R$ / month)	Wage (US$ / year)
Jumbo operator	7,822	27,171
Production drill operator	7,822	27,171
LHD / Truck operator	7,822	27,171
Semi-skilled labourer	6,509	22,611
IT operator	6,509	22,611
Unskilled labourer	5,393	18,732
Explosives charger operator	7,822	27,171
Mining supervisor	17,799	61,828
Technician	12,888	44,768
Maintenance engineer	28,953	100,574
Supervisor	17,799	61,828
Planner / Programmer	9,248	32,124
Artisans	7,822	27,171
Electricians	9,248	32,124
Geotechnical engineer	39,292	136,490
Engineer	39,292	136,490

Table 21.16 Employee benefits

Benefits	Value	Unit
Medical and dental cover cost	400	R$ per month
Profit share	2	% of Wage
Travel allowance	500	R$ per month
Meal cost	20	R$ per month
Meal and lunch ticket	20	R$ per month
Medical checks	85	R$ per month
Other personal expenses	150	R$ per month
Life insurance	170	R$ per month
13th salary provision	170	R$ per month
Benefit for dangerous work	30	% of Wage
Benefit for extended shift	25	% of Wage
Benefit for health-risk work	20	% of Wage
Benefit for night shift	10	% of Wage
Annual leave provision	8	% of Wage
FGTS	170	R$ per month
INSS	170	R$ per month

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22 Economic analysis

GPR is a producing issuer, as defined by NI 43-101 and is currently in production.

The economics regarding only the proposed underground mine are discussed further.

22.1 Underground mine economics

22.1.1 Assumptions

All currency is in US dollars ($) unless otherwise stated. The cost estimate and projected revenue for the underground mine were prepared with a base date of Year 1 and use constant Year 1 dollars (no inflation). For NPV estimation, all costs and revenues are discounted at 5% from the base date. The gold price was selected after discussion with GPR and referencing current market and recent historical prices, values used in other recent projects, and forecasts in the public domain. No depreciation, amortization or taxation has been undertaken in this evaluation. An exchange rate of US$:R$ of 1:3.8 and a royalty of 2% have been assumed. Changes to the exchange rate affect labour costs only.

Revenues from gold production have been based on:

• Metallurgical recovery of 90% (based on various metallurgical studies of the Tucano ore supported by actual recovery data from current mining and processing of the Urucum open pit ore).

• Gold price of $1,200/oz in line with recent assumptions by GPR for Mineral Reserve estimates.

• Gold royalty of 2%.

22.1.2 Economic analysis

AMC has conducted a pre-tax economic assessment of the Urucum North Underground Mine. Over an 8-year operating life, the mine is projected to generate approximately $25.5M pre-tax NPV at 5% discount rate, with a pre-tax internal rate of return (IRR) of 19.9%.

Project capital is estimated at $44M with a payback period of 4.9 years (undiscounted pre-tax cash flow from Project Year 1). Key assumptions and results of the underground mine economic assessment are provided in Table 22.1. The LOM production schedule, average gold grade, recovered metal, and cash flow forecast are shown in Table 22.2.

An operating cash cost of $802.37 per oz of gold over the LOM was calculated by taking the total operating costs (excluding Year 1) and dividing this value by the payable ounces of gold (excluding Year 1).

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Table 22.1 Key economic assumptions and results for the underground mine

Urucum North Underground project	Unit	Value
Total ore production	kt	2,378
Total waste production	kt	2,000
Gold grade (average)1	g/t	3.6
Overall gold recovery (mill) 1	%	90
Gold price	$/oz	1,200
Payable gold metal	oz	252,801
Gross revenue	$M	303
Royalty	%	2.0
Capital costs - pre-production	$M	10
Capital costs - sustaining	$M	34
Mine operating costs2	$M	170
Mine operating costs2	$/t ore	71.67
Operating costs (total)3	$/t ore	88.67
Cash cost4	$/oz Au	802.37
Undiscounted cashflow (pre-tax)	$M	42.4
Pre-tax IRR	%	19.9
Discount rate	%	5
Pre-tax NPV5%	$M	25.5

Notes: Totals do not necessarily equal the sum of the components due to rounding adjustments.

(1) LOM average.

(2) Underground mining costs only.

(3) Includes mine operating costs, milling, and G&A.

(4) Excludes off-site costs such as smelter treatment, refining, and transportation costs.

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Table 22.2 Urucum North Underground LOM production and cash flow forecast

Item	Unit	Total	2019	2020	2021	2022	2023	2024	2025	2026
Ore mined	t	2,378,113	4,861	217,801	341,481	464,935	419,166	352,862	303,141	273,865
Development - lateral	m	35,995	2,159	8,409	9,886	8,119	2,259	3,277	1,084	801
Development - vertical	m	2,270	250	698	729	374	85	133
Waste mined	t	1,999,824	159,539	559,767	613,523	432,137	90,311	95,885	32,937	15,724
Gold grade - milled	g/t	3.64	2.8	3.0	3.4	4.1	4.0	3.8	3.7	3.1
Overall gold recovery	%	90	90	90	90	90	90	90	90	90
Gold payable metal	oz	252,801	390	18,820	33,454	55,602	48,619	39,104	32,270	24,544
Total net revenue1	$M	297.3	0.5	22.1	39.3	65.4	57.2	46.0	37.9	28.9
Operating costs
Underground mining	$M	170.4	8.3	23.2	30.2	31.5	21.1	23.6	17.6	15.0
Processing and G&A	$M	40.4	0.1	3.7	5.8	7.9	7.1	6.0	5.2	4.7
Total operating cost	$M	210.9	8.3	26.9	36.0	39.4	28.2	29.6	22.7	19.7
Capital costs
Project capital	$M	9.8	9.8
Sustaining capital	$M	34.3		13.8	10.6	7.2	1.0	1.0	0.4	0.4
Total capital cost	$M	44.0	9.8	13.8	10.6	7.2	1.0	1.0	0.4	0.4
Undiscounted cash flow (pre-tax)	$M	42.4	(17.6)	(18.6)	(7.2)	18.7	28.0	15.4	14.8	8.8
Discounted cash flow (pre-tax)5%	$M	25.5	(16.8)	(16.9)	(6.2)	15.4	22.0	11.5	10.5	6.0
Cumulative pre-tax NPV5%	$M	25.5	(16.8)	(33.7)	(39.9)	(24.5)	(2.5)	9.0	19.6	25.5

Notes: Totals may not necessarily equal the sum of the components due to rounding adjustments.

(1) Gross revenue less any selling, refining, transportation, and royalty costs.

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Production costs, revenues, and cumulative pre-tax cash flow by year are shown in Figure 22.1.

Figure 22.1 Production costs, revenues, and cumulative pre-tax cash flow by year

Source: AMC

22.2 Taxes and Royalties

There are two corporate income taxes in Brazil which GPR is subject to: (a) the Corporate Income Tax (IRPJ) and (b) the Social Contribution Tax on Profits (CSLL). The tax base of IRPJ and CSLL is similar, although some specific adjustments may be applicable to one tax and not to the other.

As GPR will operate in the Amazon region of Brazil, the company is eligible for certain tax incentives.

IRPJ is a federal tax charged on net taxable income. It applies at a basic rate of 15%, plus a surtax of 10% on annual income that exceeds R$240,000 per year.

CSLL was introduced to fund social and welfare programs and is paid in addition to IRPJ. CSLL is also a federal tax levied on net taxable income and is applied at a rate of 9%.

Tax losses may be carried forward indefinitely. The offset is limited to a maximum 30% of annual taxable income. Tax losses may be lost if, between their generation and utilization, there is a change in control and change in the business activity performed by the taxpayer.

The Tucano Gold Mine is located in the state of Amapá in the Amazon region of Brazil, and as such the operation is eligible for, and has received, certain incentives under the Superintendence of Amazonas Development (SUDAM) regime.

SUDAM oversees development in the Amazon region and is administered by the Brazilian federal government. Under the SUDAM regime, the Company has received a 75% tax exemption on IRPJ payable on net taxable income generated from the Tucano Gold Mine. The SUDAM benefit applies for a 10 year period and commenced in 2013 in relation to the Tucano Gold Mine.

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Royalties are discussed in Section 4.4.

22.3 Economic sensitivity analysis

AMC has carried out a sensitivity analysis of the projection for Urucum North Underground economics. The sensitivity analysis examined the impact on pre-tax NPV (at 5% discount rate) of a 20% positive or negative change in the gold price, total production cost (all operating and capital costs), foreign exchange rate (US$:R$) and the price of power.

The results of the sensitivity analysis are shown in Table 22.3, Figure 22.2, and in Figure 22.3.

The results show that the pre-tax NPV is positive and remains so for the range of sensitivities evaluated for exchange rate and the cost of power. For the sensitivity of the gold price, the pre-tax NPV is positive when the price of gold is over $1,070/oz and for the total operating cost, the pre-tax NPV is positive when the total operating cost does not increase by more than 15%.

Pre-tax NPV is most sensitive to changes in the gold price and to changes in the total operating cost. The NPV is moderately sensitive to changes in the exchange rate. Changes in the cost of power have the least impact on NPV.

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Table 22.3 Economic sensitivity analysis (pre-tax)

Item	Value	Unit	Pre-tax NPV ($M)	Pre-tax IRR %
Base case (NPV at 5%)			25.5	19.9
Gold price - fall of 20%	960	$/oz	(22.3)	(9.4)
Gold price - increase of 20%	1,440	$/oz	72.3	44.6
Total operating cost - fall of 20%	168	$M	59.4	41.2
Total operating cost - increase of 20%	253	$M	(8.4)	0.1
Total power cost - fall of 20%	10	$M	29.0	21.9
Total power cost - increase of 20%	15	$M	21.4	17.5
Exchange rate - fall of 20%	3.0	US$:R$	13.2	12.6
Exchange rate - increase of 20%	4.6	US$:R$	33.7	25.0

Figure 22.2 Sensitivity analysis - pre-tax NPV at 5% discount rate

Source: AMC

Figure 22.3 shows the change in pre-tax NPV and IRR for different gold prices.

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Figure 22.3 Pre-tax NPV and IRR sensitivities to changes in the gold price

Source: AMC

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23 Adjacent properties

As defined in Form 43-101F1 an adjacent property means a property in which the issuer does not have an interest. The Tucano tenement ownership map is shown is Figure 23.1 and the details of that ownership are discussed in Section 4.

There are no adjacent properties.

Figure 23.1 Tenement ownership map

Sources: BDR

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24 Other relevant data and information

To the authors knowledge, there is no other relevant data or information related to the Tucano Gold property required so as to make the report more understandable or not misleading.

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25 Interpretation and conclusions

25.1 Risk assessment

The main risks identified for the operation previously have been largely reduced over the years of operation by BDR. Some key risks and controls include:

• Delays to obtain licencing and permitting; GPR is working closely with government agencies to ensure permitting is received by the required dates. No major issues identified to date.

• Wet weather impacting mine production; Wet weather has a negative effect on mine production however there has been significant improvement in recent years with an ongoing campaign of road rock sheeting and maintenance.

• Geotechnical stability of pit walls in the oxidized profile; Monitoring of pit walls and adhesion to the recommended mining parameters is enforced.

• Control of mining dilution and loss; Geology department maintains tight control over ore mark-outs, ore spotting, blast monitoring, and floor clean up to reduce ore dilution and loss.

25.2 Mineral Resources and Mineral Reserves

• In the opinion of the QPs, the drillhole database for the Tucano Project is reliable for the purpose of Mineral Resource estimation.

• Drilling on the Tap AB deposits has significantly increased the oxide resources and reserves in the AB1 area with high potential to add resources along strike of the Tap AB pit.

• Mineral Resources and Mineral Reserves have been classified according to the Canadian National Instrument 43-101 (NI 43-101) Standards of Disclosure for Mineral Projects.

• For reporting, the QPs tabulated the Mineral Resources within pit shells optimized using a gold price of US$1,500 per ounce of gold. Mineral Reserves were based on reserve modifying factors and a gold price of US$1,200 per ounce within pit designs which were based on optimized shells using US$1,150 per ounce for Tap AB and US$1,100 per ounce for Urucum and Tap C.

25.3 Mining

• BDR commenced open pit mining operations in 2011 after acquiring the former heap leach operation from New Gold in 2010.

• U&M is the main mining contractor at Tucano.

• Based on current Mineral Reserves and mine planning, the Tucano LOM production schedule extends from 2018 through to 2023 from the open pits, with a total of 1.07M oz of gold being produced during this period.

• The risks to the Tucano Mine LOM plan and operations are currently considered to be low, with outcomes being sensitive to negative commercial trends that might develop in respect of the gold price, exchange rate, and the inflationary effect on labour and consumables.

25.4 Processing

Tucano commissioned the CIL plant in November 2012 and during 2017 a total of 129,764 ounces of gold was recovered from 3.7 Mt of ore. The current plant expansion now in progress aims to increase blend ratio from 20% to 80% for sulphides. Gold recoveries are expected to improve to 93% with the ability to consistently achieve a grind sizing P80 of 75 μm following the the already commissioned ball mill and the addition of an oxygen plant and an extra leach tank in the circuit.

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25.5 Production and all-in sustaining costs

In 2017, Tucano mine produced 129,764 at an AISC of US$1,180/oz. The 2017 AISC includes all mining, treatment, general and administrative costs, sustaining capital, and royalties which are incurred at the mine site. The AISC costs exclude depreciation, amortization and corporate general and administrative costs.

25.6 Health and safety

Health and safety is actively promoted by the site Safety department through on-going education, training, monitoring, and development of a positive safety culture on site.

25.7 Underground mining

• The mine will be accessed by a single portal located at the northern end of the mine and twin declines (north and south) based on orebody strike length of >800 m.

• Level interval of 20 m (based on 15 m production blastholes and the narrow width of some lodes) with crosscuts between the decline and the mineralization, every 20 vertical metres (vm).

• AMC recommends developing the mine to use truck haulage for all material movement.

• Declines have been designed to give equipment access to each production level. On each level a crosscut will be developed from the decline to the ore lode / lodes. Declines and crosscut development has been designed to 5.0 m W x 5.0 m H, sufficient to accommodate 30 t - 40 t trucks.

• Ore drives have been designed to be 5.0 m H x 4.0 m W to accommodate production drills and 10 t load-haul-dump (LHD) units. Trucks will not enter the ore drives.

• Evaluation of the planned production rate and scheduling indicates that the deposit supports 450 ktpa to 500 ktpa. Production rates vary from year to year depending on the number of production fronts becoming available in different areas of the mine. The production rate could be further optimized as part of any future Feasibility Study.

• All waste will be tipped directly into stopes or trucked to surface.

• Approximately 2.4 Mt of ore at an average grade of 3.6 g/t Au will be mined and processed over the eight-year mine life.

• All ore production and mine development will be done by contractors; this includes all waste rock filling.

• Blasting will be undertaken primarily with ANFO and non-electric detonators. In conditions that are wet, bulk emulsion explosives will be utilized.

• Ventilation will be provided via a “pull” system with primary fans located on surface at the top of each of two primary exhaust raises.

• With the infrastructure airflow and leakage and balancing allowances, the total airflow determination based on the diesel fleet size is 350 m³/s.

• Major scheduling assumptions driving these production rates include: decline and crosscut advance of 140 m/month, ore drive advance of 60 m/month, production rate from stoping of 10,000 t/month for both Up-Hole retreat and Benching stopes, backfilling rate of 12,000 m³ per month, per stope, for Benching stopes. All scheduling assumptions are based on industry benchmarked data.

• Scheduling undertaken has been based on maintaining a near constant number of primary mobile equipment in both the North and South declines.

• Workforce size at full production is estimated to be 236 employees. Employees working underground will work 7 hr shifts, but be paid for 8 hrs on site with the additional hour allowed for meal breaks, showering etc. The estimated working time “at the face” for underground workers is estimated to be 5 hrs and 45 mins.

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• The underground workforce will consist of 6 crews working three 8 hr shifts per day Technical support staff, mining supervisors and general and administration employees will work a 5 day per week schedule.

• The mine will use modern trackless mobile equipment for the development and stoping operations.

• The underground mine would use some of the existing infrastructure built for the open pit operations such as the explosives magazines and warehouse.

• Mining costs are estimated to be US$71.67/t of ore milled and total operating costs, inclusive of processing and G&A, are estimated to be US$88.67/t of ore milled.

• The economic assessment indicates that the Project has potential to be developed into an economically viable, underground gold mine.

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26 Recommendations

26.1 Operations

The key objectives for Tucano are:

• To operate on a “Zero Harm” safety and environmental tolerance basis.

• Reduce costs and improve productivity of the operation.

• Provide on-going training for employees.

• Maintain a close reconciliation between planned production and actual production ounces.

• To extend the mine life through the focused resource targeting and definition drilling to increase reserves. With the upgraded mill expected to be completed in November 2018, the focus will shift to targeting high grade oxide and sulphide resources that can be incorporated into the mine plan.

• Maintain reconciliation studies and update the modelling factors accordingly for the determination of Mineral Resource and Mineral Reserve estimates.

• Boost company resources by new discoveries in green field areas outside of the mining concession.

• To achieve the above objectives in a socially responsible manner.

26.2 Exploration and Resource Delineation

26.2.1 Near mine and resource definition

Near mine resource definition in 2018 will focus on converting high grade Inferred Mineral Resources to Indicated Mineral Resources in areas occurring immediately outside the current reserve pits as well as locating new extensions to mineralization previously untested. With the plant upgrade planned to be complete in November 2018, the focus shifts from testing mainly oxide resources to both oxide and sulphide high grade mineralization that will increase open pit reserve and provide near term mill feed. A key area will be the poorly tested gap between the Urucum North and Urucum Central North pit areas where recent drilling has intersected strong mineralization. Other areas to extend the reserve include below the Urucum CN reserve pit, Tap AB1 Neo North, and Torres.

26.2.2 Regional exploration

Regional exploration will work through a ranked list of targets with the aim to locate a major new discovery outside the mine corridor. Exploration will range from first pass stream sediment sampling, soil and auger drilling in the green field areas and RC and diamond drilling to define mineralization at the more advanced projects. Currently a program of stream sediment sampling has been completed at Serra da Canga East and is being followed up with soil sample lines and mapping. A program of detailed airborne EM and ground IP is being reviewed for 2018. Total budget for Resource Definition, Near Mine Exploration and Regional Exploration in 2018 is US$6.02M. The majority of this cost relates to RC and Diamond Drilling as detailed in Table 26.1 below.

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Table 26.1 2018, Resource Definition, Near mine, and regional exploration budget

Summary	Description	Total m	Budget US$
Resource conversion	RC	5,000	651,515
Resource conversion	DD
Resource conversion	PW / Auger
Resource conversion	Soil
Near mine	RC	12,500	1,628,788
Near mine	DD	3,000	545,455
Near mine	PW / Auger	2,100
Near mine	Soil
Regional exploration	RC	7,700	600,000
Regional exploration	DD	2,500	390,909-
Regional exploration	PW / Auger	3,600	-
Regional exploration	Soil	9,290 (samples)
Total fixed labour			316,697
Total fixed on costs			1,258,792
Total Geology laboratory			287,273
Total Geophysical surveys			136,364
Total Archaeological surveys			212,121
Total Budget			6,027,914

RC = Reverse Circulation Drilling, DD = Diamond Drilling, PW = Rotary Air Blast Drilling (Airtrack blasthole "PW" rig), Auger = Man-portable motorized auger. PW/Auger and Soil Sampler costs captured under fixed labour costs.

26.3 Underground mining

The following recommendations are made in regard to the underground operation:

• Further drilling and investigation work aimed at upgrading Inferred Mineral Resources and increasing the geotechnical and hydrogeological understanding of the deposit is recommended to consolidate the design basis for the project.

• AMC recommends advancing the project to the FS stage. Opportunity exists to stage the project such that the second ramp is delayed until production from the first ramp proves successful and mining costs are better understood.

• AMC recommends developing and operating the underground mine with contractors, this resolves any issues regarding supply of an experienced underground work force and allows for a reduction in upfront capital expenditure for mobile equipment.

• Based on the current diamond drilling density and due to the nature of the gold mineralization (i.e. relatively coarse particle size) it is difficult to estimate gold grades within the mineralized zones accurately. AMC recommends that underground ore drive development, mapping, and sampling should be used to better define the mineralization.

• AMC recommends that further geotechnical analysis is carried out to facilitate design and the efficient recovery of pillars. Detailed geotechnical modelling is recommended during the FS stage. This will also assist with understanding of the potential dilution particularly in the narrower veins.

• Further metallurgical test work is recommended particularly on the primary sulphide material that will be extracted from underground. There is a risk that unexpected changes to metallurgical performance and gold recovery may occur.

• Further environmental work is required during the FS stage to ensure that the underground operations do not present any problems in terms of environmental compliance. Any compliance problems could inadvertently delay the project or add additional unplanned costs to the project.

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• Additional drilling is recommended to convert Inferred Resource to Indicated Resource and increase the underground Mineral Reserves. This is likely to have a positive impact on the project economics.

• Opportunities exist to reduce capital costs for the underground project through locally constructed surface change rooms and offices rather than rental and may be an inexpensive alternative. The use of less expensive Scania P420 trucks for underground haulage once the decline roadway (floor) has been prepared and graded is also an opportunity. These options should be evaluated.

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27 References

AMC Consultants 2016, Urucum Deeps Prefeasibility Study, BDR Resources Ltd. AMC Consultants: Melbourne, Australia. Report: AMC Project 115058.

Ausenco and SRK Consulting 2011, Beadell Resources Ltd Tucano Gold Project Feasibility Study Report, Ausenco Minerals and Metals & SRK Consulting: Perth, Australia. Report no. 2027-RPT-0037 Revision Number D.

Batelochi, M. 2015, 2015 Technical Mineral Resources Evaluation Report Tucano Project - Tap C & Gap Deposit, Gold Deposit in Pedra Branca Amapari - Amapa State - Brazil. M. Batelochi, Soluções em Geologia e Mineração Ltda; São Paulo Brazil.

Batelochi, M. 2016, 2016 Technical Mineral Resources Evaluation Report Tucano Project - Urucum Deposit, Gold Deposit in Pedra Branca Amapari - Amapa State - Brazil. M. Batelochi, Soluções em Geologia e Mineração Ltda; São Paulo Brazil.

Batelochi, M. 2017, 2017 Technical Mineral Resources Evaluation Report Tucano Project - Duckhead Deposit, Gold Deposit in Pedra Branca Amapari - Amapa State - Brazil. M. Batelochi, Soluções em Geologia e Mineração Ltda; São Paulo Brazil.

Costa, A. 2005. Levantamento Geofisico. Technical Report, January 2005. AFC Geofisica Ltda for MPBA.

Harrier Project Management 2017, Expansion Project Feasibility Study: Beadell Resources - Tucano Gold Project, Harrier Project Management: Perth, Australia. Report: Expansion Project_FS_Beadell Resources Tucano Mine Draft May Master.

Klinkert,P and Van den Berg, H. 2001. SPECTREM Survey of the Amapa Area, April 2001. Exploration and Acquisition Division, Anglo Operations Limited.

Potvin, Y. 1988. Empirical open Stope Design in Canada, November 1988, Universtity of British Columbia, BC, Canada.

Srivastava, M. R. and Isaacs, E. 1989, An Introduction to Applied Geostatistics. Oxford University Press, New York.

SRK Consultores do Brasil Ltda 2011, Plano de Controle Ambiental PCA (Environmental Control Plan), SRK Consultores do Brasil Ltda : Belo Horizonte, MG, Brasil. Report: SRK 031.11 30 de Setembro de 2011 - Revisão 02.

Tan, P. 2013, Urucum_East_Resource_Report_v1, Aug 2013, Beadell Internal Report.

Tomlinson, A. Jurza, P. Li, D. Trunfull, J. and Peters, P. 2015. Tucano Project Brazil - Geophysical Data Processing, 4 November 2015. Southern Geoscience Consultants: Perth, Australia. Report No. SGS3014.

Walter, B.M.T.; Cavalcanti, T.B. (eds. tec.) 2005. Fundamentos para a coleta de germoplasma vegetal. Embrapa Recursos Genéticos e Biotecnologia, Brasília, DF.

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28 QP Certificates

Brian Richard Wolfe

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Brian Richard Wolfe, MAIG, and I am the Principal Resource Geologist of the firm of International Resource Solutions Pty Ltd of 71 Watkins St White Gum Valley, WA 6162, Australia.

2. I am a practicing geologist registered with the Australian Institute of Geoscientists. My membership number is 4629.

3. I graduated with a BSc Degree (Hons) in Geology in 1992 from the National University of Ireland, Dublin, and hold a Postgraduate Certificate in Geostatistics from Edith Cowan University (2007).

4. I have practiced my profession for a total of 24 years since 1993. I have experience in exploration geology, mining geology and geostatistical modelling, and estimation of mineral resources. My previous relevant experience on deposits, which are similar in nature to that discussed in the Report, include Agbaou gold deposit (Cote d’Ivoire), Bibiani gold deposit (Ghana), Kobada gold deposit (Mali), Sabadola gold deposit (Senegal), and numerous others during more than 15 years spent working in various capacities on similar projects in the region and worldwide.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have personally visited the Tucano Project between 30 July 2015 and 3 August 2015.

7. I have previously been engaged with respect to the Tucano Gold Project by Beadell Resources Limited in my capacity as an Independent Consulting Resource Geologist to undertake Resource Estimation studies since 2015.

8. I am responsible or jointly responsible for preparing or supervising the preparation of Sections 1 (1.1 – 1.2, 1.4 – 1.9, and 1.17 – 1.18), Section 4.1, Sections 5 - 12, Section 14 (portions relating to open pit resources only), Section 23, Section 25.2, Section 26.2, and Section 27 (jointly) of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Gold Project and the Issuer (Beadell Resources Limited) pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have, nor do I expect to receive, a direct or indirect interest in the Tucano Gold Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, any securities of Beadell Resources Ltd or any associate or affiliate of such company.

Effective Date: 11 May 2018

Signed Date: 17 May 2018

Original Signed and Sealed by

Brian Richard Wolfe

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Nigel Arthur Spicer

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Nigel Arthur Spicer, C.Eng., MIMMM, and I am a Principal Mining Engineer of the firm of Minesure Pty Ltd of 12 Redmond Street, Salter Point, WA, Australia. My residential address is 12 Redmond Street, Salter Point, WA, Australia.

2. I am a practicing mining engineer registered with the Institute of Materials, Minerals and Mining (IMMM). I am a Chartered Engineer (C.Eng. MIMMM) and my membership number is 43912.

3. I am a graduate of the University of Wales and hold a Bachelor of Science in Mining Engineering.

4. I have practiced my profession continuously since 1974 during which I have estimated Ore Reserves for gold and base metal projects located in Africa, Asia, North and South America, and elsewhere. I have managed open pit and underground mining operations in Africa, Asia and Australia and compiled feasibility studies for projects in Africa, North America, South America, Asia, Europe, and Russia. My previous experience has included deposits which are similar in nature to that discussed in the Report.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have personally visited the Tucano Project area follows:

• 30 July 2015 – 1 August 2015,

• 5 September 2015 – 27 September 2015

• 14 November 2015 – 6 December 2015

• 28 January 2016 – 28 February 2016

• 24 January 2017 – 8 February 2017

7. I have previously been engaged with respect to the Tucano Project. I undertook a due diligence review of the Tucano Project in December 2009.

8. I am responsible for relevant parts of Sections 1, 3, 15, 16, 22, and 25 of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Project pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have a direct interest in the Tucano Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, securities of Beadell Resources Ltd.

Effective Date: 11 May 2018

Signed Date: 17 May 2018

Original Signed and Sealed by

Nigel Arthur Spicer

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Marcelo Antonio Batelochi

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Marcelo Antonio Batelochi, MAusIMM (CP), and I am a Principal Resource Geologist of the firm Soluções em Geologia e Mineração Ltda. My residential address is Av. Escola Politécnica, 942 – Bloco C3, Apto 41, São Paulo, São Paulo, Brazil.

2. I am a practicing geologist registered with the AusIMM - The Australasian Institute of Mining and Metallurgy. My membership number is 205477.

3. I hold a Bachelor of Honors degree from the School of Geology at UNESP - São Paulo State University, Brazil, in 1991.

4. I have practiced my profession continuously since 1992, during which I have more than 20 years of experience in the mineral resource evaluation of Iron, Copper / Gold, Nickel, Bauxite, REE, and PGE Deposits, as employee of Rio Tinto (12 years), Vale (4 years), Ferrous Resources (6 years), and five years as Independent Consultant (“MB Geologia”).

I have worked at Tucano deposits since 2012 carrying out Mineral Resources Estimates as independent consultant for Urucum, Urucum Underground, Tap-AB and Tap C + Gap, and Duckhead. Also, Tartarugalzinho Project.

My previous experience has included gold deposits of Yamana Gold (included Brio Gold), Aura Minerals and PA-Gold as a Consultant and Qualified Person. Specifically, “Morro do Ouro” deposit, where I worked for 10 years as manager of geology and mining planning (when it belonged to Rio Tinto).

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have personally visited the Tucano Project 4 times: January and May of 2012, January 2013, October 2013, February 2015, and March 2018.

7. I have not previously been engaged with respect to the Tucano Project in any capacity.

8. I am responsible for preparation of Section 20, and relevant parts of Sections 1, 4, 14, 26, and 27 of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Project pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have a direct interest in the Tucano Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, securities of Beadell Resources Ltd.

Effective Date: 11 May 2018

Signed Date: 17 May 2018

Original Signed and Sealed by

Marcelo Antonio Batelochi

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Ruy Lacourt Rodrigues

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Ruy Lacourt Rodrigues, SME – RM, and I am a Principal Mining Engineer of the firm LOM Consultants of Av. Alvares Cabral, 593 - Sala 1301 – Lourdes, 30 170-912 - Belo Horizonte – MG, Brazil.

2. I am a mining engineer, registered in the Brazilian regional council of engineering and agronomy (Conselho Regional de Engenharia e Agronomia, CREA) at the State of Bahia chapter (CREA BA 20655 D) since 1988. I am also a Registered Member of the Society for Mining, Metallurgy and Exploration, Inc. (SME), in the United States of America (SME 4172669RM) since 2014.

3. I graduated from the Universidade Federal de Ouro Preto, Brazil, in 1986 with a Bachelor of Science degree in Mining Engineering and a Master of Science degree in Mine Economics earned at the same institution in 2007.

4. I have worked as a mining engineer for 31 years since my graduation. My relevant experience to the Technical Report includes mining engineering, mine project and business development, planning and operations for gold and base metals covering mine project and planning, equipment specification and definition of personnel needs, mine operations and maintenance of mining equipment and infrastructure. My previous experience has included deposits which are similar in nature to that discussed in the Report.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have personally visited the Tucano Project in April 2018.

7. I have not previously been engaged with respect to the Tucano Project in any capacity.

8. I am responsible for preparation of relevant parts of Sections 5 and 18 of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Project pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have a direct interest in the Tucano Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, securities of Beadell Resources Ltd.

Effective Date: 11 May 2018

Signed Date: 17 May 2018

Original Signed and Sealed by

Ruy Lacourt Rodrigues

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Juan Jose Moreno Dellepiane

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Juan Jose Moreno Dellepiane, MIEAust, CPEng., and I am a Principal Geotechnical Engineer of the firm SRK Consulting of 10 Richardson Street, West Perth, WA, Australia.

2. I am a practising civil engineer registered in Australia as Chartered Professional Engineer (CPEng.) Member with the Institute of Engineers Australia (MIEA), my membership number is 3667986.

3. I am a graduate of the Pontifical Catholic University of Peru and hold a title of Civil Engineer, August 1992.

4. I have practiced my profession continuously since 1992 during which I have worked in diverse soil mechanics projects, earthworks design, design and operation of tailings facilities for projects located in Africa, Asia, North and South America, specifically in Brazil. My specific experience with the Tucano Gold Mine includes the design of the tailings storage facilities (TSF): Tap D, North Mill Pond, West Pond Phase 1, West Pond Phase 2.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have personally visited the Tucano Project area as follows:

• 6 January 2014 – 7 January 2014

• 16 November 2014 – 18 November 2014

• 11 November 2015 – 16 November 2015

• 19 October 2016 – 21 October 2016

7. I have previously been engaged with respect to the Tucano Project in any capacity.

8. I am responsible for preparation relevant parts of Section 18 of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Project pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have a direct interest in the Tucano Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, securities of Beadell Resources Ltd.

Effective Date: 11 May 2018

Signed Date: 23 May 2018

Original Signed and Sealed by

Juan Jose Moreno Dellepiane

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Peter Robin O’Bryan

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Peter Robin O’Bryan, MAusIMM (CP), and I am a Principal Mining Engineer of the firm Peter O’Bryan and Associates of 43 York Street, Subiaco, WA, Australia. My residential address is 142 Tower Street, West Leederville, WA, Australia.

2. I am a practicing mining geotechnical engineer and registered Chartered Professional with the Australasian Institute of Mining and Metallurgy (AusIMM), Membership Number 203335.

3. I am a graduate of the University of New South Wales Bachelor of Engineering (Mining) and James Cook University, Master of Engineering Science (Rock Engineering).

4. I have practiced my profession continuously since 1982 during which I have provided mining geotechnical services for investigation, mine evaluation and mine operation for projects located in Australia, Africa, Asia, Canada, North and South America, Spain, and Eastern Europe. I have contributed to compilation and due diligence of feasibility studies for projects in Africa, Australia, Canada, and Asia. My previous experience has included deposits and operations which are similar in nature to that discussed in the Report.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have personally visited the Tucano Project in September 2014 (4 days on site) and June 2015.

7. I have previously been engaged with respect to the Tucano Project. I spent 5 days on site in May 2015.

8. I am responsible for preparation of relevant parts of Section 16 of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Project pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have a direct interest in the Tucano Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, securities of Beadell Resources Ltd.

Effective Date: 11 May 2018

Signed Date: 17 May 2018

Original Signed and Sealed by

Peter Robin O’Bryan

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Raymond Henry Walton

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Raymond Henry Walton, P.Eng., and I am President and Principal Metallurgist of the firm Ray Walton Consulting Inc (RWC Inc.) of 182 DeGraaf Crescent, Aurora, Ontario, Canada, L4G 0X1.

2. I am a Professional Engineer, registered with the Association of Professional Engineers of Ontario (APGO), no. 90294521, since 1991.

3. I am a graduate of Brunel University, Uxbridge, Middlesex, 1977, where I obtained a B.Tech (Hons) degree in Materials Science and Technology.

4. I have 41 years of experience in process design and engineering, project management, commissioning and metallurgical plant operations. I have practiced my profession continuously since 1977. My relevant experience in copper and gold metallurgy, includes 9 years in operations, 20 years with engineering companies and 9 years as a Senior Director, Process Design within a large corporate Capital Projects group. I have been responsible for corporate oversight, feasibility studies, project evaluations, engineering, commissioning as well as operations on a large number of gold and copper projects around the world.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have personally visited the Tucano Project in January 2016.

7. I have previously been engaged with respect to the Tucano Project as a Consultant Metallurgist.

8. I am responsible for preparation of relevant parts of Sections 13, 17, and 25.4 of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Project pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have a direct interest in the Tucano Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, securities of Beadell Resources Ltd.

Effective Date: 11 May 2018

Signed Date: 17 May 2018

Original Signed and Sealed by

Raymond Henry Walton

amcconsultants.com 264

Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

Gary Methven

I am a Qualified Person of the report entitled “Technical Report Mineral Resource and Reserve Update for the Tucano Gold Mine, Amapa, Brazil” dated 11 May 2018, (the “Report”), and I hereby state:

1. My name is Gary Methven, P.Eng., and I am the Underground Manager and Principal Mining Engineer of the firm AMC Mining Consultants (Canada) Ltd. of 202-200 Granville Street, Vancouver, British Columbia, Canada, V6C 1S4.

2. I am a practicing mining engineer registered with the Engineers and Geoscientists of British Columbia (License #180019), a member of the Registered Professional Engineers of Queensland (License #06839), and a member of the Australian Institute of Mining and Metallurgy (CP).

3. I graduated from the University of Witwatersrand in Johannesburg, South Africa with a Bachelor of Science degree in Mining Engineering in 1993.

4. I have relevant experience precious and base metal deposits, mine infrastructure, design and planning, mine production and financial evaluation, reserve estimation, technical reviews, and technical studies.

5. I am a “qualified person” as that term is defined in National Instrument 43-101 (Standards of Disclosure for Mineral Projects) (the “Instrument”).

6. I have not visited the Tucano Project area.

7. I have previously been engaged with respect to the Tucano Project. I undertook Mineral Reserve estimation for the Urucum North underground mine as of 30 June 2017.

8. I am responsible for preparation of Sections 2, 19, 21, and 24, and relevant parts of Sections 1, 3, 15, 16, 22, 25, and 26 of the Report.

9. At the effective date of the technical report, to the best of my knowledge, information, and belief, the above parts for which I am responsible contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

10. I am independent of the Tucano Project pursuant to Section 1.5 of the Instrument.

11. I have read the National Instrument and Form 43-101F1 (the “Form”) and the parts of the Report for which I am responsible have been prepared in compliance with the Instrument and the Form.

12. I do not have a direct interest in the Tucano Project of Beadell Resources Ltd, and I do not beneficially own, directly or indirectly, securities of Beadell Resources Ltd.

Effective Date: 11 May 2018

Signed Date: 17 May 2018

Original Signed and Sealed by

Gary Methven

amcconsultants.com 265

Mineral Resource and Reserve Update for the Tucano Gold Mine Great Panther Silver Limited 718060

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