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An investigation into the use of agglomerated tailings in backfill : a potential tailings disposal option… Chovan, Karen 2001

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An Investigation into the Use of Agglomerated Tailings in Backfill: A Potential Tailings Disposal Option and Case Study for Eskay Creek Mine, British Columbia by Karen Chovan B.A.Sc , The University of British Columbia, 1998 A THESIS SUBMITTED IN PARTIAL FULFILMENT OF T H E REQUIREMENTS FOR THE DEGREE OF M A S T E R OF APPLIED SCIENCE in T H E F A C U L T Y OF G R A D U A T E STUDIES D E P A R T M E N T OF M I N I N G A N D M I N E R A L PROCESS ENGINEERING We accept this thesis as conforming to the required standard T H E UNIVERSITY OF BRITISH C O L U M B I A April 2001 © Karen Chovan, 2001 In presenting this thesis in partial fulfilment of the requirements for an advanced degree at the University of British Columbia, I agree that the Library shall make it freely available for reference and study. I further agree that permission for extensive copying of this thesis for scholarly purposes may be granted by the head of my department or by his or her representatives. It is understood that copying or publication of this thesis for financial gain shall not be allowed without my written permission. Department of M i H A . ^ £ M ><uu » | o C C S ^ C A ^ -The University of British Columbia Vancouver, Canada D a t e Af^ t\ *9 / • 2 ^ 0 \ DE-6 (2/88) ABSTRACT This paper details an investigation into the feasibility of using agglomerated tailings in a 'dry' aggregate fill for underground mining. A test program began with the agglomeration of tailings materials from Eskay Creek Mine, British Columbia, using Portland cement as a binding agent. Thereafter, a series of backfill tests were conducted in which agglomerates, river aggregate, and tailing materials were combined in various proportions to develop a high quality fill. The results of the investigation showed that it is technically feasible to utilize agglomerated tailings in an aggregate backfill used for structural support in underground mining operations. Three related factors reveal their importance: grain size distribution, the degree of compaction and the void content of the mixture. These variables, among others, significantly affect the strength and quality of the fill - for example - the lower the void ratio, the greater the strength of the fill. The strengths of the individual constituents and the water to cement ratio also play major roles in determining the fill strength, but only in those cases where the materials are compacted to minimize voids. The effects are an increase in strength with the increasing strength of the constituents and with decreasing water/cement ratio. Where the void content of the aggregate mix is high, the effect of changing the water to cement ratio is minimal. Although the process of agglomeration is technically feasible, it seems that the associated costs are in opposition to its implementation. An economic analysis of the agglomeration process has revealed that other methods of disposal such as using a tailings pipeline are significantly more attractive from a financial viewpoint mainly because such an option allows the removal of the tailings filter from the process (now in use at the mine). However, the presence of mercury, antimony and arsenic in the tailings raises health and environmental concerns; the need to insure that ARD generation and the release of heavy metals after mine closure is unresolved. In mining situations where ARD generation is an issue, the agglomeration concept may be technically feasible, although the agglomerates may require sealing with a reagent such as sodium silicate, however testwork is required to verify the efficacy of this concept. Where ARD generation is not suspected and where surface or submarine disposal are not viable, the use of agglomeration is a technically feasible alternative. ii T A B L E O F C O N T E N T S A B S T R A C T ii T A B L E O F C O N T E N T S iii LIST O F F I G U R E S ix L I S T O F T A B L E S x A C K N O W L E D G E M E N T S xi 1 I N T R O D U C T I O N 1 2 E S K A Y C R E E K M I N E 3 2.1 S E T T I N G 3 2.1.1 G E O M O R P H O L O G Y & G R O U N D W A T E R R E G I M E 3 2.1.2 C L I M A T E 3 2.2 G E O L O G Y 4 2.3 M I N I N G M E T H O D & M I L L I N G PROCESS 4 2.4 E N V I R O N M E N T A L C O N C E R N S 8 2.5 E V A L U A T I O N OF DISPOSAL OPTIONS 10 2.5.1 O N - L A N D DISPOSAL 10 2.5.2 S U B - A Q U E O U S DISPOSAL 11 2.5.3 U N D E R G R O U N D DISPOSAL AS S T R U C T U R A L B A C K F I L L 1 2 3 T H E A G G L O M E R A T I O N P R O C E S S 14 3.1 T E C H N I Q U E S 14 3.2 C O L D - B O N D A G G L O M E R A T I O N 15 3.2.1 F U N D A M E N T A L S OF B A L L I N G 16 3.2.2 OPERATION CONTROLS 16 3.2.3 M A T E R I A L REQUIREMENTS 18 3.2.4 BINDERS 18 4 R E S E A R C H P R O G R A M A T U B C 20 iii 4.1 T E C H N I C A L O B J E C T I V E S 20 4.2 T E S T P R O G R A M 21 4.2.1 A G G L O M E R A T I O N 21 4.2.2 B A C K F I L L 21 4.3 M E T H O D O L O G Y 22 4.3.1 CHARACTERIZATION OF THE M A T E R I A L S 22 4.3.1.1 Grain Size Distribution 22 4.3.1.2 Moisture Content 23 4.3.1.3 Mineralogical and Chemical Analyses 23 4.3.2 A G G L O M E R A T I O N PRODUCTION & C U R I N G 24 4.3.3 B A C K F I L L PRODUCTION A N D C U R I N G 26 4.3.3.1 Standard Procedures 26 4.3.3.2 Preliminary Tests - Baseline Testwork 27 4.3.3.3 Secondary Tests - Effect of Cement Content 27 4.3.3.4 Tertiary Tests - Effect of Water-to-Cement Ratio 27 4.3.3.5 Final Tests - NEX Materials 28 4.4 R E S U L T S 28 4.4.1 M A T E R I A L PROPERTIES 28 4.4.1.1 Grain Size Distribution and Moisture Data 28 4.4.1.2 Mineralogical and Chemical Data 29 4.4.2 A G G L O M E R A T I O N T E S T RESULTS 31 4.4.2.1 General Findings 31 4.4.2.2 Comparison of Tailing Types 32 4.4.3 B A C K F I L L T E S T RESULTS 33 4.4.3.1 Baseline Studies 33 4.4.3.2 Modes of Failure 34 iv 4.4.3.3 Proportion of Cement in Fill & Pellets 35 4.4 .3 .4 Ratio of Water to Cement 37 4.4.3.5 NEX Materials 37 DISCUSSION 38 5.1 C O N T R O L S & E F F E C T S O N A G G L O M E R A T E S T R E N G T H 38 5.1.1 M E C H A N I C A L CONTROLS OF PRODUCTION 38 5.1.2 M A T E R I A L CHARACTERISTICS 38 5.1.2.1 Mineralogy & Sorption Effects 3 9 5.1.3 B U L K I N G & C O M P A C T A B I L I T Y 3 9 5.1.3.1 Bulking 3 9 5.1.3.2 Compactability 4 0 5.1.4 SIZE & S H A P E OF PELLETS 4 2 5.1.4.1 Pellet Size 4 2 5.1.4.2 Pellet Shape 4 2 5.1.5 W A T E R TO C E M E N T R A T I O 43 5.2 C O N T R O L S & E F F E C T S O N B A C K F I L L S T R E N G T H 44 5.2.1 F I L L DENSITY 45 5.2.1.1 Component Bulk Densities & Grain Size Distributions 4 6 5.2.1.2 Relative Volumes & Resulting Mixture Grain Size Distributions .. 4 7 5.2.1.3 Relating Density to Fill Strength 4 9 5.2.1.4 Relating Density to Degree of Compaction 5 0 5.2.2 W A T E R T O C E M E N T R A T I O 51 5.2.2.1 Relating Water / Cement Ratio to Fill Strength 5 2 5.2.3 F I L L G R A I N SIZE DISTRIBUTION 53 5.2.3.1 Relating Grain Size Distribution to Fill Strength 53 5.3 S U M M A R Y 55 5.4 F I E L D CONDITIONS & O N SITE E X P E R D Z N C E 56 5.4.1 O N SITE Q U A L I T Y C O N T R O L & L A B T E S T I N G 5 6 5.4.2 M E T H O D OF P L A C E M E N T 57 5.4.3 IMPLEMENTATION 5 8 5.5 L O N G - T E R M C L O S U R E ISSUES 59 5.5.1 L O N G T E R M L E A C H I N G 6 0 5.5.2 S U L P H A T E A T T A C K 6 0 5.5.3 A C I D D R A I N A G E POTENTIAL 6 0 5.5.4 M E T A L L E A C H I N G 61 5.6 R E C O M M E N D E D F U T U R E STUDDZS 61 6 COST ANALYSIS 63 6.1 A G G L O M E R A T I O N OF T A I L I N G S & D ISPOSAL I N B A C K F D L L 63 6.2 S L U R R Y P I P E L I N E & S U B - A Q U E O U S D ISPOSAL 65 6.3 C O S T C O M P A R I S O N O F D ISPOSAL OPTIONS 67 6.4 F I N A L S E L E C T I O N O F D ISPOSAL M E T H O D 70 7 CONCLUSIONS 71 7.1 G E N E R A L 71 7.2 A G G L O M E R A T I O N 71 7.3 B A C K F H X 72 7.4 C L O S I N G 72 REFERENCES 73 APPENDIX A - Tailings Disposal Options 80 A General Tailings Disposal Options 81 A.l S U B - A E R I A L D ISPOSAL 81 A . 1.1 S L U R R Y DEPOSITION 81 A . 1.1.1 Ringed Dyke Structures 81 vi A . 1.1.2 Tailings Embankments or Dams 83 Layout 83 Raised Embankment Types 85 Conventional Water-Retention Dam Construction 91 A . l . 1 . 3 Design Considerations : 9 2 A . 1.2 T H I C K E N E D TAILINGS DEPOSITS 9 4 A . 1.3 F ILTERED TAILINGS DISPOSAL 97 A . 2 S U B - A Q U E O U S D ISPOSAL 98 A . 2 . 1 A C I D G E N E R A T I O N PROCESSES 98 A . 2 . 2 D E S I G N CONSIDERATIONS 100 A . 2 . 3 E N V I R O N M E N T A L STUDIES 102 A . 3 U N D E R G R O U N D D ISPOSAL 103 A . 3 . 1 BRIEFING 104 A . 3 . 1 . 1 History 104 A . 3 . 1 . 2 Functions of Backfill 105 Ground Support 106 Ore Recovery 106 Working Platform 106 Waste Disposal 107 A . 3 . 2 STABILIZED M I N E B A C K F I L L 107 A.3^2.1 Cemented Hydraulic Fill 108 A . 3 . 2 . 2 Cemented Rock Fill 108 A . 3 . 2 . 3 Combined Cemented Rock and Hydraulic Backfills 109 A . 3 . 2 . 4 Behaviour of the Fill 110 A . 3 . 3 H I G H D E N S I T Y P A S T E F I L L 112 A . 3 . 3 . 1 Pros and Cons 112 vii A. 3.3.2 Characteristics of Paste Fill 113 A. 3.4 A G G L O M E R A T E D T A I L I N G F I L L 116 A. 3.4.1 Process 116 A.3.4.2 Characteristics of Agglomerate Tailings Paste Fill 117 A.3.4.3 Agglomerated Tailing Aggregate Fill 118 APPENDIX B - Material Properties 119 APPENDIX C - Peptization Test Results 126 APPENDIX D - Backfill Test Results 160 • APPENDIX E - Cost Analysis 179 PHOTOGRAPHS 185 viii LIST OF FIGURES Figure 2.1: Typical Stope for Underground Mining 5 Figure 2.2: Overhand Mining 6 Figure 2.3: Underhand Mining 7 Figure 3.1: Disc Pelletizer & Flow Paths 17 Figure 4.1: Formulae for Maximum Stress on a Spherical Object 25 Figure 4.2: Uniaxial Compression of a Cylinder 27 Figure 4.3: Grain Size Distributions of Materials - As Received 29 Figure 4.4: 28 Day Backfill Strength vs. % 109 Tailings, 9% Cement 34 in Matrix Figure 4.5: 28 Day Strength vs. Cement Content, 100% Aggregate Backfill 36 Figure 4.6: 28 Day Strength vs. WatenCement Ratio for a 60R:40P Mixture 37 Figure 5.1: Packing Mechanisms of Random Particles with Clay 41 Figure 5.2: Stress Distributions in Equidimensional Objects 43 Figure 5.3: Pellet Strength vs. Water-Cement Ratio, 28 Days, Both Tailings 43 Figure 5.4: 28 Day Backfill Strength vs. Density, All mixes - 46 including variable cement Figure 5.5: 28 Day Backfill Strength vs. % 109 Tailings, 9% Cement 47 in Fill Matrix Figure 5.6: 28 Day Backfill Strength vs. Density, All mixes - 50 excluding variable cement Figure 5.7: Strength vs. Water/Cement Ratio, 28 Days 51 Figure 5.8: Strength vs. Water/Cement Ratio & Grading 54 Figure 5.9: Batching Process Control Chart, 7 Day - 1999 56 Figure 5.10: 28 Day Strength 57 Figure 6.1: Option #1: Agglomeration & Backfill 65 Figure 6.2: Option #2: Pump & Pipeline System 67 ix LIST OF TABLES Table 4.1: Elastic Modulii vs. Cement Content of Various Paste Fills 25 Table 4.2: Predominant Minerals & Minor Constituents within 30 Tailings Table 4.3: Predominant Minerals & Minor Constituents within 31 Tailings Table 4.4: Pellet Strengths at 7 & 28 Days 33 Table 6.1: Agglomeration Option Capital & Operating Costs 69 Table 6.2: Slurry Pipeline Capital & Operating Costs 69 Table 6.3: Summary of Costs 70 x ACKNOWLEDGEMENTS I would like to take this time to thank the numerous people who helped and supported me throughout my studies. First, Dr. John Meech, who as my advisor, spent many hours helping to organize the test work, providing information and assistance where his expertise was required, discussing the progress of the testing program and report writing, and reviewing the many revisions of this paper that came before the final product presented herein. Second, Brig Sigismund and Gerry Rogers of Eskay Creek Mine, as well as the personnel at Homestake Canada who initiated this project. Not only did they provide the materials and scope of work for this project, but also answered many questions about the mine site and the mining methods used on site, as well as provided a large amount of data from test work performed on site and by outside consultants for comparison of results. Dr. Mory Ghomshei commenced the original testwork, and provided a great deal of assistance throughout the remainder of the testing program, as well as his knowledge and support when it was needed. Many other individuals assisted with their invaluable expertise in backfill, tailings fill, concrete research, environmental considerations, and economics. Each specializing in their own field, these people included Dr. Rusty Morgan, Dr. Mike Davies, Dr. Rimas Pakalnis, Dr. Marcello Veiga, Dr. Nemkumar Banthia, Joe Smalcel, and my parents, Mike and Silvia Wolff (who also provided enormous support). Finally, my husband Mike Chovan, who has always supported me and loves me dearly, but who also made me sit at the computer until it was done ©. Without him, I would never have gotten this far. xi 1 INTRODUCTION Mines are a necessity of life, even though in the short term, they may not be aesthetically pleasing or particularly good for the environment. Mining is a temporary land user, engaged in the depletion of non-renewable resources and historically has often had adverse impacts on contiguous air, water, and land, as well as wildlife within a given area. Each mine is unique in terms of its deposit location, local environment, ecological setting and finite operating life. Historically, mining wastes have been disposed "at the nearest convenient location," with the consequences of such actions never considered (Klohn, 1995). Tailings, which result from the liberation of economic minerals, were dumped directly into nearby lakes and streams or earth structures of convenience without investigating the physical or chemical stability of such materials, or the effect they might have on the environment surrounding a disposal site. Mine tailings consist of barren, highly variable, fine-grained (typically > 50% finer than 200 mesh or 75 microns) and uniformly graded particles, often fluid or semi-fluid, and in many cases, toxic materials are abundant. The potential pollution hazards associated with the storage of the tailings slurry vary with different mining operations, ranging from very severe for the radioactive wastes from uranium mining, to none for those processes that grind up relatively inert ore without adding any toxic reagents. Recently, new technologies and modern milling processes have been developed such that it is now economic to mine low-grade ores that previously were left undeveloped. In extracting low-grade ores, mine tailings are now produced in much higher amounts with respect to the same production of values from the ore. Today, legislation and new regulations mandate that mines must be designed to anticipate, mitigate and resolve environmental conflicts before they arise and operate with a constant view to ultimate closure of the site. Eskay Creek Mine is a relatively small mining operation owned and operated by Homestake Canada, located in northern British Columbia, which has both high-grade and low-grade ores. The high-grade ore is mined and crushed at a rate of 250 tonnes per day and shipped directly to smelters. The low-grade ore is processed through a flotation circuit on site at a rate of 150 tonnes per day, and tailings are then disposed of sub-aqueously in a lake located 8 km away. Because the operation utilizes underground backfill for structural support during mining, the option of placing the tailings underground became an interest of Homestake officials. However, 1 due to the drift and fill method of mining utilized at the site, the use of paste fill is not applicable. For this reason, the concept of tailings agglomeration was studied. This thesis investigates the technical and economic feasibility of agglomerating tailings and using them in a 'dry' aggregate backfill for underground mining. The purpose of the research is to find a method to dispose of tailings in an underground environment, while utilizing their volume for structural support, in cases where the pumping and placement of a slurry or paste fill is not applicable. The study includes investigating various agglomeration processes; developing strength testing methods for said agglomerates; and incorporating these agglomerates into various aggregate fill materials. A number of fill mixtures were evaluated in which combinations of the agglomerates, tailings, river aggregate and cement contents were varied. All batch test results were analyzed to determine the effect of variables on fill strength. Eskay Creek Mine provided the tailings and criteria to be used for the research program. Placement options and economic scenarios were compared to the use of such backfill at this mine site. 2 2 E S K A Y C R E E K M I N E The Eskay Creek property, wholly owned and operated by Homestake Canada Inc. (HCI), hosts a medium-sized, high-grade gold, silver and base metal deposit. The mine commenced operations in January 1995 and its life has recently been estimated between another 12 to 15 years (Sigismund, 2000). At the beginning of 1998, there was a proven and probable mining reserve of 1.08 million metric tonnes grading 65.5 g Au per tonne and 2,930 g Ag per tonne (Roth et al, 1997). 2.1 S E T T I N G The mine is located in the Skeena Mining Division of British Columbia, situated between the Iskut and Unuk Rivers, approximately 83 km north east of Stewart, and 1000 km northwest of Vancouver. 2.1.1 G E O M O R P H O L O G Y & G R O U N D W A T E R R E G I M E The property lies within the Unuk River watershed on the top of the Prout Plateau, which forms the watershed divide of the Unuk and Iskut Rivers. "Prominent gossanous bluffs" rise above the headwaters of the Iskut River, with elevations ranging from less than 700 m to approximately 1400 m above sea level in the immediate vicinity of the site. To the north, the Iskut River valley drops to well below 600 m in elevation as the river begins to grow in size, while further to the west of the site, peaks rise to ultimate heights of 1800 m. There are a number of local peaks and gullies running predominantly north to south, forming numerous drainage channels, creeks and small lakes in gullies and natural depressions. The significant volume of water that collects within the large catchment area ultimately drains into the Iskut River to the north and the Unuk River to the south during the higher rainfall and snowmelt seasons. Both the Unuk and Iskut rivers originate in Canada and flow in a westerly direction through Alaska before discharging into the Pacific Ocean (Hemmera, 1997). 2.1.2 C L I M A T E The climate consists of coastal weather with short summers and long winters. Precipitation levels are moderate to high in comparison with other regions of Canada, and the snowfall season 3 extends from the beginning of October until well into June, accumulating up to 20 m in depth. The average winter temperature drops down to average 0°C, with a low of -10°C. By mid-June, typically no snow remains on the ground within camp; however, snow does stay on the slopes 100 m above the site for longer periods. The summer season entails high rainfall with moderate temperatures averaging between 15°C to 20°C. 2.2 GEOLOGY The Eskay Creek deposit is a volcanogenic massive sulphide (VMS) deposit with some later faulting and hydrothermal enrichment. The deposit is hosted by volcanic and sedimentary rocks, and "comprises an upright stratigraphic succession of andesite, marine sediments, intermediate to felsic volcaniclastic rocks, rhyolite, mudstone hosting stratiform mineralization [termed the contact mudstone], and basaltic sills and flows which are intercalated with turbiditic mudstones. The lower part of the sequence is intruded by porphyritic monzodiorite to diorite and by younger felsic dikes and sills that are feeders to the rhyolite package. Basaltic dikes and sills feeding the hanging wall basaltic package crosscut all strata in the sequence. Regional metamorphic grade in the area is lower greenschist facies" (Roth and Rogers, 1998). Mineralization is generally strata-bound, hosted in the contact mudstone. Sphalerite, pyrite, galena and tetrahedrite are the most abundant ore minerals and native gold occurs mainly as microscopic particles between sulphide grains or within pyrite. The North Extension (NEX) Zone occurs at the base of such a stratigraphic contact, with local overprints of chalcopyrite stringers. Visible gold also occurs together with coarse-grained sphalerite, galena, minor pyrite and chalcopyrite in a stockwork of quartz veins running through brecciated rhyolitic rocks beneath the contact mudstone. This is prominent in the 109 Zone (Roth and Rogers, 1998). 2.3 MINING METHOD & MILLING PROCESS All underground workings at Eskay are accessed through three portals and a series of adits and ramps. The ore deposit is mined using conventional drift and fill underground mining techniques (see Figures 2.1 through 2.3), including: 1. Overhand drift and fill mining with delayed 4% cemented rock fill. 2. Underhand drift and fill mining with delayed 9% cemented rock fill. 4 Figure 2.2: Overhand Drift & Fill 6 Sill Drift (mined out and backfilled) WASTE WASTE Access Ramp WASTE Establish new access to g am Sag undercut previous lift Cross Cut Driven at 0% Access Ramp Cemented fill wMsm WASTE Active Mine Heading Cross Cut Driven at -15% Access Ramp Figure 2.3: Underhand Drift & Fill The mine uses cemented backfill at a rate of approximately 400 tonnes per day. The backfill is produced in 3000 kg batches of river rock aggregate (retrieved from the Iskut River approximately 30 km away and screened to minus 4 inch), Portland cement and water. Water 7 addition is dependent on the moisture content of aggregate. Each batch is loaded through a GOB hopper into 15-ton underground trucks, each truck holding 3 batches. The materials are trucked to the designated stope and placed with the use of a "rammer jammer," a device on the front of the bucket of a front-end loader used for ramming purposes (Homestake, 1999). Geotechnical testing has been performed at the mine site on various mixtures of the fill, and altering proportions of cement and water are utilized in order to attain suitable fill strength for the various applications at the mine. In order to provide sufficient support for neighboring underground cuts, the mine requires minimum fill strengths of 4 MPa for its underhand fill volumes, and 2 MPa for its overhand fill volumes (Sigismund, 1998). To attain such strengths, the fill requires curing periods ranging from 24 hours up to 7 days for the overhand and neighboring sequences and up to 28 days for underhand sequences prior to the excavation of materials in neighboring stopes (Sigismund, 2000). It is the required rate of strength gain and method of fill placement that prevents the use of paste fill for this mining operation. Of the ores extracted at the mine site, approximately 250 tonnes per day is high-grade ore, which is crushed and directly shipped to smelters in Japan and Quebec, Canada. Another 150 tonnes per day is a lower-grade ore, which is processed through an on-site flotation circuit designed to operate at natural pH to maximize recovery of the sulphides and associated precious metals (Dai, 1999). The resulting tailings are thickened to approximately 55% solids in a thickener. The underflow is piped to a storage tank for up to 7 hours (capacity of tank for production of tailings at nominal plant feed rates). The tailings are then pumped to and filtered in a PF60 Larox Pressure Filter to form a 'dry cake' that is stockpiled temporarily on the floor below the filter area before being re-handled by loaders into trucks. The tailings are hauled by truck to a lake (Albino Lake) located about 8 kilometers from the mill. 2.4 ENVIRONMENTAL CONCERNS "During the mine development application review, oxidation of reactive rock and release of heavy metals into the receiving environment were identified as the main long term environmental risks associated with the mining operation and closure" (Hemmera, 1997). Heavy 8 metals such as antimony, mercury and arsenic are associated with the precious metals mined on site (Roth et al., 1997). The mine is committed to minimizing environmental impacts due to mining activities on the area, using an established monitoring program, on-site analyses, regular inspections and reports, water quality modeling, extensive recycling and training. To date, the mine has met most of the stringent effluent requirements proposed by legislation and there has been little impact of mining activities on the surrounding watercourses. In fact, "the mine has received the British Columbia Mining Association and the Ministry of Employment and Investment Exploration Award for work conducted" on site (Homestake, 1999). However, both of the two main river courses mentioned above have been designated International Waters under the Boundary Waters Treaty Act and the International River Improvements Act. "Of the two, the Unuk River is the main watershed that may be impacted by the Eskay Creek Mine operations...Discharge from the permanent waste rock storage site (Albino Lake) and the mine water discharge flow into a tributary of the Unuk, an important fishery1" (Hemmera, 1997). The site is also a very scenic locale, where it is not uncommon to see moose, black bear, grizzlies, wolves, porcupine, eagles, grouse and mountain goats. Although HCI recognizes the need to ensure that the water quality of the headwater lakes and tributaries of this system are protected, disposal of tailings into a subaqueous environment, as is the case now, seems contradictory to this concept. The tailings at Eskay Creek have high sulphide contents and low soluble carbonate contents, with overall acid-consuming to acid-generating ratios (NP/AP) of -1:6 and -1:1.5 for the 109 and NEX tailings, respectively, and are therefore potentially acid generating (Dai, 1999). Also, mercury can be largely detrimental to aquatic life and wildlife if present in abundance and in an "available" form, where it is free to react with organic and non-organic compounds in the environment, and then be taken up by aquatic plants and animals. This is a subject that was not investigated in detail at this time. If water is available as a transport medium, products of potential acid generation processes, such as elevated concentrations of metals and sulphate, could be imparted on receiving environments located down-gradient. 1 As a note, while spawning and rearing are known to extend as far upstream as Storie Creek, no fish have been found in the upper tributaries in the vicinity of the Eskay Creek Mine site, or in their headwater lakes, including Albino Lake. Fish were first found 7 to 8 km downstream of the proposed mine site during the fisheries studies conducted in support of the Mine Development Application" (Hemmera, 1997). 9 2.5 E V A L U A T I O N O F D I S P O S A L O P T I O N S Eskay Creek's mill throughput of 150 tonnes per day converts to approximately 55 thousand tonnes of tailings per year, and a total of .820 thousand tonnes of tailings over the expected life (15 years) of the mine. In terms of their size distribution, the materials are classified as fine sand to clay material. Regulatory conditions are such that the tailings may be disposed of in various fashions, given that the short and long term impacts to the environment are minimal, physical and chemical stability risks are low, and reclamation activities are carried out on an on-going basis where possible, and completely for final closure. Monitoring activities must also be carried out on an ongoing basis, and reported to governmental agencies, to prove that satisfactory conditions are being met. Various options for tailings disposal, and their design requirements and methods of construction are discussed in detail in Appendix A. Because the tailings in question are potentially acid generating (PAG) and contain heavy metals, the disposal method must ensure their complete isolation from the surrounding environment. The system must also be closed indefinitely. The options applicable at the Eskay Creek Mine site include: • on-land disposal in a tailings impoundment as a slurry or thickened deposit, • placement of filtered tailings on side-slopes or in an impoundment of reduced size, • sub-aqueous disposal in a near-by lake, and • underground disposal as a component of the backfill currently used for structural support. In the following sections, these options are discussed specifically with respect to the general site constraints and concerns for each. 2.5.1 ON-LAND DISPOSAL The on-land disposal options, including both the impoundment of tailings as slurry or as a thickened deposit and the placement of filtered tailings on side-slopes, may be dismissed almost immediately for this site. These methods of disposal face the problems of high surface runoff volumes in a large catchment area, and the problem of determining an appropriate, structurally 10 sound location for embankment construction or stacking of materials in such mountainous terrain. The location of the disposal site would need to be selected for its economic feasibility of development and its distance from the mine site and would also require sufficient area and volume for long-term tailings storage requirements. Long-term monitoring requirements would include on-going stability checks of the impoundment or tailings deposit, and extensive water quality maintenance to ensure local waterways and groundwater are not contaminated from acid generation and/or metal leaching from the tailings. 2.5.2 S U B - A Q U E O U S D ISPOSAL Sub-aqueous disposal of tailings has been proven to be one of the safest means of disposing of acid-generating tailings and waste-rock materials. This is because the submergence of the materials under water virtually eliminates oxidation of the sulphide minerals that produce the acid (see Appendix A). By removing the potential for acid generation, the release of acid-soluble heavy metals is also essentially eliminated. During the permitting stages of the mining operation, Albino Lake was approved as an appropriate site for disposal of acid generating waste rock materials that would be produced by the mine. The lake does not drain completely during the low-flow season, allowing continuous submergence of the waste materials throughout the entire year. When the mine commenced milling in 1998, this body of water also became the site for tailings disposal. The mine began hauling the filtered tailings by truck to the disposal site and dozing the materials into the lake. However, with the extended life expectancy of the mine operations, it became apparent that there would be insufficient depth and volume within Albino Lake to continue disposing waste at this location (Sigismund, 1999). Because this became an approved method of disposal at this site once before, it would seem logical that it could be approved once more at another location near the mine. However, another lake of sufficient storage capacity would first need to be located within reasonable distance from the mine site to make it economical. Due to the topography, there are a limited number of sites that fall into this category. Second, the mine would also need to carry out extensive biological and environmental studies characterizing the status of the selected lake and surrounding area, in 11 order to develop relative water quality guidelines and see that they can meet the objectives required at the new disposal site. Finally, because a 'change' in disposal was already required, the mine also wanted to consider an option where it could eliminate the Larox pressure filter from its milling process to reduce operating costs. This would require the installation of a pipeline system to pump the thickener underflow directly to the disposal site. The first important factor that could affect the costs of such an installation (due to the pumping requirements) would be the rise in elevation between the mill and the disposal site. The costs associated with pumping could render this option economically infeasible. The second factor that would require consideration would be the affect that the quality of the tailings slurry water could have on the disposal environment. The more fluid medium (than the current 'dry' filtered tailings) would take longer to settle out in an environment that was not completely still, and have a tendency to flow further. Tailings could potentially be transported out of the lake system in this way. Also, there is the potential for the slurry water to not meet the required water quality guidelines - the fluid had previously been filtered out of the tailings (for the most part) and treated and/or re-used in milling processes. The effect of the slurry water on the receiving environment was not investigated during these studies. A benefit to consider for the piping option, however, would be the relatively limited reclamation requirements for this method - only the eventual dismantling of the pipeline and deactivation of the pipeline route and service road would be required. Long-term monitoring would focus mainly on water quality issues, with respect to metals leaching and transport, as well as rehabilitation of the aquatic flora and fauna. 2.5.3 U N D E R G R O U N D D ISPOSAL AS S T R U C T U R A L B A C K F I L L The final option for tailings disposal at Eskay Creek is a method in which the tailing materials are disposed of into the underground workings, as a major component of the backfill used for structural support in the mine. Setting aside costs for the moment, this option would seem to be very appropriate for this operation, as it would allow the mine to dispose of its waste in an already disturbed area - an area which requires filling to provide support for the continuation of mining; a method in which there is no disturbance to the natural surface of the terrain. 12 Prior to contacting UBC to perform this research, Eskay Creek Mine obtained the services of various consulting and supply companies to produce and test backfill mixtures whereby filtered and thickened tailing materials were mixed with cement and/or additives and aggregate materials for placement underground. Several constraints pertaining to this method of disposal became apparent. While the idea seems logical, production of suitable backfill using tailings material is not easy, and requires a large amount of cement in comparison to the regular backfill to obtain even moderate strength values. The fineness of the materials presents problems with handling in the batch plants, trucking, and in placement; the mixture of tailings and cement requires more water than regular fill, causing it to slump with little disturbance, yet it remains very viscous, making it more susceptible to sticking. For this reason, the option to agglomerate the tailings before incorporation into the fill was considered as a possible technique to overcome these difficulties. By agglomerating the tailings, the waste material is incorporated into the aggregate component of the backfill, and handling problems are reduced. The process of agglomeration is discussed in Section 3. In Sections 4 and 5, a research program carried out to determine physical characteristics and constraints of the system is discussed in detail. Recommendations for future studies are also mentioned. In Section 6, the economic feasibility of carrying out such a plan is reported on. This analysis is compared to the current operation, and to that of pumping a thickened slurry mass to a new disposal site. Section 7 contains the conclusions of this study. 13 3 T H E AGGLOMERATION PROCESS The definition of agglomeration, according to the American Heritage Dictionary of the English Language, is the act or process of gathering into a mass. In mining, the agglomeration process is a method used to compact or consolidate fine, unmanageable materials into larger sizes that can be handled more easily. Some common examples include iron ore peptization for ease of transport and reduction in the blast furnace and agglomeration for heap leach operations, allowing for more stable heap leach stacks and recovery of ore from otherwise overly fine, unusable materials (due to associated handling and storing problems). 3.1 TECHNIQUES "Briquetting, tableting, compaction, extrusion, pelletizing, balling, sintering, nodulizing, granulation and agglomeration are words that describe methods to consolidate particulate solids" (Messman, 1977). Some of these terms are employed more or less interchangeably. There are three main techniques implemented in the consolidation of loose materials. These include: • mechanical application of pressure to compress fine materials into a dense mass, • reliance on chemical reactions between particles and/or additives, or the reliance on a physical change in a material's properties with the application of heat (both of the latter techniques cause bonding processes to develop), and • balling operations, where materials agglomerate due to the presence of weak physical forces (van der Waal bonds) between particles and water (surface tension), or by some other wetting/lubricating agent. For each of these processes, there is also the option to add some form of binder to increase the strength of the resulting product. The following paragraphs were derived from a summary of agglomeration processes, written by Messman in 1977. Mechanical consolidation mechanisms include briquetting, tableting, compaction and extrusion. Briquettes or tablets consist of small bricks of whatever shape, formed by mechanical pressure in 14 a mold, a roll briquette machine, an extruder, or some other device. Compaction is the mechanical consolidation of solid particles between rolls, or by tamps, pistons, screws or other means. Extrusion implies the forcing of a material through an aperture or die, including pelletizing by pellet mills. Pellets are generally smaller than briquettes. Granulation usually involves consolidation of loose materials into large compacts by mechanical application of pressure, followed by the milling of these compacts into grains, by knife-mills, corrugated rolls or other machines designed to generate minimum recycle fines. Sintering is heat bonding where a small portion of a particulate solid fuses to form a tacky binder at elevated temperatures (iron ore). Nodulizing is the simultaneous application of sintering and drum balling, usually in a rotary kiln. Finally, agglomeration implies consolidation of solid particles into larger shapes by means of agitation alone, without application of pressure. Snowballing is a well-known natural form of agglomeration, implemented in balling devices such as rotating discs, drums or cones. Some examples of the above, as used in the mining industry, include application of very high heat in the process of iron ore pelletization, bonding through oxidation mechanisms commonly used for heap leach stacks and some applications of underground mining fill, and cold-bond agglomeration, where cement and/or other binders are added to loose tailing materials to improve the strength of 'green' balls or agglomerates developed by balling techniques. For this project, the Eskay Creek tailing materials were consolidated using the cold-bond agglomeration technique, with addition of ordinary Portland cement as binder, and the use of rotating drums and discs. 3.2 COLD-BOND AGGLOMERATION Two commonly used methods for cold-bond agglomeration include drum and disc balling or pelletization. Drum balling is the oldest method, and is essentially the same process as that of disc balling. However, disc balling produces better ball-size classification, and involves fewer steps than does drum balling. Drum balling includes balling, double screening, crushing of the coarse fraction and recycle of the crushed material with screen fines (Messman, 1977). 15 3.2.1 F U N D A M E N T A L S O F B A L L I N G One of the earliest and most important contributions, recognized by Newitt and Conway-Jones, was that the packing characteristics of the feed material combined with the cohesive forces of moisture addition are fundamental to the theory of balling. Observations also support that "the capillary force of the added water and the mechanical force achieved by the balling operation are responsible for the binding of green balls." Today, mechanical and capillary forces are generally recognized as the two major considerations in the balling of all materials (Floyd and Engelleitner, 1977). The forming of a ball is a two-stage process - formation of a nucleus, followed by a ball growth period (Messman, 1977). The initial stage entails formation of a weakly bonded 'seed', developed within a system comprised of air, liquid and solid. "It is necessary that the individual particles be wetted and conditions be such that coalescence of the fines can occur" (Floyd and Engelleitner, 1977). Then, "adhesion forces between the particles act upon each other to create a growth process... as the particles are wetted, thin bridges form on the surface of the pellets which create bonds between the fine particles" (Pietsch, 1977). If there is a deficiency in moisture or the particles are of a nature such that a cohesive force between particles cannot develop, coalescence of the fines will not occur. The second phase of the process depends on capillary bonding within a liquid-solid system (Floyd and Engelleitner, 1977). Ball growth is attributed to coalescence (the breakup of geometrically unstable agglomerates), layering (fine particles form a layer on an existing ball to make it larger), abrasion and transference of material between balls, and fragmentation resulting from contact with parts of the balling mechanism. Size reduction mechanisms such as shattering, breakage and attrition also occur which also affect the agglomeration process (Pietsch, 1977). 3.2.2 O P E R A T I O N C O N T R O L S "Optimum operation of the disc is characterized by the formation of a distinctive pattern of free flowing fines, agglomerates and finished balls...The actual make-up of the three streams depends on the product sizing desired and the relationship between the nucleation and growth characteristics of the material" (Floyd and Engelleitner, 1977). Path #1 contains the finished and near-finished balls or pellets, path #2 contains in-between agglomerates, and path #3 holds fines 16 and seeds. The three flow paths, distributed across the surface of the disc, are illustrated below in Figure 3.1. "The disc can produce a wide range of product sizings from a given feed with the versatility required to maintain the selected sizing despite variations in feed conditions which occur in normal plant practice. Four features account for this: slope of the disc, speed of the disc, point of feed application, location of water spray(s)" (Floyd and Engelleitner, 1977). The slope and disc speed are related variables. Speed variations are required to produce the right degree of rolling action and the proper formation of flow paths while the slope of the disc face controls the volume of material on the disc - the shallower the angle of the disc, or the more horizontal its position, the more material available for agglomeration. The size of the ball or pellet is directly related to the positions of the feed and water. If these are both placed towards the outer rim of the disc, near the third path, the pellets remain small. As the pellets gain mass, they drop out of the flow path and continue to roll and tumble until flushed off the disc at the discharge side. Larger product size results when the feed point is moved to the discharge side (Floyd and Engelleitner, 1977). "The amount and location of the water addition also have a direct bearing on the size of the finished product, the efficiency of the pelletizing operation, and the quality of the green balls" (Floyd and Engelleitner, 1977). For this reason, it is beneficial to the pelletizing operation when the gradation of the feed material remains within set parameters - the pellets produced should not Rotation Discharge Figure 3.1 Disc Pelletizer and Flow Paths 17 vary significantly in strength or quality with small changes in gradation if the operator has the ability to adjust the addition of water accordingly. 3.2.3 M A T E R I A L R E Q U I R E M E N T S During their studies, Floyd and Engelleitner (1977) discovered that "extremes in particle shape (laminar, rod and spherical) often present pelletizing difficulties and result in products lacking good physical properties." They also noted such particles are inherently closely sized, often because of previous processing steps. Messman (1977) noted that "materials that are hard and brittle, or elastic, are difficult to bond without a binder, even when comminuted to relatively small particle sizes, but optimum size consistency can be helpful." The production of a quality green ball requires feed materials to have good packing characteristics - a property which is precluded by irregular particle shape (rod or laminar) and uniform size distribution. Therefore, Floyd and Engelleitner (1977) suggested grinding of the feed materials prior to agglomeration, in order to produce a greater range in particle sizes to improve the density and quality of the agglomerates. Studies in the concrete industry have however shown that packing of angular particles is less consistent and of a poorer quality than that of particles of similar size gradation that are rounded to sub-rounded (Neville, 1995). It was therefore recommended that the main proportion of the aggregate be rounded in shape. Floyd and Engelleitner (1977) determined the minimum requirements for steady state disc production of a closely sized product. For an average pellet diameter of 12 mm Q/i inch), the feed material should entirely pass through a 16-mesh sieve, contain, at least 20% passing 200-mesh and have a fairly regular gradation of sizes. The production of smaller pellets generally calls for a decrease in the allowable top size, and an increase in fines content to more than 40% passing 325-mesh. Balling can be done at any typical feed moisture for most materials, from dry to slightly above ideal pelletizing moistures (typically between 10 to 16 % by weight). 3.2.4 B INDERS "Any solid material can be compacted without binder, if sufficiently high pressure is applied...to literally smash their initially irregular surfaces together and create relatively large areas of common surface with minimum voids" (Messman, 1977). Cohesion of the individual grains of the originally bulk material is effected by van der Waals' forces binding the grain surfaces. 18 However, "the sphere of action of these forces is extremely small, making it necessary to move the grain surfaces very close together in order to produce the effect of the van der Waals' forces" (Rieschel, 1977). The addition of a binder improves the cohesiveness, provides tensile strength and increases the stiffness of a mix. "Cohesion...is caused by the admixture of an adhesive which, after careful mixing, coats every individual grain of the material with a thin film. The thickness of the binder film is many times greater than the spheres of action of the van der Waals' forces. For this reason, a lower degree of consolidation of the material is adequate for manufacturing since...only the binder films of the individual grains have to be brought into contact with each other" (Rieschel, 1977). Binders commonly used in the mining industry are Portland cement (PC), bentonite, fly ash or blast furnace slag having pozzolanic properties. 19 4 RESEARCH PROGRAM AT UBC In May 1998, The Department of Mining of the University of British Columbia was commissioned to perform preliminary testwork to investigate the incorporation of mine tailings into a cemented aggregate backfill to provide stability for Eskay Creek's underground mining operation. The mine invited UBC to find a way of using the tailings as a replacement or supplement of the river aggregate currently used in the mine backfill, while maintaining the fill's handling characteristics. Eskay Creek had already commissioned Golder Associates, Tilbury Cement, Lafarge and Target Products to carry out tests in which tailing materials were combined in a paste form with aggregate in varying combinations. Results of these preliminary laboratory tests and experience at the mine site demonstrated limitations in handling and mixing paste tailings into the backfill as well as the need for additional cement if strengths similar to that of the river aggregate fill were to be achieved. 4.1 TECHNICAL OBJECTIVES The goal of this study was to produce agglomerated tailings (also termed pellets) with sufficient strength to replace all or some of the river rock aggregate used in the mine's current backfill. At the same time, the new 'agglomerate' fill needed to meet minimum strength criteria for the underground mining operation to continue - that is, minimum strengths of 2 MPa for overhand fill and 4 MPa for underhand fill. The incorporation of agglomerated tailings into the backfill would allow for underground disposal of the tailings, reducing the need for surface and/or sub-aqueous disposal of the potentially acid-producing materials, without significantly changing the backfill handling characteristics. The research consisted of four parts: • characterization of the materials, • agglomeration of the tailings, • incorporation of agglomerated tailings into backfill mixtures, and • measurement of the strengths of these mixtures. 20 The potential for the generation of acid rock drainage (ARD) from the tailing materials was investigated in a second study at UBC (Dai, 1999), in concurrence with this work. The results of the ARD study are referred to at times here, but are not discussed in detail. 4.2 T E S T P R O G R A M 4.2.1 A G G L O M E R A T I O N Various mixtures of dried tailings, cement and water were combined in preliminary pellet production runs to observe the minimum requirements for their development. Initially, a drum pelletizer was constructed, where the speed and angle of the drum could be adjusted to various settings to control the rate and size of agglomerate production. Due to problems with producing consistent, quality pellets using the drum pelletizer, the program resumed using the disc pelletizer. The disc pelletizer could handle a higher feed rate, was easier to control and produced consistent and strong agglomerates. Other advantages of using the disc pelletizer included its ability to adjust several variables: disc rotation speed, angle of the disc surface and water addition rate. As well, the feed rate and points of feed and water addition could be easily adjusted. By trial and error, these parameters were tested to find the most efficient settings for the desired pellet size distribution and quality. Once pellets of various strengths had been developed, a number of batches were produced for use within the backfill test series. Pellet strength as a function of cement and water content was studied at a variety of curing times, and the effects of grain size distribution, average pellet size and aspect ratio were also examined. 4.2.2 B A C K F I L L Three main sets of tests were performed. 1. The prekminary backfill samples were made up of varying combinations of aggregate including: river rock, pellets and unagglomerated tailings, with the same concentration of cement. These tests were carried out to investigate the effect of replacing some of the river aggregate on the fill's strength. To minimize segregation effects on strength, the fill batches were produced at similar consistencies, measured by slump values. 2 1 2. Replicating a few of the aggregate combinations above, a second set of tests was carried out using varying cement contents, to investigate the effect of cement on fill strength. Again, the consistencies of the fills were maintained at a constant slump value. 3. On the third set of tests, one aggregate combination was produced using a constant cement concentration while the water content was altered. These tests were performed to determine the effect of the water to cement ratio on fill strength because, up to this point, constant slump values had been used for all tests. All of the tests were carried out to analyze the details of fill strength - what factors affect the strength and how. 4.3 METHODOLOGY 4.3.1 C H A R A C T E R I Z A T I O N O F T H E M A T E R I A L S Due to the filtering process, the tailings materials, as-received, were in a consolidated state. However, the clumps of material could be broken-up between the fingers, sometimes with little or no pressure. Samples were split and quartered, and then portions were dried in a vacuum chamber at ambient temperature. The materials were passed through a mechanical grinder or shaker for approximately 5 minutes with 15 to 20 steel balls on each screen to break apart the consolidated materials, and bagged in small proportions to reduce the risk of contamination, and segregation or loss of fines. Portland cement, Type #10, was chosen as the sole binder to add strength to the proposed agglomerates and backfill mixtures following consultations with personnel at Homestake. Preliminary analyses of the materials included particle size analyses and measurement of moisture contents of the tailings and river aggregate. Mineralogical and chemical analyses were also carried out on the filtered tailings. 4.3.1.1 Grain Size Distribution The particle size distributions of the materials were determined in a number of ways. The size gradation of the river rock aggregate received by UBC was measured using the dry sieve technique, as the fineness of the material was visibly very low. After drying, each sample 22 was weighed and passed through a number of progressively finer sieves. The material remaining on each sieve was measured, and the results graphed on a semi-log scale graph of cumulative percent passing versus mesh opening size in microns. Being fine-grained, the distributions of the tailing grain sizes were determined by the 'wet' sieve method. A dried sample was weighed before being washed over a Taylor series #400 sieve to remove all particles finer than 37 microns in size. The remaining sample was then re-dried in the oven and weighed before being passed through a number of progressively finer sieves, in a similar fashion as mentioned above. In order to determine the particle size distribution of the tailing particles below 37 microns, a hydrometer test was conducted. Fine materials were mixed with a dispersing agent to separate agglomerated micro-fine sized particles and after some time, water was added to the slurry. The mixture was then agitated and set aside to allow the settlement of the particles. At various stages of the settlement period, the density of the mix was measured, indicating the remaining amount of the grains in suspension. Later, this information was converted into its respective grain size data, and the grain size distribution was extended to include particles down to a size of 1 micron in diameter. The particle size distribution of the cement used was extracted from various sources used by the concrete and cement industry (Detwiler, et al, 1996). Although there is documented proof of physical and chemical variation of cement mixtures from various production mills, it was assumed that the distribution of the cement used in the test program would correspond with those in the published standards. 4.3.1.2 Moisture Content For moisture content determination, samples of suitable volume were extracted from the bulk materials as soon as they arrived. The samples were weighed and placed in a moderately heated oven to dry, then re-weighed for calculation of the as-received moisture content. 4.3.1.3 Mineralogical and Chemical Analyses The mineralogical and chemical analyses were carried out through a number of tests. Tailing samples were sent to an assay lab to perform ICP tests to determine the chemical constituents within the samples. X-Ray Diffraction (XRD) was carried out using a Siemens Diffraktometer 23 D5000, to expose the identity of the predominant minerals present in the tailings. Analyses from reflection microscopy (RM) and a Philips XL30 Scanning Electron Microscope (SEM) were used to determine the relationships of the various minerals to one another. Also performed were general fizz tests to indicate the presence of readily available carbonate minerals. 4.3.2 AGGLOMERATION PRODUCTION & CURING In 1985, Hraste observed that as the moisture content of the pellets increase, the period of growth by coalescence of primary agglomerates becomes more dominant and while fewer agglomerates form, their sizes increase. This also occurred in our studies, but the addition of too much water resulted in the material sticking to the disc surface while the addition of too little water resulted in finer pellet sizes and poor agglomeration. For these reasons, water addition was optimized to produce the best quality green balls. The effect that the water might have on long-term pellet strength was not considered during production. The dry-mixed tailing and cement was fed to the disc pelletizer using a vibrating feeder and water was added at a rate designed to produce a final green-ball pellet containing about 12% to 17% moisture, depending on the cement content and material type. The equipment was continuously monitored to regulate the development of the pellets and to avoid problems such as sticking and clumping. Each batch was processed to form agglomerates that would fall within a pellet size distribution ranging from 0.5 to 2 cm, with an average diameter of 1.0 to 1.5 cm. This range was considered appropriate to replace some of the river aggregate as well as fill the voids therein, thus forming a well-compacted, dense material when blended with unagglomerated tailings and/or river aggregate. Regardless of cement content, each batch of pellets required a similar amount of water (within a couple percent). Pellets containing 3, 4, 6, 8, and 10% cement were produced from the filtered tailings samples. Following discharge from the disc, all pellets were cured in a humidity chamber (with saturated humidity at ambient temperature) for the first 7 days, to reduce moisture loss from the pellets during the initial curing stage. After 7 days, approximately half of the pellets in each batch were removed from the chamber to cure at ambient laboratory conditions, while the remaining pellets were stored inside the humid chamber, for a period of 28 days. This approach was taken to track strength development of pellets cured inside and outside of a humid environment with time and 24 to determine whether the humid curing conditions were required for sufficient strength development after the initial 7 days. Quantitative strength analyses were performed using uniaxial compression test procedures and formulae for spherical objects, shown below in Figure 4.1, at various stages of curing (3, 7 and 28 days). A number of pellets, varying in size, from each batch were compressed at a constant rate until failure occurred, at which point the ultimate force was recorded. Sphere on a flat plate, P = Total Load I P Max ac = 0.918 3f P ' V D2[(1-Vi2)/E! + (l-v22)/E2 ] 2 D Max (Tt = 0.133 (Max <JC) ...tensile stress r Max Os = 1/3 (Max oc) ... shear stress Figure 4.1: Formulae for Maximum Stress on a Spherical Object Extracted from R.J. Roark, Formulas for Stress and Strain, 4th Ed., McGraw-Hill, 1965. For the conversion of ultimate force to a value of maximum stress, two values had to be assumed. These were the modulus of elasticity, E, and Poisson's ratio, v, of the agglomerated materials. Because these values have not been measured and / or published for an agglomerated tailing material, the values were extracted from studies based on the most similar material around - cemented paste fill. In one particular study, the elastic modulii of various paste fills were plotted against their respective cement contents (Andrieux & Brummer, 2000). Using this graph, a number of modulii were extracted for the calculation of the maximum stress, as shown above. The values that were used are listed in the following table: Table 4.1: Elastic Modulii vs. Cement Contents of various paste fills Cement Content Elastic Modulus (%) (GPa) 3 0.10 4 0.22 6 0.47 8 0.78 10 1.00 25 Results of the tests are presented in Appendix C and summarized in Section 4.4.2. 4.3.3 B A C K F I L L P R O D U C T I O N A N D C U R I N G Four sets of tests were carried out to study different properties of the backfill mixtures. For the first three sets of tests, only tailing materials from the 109 Zone were used. The final set of tests involved some of the same combinations already investigated, but were carried out using materials from the NEX Zone. These results are compared to those made from the 109 tailings. 4.3.3.1 Standard Procedures For all backfill testwork, particles above 3.8 cm (1.5") in diameter contained in the river-rock material were removed to minimize errors in the compressive strength testing. Generally all particles must be less than 1/3 of the diameter of the sample container to ensure reproducibility of test results. All materials were dry-mixed prior to adding water. Addition of water to each batch, with the exception of the third set of tests, was varied to produce a relatively constant slump of the unsupported material. The goal was to produce a backfill mix with little or no slump2 to simulate working conditions at the mine site. Batches made with 60% or greater content of agglomerated tailings would not stand up against even minor additions of water; in these cases, the loss of support was not from a slump action, but more like the dispersion of spilled marbles. The wet mixtures were placed into 10 cm (4") diameter by 20 cm (8") long cardboard tubes in three stages. To achieve quality, consistent compaction for comparable results, approximately 1/3 of the container was filled before the material was tamped or compacted into place by 'rodding' the materials numerous times, equally across the limits of the column. This was done a second and third time before the top was leveled off, and then the fill sample was placed in a humidity chamber for up to 28 days. After 7- and 28-day periods, the strengths of the backfill mixes were tested in uniaxial compression, as shown below in Figure 4.2. The fill columns were ground flat at each end to ensure that force was applied in a direction parallel to the axis of the column length. Force was 2 Slump of the mixed material was measured using ASTM standards for Hydraulic Cement Concrete. 26 applied at a constant rate until the sample failed in compression. The ultimate force at failure and the mode of failure (through clastic material or matrix) were observed and recorded. J Figure 4.2: Uniaxial Compression of a Cylinder 4.3.3.2 Preliminary Tests - Baseline Testwork The cement content chosen for the preliminary tests (9% by dry weight) was equivalent to that used at the mine in its river-rock aggregate backfill. The pellets were produced at water contents between 14% and 15% by weight, and were uniform in size gradation. The pellets were also cured in the humidity chamber for 7 days prior to use in the backfill. 4.3.3.3 Secondary Tests - Effect of Cement Content Initial tests were conducted using only river rock and cement; the level of cement was reduced incrementally from 9% to 3% by weight to determine the effect on aggregate backfill strength. The results of these tests provided a standard for strength comparison at various cement concentrations for when agglomerates are incorporated in the fill. In order to study the effect of reduced cement in backfill made with varying combinations of river rock and agglomerates, the cement content within the agglomerated tailings was held constant while the cement addition to the fill was reduced from 9% to 6% and then 5%. Further tests were carried out on similar mixtures where the cement content in the pellets was reduced while that in the total mix was held constant. 4.3.3.4 Tertiary Tests - Effect of Water-to-Cement Ratio For the third set of tests, a mixture of 60% rock and 40% pellets was chosen. This mix, when unagglomerated tailing materials were not present, packed the best in comparison to other rock-27 pellet mixtures. The cement content within the fill was held constant at 6% cement and the water-to-cement ratio was varied. 4.3.3.5 Final Tests - NEX Materials After investigating many possible combinations of materials using tailings produced from the 109 Zone, similar samples were prepared using NEX tailing materials. The strength and failure characteristics of the samples were investigated in relation to that of the 109 samples. 4.4 RESULTS In the following sections, the general findings of each of the testing stages are presented, including the material properties, and the geotechnical characteristics of the two types of agglomerates and various backfill mixtures produced during the program. 4.4.1 M A T E R I A L P R O P E R T I E S Studies were carried out on the river rock aggregate and on materials from two of the mine's mineralogically distinct zones, respectively designated as the 109 Zone and the Northeast Extension (NEX) Zone. 4.4.1.1 Grain Size Distribution and Moisture Data The river rock aggregate had a broad particle size distribution, ranging from silt-sized particles to cobbles greater than 3 inches in diameter. The particles were predominantly rounded and relatively clean, containing no apparent clay and very little silt. The moisture content of the aggregate was approximately 2.0 wt%, as received. Both of the tailings were of relatively uniform grade, being predominantly silt to clay-sized in nature. The 109 tailings tended to be a little coarser, containing a small amount of fine sand. The average moisture contents of the two tailings were 18.0 wt%, with a minimum of 16.5 wt% for one of the 109 tailings batches and a maximum of 18.5 wt% for the NEX tailings. 28 Figure 4.3 presents the average grain size distributions of the materials as measured by the techniques described in Section 4.3.1.1. See Appendix B for the results of the individual tests performed on each of the samples. 100.00 c 0 ) o d) O. > n 5 ° " E 3 o 90. 80. 70. O) 60 8 5 0 n 00 00 00 00 00 40.00 30.00 20.00 10.00 0.00 Figure 4.3: Grain Size Distribution of Materials - As Received —11 III 1 A Ml—Y i ('Average' Tailings M " I N t / —1—109 'Average' Tailings River Aggregate 1000000 100000 10000 1000 100 10 Particle Size (microns) 0.1 Note that, overall, the NEX Zone tailings are finer grained than the 109 Zone tailings (83% vs. 74% minus 200-mesh or 75 microns). They contain more clay-sized particles (9% vs. 3% minus 2 microns), while the 109 tailings have an overall more uniform size distribution. Surprisingly, the NEX tailing size distribution shows a larger maximum particle size than the 109-tailing sample. This unexpected phenomenon is due to the presence of pebble-sized rock fragments observed in the NEX tailings during sample preparation, most likely through sample contamination at the mine site. These relatively large, 'out-of-place' particles were removed from the tailing material before being used in agglomeration and backfill testwork. 4.4.1.2 Mineralogical and Chemical Data The predominant minerals and minor constituents found in the two tailing samples are summarized in Tables 4.2 & 4.3 below. For each grouping, the method of determination or source of information is indicated. 29 Table 4.2: Predominant minerals within tailings Constituent 109 Tailings NEX Tailings *Source Minerals: Quartz, K-Muscovite, Sphalerite, Pyrite Quartz, K-Muscovite, Nimite (Ni-Chlorite), Clinochlore (Fe-Chlorite), Pyrite, Stannite XRD Galena Chalcopyrite R M & S E M * XRD - X-Ray Diffraction, RM - Reflection Microscope, SEM - Scanning Electron Microscope XRD analyses conducted on the tailing samples indicate that the sulphide minerals present in the 109 samples include pyrite and sphalerite. Minor constituents containing manganese and antimony were also suggested by the analyzing software system, indicating their presence, and later confirmed by the assay results, as shown in Table 4.3. RM and SEM analyses indicate that pyrite is the dominant sulphide mineral and galena is a minor constituent associated with both pyrite and sphalerite. In the NEX sample, the only sulphide mineral indicated by XRD is pyrite. Chalcopyrite, associated with pyrite, was detected using RM and SEM analyses, but not through XRD analyses due to its low concentration in the sample. Other rock-forming minerals noted to be present in the samples include quartz and muscovite, and in the NEX sample, nimite and clinochlore. Muscovite, as well as nimite and clinochlore, two forms of chlorite, are all known to break down into various forms of clay. Finally, small amounts of carbonate in both samples were detected through fizz tests performed during ARD testing (however, this was unconfirmed by XRD analysis). As shown in Table 4.3, assay testing and results of shake flask tests (Dai, 1999) indicate the presence of metals within the tailings samples. NEX tailings contained a higher concentration of heavy metals, with the exception of zinc whose concentration was greater in the 109 tailings. Of most significance are the presence of zinc, arsenic, antimony, manganese, lead, and mercury, as they are the most toxic of the metals present. 30 Table 4.3: Heavy Metals & Other Constituents within tailings (Dai, 1999) Constituent Method Units 109 Tailings NEX Tailings (full sample) (full sample) Aluminum ICP % 3.29 7.02 Antimony - ppm 61 550 Arsenic - ppm 276 277 Barium ICP ppm 120 380 Beryllium ICP ppm 0.5 1 Cadmium ICP ppm 82 19.5 Chromium ICP ppm 187 31 Cobalt ICP ppm 1 5 Copper ICP ppm 217 535 Iron ICP % 3.2 2.25 Lead AAS ppm 1200 3660 Manganese ICP ppm 425 645 Mercury - ppb 4960 8420 Nickel ICP ppm 10 52 Sulphide (S) * % 3.77 1.79 Sulphate (S) ** % 0.07 0.11 Zinc ICP % 2.07 0.46 * Hydrochloric Acid Soluble Sulphate ** Nitric Acid Soluble Sulphate The results of the X-ray Diffraction (XRD) analyses and assay testing help explain the differing grain size distributions of the two types of materials, although the ores were extracted from their parent materials using the same mining methods. The XRD scans indicate that the NEX tailing sample contained various clay minerals not present in the 109 tailing materials, which would lead to a finer particle size range due to the friable nature of these soft minerals. Visual observations combined with some assay results also indicate a higher amount of quartz and sulphides (typically harder mineral types) present in the 109-tailing sample. The hard minerals resist the comminuting processes carried out in the mill. The combination of the above explains the coarser particle sizes of the tailings from the 109 Zone. 4.4.2 A G G L O M E R A T I O N T E S T RESULTS 4.4.2.1 General Findings In general, the results show that pellet strength rapidly increased in the first seven to ten days of curing, then only slightly thereafter (see graphs in Appendix C). Strength was not measured 31 beyond 28-days, but it is believed that >90% of the ultimate pellet strength was achieved by this time. Two sets of tests were devised for each batch of pellets. The first set was cured in a humid environment for a 28-day curing period. The second set was cured in a humid environment only for the first 7 days of curing, then removed from the humidity chamber and cured in the laboratory at ambient temperatures. The strengths of the two sets of pellets were compared and showed that the humid environment was indeed required for the initial curing stage, but was unnecessary for the final curing stages (see Table 4.4). The pellets cured outside the humidity chamber for the final curing stages (designated with a 'B' in the appended graphs) were not weaker than those cured inside the chamber for the entire 28 days. In fact, in most cases, these pellets were stronger, possibly because they had been allowed to dry out after curing and prior to testing. Pellets that were cured in a humid environment for the entire 28 days were tested immediately after removal from the chamber and may have contained interstitial pore-water that had not been consumed in cement hydration reactions. It is suggested that by not allowing the pellets to dry out prior to testing, internal pore-water remained within the pellets, potentially lowering their strengths in comparison to the pellets that were cured outside the humidity chamber. 4.4.2.2 Comparison of Tailing Types After a short curing period of 3 days, pellets made with the NEX tailings showed higher strength than those made with 109 tailings, except for those containing 10% cement which were lower in strength. After 7 days however, the allocation of relative pellet strength by cement content shifted. The strength of the NEX pellets at low cement (4 - 6 wt %) was greater than the 109 pellets. However, at higher cement contents (8 - 10 wt %), the 109 pellets developed higher strength than did the NEX pellets. The normalized pellet strength as a function of cement content after 7 days curing is shown in Table 4.4. 32 Table 4.4: Pellet Strengths at 7 & 28 Days Cement Content (%) 109 Average and Range Pellet Strength (MPa) NEX Average and Range Pellet Strength (MPa) 7 Day 28 Day 28 Day * 7 Day 28 Day 28 Day * 3 5.96 ± 0.98 Not tested 6.99 ± 0.23 Not tested Not tested Not tested 4 12.02 ± 1.06 13.58 ± 1.28 13.82 ± 1.03 14.85 ±0.90 Not tested 15.98 ± 1.51 6 23.19 ±2.79 24.59 ± 3.42 25.80 ± 3.66 24.54 ±2.17 26.52 ± 1.95 Not tested 8 46.02 ± 4.99 49.29 ± 8.45 56.95 ±7.51 43.33 ± 2.99 46.56 ± 3.28 55.37 ±4.83 10 53.99 ± 6.05 58.48 ±5.80 Not tested 52.80 ± 3.80 56.49 ± 5.86 68.15 ±4.00 Pellets were cured outside of the humidifier from day 7 to day 28. All agglomeration test results and graphs can be found in Appendix C. 4.4.3 B A C K F I L L T E S T RESULTS 4.4.3.1 Baseline Studies Preliminary tests were designed to note the effect of adding agglomerated tailings and loose tailings to the mine's rockfill; in essence, replacing some of the river rock aggregate with mine tailing waste. The agglomerates were cured for 7 days in the humidifier, for reasons stated in the previous section, and contained 8% cement by weight. At 8% cement, the pellets were of very good quality and reached strengths of close to 50 MPa, similar to that of some medium strength rock types such as sandstone, schist and shale (Hoek & Brown, 1982). In order to compare the strengths of these fills, the content of cement in the fill was held constant at 9 wt%, similar to that used at the mine. While all backfill test results and graphs may be found in Appendix D, the following was noted: 1. Strength decreased as the amount of rock in the backfill mixture decreased, until there was approximately 50% rock in the mixture. Below this point, the strength varied depending on the combination of materials used in the backfill. 2. Mixtures of river rock and agglomerates without unagglomerated tailings required a minimum of 60 wt% rocks to achieve acceptable strength levels. 33 3. Mixtures of river rock and agglomerates combined with ~ 20 wt% unagglomerated tailings showed good results, especially when the river rock was above 30 wt%. 4. In batches containing unagglomerated tailings, the strength of the fill decreased significantly when the total tailing content, including agglomerates, increased above 50 wt%, in comparison to the strength of the 100% river rock fill. 5. Mixtures containing a large proportion of tailings (60 wt%), agglomerated or not, either gained little strength, remained the same, or lost strength from 7 to 28 days. Relative strengths of the mixtures are represented in Figure 4.4. In the graph, the percent of unagglomerated tailings in the mixture represents the percent by weight of the total mixture. The remaining percentage of the mixture is made up of the listed ratios of rock and pellets. For example, a mixture of 20 wt% tailings contains an 80 wt% mixture of rock and pellets. If the ratio is 80:20 rock to pellets, then there is 64 wt% rock and 16 wt% pellets in the total mixture. Figure 4.4: 28 Day Backfill Strength vs. % 109 Tailings 9% Cement in Matrix 16 -, 4.4.3.2 Modes of Failure Each of the samples was cured in a 10 cm (4") diameter by 20 cm (8") long cylinder. On uniaxial compression of the samples, three dominant failure mechanisms were apparent: 34 1. Failure of the aggregate-cement bond, and subsequent outward buckling of the cylinder sides due to the rearrangement of aggregate particles. 2. Crumbling of the sample materials, or collapsing of the fill structure, due to insufficient material strength or high void content (in mixtures containing high volumes of agglomerates). 3. Failure (somewhat planar) through a matrix of fine materials, including agglomerates. For some mixtures, more than one failure mechanism occurred. In samples containing large volumes of rock, small amounts of pellets and up to 40 percent tailings, failure was predominantly through the aggregate-cement bond, followed by outward buckling of the column sides. As the fines content increased, samples also failed by collapse of the weaker matrix. The failure surface, however, remained very irregular due to the presence of stronger, larger particles. This irregular surface became gradually more planar as the river rock content of the sample was decreased. To better describe these failure mechanisms, pictures of some samples after failure have been attached at the end of the thesis (p. 185). In cases where the pellet content was predominant and no tailings were present, the porous structure would collapse on application of force. It was also observed that when the pellet content was above 40%, some pellets failed (fractured) during application of force. This occurred equally when either rock or tailings made up the balancing portion of the mixture. However, when the corresponding material was river rock, the pellets predominantly crushed prior to total sample failure. This allowed the harder aggregate particles to shift in the same way as if voids were present. When the balancing material was unagglomerated tailings, a planar failure surface developed through the fine matrix and pellets, as if there was no change in structure or density whatsoever between the agglomerates and the unagglomerated tailings. See attached photos (p. 185). 4.4.3.3 Proportion of Cement in Fill & Pellets The secondary tests were designed to investigate the effect of reducing the cement content in the agglomerates and the backfill. The tests began with the production of fill using river rock only (no pellets or tailings) while the cement content was reduced. These results are shown in Figure 4.5. 35 Figure 4.5: 28 Day Strength vs. Cement Content 100% Aggregate Backfill Cement Content (%) As can be seen, normal mine backfill at 9% cement provides strength of over 14 MPa. With decreasing cement content, this strength decreases linearly to 0 MPa at 2.6 wt % cement. This result is important in defining the effect of cement variation on backfill strength and to establish a baseline against which to compare strengths of fill containing tailing materials. A cemented rock fill (CRF) containing only 4.5% cement provides the acceptable minimum fill strength of 4 MPa for underhand mining sequences. Further testing was carried out on two sets of fills, one containing agglomerated tailings produced at 8% cement and the other, agglomerated tailings produced at 6% cement. Some of the test results are as follows: 1. There is significant decrease in fill strength with decreasing cement content. Satisfactory strengths (4.0 MPa) are achieved by backfills containing 6% cement after 28 days curing. 2. Fill strength remains acceptable when 6% cemented rock fill (CRF) has 40% of its river rock component replaced by pellets containing either 6% or 8% cement. 3. When the cement content of the pellets is reduced from 8% to 6%, there is a small decrease in the backfill strength, i.e., a 60% river rock, 40% pellet (60R:40P) mixture with 6% cement in the fill, changes in strength from 5.0 to 4.0 MPa when the cement content in the pellets is reduced from 8% to 6%, respectively. 36 4.4.3.4 Ratio of Water to Cement A third set of tests was carried out to determine the effect of changing the water to cement ratio between similar batches of fill, while maintaining a low slump. The cement in the fill was held constant at 6 percent, similar to that of the pellets. For this type of fill (at 6% cement), an optimal water to cement ratio of 1.2 to 1.3 was discovered (see Figure 4.6). Figure 4.6: 28 Day Strength vs. WaterrCement Ratio for a 60R:40P Mixture W 2 • 6 % C , 6 % C i n pellets • 6% C , 8% C in pellets X 6 % C , 100% R X Hi • • 0.8 0.9 1 1.1 1.2 1.3 1.4 1.5 1.6 WatenCement Ratio On investigating the optimal w/c ratio for other mixtures, it was found that this value is inversely related to the amount of cement used in the fill. As the amount of cement increases, the optimal ratio of water to cement is reduced (for 9% cement, the optimal water-to-cement ratio is ~ 0.85). 4.4.3.5 NEX Materials Only a few tests were performed on fills made with the NEX Zone tailings as all tests discussed above were performed using 109 tailing samples. After general trends were established, fills were produced with NEX material and their strengths compared to the samples made with 109 materials. The general finding was that backfill made with NEX agglomerated tailings is slightly lower in strength (on average, 0.5 to 1.0 MPa lower) than similar fill made with 109 agglomerated tailings. 37 5 D I S C U S S I O N On examining the data, there is a notable amount of scatter. A number of factors have an effect on the cemented material strengths - on the small scale for the pellets, and on the large scale within the fills. However, on comparison of these test results with those of the tests carried out by other consultants, the precision was good. In the following sections, the results of the test program are discussed in detail with respect to these strength-affecting variables. 5.1 C O N T R O L S & E F F E C T S O N A G G L O M E R A T E S T R E N G T H 5.1.1 MECHANICAL CONTROLS OF PRODUCTION During agglomerate production, it was noted that the quality, size and shape of the pellets depended largely on production controls. This included the rates at which water and dry feed were added, the location of the wet and dry feeds in relation to the disc and to each other, as well as the speed of the rotating disc. In our test work, if the water dispersal nozzle was located near the pellet discharge area, agglomerates would stick together, forming irregular, unstable shapes whereas spraying the water near the dry-feed drop point (mid-flow of path 3, Figure 3.1, p. 17) allowed agglomerates to grow gradually during tumbling. However, even with the "correct" positioning of the water nozzle, there were still cases where pellets of somewhat irregular shape would develop. The effect of shape on pellet strength is discussed in Section 5.1.4.2. The speed of the rotating disc, and the position and rates of the dry feed also affected the physical characteristics of the pellets. The slower the disc turned, the larger and wetter the pellets became. The closer the dry feed was to the discharge area, the smaller the resulting pellets. If the water addition rate was not increased with an accelerated rate of dry feed addition, the growth of the pellets typically ceased because the pellet surfaces became coated with dry material, arresting agglomeration mechanisms by hydrostatic bonding. 5.1.2 MATERIAL CHARACTERISTICS Within a margin of error, there was not much difference between the strengths of the two materials. As mentioned in Section 4.4.2.2, only two patterns emerged. At low cement content (3, 4 & 6%) and during the early curing stages (3 days), the NEX tailing agglomerates retained 38 greater strength than the 109 tailing agglomerates. At higher cement contents (8 & 10%), the opposite was true. There are a couple of ideas that may explain this, but both relate to the relative strengths and types of parent materials the tailings were derived from. First, the 109 materials were extracted from a network of stockwork veins predominantly consisting of silicified rock, quartz and sulphide minerals. The NEX tailings were derived from a unit containing high proportions of heavily altered, friable minerals and clays that are naturally softer than quartz and pyrite. 5.1.2.1 Mineralogy & Sorption Effects While the presence of clay in an aggregate mix may cause its structure to become porous due to the shape of the particles (explained in Section 5.1.3.2), clay minerals have strong adsorption properties and tend to agglomerate well due to hydrostatic forces (van der Waals' forces) acting between the clay particles and water, even when the particles are unaligned. After a short time, the 109 pellets (containing less clay) became stronger than the NEX pellets produced at the same cement content (8 and 10 wt%). One possible cause of this shift could be due to the ongoing consumption of 'bonding' water within the pellet structure through the hydration reactions of the cement, and its subsequent susceptibility to collapse on the application of pressure. This is supported by the fact that the NEX agglomerates at lower cement content continued to have greater strength than the 109 pellets after 7 days of curing. Water within these pellets would also have been consumed, but to a lesser extent (less cement consumes less water) than in those pellets containing more cement. The presence of the clay minerals and the hydrostatic forces acting within the pellets perhaps contributed to the total strength of the pellet, resulting in their greater strength at low cement contents during the early curing stage. In the same respect, the pellets containing more cement would have consumed a greater amount of the water, reducing the hydrostatic forces acting within the pellets and therefore reducing their strength. 5.1.3 BULKING & COMPACTABILITY 5.1.3.1 Bulking An effect called bulking commonly occurs when moisture is introduced to a mass of "fine" particles. Using sand as an example, Neville (1995) describes this phenomenon as an "increase 39 in the volume of a given mass of sand caused by the films of water pushing the particles apart...the bulking results in a smaller mass of sand occupying a fixed volume." He goes on to say "the extent of bulking depends on the percentage of moisture present...and on its fineness. Finer sand bulks considerably more and reaches maximum bulking at a higher water content than does coarse sand...extremely fine sand has been known to bulk as much as 40 percent at a moisture content of 10 percent." Finally, he mentions that "crushed fine aggregate bulks even more than natural sand [due to its angularity]." In light of this last statement, we can expect that tailing materials, being finer than the finest sand-size fraction (mostly silt and clay), would bulk significantly. When crushed aggregate is used in concrete, the angular particles do not compact well, in comparison to when fluvially derived or sub-rounded to rounded materials are used (Neville, 1995). The comparability and density of the aggregate mix increases slightly if there are sufficient fines (sand-sized particles) available to fill the voids between the larger particles. However, the addition of too much fine material causes the structure of the larger aggregate materials to shift and bulk slightly, resulting in an increase in the void content and in turn decreasing the density. This shows that the overall void ratio of an aggregate mix may remain relatively high, even with overly sufficient compaction efforts, for two reasons. 1. If there are too many fines, bulking occurs due to an alteration in the structure of the larger aggregate particles to accommodate the excess fines (even in the case where the "coarse" fraction is made up of sand-sized particles). 2. Where the aggregate materials (including 'void-filling' fines) are angular, the potential compactability of the mix is reduced because materials will not shift and settle into a dense formation; voids always exist. The result of these bulking effects is the production of a porous structure, typically reducing the structure's strength. 5.1.3.2 Compactability The degree to which an aggregate (combination of materials) may compact is affected by: the relative amounts of the various particle forms (angular, sub-rounded, elongate clay minerals...), their relative distributions within the mixture, and how well the particles 'fit' together when 40 mixed (see Figure 5.1 below). Of course, when cement is included as a portion of this aggregate mix, and water is added, its compactability is affected even further. Figure 5.1: Packing mechanisms of random particles with clay. Because of the stronger quality of the 109 materials, it could be expected that the particles, or a large proportion of the particles, would remain angular in shape throughout the grinding processes of milling. Therefore, the potential compaction of these mixtures would depend on the degree of angularity of the materials, and the bulking effects that this causes. Because the effects of compaction between the two material types were not compared, it is unknown to what extent these factors affected the relative strengths of the pellets or if these factors affected the pellet strength at all. However, the above theory does provide another reasonable explanation for the phenomena described in Section 4.4.2.2. At low cement contents, the early (3-day) NEX strength can be attributed to its clay content, which would provide hydrostatic bonding, and the better compaction of the mixed materials (than the 109 mixtures) due to the fuller gradation of these tailings in the fines range. At higher cement concentrations, it is possible that the cement's addition to the NEX materials caused the gradation of the mix to become too uniform, and the structure of the larger particles to shift and bulk, forming a weaker structure as described above. In the same instance, this addition of cement to the 109 tailing materials might not cause such bulking due to the greater particle sizes (than NEX materials) initially present in the sample - the fines limit would initially be less than that of the NEX materials, and the addition of fine cement particles would likely not upset the balance of coarse and fines to the bulking limit until a much greater volume of cement was added. Instead, the addition of the cement (up to 10%) to the 109 tailings would only act to improve the gradation and decrease the voids content of the mix, increasing the mixture's compactability and the resulting structure's strength. 41 5.1.4 SIZE & SHAPE OF PELLETS 5.1.4.1 Pellet Size In all cases, the force required to break a pellet increased with both the cement content as well as the mean size of the pellet. While this trend has been commonly observed in other studies of cold-bond agglomeration (L.M. Amaratunga, 1991, 1995 and 1997), conversion of the failure force to a normalized strength (N/m2) showed that, at constant cement content, strength is relatively independent of pellet size, with the exception of a few cases where strength decreased with increasing pellet diameter (see Appendix C). The fact that there was little change in strength with respect to the size of the pellet shows that the materials were well blended prior to agglomeration and the distribution of cement throughout the sample was relatively uniform. To explain why strength may have decreased with increased size, M. Hraste and Z. Nuber (1985) discovered that "in the case of monodisperse feed material there is a maximum upper limit of particles that can be agglomerated." They also found that the finer particles agglomerated first, while the larger particles tended to bond to the surface of the already formed agglomerates towards the end of the process. So it would be expected that the outer structure of said agglomerates would be looser since the particles would not fully compact due to their relative particle sizes, and there would be a lack of fines to sufficiently fill the voids. This phenomenon could be more pronounced in larger pellets, which take a longer time to form, and acquire more of the latent particles as they grow. The effect of the presence of voids within the pellet structure is typically detrimental to strength. 5.1.4.2 Pellet Shape When force is applied to an object, the stresses propagate through the material according to its shape. Stresses become evenly distributed throughout an object, given that its density remains constant throughout, when its shape is uniform or equidimensional along the axis of the application of force (see Figure 5.2 below). 42 Figure 5.2 Stress Distributions in Equidimensional Objects For this reason, one would expect that the degree of sphericity would be a contributing factor to a pellet's potential strength. However when the aspect ratio (ratio of the pellet's largest axis dimension to its smallest axis dimension) was examined, it seemed to have little or no effect on the strength of the pellet. This is shown in figures in Appendix C. 5.1.5 WATER TO CEMENT RATIO At this time, we may recall that all batches of agglomerates were produced at similar water contents. Therefore, if the cement volume in the dry-mixture increases and the water requirements remain the same, the water to cement ratio decreases. Below, in Figure 5.3, the strengths and water to cement ratios of the agglomerate batches are plotted against each other for both the 109 and NEX materials. Figure 5.3: Strength vs. Water-Cement Ratio 28 Days, Both Tailings 8 .00E+07 -i m 109 Humid Curing 7 .00E+07 A N E X Humid Curing (N 6 . 0 0 E + 0 7 2? 5 . 00E+07 £ 4 . 0 0 E + 0 7 U) c 3 . 00E+07 5J 2 . 0 0 E + 0 7 1 .00E+07 0 .00E+00 4— 1.00 1.50 2 .00 2 . 5 0 3 .00 3 .50 4 .00 4 . 5 0 Water to Cement Ratio 43 Overall, strength notably increased as the water to cement ratio decreased - this was true for agglomerates made with both tailing types. Less noticeable was the decrease in strength of the 109 agglomerates when the water to cement ratio decreased below 1.75 (see dashed trend line). This did not occur for the NEX tailings. It is suggested that this downtrend at low water to cement ratios is not representative of the material's potential strength. It is more likely that the strength of the pellets would continue to increase in strength with decreasing w/c ratios3 for the pellets that are smaller in size. Remember that as agglomerates grow in size, the growth becomes increasingly due to the agglomeration of smaller pellets. Remember also that at lower water contents, pellets do not typically grow very large. Therefore, any pellets at low water contents (low w/c ratios) that would grow to a diameter large enough to test by compression would most likely have grown in the manner just described. Such growth mechanisms would lead to apparently weaker pellet structures due to the increasing presence of voids, decreasing contact surface areas and poor bonding between these surfaces. An explanation as to why this did not occur for the NEX pellets would simply be due to the presence of more clay minerals than in the 109 materials. These minerals would have helped to improve the bonds between particles and smaller agglomerates, therefore maintaining a semblance of the strength apparent in the smaller agglomerates. 5.2 CONTROLS & EFFECTS ON BACKFILL STRENGTH Extensive testing has been carried out over the past few decades on various types of fill, including rockfill, hydraulic sandfill, cemented rock and sand fills, and paste fills. While researching data from such test programs, it became very apparent that the overall voids content and degree of compaction of a cemented aggregate material are its most important properties. However, there are three main elements, all closely related, that contribute to void content and the resulting fill strength. These factors are sample density, water to binder ratio and grain size distribution. This trend could only be expected to continue until a minimum water to cement ratio of 0.36, although this water content could be insufficient for agglomeration of the dry materials. This value is the minimum water requirement for the chemical reactions of cement hydration (Mindess, 2001). Below this value, there would no longer be sufficient water present for cement hydration, let alone agglomeration. 44 Additionally, there are a number of factors that contribute to the development of these strength-controlling elements. Based on early studies on various fill types, with the exception of paste fill, Dickhout (1973) and Thomas (1979) summarized these factors. Overall, most tests showed similar results, in that the most important contributing properties are: • solids relative density, • particle size gradation and degree of fineness of the components, • moisture content of the fill, • consolidation characteristics of the fill, • method of placement, and • binder type, its properties, and the amount used. In studies on paste fill operations, more emphasis has been placed on the pumpability characteristics and related properties such as grain size distribution, moisture content and binder type and content (if any). Most paste fills, until recently, were used to fill stopes rather than to provide structural support. Recent paste fill studies have investigated the addition of strength adding river aggregate, crushed mine rock waste or agglomerated tailings while maintaining pumpability. In our studies, the aim was to find the best quality fill, one having the ability to provide structural stability within the mine. 5.2.1 FILL DENSITY Fill density has an apparent effect on fill strength in that the pattern found during analysis indicated that an increase in density generally signified a greater strength of the fill, as shown in Figure 5.4 below. However, in many cases, an increase in density did not result in an increase in strength. The reasoning for this may be due to several factors. There are many fill properties contributing to the density of a fill, including the bulk densities of each of the material components, their individual grain size distributions and particle shapes and their relative amounts in the total mix. The ways in which these properties affect the density and strength of the fill are discussed in the following sections. 45 16.000 - i 14.000 • 12.000 • "ra Q. E 10.000 • . c 4-* 8.000 Ol c £ 6.000 5 5 4.000 -2.000 -0.000 -Figure 5.4: 28 Day Backfill Strength vs. Density All mixes - including Variable Cement § § 8 8 X < o p . f v _ \& t i — -O100R:0P O 80 R: 20 P X60 R:40P X40 R: 60 P A20 R: 80 P • OR: 100 P - Variable Cement Density (kg/m3) Note that for each ratio of rock to pellet (as shown in legend), the density changes are due to variability in tailings content. Also, all of the samples tested for the variable cement results contain no tailings. 5.2.1.1 Component Bulk Densities & Grain Size Distributions These factors played a major role in the resultant density of the fill as the three main components each had different dry bulk densities, grain size distributions and particle shapes. On average, the specific gravities (SG) and dry bulk densities (pbuik) of the materials were as follows: • Tailings: SG 2.9, pbuik =1.1 kg/L; • Pellets: SG 2.9, Pbuik(singie pellet) = 2.1 kg/L, Pbuik(peiiets) = 1-2 kg/L; and, • River rock: SG 2.6, pbuik = 1.9 kg/L. The particle shapes of the river rock ranged from rounded and subrounded to oblong with smooth corners, while those of the agglomerates were predominantly spherical with rounded bumpy surfaces. The tailings materials, as described earlier, predominantly consisted of small angular particles due to the crushing processes of their formation. The grain sizes of the river rock were well graded, distributed across a range of sizes spanning over two orders of magnitude while both the pellets and tailings materials were uniform in grain size distribution, although at differing median sizes (see Appendix B). 46 Theoretically, a full gradation of particle sizes having somewhat subrounded, non-equidimensional shape allows the settlement of particles into a more dense structure; finer particles fill void spaces developed by a coarser aggregate structure, and in doing so, reduce the void content of the mix. In opposition to this, when a uniform gradation of particle sizes having fairly spherical shape are placed in a container, a large proportion of voids is formed within the contained volume. This effect can be likened to the spaces left between marbles that are set in a jar. The dry bulk density of the river aggregate was greater than that of the agglomerated tailings, and even greater than that of the unagglomerated tailings. Both the tailings and the pellets have low dry bulk densities in relation due to the uniformity of their grain size distributions and fairly spherical shapes, because both trap significant void fractions within their structure. 5.2.1.2 Relative Volumes & Resulting Mixture Grain Size Distributions In general, adding rock to each mix has the effect of adding a greater proportion of weight per cubic meter than by adding tailings or pellets. The effect is an increase in fill density. Cleary, as the relative volume of rock declined, so did the sample density (see Figure 5.5). 2400 2300 2200 „ 2100 | 2000 >^ 1900 r 1800 w 1700 S 1600 ° 1500 1400 1300 1200 Figure 5.5 28 Day Backfill Density vs. Tailings Content All mixes - including Variable Cement 20 40 60 80 Tailings Content in Total Percent of Mix (%) • 100 R: 0 P O 80 R: 20 P X60 R: 40 P X40 R: 60 P A20 R: 80 P • 0R:100P 100 Also shown in this graph is an increase in density (to most fill mixtures) with the addition of tailings. For the fills predominantly containing pellets, the addition of tailings had the effect of 4 7 increasing the density significantly by filling large volumes of void spaces. With the continued addition of tailings, a drop in density typically followed this phenomenon. When two aggregates of differing sizes and densities are mixed and compacted, the resultant bulk density will typically range between the two primary densities or will attain a density greater than the higher primary density of the two. This can occur because a finer aggregate will settle into the voids of the coarser aggregate and so, the greater the volume of fines, the more voids filled and the denser the mixture becomes. However, there is a balance point at which the greatest density is achieved. At this point, the majority of the void spaces within the coarse aggregate structure become filled with the finer aggregate particles. If the volume of the fines exceeds the volume of the void spaces within the coarse aggregate structure, this structure becomes disturbed; the larger particles shift and 'bulk' to make room for the smaller particles. This shifting produces more voids than originally present, thereby decreasing the mix density. Continued addition of fines beyond this point eventually causes the mixture to become fines dominant and the density of the mix will approach that of the fine aggregate density prior to mixing. This was the case for those mixtures containing a high proportion of tailings. Because the density of the tailings was lowest, the mix densities dropped with each increased proportion of tailings to a point where the density of the tailings alone was reached. Note in Figure 5.5, that exceptions to this 'rule' occurred where the rock proportions were significant. In cases where the aggregate proportion of the fill was 100% river rock or a ratio of 80% rock and 20% pellets, fill density decreased with any addition of tailing materials. This is because the full gradation of the two relatively coarse samples prior to addition of fines was such that particles would settle into a dense structure containing relatively minimal void space. "When coarse aggregate is dominant, an actual mixture has more voids than that corresponding to the assumption that the voids within the coarse [aggregate structure] are filled with fines. Usually even a small amount of fine aggregate causes dispersion of coarse particles; voids within the coarse structure are usually too small to contain all sizes of a fine aggregate" (Powers, 1968). In these samples, the addition of tailings only worked to disrupt the equilibrium of the particles within the structure by making them shift to accommodate more volume than the void spaces supplied. 48 All of this is closely related to the resulting grain size distributions of the mixes. This is supported by the fact that there is a notable decrease in density when the pellet proportion in a mixture was altered; the pellets containing 8% cement were replaced with pellets containing 6% cement which had a finer, more uniform grain size distribution (smaller median size - see Appendix C). This affected the resulting gradation of the total fill, making its gradation more uniform, which in turn increased the void content. 5.2.1.3 Relating Density to Fill Strength As might be expected, the increases and decreases in density, described above and shown in Figure 5.5, are clearly reflected in the strength results (shown in Figure 4.4, p. 34). The strength of mixtures predominantly containing rock (100R:0P and 80R:20P) decreased just as the densities of the mixtures decreased in Figure 5.5. In these samples, the disruption of the aggregate structure by the addition of tailings seemingly caused the strength of the fill to decrease. Because the tailings materials were not tested by themselves at the 28-day stage, it is assumed that the strength realized when the tailings proportion became predominant was very close to that of the tailings on their own. This is supported by the consistency in strength that other mixtures developed when the tailings proportion became dominant. There was a significant decrease in strength with a decrease in rock content. This decrease was evident in any mix, whether it was due to its replacement by tailings, agglomerates, or both. Earlier, this was attributed to the self-weight of the rock and its affect on the resulting density. However, a second factor is possible. In conjunction with the reasons discussed above, this decrease in fill strength should be expected due to the relative strength of the tailings and pellets in comparison to the river rock at the time of mixing. "The compressive strength of [a fill] cannot significantly exceed that of the major part of the aggregate contained therein" (Neville, 1995). Earlier it was mentioned that fill density is only an indicator of strength and therefore should not be used to project potential strengths of undeveloped fills. This statement is verified by those cases where a greater density did not produce greater strength. Take the two mixtures: 60% rock and 40% tailings; and the second, 36% rock, 24% pellets and 40% tailings. Although the first mixture has a higher density than the second, its strength is lower. According to the above paragraphs, the expected strength of the first fill should be greater than that of the second 49 mixture simply due to the greater content of rock - both a stronger material (individual particles) and a more graded aggregate than the pellets. Although the reasoning for this phenomenon is not fully understood, it is suggested that the mixtures both have a fines content close to where maximum bulking occurs for the rock/tailing mixture before the fines become dominant with bulking diminishing to reduce voids. 5.2.1.4 Relating Density to Degree of Compaction In light of the above information, the strength and density data were analyzed in more detail, and some previously unobserved trends became apparent (see Figure 5.6). 16.000 -i 14.000 -12.000 -E 10.000 -8.000 -O) c p 6.000 -CO 4.000 -2.000 -0.000 -Figure 5.6: 28 Day Backfill Strength vs. Density All mixes - excluding Variable Cement • 100 R: 0P O80 R:20 P X60 R:40P X40 R: 60 P A20 R: 80 P • OR: 100 P O O O O O O O O O O O O O O o o o o o o o o o o o o o o <V oo ^ « iB /V co cT 5 N B * <o Density (kg/m3) The trend lines show that when each type of fill is analyzed separately, the relationships between strength and density for each are the same with the exception of the 60% rock and 40% pellet aggregate mixtures. For each group, densities and strengths begin low, followed by a gradual increase in strength with density, until the strength suddenly and rapidly multiplies with a very small increase in density. This phenomenon is attributed to the degree of compaction that each mixture has the ability to attain - its maximum potential degree of compaction. It is believed that the point where the sudden increase in strength occurs is also the point where the maximum degree of compaction was achievable for that mixture using method the compaction applied. In disagreement with these trends is the pattern depicted by the mixtures of 60% rock and 40% pellets. First, the rapid increase in strength is lacking for these batches, indicating that the mix 50 may not have the ability to compact to a greater degree possibly due to its grain size distribution, i.e. the mixture is at its maximum potential level of compaction. Second, there is a much higher 'low' strength zone in comparison to others mixes. All of the other samples have similar low strength values prior to the addition of fines and increase in density. This is a positive finding, confinning the assumption made in designing the secondary and tertiary testwork (based on observations). With little or no unagglomerated tailing material added to the fill, the mixture packed well and possibly better than other backfill mixtures at this compaction effort. 5.2.2 WATER T O CEMENT RATIO As may be seen in Figure 5.6, there was a notable degree of variation in the fill strengths, attributed so far to fill density, degree of compaction and the material properties affecting these. This graph represents only those data from the samples containing 9% cement. However, further tests were carried out on samples where the cement and water contents were varied in the fill mixtures and agglomerates. When the results from these tests were combined with the existing data, there was even further variation. In an attempt to further quantify this variance, the strengths of all of the samples were plotted against their respective calculated water to cement (w/c) ratios (Figure 5.7). A trend commonly observed in the concrete industry developed, whereby strength generally increased as the w/c ratio decreased. 20.0 Insufficiently compacted materials Figure 5.7 Strength vs. Water-Cement Ratio 28 Days 2 3 Water-Cement Ratio 51 In Figure 5.7, a trend line typically observed in the concrete industry4 is transcribed over the data measured in our test program (all notes on the graph pertain to this). Two significant points are the correlation between our data and this trend line, and the fact that all of the backfill mixtures in this program were prepared using hand compaction. It is commonly believed that the delineated peak in the graph represents strengths at the greatest degree of packing of the mixed constituents, including the water added to the mix. Since the samples were mixed and compacted without using special equipment, one would expect all data to fall in line with either the "hand compacted" or "insufficiently compacted" trend lines observed in the concrete industry. This is possibly the case, in that the increase in strength shown by this data could correspond to the peak of the hand compaction curve for cemented aggregate mixtures produced at much greater strengths (as is the case for concrete). However, it appears that from the pattern observed in our data and for higher water/cement ratios our data does correspond to that found in concrete industry with the exception that the strengths are much lower. 5.2.2.1 Relating Water / Cement Ratio to Fill Strength Studies have shown that when a cemented aggregate is fully compacted, its strength increases exponentially as its water to cement ratio decreases. "Achieving the strength corresponding to a given w/c ratio requires full compaction and this can be obtained only with a sufficiently workable mix" (Neville, 1995). In other words, there is a fine balance between the required amount of water for full compaction, and for consistencies that are within a workable range to obtain this compaction. The curve ceases to be followed only when full compaction is no longer possible; all points falling below the curve represent insufficient compaction. The fact that our data follows a similar pattern indicates that the materials were compacted to their greatest degree possible, or very close to it, especially at the higher w/c ratios. Because lower w/c ratios are supposed to give greater strengths, the position of the data that fell below the curve at such low ratios indicated that those mixtures were most likely not fully compacted. 4 The trend line, extracted from Neville's "Properties of Concrete" (1995), did not list any particular values, only the apparent pattern relating the strength of concrete to its relative water to cement ratio. 52 5.2.3 FILL GRAIN SIZE DISTRIBUTION There is a common understanding that the strength of fully compacted concrete with a given w/c ratio is independent of grading, that the only importance of grading is in its workability effects. However, "the total volume of voids in concrete is reduced when the range of particle sizes from maximum to minimum is as large as possible" (Neville, 1995). Neville also discusses the requirements of aggregate cohesiveness and workability that are opposed to the absence of segregation. For a mixture to be cohesive and workable, it must contain a sufficient amount of fines, but there is a fine line between the advantages and disadvantages of its addition. As the amount of fines increases in a mixture, the volume of water required to maintain workability increases, especially if the fines are angular. This increase in water changes the water to cement ratio, and affects the potential for strength gain, as noted above, and if too much water is added with respect to the amount of fines present, this can cause segregation problems. This is also true where there is too large a size gap between the coarse and fine components; although it is easier for particles of different sizes to pack together, this form of aggregate selection also makes it easier for fines to pass out of voids. While the gradation of the mixture does have an effect on its workable characteristics, it is believed for the reasons stated above that the quality and strength of the fill is not independent of its gradation, but in fact, very dependent on it. In order for an aggregate mixture to become fully compacted, with minimal voids content, and at reasonable compaction efforts, the materials within that mixture must be able to settle in an orderly, dense structure. This will only occur if the gradation of the mix is such that the gradation is not uniform, that there is a range of sizes where the finer particles fit into the voids of the structure formed by the coarser aggregate structure. The presence of too many voids allows for movement of the particles on the application of force, and potential failure of the structure formed by the aggregate particles. Observations made during our studies support this statement, discussed below. 5.2.3.1 Relating Grain Size Distribution to Fill Strength Test results were separated into groups based on the pattern depicted by the sample grain size distributions (full, harsh and gap gradations)5. Graphs illustrating these gradations are presented 5 In this graph the 100% rock mixture has been designated as both full grade and harshly graded. This was done because when the distribution of the aggregate grain size was analyzed, it was noted that the aggregate is fully graded within its maximum and minimum grain size range. However, when compared with the other mixtures analyzed in this study, its distribution lies in between other harshly graded and fully graded mixtures because its range of particle sizes is not large. 53 in Appendix D. These groups were then plotted on a graph (Figure 5.8) of strength versus water to cement ratio, as in Section 5.2.2. In doing this, better agreement was achieved and three trend lines were established. Figure 5.8 represents all data for the 109 and NEX fill samples containing 9% cement and shows that the mixtures of full gradation have greater strengths than those that are gap- or harsh-graded. There is also better agreement (R2 = 0.89) for the trend line pertaining to the fully graded mixtures than for the others. This is expected since the materials compacted to a higher density and there was less likelihood for materials to fail (due to the presence of large voids and the potential for materials to shift) when subjected to the application of force. re Q. 25 20 15 U) Figure 5.8: Strength vs. Water-Cement Ratio & Grading y = 9.995x"10J1 R 2 = 0.8885 y = 6.0265X"1 u " R2 = 0.6615 • Harsh Grading • Gap Grading X Full Grading y = 3.0537x2 0 1 2 9 R 2 = 0.3953 2 3 Water-Cement Ratio This observation is supported by the various failure mechanisms noted during testing. Most full grade mixtures, including those containing up to about 60% tailings, failed with a typical pattern of outward buckling and the breaking of cement/aggregate bonds. For mixtures that were poorly graded (including some gap gradations) and with low density, indicating low rock content and/or high voids content, failures were notably by collapse of the aggregate structure and/or fracturing through a fines and pellet matrix. All mixtures that were predominantly made up of tailings showed planar failure through the fine matrix of the fill. In fact, it was common to see horizontal micro-cracks developing throughout such samples prior to the 28-day testing stage. Because these fills were typically harshly or uniformly graded, this cracking has been attributed to the 54 resistance of the shrinkage mechanisms that similarly occur during the later curing stages of cement paste or mortar (Neville, 1995). In the case of the pellet mixtures, shrinkage of the paste (cement, tailings and water) bonded to the agglomerates is restricted by the relatively large size of the pellets and their inability to shift with the paste to further compact. The tensile forces developed by the shrinkage mechanisms are not strong enough to disrupt the structure of the coarser particles. In the case of the tailing mixtures, the shrinkage of the paste (cement and water) bonded to the particles (tailings) is inhibited by their angularity (to a lesser extent than in the situation above), and the resulting inability of the tailings to compact. To a large degree, the rough edges of the particles get caught on one another, and resist further movement. This was noted in the case of the 100% tailing samples produced at 9% cement. Prior to the 28-day testing stage, the samples failed in tension (indicated by a convex surface) most likely due to such labored shrinkage mechanisms. In both situations, the restraint would have caused the bond between the cement and aggregate materials to weaken or break, and the structure collapsed with the application of minor pressure. In opposition, there are occurrences in the graph above where sample strengths within the harshly graded category, at low w/c ratios, deviated from expected strength ranges. The structure of these fills, shifting from greatest strength to lowest, consisted predominantly of rock, to rock and agglomerates, to predominantly agglomerates; no tailings were present in any of these fills, and so all contained some volume of voids. For this reason, it is expected that we would see lower strengths for these samples. However, these mixtures failed at greater strengths than some of the gap-graded fills and came close to the strengths of the full grade fills. Because the strengths increased so drastically with the increase in rock content, this phenomenon was attributed mainly to the strength of the individual components. 5.3 SUMMARY In both the agglomerate and backfill testing programs, it was noted that grain size and shape are important in that they partially control the level of compaction that can be approached. Moisture content is important in that it controls the workability of the mixture, which affects its potential level of compaction, and the resulting w/c ratio, which in turn controls the ultimate strength of a fully-compacted sample. Finally, the level of compaction that is ultimately achieved within the 55 cemented aggregate is an important factor in the material's strength. In most cases, the lesser the void content, the greater the strength. 5.4 F I E L D C O N D I T I O N S & O N S I T E E X P E R I E N C E The one drawback of this testing program is that it was never tested on site, and so the effects of field conditions were never analyzed. There are a number of issues pertaining to the use and placement of fill on site. 5.4.1 ON SITE QUALITY CONTROL & LAB TESTING As part of the quality control operation at the mine, control samples6 are regularly taken from the backfill batch plant, and tested in compression at preset time periods. The mine does not regularly extract or test core samples from their underground filled stopes since experience at the site shows that the fill being produced is very competent, and stands up well during excavation of neighboring stopes. The mining operation experiences minimal dilution factors caused by blasting, and in general, the fill is "tight to the back" and very solid on inspection after such blasting (Sigismund, 1999). All of the test data is recorded in a database at Eskay Creek Mine, but as can be seen in Figure 5.9, the strength of the fill varies significantly. Figure 5.9: Batching Process Control Chart 7 Day-1999 Yield Strength •5 Day Strength 5 Day Water 30 -, 2^Jan 2-Feb 5-Mar 5-Apr 6-May 6-Jun 7-Jul 7-Aug 7-Sep 8-Oct 8-Nov 9-Dec Date (1999) A l l fill samples consist of river aggregate, cement and water only. There are no additions of fillers, tailings or agglomerated tailings in any of the graphed samples. 56 With a variety of people performing tests on the samples, it is likely that some of this variability may be attributed to human error and compaction efforts. Additionally, the aggregate source is a natural deposit, and while the large size fractions are scalped off, the material may be variable in gradation, having some effect on sample strength as described earlier. Also, slump tests are performed and measured only on quality control samples, and the consistency of the fill is generally controlled by visual inspection and the operator's experience with working in the batch plant. Water is added at a rate controlled by the operator, and so its content also becomes variable. All of these factors - gradation, water content, compaction effort, and human error affect the sample strength to some degree. As discussed previously, the water to cement ratio largely controls the ultimate strength achievable within a fully compacted cemented aggregate. Test results in the concrete industry indicate a reduction in water to cement ratio should improve strength. However, the mine has found their fill strength and quality improves with increased addition of water, as shown in Figure 5.10 below. c V </> >. a a oo CM 35.0 30.0 25.0 20.0 15.0 10.0 5.0 0.0 Figure 5.10: 28 Day Strength 0" Slump 0" Slump 0" Slump 0" Slump 0.5" Slump 35 litres 35 litres 45 litres 50 litres 60 litres 0.34 0.34 0.40 0.42 0.46 Slump/Water/Cement Ratio While this is likely due to improved workability of the fill, it shows that the w/c ratio may not be the most important factor when the fill is produced in the field, especially when there is only moderate control on the mixing operation. 5.4.2 METHOD OF PLACEMENT An important factor affecting the potential for developing a fill of low void content is the method by which it is placed. In the case of Eskay Creek Mine, compaction efforts are low and cannot 57 necessarily be improved due to the "ram into place" method used for fill placement. This method of placement is such that voids will always be present. For this reason, the strengths, proportions, gradations, and resulting structure formed by the mixed constituents mainly control the void content of the fill and its strength. With respect to the gradation, the mine would need to take some care in developing a mixture that would naturally settle into a moderately dense fill, and let underground pressures contribute to its consolidation. If the fill is 'self-healing', the fill will compress over time - if it does so slowly during the initial stages of curing, the fill will further compact, and become stronger in the long term. 5.4.3 I M P L E M E N T A T I O N In order to implement this program on site, the mine would need to take a course of action that would include development and physical characterization of agglomerates, the development and testing of various mixtures of fill based on the results of UBC's test program, but using agglomerates produced on-site, and the determination of the best water to cement ratio to use in the best compacting mixture in order to balance strength development, workability and self-settling compaction effects. Since some of these factors have already been tested, the program need not be extensive. The test results showed that agglomerated tailings could be utilized as a component of underground aggregate fill. Although the strength of the fills produced were not as great as the aggregate fill currently used at the mine site, the strengths produced did meet the design criteria for underground support. The best degree of compaction, without addition of unagglomerated tailings, was achieved where 40% of the river aggregate was replaced with pellets made with 8% cement. Optionally, it is likely that this proportion could be increased if the pellet size distribution were improved, either by increasing the range of particle sizes, or by decreasing the average sizes of the pellets and their gradations. However, this increased proportion would be highly dependent on the altered gradation of the pellets produced after the production controls were established on site. The addition of tailings materials to the mixture described above did not improve strength to a large degree during testing. However, it must be remembered that the strength of a fill made in the laboratory will never fully represent a fill made in the field. This is due to effects of large-58 scale production and placement methods. If desired, tailings or some other additive could be incorporated into the fill if its voids content seemed to be high; the addition of minor fines would help to fill the voids and limit segregation of cement slimes. However, the addition of fines would also increase the water requirements of the fill if the same level of workability were to be maintained. Therefore, it is important to find the balance between the beneficial addition of void-filling fines, which would increase the strength of the fill, and the corresponding detrimental addition of water, which would maintain the workable nature of the fill, but decrease its potential strength. In summary, the mine could use several of the fills produced during this testing program because most provided sufficient strength for underground structural requirements. Again, the most important issue would be to develop a fill that is sufficiently workable and that has self-settling characteristics to augment the development of low voids content and improved strength. 5.5 LONG-TERM CLOSURE ISSUES In the case of underground mining, long-term issues associated with the use of a cemented fill include the loss of the fill's strength, and the chemical stability of the fill, which could lead both to loss of fill strength as well as contamination of flowing groundwater and ultimately, the surface waters that receive this groundwater. Factors that are somewhat linked that could detrimentally affect the fill's long-term strength include the insitu density of the fill, the pressures of the underground environment and the groundwater conditions during and after placement. Underground mining environments are subject to variable underground pressures which can result in formation of fractures allowing concentrated flow patterns of groundwater to develop, especially in the case where fill materials are placed loosely with high voids contents. The formation of such fractures allows saturated conditions to penetrate the fill to greater depths, leading to the potential for chemical breakdown of the fill. Conditions that "can be detrimental to the chemical stability of cemented fill are: sulphate attack from sulphate rich acid rock drainage (ARD) generated within the mine or present in the tailings water used to make up the backfill mixture, and water dilution of the cement phase of incompletely cured backfill" (Bertrand, 1998). 59 5.5.1 LONG TERM LEACHING Backfill is typically exposed to the environment immediately after placement. Exposure to an aqueous leaching environment prior to being properly cured will promote the loss of binder material. This is supported by Aylmer, who in 1973 determined that if the pH of the water were below about 5.8, the cement bond would be leached. Over the long term, there is the potential that the cement coating could be entirely leached from the fill material, potentially exposing metal-bearing and/or sulphide-bearing particles to the flow of the surrounding groundwater. In the case where the groundwater contains acid, as is common in the case of a mine setting, this can lead to sulphate attack and acid generation. 5.5.2 SULPHATE ATTACK Sulphate is typically present in mine groundwater and in the wastewater generated by ore processing. Often greater than 1.5 g/1, such a solution would be classified as aggressive water in the cement industry. The problem that develops with such a solution is when free sulphate ions present in solution combine with the calcium of dissolved portlandite (in cement) to form gypsum, creating pressure that can induce cracking. A zone of lower strength is formed in the calcium-leached areas due to the increased porosity from the leaching of portlandite (Bertrand, 1998). As the degrees of cracking and porosity increase, the solution is allowed to penetrate deeper into the fill material and further deteriorate its quality. This is especially dangerous where materials within the fill consist of sulphides that become reactive on exposure to oxidative environments (potential in concentrated sulphate solutions); the sulphide materials become exposed to the penetrating solutions when cement hydrate coatings begin to deteriorate, providing the potential means for the development of concentrated acid drainage. 5.5.3 ACID DRAINAGE POTENTIAL Eskay Creek's tailings have been classified as potentially acid generating. Tailings are classified as potentially acid generating (PAG) in the short or long term when they contain reactive sulphide minerals that, on exposure to oxygen and water, will break down to produce acid through oxidation and dissolution processes. In a study on paste fill, Bertrand (1998) showed that fills containing high proportions of sulphide minerals likely require complete submergence in water at early stages to prevent oxidation of the 60 sulphide portion of the tailings; such oxidation was observed after only 14 days on one sample containing the greatest amount of sulphides and the greatest amount of cement. Bertrand also proved wrong the beliefs that cemented materials could buffer the onslaught of ARD solutions. As part of her investigations, samples were subjected to a cyclic, man-made, acidic (ferric sulphate) leaching environment. The paste fills developed a surficial coating of iron hydroxide precipitates and a second inner alteration layer of amorphous, iron sulphate precipitates, implying the penetration of the solution front into the paste. While Bertrand suggested the possibility that the precipitation of these minerals could eventually passivate the surface of the samples, the implication was that fill exposed to ARD would eventually lose its ability to buffer the drainage. 5.5.4 METAL LEACHING "Associated with the development of acidic environments is the solubilization of heavy metals such as copper, zinc, lead, arsenic, antimony, nickel and cobalt together with cations such as aluminum, iron, silica, potassium, calcium and magnesium associated with carbonate and silicate minerals" (Lawrence, 1996). Dissolution of these metals can be harmful to the environment if the amounts of these metals become too great, too quickly, in such a way that the environment cannot adapt to their presence. In essence, the concentrated metals become lethal to the biota feeding upon the groundwater containing them. 5.6 R E C O M M E N D E D F U T U R E S T U D I E S In closing, the research conducted should not be considered complete until the following items are examined: 1. Testwork should be done to characterize if the addition of cement and the process of agglomeration will impact the ARD potential of the tailings materials. 2. Testwork should be done to determine the long-term ARD and metals leaching risks associated with backfill containing agglomerated and/or unagglomerated tailings. 3. Testwork should be done to determine if the short-term quality of the fill is affected by potential leaching of cement from the fill and/or agglomerates by local groundwater flow, existing ARD at the site, or from exposure to any ARD generated by the fill material. 61 Testwork should be done to evaluate the potential of sealing the pellets from water infiltration by using a reagent such as sodium silicate. This could solve the long-term ARD-mine closure issue. 62 6 COST ANALYSIS In Section 2.5, various options for disposal were discussed with respect to site constraints at Eskay Creek. The option of subaqueous disposal is already in use at the site. The options involving on-land disposal, including slurry, thickened and filtered tailings, while not impossible, are not well suited to the site, particularly since there are available alternatives whereby land areas need not be disturbed. The option of underground disposal was suggested and a method to implement, whereby tailings are agglomerated prior to addition to backfill, was investigated. The research results suggest the procedure is technically feasible. In the following sections, the requirements to implement tailings placement as underground fill (agglomerated) or to continue the practice of subaqueous waste disposal are discussed. 6.1 AGGLOMERATION OF TAILINGS & DISPOSAL IN BACKFILL To implement an agglomeration program and combine it with the backfill plant already in use, the Larox filter press must be retained. The requirements for the system are listed below. 1. A 75-tonne cement silo, with dust collection system, hopper and feed system, to provide for the system over a one-week period. 2. A fabricated storage bin that would sit below the Larox filter to catch the filtered tailings. The bin would require a hopper bottom to regulate discharge of the materials to the mixing system. 3. An 18"-wide conveyor belt, installed on a 30% incline, to raise the elevation of the discharge point over 60 feet. 4. A pin mixer, rated at 10 stph, into which the conveyor belt would discharge, for the mixing of cemented tailing materials and for the nucleation of pellets. This mixer is to discharge onto the pelletizer. 5. A disc pelletizer, rated at 10 stph for agglomeration of the cemented tailings. A fine water spray would be attached to add water during agglomeration, as required. 63 6. An 18"-wide conveyor belt to catch the agglomerated tailings as they discharge from the pelletizer, and carry them to the storage area. Controlled, moveable ploughs would need to be installed on the conveyor belt at each desired point of discharge. 7. An 18"-wide movable conveyor with a raised end for discharge into the storage piles. To minimize the possibility of fresh pellets from sticking together, a method similar to the "Grangcold Process" can be utilized (Adnan, 1977). This has been used in the storage of agglomerated iron ore concentrate, to prevent the tendency of sticking. Grangcold Process: Pellets are mixed with fresh iron ore concentrate equivalent to one third of the pellets weight. The fresh iron ore concentrate is used as a packing material and fed together with the pellets into large storage bins. Early hardening requires 3 to 6 days minimum, during which time -70% total strength is gained. Pellets are then screened to remove the excess fines and stored in outside stockpiles to complete the 30-day setting time (when temp +0). Agglomerates could similarly be poured into the storage areas in combination with a small volume (< 20%) of unagglomerated tailings. The fine dusting of tailings would coat each pellet and provide a 'cushion' between neighboring pellets. When the pellets are batched for mixing into the fill, this fine material can be incorporated into the fill. Cured materials would then be carried over to the backfill plant for use in the agglomeration system. (The allowance for a loader operator is already in place, and can be used here). In the following figure, the above assembly line is illustrated. At this point, no mass balance has been incorporated into the figure or any calculations because the process would vary depending upon which type of fill the mine would be producing and the preferred mixtures of the mine officials. 64 FIGURE 6.1: OPTION #1 Agglomeration & Backfill Production Underground Disposal 6.2 SLURRY PIPELINE & SUB-AQUEOUS DISPOSAL The option of subaqueous tailings disposal has been used at the mine since commencement of mill operations and tailings production. The mine currently trucks the tailings to the disposal site where they are dumped. The mine could continue disposing of the tailings in this way without 65 any changes to their operation, with the exception of the need for a larger disposal area than Albino Lake, where tailings are currently being dumped. This describes the status quo option. Recently the mine began investigating if tailings could be pumped as slurry to a new proposed disposal site (Sigismund, 1999). A major benefit of this proposal would be elimination of the Larox filter press from the mill process, an action that would significantly reduce operating expenses. To carry out the cost analysis and determine an appropriate location for a slurry tailings line, Eskay Creek personnel provided maps of the site (Sigismund, 1999). A nearby body of water, Tom McKay Lake, is the likely candidate as it is the largest in the area, is relatively close to the site, and the rise in elevation between the mine and the lake is moderate in comparison to other locations. The distance is approximately 3 km, and the rise in elevation about 200 m to the disposal site. For general design purposes the incline was assumed to be gradual, with no relatively sharp bends in the pipeline to affect flow efficiency. From the underflow of the current thickener, the specific gravity of the slurry material, at 40% solids, is 1.36, and the flow rate 65 US gpm. General design parameters used by AGRA Simons (Johnston, 2000), were used to size the pumps and pipeline. These are as follows: • 3% friction head loss at a flow velocity of approximately 2 m/s, and • Break HP of approximately 45 HP - calculated based on flow rate, total head, specific gravity and efficiency. From these parameters, it was determined that 3 pumps would be required in series, each connected to a surge tank with agitator to build the required head to transport the materials to the disposal site in question. The pipeline is 3 inches in diameter, and requires insulating due to its relatively small size and the risk of materials freezing during the winter season. Figure 6.2, illustrates the system required for pumping tailings slurry to the disposal site. Again, no mass balance has been incorporated into the flow diagram. While the thickener does run at approximately 7 tonnes per hour, the pumps would only run over various periods of time, running only when each surge tank is filled with tailings slurry. 7 As requested by Homestake personnel, these maps have not been attached with this study. 66 FIGURE 6.2: OPTION #2 Pump & Pipeline Scenario Subaqueous Disposal STIR OF. TANK ST7ROE TANK PTIMPS & PIPELINE TO T A K E 6.3 COST COMPARISON OF DISPOSAL OPTIONS For the two options described, an order-of-magnitude cost estimate was carried out. In the capital cost estimate, general delivered-equipment costs were determined using various costing books (Richardson, 1992, Schumacher, 1994), and average prices developed for similar projects (Smalcel, 2000). Also utilized were supplier lists and information provided directly by those employed at supplying companies (Evans, 2000 and Wolff, 2000). Each of the items listed specifies the source of the pricing, and capital cost indexes used to upgrade researched costs to 67 current pricing were obtained through the Western Mine Engineering website (Schumacher, 1994). Operating costs were calculated based on standard hourly rates for labour, power requirements and standard equipment operating costs. Personnel on site provided estimates for cost savings pertaining to a reduction in aggregate hauling for the first option and removal of the Larox filter press for the second option (Sigismund, 2000). The net present value (NPV) was then calculated for the life of mine using a discount rate of 6 percent. A low discount rate was chosen because the disposal process must be performed. No salvage or return on investment for purchasing the equipment is expected and for all options, heavy water quality monitoring would be required throughout operations and after closure. There would be no guaranteed long-term savings for any of the disposal options except for the elimination of the filter press from the second option. Tables 6.1 and 6.2 summarize the capital and operating costs of the individual items for the agglomeration and pipeline systems described in Section 6.1 and 6.2, respectively. For full details, see Appendix E. 68 s o 2 < U o o un o r~ o OJ o o 00 co o <o un °l un i n o to o" to" CM" CO cd -a-" o" t— co un CO CM «/» m o o •>a- in i — CM_ * ^ CM" CM O O O r~ TJ- oo o ID oo o un V* *B ^ CO C O % O 0) C/3 CO 1 -2 O O o CO X CL co "cu •D o CD CD 3 I Si ui O <S>-E cu a. ™ CL E CO T3 C g 3 CL c CU O-E CO •a c o CO tfi c o t> 3 i _ in c o u •a CO o cu o c CO c 'g. 'a. E 2 >, in CO = ~ 3 CU CO CO x: o !2 - i ^ W W I 43 8-o ! •§ C/3 pj to E •o CD c c o CO ^ 3 i j o3 O) 33 cu 3 m cn q o 3 S <3- £ a c £ & .3 O 3 ® O CL 2 CD Q. O 0) "D CO O 1> <3 3> 3> 43 o _ O O OJO 2 o P -3 2 a O C/3 C/3 O o o o LO o o o CD o o CO un o CO O o un <— CM r-~ 9 co" 1— co" o" CD CM" CM_ LO 1— t— T— CO «•» v» </» </* LO ro" in CM CM to CO o o un o un CM CM CM 11 C a O e o 1 s o CO CU a. a. o x: E o o n c 00 CD CO ^ 2 «3 «^  c o 13 E CO 0) LL. O CO o O) E •§ Q) CO 5? o\ a. o a) o 0 >i x $ I O CL CU •3 5 cu o >1 cu > c o o o a? > c o o JD co ' cu > o E CD C O 43 to O U 1 C/3 c o cu 0 cu D. I > c O c ca al 0 3 O S E S E- c E a o ,aj u o c o 0 Q. O c o D) .C co S 5 U]| c CD I. 3 D •a a) 3 CO x: ui cu c c o c o (U CO ro ~ CO g_ O l/l 01 > 3 ro ro 0 1 o l/> 3 O L ± O CU Q. O k_ CU T3 CO o s s 3> >& 43 43 i i 2 2 a 5. 0 O 1 1 3 3 C/3 C/3 6.4 FINAL SELECTION OF DISPOSAL METHOD In summary, the results of the total cost analysis are shown below: Table 6.3: Summary of Costs OPTION CAPITAL COSTS ($ CAN - 2000) NPV OF TOTAL INVESTMENT STATUS QUO: AS IS $0 $0 #1: AGGLOMERATION $513,575 $ - 1,380,630 #2: PUMP & PIPELINE $ 530,662 $ - 465,035 Results of the cost comparison indicate that the most economic method for the mine is to install a pipeline to pump thickened tailings slurry to a body of water and dispose of the materials underwater. However, there are uncertainties not yet addressed that could preclude implementing such a disposal program. These include: 1. Short-term effects of mercury on the lacustrine environment and the surrounding environment, since the tailing materials contain a relatively high level of mercury (Dai, 1999). 2. Long-term leaching potential of the tailings, related to the above, but also in regards to all potentially dangerous substances and minerals within the material. 3. Time of settlement required for materials, to ensure turbidity levels are low to non-existent in the lake and waters draining there from. 4. Capacity to contain all tailings under the euphotic zone of the lake for the remaining mine life, including extensions of the mine life due to discovery of new reserves. No testing was carried out by UBC to indicate leachability of the materials in a submerged environment. However, cyclical leach tests carried out on exposed tailings did show mercury and antimony, as well as elevated sulphates, in solution (Dai, 1999). With the slurry pumping option, this could be potentially dangerous for aquatic life and local wildlife, as not only do these metals affect the water, but also the vegetation in the immediate vicinity of the water body. Because of these uncertainties, a more detailed investigation of the costs would need to be carried out before discounting the agglomeration option. 70 7 CONCLUSIONS 7.1 GENERAL There are a number of technical conclusions that can be made from this study. The most apparent observations are listed as follows: 1. With few exceptions, the strength of the products (agglomerates and backfill) increased with an increased cure period and an increase in cement content. 2. The strength of most materials increased when the water to cement ratio was decreased. 3. The effect of the water to cement ratio became apparent only after a sufficient level of compaction was achieved within the 'aggregate' mixtures. 4. The strengths of the agglomerates, while fairly strong to pressures of the hand, were much weaker than the rocks being replaced. 7.2 AGGLOMERATION The following conclusions can be made regarding the production and testing of the tailings agglomerates: 1. Both tailing types can be agglomerated with relative ease. The main factors controlling development of quality agglomerates are the amount of water and its point of addition. 2. A humid environment is not required to entirely cure the agglomerates. On average, those pellets cured outside of the humidifier (group B) after seven days, were stronger than those cured in the humidifier {group A) for the entire 28-day cure period. 3. The size distribution of the pellets produced was uniform in comparison to the river aggregate being replaced. This led to an increase in voids content and decrease in fill strength with each increment of aggregate replacement. To reduce the potential strength loss due to this increase in voids, it is advisable that the pellets be produced to a fuller grain size distribution for use in the fill. 71 4. At 8% cement, the pellets were of very good quality and reached strengths of close to 50 MPa. These strengths are similar to that of some medium strength rock types such as sandstone, schist and shale. 7.3 BACKFILL The following conclusions can be made about the production and testing of backfill: 1. Agglomerates can be used successfully in backfill as a substitute for river aggregate. However, with each increased level of replacement, a relative drop in strength proportional to the strength of the replacement materials and the gradual introduction of voids is expected. 2. The degree of compaction or final void content that develops within the 'aggregate' structure has a significant effect on strength. The lower the void content - the greater the strength. 3. The size distribution of the total mix has a significant effect on the potential degree of compaction achievable. The wider the range of particle sizes and the fuller the gradation, the more likely the particles will form a dense, highly compact structure. 7.4 CLOSING The benefit of agglomerating tailings and using them in underground fill is a reduction in the need for their disposal in a surficial environment. For Eskay Creek Mine, this is especially so. The site has the capacity to contain all of its tailings underground due to the volume of ore that is excavated and shipped directly to smelters. While the cost analysis has shown that the economic feasibility of the project is questionable for the Eskay Creek site, the test results do show it to be technically feasible. 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"A Predictive Quantitative Geochemical Assessment of the Subaqueous Stability of Acid-Generating Mine Tailings", in Proceedings of the 4th International Conference on Acid Rock Drainage, Vancouver, May 31 to June 6, 1997, p. 237-254. Nehdi, M. and S. Mindess (1999). "Microfiller Partial Substitution for Cement", Chapter in Materials Science of Concrete VI. The American Ceramic Society, Y. Skalny and S. Mindess editors, 49 p. Neville, A.M. (1995). Properties of Concrete. 4th Edition. Longman Group Limited, England, 844 p. Nicholson, R.V. (1994). "Iron-sulfide oxidation mechanisms: Laboratory studies", Jambor, p. 162-183. Ouellet, J., M. Benzaazoua, S. Serfant (1998). "Mechanical, Mineralogical and Chemical Characterization of a Paste Backfill", in Proceedings of the 5th International Conference on Tailings & Mine Waste, Fort Collins, Colorado, USA, 26-28 January 1998. Balkema, Rotterdam, 968 p. Peters and Timmerhaus (1980). 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(1965). Formulas for Stress and Strain. 4th ed.. McGraw-Hill, 1965, p. 318. Robinsky, Eli I. (1978). "Tailing Disposal by Thickened Discharge Method for Improved Economy and Environmental Control", Argall, p.75-95. Robinsky, Eli I. (1998). Paste Technology for Underground Backfill and Surface Tailings Disposal Applications. Short Course Notes. Eds. Landriault, et al., Golder Paste Technology Ltd., January 25, 1998. Roth, T., John F.H. Thompson and Tim Barrett (1997). 'The Precious Metal-Rich Eskay Creek Deposit, Northwestern BC: A Synopsis SEG Short Course at GAC-MAC", 42 p. Roth, Tina, and Jim Rogers (1998). "The Eskay Creek Gold and Silver Mine: Opportunities in a Complex Mineralized System in Northwestern British Columbia", Abstract for the 1998 Cordilleran Round-up Coreshack Display, Homestake Canada, 2 p. Sastry, K.V.S., Ed. (1977). Agglomeration 77: Proceedings of the 2nd International Symposium on Agglomeration, Volume 2, March 6-10, 1977, AIME, NY. Schumacher, Otto L., Ed. 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Planning, Design and Analysis of Tailings Dams, John Wiley & Sons, New York, 369 p. Wolff, Mike & Silvia (2000). McLean & Higgins Plumbing & Bathworks. Personal communications and quote. 79 APPENDIX A Tailings Disposal Options A GENERAL TAILINGS DISPOSAL OPTIONS A.l SUB-AERIAL DISPOSAL The most common method used in mining for the "disposal" of tailings is through their sub-aerial deposition in tailings storage facilities. In the past, tailing materials were generally deposited in the form of slurry within large impoundment structures. Today, many sites continue the deposition of tailings in this way; wet spigotting is the cheapest, least labour-intensive method of tailings disposal. However, these structures involve large water management programs, and many companies worldwide have been looking for other methods of disposal. To avoid large-scale water management issues, some mining companies have recently initiated the deposition of tailing materials in a denser state or thickened slurry, even going so far as to filter the tailings to produce 'dry' materials. These dewatered materials may be disposed of without the confines of an impoundment, or with reduced-size embankment structures, constructed mainly for sedimentation reasons. A.1.1 SLURRY DEPOSITION In the next sections, the various models for sub-aerial tailings confinement will be discussed. Although the structures may be used for thickened deposits where required, the focus of the discussion will be pertaining to the storage of slurried tailings, as they represent the most problematic form of disposal in terms of physical stability. The models include ringed dyke structures for relatively flat terrain and four types of tailings embankments for mountainous terrain, including the common upstream, downstream and centreline construction-type dams, and civil water retention-type dams. A.1.1.1 Ringed Dyke Structures Ringed dyke structures are best suited to sites where the terrain is relatively broad, and free of natural topographic depressions. The impoundment is constructed in regular geometrical fashion, the dykes constructed of earthen materials, the entire structure enclosed within perimeter ditches for drainage collection purposes (see Figure A.l). 81 7IIS A\ i l l £ FIGURE A . 1: RING D Y K E STRUCTURES; SINGLE A N D M U L T I P L E (Extracted from Planning, Design, And Analysis of Tailings Dams by Steven Vick, 1983) There are inherent advantages and disadvantages to using such a structure for the storage of tailings. Advantages: 1. As the structures are typically laid out in a regular geometrical shape, various types of liners can easily be installed where required for environmental reasons. 2. Accumulated water results only from tailing slurry waters and that which falls directly on the impoundment surface; all sides of the impoundment are enclosed, eliminating surface runoff from external factors. The extent and expense of drainage, collection and treatment systems is limited. 3. The volume of water being introduced into the system can further be reduced if the tailings are thickened before being discharged into the impoundment. 4. The impoundment may be partitioned; each segment constructed sequentially as another is filled. Although greater fill volumes are required, this pattern of construction reduces seepage volumes, allows concurrent reclamation, and defers construction costs. The structure's only disadvantages are that a relatively high quantity of fill is required in relation to the storage volume produced, and a relatively large area is required in order to store any reasonable volume of tailings. 82 A.1.1.2 Tailings Embankments or Dams Layout In mountainous terrain, there are three layouts commonly used for tailings containment, depending on specific site conditions, including topography, the size and shape of natural drainage courses, as well as availability of borrow materials. These are cross-valley, side-hill and valley bottom impoundments. Cross-valley impoundments are confined by a dam extending from one valley wall to another, and can be applied to almost any natural topographic depression in either single or multiple form. Diversion ditches may be constructed on either side of the impoundment to reduce the amount of surface runoff entering the system, but large channels required for flood peak flows are often not feasible. Also, flood runoff from very large catchment areas can be handled only by storage within the impoundment, spillways, or separate water-control dams located upstream of the impoundment to catch and redirect flows; this leads to large water management issues. To minimize this problem and reduce the volume of flood and natural water inflows, the structures are typically located near the head of the drainage basin (see Figure A.2 below). FIGURE A.2: CROSS V A L L E Y IMPOUNDMENT; SINGLE A N D M U L T I P L E (Extracted from Planning, Design, And Analysis of Tailings Dams by Steven Vick, 1983) Side-hill impoundments are enclosed on three sides, therefore requiring more fill than cross-valley structures. However, these structures may be used where incised drainages are not 83 suitable for the cross-valley option, allowing mining activities to proceed where otherwise they might not have (see Figure A.3). This type of impoundment is best suited for sites where slopes are at less than ten percent grade. FIGURE A.3: SIDEHILL IMPOUNDMENTS; FIGURE A.4: V A L L E Y B O T T O M SINGLE A N D M U L T I P L E STRUCTURES; SINGLE A N D M U L T I P L E (Extracted from Planning, Design, And Analysis of Tailings Dams by Steven Vick, 1983) A compromise between cross-valley and side-hill options, the valley bottom impoundment is well suited for cases where the drainage catchment area would be too large for cross-valley layouts, and slopes are too steep for practical application of the side-hill arrangement. The structure is enclosed on two sides, so fill requirements are intermediate between the side-hill and cross-valley options, and may be laid out in multiple form, stacked as the valley floor rises (see Figure A.4). Diversion channels are used to carry peak flood flow around the impoundment, a requirement as most valley bottom embankments are constructed in relatively narrow valleys, across original drainage channels. The diversion channel is constructed corresponding to the gradient of original stream flow, but tight against the opposing valley wall. The lower portions of the embankment are typically protected by riprap as well to protect against high velocity flow. 84 Raised Embankment Types Independent of the impoundment layout, is the method of construction chosen for the embankment. Depending on design requirements, volume of material available for construction, and the site conditions, four approaches to construction are common. These are: that of raised embankments, including the upstream, downstream, and centreline dam designs and that of conventional water retention dams. The construction of a raised embankment is staged over the life of the impoundment, beginning initially with a starter dike constructed of natural soil borrow, typically sized to impound the initial two or three years' mill tailings output as well as allowances for storage of flood inflows. Subsequent raises are scheduled to keep pace with the rising elevation of the tailings and floodwater storage allowance in the impoundment. The embankments may be constructed with a range of materials, including natural borrow soils, pit mine waste, underground development muck, hydraulically deposited tailings, or cycloned sand tailings. The advantages of raised embankments in relation to water retaining structures are significant. Expenditures are distributed over the life of the impoundment, and the initial project development costs are reduced, producing cash-flow benefits that are often important in financial considerations related to mine start-up. There is more flexibility in the selection of materials for construction; paced construction with pit development allows use of waste materials, especially if haul distances are not great. Also, raised embankments may assume many configurations, as discussed above, each with unique characteristics, requirements, advantages and pitfalls. Upstream Construction An upstream dam is constructed in the upstream direction overlying existing tailings sediments, as depicted in Figure A.5. After the starter dike is constructed, tailings are discharged from its crest to form a beach. The beach becomes the foundation for a second perimeter dike, and this continues as the embankment height rises. The tailings must form a reasonably competent beach for support, requiring that the discharged tailings contain no less than 40% to 60% sand. The major advantages of using this approach are cost and simplicity; a minimal volume of mechanically placed fill is necessary; construction and ongoing operation can be routinely performed with minimal equipment and personnel; and beach sand tailings often provide a 85 convenient source of fill for perimeter dikes, with excavation from the beach and placement by either dragline or bulldozer. Tailings Spigotted spigot Starter dike (./) FIGURE A.5: U P S T R E A M CONSTRUCTION (Extracted from Planning, Design, And Analysis of Tailings Dams by Steven Vick, 1983) This early method of upstream dam construction was limited to very specific site conditions due to the need for controlled phreatic surfaces within the impoundment, its low water storage capacity and its susceptibility to liquefaction. It has been used successfully in dry arid climates, where evaporation losses are high and a minimum of water is stored in the pond. However, under dynamic loading conditions as in the event of an earthquake, failures of these structures have occurred due to the liquefaction of the tailings materials used in construction (Klohn, 1995). Thus, in the modern upstream method of construction, one or more of the following modifications can be implemented to reduce the risk of failure: 1. The starter dam is designed and constructed as a pervious drain to control the phreatic surface in the ultimate tailings dam. 86 2. Extensive upstream drainage is provided, in the form of upstream dams and diversion ditches for surface runoff from catchment areas, to maintain a low phreatic line. 3. Wide sand beaches are maintained between the tailings dam and the pond by recycling pond waters to the mill when possible or discharging pond water to treatment systems as pond elevations rise. 4. Portions of the sand beach lying directly upstream of the starter dam may be compacted to reduce the potential for liquefaction and improve the material's effective strength for future overlying raises. 5. Finally, the downstream slopes of the dam are built to safe angles, as required by the materials used in construction, to reduce the risk of circular or planar shear failure through the dam (Klohn, 1995) Although the application of some or all of the above measures will improve the stability of the upstream raised embankment, control of ponded water and its effect on the phreatic surface remains difficult under the influence of appreciable flood or normal runoff inflows. This type of dam is not well suited for conditions where heavy water accumulation is anticipated due to flooding, long-term seasonal runoff or high rates of mill water accumulation. It should not be used for water retention applications and is also clearly inappropriate in areas of high seismic potential. Due to the low relative density and highly saturated nature of the tailing deposit, liquefaction-induced flow of tailings is highly likely in the event of any earthquake of appreciable size (Vick, 1983). Downstream Construction The downstream approach uses a large volume of embankment materials, as its overall size increases with each raise. After the starter dike is constructed and the embankment filled with tailings, subsequent raises are constructed by placing fill on the downstream slope of the previous raise. Figure A.6 represents the progressive growth of a downstream embankment. 87 Impervious starter dike Ponded water zone Internal drain _2_ FIGURE A.6: D O W N S T R E A M CONSTRUCTION (Extracted from Planning, Design, And Analysis of Tailings Dams by Steven Vick, 1983) These dams are more stable than upstream structures due in part to the fact that none of the embankment is built overtop of deposited tailings, but also because various degrees of density or competency may be achieved as required through controlled placement and compaction of the construction materials. As depicted in Figure A.6, this form of construction also allows for the incorporation of structural measures within the embankment, such as impervious cores, and/or internal drains, for positive control of the phreatic surface. Use of these measures can allow for storage of significant water volumes directly against the inner face of the embankment. To otherwise control the phreatic surface through the dam, a drainage layer is placed at the base of its downstream side. Earle J. Klohn (1995) and Steven Vick (1983) have addressed various advantages to using this system, including: 1. The method is flexible, and may be adjusted to meet changes in construction material characteristics or foundation conditions as the dam elevation increases, as well as changes in design requirements, as time passes. 2. Drainage systems are installed as the dam is built, not prior to construction, as is required for the upstream method of construction. 88 3. This is a convenient form of storage and disposal of non-acid generating waste materials from open pit operations, overburden, or cycloned sand from tailings, as these materials may be used in the construction of the dam. 4. The structures are more liquefaction resistant due to a controlled low phreatic surface within the embankment and the fact that the fill may be compacted. 5. Raising rates are essentially unrestricted because raises are structurally independent of the tailings deposit. The downstream method of construction is well suited to conditions where significant storage of water is required, as its structural soundness and behaviour are essentially equivalent to water retention type dams. However, these dams require careful planning because its toe progresses outwards as its height increases; sufficient space must be left during layout to ensure that the dam remains within property limits, and topographic constraints, as well as permanently established diversion ditches, roads or utilities on the mine site. The major disadvantage to using this approach is the large volume of fill required and the corresponding high cost. The volume of fill required to maintain the dam crest at sufficient elevations to contain both the tailings and the required flood allowances, increases exponentially with time (Vick, 1983). Also, the availability of fill materials on or near the site may impose constraints on its construction. Centreline Construction The centreline approach to construction is in some ways a compromise between the two above, sharing to a degree the advantages of the two, while mitigating their disadvantages. The starter dike is constructed and tailings are typically spigotted from the crest, as in the case for the other two. However, subsequent raises are constructed by placing fill onto both the beach and the downstream slope of the previous raise. As shown in Figure A.7, the centreline of each raise is coincident, the embankment essentially rising vertically with time. 89 Tailings (</) FIGURE A.7: C E N T E R L I N E CONSTRUCTION (Extracted from Planning, Design, And Analysis of Tailings Dams by Steven Vick, 1983) The dam's material requirements lie midway between that of the upstream and downstream construction methods, and while an above-water beach must be reasonably competent and maintained for support of the fill to be placed upon it, its width need not be large. As shown in Figure A.7, internal drainage zones may be incorporated in the design, so control of the phreatic surface is not as sensitive to the location of the ponded water as in upstream construction. The main body of the embankment (downstream portion) may be compacted and saturation levels are controlled by internal drainage, so it has generally good seismic resistance. "In the event of liquefaction of the beach tailings, limited failure of portions of the upstream fill placed upon the beach can occur, but as long as the central and downstream portions of the embankment remain intact and if water is not ponded directly against the embankment, the overall integrity and stability is generally considered to be unaffected" (Vick, 1983). Still, the dam is not well suited for permanent storage of large depths of excess water. The costs of construction are intermediate between the upstream and downstream methods, due to amount of materials required per year and the compatibility between fill requirements and fill production rates are not as great a problem as that for downstream construction. 90 Conventional Water-Retention Dam Construction "Water retention dams are constructed to their full height prior to the beginning of discharge into the impoundment and differ little from conventional water storage structures in appearance, design, or construction" (Vick, 1983). The embankment is constructed with fill that consists of native soil borrow of various types with internal zoning, shown in Figure A.8. The various layers that are typically included are an impervious core, drainage zones, appropriate filters and upstream riprap. The selection of filters, internal seepage control, and slope stability are designed according to conventional earth dam technology. However, as tailings storage dams do not experience rapid drawdown, the upstream slopes are often steeper than those of their conventional water storage counterparts. Impervious core FIGURE A.8: W A T E R - R E T E N T I O N T Y P E D A M (Extracted from Planning, Design, And Analysis of Tailings Dams by Steven Vick, 1983) Water retention structures are best suited to tailings impoundments with high water storage requirements, such as is the case for large storm runoff inputs or situations where mill process constraints prevent recirculation of discharged mill effluent. Where a relatively small amount of water storage is required, a raised embankment may be constructed. A.l.1.3 Design Considerations For all types of embankments, the standard geotechnical and environmental concerns apply, including seepage, failure, piping and seismic stability. The basic design requirements, extracted from the Keynote Address presented by E.J. Klohn at the Canadian Dam Safety Conference of October 1995, may be listed as follows: 91 /. The foundation soils at the embankment site must be competent to support the proposed structure without the danger of shear failure or excessive movements. 2. The dam, its foundations and abutments, and the reservoir must be relatively impervious to prevent excessive seepage either through the dam or into the subsoil and bedrock. 3. The structure must have stable upstream and downstream slopes. 4. The design must include adequate drainage features to control the phreatic surface within the embankment and uplift pressures in the downstream foundation soils. Internal drains may be required at areas where cracking of the embankment is considered possible, a problem that may cause concentrated seepage flows leading to the ultimate failure of the structure. 5. The structure must be designed to resist predicted earthquake forces for the region, including the effects of the probable liquefaction of the stored tailings. Where cycloned tailings are used to a large extent in construction, compaction efforts and/or special drainage measures must be realized to guard against liquefaction risks. 6. In the case of dams, the structure must be designed to handle maximum possible flows, such as flood events, or higher-than-average surface runoff from the catchment area behind the dam. Overtopping must be avoided, as it may cause failure of the dam; therefore, the dam must be constructed with sufficient freeboard to store these flows, or the design must include an appropriate spillway. 7. Fill materials suitable for construction must be available in adequate quantities. 8. Sufficient instrumentation is required to determine that the structure is performing the functions for which it was designed. In addition to monitoring water downstream of the dam to protect against pollution, settlement, pore pressures and seepage measurements should be taken from within the dam. 92 There have been many large-scale failures of dams around the world, beginning from the time that their construction was initiated until today. Such cases have been studied in detail to determine the causes of failure, in order to prevent similar occurrences from happening. Many lessons have been learned; the general issues may be subdivided under the following categories: • Geotechnics - the parameters requiring measurement for the deliberation of impoundment stability is more apparent, i.e., foundation soil parameters, including strength, compressibility, and permeability, surface water and groundwater flow patterns of the site, etc. • Flood events - the one in one hundred year flood event is no longer used as a maximum value. Instead, the probable maximum flood is determined for each site and is applied in design. • Seismic design - studied since the early 1970's, seismicity factors are now incorporated into design of the embankment using modern methods of analysis. • Site selection - the geology, hydrology and groundwater flow patterns, as well as the historical stability (previous ground movement or landslides) in and around a potential impoundment site are now investigated in greater detail. • Rehabilitation - measures must now be incorporated into the design prior to the permitting of the disposal of mine waste. New techniques are continuously being studied and applied to ensure that there is minimal disturbance to the environment following the closure of a mine. • Management - poor management practices of the past have been acknowledged, and better track records, including sampling and test results, method of placement and compaction, installation records of monitoring equipment, personnel on site, and design engineers involved in the project, are kept for accountability issues. All this being said, current legislation still requires a tailings dam inspection once every year by a geotechnical engineer with expertise in this field. This is typically done by not one person, but by a committee of specialized engineers. The inspection is carried out to ensure the quality of 93 the dams being constructed and to guard against unexpected failures of both abandoned and operating structures. A.1.2 THICKENED TAILINGS DEPOSITS "The primary aim of this form of disposal is to eliminate high and expensive dams and elevated 'slimes pond' and decant systems" (Robinsky, 1978). The approach to using this form of deposit involves the thickening of whole tailings (sand and slimes fractions) to a high pulp density, at which the tailings-water mixture behaves more like a highly viscous fluid than the liquid type slurry consistency at normal discharge pulp densities. The thickened mixture is pumped to the disposal site, and discharged from an elevated position to form a pile in the shape of an inverted cone. The system is adaptable to any topography. On flat terrain, the discharge point is located near the center of the deposit area to increase its potential capacity; the tailings line and discharge point are supported by a ramp built to an appropriate elevation. The ramp height is increased at two or three year intervals as the impoundment grows. In steep terrain, or valleys, the discharge points may be fixed permanently on high ground overlooking the disposal area. The thickened nature of the materials reduces the need for large sedimentation ponds, the reasoning discussed below. The deposit is, however, accompanied by a small pond, located at the lowest point of the disposal site, that receives both natural and tailings-water runoff. Very small fractions of tailing fines reach the pond due to the conical shape of the deposit and the inclined tailings surface, which tends to disperse runoff, and promote good drying conditions. If the site is chosen appropriately, all runoff water will reach the pond naturally, eliminating the need for extensive drainage systems. Because the pond level remains essentially unchanged throughout the life of the project, only a short overflow channel, spillway or culvert at shallow depth is necessary, and a small secondary pond for chemical treatment may be established for the recycling of water where necessary (Robinsky, 1978). It is advertised that due to the consistency of the materials, the tailings assume a slope of approximately 6%, sloping away from the point of discharge. Simply looking at large scale operations such as the Kidd Creek tailings deposit proves this notion incorrect. It is true that over a short distance (<100 m), the thickened tailings may stand up at angles close to 6% or 7% if thickened to a sufficient density. However, experience shows that over larger distances, these 94 materials settle into slopes of maximum 3%, and more typically 1-2%, as in the case of Kidd Creek (Davies, 2000). Thus, while some storage volume may be gained in comparison with the slurry tailings deposit of similar aerial dimensions, this is mostly due to the denser nature of the material rather than due to its mounding properties. Robinsky (1978) states that the principal advantage of this form of disposal is that impoundment dams are largely eliminated. He suggests the following reasons: 1. A cone-shaped hill is one of the most stable and erosion resistant structural shapes found in nature. The materials settle similarly to deltaic sediments, with slopes equal to or shallower than the material's natural angle of repose, and there is no need for confining dams. To ensure that the materials spread to the design toe or limits of the deposit, the tailings are initially controlled at minimum slurry consistency (approximately 50%), allowing the materials to flow to the design toe. As the deposit builds up, the solid-to-water ratio is increased progressively, to effectively steepen the sides of the cone while maintaining the location of its toe. 2. Segregation by sedimentation is inhibited due to the thickness of the materials and all tailings particles are deposited without sorting. "Greater than 95% of all solids remain with the deposit, eliminating the need for slimes ponds and decant systems" (Robinsky, 1978). 3. The free draining characteristic of the structure allows the tailings to naturally consolidate, increasing the density of the materials and the effective strength of the deposit, eliminating the need for containment. The escaping water is collected and treated prior to release to the environment. However, there is question of how stable a saturated material, as the thickened deposit is, will remain under dynamic loading. It is well-known that there is little risk of failure under static loading conditions, but there are differing opinions on the contentious matter of seismic loading. Eli Robinsky (1978) discussed the facts noted above, pointing out the fact that "the tailing slurry is discharged in a completely liquefied state and comes to rest at a slope determined by its angle of internal friction. By consolidation under the action of its own self-weight, the water content is reduced progressively until reaching 70 to 80 percent solids. At this water content, the angle of internal friction is expected to become at least two times as at the originally deposited 6% 95 slope...it is therefore concluded that a large reserve angle of friction is available to oppose the movement of tailing during earthquakes." Given these theoretical explanations, one would expect the deposit to be relatively safe when subjected to dynamic loading conditions. However, there are two reasons this explanation should not be taken at face value. One, as noted earlier, the materials do not stand up at the angles indicated. Two, experience indicates that the numerous layers of the deposit do not drain completely during deposition. In fact, Robinsky himself states that the deposits remain saturated, holding porewater in tension through capillary action, theoretically up to 20 meters in thickness above the water table (1998). It is well known that saturated fine-grained, relatively unconsolidated materials liquefy in the event of such dynamic conditions. Also, using natural geological occurrences as a comparative tool, it can be argued that the "liquefaction of natural slopes indicate that flow slides can develop on slopes as low as a few percent under moderate and high levels of seismic activity" Steven Vick (1983). The question of the material's stability under dynamic loadings suggests that the deposit should, in fact, be fully contained as in the case of the slurried tailings deposit. Of course, the size of the embankment is smaller than required for a slurried tailings deposit of the same tonnage due to the absence of the large volume of slurry water, reducing its construction costs, and this option would eliminate the need for a large tailings embankment in areas of low seismic risk. There are other advantages and disadvantages to using such a system that have not been discussed. Advantages: 1. Return-water pumping is reduced and smaller diameter pipelines may be used for discharge and recycling (Robinsky, 1998). 2. Reclamation of the pile may be simplified by the flatter and more uniformly graded pile slopes, and may be carried out on an on-going basis as maximum design elevations of the deposit are reached. 3. Seepage may be reduced in comparison to tailings impoundments due to the elimination of the decant pond on the tailing surface, and the material's capacity to hold pore water in tension. Disadvantages: 96 1. While the costs of an embankment are reduced, they are replaced with higher costs for thickener construction and operation, as well as the additional costs and difficulty for pumping thicker materials. 2. A greater amount of energy is required to pump thicker substances and there is the potential of greater pipe wear to consider. 3. Also, run-off handling requirements must be carefully considered. All flood runoff must be completely diverted else erosion and transport of the tailings at the toe could result. For the reasons stated above, the method is best suited to disposal sites located on relatively flat topography, where concentrated runoff does not occur, and in areas of low seismic risk. A.1.3 FILTERED TAILINGS DISPOSAL The filtered tailings disposal option is one where the tailings are first thickened, then pressed or filtered to bring the materials to a 'dry' state, of a limited water content that is typically based on the materials' ability to hold water. For example, a clayey material will hold more water due to hydrostatic forces developed between clay minerals and water. Also, the finer the material, the more water will be held in tension within the pore volume of the compressed tailing materials. Once water is removed from the tailings, they may be either conveyed or trucked to the disposal site and deposited within an impoundment or spread over low-lying slopes or in depressions. The materials may also be compacted if of trafficable quality for compacting equipment to operate on. An advantage to this system is that materials may be stacked, and sloped to safe degrees without the need for a retaining structure, and the deposit may be reclaimed in various stages. That is, once a section of the deposit has reached its maximum dimensions, reclamation activities may be carried out while deposition of further materials occur in another segment of the total deposit. A.2 SUB-AQUEOUS DISPOSAL Another form of disposal that may be used in lieu of the tailings impoundment is to submerge materials in natural bodies of water. This is especially warranted in the case of sulphide-rich tailing materials. The disposal of such tailings is an issue of environmental concern because of their potential to generate acidic rock drainage through tailings' oxidation. By placing these 97 materials under water, the exposure of reactive sulphide minerals to oxygen, an action that allows the minerals to initiate acid production, is eliminated. The oxidation and dissolution processes that generate "acidic rock drainage" (ARD) are described in the following section. A.2.1 ACID GENERATION PROCESSES When tailings are classified as potentially acid generating (PAG) in the short or long term, they are classified as such because they contain reactive sulphide minerals that, on exposure to oxygen and water, will break down to produce acid through oxidation and dissolution processes. One of the most documented (Lawrence, 1996; Ritchie, 1994; Nicholson, 1994) and significant reaction sequences occurring at exposed disposal sites involves an abundant sulphide mineral pyrite, FeS2, which is found all over the world, and is associated with many metalliferous ores. The initial reaction with oxygen and water occurs at circum-neutral pH to form ferrous ions and hydrogen ions. This is followed by the oxidation of ferrous ions to the ferric form: FeS2 + 7/2 0 2 + H 2 0 => Fe 2 + + 2 S042" + 2 H + -- decreasing pH 2 Fe 2 + + Vi 0 2 + 2 H + ^ Fe 3 + + H 2 0 For as long as the pH remains greater than three, iron precipitates will form through the following reaction: Fe 3 + + 3H 2 0 <=> Fe(OH)3 (s) + 3 H + --decreasing pH The pH continues to drop through these reactions until it becomes equal to three and lower (pending the continued presence of oxygen and water), at which point the precipitate-forming reaction reverses. Additional to this phenomenon is the development of chemolithotrophic bacteria, such as Thiobacillus ferrooxidans, in such environments. Since the oxidation reactions are exothermic, the increasing temperature tends to increase the rate of bacterial oxidation, as the optimum temperature for growth of these bacteria is around 35 degrees Celsius (Gould et al., 1994). These bacteria act to increase the rate of oxidation of sulphide minerals, their role becoming more dominant as the pH drops. Gould et al. (1994) discuss in detail the various microorganisms found in tailings, and their optimal environments of activity, as well as the actual reactions that take place due to their presence. 98 When the ferric ion is formed in such abundance, it will act as a catalyst in the following reaction with pyrite, eliminating the need of oxygen for the oxidative reaction: FeS2 + 14 Fe 3 + + 8 H 2 0 => 15 Fe 2 + + 2 S042" + 16 H + At this point, the generation of acid is no longer dependent on the presence of oxygen, but on the ferric ion, and acid development continues until the availability of reactive sulphide minerals have been depleted. One should also note that this form of sulphide oxidation produces eight times the amount of acid from the same amount of pyrite, as produced by the first reaction requiring oxygen. "Associated with the development of acidic environments is the solubilization of heavy metals such as copper, zinc, lead, arsenic, antimony, nickel and cobalt together with cations such as aluminum, iron, silica, potassium, calcium and magnesium associated with carbonate and silicate minerals. Many of the reactions of acidic drainage with other minerals in the waste are important for the neutralization of acid to maintain neutral pH conditions. The balance between the quantities of acid producing and neutralizing minerals is, therefore, critical in determining whether the final drainage or seepage water will be acceptable" (Lawrence, 1996). However, by simply removing the potential for the oxidation of the tailings prior to their contact with water, the processes of acid generation and metal solubilization may be eliminated. A.2.2 DESIGN CONSIDERATIONS "No other disposal method generates more public concern more rapidly and intensely, and mines using offshore disposal can usually anticipate that control of their tailings disposal, and therefore their entire operation, will ultimately reside in the political arena" (Vick, 1983). As such, deposition of this sort typically occurs on sites where there are major costs, and land use, topographic, population or other environmental restrictions or problems that exist for tailings disposal on land (Brawner, 1978). Given these comments, Ripley, et al. (1996) have noted that placing the materials underwater in natural depressions or lake basins appears to be the safest way to dispose of tailings. In coastal areas where the combined effects of extremely high precipitation and possible flood events, steep 99 terrain, and high seismicity make surface impoundments impossible to safely design and construct, this is the most viable option. This is also the case for tailing solids that would be reactive to weathering processes; the water cover prevents oxygen access, maintaining the tailings' physical and chemical stability, and prevents metal contamination of surface and ground waters which might otherwise result from on-land placement (Poling, 1999). Although the method has been used successfully at several mines in British Columbia, the Philippines and Central America, as well as other places around the world, neither are all tailing materials nor are all sites suitable for sub-aqueous tailing disposal. C O . Brawner (1978) suggested several possible settings that could be and have been used for sub-aqueous disposal of tailings. These are listed as follows: 1. Deposition in a lake bay with a dam constructed across the mouth of the bay to confine the deposition, as applied at Granisle Copper Ltd., B.C. Tailings are spigotted from the top of the dam, so the upstream slope of the dam can be rendered highly impervious. As for any embankment structure, the foundation stability and settlement under the dam must be evaluated in detail. Its body should be constructed with waste rock under water to protect against erosion and an upstream filter layer is required to prevent piping. 2. Deposition in one or more deep lake(s) located in close proximity with the mine. Tailings are transported to the lake and deposited via pipeline. This method was applied on Vancouver Island at Buttle Lake, B. C. by Western Mines Ltd. 3. Fill an existing lake with tailings and recreate a new lake in the open pit when it is abandoned. This may be more applicable to sites containing many neighbouring lakes, rather than only one, where environmental pressures would likely reject the suggestion. 4. Disposal in deep, ocean bays where inlets exist near the mine property. In Rupert Inlet near Port Hardy, B.C., Utah Mines Ltd. successfully used this deposition program. The salt water seemingly tends to assist in causing the tailings to flocculate and settle. 100 5. Disposal in shallow seacoast bays where large bays exist along the seacoast. Causeways, behind which tailings may be deposited, would require riprap protection. 6. Disposal in deep, sea water along the coast, where it is practical to carry the tailings in a pipeline or on a causeway to be deposited in deeper water. At Marcopper and Atlas Consolidate Mining Company, in the Philippines, tailings slide down causeway slopes to deeper water, and are piped into deep water where they slide to even greater depths. Prior to disposal in either a submarine or lacustrine setting, intensive environmental studies must be carried out, the results of which must indicate the potential aquatic damage as being negligible, or limited and temporary. Besides this, "a series of specific criteria must be met to ensure that the system will be reliable and provide minimal environmental impact" (Poling, 1999). The "effects on water quality may be limited if the chemical composition of the mill effluent is relatively innocuous" (Vick, 1983) and if the physical properties of the tailings are such that the materials "will flow as a density current away from the discharge pipe to become a stable sediment deposit" at depth (Poling, 1999). The most obvious effects of tailings disposal into aquatic environments have been sedimentation and turbidity. "Upon discharge to the hydrosphere, substances from both anthropogenic and natural sources may be acted on by diffusive and mass-flow dispersal processes. Water bodies contain currents due to the inflows and outflows of water, to the flow around rough objects, and to wave-action produced by atmospheric winds. Temperature and density changes generally ensure a turnover of lake water in spring and fall, whereas Arctic lakes and ponds may not thaw completely" (Ripley, et al., 1996). For these reasons, it is important that the point of discharge be sufficiently deep and far from the shoreline to avoid the most biologically productive and sensitive shallow-water and near-shore zones. There should also be no risk of the transport of tailings solids by current or wave action, 101 or turbidity in the upper (euphotic) zone of the water column where sensitive habitats exist. Turbidity effects reduce the intensity of incoming light, which can ultimately damage the biological productivity in the area (Ripley, et al., 1996). Additionally, there should be no risk of metals or other soluble materials diffusing into the water column, and potentially damaging other sensitive biological processes. Some final concerns to address when choosing such a form of disposal include the expected life of the project, and transport systems. The site for the tailings deposit must be of sufficient size to contain the entire volume over the life of the mine, and this capacity must remain within the "safe" zones described above. The pipeline system must be monitored and maintained to ensure that joints and pipe wear do not cause problems with leakage; and the lines must be secure against accidental rupture on shore and throughout the sub-aqueous route. A.2.3 ENVIRONMENTAL STUDIES Although disposal of tailings into marine and lacustrine environments has occurred for a long time, it is only within the last few decades that information regarding the environmental impact of this disposal method has become available. As a result of various studies on the post-depositional behaviour of submerged tailings in several Canadian lakes and fjords, precise and accurate data on submerged tailings reactivity has become available. Data was gathered using a variety of modern sampling and analytical techniques for both the water column and sediments of aquatic systems heavily impacted by ARD. So far, the data "provides convincing evidence that the reactivity of sulphide-rich mine tailings is limited by permanent underwater storage by reducing the influx of oxygen, thereby limiting the rate of acid generating reactions associated with sulphide oxidation" (Mugo, et al., 1997). The results of other general studies also indicate minimal biological and water quality consequences, and suggest quick rehabilitation after discharge ceases. An example of rapid recovery is where tailings were discharged into Rupert Inlet near Port Hardy, B.C. "Detailed examination of this site by regulators, scientists and engineers has led to the nearly unanimous conclusion that this submarine tailing placement had minimal impact on the marine ecosystem...Nearly 2V2 years after cessation of tailings deposition, the tailing sediments are rehabilitated to the same density and the same ranges of biodiversity of benthic organisms as existed in pre-mining days on 'natural' sediments" (Poling, 1999). 102 In another case regarding long-term stability, concentrations of key metals such as nickel, cobalt, copper and zinc were all markedly lower (and in some cases near or below the detection limit) in tailings interstitial waters compared to the water column. "These results suggest that the submerged tailings are undergoing only minimal reactivity and that metals are not leaching into pore waters. As a result, the tailings are unlikely to support a diffusive efflux of metals into the overlying water column" (Mugo, et al., 1997). While these results are promising, "the reactivity of tailings under water is governed by complex interactions between the geochemical characteristics of tailings and the physical, biological and chemical properties of the receiving water body" (Mugo, et al, 1997). As a result, predicting the behaviour of tailings in contact with the receiving water is not simple, and each case must be analyzed both carefully and individually. Factors applying to one situation may not necessarily apply to another. A.3 UNDERGROUND DISPOSAL C O . Brawner (1978) once stated that "one potential area [of disposal] for consideration is subsurface disposal, subsurface has been interpreted in its broadest sense meaning, below the original surface whether it be ground or water, or within stable waste material. Several alternatives bear consideration and every site will be different; therefore each must be assessed in its own setting." A.3.1 BRIEFING A.3.1.1 History Throughout history, mines have been dealing with the problem of where to put mining wastes. For some mines, limited space governed mining activities and rate of extraction, and so underground disposal operations began with the placement of unconsolidated wastes back into mined out areas as space became available. However, unconsolidated rock fill has a limited ground support capability and the freestanding height capacity is negligible, limiting its use to filling stopes that will not be exposed in future pillar recovery operations. Today, its use is generally limited to void filling and to provide some measure of passive wall support for resisting localized ground movement (Hassani and Archibald, 1998). 103 The first application of hydraulic filling (uncemented) is believed to have been used to a control ground subsidence problem, to protect the foundations of a church in Pennsylvania in 1864. In 1909, the use of hydraulic fill was initiated in South Africa, in the coalmines of Germany and various gold mines in Australia. Various other mines followed suite shortly thereafter, including some that had been using unstabilized rock fills, and the practice became standard at many mines worldwide after WW2 (Dickhout, 1973). Hydraulic fills were very practical in that they were easy to place, they fit 'tight' in the stopes and the resulting insitu densities of the fills were desirable (Chan and Hassan, 1988). Also, the method provided an environmentally acceptable method of disposing of mill tailings as they were the main component of the fill. The stabilization of fill by the addition of Portland cement and other binding materials did not become standard until the 1960's (Dickhout, 1973). "Cemented fills were a great improvement over uncemented fills, as they could stand up vertically and be used as a mucking floor, without causing excessive dilution" (Chan and Hassan, 1988). By 1973, cemented hydraulic fill was the most common type of fill being used in the mining industry. Numerous studies were being carried out to determine optimum conditions that would result in maximum strength values at the most economical cost. Many patterns were noted, including the effects of additional cement, the effect of cement replacement with other binders, the effect of grain size distribution on permeability and drainage characteristics of the materials, and the effects of segregation and mode of placement used on the fill. Stemming from these studies were the applications of mixed hydraulic fill and rock or aggregate fill, later followed by the application of paste fill technology, "first used at the Grund Mine in Germany during the 1980's" (Hassani and Archibald, 1998). Paste fill allows the placement of all tailing materials underground, not only the coarser fragments as in hydraulic fill development. However, the use of paste fill requires certain modes of mine development, as the paste requires longer cure times, and the materials must be pumped underground, as in the case of hydraulic fill. Because of these limitations, this application cannot be used in all underground mines. Another new technology, not fully explored, is the addition of agglomerated tailings into aggregate or rock fill. Not only does this method allow the placement of tailing materials underground, but the resulting fill materials are also more manageable. The agglomerated material is easier to handle than filtered or thickened tailings, and may be handled similar to any other aggregate material, allowing fill to be mixed in batches at low water contents, and 104 transported by belt or truck to the area of placement. There are no drainage requirements for the fill, and it cures in a shorter time than does paste fill. A.3.1.2 Functions of Backfill Throughout history, fill was not primarily used for subsidence control, although this and fire control are important applications. Noted as most important was the wall support it provided, including stabilization of the operating stope and adjoining pillar, and access openings and boundary pillars in small and medium mines, followed by working floor and void filling, then tailings disposal (Dickhout, 1973). Today, the many parameters of fill are well understood, and the strength and responsive characteristics of the fill may be designed to suite different applications in underground mines. The requirements and response of fill then change with its intended function. Excluding waste disposal, there are three main functions for which backfill is employed (Hassani and Archibald, 1998): 1. Provision of ground support 2. Allowance of pillar recovery 3. Provision of a working floor Ground Support "Compared to the rock mass that surrounds it after emplacement, backfill is relatively soft and does not provide significant tributary support for the rock mass. Its main stabilizing effect is to impart increased lateral confinement pressure onto the rock walls or pillars that support the rock mass' load" (Hassani and Archibald, 1998). As a structural component, however, backfill may prevent large-scale movements and collapse of openings, or accept some part of the load previously carried by the ore. The fill responds by deforming under the imposed back and wall loads, which can lead to excessive deformations if the fill is not stiff enough. Problems that have been associated with this function include liquefaction and heaving. It is important that fill placed in a saturated condition not be permitted to remain this way as the materials may liquefy in response to dynamic loading caused by blasting or rock bursts. This could impose high hydrostatic pressures in bulkheads, resulting in failure and mud-runs, and loss 105 of support. It is also conceivable that fill imposed to high lateral stresses and low vertical stresses could fail by heaving at the surface due to insufficient strength (Aitchison, et al., 1973). Ore Recovery Rock pillars are typically left in place to provide support for various underground mining methods. The trend of extracting pillars surrounded by fill for maximum recovery of ore requires a redistribution of the load carried by that pillar into the adjacent fill, therefore the fill must be strong enough to withstand these imposed loads. The removal of ore by blasting adjacent to the fill also requires it to have high tensile strength, and the ability to withstand the stress cycles induced by the firing. If the fill has low stiffness, unloading of the pillar may be difficult due to potentially high deformations in the fill. Failure of the fill also results in dilution of the ore, and economic losses, as well as the transfer of load away from the fill which could be dangerous for the underground operations (Aitchison, et al, 1973). Working Platform In overhand cut and fill mining, the fill provides support for men, drilling equipment, the ore pile and mucking vehicles. The fill must be able to sustain wheel loads from mucking vehicles, a number of load repetitions, and the action of scoops and scrapers, which could loosen the surface, causing resistance to vehicle motion or settlement. The principle mode of failure of this function is inadequate bearing capacity under the effect of surface loading. Ore fired from the back of the stope onto the surface of the fill will impose impact loads, possibly resulting in penetration of the fill; ore piles may settle into the fill under its own weight over a period of time, resulting in dilution and economic losses for the mine (Aitchison, et al, 1973). Waste Disposal The open spaces left by underground mining may be filled with waste materials, if for no other reason than simply to dispose of it. The use of waste materials in fill has been common practice, but unfortunately there always remained additional materials, predominantly the fines or slimes, that had to be exposed of in a surface impoundment or sub-aqueously. This was mainly due to the types of fill being used, predominantly hydraulic and rock fills, but there also remains another reason. Improved comminution processes used in milling stages cause the mined 106 materials to expand to many times their original volume (not as drastic in the past), and so typically there is insufficient space underground to store all of the waste produced. With the advent of paste fill and agglomerated tailings incorporated into fill, this need is reduced or eliminated for cases where sufficient material is removed from the underground operation in which to place all of the waste materials. In mines where there are ancient workings to fill, or an existing open pit space in which the materials may be disposed, there is also no problem. However, most underground mines still have a need for other modes of disposal of tailing wastes than within fill alone. A.3.2 STABILIZED MINE BACKFILL Stabilized mine backfills are similar to weak concrete materials and they consist of an aggregate such as waste rock, mill tailings or sand, a stabilizing agent such as Portland cement or alternatives, and water. The water in the mix reacts with the stabilizing agent to bond the aggregate material, giving the fill cohesion, and an increased internal angle of friction. Until recently, there were only two types: rockfill - crushed rock of size less than 15cm and very little sand or silt material; and hydraulic fill with an aggregate generally smaller in size than a coarse sand with most of the material in the silt to fine sand range (0.01 to 0.3mm). Each has its own advantages and disadvantages, discussed below. A.3.2.1 Cemented Hydraulic Fill Cemented hydraulic fills (CHF) are manufactured with tailings, sand and/or rock materials, binding agents, and water. Materials are typically placed at a pulp density of less than 70% by weight. The predominant advantages to using such a system include the minimum requirement of technical supervision, the ease of installing and operating the infrastructure, and the resulting good quality control and mixture density. Preparation of the slurry can be performed on surface or underground. Simple desliming procedures may be developed, and pumping may be avoided if the plant is positioned properly to allow for gravity flow of the slurry to the disposal site. For a long while, cemented hydraulic fills were very popular. The cost of transporting and placing the fills underground were low, and the aggregate was readily available in the form of mill tailings. However, the hydraulic transport method required that excessive amounts of water be added to the fill, increasing the water to cement ratio and reducing the effectiveness of the cement. Also, the large amount of water caused segregation of the cement and aggregate 107 particles, lowering the insitu density of the placed material and its internal angle of friction. Finally, expensive supports were needed to contain the fluid materials within a stope and extensive drainage systems were required to recover the seepage escaping from the slurry during its placement and curing stages. A.3.2.2 Cemented Rock Fi l l Cemented rock fill (CRF) is comprised of sized or unsized aggregate mixed with various types and amounts of binder materials. Waste rock is dumped into raises then distributed by trucks or conveyors to the stopes where the material will be placed. Typically, the rock aggregate is pre-mixed with cement binder, usually in slurry form, prior to entry into the stope. Alternatively, rock aggregate materials may be post-consolidated by passing a cement slurry mixture onto and through emplaced but initially unconsolidated rockfill (Hassani and Archibald, 1998). Stabilized rockfill has the advantage that it is generally stronger than stabilized hydraulic fill; strengths can be "up to two or three times as high as cemented hydraulic fill" when produced at equivalent binder contents (Hassani and Archibald, 1998). A rockfill of adequate strength can be developed with a lower cement cost than a hydraulic fill of similar strength. Such a fill will provide ground support where undermining will occur and will improve wall rock strength when exposure of fill walls is expected. 'The fill has a capacity to stand over exposure heights that are not economically possible with other types of fill" (Hassani and Archibald, 1998). Other advantages include reduced surface disposal, and a simple preparation system. There are limited drainage problems - dewatering can be avoided altogether depending on the mix proportions, and high fill quality can be achieved if materials are placed properly. Of course, there is a requirement of void filling by fines and/or cement to ensure competence; rockfill coning on placement results in segregation of coarser materials to the stope sides, reducing the potential to tightly fill stopes. Another problem surfaces when large amounts of waste rock are not available at the mine site; such sites require crushing operations for quarried rock. This leads to high transportation, surface production and haulage costs. Other economic considerations include the greater cost to develop a waste pass system to transport the rock underground than it is to install piping for hydraulic fills. These passes can become blocked, which in turn leads to time and expense to 108 clear the chutes. Finally, actual placement of the fill in the stope is often more expensive due to the labour intensive haulage methods used to place the fill. A.3.2.3 Combined Cemented Rock and Hydraulic Backfills By combining rockfill (RF) and cemented hydraulic fill (CHF), both the quality of the rockfill and the strength of cemented hydraulic fill may be improved. Rockfill may be introduced to the stope through chutes, or left in place as ore is mined out. The fill is consolidated when a slurry mixture of sand waste, binder and water are introduced into the stope, either simultaneously with the rock fill, or separately in the latter case where waste is left in place during excavation. The combined RF and CHF mixture results in complementary advantages and reduced disadvantages of the two systems. The voids within the rockfill become filled with fines, "enhancing the stability for gravity loading and blast vibration resistance during excavation of adjacent stopes or pillars" (Hassani and Archibald, 1998). The mixture typically has relatively good mobility, demonstrates less segregation potential than CRF, and is generally denser than CRF. The dewatering requirements are intermediate between the CRF and CHF systems, but can be limited by controlling the slurry pulp density. However, bulkhead control of this material is still essential due to the slurry content of the fill, and the path of the slurry flow is difficult, which can lead to zones of higher and lower strength fill. Economically, this system can lie anywhere in between, above or below the costs of the CRF and CHF systems. This is dependent on the method of mixing chosen for the mixture. If waste rock materials are left in place underground, then there is only the cost of piping the slurry mixture to the stope, or forming drop chutes to the stope of placement. However, if the rock is mixed with the slurry prior to placement in the stope, resulting in a better quality fill, then there are the considerations of larger batching equipment, more wear, larger pipes / chutes and the risk of blockage, and additional hauling costs of the rock to the mixing area. A.3.2.4 Behaviour of the Fill Many studies have been carried out on cemented hydraulic and rock fills in order to improve their strengths, and reduce costs associated with loss of strength due to segregation and bleeding (dilution of neighboring ore and drainage systems), and Portland cement. In the Proceedings of 109 the Jubilee Symposium on Mine Filling at Mount Isa Mines Ltd., a series of test results found through various studies are discussed. A summary of typical hydraulic fill behaviour and the predominant factors affecting its strength is provided below. The most obvious findings are that a fill's unconfined compressive strength increases with cement addition, age, increased pulp density and decreased moisture content. It was also noted by McCreedy and Hall in 1966 and Thomas (1971; 1979) that the permeability, moisture content and void ratio of the fill decreases while its internal angle of friction and cohesion increases with increased cement content and as curing progresses. However, results from triaxial tests showed that strengths increased not only with cement addition, but also with confining pressure. Also, Thomas (1979) notes that the addition of fines is seen to have a more marked effect in increasing friction angle than does cement addition. Corson determined that finer fill materials exhibited greater cohesion than well-graded mixes, while well-graded fills had a larger internal angle of friction (at the same cement to sand ratio). On a similar note, Thomas, 1971, found that fill strength decreases with increased void ratio, and increases with increased fines addition. This addition of fines, however, does not account for the 'slimes' or silt and clay sized fraction of a tailings material. "The presence of slimes has a detrimental effect on stiffness and strength of [hydraulic] fills, because they are more compressible than sand or silt, and leads to greater moisture retention because of the nature of clay and the reduction in permeability. High moisture content can lead to excess pore pressures and reduced effective normal stresses, and hence reduced strength under the effect of static or dynamic loading" (Aitchison, 1973). In 1966, McCreedy and Hall, and Toguri and Cerigo discovered that using various other Portland cements did not seem to produce significantly differing results than those achieved using the conventional Type 1 cement. Aylmer, 1973, proposed that the finer the grind of the cement, the faster the hydration reaction and strength development will occur, but he found that increasing the fineness of the grind is not beneficial to strength development in dilute cement-water mixtures. This is in agreement with studies carried out by Luka and Weaver, 1970, which also showed that varying the fineness of the cement does not produce great changes in fill strength. Test results presented by McCreedy and Hall suggested that the percolation of acid mine water through hydraulic fill does not affect its strength, and the use of acid mine water to transport the CHF does not seem to affect its ultimate strength. However, Aylmer, 1973, determined that if 110 the pH of the water were below about 5.8, the cement bond would be leached. He also suggested that the presence of dolomite or limestone either naturally or artificially in the fill will protect the cement bond, as will other pozzolanic materials. As discussed in the main body of the thesis, other materials exhibiting cementitious properties may also be used in partial replacement of Portland cement to reduce binder costs. Alternative agents such as natural pozzolans, blast furnace slags, and fly ash have been studied for several years. Studies show that "the reactivity of natural pozzolans is quite low. It can take several weeks before fill...exhibits adequate strength" (Chan and Hassan, 1988). Blast furnace slag acting as a pozzolan often increases the fill strength in proportion to the amount added, but this is strongly dependent on the physical condition of the slag; "the slag must be granulated, and in a metastable, non-crystalline, glassy condition" (Thomas, 1979). Fly ash can have various compositions, and only some are suitable for cement substitution. This is possibly why Luka and Weaver (1970) found that "the addition of fly ash to economize is not particularly attractive," while other studies have shown positive results in utilizing fly ash as a cement replacement. Finally, the addition of some flocculants have had beneficial effects on strength, percolation rates and in reducing losses in drainage water, while the inclusion of minor dispersant yielded a significant increase in bearing strength of the leaner sand-cement mixes. All tailing materials are different and application of such chemical constituents must be tested on each individually to know the full extent of the resulting effect on the fill. Overall, the strongest and densest mix is provided by grading and blending, resulting in increased stiffness and frictional shear strength, and may also improve the permeability of the fill. The addition of cement improves the cohesiveness, provides tensile strength and increases the stiffness of the fill (Aitchison, et al., 1973). However, both of these activities result in increased filling costs. In the field, the economic availability of the fill takes precedent over the optimization of the fill quality. As a result, most underground fill operations have concentrated on reducing costs by utilizing pozzolanic materials in place of cement wherever possible, and by decreasing the pulp density of the fill material. This activity led to the advent of high-density paste fill. I l l A.3.3 HIGH DENSITY PASTE FILL "Paste backfill is defined as an engineered mixture of fine solid particles (with or without binder) and water, containing between 72% and 85% solids by weight" (Lang, 1995). It has the appearance of 'toothpaste' due to its high pulp density and its utilization of total tailings, that is, the coarse and slimes fractions, and the mixture may also incorporate sand or waste rock for strength enhancement (Hassani and Archibald, 1998). A.3.3.1 Pros and Cons Paste fill may be prepared on and transported from the surface as "particles in a paste mixture will not settle out of the mixture if allowed to remain stationary in a tank or pipeline," and the fill "can be placed in stopes with or without binder addition depending on the strength requirements of the fill" (Lang, 1995). Some of the other advantages of the system listed by Brackebusch, 1994, Lang, 1995, and Hassani and Archibald, 1998, are as follows: 1. Higher strengths can be achieved than in hydraulic fill with equivalent cement content. 2. Shorter stope cycle times result because strength can be achieved in a shorter time with paste fill than with hydraulic fill. 3. Unclassified tailings can be used rather than just the coarse fraction as is the case for hydraulic fill, reducing the need for surface disposal. 4. Drainage of water and slimes from the fill are minimized because there is less transport water, and the water that is used is consumed by cement hydration. This reduces the need for bulkhead construction and extensive drainage works. 5. Stopes may be filled continuously until complete without worrying about liquefaction and washout of a sand barricade as in the placement of hydraulic fill. 6. Underground storage capacity is increased because paste fill systems achieve lower porosities than conventional fill. 7. Since paste fill is deposited as a non-segregated mass of fill (because cement particles are not displaced by the internal movements of the draining water), more predictable strength properties for the fill can be achieved. 112 8. A paste backfill system allows the use of mechanized undercut-and-fill mining systems, so long as the fill strength is sufficient, with nearly any ore body shape. Disadvantages of using paste fill: 1. The pumpability of a paste is very sensitive to small changes in water content and grain size distribution, bringing about the need for greater quality control. 2. Paste backfill systems typically have higher capital costs compared to conventional hydraulic backfill plants due to the need for superior dewatering facilities and greater technical precision. However, the costs are roughly similar to the investment required of rock fill systems. 3. The distribution network in the mine requires a greater level of engineering design to control pipeline pressures. A. 3.3.2 Characteristics of Paste Fill B. Lang, 1995, indicates that the key characteristics of tailings or other materials being assessed for suitability as paste backfill are the dewatering characteristics of the material, the pumpability and strength of the fill mixture and its bulk density. 'The rate of dewatering is a function of the specifications of the de-watering unit, as well as the characteristics of the material (including grain size distribution and specific gravity). The pumpability of the paste is dependent on the viscosity of the tailings [which is, in turn, dependent on the fill's pulp density, and fines and binder contents] as well as the type of pump employed and the geometry of the distribution system" (Lang, 1995). The topic of paste fill rheology, or viscous characterization, is discussed in detail by Hassani and Archibald in the Mine Backfill CD Rom developed for CIM, 1998. In summary, "paste will typically flow under gravity at an angle greater than 30 degrees" (Lang, 1995). However, when comparing to the concrete industry, one can expect that the paste fill's abrasive characteristics are likely to be much higher and the mixture will have a higher resistance to flow due to the fact that the majority of the fines are made up of angular tailings rather than cement. Material preparation of paste fill is far more complex than concrete and the fill is 113 transported over much greater distances, resulting in higher and larger ranges in estimates of pressure losses in the piping system (Hassani and Archibald, 1998). As in the case of hydraulic fills, the strength of paste fill is influenced by many parameters including pulp density and therefore specific gravity and grain size distribution, binder content and type, curing time and curing temperature. The insitu density of paste backfill is higher than conventional hydraulic fill due to the reduced water content and increased fines content of the mixture. However, the pulp density and grain size distribution of the paste is typically controlled by its rheological requirements for reduction of pipeline pressures rather than strength (Hassani and Archibald, 1998). Also, the advantages of paste fill, such as low voids ratio, can be compromised if the method of introducing the fill into the stope is not given careful consideration during the design phase of the operation. Binders, including cement, fly ash, blast furnace slag, and natural pozzolans, are used in paste fill only where strength is required and where resistance to liquefaction is necessary. At this point, it should be noted that the parameters above predominantly affect the short-term strength of the fill. As time passes, other factors that could detrimentally affect the fill's long-term strength become apparent. These include the placed insitu density of the fill, the pressures of the underground environment, the groundwater conditions during and after placement, and the chemical stability of the fill materials. There have been studies carried out where test results have showed a decrease in the fill's quality over time. In most instances, these fills contain high volumes of sulphides and/or the fill is in an environment where it is periodically or continuously exposed to highly acidic waters. For example, Bertrand (1998) carried out several tests where various paste mixtures (of different tailing materials) were exposed to leaching waters at varying degrees of acidity, including one non-acid bearing leach water. She showed that backfill "exposure to an aqueous leaching environment prior to being properly cured will promote the loss of binder material (more rapidly in an ARD environment)...which could be accentuated where slag-cement is used, as it reacts slower than OPC." She noted, however, that while a small decrease was noted in one of the samples, most held their strength regardless of the leaching environment. 'The layers of alteration noted were most likely too thin after this period of time to affect the UCS of the samples." What it did affect was the acid generating potential of the fill mixtures. 114 Bertrand stated that "binder provides some buffering capacity, but is far from sufficient to neutralize all the potential acidity of the tailings...acid base accounting (ABA) tests indicated that binder neutralization potential is readily consumed in the early stages of contact with water or ARD solution, while depletion is more severe in ARD solution for all cases. Therefore, even if cement was added to raise the neutralization potential ratio of the backfill, the NP may be depleted at a very early stage." During testing, for example, oxidation of the sulphide portion of the tailings was observed after only 14 days on one sample containing the greatest amount of sulphides and the greatest amount of cement. Therefore, for chemical stability, pastes containing high proportions of sulphide minerals would likely require complete submergence in water at very early stages. Further to the risks of ARD, Bertrand's study "suggests that in a flooded environment, where hydraulic conductivity of the saturated material is low and the backfill mixtures contain relatively high proportions of Portland cement, secondary expansive minerals may form." The sulphates present in the interstitial water and those produced by the oxidation of pyrite, in a basic medium will react with the free calcium ions produced by the dissolution of unstable portlandite hydrates Ca(OH)2 resulting in the precipitation of swelling secondary gypsum CaSO^lHjO and very expansive ettringite 3CaS04-3CaO-Al2Or32H20. These expansive minerals can generate very high crystallization pressures (70 to 200 MPa), affecting the durability of mortars. It provokes the fissuration of the concrete mass and consequently the loss of its structural strength. (Ouellet, 1998) In support of this theory, Ouellet, et al. (1998), investigated a fill where high porosity was evident and sulphuric grains were well dispersed in the material. They found that the grains were well coated by the cement, but the texture remained loose because of the structure adopted by the cement during hydration. Also noted was "the presence of a fine system of fractures in the samples accompanied by oxidation traces typical of a chemical alteration. The presence of sulphides in the tailings caused a dissolution of the calcic phases of the cement hydrates and promoted the formation of swelling phases, which in turn induced a destruction of the cemented backfill." 115 For these reasons, the long-term characteristics of backfill materials must be considered and in order to have mechanical stability in the long term, the chemical stability of the material must be insured, as chemical alterations of the cement phases can have a negative impact on the strength of the fill. A.3.4 AGGLOMERATED TAILING FILL During the past decade, a series of testing surfaced where tailing materials were agglomerated for their use as aggregate. The main objective of the agglomeration process is to reduce or eliminate the need for surface disposal of slimes. In the past decade, it was proposed that agglomerated tailings be used as supplementary additives to paste fill mixtures in an underground environment or as material for surface land cover, gravel or as building materials (Amaratunga, 1991). Just as in paste technology, the process of agglomeration precludes the use of total tailings, or the coarse and slimes fractions. Because the use of excessive fines in hydraulic backfill decreases percolation rates and increases curing times, slimes are typically disposed of in tailings ponds. This is the case even in paste fill operations where not all fines can be used. Also, supplementary materials such as gravel, alluvial sand or rock chips are used to step up the quantity of fill for underground placement, as well for strength purposes. By agglomerating tailing materials, the problems associated with handling, curing and drainage are reduced, and they provide a readily available, cheap aggregate source at the mine site. A.3.4.1 Process The process of agglomeration of mill tailings may be carried out in a number of ways, including rotating devises such as disc or drum pelletizers, or pressure and mechanical methods for briquetting, tableting, extrusion, rolling and compaction. Amaratunga (1991) proposed a technique commonly referred to as cold-bond tailings agglomeration (CBTA). The concept is that mill tailings initially be classified and then the fine materials (slimes) dewatered to form a semi-moist feed material. The filtered material is then mixed with binders prior to entering a disc or drum pelletizer used for agglomeration of the fine tailings. Following agglomeration, the pellets are cured under high humidity and stored for 3 days for strength development. The cured products are mixed with the coarse fraction of the tailings in a predetermined ratio with a cementitious binder and water to develop high quality backfill. It has been shown that this 116 process produces a strong underground fill, in fact, "significantly enhancing the strength and elastic modulus over a non-aggregate total tailings paste fill" (Amaratunga and Hein, 1998). A.3.4.2 Characteristics of Agglomerate Tailings Paste Fill Agglomerate tailings paste fill (ATPF) is of similar character to paste fill where supplementary aggregate material is added. Some of Amaratunga's (1991, 1997, 1998) findings include: 1. Increased aggregate addition leads to an increase in compressive strength (up to 200%) and the elastic modulus over a total tailings paste fill without agglomerates. 2. The compressive strength of ATPF is higher than similar fills incorporating sand and gravel. 3. Increased binder addition leads to increased strength. 4. Using fly ash as a partial substitute for Portland cement proved successful at later curing stages; strengths were lower than OPC mixtures at early curing stages. This method of utilizing all tailings in paste fill, by incorporating the fines as agglomerates in a coarser fraction, is a novel idea and has already been used successively in the field. However, not all mines have the ability to utilize a paste fill system. For this reason, a new proposal was made to incorporate total tailings agglomerates into aggregate fill systems. In this type of system, the agglomerates replace a portion of the aggregate typically used in the fill. In addition to placing the tailings underground and utilizing their strength, costs of acquiring potentially off-site aggregate are reduced. A.3.4.3 Agglomerated Tailing Aggregate Fill For the Eskay Creek project, the option of utilizing a pastefill is not an option both for economical reasons and due to the method of mining used at the site. However, the underground operation has the capacity to place all of their tailings within the excavated stopes, due to the volume of materials excavated and shipped directly as concentrate. For this reason, the company wanted to investigate a means of placing the tailings underground. In this thesis, the process of agglomeration was utilized in order to improve the handling characteristics of the total tailings material. It was thought that if the tailings materials could be 117 made more manageable, they could be utilized in the underground fill system as a source of strength. This would reduce the need for finding offsite sources of aggregate, and would also provide a means of disposal for the tailings in an environmentally friendly way. It was expected that the results of the testing program would encompass many results found in the studies of other aggregate fills. Also, although no long term studies have been documented, it is expected that similar factors affecting the fill strengths noted in the previous sections also apply here, particularly those noted for paste fill, as the matrix materials of the two fill types are of similar character. 118 APPENDIX B MATERIAL PROPERTIES Grain Size Distribution of Materials - As Received 1000000 100000 10000 1000 100 10 1 0.1 Particle Size (microns) Grain Size Distributions of Materials - As Used in Agglomeration and Backfill Production 0.0 |IHl III I |l 111 I I I I [Nil III I | l l l l I I I I | l l l l I I I I |l 111 I I I I | IIII II i ,i | 1000 100 10 1 0.1 0.01 0.001 0.0001 Particle Size (mm) 120 109 Tailings Grain Size Distributions - Measured by Various Methods 0.00 10000 1000 100 10 1 0.1 Particle Size (microns) NEX Tailings Grain Size Distributions - Measured by Various Methods 0.00 10000 1000 100 10 1 0.1 Particle Size (microns) 121 River Rock Aggregate Grain Size Distribution As-Received and After Removing Oversize 0.00 1000 100 10 1 0.1 0.01 Particle Size (mm) Various Binder Grain Size Distributions - from Various Sources 0.1 0.01 0.001 0.0001 0.00001 0.000001 Particle Size (mm) i 122 Q W co & CO < • CO O H P « H co Q W N CO E CO CO f| | | % 3 CO > Q 2 o E cu O X — 111 CO z i -•5. £ E S CO — co u_ E 5 * 8 E " CD o CD LO o o O o o LO o Oi CD r*- o O o o o cz Oi cri o CO c\i ass CO CD ass Q. vP E O Oi h- CM LO CO CM .,— c o o CM O o O o o o CD o o o O o o o o o Sere SizG E o o o O CD o o o o Sere SizG o o o o o £ ID £ E £ co — co LL. passing | o o CO CO CD o o m o o o o o o o o passing | O o o Cfj CO CO o> CO •0 r-~ cci CD CO c\i CO CO CVJ ci CM CM ci 1^ ci passing | cum 9ZIS 0.326| 0.23| 0.1491 0.11| 0.074| 0.04| 0.031| 0.022| 0.016| 0.012| 0.0085| 0.0063| 0.0045| 0.0031| 0.0021| 0.0012| Screen i |(mm) O) OO Oi h-Oi CO Oi cn 00 CM CM CO LO 3 cn o CM CM CO CO passii 00 CO passii cum 1- CO CO CM oo CO CO | 0.0024| CM Screen Size |(mm) o" o o o | 0.0244343( d d o o o o o o O o | 0.0024| O O 123 E E CM A X M LU t 2 & DC £ o io* E E •o, (S >» CO I CO CO CO CM f-O c O </> p Q. c 3 CO CO ±= a> co **= > O CO co — s s s s g <D D) 9 1 1 O LU CJ °> "g CL C E ii CO .-= CO LL. o CO CO LO CO o o o LO CO CO CM CO LO o o cri CO* CO* o cri d o o> cn CO CO -3-r- 'tn E w T _ 3 ra o a o CM r» CO m r*. o o o t CO i-~ co LO o ci d CO CO LO ai ci ci 1— CM CO LO o E <5 R "55 o CM LO CO CM CM CO o o cn CO cn CO LO o o CM cri C\i CO* ci cri o o LO o o o o CO o o CM cn o CM o CM -3* CO ci r~* CM CO •* CO CO LO CM CO •* a i CO O cn o CO O CM CO r-- LO co CO CM " , _ ***"> c Q O CO LO o o o o o O •<* CO o LO o Q 1— 1— CM CM N CO c 55 -e 2 a> OS ° E *5 1-CO > CO CO o o co CM CO •"3- CO CO o CO cn CO o o o cum % passing o o cri cn Tt* cn CO CO LO LO CM LO* ci o o TJ-LO cn o |-~-cn 1^ o cn o o cum % retain o ci  CO CO -3* CO d o o o CO CO CO LO CO CO o o cn CO o retai r o o "3* c\i CO* CM r»* CM o LO* o o o • * o CO cn CO o LO o LO o CO CO CM CO o CM CO CM •>3-r-* cn cn CO CO d LO r-* CO CO "5 CO CM CO o CO CM cn T o r~ CO LO r» CO o ons) o E ze 1 CO •"3- LO CO O o o LO o o CM o EM o o •"3-o o •>3-CO IScreen (mesh) Total CO is . c tfi &• •o Q. E E " I? X a> o o LO LO CO CM •>* CO cn LO cn CO o o cum % passing d o r>* CO CO vi CM CO CO* CO d CO CO LO £ CO CO cn CO o o cum % retain CO* cvi CM CO* CO CO CO CO * CO d o CO CO CM CO CO CM LO LO CM o CM CO CM ""Z retai r CO CO 5- CO* cri LO* LO CO* o CO o CO o LO o CM o o cn cn cn CO C\i cri LO CM CO CO cri CM CD* CO* cn CM CO •g CO CM CO o CO CM cn tt o CO LO co o (sua o E ze CO •>3- LO CO o o o LO o o CM o CM o o -400 CO Screen (mesh) Total 124 TAILINGS GRAIN SIZE DISTRIBUTIONS -AS MEASURED NEX sample, second attempt - wet sieve, shaker Filtered Tails Screen Size cum % cum % (mesh) (microns) wt(g) % retain retain passing 4 4760 0 0.00 0.00 100.00 48 326 21.65 3.14 3.14 96.86 65 230 2.25 0.33 3.46 96.54 100 149 6.43 0.93 4.39 95.61 150 110 21.91 3.17 7.57 92.43 200 74 70.60 10.23 17.79 82.21 270 53 62.90 9.11 26.90 73.10 400 37 110.00 15.93 42.83 57.17 -400 0 394.70 57.17 100.00 0.00 Total 690.44 NEX sample, second attempt - wet Filtered Tails Without the NEX sample, third attempt - wet sieve, shaker Filtered Tails Screen Size (mesh) (microns) wl(g) % retain cum % retain cum % passing 4 4760 0 0.00 0.00 100.00 48 326 15.38 4.53 4.53 95.47 65 230 1.30 0.38 4.91 95.09 100 149 3.30 0.97 5.88 94.12 150 110 9.10 2.68 8.56 91.44 200 74 21.70 6.39 14.95 85.05 270 53 42.70 12.57 27.52 72.48 400 37 52.30 15.40 42.92 57.08 -400 0 193.90 57.08 100.00 0.00 Total 339.68 sieve, shaker contamination" rocks Screen Size (mesh) (microns) wt(g) % retain cum % retain cum % passing 4 476C 0 0.00 0.00 100.00 48 326 0.00 0.00 0.00 100.00 65 230 2.25 0.34 0.34 99.66 100 149 6.43 0.96 1.30 98.70 150 110 21.91 3.28 4.57 95.43 200 74 70.60 10.56 15.13 84.87 270 53 62.90 9.41 24.54 75.46 400 37 110.00 16.45 40.98 59.02 -400 0 394.70 59.02 100.00 0.00 Total 668.79 NEX sample, third attempt - wet sieve, shaker Filtered Tails Without the "contamination" rocks Screen Size (mesh) (microns) wt(g) % retain cum % retain cum % passing 4 4760 0 0.00 0.00 100.00 48 326 0.00 0.00 0.00 100.00 65 230 1.30 0.40 0.40 99.60 100 149 3.30 1.02 1.42 98.58 150 110 9.10 2.81 4.22 95.78 200 74 21.70 6.69 10.92 89.08 270 53 42.70 13.17 24.08 75.92 400 37 52.30 16.13 40.21 59.79 -400 0 193.90 59.79 100.00 0.00 Total 324.30 RIVER AGGREGATE GRAIN SIZE DISTRIBUTIONS - AS MEASURED River Rock as Received Screen Size (microns) (mm) wt(g) % retain cum % retain cum % passing 100000 100 0 0.00 0.00 100.00 26500 26.5 4259.70 26.85 26.85 73.15 9500 9.5 3670.40 23.13 49.98 50.02 2360 2.36 3275.40 20.64 70.63 29.37 1180 1.18 969.30 6.11 76.74 23.26 300 0.3 2713.80 17.10 93.84 6.16 0 0 977.30 6.16 100.00 0.00 Total 15865.90 River Rock used in Column Tests Screen Size (mesh) (mm) wt(g) % retain cum % retain cum % passing 26.5 0.00 0.00 0.00 100.00 9.5 3670.40 31.62 31.62 68.38 2.36 3275.40 28.22 59.85 40.15 1.18 969.30 8.35 68.20 31.80 0.3 2713.80 23.38 91.58 8.42 0 977.30 8.42 100.00 0.00 Total 11606.20 - 0.3 mm 977.30 Total used 11606.20 % 8.420499 therefore, use 9% cement for this portion, but less than that for the remaining 125 APPENDIX C PELLETIZATION TEST RESULTS 109 PELLETIZATION TESTING RESULTS Force at Failure vs. Size wrt % Cement 3 Day 109 Test 1.5 2 Median Size (cm) > to I sr £ co E o p 2 ° - o c W C (0 X ra 7.00E+07 6.00E+07 5.00E+07 4.00E+07 -| 3.00E+07 2.00E+07 1.00E+07 0.00E+00 Strength vs. Size wrt % Cement 3 Day Test 109 • 3% Cement • 4% Cement A 6% Cement x8% Cement -10% Cement * 1 1 1 1-• • • ^ — m ^ ^ 0.5 1 L _ 1 1.5 2 Median Size (cm) 2.5 0) > "35 co S | co 5 I E w E C N E o CO CL 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 0.90 Strength vs. Aspect Ratio wrt % Cement 3 Day Test 109 5T +• • 3% Cement • 4% Cement A 6% Cement X8% Cement -10% Cement 1.00 1.10 1.20 1.30 1.40 1.50 Aspect Ratio (Dim 1/Dim 3) —I 1.60 1.70 a Sm ro LL +•* (C CD o 73 Q. Q. < 700.00 600.00 500.00 400.00 300.00 200.00 100.00 0.00 Force at Failure vs. Size wrt % Cement 7 Day 109 Test • 3% Cement • 4% Cement • 6% Cement x 8 % Cement -10% Cement > '35 « £ f a E OT 0) • w =j 0 <D ro 1 ^ X ro 8.00E+07 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 Strength vs. Size wrt % Cement 7 Day Test 109 0.5 • 3% Cement • 4% Cement • 6% Cement x 8% Cement -10% Cement 1.5 2 Median Size (cm) 2.5 > 'cn n a> I s C OT O <D O £ E w E CN E z i-o ro Q. 8.00E+07 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 Strength vs. Aspect Ratio wrt % Cement 7 Day Test 109 i x ^ « ^ y f g • 3% Cement • 4% Cement • 6% Cement x 8 % Cement -10% Cement 0.90 1.00 1.10 1.20 1.30 Aspect Ratio (Dim 1/Dim 3) 1.40 1.50 129 re o -~ ^ re a a < 1200 1000 -i 800 600 400 200 0 Force at Failure vs. Size wrt % Cement 28 or More Days 109 Test • 4% Cement Wet A 6% Cement Wet x8% Cement Wet -10% Cement Wet o3% Cement Dry • 4% Cement Dry A 6% Cement Dry 38% Cement Dry x g fTM U \ i X ^ H i t « i r l f l ? J - X + 0.5 1.5 Size (cm) 2.5 to o 8. 7. 6. E 5 5 f 5 •i I ! 4 ° 1. 00E+07 00E+07 00E+07 00E+07 00E+07 00E+07 00E+07 00E+07 0.00E+00 Strength vs. Size wrt % Cement 28 or More Days Test 109 • 4% Cement Wet A 6% Cement Wet x8% Cement Wet -10% Cement Wet o 3% Cement Dry • 4% Cement Dry A 6% Cement Dry g8% Cement Dry • B f l f j | | i | B - M H - 0 . a • a _^r f_B <^  iQ —0f ^ ^ iy ^ H 1 h O -(-0.5 1 1.5 2 Median Size (cm) 2.5 IT) 6 E » II S « re cu E o o 8.00E+07 7.00E+07 6.00E+07 CN 5.00E+07 E j£ 4.00E+07 £ 3.00E+07 2.00E+07 1.00E+07 0.00E+00 Strength vs. Aspect Ratio wrt % Cement 28 or More Days Test 109 0.9 • 4% Cement Wet A 6% Cement Wet x8% Cement Wet -10% Cement Wet o3% Cement Dry • 4% Cement Dry A 6% Cement Dry a8% Cement Dry 1.1 1.2 1.3 1.4 1.5 Aspect Ratio (Dim 1/Dim 3) —I— 1.6 1.7 100 0) o T 3 OJ Q. Q. < F o r c e at F a i l u r e v s . T i m e wrt S i z e 3% C e m e n t & 109 T a i l i n g s 10 20 30 Time (days) 40 50 • Ave 1.1 cm • Ave 1.5 cm A Ave 2.0 cm x Ave 2.5 cm oAve 1.1 cm B • Ave 1.5 cm B A Ave 2.0 cm B ggAve 2.5 cm B F o r c e at F a i l u r e v s T i m e w r t S i z e 4% C e m e n t & 109 T a i l i n g s z i U U -S 180 - • Ave 1.1 cm A = 160 -m A • Ave 1.5 cm A t 1 4 0 - X A . — -• 120 - ^ — A Ave 2.0 cm A ed Force (N) cn co o x Ave 2.5 cm A oAve 1.1 cm B Q. 40 -CL < 20 - n : : " " " " s ' • Ave 1.5 cm B n i i i i A Ave 2.0 cm B u i i i i 0 10 20 30 40 50 Time (days) CO LL. 01 o I— o LL •a a> "D. CL < F o r c e at F a i l u r e v s . T i m e wrt S i z e 6% C e m e n t & 109 T a i l i n g s 20 30 Time (days) 40 50 • Ave 1.1 cm A • Ave 1.5 cm A A Ave 2.0 cm A x Ave 2.5 cm A oAve 1.1 cm B • Ave 1.5 cm B AAve 2.0 cm B n LL. CD O Q. CL < F o r c e at F a i l u r e v s . T i m e wrt S i z e 8% C e m e n t & 109 T a i l i n g s 10 20 30 Time (days) 40 50 • Ave 1.1 cm A • Ave 1.5 cm A A Ave 2.0 cm A x Ave 2.5 cm A o A v e 1.1 cm B • Ave 1.5 cm B A Ave 2.0 cm B a A v e 2.3 cm B _3 o o T3 O Q. Q. < 450 400 350 300 150 F o r c e at F a i l u r e v s . T i m e wrt S i z e 10% C e m e n t & 109 T a i l i n g s A - A A _ A A-A * 1 I 1 X 1 1 10 20 30 Time (days) • Ave 1.1 cm A • Ave 1.5 cm A A Ave 2.0 cm A 40 50 S t r e n g t h v s . T i m e wrt S i z e 3% C e m e n t & 109 T a i l i n g s 10 20 30 Time (days) 40 50 • Ave 1.1 cm • Ave 1.5 cm A Ave 2.0 cm x Ave 2.5 cm oAve 1.1 cm B o Ave 1.5 cm B A Ave 2.0 cm B igAve 2.5 cm B S t r e n g t h v s T i m e w r t S i z e 4% C e m e n t & 109 T a i l i n g s 10 20 30 Time (days) 40 50 • Ave 1.1 cm A • Ave 1.5 cm A A Ave 2.0 cm A x Ave 2.5 cm A oAve 1.1 cm B • Ave 1.5 cm B A Ave 2.0 cm B S t r e n g t h v s . T i m e wrt S i z e 6% C e m e n t & 109 T a i l i n g s 20 30 Time (days) 40 50 • Ave 1.1 cm A • Ave 1.5 cm A A Ave 2.0 cm A x Ave 2.5 cm A oAve 1.1 cm B o Ave 1.5 cm B A Ave 2.0 cm B 133 S t r e n g t h v s . T i m e wrt S i z e 8% C e m e n t & 109 T a i l i n g s 20 30 Time (days) 40 50 • Ave 1.1 cm A • Ave 1.5 cm A A Ave 2.0 cm A x Ave 2.5 cm A oAve 1.1 cm B • Ave 1.5 cm B AAve 2.0 cm B ggAve 2.3 cm B 80 2 60 i— o cs E: 40 + 20 4-2 to S t r e n g t h v s . T i m e wrt S i z e 10% C e m e n t & 109 T a i l i n g s 10 + 20 30 Time (days) 40 50 • Ave 1.1 cm A • Ave 1.5 cm A • Ave 2.0 cm A cn C "in I > E 3 o 109 Pellet Size Distributions All Batches 10 Size (mm) —•— #10- 8% C, % W #11 -6%C, 11.8%W #12- 6% C, 14% W #13- 6%C, 16.4% W —*— #14- 4%C, 16.8% W —•— #15- 4%C, 14.3% W —I— #16- 4%C, 14.5% W #17- 3% C, 14.0% W #18- 3%C, 14.0% W —o— #19- 8%C, 14.0% W —O—#20-8%C, 16.0% W #21 -10% C, 14.9% W —a— #22- 10% C, 17.0% W 109 Pellet Size Distributions 100.00 90.00 cn 80.00 _c N 70.00 <n ra 0 . 60.00 50.00 > *-> 40.00 3 E 30.00 o 20.00 10.00 0.00 1000 10 Size (mm) #21 - 10% C, 14.9% W a—#22-10% C, 17.0% W 100.00 90.00 80.00 £ in 70.00 in I 60.00 o 50.00 > m 40.00 3 E 30.00 3 o 20.00 10.00 0.00 1000 109 Pellet Size Distributions •••• % \ \\ \ V #10-8%C, %W #-to_ oo/. n -i A no/ IA/ V! rs \ irio-g/o J*t.u /o vv _n— aon . «o/ r 1 A rw_ \A/ \ \ 100 10 Size (mm) 0.1 100.00 90.00 80.00 C « CO 70.00 SS0. 60.00 | 50.00 ra 40.00 3 E 30.00 3 u 20.00 10.00 0.00 1000 109 Pellet Size Distributions 100 10 Size (mm) -#11 -6%C, 11.8%W -a— #12-6% C, 14%W -X—#13 - 6% C, 16.4% W 0.1 109 Pellet Size Distributions 100.00 90.00 a> 8000 "I 70.00 co n a. 0> > I 3 E 3 u 10 Size (mm) *—#14-4%C, 16.8% W #15-4%C, 14.3% W #16-4%C, 14.5% W Eskay Creek Agglomeration project First number in each batch # indicates 109 Tailings Data age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Ratio Strength Average Diam (m) Strengths (Pa) 1st 2nd 3rd lbs N Compressive Tensile Shear 309 8% 3.1 2.8 2.7 0.87 63 280.33 0.03 3.53E+07 4.70E+06 1.18E+07 8% 2.9 2.7 2.4 0.83 63 280.33 0.03 3.68E+07 4.90E+06 1.23E+07 8% 2.8 2.6 2.5 0.89 56 249.19 0.03 3.59E+07 4.77E+06 1.20E+07 8% 2.4 2.2 2.1 0.88 31 137.94 0.02 3.29E+07 4.37E+06 1.10E+07 8% 2.3 2 2 0.87 42 186.89 0.02 3.81 E+07 5.07E+06 1.27E+07 8% 2.4 2.1 2 0.83 42 186.89 0.02 3.72E+07 4.95E+06 1.24E+07 8% 1.7 1.5 1.5 0.88 14 62.30 0.02 3.21 E+07 4.27E+06 1.07E+07 8% 1.7 1.5 1.3 0.76 16 71.20 0.02 3.43E+07 4.56E+06 1.14E+07 8% 1.7 1.4 1.3 0.76 17 75.65 0.01 3.58E+07 4.76E+06 1.19E+07 8% 1.4 1.1 1.1 0.79 11 48.95 0.01 3.56E+07 4.74E+06 1.19E+07 8% 1.3 1.1 1.1 0.85 14 62.30 0.01 3.92E+07 5.21 E+06 1.31 E+07 310 8% 2.5 2.4 2.2 0.88 33 146.84 0.02 3.22E+07 4.28E+06 1.07E+07 8% 2.7 2.5 2.1 0.78 37 164.64 0.02 3.27E+07 4.35E+06 1.09E+07 8% 2.8 2.5 2.2 0.79 37 164.64 0.03 3.23E+07 4.29E+06 1.08E+07 8% 2.2 2 1.8 0.82 25 111.24 0.02 3.29E+07 4.37E+06 1.10E+07 8% 2.8 2 1.7 0.61 39 173.54 0.02 3.66E+07 4.87E+06 1.22E+07 8% 2.7 2 2 0.74 41 182.44 0.02 3.67E+07 4.87E+06 1.22E+07 8% 2.2 1.9 1.8 0.82 34 151.29 0.02 3.70E+07 4.93E+06 1.23E+07 8% 1.7 1.5 1.4 0.82 17 75.65 0.02 3.46E+07 4.61 E+06 1.15E+07 8% 1.7 1.5 1.5 0.88 21 93.44 0.02 3.68E+07 4.89E+06 1.23E+07 8% 2 1.5 1.5 0.75 24 106.79 0.02 3.72E+07 4.95E+06 1.24E+07 311 6% 11.8% 1.97 2.5 2.5 2.4 0.96 15 66.75 0.02 1.72E+07 2.28E+06 5.72E+06 6% 11.8% 1.97 2.4 2.4 2.3 0.96 11 48.95 0.02 1.59E+07 2.12E+06 5.30E+06 6% 11.8% 1.97 2.2 2.1 2 0.91 12 53.40 0.02 1.78E+07 2.36E+06 5.92E+06 6% 11.8% 1.97 2.1 2 2 0.95 14 62.30 0.02 1.92E+07 2.55E+06 6.39E+06 6% 11.8% 1.97 1.9 1.8 1.6 0.84 10 44.50 0.02 1.87E+07 2.49E+06 6.24E+06 6% 11.8% 1.97 1.5 1.4 1.4 0.93 11 48.95 0.01 2.24E+07 2.97E+06 7.45E+06 6% 11.8% 1.97 1.5 1.4 1.3 0.87 9 40.05 0.01 2.12E+07 2.81 E+06 7.05E+06 6% 11.8% 1.97 1.5 1.4 1.3 0.87 6 26.70 0.01 1.85E+07 2.46E+06 6.16E+06 6% 11.8% 1.97 1.1 1.1 1.1 1.00 5 22.25 0.01 2.04E+07 2.72E+06 6.81 E+06 6% 11.8% 1.97 1.3 1.1 1 0.77 4 17.80 0.01 1.87E+07 2.48E+06 6.23E+06 312 6% 14.0% 2.33 2.5 2.5 2.3 0.92 23 102.34 0.02 1.99E+07 2.65E+06 6.64E+06 6% 14.0% 2.33 2.6 2.4 2.2 0.85 21 93.44 0.02 1.96E+07 2.61 E+06 6.53E+06 6% 14.0% 2.33 2.4 2.2 2.1 0.88 14 62.30 0.02 1.80E+07 2.39E+06 6.00E+06 6% 14.0% 2.33 2 2 1.8 0.90 14 62.30 0.02 1.97E+07 2.61 E+06 6.55E+06 6% 14.0% 2.33 2 1.9 1.8 0.90 14 62.30 0.02 2.00E+07 2.66E+06 6.67E+06 6% 14.0% 2.33 2 1.9 1.5 0.75 12 53.40 0.02 1.95E+07 2.60E+06 6.51 E+06 6% 14.0% 2.33 1.6 1.5 1.3 0.81 5 22.25 0.01 1.68E+07 2.23E+06 5.60E+06 6% 14.0% 2.33 1.7 1.6 1.6 0.94 6 26.70 0.02 1.67E+07 2.23E+06 5.58E+06 6% 14.0% 2.33 1.6 1.5 1.4 0.88 11 48.95 0.02 2.16E+07 2.87E+06 7.20E+06 6% 14.0% 2.33 1.2 1 1 0.83 5 22.25 0.01 2.11E+07 2.80E+06 7.02E+06 6% 14.0% 2.33 1.1 1 1 0.91 5 22.25 0.01 2.14E+07 2.85E+06 7.14E+06 313 6% 16.4% 2.73 2.5 2 2 0.80 14 62.30 0.02 1.86E+07 2.47E+06 6.19E+06 6% 16.4% 2.73 2 1.9 1.7 0.85 12 53.40 0.02 1.92E+07 2.55E+06 6.39E+06 6% 16.4% 2.73 2.5 2 1.7 0.68 10 44.50 0.02 1.70E+07 2.26E+06 5.66E+06 6% 16.4% 2.73 2 2 1.8 0.90 10 44.50 0.02 1.76E+07 2.34E+06 5.86E+06 6% 16.4% 2.73 1.7 1.5 1.3 0.76 7 31.15 0.02 1.86E+07 2.47E+06 6.19E+06 6% 16.4% 2.73 1.7 1.4 1.2 0.71 7 31.15 0.01 1.92E+07 2.56E+06 6.41 E+06 6% 16.4% 2.73 1.7 1.5 1.4 0.82 6 26.70 0.02 1.75E+07 2.32E+06 5.82E+06 6% 16.4% 2.73 1.2 1.2 1 0.83 5 22.25 0.01 1.98E+07 2.64E+06 6.61 E+06 6% 16.4% 2.73 1.2 1.1 1.1 0.92 5 22.25 0.01 2.01 E+07 2.68E+06 6.71 E+06 6% 16.4% 2.73 1.2 1.1 1 0.83 3 13.35 0.01 1.72E+07 2.29E+06 5.74E+06 6% 16.4% 2.73 1 1 0.8 0.80 5 22.25 0.01 2.25E+07 3.00E+06 7.51 E+06 6% 16.4% 2.73 1 0.95 0.85 0.85 3 13.35 0.01 1.92E+07 2.55E+06 6.39E+06 314 4% 16.8% 4.20 3.3 2.9 2.8 0.85 12 53.40 0.03 8.49E+06 1.13E+06 2.83E+06 4% 16.8% 4.20 3 2.9 2.5 0.83 11 48.95 0.03 8.54E+06 1.14E+06 2.85E+06 4% 16.8% 4.20 2.6 2.4 2.4 0.92 12 53.40 0.02 9.67E+06 1.29E+06 3.22E+06 4% 16.8% 4.20 2.5 2.4 2.4 0.96 11 48.95 0.02 9.46E+06 1.26E+06 3.15E+06 4% 16.8% 4.20 2.5 2 2 0.80 11 48.95 0.02 1.03E+07 1.37E+06 3.44E+06 4% 16.8% 4.20 2.2 2 2 0.91 11 48.95 0.02 1.06E+07 1.41 E+06 3.53E+06 4% 16.8% 4.20 1.8 1.5 1.4 0.78 8 35.60 0.02 1.15E+07 1.52E+06 3.82E+06 4% 16.8% 4.20 1.6 1.5 1.4 0.88 6 26.70 0.02 1.06E+07 1.42E+06 3.55E+06 4% 16.8% 4.20 1.6 1.5 1.4 0.88 5 22.25 0.02 1.00E+07 1.33E+06 3.34E+06 4% 16.8% 4.20 1.4 1.1 1 0.71 2 8.90 0.01 8.81 E+06 1.17E+06 2.94E+06 137 Eskay Creek Agglomeration project First number in each batch # indicates 109 Tailings Data age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Ratio Strength Average Diam (m) Strengths (Pa 1st 2nd 3rd lbs N Compressive Tensile Shear 4% 16.8% 4.20 1.3 1.1 1 0.77 3 13.35 0.01 1.02E+07 1.36E+06 3.41 E+06 315 4% 14.3% 3.58 2.4 2.1 2 0.83 8 35.60 0.02 9.21 E+06 1.23E+06 3.07E+06 4% 14.3% 3.58 2.5 1.9 1.9 0.76 9 40.05 0.02 9.89E+06 1.32E+06 3.30E+06 4% 14.3% 3.58 1.9 1.8 1.8 0.95 3 13.35 0.02 7.41 E+06 9.86E+05 2.47E+06 4% 14.3% 3.58 1.5 1.4 1.3 0.87 6 26.70 0.01 1.11E+07 1.48E+06 3.71 E+06 4% 14.3% 3.58 1.8 1.5 1.4 0.78 6 26.70 0.02 1.04E+07 1.38E+06 3.47E+06 4% 14.3% 3.58 1.5 1.4 1.4 0.93 8 35.60 0.01 1.21E+07 1.61 E+06 4.04E+06 4% 14.3% 3.58 1.1 1.1 1 0.91 5 22.25 0.01 1.25E+07 1.66E+06 4.17E+06 4% 14.3% 3.58 1.1 1.1 1.1 1.00 3 13.35 0.01 1.04E+07 1.38E+06 3.46E+06 4% 14.3% 3.58 1.2 1.2 1 0.83 3 13.35 0.01 1.01E+07 1.34E+06 3.36E+06 316 4% 14.5% 3.63 2.5 2.4 2 0.80 11 48.95 0.02 9.72E+06 1.29E+06 3.24E+06 4% 14.5% 3.63 2.2 2.2 2 0.91 9 40.05 0.02 9.58E+06 1.27E+06 3.19E+06 4% 14.5% 3.63 2.4 1.8 1.8 0.75 7 31.15 0.02 9.41 E+06 1.25E+06 3.14E+06 4% 14.5% 3.63 2.1 1.8 1.8 0.86 7 31.15 0.02 9.65E+06 1.28E+06 3.22E+06 4% 14.5% 3.63 1.7 1.5 1.5 0.88 6 26.70 0.02 1.04E+07 1.38E+06 3.47E+06 4% 14.5% 3.63 1.7 1.5 1.3 0.76 5 22.25 0.02 1.00E+07 1.33E+06 3.34E+06 4% 14.5% 3.63 1.5 1.5 1.3 0.87 4 17.80 0.01 9.51 E+06 1.26E+06 3.17E+06 4% 14.5% 3.63 1.1 1 1 0.91 3 13.35 0.01 1.09E+07 1.45E+06 3.63E+06 4% 14.5% 3.63 1.2 1 0.9 0.75 5 22.25 0.01 1.29E+07 1.72E+06 4.30E+06 4% 14.5% 3.63 1 0.9 0.9 0.90 3 13.35 0.01 1.17E+07 1.55E+06 3.89E+06 317 3% 14.0% 4.67 2.2 2.1 2 0.91 6 26.70 0.02 5.03E+06 6.69E+05 1.68E+06 3% 14.0% 4.67 2.1 2 1.9 0.90 5 22.25 0.02 4.89E+06 6.50E+05 1.63E+06 3% 14.0% 4.67 2.2 2 1.8 0.82 5 22.25 0.02 4.89E+06 6.50E+05 1.63E+06 3% 14.0% 4.67 1.5 1.4 1.3 0.87 6 26.70 0.01 6.59E+06 8.76E+05 2.20E+06 3% 14.0% 4.67 1.8 1.5 1.5 0.83 5 22.25 0.02 5.73E+06 7.62E+05 1.91 E+06 3% 14.0% 4.67 1.7 1.4 1.3 0.76 6 26.70 0.01 6.44E+06 8.56E+05 2.15E+06 3% 14.0% 4.67 1.1 1 1 0.91 3 13.35 0.01 6.44E+06 8.56E+05 2.15E+06 3% 14.0% 4.67 1 1 0.9 0.90 3 13.35 0.01 6.65E+06 8.85E+05 2.22E+06 3% 14.0% 4.67 1.2 1.1 1.1 0.92 3 13.35 0.01 6.05E+06 8.05E+05 2.02E+06 318 3% 14.0% 4.67 2.5 2.4 2.3 0.92 9 40.05 0.02 5.26E+06 7.00E+05 1.75E+06 3% 14.0% 4.67 2.5 2.5 2.3 0.92 10 44.50 0.02 5.38E+06 7.15E+05 1.79E+06 3% 14.0% 4.67 2.2 2.1 2 0.91 7 31.15 0.02 5.29E+06 7.04E+05 1.76E+06 3% 14.0% 4.67 2.1 2 1.9 0.90 6 26.70 0.02 5.19E+06 6.91 E+05 1.73E+06 3% 14.0% 4.67 2 2 1.8 0.90 8 35.60 0.02 5.81 E+06 7.73E+05 1.94E+06 3% 14.0% 4.67 1.5 1.5 1.4 0.93 3 13.35 0.01 5.05E+06 6.72E+05 1.68E+06 3% 14.0% 4.67 1.6 1.5 1.4 0.88 4 17.80 0.02 5.50E+06 7.31 E+05 1.83E+06 3% 14.0% 4.67 1.7 1.5 1.4 0.82 5 22.25 0.02 5.86E+06 7.79E+05 1.95E+06 3% 14.0% 4.67 1.3 1 0.9 0.69 3 13.35 0.01 6.33E+06 8.42E+05 2.11E+06 3% 14.0% 4.67 1.2 1.1 1 0.83 3 13.35 0.01 6.14E+06 8.17E+05 2.05E+06 3% 14.0% 4.67 1.1 1 1 0.91 3 13.35 0.01 6.44E+06 8.56E+05 2.15E+06 419 8% 14.0% 1.75 2.7 2.5 2.3 0.85 79 351.53 0.03 4.16E+07 5.53E+06 1.39E+07 8% 14.0% 1.75 2.7 2.6 2.5 0.93 107 476.12 0.03 4.48E+07 5.96E+06 1.49E+07 8% 14.0% 1.75 2.2 2.1 2.1 0.95 77 342.63 0.02 4.59E+07 6.11 E+06 1.53E+07 8% 14.0% 1.75 2.2 2.1 1.85 0.84 67 298.13 0.02 4.47E+07 5.95E+06 1.49E+07 8% 14.0% 1.75 2 2 1.8 0.90 59 262.53 0.02 4.45E+07 5.92E+06 1.48E+07 8% 14.0% 1.75 1.5 1.4 1.3 0.87 38 169.09 0.01 4.79E+07 6.38E+06 1.60E+07 8% 14.0% 1.75 1.5 1.5 1.2 0.80 43 191.34 0.01 4.94E+07 6.57E+06 1.65E+07 8% 14.0% 1.75 1.5 1.5 1.4 0.93 59 262.53 0.01 5.36E+07 7.13E+06 1.79E+07 8% 14.0% 1.75 1.2 1 0.9 0.75 15 66.75 0.01 4.33E+07 5.76E+06 1.44E+07 8% 14.0% 1.75 1 0.9 0.9 0.90 20 88.99 0.01 5.10E+07 6.79E+06 1.70E+07 8% 14.0% 1.75 1.1 1 1 0.91 30 133.49 0.01 5.45E+07 7.25E+06 1.82E+07 320 8% 16.0% 2.00 2.2 2 1.9 0.86 20 88.99 0.02 3.03E+07 4.02E+06 1.01E+07 8% 16.0% 2.00 2.3 2 1.8 0.78 29 129.04 0.02 3.43E+07 4.56E+06 1.14E+07 8% 16.0% 2.00 2.1 1.9 1.8 0.86 24 106.79 0.02 3.33E+07 4.42E+06 1.11E+07 8% 16.0% 2.00 1.7 1.6 1.5 0.88 22 97.89 0.02 3.65E+07 4.86E+06 1.22E+07 8% 16.0% 2.00 1.9 1.6 1.6 0.84 23 102.34 0.02 3.60E+07 4.79E+06 1.20E+07 8% 16.0% 2.00 1.7 1.6 1.5 0.88 16 71.20 0.02 3.29E+07 4.37E+06 1.10E+07 8% 16.0% 2.00 1.4 1.3 1.3 0.93 13 57.85 0.01 3.48E+07 4.63E+06 1.16E+07 8% 16.0% 2.00 1.5 1.3 1.1 0.73 13 57.85 0.01 3.52E+07 4.68E+06 1.17E+07 8% 16.0% 2.00 1.3 1.1 0.9 0.69 8 35.60 0.01 3.35E+07 4.45E+06 1.12E+07 8% 16.0% 2.00 1.2 1.1 1.1 0.92 13 57.85 0.01 3.88E+07 5.16E+06 1.29E+07 8% 16.0% 2.00 1.1 1.1 1 0.91 11 48.95 0.01 3.78E+07 5.03E+06 1.26E+07 308 8% 23.0% 2.87 2.95 2.6 2.4 0.81 11 48.95 0.03 2.08E+07 2.76E+06 6.93E+06 8% 23.0% 2.87 2.6 2.5 2.4 0.92 10 44.50 0.03 2.09E+07 2.78E+06 6.96E+06 138 Eskay Creek Agglomeration project First number in each batch # indicates 109 Tailings Data age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Ratio Strength Average Diam (m) Strengths (Pa) 1st 2nd 3rd lbs N Compressive Tensile Shear 8% 23.0% 2.87 3.2 2.5 2.35 0.73 16 71.20 0.03 2.36E+07 3.13E+06 7.85E+06 8% 23.0% 2.87 2.1 1.95 1.75 0.83 8 35.60 0.02 2.30E+07 3.05E+06 7.65E+06 8% 23.0% 2.87 1.9 1.8 1.7 0.89 3 13.35 0.02 1.74E+07 2.31 E+06 5.80E+06 8% 23.0% 2.87 1.8 1.8 1.5 0.83 7 31.15 0.02 2.37E+07 3.16E+06 7.91 E+06 8% 23.0% 2.87 1.6 1.3 1.3 0.81 5 22.25 0.01 2.47E+07 3.28E+06 8.23E+06 8% 23.0% 2.87 1.6 1.4 1.3 0.81 6 26.70 0.01 2.56E+07 3.41 E+06 8.54E+06 8% 23.0% 2.87 1.4 1.2 1.1 0.79 3 13.35 0.01 2.25E+07 2.99E+06 7.49E+06 8% 23.0% 2.87 1.1 1 1.1 1.00 1 4.45 0.01 1.73E+07 2.30E+06 5.76E+06 321 10% 14.9% 1.49 2.3 2 1.8 0.78 69 307.03 0.02 5.40E+07 7.18E+06 1.80E+07 10% 14.9% 1.49 2.3 2 1.9 0.83 80 355.98 0.02 5.62E+07 7.48E+06 1.87E+07 10% 14.9% 1.49 2.1 2 1.8 0.86 65 289.23 0.02 5.38E+07 7.15E+06 1.79E+07 10% 14.9% 1.49 1.6 1.5 1.4 0.88 25 111.24 0.02 4.70E+07 6.25E+06 1.57E+07 10% 14.9% 1.49 1.5 1.4 1.4 0.93 29 129.04 0.01 5.11 E+07 6.80E+06 1.70E+07 10% 14.9% 1.49 1.6 1.5 1.5 0.94 28 124.59 0.02 4.83E+07 6.42E+06 1.61 E+07 10% 14.9% 1.49 1.6 1.3 1.3 0.81 28 124.59 0.01 5.17E+07 6.88E+06 1.72E+07 10% 14.9% 1.49 1.3 1 1 0.77 18 80.10 0.01 5.26E+07 6.99E+06 1.75E+07 10% 14.9% 1.49 1.1 1 0.9 0.82 13 57.85 0.01 4.95E+07 6.59E+06 1.65E+07 10% 14.9% 1.49 1.2 1.1 1 0.83 17 75.65 0.01 5.08E+07 6.76E+06 1.69E+07 322 10% 17.0% 1.70 2.3 2.15 1.8 0.78 49 218.04 0.02 4.70E+07 6.25E+06 1.57E+07 10% 17.0% 1.70 2.3 2 1.9 0.83 43 191.34 0.02 4.57E+07 6.08E+06 1.52E+07 10% 17.0% 1.70 2.3 2.1 1.9 0.83 48 213.59 0.02 4.67E+07 6.21 E+06 1.56E+07 10% 17.0% 1.70 1.8 1.6 1.3 0.72 31 137.94 0.02 4.89E+07 6.50E+06 1.63E+07 10% 17.0% 1.70 1.8 1.6 1.4 0.78 33 146.84 0.02 4.94E+07 6.57E+06 1.65E+07 10% 17.0% 1.70 1.6 1.5 1.4 0.88 31 137.94 0.02 5.05E+07 6.71 E+06 1.68E+07 10% 17.0% 1.70 1.5 1.3 1.1 0.73 21 93.44 0.01 4.88E+07 6.49E+06 1.63E+07 10% 17.0% 1.70 1.5 1.3 1.2 0.80 29 129.04 0.01 5.36E+07 7.13E+06 1.79E+07 10% 17.0% 1.70 1.1 1 1 0.91 22 97.89 0.01 5.80E+07 7.72E+06 1.93E+07 10% 17.0% 1.70 1.1 1 0.9 0.82 19 84.55 0.01 5.62E+07 7.47E+06 1.87E+07 10% 17.0% 1.70 1.2 1.1 1 0.83 21 93.44 0.01 5.45E+07 7.25E+06 1.82E+07 709 8% 2.7 2.5 2 0.74 92 409.38 0.02 4.46E+07 5.94E+06 1.49E+07 8% 2.7 2.5 2.3 0.85 122 542.87 0.03 4.80E+07 6.39E+06 1.60E+07 8% 2.5 2.4 2.3 0.92 63 280.33 0.02 3.96E+07 5.27E+06 1.32E+07 8% 2.1 2 1.6 0.76 62 275.88 0.02 4.56E+07 6.07E+06 1.52E+07 8% 2.2 1.9 1.6 0.73 64 284.78 0.02 4.65E+07 6.19E+06 1.55E+07 8% 2.4 2 1.8 0.75 59 262.53 0.02 4.30E+07 5.73E+06 1.43E+07 8% 1.6 1.5 1.3 0.81 40 177.99 0.01 4.71 E+07 6.26E+06 1.57E+07 8% 1.6 1.5 1.4 0.88 35 155.74 0.02 4.45E+07 5.92E+06 1.48E+07 8% 1.7 1.4 1.3 0.76 35 155.74 0.01 4.56E+07 6.06E+06 1.52E+07 710 8% 2.5 2.5 2.3 0.92 79 351.53 0.02 4.21 E+07 5.60E+06 1.40E+07 8% 2.5 2.4 2.2 0.88 73 324.83 0.02 4.19E+07 5.57E+06 1.40E+07 8% 2.2 2.2 2 0.91 71 315.93 0.02 4.44E+07 5.90E+06 1.48E+07 8% 2.2 2 1.8 0.82 66 293.68 0.02 4.54E+07 6.04E+06 1.51 E+07 8% 2 1.9 1.8 0.90 62 275.88 0.02 4.60E+07 6.12E+06 1.53E+07 8% 2.2 2 1.8 0.82 68 302.58 0.02 4.59E+07 6.10E+06 1.53E+07 8% 1.8 1.5 1.4 0.78 37 164.64 0.02 4.44E+07 5.90E+06 1.48E+07 8% 1.7 1.5 1.4 0.82 30 133.49 0.02 4.18E+07 5.57E+06 1.39E+07 8% 1.5 1.5 1.4 0.93 29 129.04 0.01 4.23E+07 5.63E+06 1.41 E+07 8% 1.3 1.1 1.1 0.85 20 88.99 0.01 4.41 E+07 5.87E+06 1.47E+07 8% 1.2 1.1 1.1 0.92 15 66.75 0.01 4.07E+07 5.41 E+06 1.36E+07 711 6% 11.8% 1.97 2.7 2.5 2.5 0.93 37 164.64 0.03 2.27E+07 3.02E+06 7.58E+06 6% 11.8% 1.97 1.9 1.9 1.9 1.00 24 106.79 0.02 2.39E+07 3.18E+06 7.98E+06 6% 11.8% 1.97 1.8 1.8 1.8 1.00 18 80.10 0.02 2.25E+07 3.00E+06 7.52E+06 6% 11.8% 1.97 1.5 1.5 1.3 0.87 10 44.50 0.01 2.14E+07 2.85E+06 7.14E+06 6% 11.8% 1.97 1.5 1.5 1.5 1.00 13 57.85 0.02 2.28E+07 3.04E+06 7.61 E+06 6% 11.8% 1.97 1.7 1.5 1.3 0.76 12 53.40 0.02 2.22E+07 2.96E+06 7.41 E+06 6% 11.8% 1.97 1.2 1.1 1 0.83 20 88.99 0.01 3.24E+07 4.31 E+06 1.08E+07 6% 11.8% 1.97 1.2 1 0.9 0.75 6 26.70 0.01 2.28E+07 3.03E+06 7.58E+06 6% 11.8% 1.97 1 1 0.9 0.90 7 31.15 0.01 2.48E+07 3.29E+06 8.26E+06 712 6% 14.0% 2.33 2.6 2.6 2.2 0.85 60 266.98 0.03 2.71 E+07 3.60E+06 9.02E+06 6% 14.0% 2.33 2.6 2.5 2 0.77 23 102.34 0.02 2.02E+07 2.69E+06 6.73E+06 6% 14.0% 2.33 2 1.9 1.8 0.90 32 142.39 0.02 2.63E+07 3.50E+06 8.78E+06 6% 14.0% 2.33 2.2 2.1 1.9 0.86 29 129.04 0.02 2.40E+07 3.20E+06 8.01 E+06 6% 14.0% 2.33 2 2 2 1.00 21 93.44 0.02 2.21 E+07 2.94E+06 7.37E+06 6% 14.0% 2.33 1.5 1.5 1.4 0.93 14 62.30 0.01 2.37E+07 3.15E+06 7.89E+06 139 Eskay Creek Agglomeration project 109 Tailings Data Brazilian test First number in each batch # indicates age of pellet at time of testing. 'B' indicates curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Strength Average Strengths (Pa 1 1st 2nd 3rd Ratio lbs N Diam (m) Compressive Tensile Shear 6% 14.0% 2.33 1.7 1.5 1.4 0.82 16 71.20 0.02 2.42E+07 3.22E+06 8.07E+06 6% 14.0% 2.33 1.5 1.4 1.3 0.87 12 53.40 0.01 2.33E+07 3.10E+06 7.76E+06 6% 14.0% 2.33 1.1 1.1 0.9 0.82 7 31.15 0.01 2.36E+07 3.14E+06 7.86E+06 6% 14.0% 2.33 1.1 1 0.8 0.73 2 8.90 0.01 1.63E+07 2.17E+06 5.44E+06 6% 14.0% 2.33 1.2 1.1 1.1 0.92 9 40.05 0.01 2.45E+07 3.26E+06 8.16E+06 713 6% 16.4% 2.73 2.5 2.2 1.9 0.76 19 84.55 0.02 2.01 E+07 2.67E+06 6.69E+06 6% 16.4% 2.73 2.4 2.2 1.7 0.71 20 88.99 0.02 2.09E+07 2.78E+06 6.97E+06 6% 16.4% 2.73 2 1.9 1.7 0.85 16 71.20 0.02 2.11 E+07 2.81 E+06 7.03E+06 6% 16.4% 2.73 2 1.9 1.7 0.85 14 62.30 0.02 2.02E+07 2.68E+06 6.73E+06 6% 16.4% 2.73 1.7 1.5 1.4 0.82 14 62.30 0.02 2.32E+07 3.08E+06 7.72E+06 6% 16.4% 2.73 1.8 1.6 1.4 0.78 12 53.40 0.02 2.13E+07 2.83E+06 7.10E+06 6% 16.4% 2.73 1.5 1.5 1.3 0.87 11 48.95 0.01 2.21 E+07 2.94E+06 7.37E+06 6% 16.4% 2.73 1.1 1.1 1.1 1.00 9 40.05 0.01 2.48E+07 3.30E+06 8.28E+06 6% 16.4% 2.73 1.15 1.15 1 0.87 10 44.50 0.01 2.55E+07 3.40E+06 8.51 E+06 6% 16.4% 2.73 1.1 1 0.9 0.82 8 35.60 0.01 2.55E+07 3.39E+06 8.49E+06 714 4% 16.8% 4.20 2.7 2.6 2.6 0.96 19 84.55 0.03 1.08E+07 1.43E+06 3.59E+06 4% 16.8% 4.20 2.9 2.6 2.6 0.90 19 84.55 0.03 1.06E+07 1.41 E+06 3.54E+06 4% 16.8% 4.20 2.1 2 1.9 0.90 15 66.75 0.02 1.19E+07 1.59E+06 3.97E+06 4% 16.8% 4.20 2.1 2 2 0.95 19 84.55 0.02 1.28E+07 1.70E+06 4.26E+06 4% 16.8% 4.20 1.5 1.5 1.5 1.00 8 35.60 0.02 1.17E+07 1.56E+06 3.90E+06 4% 16.8% 4.20 1.5 1.5 1.4 0.93 12 53.40 0.01 1.36E+07 1.80E+06 4.52E+06 4% 16.8% 4.20 1.5 1.4 1.3 0.87 8 35.60 0.01 1.23E+07 1.63E+06 4.09E+06 4% 16.8% 4.20 1.2 1.1 1 0.83 5 22.25 0.01 1.23E+07 1.64E+06 4.10E+06 4% 16.8% 4.20 1.2 1.2 1 0.83 4 17.80 0.01 1.11 E+07 1.48E+06 3.70E+06 4% 16.8% 4.20 1.2 1.1 1 0.83 5 22.25 0.01 1.23E+07 1.64E+06 4.10E+06 715 4% 14.3% 3.58 2.2 2.1 1.8 0.82 11 48.95 0.02 1.06E+07 1.41 E+06 3.53E+06 4% 14.3% 3.58 2.1 2 1.6 0.76 9 40.05 0.02 1.03E+07 1.37E+06 3.44E+06 4% 14.3% 3.58 1.8 1.5 1.5 0.83 11 48.95 0.02 1.26E+07 1.68E+06 4.20E+06 4% 14.3% 3.58 1.7 1.6 1.5 0.88 9 40.05 0.02 1.17E+07 1.55E+06 3.89E+06 4% 14.3% 3.58 1.7 1.6 1.5 0.88 8 35.60 0.02 1.12E+07 1.49E+06 3.74E+06 4% 14.3% 3.58 1.1 1 1 0.91 6 26.70 0.01 1.37E+07 1.82E+06 4.57E+06 4% 14.3% 3.58 1.1 1 1 0.91 6 26.70 0.01 1.37E+07 1.82E+06 4.57E+06 4% 14.3% 3.58 1.3 1.1 1 0.77 8 35.60 0.01 1.42E+07 1.89E+06 4.73E+06 716 4% 14.5% 3.63 2.3 2.1 1.8 0.78 14 62.30 0.02 1.14E+07 1.51 E+06 3.79E+06 4% 14.5% 3.63 2.2 2.1 2 0.91 15 66.75 0.02 1.15E+07 1.53E+06 3.85E+06 4% 14.5% 3.63 2 1.9 1.9 0.95 13 57.85 0.02 1.17E+07 1.55E+06 3.89E+06 4% 14.5% 3.63 1.7 1.5 1.2 0.71 6 26.70 0.01 1.08E+07 1.43E+06 3.59E+06 4% 14.5% 3.63 1.5 1.5 1.5 1.00 10 44.50 0.02 1.26E+07 1.68E+06 4.21 E+06 4% 14.5% 3.63 1.6 1.5 1.5 0.94 8 35.60 0.02 1.16E+07 1.54E+06 3.86E+06 4% 14.5% 3.63 1.2 1.1 1 0.83 6 26.70 0.01 1.31 E+07 1.74E+06 4.36E+06 4% 14.5% 3.63 1.2 1.1 1.1 0.92 6 26.70 0.01 1.29E+07 1.71 E+06 4.30E+06 4% 14.5% 3.63 1.3 1.2 1.1 0.85 5 22.25 0.01 1.16E+07 1.55E+06 3.87E+06 717 3% 14.0% 4.67 2.6 2.4 1.8 0.69 4 17.80 0.02 4.13E+06 5.50E+05 1.38E+06 3% 14.0% 4.67 2.2 2 1.6 0.73 5 22.25 0.02 4.97E+06 6.61 E+05 1.66E+06 3% 14.0% 4.67 2.1 1.8 1.7 0.81 10 44.50 0.02 6.49E+06 8.63E+05 2.16E+06 3% 14.0% 4.67 1.8 1.7 1.6 0.89 5 22.25 0.02 5.45E+06 7.24E+05 1.82E+06 3% 14.0% 4.67 1.8 1.6 1.5 0.83 4 17.80 0.02 5.21 E+06 6.93E+05 1.74E+06 3% 14.0% 4.67 1.8 1.5 1.3 0.72 3 13.35 0.02 4.94E+06 6.57E+05 1.65E+06 3% 14.0% 4.67 1.5 1.3 1.2 0.80 3 13.35 0.01 5.42E+06 7.21 E+05 1.81 E+06 3% 14.0% 4.67 1.4 1.2 1.1 0.79 3 13.35 0.01 5.72E+06 7.60E+05 1.91 E+06 3% 14.0% 4.67 1.2 1.1 1 0.83 2 8.90 0.01 5.36E+06 7.13E+05 1.79E+06 3% 14.0% 4.67 1.1 1 1 0.91 5 22.25 0.01 7.63E+06 1.01 E+06 2.54E+06 3% 14.0% 4.67 1 0.8 0.8 0.80 5 22.25 0.01 8.65E+06 1.15E+06 2.88E+06 718 3% 14.0% 4.67 2.8 2.4 2.3 0.82 16 71.20 0.02 6.25E+06 8.31 E+05 2.08E+06 3% 14.0% 4.67 2.5 2.2 2.1 0.84 13 57.85 0.02 6.21 E+06 8.26E+05 2.07E+06 3% 14.0% 4.67 2.2 2 1.8 0.82 10 44.50 0.02 6.16E+06 8.19E+05 2.05E+06 3% 14.0% 4.67 2 1.8 1.7 0.85 6 26.70 0.02 5.52E+06 7.34E+05 1.84E+06 3% 14.0% 4.67 1.7 1.6 1.5 0.88 5 22.25 0.02 5.67E+06 7.54E+05 1.89E+06 3% 14.0% 4.67 1.7 1.4 1.3 0.76 4 17.80 0.01 5.62E+06 7.48E+05 1.87E+06 3% 14.0% 4.67 1.5 1.3 1.2 0.80 4 17.80 0.01 5.97E+06 7.94E+05 1.99E+06 3% 14.0% 4.67 1.4 1.2 1.3 0.93 4 17.80 0.01 6.12E+06 8.15E+05 2.04E+06 3% 14.0% 4.67 1.4 1.2 1.1 0.79 3 13.35 0.01 5.72E+06 7.60E+05 1.91 E+06 3% 14.0% 4.67 1.1 1 0.8 0.73 4 17.80 0.01 7.32E+06 9.74E+05 2.44E+06 3% 14.0% 4.67 1 0.8 0.7 0.70 2 8.90 0.01 6.50E+06 8.64E+05 2.17E+06 140 Eskay Creek Agglomeration project First number in each batch # indicates 109 Tailings Data age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Ratio Strength Average Diam (m) Strengths (Pa) 1st 2nd 3rd lbs N Compressive Tensile Shear 719 8% 14.0% 1.75 2.6 2.6 2.35 0.90 148 658.56 0.03 5.07E+07 6.75E+06 1.69E+07 8% 14.0% 1.75 2.6 2.2 2.1 0.81 101 449.42 0.02 4.80E+07 6.39E+06 1.60E+07 8% 14.0% 1.75 2.4 2.2 2.1 0.88 127 565.12 0.02 5.26E+07 7.00E+06 1.75E+07 8% 14.0% 1.75 2.2 2 1.8 0.82 103 458.32 0.02 5.27E+07 7.01 E+06 1.76E+07 8% 14.0% 1.75 1.8 1.6 1.6 0.89 85 378.23 0.02 5.62E+07 7.47E+06 1.87E+07 8% 14.0% 1.75 1.6 1.6 1.5 0.94 93 413.83 0.02 5.97E+07 7.94E+06 1.99E+07 8% 14.0% 1.75 1.5 1.4 1.4 0.93 58 258.08 0.01 5.45E+07 7.25E+06 1.82E+07 8% 14.0% 1.75 1.5 1.4 1.3 0.87 48 213.59 0.01 5.18E+07 6.89E+06 1.73E+07 8% 14.0% 1.75 1 1 1 1.00 16 71.20 0.01 4.50E+07 5.98E+06 1.50E+07 8% 14.0% 1.75 1.2 1 1 0.83 35 155.74 0.01 5.65E+07 7.51 E+06 1.88E+07 8% 14.0% 1.75 1.1 1 0.8 0.73 26 115.69 0.01 5.38E+07 7.15E+06 1.79E+07 720 8% 16.0% 2.00 2.2 1.9 1.8 0.82 40 177.99 0.02 3.91 E+07 5.20E+06 1.30E+07 8% 16.0% 2.00 2 2 1.8 0.90 48 213.59 0.02 4.15E+07 5.53E+06 1.38E+07 8% 16.0% 2.00 2.3 1.9 1.8 0.78 50 222.49 0.02 4.18E+07 5.55E+06 1.39E+07 8% 16.0% 2.00 1.6 1.5 1.4 0.88 31 137.94 0.02 4.28E+07 5.69E+06 1.43E+07 8% 16.0% 2.00 1.6 1.6 1.6 1.00 34 151.29 0.02 4.23E+07 5.62E+06 1.41 E+07 8% 16.0% 2.00 1.9 1.6 1.6 0.84 32 142.39 0.02 4.02E+07 5.34E+06 1.34E+07 8% 16.0% 2.00 1.5 1.3 1.3 0.87 27 120.14 0.01 4.38E+07 5.83E+06 1.46E+07 8% 16.0% 2.00 1.4 1.3 1.2 0.86 26 115.69 0.01 4.44E+07 5.90E+06 1.48E+07 8% 16.0% 2.00 1 1 1 1.00 14 62.30 0.01 4.30E+07 5.72E+06 1.43E+07 8% 16.0% 2.00 1.3 1 1 0.77 15 66.75 0.01 4.19E+07 5.58E+06 1.40E+07 8% 16.0% 2.00 1.2 1 0.9 0.75 19 84.55 0.01 4.68E+07 6.23E+06 1.56E+07 707 8% 20.0% 2.50 3.5 2.7 2.4 0.69 54 240.29 0.03 3.37E+07 4.49E+06 1.12E+07 8% 20.0% 2.50 3.1 2.9 2.3 0.74 37 164.64 0.03 2.99E+07 3.98E+06 9.98E+06 8% 20.0% 2.50 3.4 2.8 2.3 0.68 54 240.29 0.03 3.37E+07 4.49E+06 1.12E+07 8% 20.0% 2.50 2.1 1.8 1.5 0.71 18 80.10 0.02 3.16E+07 4.20E+06 1.05E+07 8% 20.0% 2.50 2.8 2.2 2 0.71 27 120.14 0.02 3.07E+07 4.09E+06 1.02E+07 8% 20.0% 2.50 2.4 2 1.5 0.63 20 88.99 0.02 3.08E+07 4.09E+06 1.03E+07 8% 20.0% 2.50 1.6 1 1.15 0.72 8 35.60 0.01 3.18E+07 4.23E+06 1.06E+07 8% 20.0% 2.50 1.6 0.95 1.4 0.88 11 48.95 0.01 3.47E+07 4.61 E+06 1.16E+07 8% 20.0% 2.50 1.5 0.9 1.35 0.90 15 66.75 0.01 3.98E+07 5.29E+06 1.33E+07 708 8% 23.0% 2.88 2.6 2.5 2.1 0.81 23 102.34 0.02 2.81 E+07 3.74E+06 9.37E+06 8% 23.0% 2.88 3 3 2.4 0.80 26 115.69 0.03 2.63E+07 3.50E+06 8.77E+06 8% 23.0% 2.88 2.2 2.1 2.1 0.95 19 84.55 0.02 2.88E+07 3.83E+06 9.60E+06 8% 23.0% 2.88 2.3 2.1 1.8 0.78 22 97.89 0.02 3.07E+07 4.09E+06 1.02E+07 8% 23.0% 2.88 2.3 1.9 1.9 0.83 16 71.20 0.02 2.83E+07 3.77E+06 9.44E+06 8% 23.0% 2.88 1.9 1.7 1.5 0.79 9 40.05 0.02 2.61 E+07 3.47E+06 8.69E+06 8% 23.0% 2.88 1.7 1.5 1.4 0.82 8 35.60 0.02 2.69E+07 3.58E+06 8.98E+06 8% 23.0% 2.88 1.5 1.4 1.2 0.80 7 31.15 0.01 2.76E+07 3.67E+06 9.20E+06 8% 23.0% 2.88 1.2 1.1 1.1 0.92 3 13.35 0.01 2.38E+07 3.16E+06 7.93E+06 721 10% 14.9% 1.49 2.2 2 2 0.91 40 177.99 0.02 4.46E+07 5.94E+06 1.49E+07 10% 14.9% 1.49 2.3 2.2 1.9 0.83 48 213.59 0.02 4.59E+07 6.11E+06 1.53E+07 10% 14.9% 1.49 2.3 2 2 0.87 50 222.49 0.02 4.77E+07 6.34E+06 1.59E+07 10% 14.9% 1.49 1.7 1.5 1.4 0.82 31 137.94 0.02 4.99E+07 6.64E+06 1.66E+07 10% 14.9% 1.49 1.8 1.5 1.3 0.72 34 151.29 0.02 5.15E+07 6.85E+06 1.72E+07 10% 14.9% 1.49 1.8 1.5 1.3 0.72 32 142.39 0.02 5.05E+07 6.71 E+06 1.68E+07 10% 14.9% 1.49 1.4 1.3 1.3 0.93 27 120.14 0.01 5.24E+07 6.96E+06 1.75E+07 10% 14.9% 1.49 1.4 1.3 1.2 0.86 26 115.69 0.01 5.24E+07 6.97E+06 1.75E+07 10% 14.9% 1.49 1.2 1.1 1 0.83 14 62.30 0.01 4.76E+07 6.33E+06 1.59E+07 10% 14.9% 1.49 1.3 1 1 0.77 15 66.75 0.01 4.95E+07 6.58E+06 1.65E+07 10% 14.9% 1.49 1.1 1.1 1 0.91 19 84.55 0.01 5.35E+07 7.12E+06 1.78E+07 722 10% 17.0% 1.70 2.6 2.1 1.9 0.73 81 360.43 0.02 5.43E+07 7.22E+06 1.81 E+07 10% 17.0% 1.70 2 1.9 1.6 0.80 60 266.98 0.02 5.47E+07 7.28E+06 1.82E+07 10% 17.0% 1.70 2 1.9 1.9 0.95 58 258.08 0.02 5.27E+07 7.01 E+06 1.76E+07 10% 17.0% 1.70 1.7 1.5 1.4 0.82 55 244.74 0.02 6.04E+07 8.04E+06 2.01 E+07 10% 17.0% 1.70 1.9 1.6 1.5 0.79 44 195.79 0.02 5.32E+07 7.08E+06 1.77E+07 10% 17.0% 1.70 1.8 1.5 1.5 0.83 60 266.98 0.02 6.09E+07 8.10E+06 2.03E+07 10% 17.0% 1.70 1.4 1.2 1.1 0.79 29 129.04 0.01 5.65E+07 7.52E+06 1.88E+07 10% 17.0% 1.70 1.6 1.3 1.3 0.81 56 249.19 0.01 6.52E+07 8.67E+06 2.17E+07 10% 17.0% 1.70 1.1 1 0.9 0.82 21 93.44 0.01 5.81 E+07 7.73E+06 1.94E+07 10% 17.0% 1.70 1.1 1 1 0.91 37 164.64 0.01 6.90E+07 9.18E+06 . 2.30E+07 10% 17.0% 1.70 1.1 1 1 0.91 21 93.44 0.01 5.71 E+07 7.60E+06 1.90E+07 2809 8% 2.7 2.7 2.4 0.89 116 516.17 0.03 4.57E+07 6.08E+06 1.52E+07 8% 2.6 2.6 2.5 0.96 140 622.96 0.03 4.93E+07 6.56E+06 1.64E+07 141 Eskay Creek Agglomeration project 109 Tailings Data Brazilian test First number in each batch # indicates age of pellet at time of testing. 'B' indicates curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Ratio Strength Average Diam (m) Strengths (Pa) 1st 2nd 3rd lbs N Compressive Tensile Shear 8% 2.2 2.1 2 0.91 78 347.08 0.02 4.65E+07 6.18E+06 1.55E+07 8% 2.1 2 2 0.95 93 413.83 0.02 5.05E+07 6.72E+06 1.68E+07 8% 1.9 1.8 1.6 0.84 57 253.64 0.02 4.68E+07 6.23E+06 1.56E+07 8% 1.7 1.5 1.3 0.76 55 244.74 0.02 5.18E+07 6.89E+06 1.73E+07 8% 1.6 1.4 1.4 0.88 33 146.84 0.01 4.47E+07 5.94E+06 1.49E+07 8% 1.2 1.2 1.1 0.92 18 80.10 0.01 4.20E+07 5.59E+06 1.40E+07 8% 1.2 1.1 1.1 0.92 17 75.65 0.01 4.24E+07 5.64E+06 1.41 E+07 8% 1.1 1 0.9 0.82 19 84.55 0.01 4.76E+07 6.33E+06 1.59E+07 2910 8% 2.4 2.2 2.2 0.92 91 404.93 0.02 4.67E+07 6.22E+06 1.56E+07 8% 2.4 2.2 2.1 0.88 61 271.43 0.02 4.12E+07 5.48E+06 1.37E+07 8% 2.4 2.3 2.2 0.92 88 391.58 0.02 4.56E+07 6.06E+06 1.52E+07 8% 1.7 1.6 1.5 0.88 43 191.34 0.02 4.57E+07 6.08E+06 1.52E+07 8% 1.8 1.6 1.5 0.83 55 244.74 0.02 4.91 E+07 6.53E+06 1.64E+07 8% 1.7 1.7 1.5 0.88 38 169.09 0.02 4.30E+07 5.71 E+06 1.43E+07 8% 1.4 1.4 1.3 0.93 20 88.99 0.01 3.92E+07 5.21 E+06 1.31 E+07 8% 1.6 1.3 1.2 0.75 42 186.89 0.01 5.08E+07 6.75E+06 1.69E+07 8% 1.3 1.3 1 0.77 34 151.29 0.01 5.05E+07 6.72E+06 1.68E+07 8% 1 0.9 0.9 0.90 14 62.30 0.01 4.53E+07 6.02E+06 1.51 E+07 8% 1 0.9 0.9 0.90 16 71.20 0.01 4.74E+07 6.30E+06 1.58E+07 2811 6% 11.8% 1.97 1.7 1.7 1.4 0.82 16 71.20 0.02 2.32E+07 3.09E+06 7.74E+06 6% 11.8% 1.97 2.4 1.5 1.5 0.63 18 80.10 0.02 2.32E+07 3.08E+06 7.73E+06 6% 11.8% 1.97 1.4 1.4 1.4 1.00 14 62.30 0.01 2.45E+07 3.26E+06 8.17E+06 6% 11.8% 1.97 1.5 1.4 1.3 0.87 18 80.10 0.01 2.67E+07 3.55E+06 8.89E+06 6% 11.8% 1.97 1.4 1.3 1.2 0.86 14 62.30 0.01 2.58E+07 3.43E+06 8.59E+06 6% 11.8% 1.97 1.1 1.1 1 0.91 12 53.40 0.01 2.78E+07 3.69E+06 9.26E+06 6% 11.8% 1.97 1.2 1.1 1 0.83 12 53.40 0.01 2.73E+07 3.64E+06 9.12E+06 6% 11.8% 1.97 1.2 1.1 1.1 0.92 11 48.95 0.01 2.62E+07 3.48E+06 8.72E+06 6% 11.8% 1.97 0.9 0.9 0.9 1.00 13 57.85 0.01 3.21 E+07 4.27E+06 1.07E+07 6% 11.8% 1.97 1 0.9 0.8 0.80 7 31.15 0.01 2.61 E+07 3.47E+06 8.71 E+06 2812 6% 14.0% 2.33 1.9 1.8 1.7 0.89 17 75.65 0.02 2.21 E+07 2.94E+06 7.37E+06 6% 14.0% 2.33 1.9 1.8 1.7 0.89 20 88.99 0.02 2.34E+07 3.11 E+06 7.78E+06 6% 14.0% 2.33 1.9 1.8 1.7 0.89 15 66.75 0.02 2.12E+07 2.82E+06 7.07E+06 6% 14.0% 2.33 1.6 1.5 1.2 0.75 12 53.40 0.01 2.27E+07 3.03E+06 7.58E+06 6% 14.0% 2.33 1.5 1.4 1.2 0.80 13 57.85 0.01 2.42E+07 3.22E+06 8.07E+06 6% 14.0% 2.33 1.5 1.4 1.4 0.93 14 62.30 0.01 2.42E+07 3.22E+06 8.08E+06 6% 14.0% 2.33 1.2 1.2 1.1 0.92 11 48.95 0.01 2.54E+07 3.38E+06 8.47E+06 6% 14.0% 2.33 1.3 1.1 1.1 0.85 8 35.60 0.01 2.32E+07 3.08E+06 7.73E+06 6% 14.0% 2.33 1 0.9 0.9 0.90 3 13.35 0.01 1.93E+07 2.57E+06 6.45E+06 6% 14.0% 2.33 1 0.9 0.8 0.80 9 40.05 0.01 2.84E+07 3.78E+06 9.47E+06 2813 6% 16.4% 2.73 3.1 2.7 2.3 0.74 24 106.79 0.03 1.89E+07 2.52E+06 6.31 E+06 6% 16.4% 2.73 2.3 2.1 1.5 0.65 21 93.44 0.02 2.21 E+07 2.94E+06 7.37E+06 6% 16.4% 2.73 1.9 1.8 1.4 0.74 17 75.65 0.02 2.28E+07 3.03E+06 7.59E+06 6% 16.4% 2.73 1.9 1.6 1.5 0.79 12 53.40 0.02 2.09E+07 2.78E+06 6.96E+06 6% 16.4% 2.73 1.8 1.5 1.3 0.72 9 40.05 0.02 2.00E+07 2.66E+06 6.66E+06 6% 16.4% 2.73 1.6 1.5 1.2 0.75 12 53.40 0.01 2.27E+07 3.03E+06 7.58E+06 6% 16.4% 2.73 1.6 1.4 1.3 0.81 11 48.95 0.01 2.24E+07 2.97E+06 7.45E+06 6% 16.4% 2.73 1.4 1.3 1.3 0.93 14 62.30 0.01 2.54E+07 3.38E+06 8.48E+06 6% 16.4% 2.73 1.4 1.3 1.2 0.86 15 66.75 0.01 2.64E+07 3.51 E+06 8.79E+06 6% 16.4% 2.73 1.3 1.1 1 0.77 18 80.10 0.01 3.08E+07 4.10E+06 1.03E+07 6% 16.4% 2.73 1.1 1 1 0.91 18 80.10 0.01 3.28E+07 4.36E+06 1.09E+07 2814 4% 16.8% 4.20 2.9 2.6 2.5 0.86 32 142.39 0.03 1.27E+07 1.69E+06 4.24E+06 4% 16.8% 4.20 2.6 2.4 2.3 0.88 24 106.79 0.02 1.23E+07 1.63E+06 4.09E+06 4% 16.8% 4.20 2.4 2.2 1.9 0.79 31 137.94 0.02 1.44E+07 1.91 E+06 4.79E+06 4% 16.8% 4.20 2.3 2 1.9 0.83 22 97.89 0.02 1.33E+07 1.77E+06 4.44E+06 4% 16.8% 4.20 2.1 2 1.8 0.86 24 106.79 0.02 1.41 E+07 1.87E+06 4.69E+06 4% 16.8% 4.20 1.9 1.8 1.6 0.84 15 66.75 0.02 1.29E+07 1.72E+06 4.30E+06 4% 16.8% 4.20 2 1.7 1.5 0.75 18 80.10 0.02 1.40E+07 1.86E+06 4.66E+06 4% 16.8% 4.20 1.7 1.6 1.5 0.88 8 35.60 0.02 1.12E+07 1.49E+06 3.74E+06 4% 16.8% 4.20 1.7 1.5 1.1 0.65 12 53.40 0.01 1.37E+07 1.82E+06 4.57E+06 4% 16.8% 4.20 1.5 1.3 1.1 0.73 14 62.30 0.01 1.55E+07 2.07E+06 5.18E+06 4% 16.8% 4.20 1.3 1.1 1 0.77 10 44.50 0.01 1.53E+07 2.03E+06 5.10E+06 2819 8% 14.0% 1.75 2.4 2.3 2.1 0.88 216 961.14 0.02 6.19E+07 8.23E+06 2.06E+07 8% 14.0% 1.75 2.2 2.1 1.9 0.86 174 774.25 0.02 6.12E+07 8.14E+06 2.04E+07 8% 14.0% 1.75 2.2 1.8 1.7 0.77 127 565.12 0.02 5.90E+07 7.85E+06 1.97E+07 142 Eskay Creek Agglomeration project First number in each batch # indicates 109 Tailings Data age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Strength Average Strengths (Pa) 1st 2nd 3rd Ratio lbs N Diam (m) Compressive Tensile 8% 14.0% 1.75 2 1.8 1.6 0.80 114 507.27 0.02 5.85E+07 7.78E+06 1.95E+07 8% 14.0% 1.75 1.8 1.7 1.7 0.94 156 694.16 0.02 6.68E+07 8.88E+06 2.23E+07 8% 14.0% 1.75 1.8 1.5 1.4 0.78 110 489.47 0.02 6.38E+07 8.49E+06 2.13E+07 8% 14.0% 1.75 1.5 1.4 1.3 0.87 73 324.83 0.01 5.96E+07 7.93E+06 1.99E+07 8% 14.0% 1.75 1.4 1.3 1.2 0.86 68 302.58 0.01 6.11 E+07 8.13E+06 2.04E+07 8% 14.0% 1.75 1.5 1.2 1.2 0.80 79 351.53 0.01 6.51 E+07 8.66E+06 2.17E+07 8% 14.0% 1.75 1.2 1.1 1 0.83 56 249.19 0.01 6.41 E+07 8.52E+06 2.14E+07 8% 14.0% 1.75 1 0.9 0.8 0.80 30 133.49 0.01 5.95E+07 7.91E+06 1.98E+07 2920 8% 16.0% 2.00 2.3 1.9 1.6 0.70 41 182.44 0.02 3.98E+07 5.29E+06 1.33E+07 8% 16.0% 2.00 1.9 1.8 1.6 0.84 27 120.14 0.02 3.65E+07 4.86E+06 1.22E+07 8% 16.0% 2.00 1.9 1.8 1.4 0.74 23 102.34 0.02 3.53E+07 4.69E+06 1.18E+07 8% 16.0% 2.00 2.2 1.7 1.5 0.68 42 186.89 0.02 4.23E+07 5.63E+06 1.41 E+07 8% 16.0% 2.00 1.8 1.6 1.4 0.78 44 195.79 0.02 4.60E+07 6.12E+06 1.53E+07 8% 16.0% 2.00 1.5 1.4 1.2 0.80 19 84.55 0.01 3.85E+07 5.12E+06 1.28E+07 8% 16.0% 2.00 1.5 1.3 1.2 0.80 24 106.79 0.01 4.27E+07 5.67E+06 1.42E+07 8% 16.0% 2.00 1.5 1.3 1.2 0.80 23 102.34 0.01 4.21 E+07 5.59E+06 1.40E+07 8% 16.0% 2.00 1.4 1.2 1.1 0.79 26 115.69 0.01 4.62E+07 6.14E+06 1.54E+07 8% 16.0% 2.00 1.2 1.1 1 0.83 31 137.94 0.01 5.26E+07 7.00E+06 1.75E+07 8% 16.0% 2.00 1 0.9 0.8 0.80 13 57.85 0.01 4.50E+07 5.99E+06 1.50E+07 2921 10% 14.9% 1.49 2.2 2 1.7 0.77 63 280.33 0.02 5.32E+07 7.08E+06 1.77E+07 10% 14.9% 1.49 2 1.9 1.8 0.90 88 391.58 0.02 6.11 E+07 8.12E+06 2.04E+07 10% 14.9% 1.49 2 1.8 1.5 0.75 67 298.13 0.02 5.83E+07 7.76E+06 1.94E+07 10% 14.9% 1.49 1.8 1.7 1.3 0.72 35 155.74 0.02 4.98E+07 6.63E+06 1.66E+07 10% 14.9% 1.49 1.6 1.5 1.4 0.88 43 191.34 0.02 5.63E+07 7.49E+06 1.88E+07 10% 14.9% 1.49 1.7 1.4 1.1 0.65 33 146.84 0.01 5.40E+07 7.18E+06 1.80E+07 10% 14.9% 1.49 1.5 1.4 1 0.67 40 177.99 0.01 5.97E+07 7.94E+06 1.99E+07 10% 14.9% 1.49 1.5 1.3 1.2 0.80 36 160.19 0.01 5.76E+07 7.67E+06 1.92E+07 10% 14.9% 1.49 1.4 1.2 1.1 0.79 18 80.10 0.01 4.82E+07 6.41 E+06 1.61 E+07 10% 14.9% 1.49 1.2 1.1 1 0.83 21 93.44 0.01 5.45E+07 7.25E+06 1.82E+07 10% 14.9% 1.49 1 0.9 0.8 0.80 30 133.49 0.01 7.02E+07 9.34E+06 2.34E+07 2922 10% 17.0% 1.70 2.3 1.8 1.7 0.74 63 280.33 0.02 5.46E+07 7.26E+06 1.82E+07 10% 17.0% 1.70 2 1.8 1.5 0.75 59 262.53 0.02 5.59E+07 7.44E+06 1.86E+07 10% 17.0% 1.70 2 1.7 1.5 0.75 57 253.64 0.02 5.63E+07 7.49E+06 1.88E+07 10% 17.0% 1.70 1.8 1.6 1.5 0.83 44 195.79 0.02 5.38E+07 7.15E+06 1.79E+07 10% 17.0% 1.70 1.8 1.5 1.3 0.72 52 231.39 0.02 5.93E+07 7.89E+06 1.98E+07 10% 17.0% 1.70 1.6 1.5 1.3 0.81 49 218.04 0.01 5.95E+07 7.91 E+06 1.98E+07 10% 17.0% 1.70 1.5 1.4 1.3 0.87 46 204.69 0.01 6.03E+07 8.02E+06 2.01 E+07 10% 17.0% 1.70 1.4 1.3 1.1 0.79 43 191.34 0.01 6.27E+07 8.34E+06 2.09E+07 10% 17.0% 1.70 1.4 1.2 1 0.71 42 186.89 0.01 6.48E+07 8.62E+06 2.16E+07 10% 17.0% 1.70 1.1 1 1 0.91 40 177.99 0.01 7.08E+07 9.42E+06 2.36E+07 10% 17.0% 1.70 0.9 0.9 0.8 0.89 23 102.34 0.01 6.55E+07 8.71 E+06 2.18E+07 3309B 8% 2.5 2.3 2.2 0.88 222 987.84 0.02 6.16E+07 8.19E+06 2.05E+07 8% 2.5 2.4 2.2 0.88 199 885.50 0.02 5.85E+07 7.78E+06 1.95E+07 8% 2 1.9 1.9 0.95 140 622.96 0.02 5.99E+07 7.96E+06 2.00E+07 8% 2.3 2 1.9 0.83 128 569.57 0.02 5.57E+07 7.41 E+06 1.86E+07 8% 2.1 2 1.7 0.81 131 582.92 0.02 5.81 E+07 7.72E+06 1.94E+07 8% 1.6 1.5 1.4 0.88 66 293.68 0.02 5.50E+07 7.32E+06 1.83E+07 8% 1.5 1.4 1.2 0.80 48 213.59 0.01 5.24E+07 6.98E+06 1.75E+07 8% 1.5 1.5 1.3 0.87 69 307.03 0.01 5.71 E+07 7.60E+06 1.90E+07 8% 1.1 1 1 0.91 20 88.99 0.01 4.76E+07 6.34E+06 1.59E+07 8% 1 0.9 0.9 0.90 19 84.55 0.01 5.02E+07 6.67E+06 1.67E+07 8% 1 1 0.9 0.90 13 57.85 0.01 4.27E+07 5.68E+06 1.42E+07 311 OB 6% 2.3 2.3 2.1 0.91 152 676.36 0.02 3.96E+07 5.26E+06 1.32E+07 6% 2.5 2.3 2.1 0.84 140 622.96 0.02 3.79E+07 5.05E+06 1.26E+07 6% 2.3 2.3 2.2 0.96 109 485.02 0.02 3.52E+07 4.68E+06 1.17E+07 6% 1.9 1.7 1.6 0.84 83 369.33 0.02 3.86E+07 5.13E+06 1.29E+07 6% 1.8 1.6 1.5 0.83 55 244.74 0.02 3.50E+07 4.66E+06 1.17E+07 6% 1.6 1.6 1.5 0.94 43 191.34 0.02 3.29E+07 4.38E+06 1.10E+07 6% 1.3 1.2 1.2 0.92 44 195.79 0.01 3.93E+07 5.22E+06 1.31 E+07 6% 1.5 1.3 1.3 0.87 42 186.89 0.01 3.62E+07 4.82E+06 1.21 E+07 6% 1.4 1.3 1.2 0.86 35 155.74 0.01 3.50E+07 4.65E+06 1.17E+07 6% 0.9 0.9 0.8 0.89 20 88.99 0.01 3.78E+07 5.02E+06 1.26E+07 6% 1 0.9 0.8 0.80 18 80.10 0.01 3.58E+07 4.76E+06 1.19E+07 3011B 6% 11.8% 1.97 1.6 1.5 1.3 0.81 19 84.55 0.01 2.62E+07 3.49E+06 8.74E+06 143 Eskay Creek Agglomeration project 109 Tailings Data Brazilian test First number in each batch # indicates age of pellet at time of testing. 'B' indicates curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Strength Average Strengths (Pa) 1st 2nd 3rd Ratio lbs N Diam (m) Compressive Tensile Shear 6% 11.8% 1.97 1.5 1.4 1.3 0.87 8 35.60 0.01 2.03E+07 2.71 E+06 6.78E+06 6% 11.8% 1.97 1.5 1.2 1.2 0.80 10 44.50 0.01 2.33E+07 3.10E+06 7.77E+06 6% 11.8% 1.97 1.4 1.4 1.2 0.86 8 35.60 0.01 2.08E+07 2.77E+06 6.95E+06 6% 11.8% 1.97 1.3 1.1 1.1 0.85 16 71.20 0.01 2.92E+07 3.89E+06 9.74E+06 6% 11.8% 1.97 1.3 1.2 1.1 0.85 10 44.50 0.01 2.43E+07 3.23E+06 8.10E+06 6% 11.8% 1.97 1 0.9 0.9 0.90 5 22.25 0.01 2.29E+07 3.05E+06 7.64E+06 6% 11.8% 1.97 1 1 0.9 0.90 4 17.80 0.01 2.06E+07 2.73E+06 6.85E+06 3012B 6% 14.0% 2.33 2 1.9 1.8 0.90 30 133.49 0.02 2.58E+07 3.43E+06 8.59E+06 6% 14.0% 2.33 2.1 2 1.7 0.81 52 231.39 0.02 3.04E+07 4.05E+06 1.01 E+07 6% 14.0% 2.33 2.2 2 1.7 0.77 43 191.34 0.02 2.83E+07 3.77E+06 9.44E+06 6% 14.0% 2.33 1.6 1.5 1.5 0.94 13 57.85 0.02 2.26E+07 3.00E+06 7.53E+06 6% 14.0% 2.33 1.7 1.6 1.4 0.82 15 66.75 0.02 2.32E+07 3.08E+06 7.73E+06 6% 14.0% 2.33 1.6 1.5 1.5 0.94 17 75.65 0.02 2.47E+07 3.29E+06 8.24E+06 6% 14.0% 2.33 1.3 1.2 1.1 0.85 11 48.95 0.01 2.51 E+07 3.33E+06 8.36E+06 6% 14.0% 2.33 1.2 1.1 1 0.83 7 31.15 0.01 2.29E+07 3.04E+06 7.62E+06 6% 14.0% 2.33 1.4 1.2 1.1 0.79 6 26.70 0.01 2.02E+07 2.69E+06 6.73E+06 6% 14.0% 2.33 1 1 0.9 0.90 7 31.15 0.01 2.48E+07 3.29E+06 8.26E+06 3213B 6% 16.4% 2.73 1.9 1.8 1.3 0.68 26 115.69 0.02 2.65E+07 3.52E+06 8.83E+06 6% 16.4% 2.73 1.8 1.6 1.4 0.78 21 93.44 0.02 2.57E+07 3.41 E+06 8.56E+06 6% 16.4% 2.73 1.7 1.6 1.4 0.82 18 80.10 0.02 2.46E+07 3.28E+06 8.21 E+06 6% 16.4% 2.73 1.7 1.5 1.4 0.82 30 133.49 0.02 2.99E+07 3.97E+06 9.95E+06 6% 16.4% 2.73 1.6 1.5 1.3 0.81 19 84.55 0.01 2.62E+07 3.49E+06 8.74E+06 6% 16.4% 2.73 1.5 1.4 1.3 0.87 14 62.30 0.01 2.45E+07 3.26E+06 8.17E+06 6% 16.4% 2.73 1.4 1.3 1.1 0.79 18 80.10 0.01 2.84E+07 3.77E+06 9.46E+06 6% 16.4% 2.73 1.2 1.1 1 0.83 19 84.55 0.01 3.19E+07 4.24E+06 1.06E+07 6% 16.4% 2.73 1.1 1 0.9 0.82 19 84.55 0.01 3.40E+07 4.52E+06 1.13E+07 6% 16.4% 2.73 1 0.8 0.8 0.80 11 48.95 0.01 3.16E+07 4.20E+06 1.05E+07 6% 16.4% 2.73 1 0.9 0.8 0.80 10 44.50 0.01 2.94E+07 3.91 E+06 9.81 E+06 3214B 4% 16.8% 4.20 2.9 2.7 2.4 0.83 28 124.59 0.03 1.21 E+07 1.61 E+06 4.03E+06 4% 16.8% 4.20 2.6 2.4 2 0.77 32 142.39 0.02 1.38E+07 1.83E+06 4.59E+06 4% 16.8% 4.20 2.3 2.2 2.2 0.96 36 160.19 0.02 1.49E+07 1.98E+06 4.96E+06 4% 16.8% 4.20 2.1 2 1.8 0.86 19 84.55 0.02 1.30E+07 1.73E+06 4.34E+06 4% 16.8% 4.20 2 1.9 1.7 0.85 16 71.20 0.02 1.27E+07 1.69E+06 4.24E+06 4% 16.8% 4.20 1.9 1.7 1.6 0.84 13 57.85 0.02 1.25E+07 1.67E+06 4.18E+06 4% 16.8% 4.20 1.9 1.6 1.5 0.79 19 84.55 0.02 1.47E+07 1.95E+06 4.89E+06 4% 16.8% 4.20 1.8 1.5 1.4 0.78 17 75.65 0.02 1.47E+07 1.96E+06 4.91 E+06 4% 16.8% 4.20 1.6 1.4 1.2 0.75 14 62.30 0.01 1.48E+07 1.97E+06 4.93E+06 4% 16.8% 4.20 1.4 1.3 1.2 0.86 10 44.50 0.01 1.39E+07 1.85E+06 4.63E+06 4% 16.8% 4.20 1 1 0.9 0.90 8 35.60 0.01 1.56E+07 2.08E+06 5.20E+06 2915B 4% 14.3% 3.58 1.9 1.8 1.7 0.89 17 75.65 0.02 1.33E+07 1.77E+06 4.45E+06 4% 14.3% 3.58 1.9 1.7 1.4 0.74 11 48.95 0.02 1.21 E+07 1.61 E+06 4.03E+06 4% 14.3% 3.58 1.7 1.5 1.4 0.82 13 57.85 0.02 1.36E+07 1.81 E+06 4.54E+06 4% 14.3% 3.58 1.7 1.5 1.3 0.76 14 62.30 0.02 1.41 E+07 1.88E+06 4.71 E+06 4% 14.3% 3.58 1.5 1.4 1.3 0.87 11 48.95 0.01 1.36E+07 1.81 E+06 4.55E+06 4% 14.3% 3.58 1.4 1.4 1.2 0.86 8 35.60 0.01 1.26E+07 1.67E+06 4.19E+06 4% 14.3% 3.58 1.5 1.3 1.2 0.80 9 40.05 0.01 1.32E+07 1.76E+06 4.41 E+06 4% 14.3% 3.58 1.3 1.3 1 0.77 9 40.05 0.01 1.39E+07 1.85E+06 4.65E+06 4% 14.3% 3.58 1.2 1.1 1 0.83 6 26.70 0.01 1.31 E+07 1.74E+06 4.36E+06 4% 14.3% 3.58 1.1 1 0.9 0.82 6 26.70 0.01 1.39E+07 1.85E+06 4.65E+06 4% 14.3% 3.58 1 0.9 0.8 0.80 5 22.25 0.01 1.41 E+07 1.87E+06 4.69E+06 2916B 4% 14.5% 3.63 2 1.9 1.6 0.80 14 62.30 0.02 1.23E+07 1.63E+06 4.09E+06 4% 14.5% 3.63 2.2 1.7 1.7 0.77 16 71.20 0.02 1.29E+07 1.72E+06 4.32E+06 4% 14.5% 3.63 2 1.6 1.4 0.70 17 75.65 0.02 1.41 E+07 1.88E+06 4.71 E+06 4% 14.5% 3.63 1.9 1.6 1.5 0.79 14 62.30 0.02 1.32E+07 1.76E+06 4.42E+06 4% 14.5% 3.63 1.7 1.6 1.5 0.88 17 75.65 0.02 1.44E+07 1.92E+06 4.81 E+06 4% 14.5% 3.63 1.8 1.4 1.3 0.72 12 53.40 0.01 1.36E+07 1.80E+06 4.52E+06 4% 14.5% 3.63 1.5 1.3 1.2 0.80 10 44.50 0.01 1.37E+07 1.82E+06 4.57E+06 4% 14.5% 3.63 1.3 1.2 1.2 0.92 10 44.50 0.01 1.44E+07 1.92E+06 4.81 E+06 4% 14.5% 3.63 1.3 1.2 1.1 0.85 13 57.85 0.01 1.60E+07 2.13E+06 5.33E+06 4% 14.5% 3.63 1.1 1 0.9 0.82 9 40.05 0.01 1.60E+07 2.12E+06 5.32E+06 4% 14.5% 3.63 1 0.8 0.7 0.70 5 22.25 0.01 1.49E+07 1.98E+06 4.97E+06 2817B 3% 14.0% 4.67 2.8 2.5 2.1 0.75 18 80.10 0.02 6.50E+06 8.64E+05 2.17E+06 3% 14.0% 4.67 2.3 2 1.8 0.78 9 40.05 0.02 5.90E+06 7.84E+05 1.97E+06 3% 14.0% 4.67 1.9 1.8 1.7 0.89 6 26.70 0.02 5.57E+06 7.41 E+05 1.86E+06 144 Eskay Creek Agglomeration project First number in each batch # indicates 109 Tailings Data age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Ratio Strength Average Diam (m) Strengths (Pa) 1st 2nd 3rd lbs N Compressive Tensile Shear 3% 14.0% 4.67 1.8 1.7 1.5 0.83 16 71.20 0.02 8.11 E+06 1.08E+06 2.70E+06 3% 14.0% 4.67 1.6 1.6 1.5 0.94 10 44.50 0.02 7.22E+06 9.60E+05 2.41 E+06 3% 14.0% 4.67 1.6 1.5 1.4 0.88 11 48.95 0.02 7.70E+06 1.02E+06 2.57E+06 3% 14.0% 4.67 1.6 1.5 1.2 0.75 7 31.15 0.01 6.77E+06 9.01 E+05 2.26E+06 3% 14.0% 4.67 1.5 1.3 1.2 0.80 7 31.15 0.01 7.19E+06 9.57E+05 2.40E+06 3% 14.0% 4.67 1.4 1.3 1.2 0.86 4 17.80 0.01 6.05E+06 8.04E+05 2.02E+06 3% 14.0% 4.67 1.3 1.1 1 0.77 6 26.70 0.01 7.62E+06 1.01 E+06 2.54E+06 3% 14.0% 4.67 1.2 1 0.9 0.75 4 17.80 0.01 7.08E+06 9.42E+05 2.36E+06 2818B 3% 14.0% 4.67 2.6 2.5 2.4 0.92 21 93.44 0.03 6.80E+06 9.04E+05 2.27E+06 3% 14.0% 4.67 2.6 2.3 2 0.77 18 80.10 0.02 6.82E+06 9.08E+05 2.27E+06 3% 14.0% 4.67 2.6 2 1.9 0.73 20 88.99 0.02 7.45E+06 9.91 E+05 2.48E+06 3% 14.0% 4.67 2 1.9 1.8 0.90 17 75.65 0.02 7.60E+06 1.01 E+06 2.53E+06 3% 14.0% 4.67 1.9 1.7 1.5 0.79 10 44.50 0.02 6.86E+06 9.13E+05 2.29E+06 3% 14.0% 4.67 1.7 1.6 1.5 0.88 11 48.95 0.02 7.38E+06 9.81 E+05 2.46E+06 3% 14.0% 4.67 1.6 1.5 1.3 0.81 9 40.05 0.01 7.28E+06 9.69E+05 2.43E+06 3% 14.0% 4.67 1.7 1.4 1.3 0.76 9 40.05 0.01 7.37E+06 9.80E+05 2.46E+06 3% 14.0% 4.67 1.5 1.4 1.3 0.87 7 31.15 0.01 6.93E+06 9.22E+05 2.31 E+06 3% 14.0% 4.67 1.3 1.2 1.1 0.85 6 26.70 0.01 7.30E+06 9.71 E+05 2.43E+06 3% 14.0% 4.67 1.1 1.1 1 0.91 3 13.35 0.01 6.24E+06 8.29E+05 2.08E+06 3919B 8% 14.0% 1.75 2.3 2.3 2 0.87 168 747.56 0.02 5.78E+07 7.68E+06 1.93E+07 8% 14.0% 1.75 2.4 2.3 2.1 0.88 256 ###### 0.02 6.55E+07 8.71 E+06 2.18E+07 8% 14.0% 1.75 2.2 2.1 1.9 0.86 136 605.16 0.02 5.64E+07 7.50E+06 1.88E+07 8% 14.0% 1.75 2 1.9 1.6 0.80 153 680.81 0.02 6.33E+07 8.42E+06 2.11 E+07 8% 14.0% 1.75 1.8 1.8 1.7 0.94 198 881.05 0.02 7.09E+07 9.44E+06 2.36E+07 8% 14.0% 1.75 1.5 1.4 1.2 0.80 99 440.52 0.01 6.68E+07 8.88E+06 2.23E+07 8% 14.0% 1.75 1.5 1.3 1.2 0.80 78 347.08 0.01 6.32E+07 8.41 E+06 2.11 E+07 8% 14.0% 1.75 1.4 1.3 1.2 0.86 88 391.58 0.01 6.66E+07 8.86E+06 2.22E+07 8% 14.0% 1.75 1.2 1.2 0.9 0.75 91 404.93 0.01 7.42E+07 9.87E+06 2.47E+07 8% 14.0% 1.75 1.2 1.1 1 0.83 56 249.19 0.01 6.41 E+07 8.52E+06 2.14E+07 8% 14.0% 1.75 1.1 1 0.9 0.82 52 231.39 0.01 6.66E+07 8.86E+06 2.22E+07 145 CO co -BI ° c E ca 3 3> 3 ? CO I f CM CO o o ci 5 LO CO CM LO cd o CO LO ci HI p is o o ci o o ci E S 3 ? LO cd CM 31 IW Is CO o co CO o o d o co co o ro cd co co CM m CM CM 1^  o ro cd eg o I-3 B o -\ O 3 B.-S < CO « 5 a 0 a) °> 6 - = u co = co .E H •* « cn 0 1 o UJ CS r-E S 3 ? co r- CO 3 CO CD .a 3 ? CO CO -SI 5 i i c 'eg 3 ? CO o o ci CO o o ci co co ro co LO cd ro co cd ro LO CD CM \4L cum % passing 100.001 98.08| 96.701 90.66I 80.04I 15.301 0.00| cum % passing 100.00| 100.001 100.001 96.281 71.851 9.081 o.ool cum % passing 100.00| 100.00| 98.07| 92.44| 79.88| 9.191 0.00| cum % retain 0.00 1.92 3.30 9.34 19.96 84.70: 100.00| cum % retain 0.001 o.ool o.ool 3.721 28.151 90.921 100.00| cum % retain o.ool o.ool 1.93| 7.561 20.121 90.811 100.00| % retain 0.00 1.92 1.38 6.04 10.63 64.74| 15.30 % retain 0.00| o.oo| 0.00| 3.721 24.431 62.781 9.08| % retain 0.00| o.ool 1.931 5.631 12.57| 70.681 9.19| wt(g) o 46.11 33.04 144.85 255.00 1553.40| 367.001 2399.401 % o 0.00| 0.00| 97.951 642.75| 1651.80| 238.80| 2631.30| wt(g) o o.ool 50.201 146.51| 327.19| 1840.40| 239.401 2603.701 cumulative wt(g) o 46.11 79.15 CM CM cn 2032.4| 2399.41 cumulative wt(g) o o o 97.95| 740.71 2392.5| 2631.31 cumulative wt(g) o o 50.21 196.71| 523.91 2364.3| 2603.71 (mm) o LO 26.5 22.4 cn 13.5 4.76 o 8%,16%W (mm) o CO 26.51 22.41 cn 13.5| 4.76| o 10%, 14.9% W (mm) 09 26.51 22.41 o> 13.5| 4.761 o Screen Size (mesh) CM 11.06" |7/8" |3/4" |0.530" 3 |#4 minus |Total | 109 #20 Screen Size (mesh) CM |1.06" \ |7/8" | |3/4" | |0.530" | |#4 minus | |Total \ 109 #21 Screen Size (mesh) CM |1.06" | |7/8" | |3/4" | |0.530" | * | #4 minus | Total S S = 8 .2 3 <o O II O 3 O) ~ O) < CO "55 = co lis ro cn o | 3 3 ? CO ra 3 I c CD CO CO £.1 i a, E 3 3 $ 3 % CD N co c CD CD co £-1 cn LO cn o c> CM cri co o o ci o CO CM CM CM 3 CO o ci o o d o co CM 05 CM LO CM o CO CM CO 3 ? LO CO CO 1^ CO 03 d LO CO co c 0) 0) CO £.1 146 NEX PELLETIZATION TESTING RESULTS X ! CD a < 500.00 450.00 400.00 350.00 300.00 250.00 200.00 150.00 100.00 50.00 0.00 Force at Failure vs. Size wrt % Cement 3 Day NEX Test __ » 4 % Cement __ B 6 % Cement A 8% Cement x 10% Cement X A A > T x A A^P*1^ WW x • ^ — ' • "~^Z~— -) p r r z 1 1 1 1 0.5 CD > '55 £ CN c" co E O CD Z c 0) ° c co i X CD 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 1 1.5 2 Median Size (cm) Strength vs. Size wrt % Cement 3 Day Test NEX 2.5 • 4% Cement • 6% Cement • 8% Cement x 10% Cement -4 • » » 0.5 1.5 Size (cm) 2.5 CD > CO (0 £ 0 CD Z " S o c W 1 & 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 Strength vs. Aspect Ratio wrt % Cement 3 Day Test NEX x x A A A •1 1 + • 4% Cement • 6% Cement A 8% Cement x 10% Cement 0.90 1.00 1.10 1.20 1.30 1.40 1.50 Aspect Ratio (Dim 1/Dim 3) 1.60 1.70 CO i_ D ' r a LL +-* CO CD o TJ O Q. < 500.00 450.00 400.00 350.00 300.00 250.00 200.00 150.00 100.00 50.00 0.00 Force at Failure vs. Size wrt % Cement 7 Day NEX Test _. • 4% Cement -. B 6 % Cement -- A 8% Cement -• x 10% Cement 1—— 1- i 1 1 0.5 1 1.5 2 Median Size (cm) > 'cn co 01 Q. -C E o> o c O o> |5> E x co 7.00E+07 6.00E+07 ^ 5.00E+07 | 4.00E+07 ° 3.00E+07 ra 0. w 2.00E+07 1.00E+07 0.00E+00 Strength vs. Size wrt % Cement 7 Day Test NEX 0.5 1 1.5 2 Median Size (cm) 2.5 • 4% Cement • • 6% Cement A 8% Cement X 21 K X x 10% Cement a 1- T X A T -*• A A A 1 f A X X Z A A A • - • • * • • • • v .4. -• • 1 1 i i i 2.5 > 'to (0 S: CM a. .c r o c z " S> S | 55 « E x ro 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 Strength vs. Aspect Ratio wrt % Cement 7 Day Test NEX A 3 Z • 4% Cement • 6% Cement A 8% Cement x 10% Cement 0.90 1.00 1.10 1.20 1.30 1.40 Aspect Ratio (Dim 1/Dim 3) 1.50 1.60 151 _ TO LL _• CO <D o •o as Q. < 800.00 700.00 600.00 500.00 400.00 300.00 200.00 100.00 0.00 Force at Failure vs. Size wrt % Cement 28 Day NEX Test • 4% Cement Dry • 6% Cement Wet •j A 8% Cement Wet A 8% Cement Dry x 10% Cement Wet 310% Cement Dry A a y / _/ y S yS* a uf^S^ a ^ j ^ K ^ ^ ^ ^ (0 > M (0 S. CM O. £ c C ** _ • i If z <•> 2 S rs 8.00E+07 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 0.5 1 1.5 2 Median Size (cm) Strength vs. Size wrt % Cement 28 Day Test NEX 2.5 • 4% Cement Dry • 6% Cement Wet " A 8% Cement Wet A 8% Cement Dry x 10% Cement Wet a 10% Cement Dry ^ a - ^ a - ^ 3 ^ ^ ^ ^ 3 v a a ^ x a a A * v A—y A A x A . A S . _____ x A A ' * A m _ • • _ * _.. 1 1 1 1 1 0.5 1 1.5 2 Median Size (cm) 2.5 cn c E co E I ro > <fl 0) (0 !& E o o 8.00E+07 7.00E+07 6.00E+07 5.00E+07 4.00E+07 3.00E+07 2.00E+07 1.00E+07 0.00E+00 Strength vs. Aspect Ratio wrt % Cement 28 Day Test NEX • 4% Cement Dry • 6% Cement Wet A 8% Cement Wet A 8% Cement Dry x 10% Cement Wet a 10% Cement Dry 4 3 a „ a «L a . a i _____ • • • " j—*—m—m—| ~: « • 1 1 1 1 1 1 1.00 1.10 1.20 1.30 1.40 1.50 Aspect Ratio (Dim 1/Dim 3) 1.60 1.70 152 800 ,? 600 + " z 400 4-•a = 200 + a. a. < F o r c e at F a i l u r e v s . T i m e wrt S i z e 8% C e m e n t & N E X T a i l i n g s A S3 • — — • D • r r i i _ f 1 1 1 1 10 20 30 Time (days) 40 50 • Ave 1.1 cm • Ave 1.5 cm • Ave 1.9 cm x Ave 2.3 cm ©Ave 1.1 cm B o Ave 1.5 cm B A Ave 1.9 cm B ggAve 2.6 cm B 250 Force at Failure vs. Time wrt Size 6% Cement & NEX Tailings 10 20 30 Time (days) 40 50 • Ave 1.1 cm • Ave 1.5 cm A Ave 1.8 cm x Ave 2.2 cm Strength vs. Time wrt Size 10% Cement & NEX Tailings 10 20 30 Time (days) 40 50 • Ave 1.1 cm • Ave 1.5 cm A Ave 1.9 cm x Ave 2.6 cm o Ave 1.1 cm B • Ave 1.5 cm B A Ave 2.0 cm B S t r e n g t h v s . T i m e wrt S i z e 8% C e m e n t & N E X T a i l i n g s 20 30 Time (days) 40 50 • Ave 1.1 cm • Ave 1.5 cm A Ave 1.9 cm x Ave 2.3 cm ©Ave 1.1 cm B D Ave 1.5 cm B A Ave 1.9 cm B BAve 2.6 cm B 40 S t r e n g t h v s . T i m e wrt S i z e 6% C e m e n t & N E X T a i l i n g s CM E z S 30 -F A j i (MPa o 20 -• Mve i.i cm • Ave 1.5 cm A Ave 1.8 cm -1-—-O) 10 -x Ave 2.2 cm c 0) — CO o I I I 1 ( 1 | 1 1 I 10 20 30 40 Time (days) 50 154 Force at Failure vs. Time wrt Size 4% Cement & NEX Tailings CJ — TO 250 200 150 u Z o L L T3 CJ Q. Q. < 100 • Ave 1.1 cm • Ave 1.5 cm _ Ave 1.9 cm x Ave 2.3 cm 10 20 30 Time (days) 40 50 Strength vs. Time wrt Size 4% Cement & NEX Tailings • Ave 1.1 cm • Ave 1.5 cm - Ave 1.9 cm x Ave 2.3 cm 10 40 50 Note: 20 30 Time (days) All 28 day samples were cured outside of the humidifier after 7 days. NEX Pellet Size Distributions 100.00 90.00 - • — N1 - 10% C, 15.4% W -m~ N2 - 8% C, 14.1 % W N3 - 6% C, 16% W H * — N4 - 10% C, 17.4% W N5- 8% C, 17.1% W - * — N6 - 4% C, 15.5% W 1000 100 10 Size (mm) 0.1 Eskay Creek Agglomeration project First number in each batch # indicates NEX Tailings age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Strength Average Strengths (Pa) 1st 2nd 3rd Ratio lbs N Diam (m] Compressive Tensile Shear 3N1 10% 15.4% 1.54 2.7 2.6 2.5 0.93 92 409.38 0.03 5.03E+07 6.70E+06 1.68E+07 10% 15.4% 1.54 2.7 2.1 2.1 0.78 40 177.99 0.02 4.19E+07 5.59E+06 1.40E+07 10% 15.4% 1.54 1.9 1.9 1.8 0.95 20 88.99 0.02 3.76E+07 5.01 E+06 1.25E+07 10% 15.4% 1.54 1.9 1.8 1.7 0.89 16 71.20 0.02 3.59E+07 4.78E+06 1.20E+07 10% 15.4% 1.54 1.9 1.9 1.8 0.95 45 200.24 0.02 4.93E+07 6.57E+06 1.64E+07 10% 15.4% 1.54 2.1 1.5 1.3 0.62 43 191.34 0.02 5.39E+07 7.19E+06 1.80E+07 10% 15.4% 1.54 1.5 1.5 1.4 0.93 39 173.54 0.01 5.51 E+07 7.35E+06 1.84E+07 10% 15.4% 1.54 1.7 1.5 1.3 0.76 32 142.39 0.02 5.10E+07 6.80E+06 1.70E+07 10% 15.4% 1.54 1.3 1.2 1.2 0.92 18 80.10 0.01 4.82E+07 6.43E+06 1.61 E+07 10% 15.4% 1.54 1.2 1.1 1.1 0.92 30 133.49 0.01 6.05E+07 8.06E+06 2.02E+07 10% 15.4% 1.54 1.1 1 1 0.91 20 88.99 0.01 5.62E+07 7.50E+06 1.87E+07 3N2 8% 14.1% 1.76 2.6 2.4 2.2 0.85 82 364.88 0.02 4.32E+07 5.76E+06 1.44E+07 8% 14.1% 1.76 2.4 2.2 2 0.83 53 235.84 0.02 3.96E+07 5.28E+06 1.32E+07 8% 14.1% 1.76 2 2 1.9 0.95 64 284.78 0.02 4.53E+07 6.04E+06 1.51 E+07 8% 14.1% 1.76 2.2 1.9 1.8 0.82 72 320.38 0.02 4.76E+07 6.34E+06 1.59E+07 8% 14.1% 1.76 1.5 1.5 1.4 0.93 34 151.29 0.01 4.46E+07 5.95E+06 1.49E+07 8% 14.1% 1.76 1.6 1.5 1.4 0.88 26 115.69 0.02 4.03E+07 5.38E+06 1.34E+07 8% 14.1% 1.76 1.7 1.4 1.3 0.76 31 137.94 0.01 4.38E+07 5.83E+06 1.46E+07 8% 14.1% 1.76 1.2 1.1 1 0.83 12 53.40 0.01 3.83E+07 5.11E+06 1.28E+07 8% 14.1% 1.76 1.3 1.1 1.1 0.85 25 111.24 0.01 4.75E+07 6.34E+06 1.58E+07 8% 14.1% 1.76 1.2 1.1 1.1 0.92 21 93.44 0.01 4.55E+07 6.07E+06 1.52E+07 3N3 6% 16.0% 2.67 2.9 2.8 2.3 0.79 40 177.99 0.03 2.25E+07 2.99E+06 7.48E+06 6% 16.0% 2.67 2.8 2.2 2.2 0.79 39 173.54 0.02 2.44E+07 3.26E+06 8.14E+06 6% 16.0% 2.67 2.3 2 1.7 0.74 34 151.29 0.02 2.60E+07 3.46E+06 8.66E+06 6% 16.0% 2.67 2.2 1.9 1.8 0.82 27 120.14 0.02 2.45E+07 3.26E+06 8.16E+06 6% 16.0% 2.67 1.9 1.8 1.7 0.89 25 111.24 0.02 2.52E+07 3.35E+06 8.38E+06 6% 16.0% 2.67 1.7 1.7 1.5 0.88 15 66.75 0.02 2.25E+07 3.00E+06 7.49E+06 6% 16.0% 2.67 1.7 1.6 1.3 0.76 18 80.10 0.02 2.49E+07 3.32E+06 8.30E+06 6% 16.0% 2.67 1.7 1.7 1.4 0.82 26 115.69 0.02 2.73E+07 3.64E+06 9.09E+06 6% 16.0% 2.67 1.3 1.1 1.1 0.85 11 48.95 0.01 2.58E+07 3.44E+06 8.60E+06 6% 16.0% 2.67 1.3 1.1 1.1 0.85 10 44.50 0.01 2.50E+07 3.33E+06 8.33E+06 6% 16.0% 2.67 1.1 1.1 0.9 0.82 11 48.95 0.01 2.74E+07 3.65E+06 9.13E+06 3N4 10% 17.4% 1.74 2.3 2.2 2 0.87 58 258.08 0.02 4.86E+07 6.47E+06 1.62E+07 10% 17.4% 1.74 2.3 2.1 2 0.87 50 222.49 0.02 4.69E+07 6.26E+06 1.56E+07 10% 17.4% 1.74 2.2 1.9 1.5 0.68 43 191.34 0.02 4.85E+07 6.47E+06 1.62E+07 10% 17.4% 1.74 2.3 2 1.9 0.83 42 186.89 0.02 4.54E+07 6.05E+06 1.51 E+07 10% 17.4% 1.74 2.2 1.9 1.7 0.77 41 182.44 0.02 4.69E+07 6.25E+06 1.56E+07 10% 17.4% 1.74 1.8 1.6 1.5 0.83 29 129.04 0.02 4.68E+07 6.24E+06 1.56E+07 10% 17.4% 1.74 1.7 1.5 1.4 0.82 31 137.94 0.02 4.99E+07 6.66E+06 1.66E+07 10% 17.4% 1.74 1.6 1.6 1.3 0.81 23 102.34 0.02 4.52E+07 6.03E+06 1.51 E+07 10% 17.4% 1.74 1 1 1 1.00 18 80.10 0.01 5.52E+07 7.36E+06 1.84E+07 10% 17.4% 1.74 1.2 1 1 0.83 19 84.55 0.01 5.44E+07 7.25E+06 1.81 E+07 10% 17.4% 1.74 1.1 1 0.9 0.82 11 48.95 0.01 4.68E+07 6.24E+06 1.56E+07 3N5 8% 17.1% 2.14 2.4 2.2 2.1 0.88 66 293.68 0.02 4.23E+07 5.64E+06 1.41 E+07 8% 17.1% 2.14 2.4 2 2 0.83 46 204.69 0.02 3.90E+07 5.20E+06 1.30E+07 8% 17.1% 2.14 2 2 2 1.00 40 177.99 0.02 3.84E+07 5.12E+06 1.28E+07 8% 17.1% 2.14 2 1.9 1.9 0.95 38 169.09 0.02 3.88E+07 5.17E+06 1.29E+07 8% 17.1% 2.14 1.5 1.4 1.3 0.87 18 80.10 0.01 3.74E+07 4.98E+06 1.25E+07 8% 17.1% 2.14 1.5 1.5 1.3 0.87 20 88.99 0.01 3.78E+07 5.04E+06 1.26E+07 8% 17.1% 2.14 1.4 1.4 1.3 0.93 18 80.10 0.01 3.78E+07 5.04E+06 1.26E+07 8% 17.1% 2.14 1.1 1 0.9 0.82 8 35.60 0.01 3.57E+07 4.76E+06 1.19E+07 8% 17.1% 2.14 1.1 1 1 0.91 11 48.95 0.01 3.90E+07 5.20E+06 1.30E+07 8% 17.1% 2.14 1.2 1 1 0.83 10 44.50 0.01 3.72E+07 4.96E+06 1.24E+07 3N6 4% 15.5% 3.88 2.5 2.5 2.2 0.88 22 97.89 0.02 1.19E+07 1.59E+06 3.97E+06 4% 15.5% 3.88 2.2 2.2 2.2 1.00 33 146.84 0.02 1.46E+07 1.94E+06 4.85E+06 4% 15.5% 3.88 2.1 2.1 1.8 0.86 27 120.14 0.02 1.44E+07 1.92E+06 4.79E+06 4% 15.5% 3.88 2.3 2 2 0.87 26 115.69 0.02 1.40E+07 1.86E+06 4.66E+06 4% 15.5% 3.88 1.6 1.6 1.5 0.94 17 75.65 0.02 1.46E+07 1.94E+06 4.86E+06 4% 15.5% 3.88 1.5 1.5 1.5 1.00 18 80.10 0.02 1.53E+07 2.O5E+06 5.12E+06 4% 15.5% 3.88 1.6 1.5 1.4 0.88 17 75.65 0.02 1.51 E+07 2.01 E+06 5.02E+06 4% 15.5% 3.88 1.3 1.2 1 0.77 8 35.60 0.01 1.38E+07 1.84E+06 4.59E+06 4% 15.5% 3.88 1.4 1.1 1.1 0.79 5 22.25 0.01 1.18E+07 1.57E+06 3.93E+06 4% 15.5% 3.88 1.3 1.2 1 0.77 10 44.50 0.01 1.48E+07 1.98E+06 4.95E+06 7N1 10% 15.4% 1.54 2 1.9 1.8 0.90 0.00 0.02 0.O0E+O0 O.00E+0O 0.00E+00 10% 15.4% 1.54 2.2 1.9 1.7 0.77 55 244.74 0.02 5.18E+07 6.90E+06 1.73E+07 10% 15.4% 1.54 2.2 2 1.8 0.82 66 293.68 0.02 5.36E+07 7.15E+06 1.79E+07 156 Eskay Creek Agglomeration project First number in each batch # indicates NEX Tailings age of pellet at time of testing. 'B' indicates Brazilian test curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Strength Average Strengths (Pa) 1st 2nd 3rd Ratio lbs N Diam (m) Compressive Tensile Shear 10% 15.4% 1.54 1.5 1.5 1.2 0.80 44 195.79 0.01 5.87E+07 7.83E+06 1.96E+07 10% 15.4% 1.54 1.5 1.5 1.4 0.93 44 195.79 0.01 5.74E+07 7.65E+06 1.91 E+07 10% 15.4% 1.54 1.5 1.4 1.3 0.87 19 84.55 0.01 4.49E+07 5.99E+06 1.50E+07 10% 15.4% 1.54 1.4 1.2 1 0.71 20 88.99 0.01 5.06E+07 6.75E+06 1.69E+07 10% 15.4% 1.54 1.3 1.3 1.2 0.92 38 169.09 0.01 6.02E+07 8.03E+06 2.01 E+07 10% 15.4% 1.54 1.2 1 1 0.83 14 62.30 0.01 4.91 E+07 6.55E+06 1.64E+07 10% 15.4% 1.54 1 1 0.8 0.80 13 57.85 0.01 5.12E+07 6.83E+06 1.71 E+07 7N2 8% 14.1% 1.76 2.6 2.6 2 0.77 75 333.73 0.02 4.14E+07 5.52E+06 1.38E+07 8% 14.1% 1.76 2.2 2 1.8 0.82 71 315.93 0.02 4.65E+07 6.20E+06 1.55E+07 8% 14.1% 1.76 2 1.9 1.8 0.90 48 213.59 0.02 4.23E+07 5.63E+06 1.41 E+07 8% 14.1% 1.76 2 1.9 1.8 0.90 46 204.69 0.02 4.17E+07 5.56E+06 1.39E+07 8% 14.1% 1.76 1.5 1.3 1.3 0.87 23 102.34 0.01 4.15E+07 5.54E+06 1.38E+07 8% 14.1% 1.76 1.5 1.5 1.5 1.00 44 195.79 0.02 4.81 E+07 6.41 E+06 1.60E+07 8% 14.1% 1.76 1.4 1.4 1.2 0.86 29 129.04 0.01 4.49E+07 5.98E+06 1.50E+07 8% 14.1% 1.76 1.2 1.2 1 0.83 13 57.85 0.01 3.82E+07 5.10E+06 1.27E+07 8% 14.1% 1.76 1.3 1.1 1.1 0.85 16 71.20 0.01 4.10E+07 5.46E+06 1.37E+07 8% 14.1% 1.76 1 1 1 1.00 21 93.44 0.01 4.92E+07 6.56E+06 1.64E+07 8% 14.1% 1.76 1 1 0.9 0.90 15 66.75 0.01 4.48E+07 5.97E+06 1.49E+07 7N3 6% 16.0% 2.67 2.7 2.1 2.1 0.78 39 173.54 0.02 2.51 E+07 3.35E+06 8.38E+06 6% 16.0% 2.67 2.9 2.1 2 0.69 40 177.99 0.02 2.52E+07 3.35E+06 8.39E+06 6% 16.0% 2.67 1.9 1.9 1.5 0.79 35 155.74 0.02 2.81 E+07 3.75E+06 9.38E+06 6% 16.0% 2.67 1.8 1.6 1.4 0.78 12 53.40 0.02 2.13E+07 2.84E+06 7.10E+06 6% 16.0% 2.67 1.6 1.5 1.3 0.81 11 48.95 0.01 2.18E+07 2.91 E+06 7.28E+06 6% 16.0% 2.67 1.6 1.5 1.3 0.81 19 84.55 0.01 2.62E+07 3.49E+06 8.74E+06 6% 16.0% 2.67 1.5 1.3 1.3 0.87 15 66.75 0.01 2.57E+07 3.43E+06 8.57E+06 6% 16.0% 2.67 1.3 1.2 1.2 0.92 11 48.95 0.01 2.47E+07 3.30E+06 8.24E+06 6% 16.0% 2.67 1.1 1 0.9 0.82 7 31.15 0.01 2.44E+07 3.25E+06 8.12E+06 6% 16.0% 2.67 1.1 1 1 0.91 5 22.25 0.01 2.14E+07 2.85E+06 7.14E+06 6% 16.0% 2.67 1 1 0.9 0.90 8 35.60 0.01 2.59E+07 3.45E+06 8.63E+06 7N4 10% 17.4% 1.74 3 2.7 2.5 0.83 81 360.43 0.03 4.67E+07 6.23E+06 1.56E+07 10% 17.4% 1.74 2.7 2.4 2.2 0.81 72 320.38 0.02 4.85E+07 6.47E+06 1.62E+07 10% 17.4% 1.74 2.6 2.1 2 0.77 69 307.03 0.02 5.11 E+07 6.81 E+06 1.70E+07 10% 17.4% 1.74 2.3 1.9 1.8 0.78 65 289.23 0.02 5.38E+07 7.17E+06 1.79E+07 10% 17.4% 1.74 1.9 1.8 1.8 0.95 56 249.19 0.02 5.39E+07 7.19E+06 1.80E+07 10% 17.4% 1.74 1.8 1.7 1.6 0.89 42 186.89 0.02 5.14E+07 6.85E+06 1.71 E+07 10% 17.4% 1.74 1.7 1.5 1.3 0.76 39 173.54 0.02 5.45E+07 7.26E+06 1.82E+07 10% 17.4% 1.74 1.6 1.5 1.4 0.88 34 151.29 0.02 5.21 E+07 6.94E+06 1.74E+07 10% 17.4% 1.74 1.7 1.3 1.2 0.71 36 160.19 0.01 5.62E+07 7.50E+06 1.87E+07 10% 17.4% 1.74 1.4 1.3 1.3 0.93 34 151.29 0.01 5.66E+07 7.54E+06 1.89E+07 10% 17.4% 1.74 1 1 0.9 0.90 16 71.20 0.01 5.40E+07 7.19E+06 1.80E+07 7N5 8% 17.1% 2.14 2.9 2.8 2.7 0.93 82 364.88 0.03 3.90E+07 5.20E+06 1.30E+07 8% 17.1% 2.14 2.7 2.2 2.1 0.78 56 249.19 0.02 3.92E+07 5.22E+06 1.31 E+07 8% 17.1% 2.14 2.4 1.9 1.8 0.75 65 289.23 0.02 4.52E+07 6.02E+06 1.51 E+07 8% 17.1% 2.14 1.8 1.7 1.5 0.83 56 249.19 0.02 4.84E+07 6.45E+06 1.61 E+07 8% 17.1% 2.14 1.6 1.5 1.4 0.88 29 129.04 0.02 4.18E+07 5.58E+06 1.39E+07 8% 17.1% 2.14 1.6 1.5 1.3 0.81 29 129.04 0.01 4.23E+07 5.64E+06 1.41 E+07 8% 17.1% 2.14 1.5 1.4 1.3 0.87 28 124.59 0.01 4.33E+07 5.77E+06 1.44E+07 8% 17.1% 2.14 1.4 1.3 1 0.71 22 97.89 0.01 4.31E+07 5.74E+06 1.44E+07 8% 17.1% 2.14 1.3 1.2 1.1 0.85 24 106.79 0.01 4.56E+07 6.08E+06 1.52E+07 8% 17.1% 2.14 1.5 1.1 1 0.67 19 84.55 0.01 4.28E+07 5.70E+06 1.43E+07 8% 17.1% 2.14 1.2 1 0.8 0.67 14 62.30 0.01 4.30E+07 5.73E+06 1.43E+07 7N6 4% 15.5% 3.88 2.7 2.5 2.3 0.85 29 129.04 0.03 1.28E+07 1.71 E+06 4.27E+06 4% 15.5% 3.88 2.3 2.2 2 0.87 33 146.84 0.02 1.47E+07 1.95E+06 4.89E+06 4% 15.5% 3.88 2.5 2.1 1.9 0.76 28 124.59 0.02 1.40E+07 1.86E+06 4.66E+06 4% 15.5% 3.88 2.2 2 1.9 0.86 34 151.29 0.02 1.55E+07 2.07E+06 5.18E+06 4% 15.5% 3.88 2.1 1.9 1.7 0.81 24 106.79 0.02 1.44E+07 1.92E+06 4.81 E+06 4% 15.5% 3.88 1.9 1.8 1.7 0.89 26 115.69 0.02 1.54E+07 2.05E+06 5.12E+06 4% 15.5% 3.88 1.8 1.5 1.3 0.72 17 75.65 0.02 1.49E+07 1.99E+06 4.96E+06 4% 15.5% 3.88 1.7 1.5 1.4 0.82 19 84.55 0.02 1.55E+07 2.06E+06 5.15E+06 4% 15.5% 3.88 1.6 1.4 1.3 0.81 18 80.10 0.01 1.59E+07 2.12E+06 5.29E+06 4% 15.5% 3.88 1.4 1.3 1.1 0.79 14 62.30 0.01 1.57E+07 2.10E+06 5.24E+06 4% 15.5% 3.88 1.2 1.1 0.9 0.75 8 35.60 0.01 1.46E+07 1.95E+06 4.88E+06 28N1 10% 15.4% 1.54 3.1 2.8 2.4 0.77 130 578.47 0.03 5.40E+07 7.20E+06 1.80E+07 10% 15.4% 1.54 2.8 2 1.8 0.64 51 226.94 0.02 4.69E+07 6.25E+06 1.56E+07 10% 15.4% 1.54 2.4 1.8 2 0.83 87 387.13 0.02 5.88E+07 7.84E+06 1.96E+07 10% 15.4% 1.54 2.3 1.8 1.5 0.65 94 418.28 0.02 6.35E+07 8.47E+06 2.12E+07 157 Eskay Creek Agglomeration project NEX Tailings Brazilian test First number in each batch # indicates age of pellet at time of testing. 'B' indicates curing outside of a humidifier after the first 7 days. Batch Cement Water W:C Ratio Dimensions (cm) Aspect Strength Average Strengths (Pa) 1st 2nd 3rd Ratio lbs N Diam (m) Compressive Tensile Shear 10% 15.4% 1.54 2 1.7 1.7 0.85 76 338.18 0.02 6.08E+07 8.11 E+06 2.03E+07 10% 15.4% 1.54 1.9 1.6 1.3 0.68 54 240.29 0.02 5.82E+07 7.76E+06 1.94E+07 10% 15.4% 1.54 1.8 1.5 1.2 0.67 44 195.79 0.02 5.67E+07 7.56E+06 1.89E+07 10% 15.4% 1.54 1.7 1.4 1.1 0.65 36 160.19 0.01 5.56E+07 7.41 E+06 1.85E+07 10% 15.4% 1.54 1.7 1.3 1.2 0.71 45 200.24 0.01 6.06E+07 8.07 E+06 2.02E+07 10% 15.4% 1.54 1.2 1.1 0.9 0.75 17 75.65 0.01 5.16E+07 6.88E+06 1.72E+07 10% 15.4% 1.54 1.1 0.9 0.8 0.73 15 66.75 0.01 5.47E+07 7.29E+06 1.82E+07 28N2 8% 14.1% 1.76 2.8 2.7 2.4 0.86 94 418.28 0.03 4.24E+07 5.65E+06 1.41 E+07 8% 14.1% 1.76 2.6 2.5 2.2 0.85 118 525.07 0.02 4.82E+07 6.42E+06 1.61 E+07 8% 14.1% 1.76 2.5 2 1.9 0.76 52 231.39 0.02 4.06E+07 5.41 E+06 1.35E+07 8% 14.1% 1.76 1.9 1.8 1.5 0.79 63 280.33 0.02 4.89E+07 6.52E+06 1.63E+07 8% 14.1% 1.76 1.9 1.6 1.4 0.74 42 186.89 0.02 4.49E+07 5.98E+06 1.50E+07 8% 14.1% 1.76 1.7 1.5 1.5 0.88 44 195.79 0.02 4.70E+07 6.27E+06 1.57E+07 8% 14.1% 1.76 1.6 1.5 1.5 0.94 48 213.59 0.02 4.89E+07 6.52E+06 1.63E+07 8% 14.1% 1.76 1.5 1.4 1.3 0.87 38 169.09 0.01 4.79E+07 6.39E+06 1.60E+07 8% 14.1% 1.76 1.5 1.4 1.2 0.80 34 > 151.29 0.01 4.67E+07 6.23E+06 1.56E+07 8% 14.1% 1.76 1.5 1.3 1.1 0.73 26 115.69 0.01 4.44E+07 5.92E+06 1.48E+07 8% 14.1% 1.76 1.1 1 0.9 0.82 25 111.24 0.01 5.22E+07 6.95E+06 1.74E+07 28N3 6% 16.0% 2.67 2.6 2.4 2.1 0.81 45 200.24 0.02 2.54E+07 3.39E+06 8.48E+06 6% 16.0% 2.67 2.3 2.2 1.7 0.74 26 115.69 0.02 2.30E+07 3.07E+06 7.67E+06 6% 16.0% 2.67 2.5 1.9 1.8 0.72 38 169.09 0.02 2.67E+07 3.56E+06 8.91 E+06 6% 16.0% 2.67 1.9 1.8 1.6 0.84 38 169.09 0.02 2.92E+07 3.89E+06 9.73E+06 6% 16.0% 2.67 1.9 1.7 1.3 0.68 23 102.34 0.02 2.59E+07 3.46E+06 8.64E+06 6% 16.0% 2.67 2 1.6 1.4 0.70 23 102.34 0.02 2.59E+07 3.46E+06 8.64E+06 6% 16.0% 2.67 •1.8 1.6 1.4 0.78 34 151.29 0.02 3.01 E+07 4.02E+06 1.00E+07 6% 16.0% 2.67 1.8 1.5 1.2 0.67 17 75.65 0.02 2.50E+07 3.33E+06 8.33E+06 6% 16.0% 2.67 1.4 1.4 1.2 0.86 18 80.10 0.01 2.73E+07 3.64E+06 9.10E+06 6% 16.0% 2.67 1.4 1.2 1.1 0.79 13 57.85 0.01 2.61 E+07 3.49E+06 8.71 E+06 6% 16.0% 2.67 1.1 1.1 1 0.91 11 48.95 0.01 2.70E+07 3.60E+06 8.99E+06 28N4B 10% 17.4% 1.74 2.6 2.2 1.6 0.62 146 649.66 0.02 6.66E+07 8.87E+06 2.22E+07 10% 17.4% 1.74 2.5 2 1.7 0.68 117 520.62 0.02 6.38E+07 8.51 E+06 2.13E+07 10% 17.4% 1.74 2 1.9 1.6 0.80 153 680.81 0.02 7.47E+07 9.96E+06 2.49E+07 10% 17.4% 1.74 1.9 1.7 1.7 0.89 95 422.73 0.02 6.62E+07 8.82E+06 2.21 E+07 10% 17.4% 1.74 1.7 1.6 1.4 0.82 68 302.58 0.02 6.35E+07 8.46E+06 2.12E+07 10% 17.4% 1.74 1.9 1.5 1.4 0.74 103 458.32 0.02 7.29E+07 9.72E+06 2.43E+07 10% 17.4% 1.74 1.6 1.5 1.4 0.88 63 280.33 0.02 6.39E+07 8.52E+06 2.13E+07 10% 17.4% 1.74 1.6 1.4 1.4 0.88 84 373.78 0.01 7.20E+07 9.60E+06 2.40E+07 10% 17.4% 1.74 1.4 1.2 1.2 0.86 58 258.08 0.01 7.02E+07 9.36E+06 2.34E+07 10% 17.4% 1.74 1.4 1.1 1 0.71 40 177.99 0.01 6.56E+07 8.75E+06 2.19E+07 10% 17.4% 1.74 1.1 1 0.9 0.82 37 164.64 0.01 7.02E+07 9.35E+06 2.34E+07 28N5B 8% 17.1% 2.14 2.7 2.6 2.4 0.89 133 591.82 0.03 4.85E+07 6.46E+06 1.62E+07 8% 17.1% 2.14 2.7 2 2 0.74 160 711.96 0.02 5.77E+07 7.69E+06 1.92E+07 8% 17.1% 2.14 2.4 1.8 1.6 0.67 114 507.27 0.02 5.64E+07 7.52E+06 1.88E+07 8% 17.1% 2.14 1.8 1.8 1.7 0.94 83 369.33 0.02 5.31E+07 7.08E+06 1.77E+07 8% 17.1% 2.14 1.9 1.7 1.5 0.79 101 449.42 0.02 5.83E+07 7.78E+06 1.94E+07 8% 17.1% 2.14 1.8 1.6 1.5 0.83 83 369.33 0.02 5.63E+07 7.51 E+06 1.88E+07 8% 17.1% 2.14 1.8 1.4 1.4 0.78 56 249.19 0.02 5.21 E+07 6.94E+06 1.74E+07 8% 17.1% 2.14 1.5 1.3 1.2 0.80 39 173.54 0.01 5.02E+07 6.69E+06 1.67E+07 8% 17.1% 2.14 1.2 1.2 1.1 0.92 35 155.74 0.01 5.24E+07 6.99E+06 1.75E+07 8% 17.1% 2.14 1.4 1.1 1 0.71 46 204.69 0.01 5.82E+07 7.76E+06 1.94E+07 8% 17.1% 2.14 1 0.9 0.9 0.90 43 191.34 0.01 6.58E+07 8.78E+06 2.19E+07 28N6B 4% 15.5% 3.88 2.4 2.2 1.9 0.79 34 151.29 0.02 1.48E+07 1.97E+06 4.94E+06 4% 15.5% 3.88 2.3 2.2 2 0.87 36 160.19 0.02 1.51 E+07 2.01 E+06 5.03E+06 4% 15.5% 3.88 2.2 2 1.9 0.86 32 142.39 0.02 1.52E+07 2.03E+06 5.07E+06 4% 15.5% 3.88 2 1.9 1.6 0.80 31 137.94 0.02 1.60E+07 2.13E+06 5.33E+06 4% 15.5% 3.88 1.8 1.7 1.4 0.78 23 102.34 0.02 1.56E+07 2.08E+06 5.21 E+06 4% 15.5% 3.88 1.8 1.8 1.5 0.83 25 111.24 0.02 1.56E+07 2.08E+06 5.20E+06 4% 15.5% 3.88 1.4 1.4 1.2 0.86 24 106.79 0.01 1.81 E+07 2.42E+06 6.04E+06 4% 15.5% 3.88 1.4 1.2 1 0.71 13 57.85 0.01 1.60E+07 2.13E+06 5.33E+06 4% 15.5% 3.88 1.4 1.3 1.2 0.86 9 40.05 0.01 1.34E+07 1.79E+06 4.47E+06 4% 15.5% 3.88 1.1 1.1 0.9 0.82 13 57.85 0.01 1.75E+07 2.33E+06 5.82E+06 4% 15.5% 3.88 1.1 0.9 0.8 0.73 12 53.40 0.01 1.85E+07 2.47E+06 6.17E+06 158 cum % passing 100.00| 99.241 97.25I 91.641 74.261 8.95I o.ool cum % passing 100.001 99.141 96.311 92.151 79.731 18.191 o.ool cum % passing 100.001 96.49I 82.92I 65.46| 43.46I 3.20| o.ool cum % retain 0.00 0.76 2.75 8.36 25.74 91.05 100.00 cum % retain o.ool 0.861 3.691 7.85| 20.271 81.811 100.001 cum % retain o.ool 3.511 17.081 34.54| 56.54I 96.80| 100.00I % retain 0.00 0.76 1.99 5.61 17.38 65.31 8.95 % retain o.ool 0.861 2.83| 4.161 12.42| 61.53| 18.19| % retain o.ool 3.511 13.57| 17.461 22.00| 40.271 3.201 § 0 17.44 45.40 128.17; 396.99| 1492.10| 204.501 2284.60! wtfg) 0 18.23| 59.781 87.901 262.39| 1300.00| 384.40| 2112.70| 3 ? 0 79.321 306.321 394.061 496.60| 909.00| 72.20| 2257.501 cumulative wt(g) 0 17.44 62.84 191.01 GO CO LO 2080.1 2284.6 cumulative wtfg) 0 18.231 78.011 165.91| 428.31 1728.3| 2112.7| cumulative wtfg) 0 79.321 385.641 779.701 1276.30| 2185.30| 2257.50| (mm) 0 LO 26.5 22.4 cn 13.5 4.761 0 8%, 17.1% W (mm) 0 LO 26.5I 22.4| cn 13.5| 4.761 0 4%, 15.5% W (mm) 0 LO 26.5| 22.41 05 13.5| 4.761 0 Screen Size (mesh) 11.06° |7/8° |3/4° |0.530" -f =* |#4 minus |Total | N5 Screen Size (mesh) CM 11.06" | |7/8" —| |3/4" | 10.530" | |#4 minus | |Total \ CO Z Screen Size (mesh) CM |1.06" | |7/8" | |3/4" | |0.530" 1 * |#4 minus | |Total | o o ci 8 E 1 00 CM CO CO CM a E E co c CD CU CO 1^ CO LO CO LO CM o o> LO-CO LO CM CO =• CO, o Q.I 3 c Q) CD k-u CO £-1 a> N CO I f o o 0 o o ci LO CO CM LO co APPENDIX D BACKFILL TEST RESULTS 7 Day Backfill Strength vs. %109 Tailings (of Rock, Pellet + Tailings) 9% Cement in Matrix 10.000 _ _ -—• c — CO 8.000 6.000 4.000 2.000 0.000 30 40 50 60 70 % Unagglomerated Tailings 80 90 -•—100 R:0P -e— 80 R: 20 P - * — 60R:40P -X—40 R: 60 P -6—20 R: 80 P -B—OR: 100 P 100 12.000 10.000 8.000 Q_ £ 6.000 c S m 4.000 2.000 0.000 7 Day Backfill Strength vs. % 109 Tailings (of Rock, Pellet + Tailings) 9% Cement in Matrix % Tail, ..CP • 100 R: 0 P • 80 R: 20 P • 60 R: 40 P • 40 R: 60 P • 20 R: 80 P • 0R100P 1°0R:0p ings 161 28 Day Backfill Strength vs. % 109 Tailings (of Rock, Pellet + Tailings) 9% Cement in Matrix - e — 100 R: 0 P - © — 80 ft 20 P - * — 6 0 ft 40 P - X — 4 0 ft 60 P 20 ft 80 P - B — O R : 100 P 10 20 30 40 50 60 70 % Unagglomerated Tailings 80 90 100 28 Day Backfill Strength vs. %109 Tailings (of Rock, Pellet + Tailings) 9% Cement in Matrix 100 R; o p 60 R; 40 p 20 R; so P fa 100 R: 0 P • 80 ft 20 P • 60 ft 40 P O 40 ft 60 P • 20 ft 80 P • 0 R : 1 0 0 P ' Tailing 162 28 Day Strength vs. Cement Content 100% Aggregate Backfill 16.00 14.00 12.00 CL E £ 8.00 o> c cu 6.00 + 4.00 2.00 + 0.00 • 100% Rock o 60% Rock, 40% Pellet (8%C) 109 + 60% Rock, 40% Pellet (6%C) 109 • 60% Rock, 40% Pellet (6%C) NEX • -+-o + H 1 1— 5 6 7 Cement Content (%) 10 163 Eskay Creek Backfill Strengths 20% Tailings with Pellet:Rock Combinations 7 Day Strengths 7.000 T — 6.000 5.000 •s 4.000 4-J Z c 2 co 3.000 + 2.000 1.000 * 0.000 • 109 Tailings, 9% C • NEX Tailings, 9% C 0 :80 16:64 32 :48 48 :32 Rock : Pellet Percentages 64 : 16 80 : 0 0 0 : 80 Eskay Creek Backfill Strengths 20% Tailings with Pellet:Rock Combinations 28 Day Strengths - • — 109 Tailings, 9% C - * — NEX Tailings, 9% C 16 : 64 32 : 48 48 : 32 Rock : Pellet Percentages 64 : 16 80 : 0 164 f 2400 2300 2200 6 2100 |[ It 2000 m E 1900 X 1800 g c 1700 cu ° * 1600 * 1500 1 1400 1300 1200 28 Day Backfill Density vs. Tailings Content All mixes - including Variable Cement x T 20 40 60 Tailings Content in Total Percent of Mix (%) 80 O100R0P O 80 R 20 P X 60 ft 40 P X 40 ft 60 P A 20 ft 80 P • 0 ft 100 P 100 28 Day Backfill Strength vs. Density All mixes - including Variable Cement 16.000 14.000 12.000 •J- 10.000 a. E £ 8.000 s i c £ 6.000 4.000 2.000 0.000 © o O X © X X * ( o o >° © A • A ** " x Sc K • \y. >p c m m I x « I -i • Variable Cement 1200 1300 1400 1500 1600 1700 1800 1900 2000 2100 2200 2300 2400 2500 Density (kg/m3) Note that for each ratio of rock to pellet (as show n in legend), the density changes are due to variability in tailings content. A lso , all of the samples tested for the variable cement results contain no tailings. 165 1 2 . 0 0 0 1 0 . 0 0 0 ^ 8 . 0 0 0 CD Q. £ 6 . 0 0 0 O) c CO CO 4 . 0 0 0 2 . 0 0 0 0 . 0 0 0 Strength vs. Water-Cement Ratio 7 Days • 9C, 109 • 9C, NEX A V C , 109 • • • A * * — S 2 3 4 Water-Cement Ratio 2 5 . 0 0 0 2 0 . 0 0 0 1 5 . 0 0 0 U) _ 1 0 . 0 0 0 CO 5 . 0 0 0 0 . 0 0 0 Strength vs. Water-Cement Ratio 28 Days 2 3 4 Water-Cement Ratio • 9C, 109 • 9C, NEX A V C , 109 • • • y • \ . — t t 6 16( 25.000 20.000 Strength vs. Water-Cement Ratio 28 Days • Harsh Grading • Gap Grading X Full Grading CO a c o 15.000 2 10.000 co 5.000 0.000 y = 9.995x' " 3 , ° R 2 = 0.8885 y = 3 . 0 5 3 7 x 2 ° , 2 s R 2 = 0.3953 y = 6.0265x 1 0 7 , 8 R 2 = 0.6615 X 0.5 1.5 2 2.5 3 Water-Cement Ratio 3.5 4.5 CO Q. 25.000 20.000 15.000 cn c CD mm CO 10.000 5.000 0.000 -I Strength vs. Density 28 Days I • Harsh Grading • Gap Grading 2 = 0.9458 A r u i uraaing SE * • v y = F 1 E . l g x 5 9MS 2 =0.5689 y = 7E-24x 7 R 2 =0.76f * x^5 942 * 6 x Y ^ . > < " 2400.000 2200.000 2000.000 1800.000 1600.000 1400.000 1200.000 1000.000 Density (kg/m3) 167 - - o o o —- - - — O ) C O L O 0 . 0 . 0 - 1 X 0 . 0 . —» i i i Q . - = 3 - C M Q . 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W M - Wes te rn Mine Engineering H E - Husky Excavation Units Price per unit ** Total $ C A D 2000/06 3 E A $25,500.00 H G $76,500 3 E A $4,300.00 H G $12,900 3 E A $2,715.00 H G $8,145 2 E A $19,290.00 W M $38,580 10000 F T $14.00 M H $140,000 100 E A $23.12 M H $2,312 100 E A $22.25 M H $2,225 $280,662 L S H E $250,000 $530,662 181 Eskay Creek Mine Order of Magnitude Operating Cost Estimate Agglomeration Option Power Hopper Mixer Pelletizer Conveyors Manpower Controls & Monitoring Cement Quantities . Addition to 150 tpd agglomeration operation (+1%) Equipment Cost Savings Truck Hauls - for aggregate (Cost of savings per elimination of tonnes hauled) Loader operator Incremental Operating Cost per Day (24 hr) Incremental Operating Cost per Year * Less hours required in comparison to status quo ** Represents source of information on pricing: EC - Eskay Creek Mine AS - AGRA Simons personnel Units Price per unit ** Total SCAD 2000/06 2400 kWH $0.09 EC $216 24 manhours $28.00 EC $672 1 T $106.42 AS $106 -150 T $5.00 EC -$750 10 hours -$25.00 EC -$250 $244 $89,274 182 Eskay Creek Mine Order of Magnitude Operating Cost Estimate Slurry Pipeline Option Units Price per unit ** Total $CAD 2000/06 Power Pumps 1550 kWH $0.09 EC $140 Larox Filter Press * 150 T -$15.00 EC -$2,250 (quoted cost of savings per tonne of filtered material) Manpower Controls & Monitoring 24 manhours $28.00 EC $672 Pump Regulation 24 manhours $35.00 EC $840 Mechanic 24 manhours $45.00 EC $1,080 Loader Operator * 20 manhours -$25.00 EC -$500 Total Relative Operating Cost per Day (24 hr) -$19 Total Relative Operating Cost per Year -$6,757 * not required here in comparison to other options ** Represents source of information on pricing: EC - Eskay Creek Mine 183 Eskay Creek Mine Order of Magnitude Cost Comparison OPTION Capital Costs Year Annual Savings NPV Status Quo $0.00 $0.00 $0.00 Agglomeration $513,575.00 -$513,575.00 Pelletizing System 1 -$89,274.41 -$84,221.14 Storage Tanks 2 -$89,274.41 -$79,453.90 Transport Conveyors 3 -$89,274.41 -$74,956.51 4 -$89,274.41 -$70,713.69 Discount Fac to r for N P V 6% 5 -$89,274.41 -$66,711.03 6 -$89,274.41 -$62,934.93 7 -$89,274.41 -$59,372.58 8 -$89,274.41 -$56,011.87 9 -$89,274.41 -$52,841.38 10 -$89,274.41 -$49,850.36 11 -$89,274.41 -$47,028.64 12 -$89,274.41 -$44,366.64 13 -$89,274.41 -$41,855.32 14 -$89,274.41 -$39,486.16 15 -$89,274.41 -$37,251.09 Total NPV -$1,380,630.25 Slurry Pipeline Piping Pumps Dismantling of Filter Press Discount Fac to r for N P V $530,662.00 -$530,662.00 1 $6,757.13 $6,374.65 2 $6,757.13 $6,013.82 3 $6,757.13 $5,673.41 4 $6,757.13 $5,352.28 5 $6,757.13 $5,049.32 6 $6,757.13 $4,763.51 7 $6,757.13 $4,493.87 8 $6,757.13 $4,239.50 9 $6,757.13 $3,999.53 10 $6,757.13 $3,773.14 11 $6,757.13 $3,559.57 12 $6,757.13 $3,358.08 13 $6,757.13 $3,168.00 14 $6,757.13 $2,988.68 15 $6,757.13 $2,819.51 Total NPV -$465,035.12 184 PHOTOGRAPHS 100% Tailings (unagglomerated) 28 - Day Test Stage The second sample, shown to the right, failed before the 28-day cure stage was complete,on i-rs The exact length of time that passed prior to its failure is unknown. The failure seems to be extensional or tensile, as apparent by the concoidal shapes of the failure surfaces. 100% Tailings (unagglomerated) 7 - Day Test Stage The sample failed in a shearing fracture, the failure surfaces having a somewhat planar feature due to the homogeneity of the mixture. Note the two failure planes are at similar, although opposite, angles to the sample's vertical axis (~ 45°). Samples made up of 100% tailings. Tailings were extracted from the 109 Zone. 80% tailings (unagglomerated) 20% tailing pellets (agglomerated) 28 - Day Stage The sample failed by buckling at the center, expanding in the middle with the appearance of micro-vertical fractures prior to total failure. The failure surfaces are curvilinear, cutting predominantly through the fine tailing matrix and oriented sub-vertically. 60% tailings (unagglomerated) 40% tailing pellets (agglomerated) 28 - Day Stage The sample failed brittly, buckling in a similar fashion to that described above - cracks propagated through ~ 2/3 of the sample height. Note that while some of the pellets within the matrix split in half, others held their shape and the matrix failed around them. Samples made up of combined unagglomerated tailings and pellets (agglomerated tailings). Tailings were extracted from the 109 Zone. I B ? 40% tailings (unagglomerated) 60% tailing pellets (agglomerated) 28 - Day Stage The sample failed by buckling in a similar fashion to that described above - cracks propagated throughout the sample height. Note the expansion of the column's diameter in the central region and the cracks that run sub-vertically along the outside of the sample. 40% tailings (unagglomerated) 60% tailing pellets (agglomerated) 7 - Day Stage The sample failed by buckling at the top, expanding close to the middle with the appearance of micro-vertical fractures prior to total failure. The sample failed both around and through the pellets, as above. Samples made up of combined unagglomerated tailings and pellets (agglomerated tailings). Tailings were extracted from the 109 Zone. 20% tailings (unagglomerated) 80% tailing pellets (agglomerated) 28 - Day Stage The sample failed by shearing fracture, the failure surface being somewhat planar and oriented at ~ 20° to the vertical axis of the sample. The failure surfaces cut through both the pellets and the matrix of the fill. 20% tailings (unagglomerated) 80% tailing pellets (agglomerated) 28 - Day Stage The sample failed both by buckling and in shear failure, but at a much greater angle to the axis of the sample - the sample was not compacted as well and more porous. Again, the failure surfaces were predominantly planar, cutting through pellets and the tailing matrix alike. Samples made up of combined unagglomerated tailings and pellets (agglomerated tailings). Tailings were extracted from the 109 Zone. 189 100% River Rock 28 - Day Test Stage The sample failed by buckling of the central portion, with the appearance of cracks about the sides of the sample prior to failure. The sample failed around the individual rock particles, through the finer matrix of sand, silt and cement. See failure surface to the left. 80% River Rock 20% Tailings (unagglomerated) 28 - Day Test Stage The sample failed by buckling of the central portion of the sample. Note the vertical micro-fractures visible on the sides of the sample. As in the 100% rock fill, this sample held together quite well after the maximum load was relieved. Samples made up of combinations of river rock and unagglomerated tailings. Tailings were extracted from the 109 Zone. no 60% river rock 40% tailings (unagglomerated) 28 - Day Stage The sample failed by buckling at the center, expanding in the middle with the appearance of sub-vertical fractures prior to total failure. The propagation of cracks through this sample was more extensive than the previously described failed samples. 60% river rock 40% tailings (unagglomerated) 28 - Day Stage This sample also failed due to outward buckling of the central region. The propagation of cracks extended sufficiently for the sample to break apart during the application of force. Note that the sample failed through the matrix, and around the individual particles of rock. Samples made up of combined river rock and unagglomerated tailings. Tailings were extracted from the 109 Zone. 191 60% river rock 40% tailing pellets (agglomerated) 28 - Day Stage The sample failed due to the expansion of the middle with the propagation of micro-vertical fractures. The sample held together on the relief of the maximum load. 40% river rock 60% tailing pellets (agglomerated) 28 - Day Stage The sample failed due to the propagation of micro-vertical fractures. The cracks propagated through and around pellets, and around the individual rock particles. The sample held together on the relief of the maximum load. Samples made up of combined river rock and pellets (agglomerated tailings). Tailings were extracted from the 109 Zone. \1Z 40% river rock 60% tailing pellets (agglomerated) 28 - Day Stage Same sample shown above, close up to better view the micro-fractures. Note that the void content of the fill is relatively high. 20% river rock 80% tailing pellets (agglomerated) 28 - Day Stage Sample failed due to the propagation of cracks through the pellets and cement matrix, similar to description above. Note the void content of the fill - it is very significant. Samples made up of combined river rock and pellets (agglomerated tailings). Tailings were extracted from the 109 Zone. 193 16% River Rock 64% Tailing Pellets & 20% Tailings (unagglomerated) 28 - Day Test Stage The sample failed somewhat brittly, by outward buckling of the central portion causing mass failure in many failure planes oriented in all directions. The sample failed around the individual rock particles, and through the finer matrix of sand, silt and cement, as well as the pellets within the fill. 32% River Rock 48% Tailing Pellets 20% Tailings (unagglomerated) 28 - Day Test Stage The sample failed similarly to that above, although the failure planes seem to be oriented more sub-vertically and less randomly. Again, samples failed through pellets and the fine matrix alike. Samples made up of combined river rock, pellets (agglomerated tailings) and unagglomerated tailings. Tailings were extracted from the 109 Zone. 194 48% River Rock 32% Tailing Pellets 20% Tailings (unagglomerated) 28 - Day Stage The sample failed brittly, as above. The failure surfaces did not propagate throughout the entire sample. From the existing portion of the sample, there seems to be two potentially significant failure plane orientations (~ 45° and ~ 25°). made up of combined river rock, pellets (agglomerated tailings) and unagglomerated tailings. Tailings were extracted from the 109 Zone. 

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