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An assessment of acid rock drainage potential of waste rock and implications for long term weathering… Lister, Diane 1994

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AN ASSESSMENT OF ACID ROCK DRAINAGE POTENTIAL OF WASTE ROCK AND IMPLICATIONS FOR LONG TERM WEATHERING OF THE NORTH DUMP AT ISLAND COPPER MINE, PORT HARDY, B.C.  by  DIANE LISTER B.A.Sc., The University of British Columbia, 1989  A THESIS SUBMITTED IN PARTIAL FULFILLMENT OF THE REQUIREMENTS FOR THE DEGREE OF MASTER OF APPLIED SCIENCE in THE FACULTY OF GRADUATE STUDIES (Department of Mining and Mineral Process Engineering)  We accept this thesis as conforming to the required standard  THE UNIVERSITY OF BRITISH COLUMBIA March 1994 © Diane Lister, 1994  In presenting this thesis  in  partial fulfilment of the requirements for degree at the University of British Columbia, I agree that the Library freely available for reference and study. I further agree that permission copying of this thesis for scholarly purposes may be granted by the department  or  by  his  or  her  representatives.  It  is  understood  that  an advanced shall make it for extensive head of my copying  or  publication of this thesis for financial gain shall not be allowed without my written permission.  (Signature)  Department of Mining & Mineral Process Engineering The University of British Columbia Vancouver, Canada Date  DE-6 (2/88)  Ari1 20. 1994  11  ABSTRACT  Island Copper Mine, owned by BHP Minerals Canada Ltd., is located at the north end of Vancouver Island, British Columbia. The mine, a copper-molybdenum porphyry deposit and one of Canada’s largest open low grade copper mines, commenced operation in 1971 and with current reserves, mining is expected to continue until late 1995. Over the years of operation, approximately 616 million tonnes of waste rock has been placed both on land (76 million tonnes) and into Rupert Inlet (540 million tonnes), an adjacent marine fjord.  Acid rock drainage, the term used for contaminated drainage resulting from oxidation of certain sulfide minerals, was first detected in the open pit area in 1982, and from the on land waste rock dumps in 1985. Currently, all drainage from the mine area is directed to a water management pond for recycling to the concentrator and discharge when all provincial effluent standards are met. In comparison with other Canadian mines with acid rock drainage, effluent from Island Copper’s on land dump has relatively low concentrations of contaminants.  The objective of this study was to assist Island Copper in long term prediction of water quality emanating from the North dump, the largest on land dump at the mine. The study involved characterization of both the waste rock dump as a whole, and of the various waste rock types comprising the dump.  Waste rock dump characterization entailed examination of existing data coupled with field measurements. Results indicate that there is sufficient oxygen in almost all areas of the on land dumps for pyrite oxidation. There is also evidence of convective transport of air through the dump. In addition, elevated temperatures, indicative of increased sulfide oxidation rates, have been documented in several of the drill holes through the waste rock dumps.  The limited  historical monitoring of one dump indicates that temperatures have increased over the last five years, but may be stabilizing in the low 20° C range.  111  Waste rock characterization involved geological, mineralogical, geochemical and physical assessment, as well as kinetic testing of samples from eight selected sites on the North dump. From this work, three categories of waste rock were derived:  i) type I rocks, which are  interpreted to have sufficient excess alkalinity to do some degree of buffering on infiltrating acidic drainage, ii) type II rocks, which although possibly generating alkaline leachate at present, are not regarded to have sufficient excess alkalinity to adequately buffer infiltrating acidic drainage, but are not expected to significantly contribute to net acidity of the waste rock dump; and iii) type III rocks, which are presently producing leachate with high net acidity. Type III rocks are of variable lithology, strongly hydrothermally altered, and contain elevated levels of fine grained disseminated pyrite. The dumps or dump areas documented to be producing acidic effluent contain at least 14 percent of type III and 21 percent of type II material.  Comparison of leachate quality from laboratory kinetic tests and waste rock dump effluent indicates that the kinetic tests significantly overestimate actual contaminant loads, and only net acid load and molar calcium to sulfate can be confidently scaled from laboratory to field conditions.  The model derived for prediction of future effluent quality indicates that when dump temperatures stabilize, effluent quality will also stabilize and contaminant concentrations will gradually decrease over time.  iv TABLE OF CONTENTS Abstract Table of Contents List of Figures List of Tables Acknowlegements  ii iv vii xi xiii  1.0 1.1  Introduction and Research Plan Introduction Background 1.2 1.2.1 Acid Rock Drainage 1.2.2 Relevant History of Island Copper Mine 1.2.3 Other Relevant Research Research Objectives 1.3 Research Plan 1.4  1 3 3 4 7 9 9  2.0  Site Characteristics Setting 2.1 2.1.1 Location and Access 2.1.2 Mine History 2.1.3 Local Geography and Climate Geology 2.2 2.2.1 General 2.2.2 Bonanza Group Volcanic 2.2.3 Rhyodacite Porphyry 2.2.4 Hydrothermal Breccia 2.2.5 Hyrdrothermal Alteration 2.2.6 Sulphide Mineral Occurence 2.2.7 Acid Consuming Mineral Occurrence 2.2.8 Discussion Waste Rock Dump Characterization 2.3 2.3.1 Objectives 2.3.2 Overall Effects of Development 2.3.3 Construction 2.3.3.1 Old North Dump 2.3.3.2 Eastern Most Outsiope 2.3.3.3 Cap 2.3.4 Hydrology 2.3.5 Water Quality 2.3.6 Drilling 2.3.7 Acid Base Accounting 2.3.8 Oxygen and Carbon Dioxide Monitoring 2.3.9 Temperature Monitoring 2.3.10 Summary  11 11 11 12 15 15 15 16 17 17 18 20 22 24 24 24 25 26 28 28 29 30 30 32 33 34 36  V  3.0  4.0  5.0  Waste Rock Characterization Objectives 3.1 Methods and Procedures 3.2 3.2.1 Sampling 3.2.2 Date of Mining 3.2.3 Rock Type and Mineralogy Assessment 3.2.4 Geochemistry 3.2.5 Physical Analysis 3.2.6 Replicate Analyses Results 3.3 3.3.1 Overall Observations 3.3.2 Date of Mining 3.3.3 Rock Type and Mineralogy Assessment 3.3.4 Mineralogy 3.3.5 Elemental Analysis 3.3.6 Acid Base Accounting 3.3.7 Physical Analysis Summary 3.4  40 40 40 44 44 45 46 46 47 47 50 50 53 55 57 60 62  Kinetic Test Work Introduction and Background 4.1 Column Tests 4.2 4.2.1 Objectives 4.2.2 Equipment and Procedure 4.2.3 Results General Observations 4.2.3.1 Weekly Leachate Quality 4.2.3.2 Loads and Cumulative Loads 4.2.3.3 Pore Gas Analysis 4.2.3.4 Pre- and Post- Test Analysis of Column Material 4.2.3.5 Tests Cell Humidity 4.3 4.3.1 Background 4.3.2 Objectives 4.3.3 Equipment and Procedure 4.3.4 Results General Observations 4.3.4.1 Weekly Leachate Quality 4.3.4.2 Loads and Cumulative Loads 4.3.4.3 Discussion of Results 4.4 4.4.1 Summary 4.4.2 Kinetic Test Rate Constants 4.4.3 Kinetic Test Neutralization Potential Depletion 4.4.4 Kinetic Test ARD Potential Classification  64 67 68 68 73 73 74 85 88 90 92 92 93 93 95 95 95 100 100 100 108 109 112  Prediction of Effluent Quality Introduction 5.1 Characterization of ARD Potential Categories 5.2  116 117  vi  5.3 5.4  5.5  6.0  5.2.1 Lithology 5.2.2 Alteration 5.2.3 Pyrite 5.2.4 Carbonate Minerals 5.2.5 Acid Base Accounting 5.2.6 Elemental Analyses 5.2.7 Rock Competence 5.2.8 Discussion Assigning ARD Potential Categories to Waste Rock Dumps Prediction of Present Effluent Quality from Dump Areas 5.4.1 Method 5.4.2 Leachate Quality of ARD Potential Categories 5.4.3 Actual Effluent Quality of EMO and NWD Dumps Calculation of Actual Loads 5.4.3.1 Calculation of Sulfide Oxidation Rate Constants 5.4.3.2 Estimated and Actual Effluent Quality of Comparison 5.4.4 from EMO and NWD 5.4.5 Discussion of Results Temporal Modeling of EMO and NWD Dumps 5.5.1 Introduction 5.5.2 Limitations 5.5.3 Method 5.5.4 Results 5.5.5 Validation and Calibration 5.5.6 Discussion  Conclusions and Recommendations Conclusions 6.1 Recommendations 6.2  Appendix 2 Appendix 3 Appendix 4 Appendix 5 Appendix 6 Appendix 7 Appendix 8 Appendix 9 Appendix 10 Appendix 11  139 139 139 141 141 146 151 152  155 158 160  References Appendix 1  117 117 117 119 120 123 124 124 126 129 129 131 131 133 134 135  Petrographic Report on Ten Polished Thin Sections from the Island Copper Deposit, B.C. for Acid Rock Drainage Studies List of Sample Numbers and Description Elemental Analyses of Waste Rock and Till Samples 2 Calculation Method for APP and APPS Sieve Analysis of Waste Rock Samples Moisture Contents, Sites 1 through 4 Sample Pits Leachate Quality Analytical Techniques Leachate Quality Analytical Replicates Column and Humidity Cell Leachate Quality Column Pore Gas Analyses Beach Dump Characterization  164 176 177 179 180 190 191 193 199 215 216  vii LIST OF FIGURES Figure 1.1  Location of Island Copper Mine  2  Figure 1.2  Plan view of Island Copper Mine and waste rock dumps  5  Figure 1.3  Regions of the North dump  6  Figure 2.1  Normal monthly temperature and total precipitation, Port Hardy  13  Airport Figure 2.2  Location of watersheds in Island Copper Mine area  14  Figure 2.3  Geological cross section, Island Copper pit  19  Figure 2.4  Location of perimeter ditch drainage sample sites and North  31  dump drill holes Figure 2.5  Average dump temperature versus time for North West dump  35  Figure 2.6  Down-hole temperatures for AMD #4 (east end of Old North  37  Dump) and AMD #7 (upper Cap) Figure 2.7  Down-hole temperatures for AMD#1 and #2 (lower EMO), and  37  AMD #3 (upper EMO) Figure 2.8  Down-hole temperatures for AMD#5 and #6 (west OND)  38  Figure 3.1  Location of waste rock sample sites  42  Figure 3.2  Sample pit profile, Site 4, lower EMO  48  Figure 3.3  Indurated horizon below highly oxidized zone on upper Cap  49  Figure 3.4  Sericite-chlorite-clay (SCC) altered volcanic? on Upper Cap  51  Figure 3.5  Weakly altered volcanic on Upper Cap  51  Figure 3.6  Acid consuming potential (ACP) versus acid-volatilized carbon  59  dioxide for the waste rock samples Figure 3.7  Range of particle size for waste rock samples  61  Figure 3.8  Particle size distribution of till samples  61  Figure 4.1  Column test apparatus  69  Figure 4.2  Column test leachate pH  76  Figure 4.3  Column test leachate Eh  76  viii Figure 4.4  Column test leachate conductivity  77  Figure 4.5  Column test leachate sulfate  77  Figure 4.6  Column test leachate alkalinity  78  Figure 4.7  Column test leachate acidity  78  Figure 4.8  Column test leachate dissolved metals: a) aluminum, b) calcium,  80  c) cadmium, and d) copper Figure 4.9  Column test leachate dissolved metals: a) iron,  81  b) magnesium, c) manganese, and d) molybdenum Figure 4.10  Column test leachate dissolved metals: a) sodium,  82  b) nickel, c) phosphorus, and d) strontium Figure 4.11  Column test leachate dissolved metals: a) titanium, and b) zinc  83  Figure 4.12  Column test leachate sulfate loadings: a) column 1, b) column 2,  86  c) column 3, and d) column 4 Figure 4.13  Column test leachate alkalinity and acidity loadings: a) column 1,  87  b) column 2, c) column 3, and d) column 4 Figure 4.14  Oxygen content of column pore gas, week 39  89  Figure 4.15  Carbon dioxide content of column pore gas, week 39  89  Figure 4.16  Carbon dioxide content of column 4 pore gas, weeks 39 to 44, and  90  July 28, 1993 Figure 4.17  Pre- and post- test acid base accounting of column test material  91  Figure 4.18  Humidity cell leachate pH  97  Figure 4.19  Humidity cell leachate Eh  97  Figure 4.20  Humidity cell leachate conductivity  98  Figure 4.21  Humidity cell leachate sulfate  98  Figure 4.22  Humidity cell leachate alkalinity  99  Figure 4.23  Humidity cell leachate acidity  99  ix Figure 4.24  Selected humidity cell loads: a) cell 1 sulfate, b)cell 6 sulfate, c)  101  cell 1 alkalinity, and d) cell 6 acidity Figure 4.25  Leachate sulfate loads of humidity cell replicates:  102  a) cell 2, b) cell 3, c) cell 4, and d) cell 6 Figure 4.26  Leachate acidity loads of humidity cell replicates:  103  a) cell 2, b) cell 3, c) cell 4, and d) cell 5 Figure 4.27  Ranges of kinetic test molar calcium and magnesium, alkalinity,  105  and acidity to sulfate load ratios Figure 4.28  Comparison of replicate column and humidity cell sulfate,  107  alkalinity and acidity loads Figure 5.1  Lithological characteristics of ARD potential categories  118  Figure 5.2  Visual pyrite estimates for each ARD potential category  119  Figure 5.3  Acid base accounting parameter characteristics of ARD potential  121  categories, a) total sulfur, b) sulfate, c) acid consuming potential (ACP), d) acid-volatilized carbon dioxide Figure 5.4  Acid base accounting parameter characteristics of ARD potential  122  categories, a) net neutralizing potential using total sulfur (NNPS), b) net neutralizing potential using sulfide sulfur , c) acid consuming potential to acid production (NNPS j 2 potential (ACP:APP), d) paste pH (ppH) Figure 5.5  Estimated proportions of rock in each ARD potential category in  127  the North dump Figure 5.6  Estimated proportions of rock in each ARD potential category in  128  the North West and Beach dumps Figure 5.7  Flowchart for prediction of present effluent quality from dump  130  areas Figure 5.8  EMO predicted sulfate and acidity concentrations, temperature scenario 1  147  x Figure 5.9  EMO predicted sulfate and acidity concentrations, temperature  147  scenario 2 Figure 5.10  EMO predicted sulfate and acidity concentrations, temperature  148  scenario 3 Figure 5.11  Comparison of EMO average annual sulfate and acidity  148  concentrations with predicted values using temperature scenario 2 Figure 5.12  NWD predicted sulfate and acidity concentrations, temperature  149  scenario 1 Figure 5.13  NWD predicted sulfate and acidity concentrations, temperature  149  scenario 2 Figure 5.14  NWD predicted sulfate and acidity concentrations, temperature  150  scenario 3 Figure 5.15  Comparison of NWD average annual sulfate and acidity concentrations with predicted values using temperature scenario 2  150  xi LIST OF TABLES Table 1.1  Comparison of typical drainage quality between Island Copper and  8  other selected mines in Canada Table 2.1  Characteristics of the North dump and its major regions  27  Table 2.2.  Mean and standard deviations of acid base accounting analyses  33  from the North dump and its various areas Table 3.1  Waste rock sample summary  41  Table 3.2  Calculated precision of selected chemical analyses of waste rock  47  Table 3.3  Estimated date of mining and in-pit source for waste rock sample  48  sites Table 3,4  Rock classification summary of the eight waste rock sample sites  52  Table 3.5  X-Ray diffraction results from rock surface material  55  Table 3.6  Acid-Base accounting and acid-volatiliized carbon dioxide analyses  58  of the eight waste rock samples and two till composites Table 4.1  Distilled water infiltration schedule for colunm tests  71  Table 4.2  Waste rock conditions in column tests at start-up  74  Table 4.3  Waste rock samples tested in humidity cells  94  Table 4.4  Kinetic test periods used for data analysis  104  Table 4.5  Minimum and maximum general leachate quality parameters from  104  kinetic tests Table 4.6  Kinetic test rate constants and half-lives  110  Table 4.7  Kinetic test estimated times to neutralization potential depletion  114  Table 4.8  Delineating parameters for ARD Potential classification of kinetic  114  tests Table 4.9  Classification of kinetic tests by ARD Potential category  115  Table 5.1  Dominant alteration characteristics of ARD potential categories  119  Table 5.2  Rock matrix reactivity to hydrochloric acid of ARD potential  120  categories  xli  Table 5.3  Vein or fracture reactivity to hydrochloric acid of ARD potential  120  categories Table 5.4  Rock competence characteristics of ARD potential categories  124  Table 5.5  Tentative guidelines to field classification of Island Copper waste  125  rock by ARD potential category Table 5.6  Tentative acid base accounting criteria for ARD Potential categories  126  Table 5.7  Characteristics of kinetic test leachate chemistry by ARD Potential  132  category Table 5.8  Characteristics of kinetic test sulfide oxidation rate constants by  133  ARD Potential Category Table 5.9  Comparison of calculated versus actual effluent quality of EMO and  136  NWD drainages Table 5.10  Adopted scaling factors for calculating actual effluent conditions  138  from estimated values Table 5.11  Estimated current effluent conditions in selected dumps and dump  138  areas at Island Copper Mine Table 5.12  Available alkalinity in EMO and NWD dump areas  143  Table 5.13  Input parameters for temporal modeling of EMO and NWD dump  144  effluent Table 5.14  Valid prediction periods for model trials  146  xlii  ACKNOWLEDGEMENTS  I would like most of all to thank my advisor Dr. George Poling for obtaining the necessary cooperation and funding for this project, and for his confidence and never-ending support in my work. My committee members have been very flexible and have given me many helpful suggestions as to the direction of the study, and I am grateful to them as well. BHP Minerals Canada Ltd., Island Copper Mine provided funding and conducted some of the analytical work for the project. In particular, Ian Home has done his best to accomodate the needs of this project and his interest and unbounded enthusiasm has provided me with much motivation. I have also been given tremendous support from the Department of Mining and Mineral Process Engineering staff and fellow students. Sally Finora, Pius Lo, and Frank Schmidiger were always able to help me with my technical problems, and Marina Lee and Gordie Lagore often went out of their way to deal with administrative issues. My fellow students seemed to have a knack of being there to help me just when I really needed it, and I only hope that I can return the numerous favours someday. My family and friends have been enthusiastic supporters of my work, and I am grateful to them as well. Finally, I would like to thank the Science Council of British Columbia for providing additional funding in the form of a G.R.E.A.T. scholarship.  1 1.0  INTRODUCTION AND RESEARCH PLAN  1.1  Introduction  Island Copper Mine, owned by BHP Minerals Canada Ltd., is one of Canad&s largest open pit, low-grade copper mines. The mine is located on the north shore of Rupert Inlet, about sixteen kilometres south of the town of Port Hardy at the north end of Vancouver Island (Fig. 1.1). Mining commenced in 1971, and with current reserves is expected to continue until late 1995.  In 1985, seepages from one of the on-land waste rock dumps were found to contain elevated zinc levels and prompted the establishment of an extended monitoring network and the finding of a few areas of the dump where sulfide minerals in the waste rock were oxidizing and producing acidic, metal-contaminated, drainage (termed “acid rock drainage”, or “ARD”). Currently, all seepage emanating from the on-land waste rock dumps is collected by a system of drainage ditches and culverts which direct the water to a water management pond where it can be recycled through the concentrator as process water. Excess water in this pond can be released into Rupert Inlet when all provincial water effluent standards of the permit are met.  Although acid rock drainage is adequately managed at Island Copper Mine, long term water quality trends must be predicted in order to plan for future mine decommissioning. This thesis details waste rock characterization of primarily the North dump, the largest on-land waste rock dump at Island Copper Mine. The study is one of a several conducted and/or funded by the mine to provide the necessary data needed to plan for an appropriate and cost effective  decommissioning of operations.  Figure 1.1  Location of Island Copper Mine  3 The project was funded jointly by BHP Minerals Canada Ltd. Island Copper Mine and a Science Council of British Columbia G.R.E.A.T. Award, with research conducted by Diane Lister, supervised by Dr. George W. Poling (UBC Department of Mining and Mineral Process Engineering), and in collaboration with Ian Home (co-ordinator, environmental closure plan, Island Copper Mine). Work was conducted from May 1992 to March 1994.  1.2  Background  1.2.1  Acid Rock Drainage  The prevention, control, and treatment of acid rock drainage (ARD) is a significant environmental challenge facing the mining industry today. In a 1987 questionnaire circulated to British Columbia’s operating metal mines, six of the sixteen respondants reported the presence of ARD in varying degrees of severity at their site. In addition, at least five abandoned mines in British Columbia reportedly have ARD (Steffen, Robertson and Kirsten (SRK), 1989).  Acid rock drainage results from the spontaneous weathering of certain sulfide minerals (most commonly pyrite due to its ubiquitous presence in the geological environment).  Sulfide  minerals are formed under reducing conditions and when exposed to both oxygen and water can become chemically unstable. The oxidation of pyrite by oxygen can be shown as: 2 +7/202+1120 Fe5  —*  2 + 25042 Fe  +  2H  (1.1)  The dissolved ferrous, sulfate and hydrogen ions result in an increase in total dissolved solids and, unless the surrounding solution is well-buffered, an increase in acidity and subsequent decrease in pH. The reduction in pH can cause solubilization of heavy metals from either the source rock or any medium along its seepage path. If sufficient acidity and heavy metals are introduced into the effluent, the result is a low pH, metalliferous water that has potential to adversely impact aquatic life (SRK, 1989).  4 Mine waste rock piles and tailing impoundments may contain material with sufficient reactive sulfide and insufficient buffering capacity to produce net acidic and metal contaminated effluent. The impacts of ARD left unchecked in Scandanavia and eastern United States have been welldocumented (SRK, 1989).  Equation 1.1 represents just one of many possible reactions. A more rigourous discussion of the chemistry of ARD is presented in SRK (1989), Morin et. al (1991), and Li (1991).  1.2.2  Relevant History of Island Copper Mine  At Island Copper, waste rock has been placed in rock piles at various locations near the pit, both on land and into the sea (Li, 1991). In the last decade, several small rock dumps have been developed in the bottom pit area as well. Construction of the on land dumps started at the commencement of mine operations and continued until 1987. The four on land dumps are: the North dump (76.4 x 106 tonnes), the North West dump (0.95 x 106 tonnes), the West dump (3.0 x 106 tonnes), and the South dump (6.0 x 106 tonnes) (Figure 1.2). The Beach dump, located along the north shoreline of, and extending into Rupert Inlet was also started in the early 70’s. Since 1987, all waste rock produced has been placed on this dump, making it by far the largest at the mine (540 x 106 tonnes),  Despite pre-operational testing in 1968 for heap leach potential that indicated amenability of  both ore and waste rock to bacterial leaching (BHP and Rescan, 1988), potential for ARD at Island Copper was not fully realized until 1982. At this time it was observed that pyrite-rich rocks on the west wall of the pit were iron-stained. Seeps from this area had a pH of 3.5.  In 1985, seeps from the Old Marginal dump, forming the northwestern part of the North dump, began to show elevated zinc levels. This prompted the expansion of the on land dump drainage  •t\)  I C)  I  S  —.  0 Cl)  0  0  9  7 monitoring program and resulted in the delineation of two additional areas of the North dump showing evidence of acid production: the Cap and Eastern Most Outsiopes (EMO) (Figure 1.3). In addition, seepage from the North West dump (NWD) was also found to be net acidic. In 1986, Island Copper commenced construction of perimeter ditches to collect seepage from the North and North West dumps.  Currently, all drainage collected is directed to the water  management pond where water is contained, and either recycled to the mill, or released into an adjacent exfiltration pond when all effluent quality standards of the operating permit are met.  In recognition of the potential for ARD from waste rock, two on land waste rock dumps (the South and West dumps) were constructed from 1986 to 1987 from waste rock designated from production hole acid base accounting analyses to be potentially non-acid generating (BHP and Rescan, 1988). Rock deemed net acid producing was placed in the Beach dump.  Table 1.1 shows typical drainage quality from EMO station, considered to be the most contaminated effluent emanating from the waste rock dumps, and from station WME, which is the cumulative drainage from North and North West dumps. Also shown are typical ARD drainage quality values from selected Canadian mines.  1.2.3  Other Relevant Research  Since the onset of acid rock drainage at Island Copper, the mine has conducted and/or funded considerable related studies including: i) over 300 static tests to determine acid generation potential of waste rock produced by the pit expansion into the south wall (BHP and Rescan, 1988), ii) a drill program on the North and North West dumps to gather samples for static testing, install instrumentation gas monitoring, and allow ongoing temperature and water level monitoring of the core of the dumps (UBC MMPE 1990a,b,c, and Li (1991)), iii) ongoing assessment of leaching from pit walls (Morin, 1992),  8 Table 1.1  Comparision of  typical drainage quality between Island Copper and other  selected mines in Canada (adopted from Rescan, 1992 and Steffen Robertson and Kirsten (1989))  Island Copper  Waste Rock  Underground  Island Copper  Overall Dump  Tailings Pond  Dump Seepage,  Minewater,  ARD Dump  Drainage  Seepage,  British  British  (EMO)  (WME)  Ontario  Columbia  Columbia  pH  4.0  6.5  2.0  2.8  3.5  Sulfate  1800  650  7440  4650  1500  600  40  14600  43000  Not Available  100000  200  588  359  Not Available  2000  60  3600  89800  16500  90  50  50  500  143  1000  Not Available  3200  1190  10.6  18000  2500  11400  53200  28500  Parameter  (mg/I) Acidity l) CaCO / (mg 3 Al (jig/I) Cu (jig/i) Cd (jig/i) Fe (mg/I) Zn (jig/i)  9 iv) hydrological and metal loadings study of North and North West dump drainages (Rescan, 1992), and v) prediction of minewater chemistry from the on-land waste rock dumps (Morin, 1994)  1.3  Research Objectives  The primary objective of this work is to assist Island Copper Mine in long term prediction of water quality emanating from the North Dump by: i)  determining weathering characteristics of the various dump rock types using mineralogy, geochemistry, physical characterization, and controlled laboratory weathering (kinetic testing),  ii)  correlating waste rock characterization fmdings with current and historic waste rock dump conditions, and  iii)  assisting in evaluating alternative control and treatment strategies for decommissioning.  Data derived from this study will be used by Morin (1994), in prediction of metal concentrations from the on-land waste rock dumps.  1.4  Research Plan  In order to achieve the research objectives, work was divided into the following two components: i)  waste rock characterization, and  ii)  dump characterization.  Waste rock characterization comprised first obtaining a suite of samples from eight sites in the North dump deemed representative of the various waste rock units, or mixtures of rock units.  10 Samples were first classified geologically, mineralogically, geochemically (including multielement and acid base accounting analyses) and physically (Chapter 3.0). Laboratory controlled weathering experiments in the form of column leach and humidity cell tests were conducted to determine resultant effluent quality from each sample (Chapter 4.0).  Much of the dump characterization was based on measurements and work conducted by previous graduate student M.G. Li (UBC MMPE 1990a, 1990c, and Li, 1991), but also included recent measurements and observations of the North and other dumps at Island Copper Mine (Chapter 2.0).  Findings from the waste rock characterization study were then related to present and historical conditions mainly in the North dump, but with expansions to other dumps as data and relevance allowed (Chapter 5.0). Implications of the study’s findings for future water quality trends are also presented.  Finally, conclusions of this study, and recommended additional work are discussed in Chapter 6.0.  11 2.0  SITE CHARACTERISTICS  2.1  Setting  2.1.1  Location and Access  Island Copper Mine is located at latitude 50°32’N and longitude 127°37’W on the north shore of Rupert Inlet, a marine fjord. Access to the mine is by paved road from Port Hardy. Separate barge and ship docks at the mine provide moorage for incoming freight and outgoing concentrate (Perelló et al., 1994).  2.1.2  Mine History  Island Copper is a calc-alkaline copper-molybdenum-gold porphyry deposit (Cargill et al., 1976). The initial discovery in the area was made in the mid- 1960’s when prospector Gordon Milboume located minor amounts of native copper and chalcopyrite on roads, outcrops and test pits north of what is now Island Copper. In 1966 Utah Construction and Mining Co. entered into an option agreement with Mr. Milboume and commenced an aggressive exploration program consisting of drilling, geological mapping, soil sampling and geophysics, Drilling in February 1967 testing a large soil copper anomaly several kilometres southeast of the original showings intersected 88 metres of 0.88 percent copper  -  the Island Copper deposit had been discovered.  Drilling  continued and by 1969 reserves of 257 million tons (233 million tonnes) of 0.52 percent copper and 0.0 17 percent molybdenum had been delineated (Perelló et al., 1994).  Project approval was given later in 1969, and construction completed in 1971. Major facilities consisted of:  a 30,000 tonne per day concentrator including tailings thickeners and outfall  pipeline to Rupert Inlet, docking facilities for concentrate shipment, a 19 kilometre water supply pipeline and pump station from Alice Lake (east of the mine), a 194 kilometre 138 kV transmission line from Strathcona generating station, assay and environmental lab facilities, and administration, warehouse and shop buildings.  12  Contingent upon obtaining a permit to discharge tailing effluent into Rupert Inlet, Island Copper committed to conducting an extensive, continuous marine monitoring program covering Rupert Inlet and adjoining fjords. The various surveys, performed on seawater, bottom sediments and marine organisms are conducted on a monthly, quarterly or yearly basis to determine tailings deposit depths and migration, seawater turbidity and metal content, and heavy metal uptakes in marine organisms. In addition, monitoring of freshwater quality of drainages in and adjacent to the mine area is regularly conducted. As a result of this intensive program, the mine operated with five full-time environmental employees and full laboratory facilities. In 1971, this was very much a precedent, and even today the number of staff and facilities at Island Copper’s environmental department are unmatched by any British Columbia mine.  2.1.3  Local Geography and Climate  The area is characterized by low hills up to 150 metres elevation which are overlain by 1 to 20 metres (locally up to 75 metres) of overburden composed of glacial till, colluvium, peat and moss. Outcrops are sparse with exposures limited to road cuts, streams, shorelines, and rare cliffs (Cargill, 1975). A dense growth of timber and undergrowth covers most areas.  Although climate data at the mine site have been collected since commencement of operations, the data, unlike Port Hardy airport’s (17 km northeast of the mine), have not been statistically compiled. Comparisons between Island Copper’s and Port Hardy airport’s precipitation records indicate that the two are sufficiently correlated to be used interchangeably.  Island Copper and the Port Hardy area have a normal annual precipitation of about 1780  millimetres with over 50 percent of this precipitation occurring from October through January, where on average 244 millimetres precipitation occurs per month. Snowfall makes up only about 76 millimetres (4.3 percent) of the annual precipitation, and typically does not accumulate for more than a few weeks during the year. Climatic records for Port Hardy airport indicate that the  13 maximum precipitation recorded in a 24 hour period is 153.8 millimetres which occurred on December 10, 1980. The driest months in the area are May through August with about 65 millimetres monthly precipitation.  Temperatures at the site range from -7°C to 27°C. The mean normal annual temperature, taken from Port Hardy airport meteorological records, is 8°C.  Figure 2.1 shows monthly normal  temperature and precipitation values for Port Hardy airport.  20 C) 0  a?  15 E S  10  C  I-  0)  o  5  CU .1—i  S I—  0  C,  a?  CU  0  o z  -  I— CU  E  I  0  z ._  Figure 2.1  temperature  • precipitation  Normal monthly temperature and total precipitation, Port Hardy Airport (adapted  from BHP, 1988)  Most of the Island Copper Mine and its associated workings are located within two small hydrological basins:  , and Trey Creek km ) the historic End Creek (approximately 3 2  (approximately 1.5 km ) watersheds (Figure 2.2). A portion of the on-land waste rock dumps 2 northwest of the current open pit are located within the much larger Stephens Creek watershed which flows westward into Francis Bay and contains productive fish habitat and a salmon hatchery (Island Copper Mine, 1 988a).  14  C  U  I I ..  () 0  r4  •—  15 Monitoring of freshwater quality began in 1970, one year prior to mine production, and primarily focussed on End Creek, Trey Creek and Bay Lake (located at the headwaters of Stephens Creek). No natural acid rock drainage prior to operations is documented.  2.2  Geology  2.2.1  General  The Island Copper deposit is part of the Island Copper Cluster, a series of five calc-alkaline porphyry copper-molybdenum-gold systems genetically associated with Jurassic rhyodacitic porphyry stocks (approximately 180 Ma) that intruded calc-alkaline basalts, andesites, and pyroclastic rocks of the comagmatic Bonanza Group (Perelló et a!., 1994).  Ore minerals  (chalcopyrite and molybdenite) occur with pyrite in breccia and fracture stockworks immediately adjacent to an approximately 200 metre wide steeply dipping rhyodacitic dyke intruding the volcanics. As is typical of porphyry deposits, rock units have been altered by several stages of hydrothennal fluid intrusions both during and after emplacement of economic minerals (Lister et al., 1993, Perelló et al., 1994).  The three main lithological units recognized in the Island Copper pit area are: i) upper member Bonanza Group Volcanic, ii) Rhyodacite Porphyry, and iii) Hydrothermal Breccia.  2.2.2  Bonanza Group Volcanic  Bonanza Group Volcanics are the most common lithology in the pit area and are comprised of lithic tuffs, breccias and interbedded andesitic and basaltic flows, and form a belt striking approximately N70°W and dipping 25° to 30° to the southwest (Cargill, 1975).  16 Tuffs are massive to finely bedded, are of variable grain size, and contain lithic and volcanic fragments as well as plagioclase and quartz crystals. Other features of the tuffs include graded bedding, fine layering and the occasional presence of bivalve fossils (Perelló et al., 1994).  Flows tend to be massive and have aphanitic to medium-grained porphyritic and brecciated textures.  Primary mineralogy includes plagioclase (labradorite to bytownite composition),  augite, hypersthene and amphibole (Perelló et al., 1994).  2.2.3  Rhyodacite Porphyry  The Rhyodacite Porphyry is subdivided into: i) the Main Porphyry which is associated with economic mineralization, and ii) Intermineral and Late Mineral phases which crosscut the main porphyry and are not associated with economic mineralization.  The Main porphyry forms an elongate dyke intruding the Bonanza Volcanics along a strike of N70°W and a dip of 60° to the northeast. It is over 1200 metres in length, has an average width of 200 metres and extends at least 350 metres below sea level. Much of the dyke has been intensely altered and thus its original composition is difficult to interpret (Cargill, 1975). The minimally altered dyke core consists of a fine-grained predominantly quartz and feldspar groundmass with 15 to 30 percent of 0.4 to 1.0 centimetre subrounded quartz phenociysts, 15 to 30 percent 0.2 to 0.5 centimetre plagioclase (oligoclase to andesine composition) phenocrysts, and 5 percent or less chloritized biotite phenocrysts up to 0.5 centimetres across (Cargill, 1975, Perelló et al., 1994).  Intermineral and Late Mineral phase porphyry units tend to be less hydrothermally altered, are often coarser grained, and tend to have more quartz phenocrysts than the Main porphyry (Perelló Ct  al., 1994).  17 2.2.4  Hydrothermal Breccia  A substantial volume of Hydrothermal Breccias were formed along the Main porphyry and Bonanza Volcanic contact.  Two main types of breccias are recognized in the pit area:  i)  Marginal Breccia, and ii) Pyrophyllite Breccia.  Most of the Marginal Breccias are classified as either i) crackle breccias usually composed of either porphyry or volcanic unrotated fragments in a stockwork of quartz-amphibole-magnetite veinlets, and ii) rotational breccias composed of mineralized rotated fragments of volcanic and/or porphyry in a rhyodacitic, locally quartz-flooded matrix (Perelló et al., 1994).  In the pit, Pyrophyllite Breccia occurs as a tabular body capping the northwest end of the porphyry dyke where it is more than 100 metres wide and gradually tapers off, wedge-like, to the northwest (Cargill, 1975). Poorly sorted, angular to subrounded fragments of volcanic, porphyry and vein quartz are supported by a matrix thought to have been derived from intensely brecciated “rock flour” (Perelló et al., 1994). Subsequent hydrothermal alteration then later transformed the flour to an assemblage of pyrophyllite, kaolinite, sericite, and dumortierite, with variable amounts of pyrite.  2.2.5  Hydrothermal Alteration  The three lithological rock units at Island Copper have been subsequently subjected to hydrothermal alteration, causing modification of original, or primary mineralogy, and in extreme cases, destruction of original textures to the point that the primary rock type is indistinguishable.  The current conception of Island Copper geology identifies three distinct phases of alteration: i) early stage, ii) intermediate stage, and iii) late stage (Perelló et al., 1994).  18 Early stage alteration occurred upon and just after intrusion of the porphyry dyke is and considered by Cargill (1975) to be mainly a contact, or thermal, metamorphism phenomena. Perelló et al. (1994) however believe it to be hydrothermal, and divide the effects of early stage alteration into four outwardly-progressing zones: i) a stockwork core of quartz-amphibole magnetite, ii) a biotite-magnetite zone, and iv) an epidote zone. The economic mineralization, thought to be emplaced mainly during this phase, tends to be contained within the biotite magnetite zone.  Additional economic mineralization is thought to have been emplaced during the intermediate stage of alteration. This stage is mineralogically characterized by secondary quartz, sericite, kaolinite, illitic clays, and chlorite, accompanied by pyrite, molybdenite and minor chalcopyrite. The two types of intermediate alteration distinguished are: i) quartz-sericite-pyrite, affecting a relatively small volume of rock, and ii) pervasive sericite-clay-chiorite-pyrite (SCC) which has affected large volumes of both ore and waste rock. The SCC alteration caused total or partial destruction of feldspars, biotite and amphiboles, although original rock texture was retained. The introduction of fine-grained, disseminated pyrite is also a significant effect of the SCC alteration.  Emplacement of the Pyrophyllite Breccia and later episodes of calcite, ankerite and zeolite veining mark the late stage alteration phase at Island Copper.  Figure 2.3, a typical geological cross section through the deposit, illustrates the complexity of Island Copper geology.  2.2.6  Sulfide Mineral Occurrence  The major sulfide minerals at Island Copper are chalcopyrite, molybdenite, sphalerite and pyrite. Only chalcopyrite and molybdenite are recovered for economic purposes.  Figure 2.3  U.’  v  7  Geological cross section, Island Copper pit (PerellO et al., 1992)  ---120 M)  400 Feet  -240 Bench (1240’below sea level)  0 Bench  560 Bench 480 Bench  F’  level)  [A  L-I  Bonanza Volcanics  // auartzMagnetite Stockwork I I Quartz-’Serlcite Stockwork 1* ‘1 Marginal Breccias 1+ +1 flhyodacite Porphyrles  PROPOSED ULTIMATE PIT  20 Chalcopyrite occurs as veinlets (0.1 mm thick) and disseminations on fractures and slip surfaces. Minor amounts occur with molybdenite on slip surfaces and with sphalerite in late stage carbonate-zeolite veins (Cargill, 1975).  Molybdenite occurs in quartz veins and on fractures and slip surfaces.  The molybdenite  recovered at Island Copper has economically significant rhenium content.  Iron-rich dark brown to black sphalerite occurs as one millimetre crystals with pyrite in carbonate-zeolite veins, both within and adjacent to the ore zone. Trace amounts of galena have also been reported with the sphalerite (Cargill, 1975).  Pyrite is the most common sulfide mineral at Island Copper and occurs both in and adjacent to the ore zone, with modal estimates generally ranging from 2 to 5 and locally up to 15 percent (Cargill, 1975). Within the ore zone, pyrite is associated with chalcopyrite and molybdenite in veinlets, and within chioritized mafic minerals (SCC alteration). In waste rock, fine-grained (1 mm or less) pyrite is disseminated in the porphyry dyke, and also occurs as a disseminated secondary alteration mineral with chlorite in both the early stage chlorite-magnetite and intermediate stage sericite-chlorite-clay (SCC) alteration in all lithological units (Cargill, 1975). Relatively coarse-grained (2 mm) pyrite also occurs with sphalerite in late stage carbonate-zeolite veins.  2.2.7  Acid Consuming Mineral Occurrence  Although calcite, aragonite and dolomite are the most well known acid consuming minerals, Kwong (1994), quoting a study of acid neutralizing capacity of various silicates, proposes that fast and intermediate weathering silicate minerals, if occurring in excess of approximately 10 and  15 percent respectively, may also contribute to acid neutralization. Minerals on Kwong’s list occurring in significant amounts at Island Copper are anorthite (fast weathering), and epidote, pyroxene group minerals, chlorite, and biotite (intermediate weathering). Not mentioned on the  21 list but relatively common at Island Copper are zeolite minerals, which although not acid consuming, can be effective in heavy metal removal (Vos and O’Hearn, 1993).  Calcite is the most common carbonate mineral at Island Copper, but brown-weathering dolomite or ankerite are also observed (Leitch, 1988). Calcite most commonly occurs in veins up to one or two centimetres across, down to tiny veinlets less than 0.5 millimetres wide, and occasionally as irregular alteration patches with sericite. Modal composition is highly variable from virtually none to 5 percent (BHP-Utah and Rescan, 1988). Although not a rule, calcite tends to be more prevalent in the less altered rocks.  Anorthite, a calcium-rich feldspar, has not been identified at Island Copper, but slightly more sodic feldspars, labradorite and bytownite, are present in weakly or unaltered Bonanza Volcanics. Petrographic reports from BHP-Utah and Rescan (1988), however report that although original feldspar contents in the rock are estimated to be 45 to 55 percent in Bonanza Volcanic units and 25 to 45 percent in Rhyodacite Porphyry units, in many samples examined, primary plagioclase has been completely replaced by sericite or sericite-clay (Leitch, 1988).  Overall, feldspar  minerals, because of their replacement by secondary minerals, are not considered significant in acid neutralization.  Epidote and pyroxene minerals, though fairly common in weakly altered Bonanza Volcanic, generally account for less than 10 modal percent of the rock, and are thus not considered to be significant contributors to acid neutralization. Similarly, biotite, though readily recognized by its distinctive purple-brown colour in hand specimens of volcanic, rarely accounts for more than 10 percent of the rock, and thus does not likely play a significant buffering role.  Chlorite minerals are ubiquitous at Island Copper as a replacement of most primary mafic minerals, and are present in most alteration zones (Cargill, 1975).  Leitch (1988) identified two  varieties of chlorite present at Island Copper: i) a possibly higher magnesium variety found only  22 in Bonanza Volcanics, and ii) a possibly lower magnesium variety found only in the rhyodacite porphyry.  Most petrographic reports reviewed indicate that significant amounts of chlorite  (greater than 15 percent) occur only in Bonanza Volcanic units. Estimates of chlorite content in volcanics ranges from 10 to 20 percent in highly altered sericite-chiorite-clay zones, to between 35 and 50 percent in weakly altered chiorite-magnetite and epidote zones. Based on Kwong’s (1994) criteria, chlorite may be a significant buffering mineral at Island Copper.  Although zeolite minerals, occurring as late stage veins with calcite may be of local significance in heavy metal attenuation, they are not considered to be sufficiently abundant to effect overall rock dump water quality.  In summary, calcite and possibly chlorite are the two minerals considered to be present in sufficient quantity and occurring in a variety of rock types to be significant acid neutralizers at Island Copper.  However, because of the lack of substantiated data on the role of chlorite  minerals as ARD neutralizers, for this study only calcite will be considered as the active neutralizing mineral at Island Copper. Dolomite or ankerite, calcium-sodium feldspars, and zeolite minerals though possibly of sufficient quantity in localized areas, likely play an insignificant buffering role.  2.2.8  Discussion  In assessing acid rock drainage potential of Island Copper waste rock, it is of relevance to consider: i) the nature of the circulating hydrothermal fluids during the various alteration phases, ii) the type, habit and reactivity of sulfide minerals, and iii) the type, habit and reactivity of acid consuming minerals.  Early stage alteration is thought to be in large part a “thermal’ phenomena with limited circulation of fluids only where stockwork can be observed in the rock.  23 The intermediate stage pervasive alteration of primary biotite, feldspars and amphiboles to sericite, clay, and chlorite with pyrite is considered to be due to acidic, reducing, sulfur-rich water (Cargill, 1975). Ostatenko and Jones (1976) postulate that the presence of the sericite pyrite assemblage implies a sulfur rich, reducing solution with an estimated pH and temperature of 1.7 and 260°C, respectively.  These fluids, in addition to depositing pyrite, also were  sufficiently acidic to leach strongly the original rock, leaving it with virtually no neutralizing capacity. This concurs with Kwong (1994), which ranks caic-alkaline suite porphyry deposits as more susceptible to ARD than the alkaline suite due to the much more intensive hydrothermal alteration.  Fluids causing late stage pyrophyllite alteration were likely similarly acidic, but with less sulfur, since Pyrophyllite Breccias tend to be low in pyrite. Kwong (1994) notes that pyrophyllite is a relatively unstable mineral under atmospheric conditions and rapidly weathers to clay minerals, releasing aluminum and exposing fresh sulfides.  A substantial change in thermal gradients occurred before fmal stage alteration (Cargill, 1975) since the zeolite-carbonate mineral assemblage is indicative of low temperature, alkaline conditions.  Sulfides present in waste rock include 2 to 5 percent pyrite with rare sphalerite, chalcopyrite, and molybdenite. Pyrite can occur both finely disseminated throughout the three lithological units, and as fracture fillings with chalcopyrite and molybdenite, or with quartz, calcite  ±  sphalerite.  Pyrite contents of up to 15 percent have been observed in occasional wide veins and stockwork.  For this study, calcite is the only mineral considered to have a potential significant contribution to acid neutralization.  24 In summary, the original textures and mineralogy of the three main lithological units at Island Copper proximal to the ore zone have in many cases been totally or partially destroyed by later hydrothermal alteration. In terms of acid rock drainage, the most significant alteration occurred during the intermediate stage when high temperature, reducing, sulfur rich, and acidic solutions infiltrated through the rock mass causing both: i) pervasive leaching of primary feldspar and mafic minerals and replacement with a sericite-chlorite-clay (SCC) assemblage, and ii) emplacement of fine-grained disseminated pyrite. Final phases of late stage alteration emplaced potentially acid-buffering carbonate with zeolite minerals along fractures.  2.3  Waste Rock Dump Characterization  2.3.1  Objectives  The primary objective of the waste rock dump characterization was to assess the macroscopic ARD-controlling features, including: i) flow of water through and out of the waste rock dump, ii) amount of sulfide and acid consuming potential in the waste rock dump, and iii) supply of oxygen to the interior of the waste rock dump.  In addition, current and historic conditions of water quality and waste rock dump temperature are discussed as they apply to prediction of future conditions. 2.3.2  Overall Effects of Development  Local on-land conditions at Island Copper have been affected both by excavation of the open pit, and construction of the waste rock dumps.  Although there is very limited knowledge of pre-mining groundwater conditions, excavation of the open pit created a discharge area, and likely resulted in a lowering of the groundwater table in areas upgradient, or north of the pit. Because the pit and most of the North Dump are within the  25 End and Trey Creek wastersheds, and also that the underlying Bonanza Volcanics dip south, it logically follows that if drainage from the North dump enters the groundwater system, the majority will ultimately seep into the pit.  2.3.3  Construction  Based on observations of current dumping practice and analysis of annual aerial photography, waste rock dumps at Island Copper were constructed using both the push dumping and free dumping methods described in Morin et al. (1991). In push dumping, waste rock is dumped near the dump crest, and a bulldozer is used to push material over the crest. In free dumping, waste rock is deposited in closely spaced truckload-sized piles approximately three metres in height across the level surface of the dump. The material is then leveled using a bulldozer before the next level of dumping proceeds.  The Beach dump is currently the only active waste rock dumping area at the mine and is mainly constructed using the push dumping method. Marginal ore was also deposited on the top of the Beach dump prior to 1985 using both push and free dumping, and is currently being re-handled and processed with freshly mined ore (Perelló et al., 1994).  Similar construction techniques were used in the North, North West, South and West dumps. However, it appears from airphotos that the accumulation of too many truckload piles adjacent to the dump crest resulted in most of the waste rock being pushed over the crest into lifts up to 20 metres high. The remaining waste rock, deposited some distance from the crest was smoothed over the top of the dump, free-dumping style, creating a very short lift.  The overall technique of dumping has had a profound effect on the mixing of the pile material and the distance between subsequent lifts.  Dump scarps on the marginal ore beach dump  illustrate the various patterns that have arisen. Push dumping has resulted in an approximately 8 to 15 metre high lift with a cross section showing the individual truck load piles distributed in a  26 series of moderately steeply dipping bands up to two metres wide. Free dumping has resulted in a much shorter lift height of about two and a half metres, and the waste rock relatively well mixed.  Although push dumping appears to have been most frequently used in dump  construction, the intermittent use of free dumping makes it difficult to predict lift height, and hence water flow paths, in the dump.  Table 2.1 summarizes some important characteristics of the North dump. The table, and much of the following information on the North dump construction was taken from UBC MMPE (1990c). For this study, the North dump consists of three distinct regions: the Old North dump (OND), the Eastern Most Outslopes (EMO), and the Cap.  Prior to dumping, the area north of the pit was logged.  During initial years of mining,  considerable till from overburden stripping in addition to waste rock was deposited on the dump. The till by nature contains significantly more sand to silt sized particles than does waste rock and has net acid consuming properties (Section 3.4). As a result the basal ten to twenty metres of almost all of the North dump forms a relatively low permeability, acid consuming horizon. The EMO and Cap areas constructed later in the mine’s life contain little or no till and this factor is a likely contributor not only to their high permeability, but also to their current net acid-generating character.  2.3.3.1 Old North Dump The Old North dump was constructed from 1971 to approximately 1981 and is a mixture of about 44 percent till and 56 percent waste rock (UBC MMPE, 1990c). Some previous studies (UBC MMPE 1 990c and Li, 1991) have defined the Old North as being only the westernmost region of  the North dump, however for this study OND is defined as the lower bench of waste rock  36.5  140  76.4  Aug. 1985  April 1971  North  0.0  40  9.3  Aug. 1985  June 1984  CAP  44.4 2.4  125  (wt.%)  Till  15  5.0  Dec. 1981  April 1981  EMO  62.1  (Ha)  (x 106 tonnes)  <June 1984?  To  Area  Tonnage  April 1971  From  Construction Period  Characteristics of the North dump and its major regions (from UBC MMPE, 1990c)  OND  Dump  Table 2.1  0.21  0.27  0.30  0.19  Bulk Density  Porosity  1965  1928  1840  1981  ) 3 (kg/rn  Estimated  Estimated  28 covering an area extending from the western limit of the North dump east to the Eastern Most Outsiope dump.  From 1972 to 1975, approximately 830,000 tonnes of marginal ore was  stockpiled on the west end of the North dump. About 440,000 tonnes of till were later dumped in the same area (now known as the Old Marginal dump) when it was apparently decided to not process the marginal ore. Since the exact location of the Old Marginal dump is not clear from previous reports, for this study the Old Marginal dump is referred to as being part of the Old North dump. Initial reclamation of most of the Old North dump was conducted in 1984, and trees planted in 1987. Numerous truckloads of till were deposited on the top of the eastern end of OND during the late 1980’s for later reclamation work.  2.3.3.2 Eastern Most Outsiope The Eastern Most Outsiope was built off of the outer southeast margin of OND from April to December 1981. The dump has two levels; the lower EMO (12.08 Ha) and upper EMO (3.16 Ha). EMO material is estimated to contain less than three percent till. A water quality seepage survey during the summer of 1987 showed several low-pH seepages, and a ditch was immediately constructed to collect drainage and direct it to the existing Eastern Drainage ditch (or EDD) (BHP-Utah, 1988). Reclamation activities included seeding with a grass and legume mixture and planting alders in 1987.  2.3.3.3 Cap Material placed on the central portion of the Old North dump from June 1984 to August 1985 forms what is known as the Cap. The Cap covers 40 hectares and is comprised of the westerly Lower Cap, and the easterly Upper Cap. The Cap is composed exclusively of waste rock. Its surface has numerous oxidized patches up to twenty metres across, and some seepages from the base of the Cap are acidic. The mine currently has a permit to dump wood and scrap metal into a pit near the eastern end of the Upper Cap. Heavy grease products are disposed in a small soil lined pit a short distance to the west. Reclamation work on the Cap has been limited due to uncertainty of decommissioning strategies.  Till from South Wall Pushback pit extension  29 stripping was deposited in the late 1980’s along the south margins of both the Upper and Lower Cap in 1988. The spreading of this material across the southern slope of the Cap commenced in early 1993.  2.3.4  Hydrology  As previously mentioned, most of the North dump area (and almost the entire pit) are located within the historic End and Trey Creek watersheds (Li, 1991, Fig. 2.2). A portion of the western end of the Old North dump, and the entire North West dump are within the Stephens Creek watershed.  The development of the North dump has most significantly affected the End Creek watershed by restricting flow down the historic drainages. Although subterranean flows thought to be End Creek and Trey Creek are observed on the south margin of the North dump, the development of the dump has resulted in the formation of several ponds along its northeast margin.  Construction of the perimeter ditch system around much of the North and North West dumps from 1986 through to the present has directed all drainage southeastward. Drainage from the North West dump, situated at slightly lower elevation than the North dump, is pumped approximately 500 metres east and fed into the North dump drainage ditch.  Currently, all  drainage flows to a sump southeast of the North dump, which is drained by a culvert leading to the synthetic membrane-lined Water Management Pond situated on the Beach dump. Water Management Pond’s water is pumped to the mill for use as process water. If necessary, excess water may be released into the adjacent Exfiltration Pond when all water quality objectives of the operating permit are met. Tidal action allows gradual mixing of sea water and effluent as the water exfiltrates through the Beach dump.  The total basin area drained by the perimeter ditch system (not including the North West dump), is estimated at 279 hectares (Rescan, 1992), twice the area of the North dump.  30  2.3.5  Water Quality  Detailed water quality monitoring results of potentially impacted sites and waste rock dump perimeter stations are given in annual environmental assessment reports.  These reports,  compiled by Island Copper Mine, are reviewed by the Island Copper Mine environmental technical advisory committee.  The location of current perimeter ditch sample sites are given in Figure 2.4. Acidic effluent comparable in magnitude with typical drainage from station EMO (Table 1.1) occurs at stations NWD, NDD, EMO, EDD, and EDT.  Water samples are routinely measured for pH, total dissolved, fixed and volatile solids, alkalinity, acidity, turbidity, sulfate  nitrate, calcium and magnesium.  Metals determined  include dissolved cadmium, copper, iron, lead, manganese, molybdenum, and zinc.  These  parameters were short-listed to pH, alkalinity, acidity, sulfate, calcium, magnesium, and dissolved aluminum, cadmium, copper, and zinc for the recent hydrology and metal loadings study (Rescan, 1992). Detailed analytical procedures are given in BHP (1986).  2.3.6  Drilling  In 1988, a total of ten drill holes were completed on the North (7 holes) and Beach (3 holes) dumps using a Becker 505 percussion drill (UBC MMPE, 1990c).  Eight foot composites of drill cutting splits were analyzed for acid base accounting parameters of total sulfur and acid consuming potential. No sulfate or paste pH analyses were performed. Following completion of each hole, a slotted PVC pipe was installed to maintain the opening for future monitoring (UBC MMPE, 1990c).  Figure 2.4  Location of perimeter ditch drainage sample sites and North dump drill holes  32 In 1989, seven Becker percussion drill holes were completed on the North West dump. In addition to performing acid base accounting on drill cuttings and installing slotted PVC pipe, a basal piezometer and gas monitoring tubes at 2 metre intervals were installed down each hole.  The locations of the seven North dump drill holes are shown in Figure 2.4. AMD #1 through #3  are located on the EMO dump area, AMD #4 at the eastern end of OND was intended to intersect the old Trey Creek streambed, holes AMD #5 and #6 tested the Old Marginal dump area and western end of OND respectively; and hole AMD #7 was intended to test the Cap and possibly intersect a stream bed in the End Creek watershed (UBC MMPE, 1990c).  2.3.7  Acid Base Accounting  A total of 188 acid base accounting analyses were performed on drill cuttings from the North (68 samples), North West (66 samples), and Beach (54 samples) dump drill holes. Detailed results are given in UBC MMPE (1990a, 1990c) and Li (1991).  Assuming normal distributions, mean and standard deviations of acid base accounting results are given in Table 2.2 for both the North dump and its various areas.  The high standard deviations with respect to the means illustrate the heterogeneity of the dump material. Based on the significant variability, it is questionable whether the amount of sampling is adequate for an 84 million tonne, 140 hectare waste rock dump. The lack of adequate data is most obvious in the Cap area (7 samples from one drill hole). Nevertheless, the drill holes represent the only direct information on dump composition (UBC MMPE, 1 990c).  Overall, the acid base accounting analyses do confirm: i) the net acid generating potential of the EMO and Cap areas, and ii) the net (albeit marginal) acid consuming character of OND.  33 Table 2.2  Mean and standard deviations of acid base accounting analyses from the North  dump and its various areas  Area OND  APP  ACP  NNP  t) CaCO / (kg 3  (kg CaCO /t) 3  t) CaCO / (kg 3  ACP:APP  Mean  Std.Dev.  Mean  Std.Dev.  Mean  Std.Dev.  Mean  Std.Dev.  22.6  14.0  30.5  17.0  +7.9  15.4  1.72  1.65  51.1  23.5  18.7  11.7  -32.5  29.1  0.48  0.46  52.3  15.3  43.5  10.4  -8.8  19.5  0.91  0.41  33.6  22.1  28.5  16.6  -5.1  27.0  1.29  1.43  (n = 42) EMO (n=19) CAP (n = 7) North Dump (n  =  2.3.8  68)  Oxygen and Carbon Dioxide Monitoring  Monitoring of oxygen and carbon dioxide levels in the seven North West dump drill holes was conducted in October and November 1989, and February 1990 (UBC MMPE, 1990a). The tubing leading out of the holes, exposed to light and variations in temperature, has since disintegrated and cracked, and the system will require some refurbishing for future monitoring.  Detailed monitoring results are given in UBC MMPE (1990a). Decreased oxygen and elevated carbon dioxide levels were encountered down a number of drill holes. Very low (less than 1%) oxygen levels were consistently encountered only in one six metre interval down NWD #4; the oxygen levels in the remaining areas generally ranged from 10 to 21 percent.  Morin (1990), quoting work from Ohio State University, indicates that pyrite oxidation by oxygen is limited by oxygen levels below 2 percent. Based on the North West dump monitoring results, it is concluded that although there are occasional pockets within the dump that have oxidation rate-limiting levels, the majority of the dump contains oxygen levels that are not  34 sufficiently low to limit the rate of pyrite oxidation. Since similar waste rock and construction methods were used, the same conclusion is reached for the North dump.  2.3.9  Temperature Monitoring  Harries and Ritchie (1981) justify the use of waste rock dump temperature as an ideal method of determining pyrite oxidation rate because: i) temperatures are easily and accurately measured with relatively simple instrumentation, ii) temperatures are a measure of average heat production over a large volume of surrounding dump material, and iii) waste rock dump temperatures respond quickly (within a few weeks) to changes in oxidation rates.  Down hole temperatures were monitored using multi-thermistor strings and a portable readout device (UBC MMPE, 1 990a). Temperature monitoring was conducted on the North West dump concurrent with oxygen and carbon dioxide monitoring. At time of writing, Island Copper Mine has commenced routine temperature monitoring down the fourteen on land dump drill holes. Recent results from this dump allow the construction of a temperature time gradient to assist in -  prediction of future temperature trends.  Temperature monitoring at both the Rum Jungle site, Australia, (Harries and Ritchie, 1981) and Mine Doyon, Quebec (Université Laval, 1991) showed that at least the top four metres of the waste rock dump is subject to temperature variations caused by seasonal climatic changes. Therefore, to eliminate the effect of seasonal variation, only temperatures below four metres depth were used to calculate average down hole temperatures on Island Copper waste rock dumps.  Solely for the purpose of determining overall temperature change with time, temperature measurements below four metres depth from all seven North West dump drill holes were  35 averaged to give a single value for the monitoring period. Because they showed negligible variation, readings from October 1989 to February 1990 were taken as a single reading taken on December 1, 1989. Assuming that the rock temperature at time of dumping (January 1, 1983) was equal to the average annual temperature of 8°C, a temperature versus time relationship can be estimated (Figure 2.5). The results show a time of accelerated temperature increase between 1983 and 1989, and a relatively lower rate of increase from 1989 to 1994.  25  20  E 15 a, I C,  a, 10  5— Jan-82  Figure 2.5  Jan-85  Jan-88 Date  Jan-91  Jan-94  Average dump temperature versus time for North West dump  From the measurements, the North West dump calculated temperature gradient from December 1983 to December 1989 is +1.54°C per year, and from December 1989 to February 1994 is +0.37°C per year.  Although temperature measurements down North dump drill holes were reportedly taken in late 1989 or early 1990 (Ian Home, pers. comm.), results have not been located to date. Results of February 1993 monitoring are given in figures 2.6 to 2.8.  36  Significantly elevated temperatures were found in holes through the Cap and EMO areas, and warm moist air could be easily felt moving out of the holes. Hole AMD #7 through the Cap showed the highest temperature of 24.3°C, similar to elevated temperatures recorded in NWD #2 during the 1989 and 1990 monitoring. Holes through the OND areas showed the least elevated temperatures.  Assuming an initial dump temperature of 8°C and using an average temperature of 17.0°C obtained in March 1993, the temperature gradient for EMO from January 1981 to March 1993 is +0.74°C per year.  2.3.10 Summary  The development of the Island Copper Mine has affected local hydrology by: i) creating a discharge area by excavating the open pit, ii) impeding flow of End and Trey Creek drainages, thus creating several ponds along the northern boundary of the dump, and iii) directing surface water from the End and part of the Trey Creek drainage basins into the perimeter ditch system.  Waste rock dump construction used mainly push dumping with occasional use of free dumping methods (Morin et a!., 1991). The intermittent use of free dumping makes it difficult to predict water flow paths in the waste rock dumps.  37 30  0  ci)  a. E (1) 10 I—  0 10  5  0  --  20 15 Depth (m)  AMD#4 (east OND)  --  25  30  35  AMD#7 (upper Cap)I  Down-hole temperatures for AMD #4 (east end of Old North Dump) and AMD Figure 2.6 #7 (upper Cap)  30  0  ci) 0.  E  l10  0 0  --  Figure 2.7 EMO)  5  10  20 15 Depth (m)  25  30  35  AMD#1 (lower EMO)—_ AMD#2 (lower EMO)-E-. AMD#3 (upper EMO)  Down-hole temperatures for AMD#1 and #2 (lower EMO), and AMD #3 (upper  38  30  C) o  20  0-  0  I  •  5  10  I  20  15  I  25  I  30  35  Depth (m)  -  Figure 2.8  AMD#5 (west OND)  --  AMD#6 (west OND)  Down-hole temperatures for AMD#5 and #6 (west OND)  Perimeter ditch water quality stations consistently returning contaminated effluent are: i) NWD (draining the North West dump), ii) NDD (likely draining the Old Marginal dump), iii) EMO (draining the EMO dump area), iv) EDD (draining EMO and the eastern flank of OND and the Cap dump areas), and v) EDT (combination of EDD and Trey Creek subterranean flow).  Acid base accounting results indicate that the OND dump area has the lowest potential for net acid generation. This region contains significant till, thus creating a potentially acid consuming, and relatively low permeability horizon in the basal ten metres of the North dump. The EMO dump area has virtually no till and acid base accounting analyses show it to have net acid generation potential. The Cap region apparently has less net acid generation potential, but limited data are available for this area.  39 It is concluded that, based on oxygen measurements from North West dump montioring, that there is sufficient oxygen in almost all areas of the dump for pyrite oxidation. There is also evidence of convective transport of air through the dump.  Elevated temperatures (indicative of enhanced pyrite oxidation rates) have been documented down several North and North West dump drill holes.  Calculations of North West dump  monitoring data spanning five years indicates that the temperature gradient is decreasing. Due to lack of historic monitoring data, it is uncertain whether this gradient can be applied to the Cap and EMO regions of the North dump. Future temperature monitoring of the waste rock dumps will provide valuable data for modeling of future effluent quality trends.  40  3.0  WASTE ROCK CHARACTERIZATION  3.1  Objectives  The objectives the waste rock characterization study were: i) to characterize waste rock material with respect to rock type and alteration, geochemistry, and physical characteristics such as grain size distribution, ii) to determine degree of weathering that had already occurred in the rock after 8 to 12 years in the dump, and iii) to determine effluent quality from various rock units under controlled laboratory weathering conditions.  3.2  Methods and Procedures  3.2.1  Sampling  A total of eight waste rock and two till samples were taken from the North dump in 1992. Details of each sample site are given in Table 3.1, and site locations are shown in Figure 3.1.  During phase 1 sampling in April 1992, four 100 to 140 kilogram samples were collected for column test work. The Cap and EMO dump areas were selected as source areas because based on dump seepage monitoring, these areas were currently generating the most significant ARD. Three out of four sites were located adjacent to existing drill holes, as the drill holes provided acid base accounting values to guide in selection of a sample each of material currently acid generating, and material with some acid-generating potential from the two dumps.  10  10  5  5 5  Sept. 1992 Sept. 1992 Sept. 1992 April 1992  April 1992  26300 28300 30000 NA;composite sample NA;composite sample  8300 8300 7100 NA;composite sample NA;composite sample  OND-W  CAP-L  OND-E  CAP-U  OND-E  7  8  Ti  T2  Abbreviations: EMO CAP-L CAP-U  Easternmost Outsiope Lower Cap Upper Cap  10  10  10  OND-E Old north dump, east end OND-W Old north dump, west end NA Not Applicable  -  -  -  10  6  -  5  Sept. 1992  25200  8600  5  OND-W  EMO EMO  3 4  -  100- 140 100-140  -  April 1992 April1992  CAP-U CAP-U  1 2 31600 32500  Dump Area  Site  6100 6100  Date Sampled  Approx. Sample Mass (kg) 100 140 100 140  April 1992 April 1992  Waste rock sample summary Island Copper Grid Coordinates Easting Northing (feet) (feet) 30100 7700 29700 7800  Table 3.1  1-1 .5m pits; column testing 1-1.5m pits; column and humidity cell testing 1-1.5m pits; columntesting 1-1.5mpits;columnand humidity cell testing 0.30m pits; humidity cell testing 0.30m pits; humidity cell testing 0.30m pits; humidity cell testing 0.30m pits; humidity cell testing composite of >10 till piles on dump; chem. & phys. characterization composite of >10 till piles on dump; chem. & phys. characterization  Comments  Figure 3.1  Location of waste rock sample sites  43 A “grade-all” excavator was used to obtain the four samples and place them in barrels. Depths of sites 1, 3, and 4 sampling pits were about ito 1.5 metres, and only 0.3 metres for site 2.  The material was placed into barrels double-lined with heavy duty plastic bags, sealed and immediately shipped to the UBC Department of Mining and Mineral Process Engineering.  Photographs were taken of each excavation and of the waste rock profile in each pit. Continuous channel samples were taken down one or more profiles in the pit and samples were immediately sealed in plastic for subsequent moisture content determination. Three to four hand specimens of the rock types at each site were taken for identification and further study.  Two ten kilogram composite till samples were also taken in April 1992 from 10 to 15 piles each on the Upper Caps (sample Ti) and east end of Old North dump (sample T2).  The second phase of sampling consisted of four additional waste rock samples (sites 5 through 8) for humidity cell testing, and were obtained in September 1992 after material from the first four sites had been examined and column testing had commenced. Selection criteria was based on both obtaining better representation of the various Island Copper waste rock units, and better spatial representation of the dump as a whole. Samples were generally taken from the top 0.3 metres of the dump surface on sideslope areas where there was minimal vegetation.  Selected samples from sites 1 through 4 and site 6, and from waste rock taken from the southeast area of the 640 bench in the open pit were used in a study relating field weathering and ARD prediction by acid-base accounting (Lister et a!,, 1993).  44 3.2.2  Date of Mining  The date of mining was determined by referring to waste rock dumping records compiled by Li (1988). The compilation consists of a map of the dump and pit area for every month since commencement of operations, and based on availability of information, shows both source areas in the pit, and destination areas in the dump. For some months, poor records were kept, and the source and/or destination information is missing.  3.2.3  Rock Type and Mineralogy Assessment  The rock type and mineralogy of the sample was assessed through hand sample and limited thin section and x-ray diffraction (XRD) analysis on selected samples.  Hand specimens taken from sites 1 through 4 at time of sampling were examined and described. A selection of these rocks were also examined for confirmation purposes by J. Fleming, chief geologist at Island Copper. In addition, a number of rocks were randomly selected from each site’s sample upon its arrival at UBC (10 samples each for sites 1 through 4, and 5 samples each for sites 5 through 8). Each sample was classified as to rock type and alteration, amount and mode of occurrence of sulfide minerals, amount and occurrence of calcite, colour of weathered products, rock competence, and presence of zeolite, gypsum, feldspar, or other accessory minerals of interest.  Ten polished thin sections were made from a suite of mainly highly oxidized samples selected from the various sites.  These were microscopically examined by Dr. C. Leitch (an ore  petrologist with extensive experience in Island Copper geology) for similar characteristics as the hand sample classification. In addition, modal estimates of each sample’s mineral assemblage and an estimate of degree of pyrite oxidation were compiled.  The interaction with both Mr.  Fleming and Dr. Leitch provided a solid basis for the author’s own interpretation of the somewhat complex alterations observed in many of the hand samples.  45 Limited XRD analysis (four samples) were conducted on weathered products from surfaces of the hand samples. Analyses were done at the UBC Department of Geological Sciences using a Siemens D5000 x-ray powder diffractometer.  3.2.4  Geochemistry  Geochemical analyses consisted of 30-element inductively coupled plasma (ICP), acid-base accounting (ABA), and acid-volatilized carbon dioxide.  2 analyses were initially crushed, split if Samples for ICP, ABA, and acid-volatilized CO required, and pulverized using UBC MMPE facilities. Pulverized samples were stored in either kraft or plastic bags.  A 10 to 20 gram pulp of each sample was submitted to Acme Analytical Laboratories in Vancouver who conducted a thirty-element ICP scan on a 0.500 gram sample digested in aqua regia. The aqua regia leach is considered partial for Mn, Fe, Sr, Ca, P, La, Cr, Mg, Ba, Ti, B, W, and limited for Na, K, and Al.  Acid-base accounting was conducted at the UBC MMPE department. Acid consuming potential (ACP) was conducted according to BHP (1986) procedures, which are an adapted version of the Sobek et al. (1978) procedure. Total sulfur was gravimetrically determined by oxidation to sulfuric acid using nitric and hydrochloric acid digestion, followed by precipitation with barium chloride. Sulfate was similarly determined, with the exception that a weak acid (10 percent HC1) digest solution was used to extract soluble sulfate for barium chloride precipitation.  46  Paste pH was determined using a 2:1 soil to water mixture.  2 (as carbonate minerals) was determined using a Coulometrics Coulometer. Acid-volatilized CO 2 gas released when 2N perchioric acid was added This instrument measured the quantity of CO to a pre-weighed mass of sample.  3.2.5  Physical Analysis  Air-dried moisture content was determined for profile samples taken from sites 1 through 4 by recording initial mass, air-drying the sample for 36 hours at approximately 30°C, and re weighing.  Dry sieve analysis was conducted on samples from all sites using a .ñ series of sieves.  3.2.6  Replicate Analyses  Replicate geochemical and physical analyses were conducted on the waste rock samples to determine variability due to sampling and laboratory technique. Where possible, estimates of precision were made on the replicate data set. The method adopted involved first calculating the absolute value of the difference in result between each replicate set replicates (Lix), and then calculating the standard deviation of the replicate data set. The precision values presented in Table 3.2 are two standard deviations of the summed Lx values; thus, nineteen times out of twenty (or 95 percent of the time), a duplicate analysis will yield a value similar to the original result within plus or minus the precision indicated.  47 Calculated precision of selected chemical analyses of waste rock  Table 3.2  Parameter  # Replicates  Total Sulfur  2  ±O.4wt%  Sulfate  2  ±O.lwt%  Paste pH  11  ±  0.22 pH units  Acid Consuming Potential  7  ±  t CaCO / 10 kg 3  2 Acid-Volatilized CO  5  ±  2 0.03 wt% CO  3.3  Results  3.3.1  Overall Observations  Calculated Precision  Figure 3.2 shows the profile of the 1.2 metre deep sample pit at Site 4 on lower EMO. Evidence of sulfide oxidation is indicated by yellow and dark rusty brown staining. At most sample sites, a considerable amount of fines were agglomerated onto the surfaces of the larger rock particles.  Figure 3.3 shows an attempt at obtaining a sample in an obviously highly oxidized zone on the Upper Cap.  The excavator encountered an indurated horizon approximately 15 centimetres  below the surface. The horizon could be seen on a dump scarp about ten metres from the site as a distinct oxidation “front” into the pile, with relatively fresh, unoxidized material below. A second attempt in another oxidized area on the Upper Cap produced similar results, only the indurated horizon was 30 centimetres below the surface.  As the excavator was unable to  penetrate further, the site 2 sample was taken from the upper oxidized horizon.  00 ‘1-  .4.  Q  0  Q  -4-.  E  0 0  a)  I)  d  a) —  0  -  I  a) a)  —  :. c  -  -4-.  a)  I  49  Figure 3.3  Indurated horizon below highly oxidized zone on upper Cap  50 Reconnaissance on the Cap area during sampling revealed that the most of the acid-generating rocks were highly altered with fine-grained disseminated pyrite.  In contrast, rocks with less  pervasive hydrothermal alteration, but still with significant pyrite content, showed relatively little evidence of pyrite oxidation (Figures 3.4 and  3.5).  In addition, the acid-generating rock  types were frequently found intermingled with Pyrophyllite Breccia, which tends to contain little pyrite, but has negligible acid-neutralizing capabilities (BHP  3.3.2  and Rescan,  1988).  Date of Mining  Estimated dates of mining, and in-pit source for the eight sites are given in Table 3.3.  Table 3.3  Site  Estimated date of mining and in-pit source for waste rock sample sites  Estimated  August 1984  -  Estimated In-Pit Source(s)  Mining  Number 1  Date of April 1985  2  August 1984  3  Sept 1981  Northwest 1040,1080,1120 benches Northwest 1120 bench  West 600,800,840,1280,1320 benches East 600, 1080, 1120 benches  4  West 600,800,840,1280,1320 benches  Sept 1981  East 600, 1080, 1120 benches  1983  5  Sept  6  Oct 1983  7 8  Jan June  -  -  May  Northwest 1160 bench Northwest 1160, 1200 benches  1984  August  1985  Northwest and West 1160 bench West 1040 and 1080 benches  51  Sericite-chiorite-clay (SCC) altered volcanic? on Upper Cap, Figure 3.4 disseminated pyrite that is obviously oxidizing.  Weakly altered volcanic on Upper Cap, with Figure 3.5 controlled pyrite showing little evidence of oxidation.  medium-grained  with finely  ,  fracture  yellow  white  white,  2  3  4  yellow  Lithology BVAN BVAT BXBV BXPY PPQF VEIN  VEIN-40%  BVPY-60%  PPQF-40%  Bonanza Voic., Undiff. Bonanza Voic., Tuff Bonanza VoIc., Breccia Pyrophyllite Breccia Rhyodacite Porphyry Vein material  yellow  white,  100%  rust,  7  8  BVAN-  yellow  6  100%  white  BVAN-  BVAN-80%  BXPY-70%  BVAN-60%  100%  5  yellow  white  BVAN-  Lithology  Surface  Colour  Dominant  Weathered  Minerals blo chi ep dumort ep he pyroph qz  pyroph  SCC  SCC  ser-clay  qz-ser  bio-mt  pyroph  SCC  epidote  Alteration  Dominant  biotite chlorite chalcopyrite dumortierite epidote hematite pyrophyllite quartz  BVAN-20%  PPQF-20%  BVAT-20%  PPQF-20%  PPQF-30%  BXPY-10%  BXBV-1 0%  BVAT-20%  BVAT-100%  Lithologies  Other  2.0  1.0  12  0.5  1.5  tr  3  diss  vein  diss>  vein  diss>  nil  nil  nil  nil  weak  nil>  nil  mod  comp  mod  soft>  v.comp  comp  mod>  mod  soft  mod>  v.comp  comp>  Competence  Physical  Sulfide habitlabundance tr trace diss disseminated > more common than HCI Reaction str strong mod moderate weak weak nil no reaction  dumort  ep  gypsum  sp-0.5%  diss  weak  zeo-veins  mod  zeo-veins  gypsum  chl  bio, he, ep,  zeo-veins  Noted  Minerals  Other  cp-trace  weak  nil  nil  str  Vein  vein>  mod  nil  weak  mod  Matrix  ep,pyrp,bio  cp-trace  cp-trace  Sulfides  Other  vein  diss>  diss  diss  diss  Habit  Avg. % tr  Pyrite  Primary  Pyrite  Visual  HCI Reaction  Alteration ser sericite bio-mt biotite-magnetite sphalerite pyroph pyrophyllite sp zeolite zeo qz-ser quartz-sericite Competence SCC sericite-chl-clay v.comp very competent camp moderate mod. moderately competent soft soft  SCC  qz-ser  SCC  qz-ser  pyroph  qz-ser  qz-ser  Alteration  Other  Rock classification summary of the eight waste rock samples  I  Site  Table 3.4  53  3.3.3  Rock Type and Mineralogy Assessment  Table 3.4 summarizes rock characteristics obtained from hand specimen examination of several randomly selected samples from each site.  As mentioned, all specimens had a considerable amount of fine-grained particles agglomerated to their surfaces.  Colours of both the agglomerated fines and the actual weathered sample  surface were either white-gray, yellow, or dark brown rust. Numerous specimens, especially those from sites 4 and 8 had both white and yellow mottled surfaces. Specimens from sites 2, 6, and 7 had entirely yellow or dark brown rust surfaces, indicative of significant sulfide oxidation.  It can be seen that Bonanza Volcanic lithologies predominate the sample suite, followed by Hydrothermal Breccias, and finally, Rhyodacite Porphyry material. The specimens examined were variably altered.  Pyrite occured mainly in disseminated form while chalcopyrite and  molybdenite were rarely observed.  Significant quantities of vein-occurring sphalerite were  found in two of the five hand specimens from site 5. The highest amount of pyrite (12 percent on average) was observed in specimens from site 6.  With the exception of sites 1, 4, and 5, the specimens examined generally showed weak to nil effervescence with ten percent hydrochloric acid. Accessory minerals observed included zeolite in veins in site 1, 4 and 5 material, and two millimetre long gypsum needles on highly oxidized surfaces of sites 2 and 7 specimens.  Competence of the rock, qualitatively assessed by its hardness and ease of breaking, varied considerably.  3.3.4  Mineralogy  Petrographic descriptions of a suite of mainly highly oxidized samples is given in Appendix 1.  54  The study confirmed the primary habit of pyrite in waste rock as disseminated subhedral to euhedral one to two millimetre crystals. Modal estimates of pyrite varied from zero (in one site 8 specimen) to 75 percent (in a specimen of site 6 vein material), but generally ranged from two to ten percent.  The most common mineral in the highly oxidized suite was sericite, followed by chlorite, clay minerals, and quartz. Traces of rutile occurred in many samples, and occasional dumortierite, apatite, magnetite, zeolite and carbonate were observed in one or two samples. Significant quantities of feldspar were only found in a sample from site 8; in all others primary feldspar was totally replaced by sericite.  Samples of agglomerated fines adhering to outer surfaces of four samples were analysed by x ray diffraction. Results are given in Table 3.5.  Unexpectedly, no jarosite (iron sulfate hydroxide) minerals were found in the yellow to brown weathering material. It is possible that amorphous iron hydroxide minerals (undetectable by x ray diffraction) are instead the main product of sulfide oxidation.  With the exception of  gypsum, the suite of minerals adhering to the rock surfaces are identical to those found in the rock mass.  Therefore, the agglomerated fines appear to be mainly derived from either  mechanical breakdown of rock due to mining or physical weathering, and are not chemical weathering products.  Both petrographic and x-ray diffraction studies indicate that although some samples show considerable surface oxidation, negligible pyrite and other mineral weathering has occurred within the rock mass.  55  Table 3.5  X-ray diffraction results from rock surface material  Sample  Site  Appearance  Minerals Identified  X3-1A  3  gray-white  quartz pyrophyllite nacrite (clay mineral)  X4-2A  4  medium brown  quartz muscovite (sericite) gypsum clinochlore (chlorite)  X4-3A2  4  white  muscovite (sericite) quartz laumonite (zeolite) nimite (chlorite) gypsum  X4-2C2  4  yellow  quartz gypsum muscovite (sericite) clinochiore (chlorite)  3.3.5  Elemental Analysis  Results of thirty-element analyses performed on samples from the eight waste rock sites and two till composites are given in Appendix 3.  The major mineral-forming elements analysed included aluminum, calcium, magnesium, potassium, and sodium. The highest aluminum values were from sites 1 and 7 (3.5 to 4.2 percent), with the lowest, site 8 at 0.8 percent. Pyrophyllite, the primary mineral in site 8 rock, is an aluminum-poor mineral relative to feldspar and sericite, which are primary constituents of site 1 and 7 rock.  Highest calcium values were found in sites with little or no evidence of sulfide oxidation (sites 1 and 5, till composites), and were in the range of 2.1 to 3.5 percent. Lowest values, less than 0.7  56 percent occurred in site 2, 6 and 8 samples.  These results correlate well with the dilute  hydrochloric acid reactivity of each site (Table 3.3), and indicate that most calcium probably occurs as calcite or dolomite. Magnesium values follow a similar trend as calcium, except that site 7 returned high magnesium values.  This may be attributable to the presence of  ferromagnesium minerals.  Highest potassium values (0.12 to 0.2 1%) tended to occur in the obviously acid generating sites 2 and 6. Lowest values (0.04 to 0.06%) were from sites 3, 5 and 8 samples. The opposite appeared to be true for sodium; slightly higher sodium values (greater than 0.1%) occurred in sites 1 and 5, and the till composites, while the highly oxidized sites 2, 6 and 7 were relatively depleted of sodium. High potassium values can be attributed to the presence of considerable sericite, while sodium values may be due to the presence of plagioclase feldspar.  Heavy metals of significance to Island Copper include copper, iron, manganese, molybdenum, nickel, cobalt, zinc, and cadmium.  Highest copper concentrations (427 to 847 ppm) were  encountered in sites 3, 4 and 6; site 8 contained the lowest amount of copper (65 ppm). Highest iron values (13.14 to 13.87%) were found in the highly pyritic sites 6 and 7, with the lowest in sites 3 and 8 (1.56 to 1.59%). High manganese, probably indicative of mafic minerals, occurred in sites 1, 5 and 7 (3192 to 6709 ppm). The lowest manganese levels (401 to 694 ppm) occurred in sites 3, 4 and 8. Molybdenum values were low in all samples, ranging from 3 to 32 parts per million. High nickel (85 ppm) concentrations were found in site 5 material, with the remaining sites ranging from 9 to 30 parts per million. Cobalt levels were greater than 20 parts per million in sites 1, 2, 5, 6 and 7. Highest zinc (5225 ppm) occurred in sphalerite-rich site 5 material, and site 1 also returned anomalous values of 740 to 1125 parts per million. Lowest values (83 to 180 ppm) were found in the till composites and site 4 material. Cadmium, occurring as a trace element in sphalerite, had a similar pattern as zinc, but with concentrations ranging from 0.2 to 32.3 parts per million.  57 Scatter plots of the various heavy metals indicate strong positive correlations between cadmium and zinc, iron and cobalt, iron and nickel, and nickel and cobalt.  This suggests that i) as  mentioned above, cadmium occurs as a trace element in sphalerite, and ii) pyrite contains traces of nickel and cobalt as either inclusions or in solid solution.  3.3.6  Acid Base Accounting  Acid base accounting parameters analysed for included total sulfur, sulfate, acid consuming potential, and paste pH, and acid-volatilized carbon dioxide. Results are given in Table 3.6.  Anomalously high total sulfur values occurred in site 6 material, and the remainder of the waste rock samples sites had total sulfur contents ranging from 0.59 (site 3) to 4.3 (site 7) percent. No detectable sulfate was found in samples from sites I and 3; sites 6 and 7 had significant sulfate contents (over 1%). Two types of acid producing potential (APP) were determined: i) from total . The calculation methods for (APPS ) sulfur (APPS), and ii) from total sulfur minus sulfate 2 2 are given in Appendix 4. APPS and APPS  Acid consuming potential (ACP) varied considerably. Negative values were encountered in a sample from site 6, indicating accumulation of considerable acidic products on the material. tonne. CaCO / The highest ACP value, found in siteS material, was 110.2 kg 3  Net neutralization potential (NNP, the difference between ACP and APP) of the samples tended to be low. The NNP values obtained for samples with low paste pH may be slightly inaccurate, since the oxidation products contain sulfate and are accounted for in the APPS calculation, and also cause a decrease in the ACP value due to their acidic nature. Only two of the eight waste rock sites and the two till composites had positive NNP values. kg CaCO /tonne) was obtained from site 6 material. 3  2 The lowest NNPS  (- 261.4  NNPS  579 3.6 8.2 6.7 21.8 17.8  40.6 93.3 76.6 77.9 18.4 29.4  40.6 101.6 82.8 84.4 18.4 29.4 22.5 40.6 71.9  <0.02 0.26 0.20 0.21 <0.02 <0.02 <0.02 0.11 0.17 0.14  1.30  3.25  2.65  2.70  0.59  0.94  0.72  1.30  2.30  1.60  2.95  1  2  2  2  3  3  3  4  4  4  2  4  3  4  5  6  7  8  9  10  11  12  22  23  24  0.04  -68.4 -71.2  -74.6 -77.6  -11.9 9.5 10.5  53.5 249.5 120.3 29.4 8.5  54.7 278.1 134.4 31.3 9.4  0.04 0.92 0.45 0.06 0.03  1.75  8.90  4.30  1.00  0.30  5  6  7  8  Till  25  26  27  28  OND  Till  14.4  110.2  52.3  59.4  0.23  1.90  15.0  -19.5  -26.5  32.8  85.2  92.2  0.22  0.02  -77.2 -84.2  22.9  45.5  50.0  0.48  0.50  29  0.44 -22.7 -27.1  29.1  66.5  U.Cáp  0.59 -37.4 -42.8  21.9  37.3  44.1  44.6  36.1  -20.8 35.2  29.7  -18.9  -124.9  29.1  -0.05 -110.8  -290.0  3.06  5.25  0.36  0.08  2.06  56.6 -261.4  55.5  0.63  0.09  0.92 -15.4  -18.7  20.7  22.5  8.0  0.61 -1.8  -1.8  8.07  7.85  4.84  2.87  2.38  7.77  6.60  3.50  6.62  6.38  6.82  7.08  6.76  6.99  1.18 -11.6  3.46  3.73  3.38  7.96  7.76  0.19  0.11  1.43  -89.7  -98.0  1.60  17.3  17.3  -11.6  62.5  39.1  39.1  <0.02  7.91  1.65  23.4  pH  2 APPS  (kg CaCO3/t) 29.4  Paste  ACP:  2 NNPS  23.4  (kg CaCO3It) (kg CaCO3/t) 29.4 747  1.25  (kg CaCO3It)  ACP  1  2  (kg CaCO3/t)  2 APPS 453  as S)  APPS 453  (wt%  4 SO <0.02  1  1  (wt%)  Total S  0.79  1.18  0.04  0.01  0.02  3.54  0.56  0.04  0.38  0.41  0.43  0.12  0.06  0.12  0.08  0.1  0.06  1.39  0.96  1.95  CO 2 (wt%)  Acid base accounting and acid-volatilized carbon dioxide analyses of the eight waste rock samples and two till composites  1.45  Site #  Samp#  Table 3.6  Co  59  On average, 93 percent of total sulfur is estimated to be in sulfide form with the remainder as 2 is about 85 percent of the mean NNPS. sulfate. Carrying this to NNP values, the mean NNPS  Paste pH is considered to be indicative of current acid generating conditions in the sample, and if a sample has a paste pH of less than 4.5, it is considered to be generating net acidity. Paste pH values were below 4.5 in sites 2, 6, 7 samples, and a marginally acid generating pH value of 4.8 was obtained from site 8 material. The highest paste pH was from site 1 material (7.98 pH).  , and 1 (0.96 to wt%C0 ) Acid-volatilized carbon dioxide analyses were highest in sites 5 (3.54 2 . Low, but detectable values were encountered in the strongly oxidized sites 2, 6 1.95 2 wt%C0 ) . wt%C0 ) and 7 samples (0.01 to 0.10 2  There is a reasonable correlation between acid-  volatilized carbon dioxide and acid consuming potential (Figure 3.6). Linear regression of this relationship gives the following equation:  mol/kg) C0 ( 2  =  -  -20  •,  0  20  60  40  80  100  120  ACP (kg CaCO3It) • ACP vs. C02  Figure 3.6  (3.1)  t)} 0.055 CaCO / 0.0066[ACP(kg 3  —  (Linear Fit)  Acid consuming potential (ACP) versus acid-volatilized carbon dioxide for the  waste rock samples  60  3.3.7  Physical Analysis  Physical characterization included particle size analysis, moisture content, and surface area.  Sampling of sites 1 through 4 was constrained to take that portion of the dump material that could easily be placed in the sample barrels, and the material obtained contained fragments up to 5 inches (12.70 cm) in size. During sample preparation for column testing, the minus one inch fraction was segregated from the rest of the waste rock sample. The amount of material between one and five inches (2.54 and 12.70 cm) in size was measured to range from 42.0 to 50.6 percent of the total sample mass. Extrapolating this to the entire size range of particles in the waste rock dump, it is conservatively estimated that 50 weight percent of the dump is comprised of fragments coarser than the one inch (2.54 cm) and less fraction used in kinetic testing.  Detailed particle size analysis results for the eight waste rock sites and two till composite samples are given in Appendix 5. Only the minus one inch (2.54 cm) portion of the waste rock sample was used in sites 1 through 4 analysis, the minus one half inch (1.27 cm) portion was used for sites 5 through 8, and the entire sample (up to 2 inch or 5.08 cm particles) was used for the till composites. Problems of static agglomeration of fines and blinding of mesh openings were encountered during dry sieving of the till composite samples for mesh sizes of 70# (0.212 mm) and below. Figure 3.7 and 3.8 show the range of particle sizes obtained from the waste rock and till composites, respectively.  Comparisons between Figures 3.7 and 3.8 indicates that the till samples have significantly more material finer than 0.5 millimetres than do the waste rock samples. Wet sieving, or another appropriate method, is required to obtain the lower portion of the till particle size distribution curve.  •00  •0  C)  -5  CD  C  z  0  -I  CD  0  0  0  0  -a  p  0---  p  —  I 11111 11111 11111 11111  I 11111 I 11111 11111 11111  I I  I 11111 I 11111 11111 11111 It III 11111  I I I  I I I I  IltIll I  I I I  I I I I  I I I I I I  I I  111111 I 11111 111111 111111 111111  I I I I  11111 lull 111111 111111 I 11111 I 11111 111111 I  I  I  I  I I I I  I  I- —1—4-4-1-14-i— ——4  thu  Illill  I  Itt  I I  lilt lIlt lilt  lilt I I  I  I  11%  lIttJ  11111 11111  I Ill 11111  lilill 11111  I  I  I  I I  I I  II II  -  11111 I  11111  -a 0 0  111111 1111111 II llll 111111 II ttiil 1111111 ll 11111 111111 111111 11111) 11111 11111 111111  4--I  3 j3  I  I I I I t I I I I I I  __  I  I I  I I I  I  I I I I i  I- II II Ill II Ill II Ill II lilt 1111 I Ill liii I I Ill 11111 I 11111 I 11111 I I 1111  1111111 I 11111111 1111111 I I I 11111 iJi_J_J_I_lJLll I...J_ I I 111111 Ill I I lilt I I I I I I I Ill I 111111 I I lllli 1111111 I llttIlII I I IIII I I 11111 I 11111 111111 I ‘ I 11111 I I 11111 11111 I I I 11111 11111 lilt II St 11111 II II 11111 I I Ill I II tlIl t 1 1 L ! 411111 II iCit1— I 11111 II II I >I CD )I tlll II It I II It I Ill l I I I I Il Ill I II II 0 II ti I II It II () CDt I It II Ii tIl I I ltl II I I I I I tlI I II It 1111111 II I I I I tlt I II It I I I II 111111 14  I I  1111111  I  I 11111 I 11111 11111111 t I  1 I 11111 1111111 1111111 I I 11111 I I 11111  I I I I I  I 11111 1111111  11111111  III  11111  0  -.  Wt. % Passing  I  I  CD  -4  T1  -5  -  3  I-  CD  0  I-  0)  -5  CD  C  CD  CD  0  -a 0 0  0  -a  p -a  0— -a  p  lilitI.  IlIltI  5  51  %I  I  ttttit  111111  I  I  I  I I  1111111 1111111 1111111  1111111  1111111  1111111 1111111 t 1111111 1111111 1111111  I  I I I I I  I  I  I  I 511111 I lltIt!S 1411111 I 11111111k SI 1111111 I 151111 lillill t I 1111111 I S 15111 t IllIlt I 11111 t\t ttt4tit 1511 ISlI I tltltt. •It 5tt I I lllIllj lr.I l’4t 11111111 I  I 1111111 lSl lltIt 1111111 I i\l I lIlj 1111111 411 lt, I 1 1111111 t\llt tl\ I I 1111111 15111! I I I I ItllI 1111111 I St I 151111 l 1111111 I I IitIt 1111111 SI t I tISIlItI I I 1111111 lt5lll 4 1111111 h 151 111111 I I l\l 111111 I I tliI\tl _ii_l_II.I_tJ. I lIill!u ftltltIl I IllIllIl I 1111111% t 14111111 I I ttltli! I  l1—I4-IH4-  141111  t\lltIl  thu l%lhhttI 4,t  I I I I I I I  0  -a 0  _LLL1LlJL___l_J_I_I_tiLt__J_LJ4JJ. l 1111111 I I I lilt 1111111 I I I lt\llt I 1111111, I IltIll I I I 11151 I IllIll I llthll I IttlIlt 1111111 I t tll!,j I I t l!lIll I tiltiS I hull 1111111 I It IllIll I 111111 1111111 I I l1llll 1111111 I I 1111111 1111111 I 111111 I I I I 111111 I I 1111111 I I 111111 I 111111 I I I IllutI I I I 1111111 11111 I 1111111 I I I IIIIIt  I Ittilt I ttlItI 1111111 1111111  1111111 1111111 I IttitI  1111111 1111111 1111111 I Iltltt  I  111111 1111111 1111111 1111111 1111111 1111111 1111111 1111111 I Ittlil 1111111  -a 0 I lltIll 1111111 I 111th I lilItl titttitl 1111111 I hull  I I’l  I  I  I  I  I  —t—-ti-t-I1-———Ss—_4  I  lllltt 1111111  I  1111111 tl’I4.tI  1111111 1111111  1111111 1111111  llIIj  -a  Wt. % Passing  0  62 Moisture contents from sites 1 through 4 sample pit profiles are given in Appendix 6. Site 2 material had the highest moisture content of 6.9 weight percent in the top oxidized 30 centimetres, and 9.3 weight percent from the underlying indurated horizon.  High moisture  contents in the top 8 centimetres of the two site 3 profiles (7.9 and 9.8 wt%) are likely a result of high water retention by organic soil and root matter on the EMO dump. Below this horizon, and in sites 1 and 4 profiles, moisture contents ranged from 1.5 to 4.1 weight percent.  No  relationship between moisture content and depth was observed.  3.4  Summary  Based on paste pH analysis results, sites 2, 6, and 7 are deemed to be currently net acid generating.  Site 8, at pH 4.8, is classified as marginally acid generating.  The three acid  generating sites all have obvious yellow to rust brown iron oxide staining completely covering exposed surfaces. Sulfur content of site 6 and 7 material is high in comparison with the rest of the waste rock sample suite, however sulfur content of site 2 material is only marginally anomalous. Samples from all of the acid generating sites appear to be strongly hydrothermally altered, and based on acid-volatilized carbon dioxide analysis, contain very low levels of carbonate minerals.  Sites not currently acid generating vary greatly in mineralogy and geochemistry. Sites 3 and 8 have acid-volatilized carbon dioxide contents comparable to the low levels found in the acid generating sites, and are interpreted to have little excess buffering capacity.  Petrographic analysis of selected samples indicates that sulfide oxidation is restricted to exposed rock surfaces and is rarely present in fractures. Therefore, a very small proportion of the total contained pyrite is presently oxidizing.  63 Variable levels of heavy metals are present in each of the eight sites’ samples.  Based on  correlations from elemental analysis, traces of cadmium are present in sphalerite, and traces of cobalt and nickel are occurring in pyrite.  Geochemical analysis of the two till composites indicates that they are low in sulfur (0.30 to /tonne), and moderate to 3 0.48 wt %), have moderate acid consuming potential (44 kg CaCO . In comparison with the eight waste C0 ) high carbonate mineral content (0.70 to 1.18 wt% 2 rock sites, the till is considered to have sufficient excess alkalinity to some degree of buffer infiltrating acidic drainage.  64 4.0  KINETIC TEST WORK  4.1  Introduction and Background  Unlike static tests which attempt to predict ultimate acid generation potential, kinetic tests attempt to predict the longer term weathering characteristics of a waste material and leachate water quality as a function of time (Lawrence, 1990b).  The objectives of the Island Copper waste rock kinetic test program were: i)  to quantify aqueous chemical loads of eight waste rock samples from the North dump,  ii)  to give indication of variation of water quality with infiltration rate, and  iii)  to document temporal variation of leachate quality over the test duration.  Waste rock from the eight sample sites described in section 3.0 were weathered in four column test and eight humidity cell experiments. The experiments produced a significant amount of data, some of which are extraneous to the specific goals of this study. All data obtained from kinetic testing is presented in this report, making this particular chapter a considerable length. However, the kinetic test results are considered to be a major contribution to the overall understanding of acid rock drainage at Island Copper, and will likely be used in subsequent studies.  The overall objectives of ARD kinetic tests are (Steffen Roberston and Kirsten, 1989):  i)  to establish rate and temporal variation of acid generation and water quality of a sample on a continuous basis,  ii)  to confirm static test results, and  iii)  to test treatment and mitigation option options.  65 Because acid generation is a reaction which often takes years to evolve in the field, workers conducting laboratory kinetic tests attempt to accelerate the process in order to collect the maximum data on long term weathering in the shortest amount of time. Several enhancement methods can be considered:  i)  exposing higher surface area (such as crushing sample),  ii)  increasing infiltration rate (Ritcey and Silver, 1982),  iii)  pre-acidification of test (Lawrence, 1990),  iv)  inoculation of test with T.ferrooxidans (Lawrence, 1990),  v)  increasing temperature of test,  vi)  increasing oxygen concentration in surrounding atmosphere, and  vii)  supplementing test with humidified air.  While most of the above mentioned implementations do appear to increase the rate of acid production, it is difficult to quantify exactly how much or how uniformly the oxidation reaction has been accelerated.  Ritcey and Silver (1982) contend that their lysimeter test work on  uranium tailings was accelerated ninefold by infiltrating the tests at nine times the average rainfall rate, thus implying that in their case, removal of oxidation products is the controlling factor in the rate of acid production. Caruccio and Geidel (1981) performed small scale (300500 gram samples) leaching tests subjecting the sample to humidified air and rinsing with deionized water every 3 to 5 days. No estimate of temporal acceleration was given.  Other workers in the field of dump leach modeling contend that the rate determining factor in acid generation is oxygen availability (Davis and Ritchie, 1986). T.ferrooxidans also appears to have a direct effect on oxidation of pyrite.  The presence of Based on these  contentions it seems that predictable test acceleration is complex and at the present time incompletely understood.  66  Perry (1985) is skeptical of accurate interpretation of leaching tests for several reasons: i)  test conditions are generally designed to promote maximum rates and amounts of pyrite weathering, thus giving “worst case” results,  ii)  actual period of weathering simulated is uncertain and might only be simulating short term conditions,  iii)  results are method specific and cannot be compared directly with other leaching tests,  iv)  scaling factors (method and site specific) must be applied to correlate lab results with field conditions, and  v)  leaching tests were originally designed for studies of solid and hazardous waste landfills and as such are not necessarily applicable to waste rock studies.  Lapakko (1990) recognized that due to the expense and time required for kinetic test work, tests may be run for durations that fall short of the time required for depletion of acid neutralization potential. Nonetheless, he showed that once leachate concentrations reached steady state, the test time to depletion of acid neutralization potential can be calculated. However, Lapakko did not predict actual field weathering rates from the kinetic test rates. Ferguson and Morin (1992) presented a similar method of calculating time to sulfide and neutralization potential depletion. In addition, they compared results from a 30 tonne waste rock test pile with humidity cell test results from the same material and found that the time to peak sulfate concentration was much shorter for the humidity cell, and that the humidity cell released sulfur at a much higher rate. However, results suggested that the test pile would ultimately produce more sulfate, possibly due to field conditions causing more physical breakdown of particles than occurs in humidity cells.  Morwijk (1993) compared results from Bell Copper’s humidity cells, 4 inch diameter columns, 24 inch diameter columns and 10 tonne on-site waste rock pads, and found that chemical loading rates for the different types of tests were comparable.  Renton et al. (1988) found that the  reverse of laboratory acceleration occurred; their field tests weathered almost 9 times faster than  67 their bench scale tests. These two examples illustrate the variety of results that can be obtained and the potential difficulties in extrapolating kinetic test results to waste rock dumps.  As waste rock material weathers in the field, individual rock particles slake, or break down, exposing more surface to oxidizing conditions, and thermodynamically unstable minerals dissolve or alter to more stable compounds. This process occurs at widely varying rates for different rock types and is an inherent part of the soil forming mechanism, taking tens of years, or even centuries to occur to any appreciable extent. As the residual mineral composition of the rock changes, leachate from the waste rock changes as well. From the literature reviewed and consulting various workers in the ARD field, insignificant physical weathering has been observed during kinetic tests, and thus it appears that the tests do not simulate this long term aspect of weathering. Given this, one should bear in mind that at best ARD kinetic tests can optimistically simulate only the initial stages of waste rock weathering.  The limitations of unknown acceleration of field conditions by kinetic tests, and the short term weathering simulated must be kept in mind when considering the design, results, and interpretation of Island Copper waste rock kinetic tests.  4.2  Column Tests  Column testing was conducted on the four large scale waste rock samples from the Caps and EMO dumps (sites 1 through 4). The experiment was initiated on September 4, 1992 and was terminated on June 11, 1993 for three of the four columns, and on July 9,1993 for the final column.  68 4.2.1  Objectives  The objectives of the Island Copper Mine column test work were: i)  to quantify aqueous chemical loads of four waste rock samples from the Caps and EMO areas of the North Dump, and  ii)  to give indication of variation of water quality with infiltration rate, and  iii)  to document temporal variation of leachate quality over the test duration.  4.2.2  Equipment and Procedure  Clear plexiglass columns 1.9 metres high and 15 centimetres outside diameter (14 cm inside diameter) (Fig. 4.1) were constructed for the tests. The rock was supported at the base with a perforated PVC plate lying across stainless steel rods. The entire column was suspended from the top from two additional stainless steel rods. One half inch (12.5 mm ) diameter ports were installed at 30 centimetre intervals, with rubber septums sealing the ports.  Distilled water was dripped into each of the columns at a slow rate (0.36-1.13 mI/mm) using a Masterfiex low rpm multi-channel peristaltic pump. Each of the four columns had its own source container in order to monitor accurately water influx to each sample. Leachate was collected in plastic pails placed at the base of each column.  The experiment was intended to be run for approximately 40 weeks. At the end of 40 weeks, column tests 1 through 3 were terminated, while column test 4 was continued for another four weeks.  Since the experiment was designed to determine present effluent emanating from the waste rock samples, no additional crushing was performed. To reduce channeling due to large particles (Van Zyl et al., 1988), only the minus one inch (2.5 cm) portion of each waste rock sample was placed in the columns. Approximately 50 percent of the waste rock dump’s mass is estimated to  69  Distilled Water  1/32” LD. Tygon tubing rods  wall X 6’ plexiglass tubing 6” l.D.  1/411  Waste rock material  K  —  Septums for gas sampling  Glass wool Plexiglass I PVC perforated plate 2 X 1/211 stainless steel rods V  Collection vessel Approx. Scale: 1”  Figure 4.1  Column test apparatus  =  10”  70 be above 1 inch in size (Section 3.3.6), however, due to their relative surface areas, the amount of weathering products produced by the coarse fraction is negligible compared with the finer fraction. Therefore, barring any other scaling and laboratory factors, the calculated weekly loads per unit mass from the column (and humidity cell) tests will approximately double the amounts obtained on a sample of the entire grain size distribution of the waste rock.  Prior to testing, samples were characterized chemically, and physically, using three representative two to four kilogram splits from each of the pre-test samples. A fourth split of similar size was sealed in plastic and archived for possible future reference. Detailed procedures and results of the pre-test characterization are given in chapter 3.0.  Water infiltration rates for the column tests were determined through literature surveys, analysis of Island Copper and Port Hardy airport precipitation data, and consultation with UBC faculty and Island Copper Mine staff.  The waste dumps are subject to varied intensities of rainfall (see Section 2.1.3), but due to the site’s latitude and elevation, no snowpack accumulation occurs during winter months. The rainfall “seasons” were simulated in the column test with periods of intense, moderate, and low infiltration rates. This ensured that both maximum flushing of oxidation products occurred occasionally through all possible flow paths in the column, and that accumulation of oxidation products was possible during moderate and low infiltration intervals.  At the same time,  infiltration rates were not varied so much as to make results difficult to interpret.  Examination of the annual precipitation patterns in the Island Copper area showed that the year can be, without excessive distortion, divided up into time periods of moderate (January to June), light (July to September), and heavy (October to December) rainfall (see Fig. 2.1). Infiltration into the colunms was varied in this manner as well, compressing one year’s precipitation into a four week cycle. This enabled the experiment to run through at least ten yearly precipitation  71 cycles in the 40 week test period. The infiltration schedule is shown in Table 4.1. As discussed in Section 4.1, it is difficult to quantify the amount of reaction acceleration, if any, as a result of the tenfold flow rate increase over natural conditions. The high flow rate does however, reduce the chance of reactions being limited due to solubility constraints.  Table 4.1  Distilled water infiltration schedule for colunm tests  Weekly Cycle #  Infiltration Level  Monthly Field  Column  Total Infiltration  Rainfall Simulated  Infiltration Rate  Volume  (mm)  (mI/mill)  (I)  1  Moderate  140.2  0.58  5.8  2  Moderate  140.2  0.58  5.8  3  Low  65.1  0.36  3.6  4  Heavy  248.2  1.13  11.4  Using an average porosity of 0.32 (see Table 4.2) and knowing that the rock in each column occupied a volume of 0.028 cubic metres, each four-week cycle infiltrated almost three pore volumes of water through the column.  Pore gas samples were extracted in 60 cubic centimetre plastic syringes via the sampling ports during the final weeks of testing (weeks 39 and 40 for columns 1 through 3, and weeks 39 through 44 and July 28, 1993 for column 4). Oxygen analysis was conducted on the samples using a PE Series 104 gas chromatograph, and a Coulometrics Coulometer was used to determine carbon dioxide content of the sample.  Average moisture content was determined by weighing each column using a heavy duty overhead spring balance at leachate sampling times. Weighing was initially conducted on a  72 weekly basis up to week 26, after which the frequency was reduced to approximately every four weeks.  The experimental work was conducted at the ambient temperature in the Centre for Coal and Mineral Processing (CMP) building (average 21°C). Due to the relatively small diameter of the column, core temperatures were not expected to be significantly elevated, and were not monitored.  Before filling, the columns were thoroughly cleaned first using laboratory detergent, followed by distilled water rinsing, rinsing with a ten percent nitric acid solution, and final distilled water rinsing. Once dry, the columns were weighed. Waste rock was then placed in columns using a 4” PVC pipe insert to minimize damage to the inside of the columns. The columns were filled in about one foot lifts, and the outside of the columns pounded with a rubber mallet to help settle the material.  Glass wool mats were placed above and below the waste rock. The lower mat acted as a fines filter for the infiltrating water, and the upper mat helped to disperse the distilled water as it dripped into the column.  The dry weight of each column’s material was determined by re-weighing the column after filling. Water infiltration at 0.58 ml/min began on September 4, 1992. Initial wetting of the columns was closely monitored by taking periodic photographs and measurements during the first 72 hours of infiltration.  Columns were monitored at least every other day and the volume of water pumped recorded. At the end of each week (infiltration cycle), the columns were weighed and the leachate collected and its volume recorded.  73 Less than two litres of leachate were required for all analytical work. Approximately 1.5 litres was filtered through 0.45pm cellulose nitrate, transferred to 1 litre bottles, preserved with concentrated nitric acid, and shipped to either Island Copper Mine or Analytical Services Laboratory, Vancouver for dissolved metals analysis. The remaining 0.5 litres of the sample was analyzed at the UBC Department of Mining and Mineral Process Engineering laboratory facilities for sulfate, alkalinity, acidity, conductivity, pH, and Eh. Details of analytical methods and estimated precisions are given in Appendices 7 and 8, respectively.  Leaching of columns 1 through 3 was terminated on June 11, 1993 and on July 9 for column 4. The columns were allowed to stand in place for approximately 2 weeks to allow accumulated water to drain.  Contents of each colunm were then emptied into semicircular troughs  constructed of 10 inch (25 cm) PVC pipe cut in half lengthwise, and markers put in every 25 centimetres. The material in the troughs was photographed immediately following placement.  Three two kilogram post-test composite samples were collected over the entire length of each column. The material was split twice using a Jones riffle, one fraction pulverized for chemical analysis, and one fraction stored in plastic for future reference. Post-test chemical analyses included acid-base accounting, and 30-element ICP,  4.2.3  Results  4.2.3.1 General Observations Initial conditions of the material each colunm are given in Table 4.2.  Overall, the experiment was relatively simple to run and maintain. There were problems in calibrating the peristaltic pump precisely to the design flow rate, but frequent monitoring allowed adjustments in flow to be made throughout the week.  74 At an initial flow rate of 0.58 millilitres per minute, columns I and 4 took approximately 90 hours for effluent “breakthrough” at the bottom of the column.  Columns 3 and 2 took  significantly longer (120 and 140 hours, respectively).  Table 4.2  Waste rock conditions in column tests at start-up  Mass Dry Waste  Column 1  Column 2  Column 3  Column 4  53.3  55.8  48.9  49.4  2785  2750  2725  2730  0.019  0.020  0.018  0.018  0.028  0.028  0.028  0.028  0.32  0.29  0.36  0.36  Rock (kg) Waste Rock ) 3 Density (kg/rn CaIc. Waste Rock ) 3 Volume (m Volume Occupied ) 3 in Column (m CaIc. Porosity  Logically, the time to effluent breakthrough for each of the columns is directly related to their respective moisture contents during the experiment. Columns I and 4 had the lower moisture contents ranging from 6.0 to 8.5 weight percent (column 1), and from 6.4 to 8.2 weight percent (column 4). Moisture contents for column 2 ranged from 9.5 to 11.0 weight percent, and for column 3 from 13.0 to 15.1 weight percent. content as the experiment progressed.  All columns showed an increase in moisture  75  4.2.3.2 Weekly Leachate Quality Tabulated leachate quality results for the column tests are given in Appendix 9.  Leachate pH conditions of each column remained approximately constant throughout the duration of testing (Fig. 4.2). Leachate from columns 1, 3 and 4 returned near- or above neutral (6.91 to 8.76) pH levels, while column 2 leachate remained acidic (2.13 to 2.82). Although pH values for each column appeared to fluctuate, no pattern of variation with infiltration rate was observed.  Measurement of leachate Eh did not commence until week 5 due to equipment availability. Similar to pH, leachate Eh remained approximately constant for each column during the course of the experiment, with column 2 returning consistently high values (767 to 921 mV H°), and columns 1, 3 and 4 remaining in the more moderate range (413 to 640 mV H°) (Fig. 4.3). The considerable variation in the values is likely more reflective of the difficulty in measuring Eh consistently, rather than an actual change in leachate quality.  Leachate conductivity values for each column varied from the beginning to end of the testing (Fig. 4.4), with values initially high, then plateauing as the experiment progressed. Levels in columns 1 through 3 leachate became more consistent by week 12, while column 4 levels remained high until week 25 before sharply dropping. All column leachate showed highest conductivity values during low infiltration periods in the four week cycle. Lowest values were observed from leachate collected the week after the high infiltration period.  Leachate sulfate concentrations appear to mirror conductivity trends for the columns (Fig. 4.4 and 4.5).  Sample conductivity was used to determine the appropriate dilution for sulfate  analyses which considerably reduced lab work. As with conductivity, sulfate concentrations were highest during low infiltration periods, and lowest in the week following high infiltration. Sulfate concentration values for all four columns show a decreasing trend during the experiment.  76 10  50  8  40  6  30  4  20  -  Week I0h1 _+_coI2_’i._coI3__CoI4  Colunm test leachate pH  Figure 4.2  50  1000  Z  I  700J  600  0  ___  -  2  20.  LU 500  _  10 400  --  ——  j 1’1’12  Week --coI1 ._.—coI2 ——coI3 .-—co(4  Figure 4.3  Column test leachate Eh  UI  CD  I  8  CD CD  -4  C.)  C.) C.)  7’) IC  7’) 0  1%)  0  I I  I  I  0  —  11111 I lillil I 11111 •Thi  IH’’  111111  till 111111 IllI I 11111  I I lIlt 1111111  1111111  11111  I I I I I I  I I I I I  I  I I I I I I I I I II I  1111111 1111111 1111111 I I 111111  II I I I 111111 I  CO  m  7’) 0  C.) 0  I I I I lIIl I I II I III I I II I III I I II III I I I I I II I I II I I II III 11 I I Ill I I I iiiti% Ill III IIIIII tllIl II HI Ill llIlIl II IIlIlI Illil ii I I dr1IlII I I4I hull Ill hlII l6IillllI II i),iiiiiit 1 ii iIlIIIII ii. I Iii 1lIIlIIIl 1 II II l I I II II IlIllIIII IIlIllIl liii lIIIlIIlI IIIIIIII II hIll tI I I IIIIII i) IIIIIII I I II hill  11111111 I 1111111 III II I I III I IIIIlI. I 1111 I 11111 I 11111 lIlt II II I III I I Il lI I I I I II I II I I IlI  r’3  m  I I  I  I  I  .H.  m  I  I 111111  I  I I I I I I  I 1111111  I 1111111 1111111 I IllIlhI I 1111111 I 1111111 I hIlhIll  0  111111 111111  1111111 I 111111 III hilt 1111111 1111111 111111 I 1111111  I I  I I I I I I I I I I I I I I  111111 Illill Iii lilt  I I  liii Ill lilt  I  I I  I  111111  I  I I I I  111111  111111  II till I 11111  Ill III  111111 111111 111111 111111 111111 111111  111111 I 11111 111111 111111 111111  I I 111111 I I 111111 I I 111111  0  C,’  Cfl  m  I 111111  I I I I I I I I I I I I I I  I I 11111  I I I I I I I I I I  I I I I I I I  I hIll ii liii IIIIII hill 11111  I  I  I  I I  I I  I  II lIlt  I  I lIhill I III lilt  I IhlIlIl  I Ill IIIl.,.—TiI Ilill I I III’ I I I hIll I I III lIlt I I IllIllI I I  Infiltration Volume (I)  I  I  I  I  111111 III II 1111111  111111 111111 I 1)11111  IC  C.  I III I 11111 •I 111111  11111 111111  ;:  —  m  C,  0  m  Sulfate (mgIl)  C.)  0  CD C.)  CD  CD  0  (Th  -  CD  8  8  c3  CD CD  C.)  is.)  (.11  Cl  C’  V  •  1mw  -  p  I I I I I  iii,, ii till iii 1111 I liii hIll liii  -  it  IHI  0  —  Conductivity (mS/cm)  Infiltration Volume (I)  Djiii 1 i  fiui  liii liii iii I 11111 11111 11111 1111 11TH 11111 1111111 1111 I 111111 Huh 11111111 1111111 I. I 111111 S. Hilili IIIIII IllIll I HIll 111111 111111 111111 I I 111111 11111111 iiIIIhl 11111 1111111 I I 1111111 111111 1111111 1111111 111111 I III I I I liii I lii 11111111 111111 11111 I I II IlIluui I 111111 1111111 11111114 1111111 SwSIIulIIIl 1111111 I 111111 11111111 1111111 I .11111 ‘H.i 1111111 111111 I 1111111 1111111 1111111 11111111 1111111 111111 I 1111111 111111 Ii 11111 h%’ 11111 1111111 lI I lull 1111111 HIIIiiI I 11111 • 1111 itiI4 I I ill I 111111 1)1111 III III I I 11111 11111 IIIll I I IIIil 1111111 II 111111 1111111 I Il_l 111111 I I 11111111 iuff I it 1111 IlIl 11111 IIIil I I I I I H ill 1111111 I I I 111111 111111 IIIl I I III: I_JI Iii 111111 • I I I 1111111 I I IJIII I I 1111111 1111111 I 11111111 I I IIIII  0 0  II  111111 1111111 I 11111 1111111 I 111111 1111111 1111111 1111111  I I  I  I 111111  1111111  I  I  I I 111111 I I 111111 I 1111111 I 1111111 I I 1111111 I I III I I 111111 I 1111111 11111111 I I 111111 I 1111111 I 1111111 I I 11111 I I 11111 I 1111111 I I 111111  I I I  I  1111111  I  Ii 11111 I  II I I  I 11111 11111 1111111  I I I I I  I  I I I I  I  I-I  ‘I’m  0  78  250  200  —  —  .40  c)  o C-) (0  C_)  ci) .30  150  >  0)  E  100  -  --  -  .20.2  -  ___  50  —  21  22’337’’41  Week  I Column test leachate alkalinity  Week  21 Figure 4.7  10  0  0  Figure 4.6  E  -  Column test leachate acidity  79  Leachate alkalinity analyses were performed on columns 1, 3 and 4, are shown in Figure 4.6. Values varied considerably during the course of the experiment, with a general trend towards decreasing alkalinity for each of the three columns. Column 4 leachate showed the largest drop /l, then leveling off towards the end of the test to 3 in alkalinity, starting at 240 mg CaCO 1. Leachate alkalinity for each column had a distinct trend CaCO I between about 70 to 110 mg 3  with the four week infiltration cycle. Columns 1 and 4 leachate showed a similar pattern of peak alkalinity during high infiltration, followed by a drop during the first week of moderate infiltration, followed in the next week by a slight increase, and finally sharply dropping to its four-week trough during low infiltration. Column 3 leachate showed a different and much more subtle pattern with the infiltration cycle. Peak levels of alkalinity consistently occurred during the low infiltration week, and the lowest alkalinity concentrations usually occurred in leachate from the week following the high infiltration period.  Leachate acidity was determined only for column 2 leachate, due to the high pH levels of the l) for the first CaCO / other three columns. Initial acidity levels were extremely high (26,000 mg 3 four weeks of testing (Fig. 4.7).  Leachate acidity concentrations were highest during low  infiltration periods and lowest in the week following high infiltration periods. Acidity values appear to gradually decrease from about week 25 to the end of the experiment.  Leachate dissolved metal concentrations for fourteen selected metals are shown in Figures 4.8 to 4.10. Leachate from the first 12 weeks of testing was analysed by atomic absorption at 1CM environmental laboratory, and subsequent weeks by 30-element ICP at Analytical Service Laboratory (ASL) in Vancouver.  The initial analyses, though a smaller metal suite, had  considerably lower detection limits than the subsequent ICP analyses.  600  8  12  12  16  16  24  20  24  Week  20  28  26  32  32  36  36  40  40  44  44  d)  C)  —  6  8  10  0  0.1  0  4  4  8  8  I_-co  12  12  24  24  Week  20  Week  20  28  28  32  32  1__CoI2YCoI3..x_CoI4I  16  16  Column test leachate dissolved metals: a) aluminum, b) calcium, c) cadmium, and d) copper  Week  0 0  0  2  04  4  8  100  200  300  400  •0  T  4  E  500  600  E  E  0  b)  C)  0.2  0.3  0  ‘:  C)  800  1000  Figure 4.8  c)  a)  36  36  40  40  44  44  C  00  E  C  0)  0  16  .24  32  120  U.. 240  C)  E a,  360  480  600  0  4  4  i.  Figure 4.9  c)  a)  12  16  24  Week  20  12  16  24  Week  20  .  .  28  28  32  32  36  36  40  40  44  44  d)  E  0)  0)  E  0)  0  0.02  0.04  0.06  0.08  0.1  10  20  40  0•  a  x  4  8  IIIT  -.-  12  Coil  16  —  24  Col 2  Week  Week  20  32  Col 3 -x- Ccl 4  28  Column test leachate dissolved metals: a) iron, b) magnesium, c) manganese, and d) molybdenum  8  1A.__._EIETIEII]IEIE1EXI.  8  b)  36  40  44  00  d)  b)  0  0.4  4  8  Iul  12  16  24  Week  20  Week  Column test leachate dissolved metals: a) sodium, b) nickel, c) phosphorus, and d) strontium  Week  0. 44 36 40 28 32 20 24 16 12 6 4 0  I____  Figure 4.10  c)  Week  a)___  28  32  36  40  44  I-  0.01  0.015  0.02  0.025  0.03  0  Figure 4.11  a)  4  12  16  24  Week  20  28  32  36  40  44  E  30  10  N 20  0)  40  50  0  Co1unin test leachate dissolved metals: a) titanium, and b) zinc  8  b)  4  8  16  24  Week  20  28  32  36  j...-CoIl _i_CoI2.._CoI3x..CoI4j  12  40  44  Co  84 Relative to columns 1, 3 and 4, column 2 leachate consistently returned anomalous concentrations of dissolved aluminum, copper, cadmium, iron, nickel, and zinc. Concentrations gradually decreased over the course of the experiment.  Additional anomalous metal  concentrations for column 2 leachate not shown in the figures include cobalt, chromium and lithium. With the exception of dissolved zinc, column 1, 3 and 4 leachate tended to have concentrations of these nine metals below the ICP detection limit. Of these three columns, dissolved zinc concentrations were highest in column 3 (0.03 to 0.05 mg/l range), followed by column 1 (<0.005 to 0.045 mg/i).  Dissolved zinc levels in column 4 leachate were low;  dropping below the detection limit from week 31 to the termination of the experiment. In contrast to column 2, column 1, 3 and 4 leachate contained measurable, though still low, concentrations of molybdenum (up to 0.109 mg/i).  Other dissolved metals of interest include calcium, magnesium, manganese, sodium, phosphorus, strontium, and titanium.  Column 2 and 4 leachate calcium levels were initially high, corresponding with high sulfate concentration. These levels are likely reflective of dissolution of accumulated gypsum and other soluble sulfate minerals. Towards the end of the test, calcium concentrations reduced to levels in the same order of magnitude as columns 1 and 3.  Similar initial trends can be seen for dissolved magnesium concentration in the column leachate, however column 2 leachate concentrations began to rise again after week 18, and continually increased until the end of the experiment. Concentrations of dissolved titanium in colunm 2 also show a corresponding dramatic increase after week 18.  Dissolved manganese was only present in consistently detectable quantities in column 2 leachate, and showed a slight increasing trend towards the end of the experiment.  85  Detectable levels of dissolved phosphorus were present only in column 2 leachate, showed maximum concentrations during the middle period of testing, and again had reduced concentrations from week 36 to 40.  Dissolved strontium concentrations remained below 1 mg/i for column 1 through 3 leachate. Only column 4 leachate had high initial levels of dissolved strontium.  Dissolved sodium was one of the few parameters that showed near equal concentrations in each column’s leachate. Bonn et. al (1985) note that during initial silicate weathering processes, alkali and alkaline earth ions are released into the soil solution.  Providing that there is  sufficient water influx, both sodium and potassium tend to not re-precipitate as secondary minerals and thus may be detectable in the resulting leachate. In the case of the column tests, the resulting concentrations of these two metals is considered to be reflective of the weathering rates of the primary silicate minerals, namely feldspars.  Although dissolved sodium  concentrations were consistently above the ICP detection limit of 2.0 mg/i, potassium values were not.  Leachate analyses using atomic absorption methods from weeks 10 through 12  returned dissolved potassium concentrations between 0.1 and 0.5 mg/i for the four columns.  4.2.3.3 Loads and Cumulative Loads Calculated loading and cumulative loading for sulfate and alkalinity or acidity are plotted in Figures 4.12 and 4.13.  Columns 1 and 3 both produced low sulfate loading rates and cumulative loads (less than 10 mg/kg/week, and less than 600 mg/kg over 40 weeks). Column 2 consistently produced high sulfate loading (more than 200 mg/kg/week).  Column 4 was transitional, and normally  produced from 100 to 350 mg/kg/week up to week 23. After this time, sulfate production sharply declined, and in the final four week cycle, only 16 to 37 mg/kg/week was produced.  0  100  C/)  50  0 a,  100  0 a, o 150 -J a) a)  U)  a,  C  8)  150  0  Figure 4.12  c)  a)  4  12  16  24  Week  20  28  32  2’2” 3  Week  2:  36  40  h’  44  0  d)  0 4  8  12 16  24  28  32  36  40  44  Sulfate Load (mg/kg/wk)  Week  Week  20  Sulfate (mg/I) Cumm. Sulfate Load (mg!kg)  0  500  1000  1500  2000  0  2000  4000  6000  8000  -v-.  C/)  0 a’  a, a,  a,  U)  a) (  U,  a)  -J  10000  Column test leachate sulfate loads: a) column 1, b) column 2, c) column 3, and d) column 4  8  “WI”  100  E E  200w  0  I  400  b)  a, 0  E E  C/)  a) a,  1000 C)  2000  3000  4000  5000 D  RflC)fl  00  Cu  t  >.  0  .  -J  Cu 0  0  .3  0  150  200  LQV  100  150  4  Figure 4.13  c)  a)  12  16  24  Week  Week  20  26  32  36  40  44  400E  5003  600  IVY  C)  < E  Cu  C  ‘S  -J  0  -,-  .  d)  cu  0  50  100  150  200  250  0  2000  4000  6000  8000  0  4  8  12  16  Concentration (mg CaCO3II) Cumm. Load (mg CaCO3Ikg)  IC  Cu  C  ‘  0  -J  0  V  0  >.  •0 C)  -J  Cu 0  10000  24  -  28  32  36  40  44  Cu  t  cu 0  0  ‘S  E E z  cu  C  ‘S  -J  C-)  ‘AnD  4000  6000  8000  ioooo3  12000  14000  Weekly Load (mg CaCO3IkgIwk)  Week  20  Column test leachate alkalinity and acidity loads: a) colunm 1, b) colunm 2, c) column 3, and d) column 4  8  200  300  500  b)  —a  00  88 Column 2 weekly acidity loads were in the 150 to 350 mg 3 /CaCO kg/week range.  The  alkalinity loading plots (Fig. 4.13a, 4.13b, and 4.13d) for columns 1, 3 and 4 are plotted on identical scales. Weekly loads were similar for columns 1 and 4 (5 to 25 mg CaCO Ikg/week), 3 /kg/week. 3 while column 3 had a significantly lower alkalinity loading rate of 3 to 10 mg CaCO The contrast between acidity and alkalinity loads illustrates the low buffering potential of the non-acid generating waste rock.  4.2.3.4 Pore Gas Analysis Selected results of pore gas analysis are shown in Figures 4.14 and 4.15, and tabulated results are given in Appendix 10.  Measured oxygen levels between weeks were similar for each  column. Oxygen levels decreased with depth in the top half of all columns, and with the exception of column 4, values leveled off or increased towards the bottom of each column. Column 2 had markedly decreased oxygen levels (approximately 10 mol %) within 0.2 metres from the top, and decreased to below detection limit (approximately 0.5 mol %) at 0.5 metres. Oxygen levels gradually increased down the column to about 18 percent at 1.6 metres depth. Columns 1 and 3 had only slightly reduced oxygen levels (approximately 19.5 mol %) in the centre region of each column.  Column 4 had steadily decreasing oxygen levels down the  column, with the minimum value of 13.6 mole percent measured at the bottom sampling port.  Carbon dioxide levels in cells were highest during initial sampling in week 39 (especially in Columns 3 and 4), and subsequently declined. Column 4 pore gas was sampled on July 28 (19 days after previous sampling) to determine whether or not carbon dioxide levels were depleting as a result of sampling, or by reasons related to weathering reactions within the column. Carbon dioxide levels on July 28 were significantly higher than those measured in weeks 40 through 44, but were lower than the initial levels in week 39 (Figure 4.16). This indicates that levels were probably depleting as a result of sampling, and that four to six weeks is likely the maximum sampling frequency for meaningful carbon dioxide pore gas measurement.  89 25  2O!  15.  Distance from Column Top (m) --CoIl _--Co12 -—CoI3 -—CoI4 1  Oxygen content of colunm pore gas, week 39  Figure 4.14  4,  0  E  ci)  0  , 2 x 0  0 .0 I a)  C-)  ---  0.2  0.4  0.6  0.8  1  1.2  1.4  Distance from Column Top (m) --CoIl __CoI2 --CoI3 -—CoI4  Figure 4.15  Carbon dioxide content of column pore gas, week 39  1.6  1.8  90 2.5  2 0  E  1.5 x  0  121 C 0 .0 1 Cu  () 0.5  0 0.8  0.5  0.2  1.1  1.4  1.7  Distance from Column Top (m)  __  Week 39  Figure 4.16  -+—  Week 40 -—Week41 -*-Week42 —*—Week43 -x_Week44  Carbon dioxide content of column 4 pore gas, weeks 39 to 44, and July 28, 1993  4.2.3.5 Pre- and Post- Test Analysis of Column Material Comparisons between pre- and post- test acid-base accounting analyses are shown in Figure 4.17. Obvious visual differences between pre- and post- test results are apparent for sulfate, paste pH, and acid consuming potential (ACP).  Sulfate content of column 2 waste rock appears to have increased as a result of kinetic testing, however the reverse appears true for column 4 material. Column 1 and 3 pre- and post-test samples had sulfate levels below the detection limit.  Sulfate content of the three pre-test samples of column 2 waste rock appear to be significantly less than the post-test samples.  Conversely, column 4 waste rock pre-test samples are  significantly higher in sulfate than the post-test samples. Neither pre- nor post- test samples of columns 1 and 3 contained detectable levels of sulfate.  8  0  a-  CU  (1) 6  I—  2  3  .  •  I  I  Post  Col 2  Pre  I  I  Post  Col 3  Pre  III  1  I  I  —  Post  Col 4  Pre  1’  Cot 2 Col 3  Cot 4  d)  b)  0.2  ()  0  <20  C.)  0)  C.)  ()  0  60  80  v  Cl) 0.1  4  Pre- and post- test acid base accounting of column test material  Cot I  re ost Pre P 1 P  .  .  Post  Co! I  Pre  Figure 4.17  c)  a)  .  I  i  Post  Col 2  Pre  I  I  Post  i  Col 3  Pre  I  Col 4  Pre  Post  Col I  Col 2  Col 3  Cot 4  re ost ost’Pre Pre P P 1 P  I  I  Post  Col I  Pre  92  Post-test paste pH values for column 4 are significantly higher than corresponding pre-test results. Column 2 shows a similar trend, however a replicate reading of a pre-test sample (done concurrent with the post-test analyses) returned a higher pH value than previously recorded. No significant change in paste pH was noted for columns 1 and 3.  Column 4 post-test acid consuming potential (ACP) values were slightly higher than pre-test values. In view of the higher paste pH and lower sulfate content in the post- test samples this indicates that low pre-test ACP values are due to considerable accumulated acidic products on the samples before column testing.  4.3  Humidity Cell Tests  Humidity cell testing was conducted on eight samples of waste rock from six different sites. Testing was initiated on August 3, 1993 and terminated on September 14 after 42 days of continuous operation.  4.3.1  Background  Humidity cells are an industry-accepted method of determining acid generation potential of waste rock. Although cell design varies considerably, the underlying principle of the test is the same; waste rock is subjected to three different conditions, usually over a 7 day period: i) dry air passed through the sample for 3 days, ii) humid air passed through the sample for 3 days, and ii) leaching of the sample in the cell on the final day of the cycle (Lawrence, 1990).  Previous humidity cell testing on Island Copper waste rock includes five samples of cuttings from drill holes from the North West dump which were tested for 10 weeks in 1990. Results of this work are documented in UBC MMPE (1990a).  93  4.3.2  Objectives  The objectives of the humidity cell tests were: i) to indicate leachate quality currently being produced by various Island Copper waste rock types, and ii) to document temporal variation of leachate quality over the test duration.  4.3.3  Equipment and Procedure  Equipment for the testing included the humidity cell, an air humidifier, and a compressed air source. Both cells and humidifier were constructed according to specifications described in Lawrence (1990).  The humidity cells used were constructed using an 8 inch length of 4 inch diameter plexiglass tubing. The cell had a fixed base plate with a leachate drainage hole fitted with a tubing nipple. The waste rock sample was placed inside the tubing on a perforated support plate about 1 inch above the base plate. Air entered the cell via a side port located between the base and support plates, and exited via a hole in the top cover of the cell. Several layers of screen cloth were placed on the support plate prior to putting the waste rock sample in the cell to impede the migration of fmes into the leachate.  The air humidifier used was an approximately 3 foot length of 4 inch diameter plexiglass tubing sealed at both ends, and operated in the horizontal position. The tube was filled approximately half full of water and contained two 15 centimetre long aeration stones connected via tubing to the air source and a submersible aquarium water heater laying along the length of the humidifier. Air outlets to the humidity cells were spaced equidistantly along the top surface of the humidifier and were fitted with tubing nipples. Air temperature in the cells was 23°C.  A Perkin-Elmer oil free air compressor was used to deliver air to the humidity cells.  94  One (±0.004) kilograms of minus one half inch waste rock was placed in each cell. About 0.75 kilograms of additional sample was also prepared for physical and geochemical analysis purposes. Samples tested are given in Table 4.3.  Table 4.3  Waste rock samples tested in humidity cells  Cell Number  Site Number  HC1  5  HC2  8  HC3  8  HC4  7  HC5  7  HC6  6  HC7  2  HC8  4  Descriptions of the samples are given in Chapter 3.0. Cells 3 and 5 were duplicates of cells 2 and 4 and were intended to qualitatively indicate variance due to sampling and testing. Cells 7 and 8 contained samples from sites 2 and 4, which were column-tested, and were intended to assist in correlation between column and humidity cell test results.  Testing was conducted for six weeks, thus completing six weekly cycles of alternating dry and wet air, and sample leaching. The leaching phase consisted of adding 500 millilitres of distilled water to the cell and allowing it to stand for one hour with the leachate drain closed. The leachate was then drained into a tared sample bottle which was then weighed to determine leachate volume.  95  An intense pre-flushing of each sample was conducted prior to starting the first humidity cell cycle.  This was done in order to remove accumulated soluble weathering products on the  sample. After samples were placed in the cells, the procedure consisted of adding a total of 1500 millilitres of distilled water to the cells in 500 millilitre increments.  The water was  allowed to stand for a minimum of one hour before draining and repeating the cycle a second and third time.  Pre-flushing leachate was analysed for pH, conductivity, and sulfate. The weekly leachate was analysed identically to the column tests for pH, Eh, conductivity, acidity and/or alkalinity, and sulfate using the UBC Department of Mining and Mineral Process Engineering facilities. Thirty element dissolved metals analyses (via ICP) was performed on two-week composites by Analytical Service Laboratory, Vancouver.  Testing commenced on August 3, 1993 and was terminated on September 14 after 42 days of continuous operation, with leachate samples being taken weekly.  4.3.4  Results  4.3.4.1 General Observations Overall, it was found that air flow conditions in the humidity cell were difficult to control, possibly causing varying concentrations of soluble salts to occur. Each sample in its cell had a distinct air permeability, and thus different air flow rates were required to sufficiently supply each cell. Varying amounts of water (up to 75 ml) often condensed in the base of the cell during the humidified air cycle, thus causing dilution of the leachate. Mechanical problems were also encountered with the compressor, however these were resolved early in the testing.  96 4.3.4.2 Weekly Leachate Quality Tabulated leachate quality results for the humidity cell tests are given in Appendix 9.  Similar to the column tests, leachate pH conditions for each cell remained approximately constant throughout the duration of testing (Fig. 4.18).  Only cells 1 and 8 had pH-neutral  leachate, cell 2 and 3 leachate had weakly acidic (jH 4 to 5 range) values, and cell 4 through 7 leachate was less than pH 3.4.  Leachate Eh values (Fig. 4.19) did not show any trend over the course of the experiment. Cell 1 leachate returned low Eh values consistent with non-acidic leachate, and cell 5 through 7 leachate Eh values were high (over +700 mV H°). Leachate Eh from cells 2, 3, and 8 were intermediate (+550 to +675 mV H°).  Leachate conductivity was initially high for all cells during the pre-flushing phase, but tended to plateau out once weekly leaching commenced. Some erratic values were observed (Fig. 4.20).  Leachate from all eight cells had elevated sulfate levels in the pre-flushing phase. The most acidic cells (4 through 6) produced extremely high (48,500 to 190,000 mg/I) concentrations (Fig. 4.2 1). Cells 2 and 6 had considerable variance in sulfate concentrations following the pre flushing phase, while the other six cells had relatively consistent concentrations.  Leachate alkalinity, measured only for cells 1 and 8, is shown in Figure 4.22. Cell 1 leachate l) than cell 8 (5.5 to 16.0 mg CaCO / was consistently higher in alkalinity (14.5 to 26.5 mg 3 l). No obvious temporal trends were observed for either cell. CaCO / 3  Acidity values for cell 4 and 5 leachate were both relatively high in week 1 before leveling off for the rest of the test (Fig. 4.23). Cell 6 leachate acidity appeared to be taking off in the final I for both weeks 5 and 6. CaCO / two weeks, increasing by 1000 mg 3  97 9  7  E3  C  -  -  6 0 — —  4  — _-__ ___—  2 0  = I 1  2  I 3  I 4  I 5  6  Week HC2 ....HC3  -  HC4 HC5 -HC6 --HC7 ...-HC8  Humidity cell leachate pH  Figure 4.18  1000 900  I  Week  , 1 I— -.HC2-..HC3., Figure 4.19  Humidity cell leachate Eh  I!  —  C) CD  I  I  C)  =  :r  C)  C)  a  z C) r%3  •1’  C)  *  c.  0)  0i  CDC CD  -  0  Sulfate (mg/I)  CD C)  C.) CD  C  CD  I  I C)  2: C)  C)  I C)  a  I C)  C)  II  CD  0)  01.  0  p I  )IISj4II IIIIIIIIjII  IIIIIIII 111111111411 II III IIIJI 1/ II I U II II I II I I iiili IJI JIIIII  I  f  iirt 111111111  I  iii 1111111 PI_IIllllI  4111111 11111111 I 11111 IfI 11111 I ji 11111 ifi 11111  II I lI’I I j I I J I II I L I 1 .  if  lIIIII  ILI.iIIIIII  fi  Ii  I 1111111 ,I 11111111 1/IIIIIIfII’ 11111111 I I l I 1111111 IfIIIIIlI4” I I) III_I’Il I 1111111 I II,).I I II I I I II III iftiiiiIpi I 1111111 IIlIIIIIII I 1111111 ViIi1ii 1111W I IlIIIIII 1111111 ii I I I I 1111 I I Ii I I llI Iit’IIiI 1111111 I I’ I. I’III’I.III 1111111 II III I I l•II I I I I 11111 IIIIIIl,4 I 1111111 I I I 111111 111111 I IIIjJII P i 1111111 I 11111111 ‘1111111111 II ?lIlIl II lIIlllIl 111111111111 IIIIIIIII I SI I 11)11 I 1l I 11111 I I jI II II lIl I Ii II I I 1111 ilIllIllIll II 11,111111 II ‘i)IIIIII lI_IlIlill  •i  I I lilt I I I Sill  I I I I  I I  I I I I I I I I I I I I I I I I I I I I I I I  I I I I  iII(tit ill liii I I  I III III I  I I  I I I I) I lI1(1 I  1  I i I I  I  I I  lIlIlil  I  I I  I  :1::  .1 4iu.L  I I I  !IIILII  I  I iijii\i/ 41!111111 i I I IIiiiiS( 11111111 111111 I I II 14111 I ji iij\ I I I I 14111111 III iki$i 1111111 I I igiiiili I 11111111 I I III 1111 I II I I 11111 II I I I liii I I III 111111  I  r:::::  0  -  Conductivity (mS/cm)  I  I I I I I Ill  I I I I  I  I I I I  I I I Iii I I III 11111  111111  1111111 I I 111111 I 111111  I 111111 I 111111 I I I liii I I 111111 I I 111111 I I 11111 I I I I 1111 1111111 I I 11111 I 111111 I 111111 I 111111 1111111 I I I 1111 I I I I III  I 1111111 111111 1111111  I I  1111111 I I I liii I I I liii  I  111111 1111111  -  0 0  00  99 50  —  40  -  c)  0  C-)  30  -  Week HC1 -,-HC2 -..HC3  Figure 4.22  Humidity cell leachate alkalinity  3500  Week  --HC2-..HC3-. HC4.HC5HC6HC7.HC8I  Figure 4.23  Humidity cell leachate acidity  100 Leachate dissolved metals concentrations are given in Appendix 9. Because only three bi weekly composites were analysed for each cell, results are not presented in graphical form. Similar to the column tests, leachate from the highly acidic cells (HC4 through HC7) returned consistently anomalous concentrations of dissolved aluminum, copper, cadmium, cobalt, iron, nickel and zinc.  The two weakly acidic cells (HC2 and HC3) produced concentrations of  cadmium, nickel, cobalt and zinc comparable to the highly acidic cells, but relatively reduced levels of copper, and significantly less iron and aluminum. The non-acidic cells (HC 1 and HC8) produced very low levels of heavy metals.  4.3.3.3 Loads and Cumulative Loads Figure 4.24 shows the range of sulfate, alkalinity and acidity loads obtained from the humidity cell tests. Cell 1 had the lowest sulfate and highest alkalinity loads, and cell 6 had the highest sulfate and acidity loads.  Results of the two replicate cell pairs are given in Figure 4.25 and 4.26. Similar average weekly loads were obtained for each pair, however cumulative loads between each replicate varied considerably. This is attributable to the large initial concentrations obtained during pre-flushing and the first week of the regular humidity cell cycle.  4.4  Discussion of Results  4.4.1  Summary  Twelve kinetic tests consisting of four column leach and eight humidity cells tested the weathering behavior of eight samples from various areas of the North dump.  Cl)  023456  60  —  Week  Cumm. Sulfate Load (mg/kg)  Week  70.0  C.)  E  E  Alkalinity Load (mgCaCO3/kglwk  Sulfate Load (mg/kg/week)  150  a)  4!  a)  •0 (0 0  d)  b)  .  3  3  5  -  6  •0 a)  4000  —x Acidity Load (mgCaCO3IkgIwk)  Week • Acidity (mg CaCO3/l) Cumm. Acid. Load (mgCaCOSikg) .  C  .4000  .  5000  • 5000  -  gnnn  0  4  —x Sulfate Load (mglkg!wk)  Week • Sulfate (mg/I) Cumm. Sulfate Load (mg/kg)  2  nnn  Cl)  4!  0 a)  C,,  4!  (0 0 -J a)  Selected humidity cell loads: a) cell 1 sulfate, b)cell 6 sulfate, c) cell 1 alkalinity, and d) cell 6 acidity  ._  Alkalinity (mg CaCO3/l) Cumm. PJk. Load (mgCaCO3Ikg)  -  * Sulfate (mg/I)  5 1001  a) (0  200  250  300  o 200  I-  C’)  .  -J a)  300  400  500  Figure 4.24  c)  a)  a  c 1000’  000  4000  5000  1500  0  -  -  1  =  Week  I 5  I 4  --  I  = --  = 6  oooo  52000  E E  5400  5600  560OQ  1400  d)  b)  .  0  1000  •2000  X00  4000  5000  1500  1  2  -  I  ._  Week  3  • Sulfate (mg/I) Cumm. Sulfate Load (mg/kg)  0  =  .  -..  Week  4  I  58000  1400  I 5  I  6  50000  2000  54000  E E  6000  Sulfate Load (mg/kgtwk)  I  Leachate sulfate loads of humidity cell replicates: a) cell 2, b) cell 3, c) cell 4, and d) cell 5  2  ---  ;:-.:  3  Week  -:  Figure 4.25  c)  a)  0  3000  4000  : :  3  100  0  0  Figure 4.26  c)  a)  1  1  I  Week  3  Week  3  I  4  4  I  5  5  I  —  —  —  6  6  !  (0  2000 0)  •0  (0 0 -J 300°  AnnA  d)  4000  A  Week  Week  4  5  6  E  0 (0  3  0  (0  2000 0)  -J 3°00’  4000  A  :°  F° E  40  80  Inn  • Acidity (mg CaCO3/I) —x Acidity Load (mgCaCO3lkglwk) Cumm. Acid. Load (mgCaCO3Ikg) -  2 1000  •0 <2000  33000  to  D  0 25  So  0 I  75  3  (0  Leachate acidity loads of humidity cell replicates: a) cell 2, b) cell 3, c) cell 4, and d) cell 5  I 2  2  100  b)  104 As expected, the pre-oxidized nature of the samples produced initially high ionic strength solutions in the first stages of testing. This is most obviously reflected in conductivity, sulfate, acidity, alkalinity and metal levels. In order to best analyse results, only leachate quality from the period after each test reached “pseudo-steady state” is considered in the following discussion. The periods used in analysis are shown in Table 4.4.  Table 4.4  Kinetic test periods used for data analysis  Weeks Used for Data Analysis  Test Coil  21-40  Co12  17-40  Coi3  21-40  Co14  37-44 1  Eight humidity cell tests conducted August -_September_1993  -  6  Leachate from sites 1, 3, 4, and 5 were pH neutral or above. Leachate from site 8 was weakly acidic (greater than pH 4.0), and the remaining three sites (2, 6, and 7) produced strongly acidic (less than pH 4.0) leachate. Table 4.5 shows the range of pH, Eh, conductivity, and sulfate, alkalinity and acidity loads produced from the twelve tests after they reached pseudo-steady state.  Table 4.5  Minimum and maximum general leachate quality parameters from kinetic tests  pH  Eh (mV HO)  Cond  Alkalinity  Acidity  kg/wk) (mgCaCO / kg/wk) 3 (mgCaCO / (mS/cm) 3  Sulfate (mg/kg/wk)  Iinimum  2.13  384  0.11  0  0  3  Iaximum  8.51  860  4.79  23  1737  2458  105  The molar ratios of alkalinity, calcium and magnesium and acidity to sulfate loads can give considerable insight on the acid generation and neutralizing reactions taking place within the sample (Ferguson and Morin, 1992, UBC MMPE, 1 990a). Molar alkalinity or calcium and magnesium to sulfate ratios indicate the amount of excess alkalinity in leachate. The decline of this ratio to unity during kinetic testing is often used as an early warning to imminent net acidity production. Molar acidity to sulfate ratios indicate the amount of acidity not internally buffered in the sample. Assuming sulfate load is reflective of the total acidity released by the rock sample, a ratio of unity indicates no internal buffering, while a near zero ratio indicates significant buffering. Ranges of these ratios obtained in the kinetic tests are shown in Figure 4.27.  3  2.5  2  1.5  I  0.5  0  Figure 4.27  mol((Ca+Mg)1S04)  mol(AIK1SO4)  mol(AcIdISO4)  Ranges of kinetic test molar calcium and magnesium, alkalinity, and acidity to  sulfate load ratios  High (greater than two) molar calcium and magnesium to sulfate load ratios consistently occured in column 1 leachate, and occasionally from column 3 and HC 8. Lowest ratios (less than 0.5) were found in column 2, HC 4, HC 5, and HC 6 leachate.  106 Molar alkalinity to sulfate load ratios were calculated only for leachate samples with measurable alkalinity.  Highest ratios (above 1.5) were found only in column 1 leachate samples.  A  considerable number of leachate samples with net neutral pH from columns 3 and 4 and HC 8 had calculated molar ratios of less than one.  This may be indicative of soluble sulfate  dissolution contributing to total measured sulfate in solution.  Molar acidity to sulfate load ratios were calculated only for leachate samples with measurable acidity. The highest ratio was 1.08 and was obtained from the first week’s leachate from HC 4; this greater than unity value may be due to stored soluble acid products still being released from the sample despite the pre-flushing. Subsequent weekly leachate from HC 4 had calculated ratios of 0.75 or less. Leachate from column 2, HC 4, HC 5, and HC 6 consistently had ratios of over 0.50, illustrating the negligible remaining neutralizing capacity of the samples. On the other hand, the very low ratios (less than 0.10) from HC 2, HC 3, and HC 8 indicate that most acidity produced by these samples is internally buffered.  Samples from sites 2 and 4 were tested both in column and humidity cells. Figure 4.28 depicts the ranges and means of alkalinity, acidity and sulfate loads for the two replicates. Overall, acidity and alkalinity loading in leachate from both column tests was higher than in their corresponding humidity cells. Conversely, sulfate loading was higher for the humidity cells.  The difference in alkalinity loading may be in part attributable to the high aeration conditions in humidity cell tests which rapidly remove carbon dioxide volatilized during acid neutralization. This favors the following reaction to proceed: 0 3 C 2 H °  —*  0 2 2 (g) + H CO  —>  3 H + HCO  rather than 0 3 C 2 H °  which would tend to dominate in a closed system and slow the rise of pH to levels above 6.0 (Ferguson and Morin, 1992).  Since leachate alkalinity is a measure of bicarbonate ion in  0)  HC7  10:  <  Figure 4.28  :  10  mean  maximum  Co14  minimum  Co14  Comparison of replicate column and humidity cell sulfate, alkalinity and acidity loads  0  V  200  . >.  0)  0 0  Co12  20  400  15  25  500  0  HC7  0  Co12  50  100  Ca 0 —‘  50  100  300  Ca o —I  150  ,  200  200  150  250  250  HC8  HC8  -—  —  108 solution, the resulting alkalinity values for humidity cell would be less than for a column test on similar material. The open, aerated system of the humidity cell also results in a more efficient neutralization of hydrogen ion by calcite (close to 2:1 molar ratio), whereas in a more closed system such as the columns, a 1:1 acid to carbonate molar ratio is dominant. This may also explain the comparatively lower acidity loads produced in site 2 humidity cell test; the humidity cell allows available neutralizing capacity to be used in close to a 2:1 acid to carbonate ratio, thus the resulting humidity cell leachate should of lower net acidity than a column test.  The higher sulfate loading in humidity cells of both replicates may indicate that either more aggressive oxidation occurred in the aerated humidity cell tests, or that significant soluble sulfate was still leaching from the material despite the pre-flushing.  Pore gas analysis of column 2 indicates that highly acid generating material may have oxygen depletion within the top 30 centimetres of the waste rock dump. This explains the occurrence of fresh, unoxidized material beneath a 15 to 30 centimetre layer of highly oxidized material on the dump surface (Fig. 3.3). Oxygen availability in such rapidly oxidizing waste rock is obviously limited both by diffusion and consumption by pyrite oxidation.  4.4.2  Kinetic Test Rate Constants  Renton et al. (1988) noted that plots of unreacted sulfide versus time for coal refuse kinetic tests resembled a first order decay curve, in which the rate of sulfide oxidation is assumed to be directly related to the amount of unreacted sulfide remaining. By applying the relatively simple first order equation to laboratory bench scale tests, they were able to successfully model sulfide oxidation of a 350 ton waste rock kinetic test. Steffen, Robertson and Kirsten (1992) also used the first order equation to predict not only sulfate, but also various metal loads.  109 The first order equation can be expressed as: (4.1)  1 2 = initial oxidation rate (mollkg/wk), and k where dS/dt = rate decrease of solid phase sulfide, k =  sulfide oxidation rate constant (wlcl).  1 is equal to: For a first order reaction, the rate constant k  /dt) -2.303(dlog[S 1 =2 1 k  (4.2)  (Russell, 1980)  This is simply 2.3 03 times the slope of the log of sulfide remaining versus time plot, and can easily be calculated for all the kinetic tests using only the weeks after pseudo-steady state was obtained (Table 4.6).  Problems arise in applying equation (4.1) to the kinetic tests because the material had weathered , and the (k ) at least eight years before sampling (Table 3.3). Therefore, the initial oxidation rate 2 time of commencement of reaction, or time of ARD onset, are unknown. Numerous attempts were made to apply the model to this study’s kinetic tests, but through sensitivity analysis it was found that the assumptions that had to made resulted in significant variation of results.  ) can 1 2 and the time of ARD onset are unknown, the sulfide oxidation rate constant (k Although k be calculated, and are presented in Table 4.6. For clarity, the half-life, or amount of time required for half of the material to oxidize (equal to [2.3031og(2)/kj]  ), is also given.  110 Table 4.6  Kinetic test rate constants and half-lives  Kinetic Test  Site #  Rate  Half-Life  Constant*  (years)  (wk- 1)  *  Col 1  1  0.00005  245.1  Co12  2  0.00158  8.4  Col 3  3  0.00009  154.3  Col 4  4  0.00027  50.3  HC 1  5  0.00013  100.8  HC 2  8  0.00057  23.6  HC 3  8  0.00076  17.5  HC4  7  0.00184  7.2  HC5  7  0.00131  10.2  HC 6  6  0.00168  7.9  HC 7  2  0.00065  20.6  HC8  4  0.00051  26.1  rate constants temperature-adjusted to 8°C  4.4.3  Kinetic Test Neutralization Potential Depletion  In order to classifSr samples as to their long term acid buffering ability, times to depletion of available neutralization potential (NP) were calculated for each kinetic test. The time to NP depletion uses weekly sulfate loads coupled with either weekly alkalinity or calcium and magnesium loads to determine both the rate of sulfide depletion and the rate of consumption of alkalinity. If the available NP in the sample is known either through acid base accounting, acid volatilized carbon dioxide, or elemental analysis, the time to NP depletion can be calculated. Although Lapakko (1990) showed that the time of alkalinity depletion could be predicted in  111 short term kinetic tests, in this study, the times estimated by these calculations are used strictly to classify samples as to their long term buffering capacity. Because the calculation uses a linear extrapolation of short term kinetic test results, the times yielded by the calculation, especially if greater than two years, are not considered to accurately indicate true depletion times.  As mentioned, NP depletion can be calculated in two ways: i) using molar sulfate to alkalinity /Alk) (Ferguson and Morin, 1992), or ii) using molar sulfate to calcium and 4 j (S0 3 (as HCO /(Ca+Mg)) (Lapakko, 1990). Both methods were used and compared on 4 magnesium ratios (S0 Island Copper kinetic tests.  /(Ca+Mg) ratios were comparable. 4 IA1k ratios and S0 4 It was found that for the columns, SO However for the humidity cells, measured leachate alkalinity was significantly lower than those /Alk ratios were obtained. 4 obtained from column tests, and as a result, extremely high S0 Reasons for this are related to the low alkalinity loading observed in humidity cell leachate and are discussed in section 4.4.1.  For this study, the method of Lappakko (1990) was used to calculate time to NP depletion. The assumptions used in the calculations are that: i) all acid neutralization will be done by carbonate minerals (calcite and/or dolomite); ii) all carbonate is available for neutralization; iii) all sulfide sulfur is available for oxidation; iv) all sulfate released to solution is from iron sulfide oxidation; v) all of this sulfate remains in solution; and vi) the rates of carbonate and sulfide depletion is constant.  To test assumption v), actual molar sulfate to calcium ratios were compared with theoretical values derived from the formula for gypsum equilibrium given in Ferguson and Morin (1992). All leachate samples from the data analysis periods specified in Table 4.4 had much higher  112 molar sulfate to calcium ratios than the theoretical level, indicating that they were unsaturated with respect to gypsum; thus, assumption v) is valid.  The equation used for NP depletion calculation is: 0 Cc  43  Cd[Ca2++Mg2+1  dt  where t,  =  0 time to carbonate depletion (weeks), Cc  =  amount carbonate in sample at test start  2]/dt = leachate calcium and magnesium load (mol/kg/week). 2 + Mg (mol/kg), and d[Ca  The total available carbonate in each sample expressed in mol/kg units was determined through acid-volatilized carbon dioxide analyses. In the case of the columns where three separate analyses were performed on each sample, the geometric mean was used.  NP depletion calculations were performed on data from each week of the humidity cell tests, and for the columns, weekly loads were summed into a four-week load.  The resulting ranges of estimated time to NP depletion for each kinetic test are given in Table 4.7.  4.4.4  Kinetic Test ARD Potential Classification  In examination of a few key water quality parameters from all kinetic tests, certain groupings became apparent. These, combined with time to neutralization depletion calculations, served to c1assifr the twelve kinetic tests into one of three categories: i) Type I rocks, which are interpreted to have sufficient excess alkalinity to do some degree of buffering of infiltrating acidic drainage,  113 ii) Type II rocks, which although possibly generating alkaline leachate at present, are not regarded to have sufficient excess alkalinity to adequately buffer infiltrating acidic drainage, and are also not expected to significantly contribute to net acidity of the waste rock dump; and iii) Type III rocks, which are presently producing leachate with high net acidity. The delineating parameters for each category are given in Table 4.8. The category assigned each kinetic test is shown in Table 4.9.  114  Table 4.7  Kinetic test estimated times to neutralization potential depletion  Predicted Time to  Kinetic Test  Neutralization  Potential Depletion* (years) Coil  38-44  Co12  <1  Co13  4-5  Co14  5-6  HC1  40-53  HC2  <1  HC3  <1  HC4  <1  HC5  <1  HC6  <1  HC7  <1  HC8  1-3  * assuming laboratory temperatures  Table 4.8  (21 to 23°C)  Delineating parameters for  ARD  Potential classification of kinetic tests  Time to Alkalinity  >25  years  Type III  Type II  Type I  <1  25  to  <1 year  years  Depletion pH Sulfate Load  >7.0 <30  mg/kg  4.3  >4.3 160  mg/kg  >  160  mg/kg  115 Table 4.9  Classification of kinetic tests by ARD Potential category  ARD Potential Category Type II  Type III  Site  Test  Type I  1  Coil  .7  2  Co12  .7  2  HC7  .7  3  Co13  .7  4  Co14  .7  4  HC8  .7  5  HC1  6  HC6  .7  7  HC4  .7  7  HC5  .7  8  HC2  .7  8  HC3  .7  .7  116 5.0  PREDICTION OF EFFLUENT QUALITY  5.1  Introduction  This chapter links results of laboratory scale waste rock characterization studies with measured characteristics of the North dump. At the conclusion of Chapter 4.0, the eight sites sampled were categorized into one of three ARD potential categories on the basis of their kinetic test performance: i) Type I rocks, which are interpreted to have sufficient excess alkalinity to do some degree of buffering on infiltrating acidic drainage, ii) Type II rocks, which although possibly generating alkaline leachate at present, are not regarded to have sufficient excess alkalinity to adequately buffer infiltrating acidic drainage, but are also not expected to significantly contribute to net acidity of the waste rock dump; and iii) Type III rocks, which are presently producing leachate with high net acidity.  In section 5.2, rock type, alteration, mineralogy, geochemical and water quality data from the eight sites are re-examined to determine if there are any delineating characteristics for each category.  In section 5.3, selected dumps and dump areas are quantified with respect to the amount of material in each ARD potential category.  In section 5.4, characteristics of present effluent quality in various dump areas are predicted using relationships between ARD potential categories and dump characteristics.  117 In section 5.5, a simple, temporal model is used to predict future effluent quality from EMO and North West dumps.  5.2  Characterization of ARD Potential Categories  5.2.1  Lithology  Estimated proportions of each lithology found in samples from the three ARD potential categories are shown in Figure 5.1. Data from Table 3.4 was used to arrive at these estimates.  Type II and III samples are composed of a variety of lithologies, however type I samples contain only Bonanza Volcanic rocks. Type II samples contain a significant amount of pyrophyllite breccia.  5.2.2  Alteration  Table 5.1 depicts the dominant alteration of the samples from each ARD potential category. Although type I and II samples have overlapping alteration characteristics, type III samples are all dominantly SCC (sericite-chlorite-clay-pyrite) altered. Type II samples contain dominantly pyrophyllite-altered material.  5.2.3  Pyrite  Visual estimates of pyrite for each ARD potential category are shown in Figure 5.2. There is significant overlap in the visual pyrite estimates in each category.  75.  Figure 5.1  0.  25.  50.  (U  E  w a.  100  0.  onanza V  P,ropliytlite Brecca Hydrothermal Breccia ,..hyodacite Porphyry  Vein  -—  Lithological characteristics of ARD potential categories  I ,rophyflite BreJia 1 P onanza VoIca Hydrothermal Breccia Rhyodacite Porphyry  Type III  Type II  _nanza Volcanic Rhyodacite Pocpt  rj  0 •  25  75  25.  (U Cl)  E  a.  .  50  75.  100  50.  a)  E  CI) 0.  100  Type I  I. Vein Breccia Hydrothermal Breccia  00  119 Table 5.1  Dominant alteration characteristics of ARD potential categories  ARD Potential Category III  II  I  Alteration sericite-  1 V’  chlorite-clay  ‘V  quartz-sericite  pyrophyllite  •1  ‘/‘  biotite-  ‘/‘  magnetite epidote  chlorite  12  10 8  I I  -I ]  I .1I  I -l  6 4 2 0  I  Type I  I  —  I  Type II  I  Type flI  . pyrite (°“) Figure 5.2  Visual pyrite estimates for each ARD potential category  ‘V  1  120  Carbonate Minerals  5.2.4  Tables 5.2 and 5.3 illustrate degree of effervescence from dilute hydrochloric acid for rock matrix and veins or fractures for each ARD potential category. While type I samples generally show greater effervescense to dilute acid, no one ARD potential category is distinguishable from the other based on effervescence alone. Table 5.2  Rock matrix reactivity to hydrochloric acid of ARD potential categories ARD Potential Category III  II  I  Reactivity Strong Moderate  4/  4/  4/  Weak  4/  Table 5.3  4/  4/  4/  Nil  4/  Vein or fracture reactivity to hydrochloric acid of ARD potential categories ARD Potential Category I  Reactivity Strong  II  III  4/  7  4/  Moderate Weak  Nil  5.2.5  4/  4/  4/  4/  4/  Acid Base Accounting  Type III samples are distinctly higher in total sulfur than the type I or II samples (Fig. 5.3a). Based on this relatively small dataset, it appears that samples with greater than 2 percent total sulfur tend to be in the strongly acid generating category.  -20  0  20  40  60  80  100  120  0  2  Type I  Type I  I  —  Type Ill  I  Type II  I Type lii  I.’.  “r  Total Sulfur (wt%)  Type II  •  • ACP (kg CaCO3It)  I  -1----.  [L  4  I  I  I  6  8  10  d)  b)  0  1  2  3  4  0  0.2  0.4  0.6  0.8  I  .  I-  I  Type II  I  F  Type Ill  Type Ill  I — Acid-volatilized C02 (wt%)  Type I  — Type II  Sulfate (wt% as Sulfur)  J  F •I  Type I  I I  I  Acid base accounting parameter characteristics of ARD potential categories, a) total sulfur, b) sulfate, c) acid consuming Figure 5.3 potential (ACP), d) acid-volatilized carbon dioxide  c)  a)  I ACP:APPI  Type II  Type III  2  Type I  -0.5  4 3  _  0  0.5  5 .----  7  1  --------------  8  6  ----a  d)  1.5  2  2.5  NNPS (kg CaCO3It)  Type I  Type I  I  I.PasteP’  Type II  I  —--H •1  I  NNPS2- (kg CaCO3It)  Type II  Type III  Type III  Acid base accounting parameter characteristics of ARD potential categories, a) net neutralizing potential using total sulfur Figure 5.4 j, c) acid consuming potential to acid production potential (ACP:APP), d) 2 (NNPS), b) net neutralizing potential using sulfide sulfur (NNPS paste pH (ppH)  c)  -300  -300 Type III  -250  -250  Type II  -200  -200  Type I  -150  -150  0 -100  —  -100  .  50  100  -50  0  50  100  b)  -50  a)  123 Type III samples are also significantly higher in sulfate than types I or II (Fig. 5.3b). Samples show distinct acid consuming potential (ACP) characteristics for each ARD potential category (Fig. 5.3c). Acid-volatilized carbon dioxide levels are distinct between type I and II samples, but there is overlap between types II and III (Fig. 5.3d).  Similar trends for net neutralization potential using total sulfur (NNPS) and sulfide sulfur j (Fig. 5.4a and 5.4b). 2 (NNPS  Samples in each ARD potential category also have distinct acid consuming potential to acid producing potential (ACP:APP) ratios (Fig. 5.4c).  Paste pH of the samples also shows a  decreasing trend from type Ito type III category rocks (Fig. 5.4d).  5.2.6  Elemental Analyses  Overall, only a few distinguishing elemental characteristics of the ARD potential categories were delineated. This is in part due to the small sample size.  Type II samples are distinct in that they are relatively low in manganese, iron, magnesium and aluminum. As mentioned in Section 3.3.3, this is reflective of the considerable amount of pyrophyllite-rich rocks in the type II samples.  Calcium is the only element with relatively distinct groupings for each ARD potential category which closely resemble that of acid-volatilized carbon dioxide (Fig.5.3d).  The two type I  samples are over 2.5 percent calcium, the three type II samples are between 0.5 and 1.5 percent calcium, and the three type III samples form a tight grouping between 0.5 and 0.75 percent.  Type I samples were relatively high in zinc and cadmium. This is likely due to the association of sphalerite with carbonate minerals.  124 5.2.7  Rock Competence  The sample’s rock competence, qualitatively determined by the number of hammer blows required to fracture a given specimen are shown in Table 5.4. Type I samples tend to be more competent than type II and III samples.  Table 5.4  Rock competence characteristics of ARD potential categories  ARD Potential Category I  Competence  Very Competent Competent  III  II  7 ../  Moderate  1 ..7  v_F  Soft  5.2.7  Discussion  Based on observations of lithology, alteration, pyrite content, reaction to dilute hydrochloric acid, and rock competence, tentative guidelines to field classification of Island Copper waste rock into ARD potential category are given in Table 5.5.  Table 5.5 also indicates the rank, or weighting that each parameter should be given when classifying a given rock. For example, if considering an SCC-altered rock with less than 1 percent estimated pyrite, it would be classified as type III because of the higher weighting given to alteration.  More sampling in the form of reconnaissance work on the dumps would further improve this classification.  125  Table 5.5  Tentative guidelines to field classification of Island Copper waste rock by ARD  potential category  ARD Potential Category Dharacteristic .ithology  Rank 4  Type II  Type I  Bonanza Volcanic All Lithologies, but predominantly  Type III All Lithologies  pyrophyllite )ominant Alteration  1  non-SCC non-pyrophyllite  non-SCC, usually pyrophyllite or quartz-sericite  SCC  Pyrite visual estimate) Reactivity to 10% Rydrochioric Acid  2  <1%  <2%  >1%  3  moderate to strong  nil to weak  nil to weak  Rock Competence  5  very competent  moderate  variable  The analysis of acid base accounting characteristics of the ARD potential categories has demonstrated that samples indicated by acid base accounting to be potential acid generators are not doing so despite up to twelve years of weathering on the dump surface. The criteria in Table 5.6 are based on results presented in Tables 5.3 and 5.4.  Worthwhile to note is that based on this work, an ACP:APP ratio of greater than 1.3 is considered indicative of type I, or acid consuming waste rock. This is in contrast to recent guidelines suggesting minimum ratios of 4:1 (Ministry of Energy, Mines and Petroleum Resources, 1993) for initial screening of potentially non-acid generating samples.  Further  sampling on the waste rock dumps and integration of existing data (for example BHP and Rescan, 1988) may further refine the ABA criteria given in Table 5.6.  126 Table 5.6  Tentative acid base accounting criteria for ARD Potential categories  ARD Potential Category Type II  Type I NNP(total sulfur)  Type III  NNP> 10  -25< NNP  10  NNP  -25  ACP> 50  15  ACP  50  ACP  15  t) CaCO I (kg 3 ACP  <  t) CaCO I (kg 3 ACP:APP  ACP:APP >1.3  0.3  <  ACP:APP  1.3  ACP:APP  0.3  (using total sulfur NNP)  5.3  Assigning ARD Potential Categories to Waste Rock Dumps  Based on the acid base accounting (ABA) criteria in Table 5.6 and using the drill hole ABA analyses, the amount of waste rock in each ARD potential category was calculated for the Old  North, EMO, and Cap regions of the North dump, as well as the North West and Beach dump. Each ABA analysis was designated to the appropriate ARD Potential category using criteria for NNP(total sulfur). Assuming equal weighting to each analysis, the percent of waste rock in each ARD potential category was calculated. For example, out of a total of 19 samples from EMO, 8 were categorized to be marginal. Therefore, 42 percent of waste rock in EMO is estimated to be in this category.  The results of the waste rock dump classification are shown for the North (Figure 5.5), North West (Figure 5.6), and Beach (Figure 5.6) dumps. It can be seen that the waste rock regions or  —Type 1(47.60%)  —Type 11(42.10%)  1(5.30%)  Type 11(50.44%)—’  Type III  L_Type 11(71.40%)  Type 1(40.27%)  ,.—.-Typel(14.30%)  Entire North Dump  Type III (14.30%)—\  CAP (n =7)  Estimated proportions ofrock in each ARD potential category in the North dump  (50.00%)___I  Figure 5.5  Type II  OND(n=42)  Type III (2.40%).1  Type III (52.60%)—i  EMO(n=19)  L’J  Figure 5.6  (42.60%)___1  [. Type 11(53.70%)  Type 1(53.00%)  Estimated proportions of rock in each ARD potential category in the North West and Beach dumps  Type III  (n 66)  BCH(n=54)  NWD  00  129 dumps known to be net acid generating are estimated to contain from 14 to 56 percent of type III waste rock. In contrast, the Old North dump is estimated to contain just over 2 percent (1 sample) of strongly acid generating material. There is also a problem of adequate sampling in some areas, particularly from the Cap region (7 samples). The entire North dump is estimated to contain about 9 percent of type III material.  5.4  Prediction of Present Effluent Quality from Dump Areas  5.4.1  Method  A simplified flow chart of the prediction process is shown in Figure 5.7.  The kinetic test leachate loads are first examined according to their assigned ARD potential category (box 10 on Figure 5.7), and simple statistics were performed to arrive at mean values of alkalinity, acidity, molar calcium and magnesium, and sulfate loads (box 9). In addition, the “net acid load” for each category is also calculated. The net acid load is equal to the difference between molar sulfate and calcium and magnesium loads, and is indicative of the amount of acid produced that is not neutralized by available carbonate minerals.  The assumptions in the  calculation are that: i) sulfate production is representative of acid generation, ii) calcium and magnesium in effluent is most reflective of carbonate mineral dissolution, and iii) carbonate minerals play the major role in acid neutralization.  For each dump or region of dump quantified with respect to ARD potential, typical acidity or alkalinity, sulfate, molar calcium and magnesium, and net acid load values are assigned based on the percent of material in each ARD potential category (box 11 on Figure 5.7).  130  Waste Rock Dumps  ERock Samples  I  2  1 Kinetic test leachate quality  3  (Ch.4.O)  acid base accounting  5  (Table 3.6)  Kinetic test leachate quality  9  (Ch.4.O)  Derive ARD Potential Categories 4  (Table 4.8)  ABA Criteria for ARD ARD Potential Categories 6  (Table 5.6)  Amt. of Each ARD Potential Cat. in Waste Rock Dumps 7  (Fig.5.5,5.6)  Estimated Effluent Quality for EMO and NW dumps ii  (Table 5.9)  acid base accounting  (Li,1 991)  8  Actual Effluent Quality from EMO and NW dumps 12  (Table 5.9)  Compare Estimated and Actual Effluent Quality 13  flow of prediction process source of data  Estimate Effluel Quality for OND Cap, and Beach dump areas 14  Figure 5.7  (Table 5.9)  (Table 5.11)  Flowchart for prediction of present effluent quality from dump areas  131  Actual effluent quality loads from the EMO and NWD dumps are calculated so that comparisons between kinetic test and field-derived data can be made (box 12 on Figure 5.7)  Estimated results are then compared with field water quality measurements.  Factors  contributing to discrepancies and agreement in the results are then discussed.  5.4.2  Leachate Quality of ARD Potential Categories  For the purposes of effluent quality prediction, only selected leachate quality parameters will be discussed in detail. Gaussian statistics for these selected parameters by ARD Potential Category are given in Table 5.7.  As shown, types I and II leachate have negative net acid loads (or net  alkalinity). Although the net acid load values for the two categories are similar in magnitude, alkalinity depletion calculations (section 4.4.4) demonstrates the lower ultimate buffering capacity of the category II material.  Sulfide oxidation rate constants for the kinetic tests are tabulated by ARD potential category in Table 5.8. Because only one rate constant was calculated per kinetic test, standard deviations are not indicated for category I due to its small sample size.  5.4.3  Actual Effluent Quality of EMO and NWD Dumps  The EMO and NWD dumps were selected to compare loads predicted from kinetic tests with field conditions because: i) water quality data is available from ditches exclusively draining each of these dump areas, and ii) there is sufficient acid base accounting information on each of these dump areas.  -  (mg CaCO 3 equivJkg/wk) -8.2 -8.7 404.0  73.6 452  46.3 574  436.0 2458  3.0  155  II  III  Mean 10.5  (=  Net Acid Load mean molar (sulfate (Ca+Mg)])  0.005440  0.000347  329  329  386  1737  Standard Deviation 8.1  -  0.002870  0.000054  9.0  3.4  0.000385  0.000080  53  Max.  Mm.  0  Standard Deviation -  32.0  Max.  -  Mean -  -  Max.  0.001941  0.000567  0.000192  Mean  Molar (Ca + Mg) Load (mollkg/wk)  -  Mm.  4.0  Mm.  Sulfate Load (mg/kglwk)  -  5.0  Standard Deviation 4.0  Acidity Load /kg/wk) 3 (mg CaCO  I  Category  ARD Potential  -  5.0  23.0  0  II  III  9.8  Mean  19.0  Max.  4.0  Mi  Alkalinity Load (mg CaCO /kg/wk) 3  Characteristics of kinetic test leachate chemistry by ARD Potential category  I  Category  ARD Potential  Table 5.7  0.001506  0.000708  Standard Deviation 0.0001 10  133 Table 5.8 Category  Characteristics of kinetic test sulfide oxidation rate constants by ARD Potential  ) 1 Sulfide Oxidation Rate Constant (k (wk-’)  ARD  Potential Category  I  Mean  Standard Deviation  Mm.  Max.  0.00005  0.00013  0.00093  0.00009  0.00076  0.000438  0.000264  0.00065  0.00184  0.001411  0.000469  (n=2)  II (n=5) III  (n=5)  To best compare kinetic test and field-derived data, effluent quality for both dumps was calculated using data for one year from September 1992 to August 1993, as during this period the majority of kinetic testing was conducted.  5.4.3.1 Calculation of Actual Loads The method of calculation of actual dump loads used total precipitation minus losses to evaporation and groundwater to estimate the volume of water reporting to the drainage ditches. An evaporation loss of 15 percent of total precipitation was used. Li (1991) calculated a 21.7 percent evaporation loss based on pit dewatering data; it was believed, however, that this estimate was too high. Assuming this evaporation loss, Li also calculated losses to groundwater in two areas of the North dump to be 7 and 17 percent respectively. A loss to groundwater of 10 percent of total precipitation has been adopted for this study.  The total annual precipitation was broken down into monthly values, and based on horizontal dump area and accounting for losses to evaporation and groundwater infiltration, monthly volumes of water reporting to the drainage ditch were calculated,  These volumes were  multiplied by corresponding average monthly sulfate, acidity, calcium and magnesium concentrations to get the monthly mass of each parameter reporting to the drainage ditch. The  134  amount calculated to be lost to groundwater was then added to these values.  The twelve  monthly loads were then summed to yield the annual mass of each parameter released from the dump.  To best compare kinetic test and field data, the “active” dump mass was considered to be the portion of material in the dump of similar particle size distribution to the kinetic test material. As discussed in Section 3.3.5, about 50 weight percent of the dump is estimated to be greater than the size fraction used in the tests, therefore, the active dump mass is assumed to be 50 percent of the total dump mass.  Calculation of load per unit active mass of the dump was then obtained by dividing the annual mass of sulfate, acidity, calcium and magnesium released by each dump by its active dump mass.  The loads for EMO and NWD are presented in Table 5.9.  5.4.3.2 Calculation of Sulfide Oxidation Rate Constants Sulfide oxidation rate constants from September 1992 to August 1993 for EMO and NWD were calculated using: i) the annual total sulfur oxidized and released as sulfate (derived according to the method given in section 5.4.3.1), and ii) the total available sulfur in. each dump as calculated from drill hole acid base accounting values.  Total sulfur available in each dump was calculated by using the average APP value expressed as kilograms sulfur per tonne, and multiplying this by the dump’s active mass (one half of its total  135 tonnage). The estimated percent sulfur released from the dump during the one year period was calculated using:  %S Released  [STR(92.93)] / [ST] x 100  (5.1)  where STR(92..93) is the total amount of sulfur released as sulfate from the dump from September 1992 to October 1993 (mol/kg), and ST is the estimated total sulfur contained in the waste rock dump.  Using the percent sulfur released from 1992 to 1993 and knowing the year of the collection of the acid base accounting samples from the dump, the approximate sulfur content of the dump in September 1992 was calculated. The sulfur content of the dump in August 1993 was then determined by subtracting the annual percent sulfur released determined from equation 5.1. The )was calculated using: 1 1 (in year annual sulfide oxidation rate constant, k  1 k  =  [log(ST(92)) log (ST(93))] / [1 year] -  (5.2)  Rate constants for EMO and NWD are given in table 5.9.  5.4.4  Comparison of Estimated and Actual Effluent Quality from EMO and NWD  Using the mean leachate quality parameters given in Tables 5.7 and 5.8, coupled with the proportion of each ARD potential category in the EMO and NWD dumps, an estimate of dump effluent quality was calculated for each parameter.  A comparison of estimated and actual effluent quality for the selected parameters is given in table 5.9.  Overall, loads predicted by kinetic testing and ARD potential category were  significantly higher than actual field conditions. Net acid load gave the largest differences  136 between actual and estimated values, while the smallest differences were obtained for the dimensionless molar calcium and magnesium to sulfate and net acid load to sulfate ratios.  Table 5.9  Comparison of calculated versus actual effluent quality of EMO and NWD  drainages  0.0000637  Scale Factor 0.125  % Difference in Scale Factors 70  0.000723  0.000049  0.067  70  0.0086  98.0  0.8  0.0082  5  4.2  0.013  163.6  5.5  0.033  60  0.38  0.59  1.67  0.42  0.86  2.00  15  0.62  0.41  0.67  0.58  0.14  0.24  60  NWD  EMO  Rate Constant ) 1 (wk Mo! (Ca+Mg) (mo!Ikglwk) Net Acid Load 3 (mg CaCO equivikglwk) Sulfate Load (mg/kglwk) molar [(Ca+Mg):Sulfate] molar [Net Acid:Sulfatel  Estimated  Actual  0.0000368  Scale Factor 0.038  0.0005060  0.001271  0.000026  0.020  208.4  1.8  322.1  Estimated  Actual  0.00093 15  The discrepancies between actual and estimated effluent quality may be attributable to the following factors: 1) the kinetic tests significantly accelerating weathering reactions, ii) a significant error made in the estimate of actual active leaching mass of each dump, and iii) precipitation of gypsum and other sulfate minerals within dump, thus reducing effluent sulfate values.  The acceleration of weathering reactions within kinetic tests may have occurred because of the comparatively higher temperatures and the aggressive flushing regime used in the experiments. However, as discussed in Section 4.1, kinetic test acceleration of field behavior is poorly  137 understood. Nonetheless, the estimates of the existing acceleration of field weathering are less than an order of magnitude, thus not entirely accounting for the discrepancies observed.  The actual active leaching mass of each of the dumps is probably the least known factor in the prediction calculation. As discussed in Section 2.3.4, flow of water through the Island Copper and other mine dumps is at present poorly understood. However, even if the 50 percent active leaching mass assumed for the Island Copper waste rock dumps has overestimated the actual value by 100 or 200 percent, the large discrepancy in loads is still unexplained.  Although occasional gypsum crystals were observed on surfaces and fractures of strongly acid generating samples, no other sulfate minerals were noted on rock surface coating analysed by x ray diffraction (Section 3.3.4). Li (1991) concluded from analysis of EMO drainage chemistry that conditions indicative of gypsum precipitation only occurred occasionally in the data. Although the precipitation of gypsum and other sulfate minerals may be occurring in localized areas of the dumps, it cannot by itself account for the large discrepancy between estimated and actual sulfate loading.  Overall, the significant discrepancies between estimated and actual effluent loads of the EMO and NWD drainages remain largely unexplained. This is likely due to our limited knowledge of dump hydrology, the chemistry of rock-water interactions, and the design of laboratory scale simulations. A better understanding of speciation and precipitation reactions within the dumps would no doubt make correlations between laboratory and field-derived data better. The study of the on-land dump effluent chemistry being conducted by Morin (1994) may provide some insight into this phenomonen.  Despite these discrepancies, if the ratio between actual and estimated conditions is known, the estimates can be scaled with reasonable confidence. By comparing actual and estimated scale factors for the two dump areas, an indication of the reliability of the factor can be obtained. The  138 far right column of Table 5.9 shows the percent difference between the scale factors obtained for the EMO and NWD drainages. It can be seen that reasonable agreement (within 5 and 15 %, respectively) was obtained from both the net acid load and molar calcium and magnesium to sulfate ratio scale factors. Adopted scale factors for these two parameters were obtained by taking the mean of the values from EMO and NWD, and are shown in Table 5.10. Reliable scaling of the other parameters shown in Table 5.9 is not considered feasible.  Table 5.10 values  Adopted scaling factors for calculating actual effluent conditions from estimated  Parameter Net Acid Load 3 equiv./kg/wk) (mg CaCO molar [(Ca + Mg):Sulfate]  Adopted Scaling Factor Estimated —* Actual 0.0084 1.8  Using these scale factors, estimated current effluent conditions for the OND, Cap, and Beach dumps are given in table 5.11. Actual effluent conditions for EMO and NWD are also given. Estimated current effluent conditions in selected dumps and dump areas at Island Table 5.11 Copper Mine Dump or Dump Area  *  Net Acid Load 3 equiv.Ikglwk) (mg CaCO  <0.1 OND EMO* 1.8 0.4 Cap NWD* 0.8 1.4 Beach actual values based on measured effluent quality  molar [(Ca + Mg):Sulfate] 5.0 0.6 1.0 0.9 0.7  139 5.4.5  Discussion of Results  The net acid consuming potential of the OND dump area is reflected in its estimated low net acid load and high molar calcium and magnesium to sulfate ratio (Table 5.11).  There is inadequate waste rock data on the Cap (only 7 samples from one hole), and the calculated amount of material in the strongly acid generating category (type III) is considered to underestimate actual amounts. This is reflected in the low net acid load and the high molar calcium and magnesium to sulfate ratios estimated for Cap effluent. However, this result is belied by field evidence indicating significant acid generation in the Cap, including: 1) the presence of low pH effluent from several seepages from the toe of the Cap, ii) the numerous surface patches of highly oxidized material on the dump, and iii) elevated down-hole temperatures greater than those obtained in the EMO dump holes. Based on this evidence, the degree of acid generation presently occurring in the Cap is considered to be equal in magnitude to the EMO or Beach dump areas, despite the above estimates to the contrary.  EMO and Beach dump areas have similar predicted effluent quality characteristics. However, the Beach dump, being partially submerged in tidewater, has oxygen availability and temperature distributions different from the on land dumps. Since the predictions are based on on-land dump scale factors, the accuracy of the Beach dump predictions are the least certain of all the dump areas. Further discussion of the Beach dump predictions is given in Appendix 11.  5.5  Temporal Modeling of EMO and NWD Dumps  5.5.1  Introduction  Given that ARD is already occurring within some regions of the Island Copper on-land waste rock dumps, the purpose of temporal modeling of the dumps is to help answer the following questions:  140 i) will the ARD problem get better or worse? ii) what concentrations of contaminants are expected from the dumps in the future?  Some previous studies have also focused on the question of time to sulfide depletion (Morwijk, 1993). However, in terms of bond-setting and design, the question of exactly how long to sulfide depletion becomes purely academic if our current unrefined estimates give times in the order of several hundred years or longer, for example the 610 years calculated by Li (1991) for the EMO dump.  Indeed, the actual existence of a sulfide-depleted waste rock dump is  debatable; to the writer’s knowledge, there are no waste rock piles in existence whose sulfide reserves are depleted and are no longer producing acidic leachate. This includes the several hundred year old ARD sites in Sweden.  As discussed in Section 2.2.7, field observations on waste rock dumps at Island Copper Mine indicate that oxygen unavailability, though occurring in localized zones in the dumps, is likely not controlling the rate of acid generation. However, the documented temperature increases within the dump do have the potential to significantly increase sulfide oxidation, and hence acid generation rates (Marries and Ritchie, 1981).  The temporal model assumes that acid generation kinetics can be approximated using a first order reaction. This method, as discussed in Section 4.4.2, has been used by Renton et. al., (1988) and Steffen Robertson and Kirsten (SRK)(1993) for bench and field scale kinetic tests. In modeling copper dump leaching, Cathies and Apps (1975) also assumes that pyrite oxidation by ferric ion is a first order reaction. In addition, similar decay curve equations have been used in other studies (Norecol, 1988, and the Dominique-Janine extension mentioned in SRK, 1993).  The model allows for increased oxidation rate due to temperature rises by using the Arrhenius equation (Russell, 1980) to recalculate rate constants for each time step.  141  5.5.2  Limitations  In considering future effluent trends, predictions made in this study are based on current rate constants from EMO and NWD and are thus limited to estimating effluent quality assuming that site conditions remain the same in the future. The addition of till covers and other progressive reclamation activities will reduce both water infiltration and oxygen availability within the dump, and the effluent quality changes as a result of these activities is beyond the scope of this study.  5.5.3  Method  As mentioned in Section 4.4.2, the first order reaction equation is:  (5.3)  =  where dS/dt  and kj  =  2 rate decrease of solid phase sulfide (mol/yr), k  =  initial oxidation rate (mol/yr),  sulfide oxidation rate constant (yr I).  2 are required for this model. Only two input parameters, kj and k  Sulfide oxidation rate  constants for EMO and NWD used in the model are given in Table 5.9. By integrating equation  5.3, sulfide remaining at time t can be expressed as: (54) 1 Ic  0 is therefore given by: At time t = 0, the original sulfide content S !2_ S = 0  and hence:  (5.5),  142 (5.6)  S 0 = 2 1 k  Since larger rock particles are expected to break down in size during the modeled time period, the entire dump is considered to be the “active” mass. Although rate constants given in Table 5.9 are calculated using only 50 percent of the dump as the active mass, check calculations indicate that insignificant error is introduced by using these values for the temporal modeling.  Total sulfur content at dump construction (1=0) was back calculated from time of waste rock dump drilling (collection time of acid base accounting samples) using the annual estimate of annual sulfur released from dump from September 1992 to October 1993.  Total available alkalinity in each dump area is assumed to be only the estimated carbonate mineral proportion of average acid consuming potential (ACP).  From Section 3.3.6, the  relationship at Island Copper between acid-volatilized carbon dioxide and ACP is:  3 It)— 0.055 mol I kg) = 0.OO66ACP(kgCaCO C0 ( 2  (5.7)  Table 5.12 gives calculations of available alkalinity in the EMO and NWD dump areas.  The model assumes that at dump construction, average temperature of the dump is equal to the normal average annual air temperature at the minesite (8°C).  Since future temperature gradients are uncertain, three trials using different temperature scenarios are given. The first two scenarios use calculated gradients based on actual temperature measurements of EMO and NWD drill holes. For extrapolation into the future, the North West dump calculated gradient of +0.37°C per year from 1989 to 1994 is adopted (Section 2.3.9). The two scenarios, test both a constant and decaying temperature gradient. The third scenario  143 tests the effect of constant dump core temperature as measured in 1993 (for EMO) and 1994 (for NWD). The three scenarios are: i) scenario 1, assuming that the temperature gradient will remain constant at +0.37°C per year to a maximum average dump temperature of 50°C (above which bacterial activity is substantially reduced and sulfide oxidation rates decrease (Harries and Ritchie, 1981)), ii) scenario 2, assuming that the initial temperature gradient of +0.37°C per year will decrease at a rate of 0.5 percent per year to a minimum average dump temperature of 15°C, and iii) scenario 3, that temperatures in the waste rock dumps will remain at levels of 17.0°C for EMO and 20.2°C for North West dump.  Table 5.12  Available alkalinity in EMO and NWD dump areas  Estimated  Portion of ACP  Available  Mass of  Available  as Available  Average ACP  Alkalinity  Dump*  Alkalinity  Alkalinity  (kg CaCO3It)  /kg) 2 (mol C0  (tonnes)  (mol)  (%)  EMO  18.67  0.0675  4.670(106)  3.152 (10)  36  NWD  49.16  0.2675  9.S43(10)  2.552 (108)  54  Dump Area  *  Estimated  from Li (1991) and UBC MMPE (1990a)  The change in sulfide oxidation rate constant is related to the change in temperature by the Arrhenius equation (Russell, 1980):  1ogk  =  logka  +  Ea (.i__!_) 2.303R 1 7  (5.8)  144 where ka and kb are rate constants at temperatures Ta and Tb (temperatures in K), Ea is the activation energy for pyrite (taken as 14,000 cal/mo! or 58,576 joules/mol (Cathies, 1975)), and ol’). JK m R is the ideal gas constant (8.3 14 1  Table 5.13 gives the input parameters for modeling of the EMO and NWD dump effluent.  Table 5.13  Input parameters for temporal modeling of EMO and NWD dump effluent  Dump  1 k  Avg. Dump Temp.  1 k  Area  Sept/92 to Aug/93  Sept/92 to Aug/93  (8°C, 1=0)  ) 1 (yr  (°C)  ) 1 (yr  EMO  0.001915  17.0  0.000880  NWD  0.003313  19.8  0.001208  Estimated Total Dump  Year of Dump  Year of ABA  Estimated Total  0 at Dump Sulfur S  Area  Construction  Sampling  Sulfur at Sampling  Construction  (1=0)  (moles)  (moles)  (mol/yr)  (t=0) EMO  1981  1988  2.432(10)  2.458 (109)  2,163,000  NWD  1983  1989  3.719(108)  3.775(108)  456,100  For each one year time step, the model outputs the rate of sulfide oxidation dS/dt. Assuming an average annual total precipitation of 2000 millimetres and that 75 percent of total precipitation reports each dump’s drainage ditch, and knowing the horizontal surface area of the dump, an average annual sulfate concentration in the ditch drainage is calculated. The sulfide remaining in the dump at each time step (S(i9) is determined by subtracting the sulfide consumed during the one year time step from the previous S(t).  145 1 and sulfide oxidation rate k 2 are For each time step, a new sulfide oxidation rate constant k  calculated. The new sulfide oxidation rate constant is determined by using the new temperature for the time step and applying the Arrhenius equation (5.8), and the new sulfide oxidation rate is . 0 determined from equation 5.5, substituting S(t) for S  Sulfate concentration is determined  using:  504 (mg/i)  =  [](96ooo) / [(0. 75)(2.0)(A)(1 000)]  (5.9)  where 96,000 is the moles to milligrams sulfate conversion factor, 0.75 is the proportion of total precipitation reporting to the drainage ditch, 2.0 is the estimated total annual precipitation in the mine area, and 1,000 is the cubic metres to litres conversion factor. A is the horizontal area of the dump in square metres.  The modeling of both alkalinity consumption and acidity concentration is linked to modeled sulfide oxidation by the molar calcium and magnesium to sulfate ratios given in Table 5.9. It is assumed that this ratio will be constant over the modeled time period and that alkalinity consumption will be equal to sulfide consumption times this ratio. The amount of unbuffered acidity, that is net acidity in each dump’s effluent, is estimated using an equation similar to 5.9:  —  3 / 1) Acidity (mgCaCO  dAlk  ](i 00000)/ [(0. 75)(2. 0)(A)(1 000)]  (5.10)  =  where dAik/dt is the rate of alkalinity consumption, and 100,000 is the moles to milligrams calcium carbonate conversion factor.  Renton et. a! (1988) noted that in chemical kinetics, reactions exhibit first order behavior only until approximately 63 percent (1 oo[i eq]) of the reactant is consumed. This was validated for —  ARD applications in small scale field experiments. Therefore, this model is considered to be  valid for amounts of total sulfide consumed between 0 and 63 percent.  146  The model was formulated and tested using a spreadsheet program, where graphical results could be easily generated.  5.5.4  Results  Figures 5.8 to 5.10 and 5.12 to 5.14 graphically show EMO and North West dump temporal modeling results for the three temperature scenarios.  As mentioned in Section 5.5.3, the first order model is considered valid only until 63 percent of the original sulfide is consumed.  Because of their different rate constants and the two  temperature scenarios tested, each trial of the model has a unique prediction period, and these  are given in Table 5.14.  Table 5.14  Valid prediction periods for first order model trials  Years of Valid Prediction Temperature  Temperature  Temperature  Scenario 1  Scenario 2  Scenario 3  EMO  108  410  551  NWD  99  298  387  Dump Area  As expected, temperature scenario 1 shows the higher ultimate sulfate and acidity concentrations for both EMO and NWD. Given this temperature regime, peak concentrations are expected to occur between the years 2060 and 2080 and be in the range of over 10,000 mg/l sulfate and over I CaCO / 1 acidity for EMO, and just under 5,000 mg/I sulfate and 750 mg 3 CaCO / 4000 mg 3 acidity for NWD. Peak values for all trials correlate well with peak temperatures in the dump.  148  100  2000  90 80 1500 70 0  ‘  •0 C,  60  —  < 50  1000  40  E  30 20 50:  10 n  I  2000  1900  2100  2200  2300  2400  2500  2600  Year  L— Sulfate (mgIl)  —  Temperature  —  Acidity (mg CaCO3II)  EMO predicted sulfate and acidity concentrations, temperature scenario 3  Figure 5.10  100  5000  90 80  4000  70 60  3000  50  o  I  2000  100:  1975  1980  1990  1985  1995  2000  Year Sulfate (mg/I) Acidity (mg CaCO3/I) -A- Actual Avg. Annual Acidity  —  —  --  Temperature Actual Avg. Annual S04  Comparison of EMO average annual sulfate and acidity concentrations with Figure 5.11 predicted values using temperature scenario 2  149  5000  100 90  4000  80 70 C.) 60  3000  z  50  41)  ! 2000  0.  40  0  E  30 20  1000  10 A  0 1980  2020  2000  2080  2060  2040  Year SuIfate (mg/I)  —  —Acidity (mg CaCO3/I)I  Temperature  NWD predicted sulfate and acidity concentrations, temperature scenario 1  Figure 5.12  100  2500  90 .80  2000  70 C.)  2 1500  60  I-  o 40  1000  ci)  Cl)  30  .20  500  0 1950  —  Figure 5.13  1’  :0  “  I  2000  Sulfate (mg/I)  2050  2150  2100  I 2200  I  I 2250  2300  Year  —  Temperature  —  Acidity (mg CaCO3/l)  NWD predicted sulfate and acidity concentrations, temperature scenario 2  150  1200 90 1000  80 70  >.  800  C-) 60  C., 0 1)  50  600  40 C’, 400  E  30 20  200 10 0 1900  2100  2000  [ Sulfate (mg/I)  —  Year  2300  2200  Temperature  —  2400  Acidity (mg CaCO3/I)  NWD predicted sulfate and acidity concentrations, temperature scenario 3  Figure 5.14  IflI)  5000  90 80  4000.  70 3000.  60  o  50  1)  1975  — — --  1980  1990  1985  1995  2000  Year  Sulfate (mg/I) Acidity (mg CaCO3/I) Actual Avg. Annual Acidity  —  Temperature Actual Avg. Annual S04  Comparison of NWD average annual sulfate and acidity concentrations with Figure 5.15 predicted values using temperature scenario 2  151 If the more conservative scenario 2 temperature gradient is adopted, the model predicts that peak concentrations will occur sooner, but will be not as high in magnitude as the scenario 1 results. The peak values are predicted to occur during the years 2040 to 2060. At this time sulfate and acidity values in EMO drainage are predicted to be over 4000 mg/l and just under 2000 mg i, respectively. NWD drainage sulfate and acidity concentrations are predicted to be CaCO / 3 l acidity, respectively. CaCO / over 2000 mg/l and over 250 mg 3  If, as in scenario 3, dump core temperatures do not increase in the future, sulfate and acidity concentrations will gradually decline, with peak concentrations occurring in 1993 and 1994 of 1 acidity for EMO drainage and CaCO / just under 2000 mg/i sulfate and approximately 750 mg 3 l acidity for NWD drainage. CaCO / under 1200 mg/i sulfate and less than 200 mg 3  5.5.5  Validation and Calibration  Validation of acid rock drainage weathering models is difficult because of the general lack of long-term field data (Steffen, Robertson and Kirsten, 1989).  Some validation, and calibration of this model was derived using historical effluent quality data. Annual averages of both sulfate and acidity concentrations of EMO and NWD sampling stations were used.  Figures 5.11 and 5.15 show actual versus predicted data for EMO and NWD. In both dumps, the model has underestimated sulfate concentration and overestimated acidity.  Reasons  speculated for this discrepancy are: i) an inappropriate value for pyrite activation energy (Ea, 58,576 J/mol used), ii) an underestimation of dump temperature at time of construction, and iii) the relatively low sulfate loads calculated from the 1992 to 1993 data compared with previous data.  152  Based on literature surveyed, pyrite activation energies can range from 10.7 to 25.6 kcallmol (44,769 to 107,100 J/mol, respectively) (Halbert et. al, 1983). These extremes of activation energies were used instead of the 58,586 J/mol originally assumed in the initial trials. The high activation energy trial gave a much poorer fit to the actual data than the original trial, but the low activation energy trial did not show any improved fit to the actual data. It appears that i) the actual activation energy of pyrite in the dump is in the low range, and ii) no significant improved fit to actual data is realized with changes in activation energy.  The validity of the initial dump core temperature assumption of 8°C was tested by assuming initial temperatures of 9, 10 and 12°C. The initial temperature gradient was recalculated to account for the new initial temperature. A noticeably better fit of actual to modeled sulfate concentrations was obtained, however acidity concentrations showed more deviation.  A third possible explanation for the difference is that the model is based on calculated sulfide oxidation conditions from September 1992 to August 1993, and sulfate concentrations for EMO in this time period were only slightly higher than the year before. Similarly, lower sulfate concentrations than the previous year were calculated for NWD.  This resulted in a lower  calculated rate constant for both dumps than would have been obtained from earlier data. It is too soon to tell if the one year’s data is indicative of an improving trend for the Island Copper waste rock dump effluent quality, but based on both examples from other dumps and the significant sulfur reserves still within the Island Copper piles, it is unlikely that sulfide oxidation will begin to abate at this early date.  5.5.6  Discussion  This temporal model, although more sophisticated than a linear estimation, is very simple compared with some other models which attempt to predict oxygen transport and heat production processes, for example, Cathles and Apps (1975). Several factors should be kept in mind when considering this model:  153 i) the model is based on the premise that there is an existing waste rock dump draining measurable and net acidic effluent, and ii) oxygen availability and intra-dump water transport characteristics, although not quantified, are “pseudo steady state” for the particular dump and the corresponding sulfate and acidity  loads have been measured. As long as neither oxygen availability nor intra-dump water transport conditions change significantly, they do not need to be known or estimated.  Rather than attempting to predict conditions of a new dump, this study addresses the simpler problem of predicting future behavior of already net acid-generating dumps, based on present and historical conditions.  Thus, a simpler model may be all that is necessary.  To quote  Nicholson (1992): “it is better to pose our uncertainties in the context of a simple model than a complex one.”  The temporal sulfide oxidation model illustrates the dependence of sulfide oxidation rate on average dump core temperature. This relationship has been verified by others in the field (Cathies and Apps, 1975, Harries and Ritchie, 1981). Unfortunately, present temperature data are considered insufficient to confidently predict future trends, and at best, a range of effluent  quality can be currently predicted.  The worst case is considered to be temperature scenario 1 which ultimately plateaus at 50°C average dump temperature. Such high temperatures have been measured at Equity and Mine Doyon operations (Morin, 1991, and Université Lava!, 1991), however, a relatively short period of time was required to achieve these temperatures. The low temporal temperature gradient measured in the North West dump may indicate that Island Copper dumps will never attain the heat production and correspondingly high sulfide oxidation rates that other dumps have.  However, based on ARD potential categorization, the North West dump has significantly less strongly acid generating material (type III) than EMO or Beach dumps (Figure 5.5 and 5.6). The Cap, with the highest down-hole temperatures of all the on land dump drill holes, is also  154 considered to have a large amount of strongly acid-generating material.  Heat production  behavior of the North West dump, therefore, may not be directly applicable to other net acid generating dumps at the minesite.  As discussed, the model assumes a constant ratio of sulfate to acidity, meaning that the dump will continue to buffer acidity at the same rate as measured in 1992 to 1993. In reality, the ability of the dump to buffer acidity will likely decrease over time. This implies that actual acidity levels will be higher than predicted.  When temperatures stabilize, concentrations of sulfate and acidity will gradually decrease. This is supported by the decrease in leachate sulfate and acidity concentrations during the 40 week colunm tests conducted at constant temperature.  However, temperature monitoring results  indicate that a low, positive temporal temperature gradient is plausible in the North West dump, EMO and Cap areas, therefore there is potential for increases in effluent sulfate and acidity concentrations. The magnitude of the increase can not be determined at this time due to lack of dump temperature data, and some unanswered questions on model accuracy.  155  6.0  CONCLUSIONS AND RECOMMENDATIONS  6.1  Conclusions  The heterogeneity of the waste rock dump material is reflected by the high variances in the drill hole acid base accounting analyses.  The results do confirm the significant  amount of potentially buffering till which comprises about one third of the material underlying most of the North dump. Insufficient sampling is evident in some areas, particularly the Cap, with only one drill hole and seven acid base accounting analyses.  The combination of push and free dumping methods has contributed to dump  heterogeneity, and has also made prediction of water flow through the dump difficult.  Oxygen levels in 1989 in the North West dump indicate that nearly all locations in the pile had sufficient oxygen at that time for sulfide oxidation. Pore gas oxygen levels in one of the column tests indicates that low oxygen levels in some areas may be limiting ARD generation in highly oxidized zones and within 15 to 30 centimetres of the surface. However, on a dump scale, very few regions are thought to have significantly depleted oxygen levels. Convective flow of air through the dump is evidenced by perceptible warm, moist air flow out of drill holes in the Cap and EMO dump areas.  Elevated temperatures in the EMO, Cap and North West dump areas are indicative of significant pyrite oxidation. The maximum measured temperature, from the Cap drill hole, is 24.3°C. Available data from the North West dump indicates that maximum temperatures in that area are stabilizing in the low 20°C level.  Due to the lack of  historical data for the Cap and EMO, conclusions on temperature stability cannot be made at this time. Compared with other monitored ARD sites (for example Equity and Mine  156  Doyon), the Island Copper on land dumps are relatively “cool”; this is reflected in acidity  concentrations from the EMO and NWD drainages being lower than typical ARD by more than an order of magnitude (see Table 1.1).  Overall, kinetic test work of the eight waste rock samples did not indicate worsening of water quality with time. Samples which produced net acidic leachate at test initiation continued to do so throughout the course of the experiment, and samples initially producing net alkaline leachate showed no indication of increasing sulfide oxidation with time. Based on the kinetic test work, rock at Island Copper has been classified into three categories: i) type I rocks, which are interpreted to have sufficient excess alkalinity to do some degree of buffering on infiltrating acidic drainage, ii) type II rocks, which although possibly generating alkaline leachate at present, are not regarded to have sufficient excess alkalinity to adequately buffer infiltrating acidic drainage, but are not expected to significantly contribute to net acidity of the waste rock dump; and iii) type III rocks, which are presently producing leachate with high net acidity.  Type III rocks tend to be of variable lithology but in the samples are strongly hydrothermally altered to a sericite-chlorite-clay (SCC) assemblage, and have estimated pyrite contents of greater than 2 percent. The hot, acidic fluids which circulated along the porphyry dyke and volcanic contact zone during and after ore emplacement not only deposited pyrite but caused extensive leaching of primary feldspars and carbonate minerals, thus leaving the rock mass with both acid generation potential and little neutralizing capacity. Type II rocks are also of variable lithology, but are less altered than type III rock. Sampling thus far indicates that pyrophyllite breccia comprises most  157  of the material in this category.  Type I rocks are exclusively of Bonanza Volcanic  lithology, are less altered than type III rock, and contain carbonate minerals.  The acid base accounting boundary criteria between type II and III samples are a net t, and an acid consuming to acid producing CaCO / neutralization potential of -25 kg 3 potential ratio of 0.3:1.  Applying the acid base accounting criteria of the ARD potential categories to the waste rock dumps, the dumps or dump areas considered to be producing net acidic leachate are estimated to contain a minimum of 14 percent of type III material. The EMO and Beach dump areas contain the highest amounts of type III material (53 and 43 percent, respectively). The Old North dump is estimated to contain only 2 percent, while the entire North dump is estimated to contain just over 9 percent of type III material.  Comparisons between water quality estimated from kinetic tests and actual water quality from the EMO and North West dump areas indicate that kinetic test p4rameters such as sulfide oxidation rate constant, sulfate loads, and molar calcium and magnesium loads cannot be confidently scaled to field conditions. However, estimated net acid load and molar calcium and magnesium to sulfate ratios do appear to be scalable, and therefore, can be estimated for the Old North, Cap, and Beach dump areas. Results, given in Table 5.11, indicate that the effluent contributed by Old North dump material has significant excess alkalinity. The estimate for the Cap returned effluent quality more alkaline than expected (its actual effluent quality is thought to be comparable to EMO drainage), and it is believed that the limited number of samples from this dump have resulted in an underestimate of the amount of type III material. Effluent quality from the Beach dump is estimated to be similar in magnitude to EMO, however, due to the unique oxygenation  158  conditions and temperature distributions of the Beach dump, this prediction may not be accurate.  Temporal modeling of the EMO and NWD drainage effluent quality indicates that if the 1988 to 1994 calculated temperature gradient for North West dump (+0.37°C/year) persists, this will be reflected in significant increases in sulfate and acidity concentrations. However, if temperatures stabilize, concentrations are predicted to similarly stabilize and then slowly decrease. Although the principal of the model is believed to be correct, some calibration work is still required to have confidence in the quantities predicted.  6.2  Recommendations  Continued monitoring of down hole temperatures on all of the on-land waste rock dumps is strongly recommended. Sampling frequency should be approximately three months, and a minimum of two year’s monitoring (from January 1993, the initiation of monitoring of the North dump drill holes) is required to assess temperature stability of the North dump.  This study has suffered in some areas due to lack of sampling data in the Cap region of the North dump. However, this may be adjusted for by assuming that the Cap has rock, and hence drainage quality, similar to the EMO dump area.  Decommissioning of the waste rock dumps will likely include a combination of covering or relocating the rock, and continued collection and perhaps treatment of effluent.  159  Covers in the form of spreading 0.3 to 0.5 metres of till on the dump surface will reduce infiltration and can, based on literature, result in decreased oxygenation of the waste rock dumps. Almost all reclaimed areas of the North dump were covered with till for the purpose of aiding revegetation, and there is sufficient till remaining to cover the rest of the dump. The amount of improvement in dump drainage quality as a result of applying a waste rock dump soil cover is still debatable (Morin et al., 1991). This may be in part due to accumulated acidic products within covered dumps delaying a perceivable improvement in effluent.  The waste rock dump literature review of Morin et a!. (1991) proposes that primary air transport into the dump is from the base and pile edges. Therefore, if excess material is available, a thicker till cover should be applied on sides and base of the Cap, the only unreclaimed dump area known to be net acid generating. Careful infilling of the metal and wood refuse pit on the Cap may significantly reduce oxygen transport in this area.  Where possible, the segregation of drainage from the known acid generating regions of the on land dumps will considerably improve the remaining drainage water quality.  160 REFERENCES BHP-Utah Mines Ltd., 1986. Mine, 106 p.  Methods Manual, Environmental Department, Island Copper  BHP-Utah Mines Ltd., and Rescan Environmental Services, 1988. South Wall Pushback: Description of Project and Related Environmental Issues. Prepared for Vancouver Island Regional Reclamation Advisory Committee. BHP-Utah Minerals International, 1990. Island Copper Mine Closure Plan. Bonn, H.L., McNeal, B.L., and O’Connor, G.A., 1985. Soil Chemistry, 2nd ed. New York, John Wiley & Sons. Cargill, D.G, 1975. Geology of the “Island Copper” Mine, Port Hardy, British Columbia, Ph.D. Thesis, University of British Columbia, 25Op. Cargill, D.G., Lamb, J., Young, M.J. and Rugg, E.S., 1976. “Island Copper”, in Brown, A.S.,ed., Porphyry Deposits ofthe Canadian Cordillera, CIM Spec. Vol. 15, pp. 206-226. Caruccio and Giedel, 1981. “Estimated the minimum acid load that can be expected from a coal strip mine”, in Proceedings ofthe Symposium on Surface Mining Hydrology, Sedimentology and Reclamation, Lexington, KY., Dec. ‘7-ll,l981,pp.17’7-l22. Cathies, L.M., and Apps, J.A., 1975. “A model of the dump leaching process that incorporates oxygen balance, heat balance, and air convection”, Met. Trans. B, Vol. 6B, pp.617-624. Davis, G.B., and Ritchie, A.I.M., 1986. “A model of oxidation in pyritic mine wastes: part 1 equations and approximate solution”, Appl. Math. Modelling, Vol.10, . 329 314 pp. Ferguson, K., and Morin, K., 1992. “The prediction of acid rock drainage lessons from the database”, in Procedings ofInternational Symposium of the Abatement ofAcid Mine Drainage, Montreal, Que., pp.85-106. -  Harries, J.R., and Ritchie, A.I.M., 1981. “The use of temperature profiles to estimate the pyritic oxidation rate in a waste rock dump from an opencut mine”, Water, Air, and Soil Pollution, V.15, pp.405-423. Halbert, B.E., Scharer, J.M., Knapp, R.A., and Gorber, D.M., 1983. “Determination of acid generation rates in pyritic mine tailings”, Presented at 56th Annual Conference of the Water Pollution Control Federation, Oct. 2-7, Atlanta, Ga., 15 p. Island Copper Mine, 1988a. 1987 Annual EnvironmentalAssessment Report, VoLI.  161 Island Copper Mine, 1 988b. 1987 Annual Environmental Assessment Report, Vol.11 Kwong, J.Y.T., 1994. Prediction and Prevention ofAcid Rock Drainage from a Geological and Mineralogical Perspective. MEND Report 1.32.1,47 p. Lapakko, K., 1990. “Solid phase characterization in conjunction with dissolution experiments for prediction of drainage quality”, in Doyle, F.M. ed., Mining and Mineral Processing Wastes, Proceedings of Western Regional Symposium on Mining and Mineral Processing Wastes, Berkeley, Ca.. Littleton, Co., AIME, pp.81-86. Lawrence, 1 990a. The Humidity Cell: Principals, Operation and Data Interpretation. Lawrence, Marchant Ltd., the Triton Group, Vancouver B.C. Lawrence, R.W., 1990b. “Laboratory procedures for the prediction of long term weathering characteristics of mining waste, in acid mine drainage”, Designingfor Closure, papers presented at the GAC/MAC Joint Annual Meeting, Vancouver, B.C., BiTech Publishers, Vancouver, pp.13 i-i40. Leitch, C.H.B., 1988. “Brief summary of mafic-potassic style alteration as applied to the Island Copper porphyry deposit, British Columbia”, Unpublished report, 5 p. Li, M.G., 1988?. “Locations of origin and destination of materials hauled to North dump”, Set of maps prepared for North Dump study (UBC MMPE, 1990c), approx. 200p. Li, M.G., 1991. Chemistry of the Drainage from a Waste Dump at BHP-Utah Mines Ltd., Island Copper Mine, M.A.Sc. thesis, University of British Columbia, 201 p. Lister, D., 1993. Modelling of Long Term Acid Rock Drainage from Waste Dumps at Island Copper Mine, Port Hardy, B.C.: Year-End Progress Report, University of British Columbia, unpublished report prepared for G.W. Poling and BHP Minerals Canada Ltd. Lister, D., Poling, G., Home, l.A., and Li, M.G., 1993. Prediction and Reality: “Static analyses versus actual rock weathering in waste dumps at Island Copper Mine, Port Hardy, B.C.”, in Proceedings of 17th Annual B. C. Mine Reclamation Symposium, Port Hardy B.C., pp. 1 09-118. Ministry of Energy, Mines and Petroleum Resources, 1993. Interim Policy for Acid Rock Drainage at Mine Sites, Prepared by B.C. Reclamation Advisory Committee, 13 p. Morin, K.A., 1990. Acid Drainage from Mine Walls: the Main Zone Pit at Equity Silver Mines. Prepared for British Columbia Acid Mine Drainage Task Force, 109 p. Morin, K.A., Gerencher, F., Jones, C.E., Konaseuish, D.E., and Harries, J.R., 1991. Critical Literature Review of Acid Drainage from Waste Rock. MEND Report 1.11.1.  162 Morin, K.A., 1992. Draft Report: Pit-Wall Assessment at Island Copper Mine: Implications for Mine Closure. Morwijk Enterprises, report prepared for BHP Minerals Canada Ltd. Morin, K.A., 1994. Prediction of Minewater Chemistry at Island Copper’s On-Land Dumps. Prepared for BHP Minerals Canada Limited, Island Copper Mine, in press. Morwijk Enterprises Ltd., 1993. Mine Rock and Tailings Geochemistry and Prediction of Water Chemistry. Bell 92 Project, Closure Plan, Supp. Doe. E, 189 p. Nicholson, R.V., 1992. “A review of models to predict acid generation rates in sulphide waste rock at mine sites. in draft proceedings, International Workshop on Waste Rock Modelling, Sept.29-Oct. 1, Toronto. Norecol Environmental Consultants Ltd., 1988. Cinola Gold Project Stage II Report, Volume V, Environmental Research and Special Studies, Prepared for City Resources (Canada) Limited. Osatenko, M.J., and Jones, M.B., 1976. “Valley Copper”, in Brown, A.S.,ed., Porphyry Deposits ofthe Canadian Cordillera, CIM Spec. Vol. 15, pp.130-143. Perelló, J.A., Arancibia, O.N, Burt, P.D., Clark, A.H, Clarke, G.A., Fleming, J.A., Himes, M.D., Leitch, C.H.B., and Reeves, A.T., 1992. “Porphyry copper-molybdenum-gold mineralization at Island Copper, Vancouver Island, B.C.”, presented at Northwest Mining Association Short Course “Porphyry Copper Model Regional Talks and Settings”, Tuscon, Arizona and Spokane, Washington, Nov.28-Dec. 1, 1992. -  Perelló, J.A., Fleming, J.A., O’Kane, K.P., Burt, P.D., Clarke, G.A., limes, M.D., and Reeves, A.T., 1994. “Porphyry copper-gold-molybdenum mineralization in the Island Copper cluster, Vancouver Island”, in Schroeter, T., ed., Porphyry Deposits of the Northwestern Cordillera of North America, CIM Spec. Vol. 46, in press. Perry, E., 1985. “Overburden analysis: an evaluation of methods”, Proceedings, Symposium on Surface Mining Hydrology, Sedimentology and Reclamation, Lexington, Kentucky, College of Engineering, U. of Kentucky, pp.369-375. Renton, J.J., Rymer, T.E., and Stiller, A.H., 1988. “A laboratory procedure to evaluate the acid producing potential of coal associated rocks”, Mining Sci. and Tech., Vol.7, pp.227-235. Rescan Consultants Inc., 1992. Island Copper Mine Decommissioning and Closure Plan, Hydrological and Metal Loadings Study, Phase I Report, Prepared for BHP Minerals Canada Limited, Island Copper Mine.  163 Ritcey, G.M., and Silver, M., 1982. “Lysimeter investigations on uranium tailings at CANMET”, CIM Bulletin, Vol.75, No.846, pp. 1 34-143. Russell, J.B., 1980. General Chemistry, New York, McGraw-Hill, 797 p. “Phosphorus transformations in a soil Singleton, G.A. and Lavkulich, L.M., 1987. chronosequence, Vancouver Island, British Columbia”, Can. J. Soil Sci. 67: pp.787-793. Sobek, A.A., Schuller, W.A., Freeman, J.R., and Smith, R.M., 1978. Field and Laboratory Methods Applicable to Overburden and Mine Soils, Cincinnati, Ohio, U.S. Environmental Protection Agency, Report EPA-600/2-78-054. Sobek, A.A., Bambenek, M.A., and Meyer, D., 1982. “Modified soxhiet extractor for pedologic studies”, Soil Sd. Soc. Am. J. 46: pp 1340-1342. Steffen Robertson and Kirsten (B.C.) Inc., 1989. Draft Acid Rock Drainage Technical Guide, VoL I, Prepared for the B.C. AMD Task Force, BiTech Publishers, Vancouver. Steffen Robertson and Kirsten (B.C.) Inc., 1993. Rock Pile Water quality Modelling, Phase I Final Draft Report. Unpublished report to MEND prediction committee, August 1993.  -  Université Laval Groupe de Recerche en Geologie de L’Ingenieur, 1991. Acid Mine Drainage Generation from a Waste Rock Dump and Evaluation ofDry Covers using Natural Materials: La Mine Doyon Case Study, Quebec. Executive Summary (en anglais) Prepared for Service de la Technologie Miniere Centre de Recherches Minerales, 22 p. University of British Columbia, Dept. of Mining and Mineral Process Engineering (UBC MMPE), 1 990a. Island Copper Mine North West Dump Study, Stage Report II, Humidity Cell Test Results, Prepared for Island Copper Mine. -  University of British Columbia, Dept. of Mining and Mineral Process Engineering (UBC MMPE), 1 990b. Island Copper Mine North West Dump Study, Stage Report I, Prepared for Island Copper Mine. -  University of British Columbia, Dept. of Mining and Mineral Process Engineering (UBC MMPE), 1990c. Acid Mine Drainage Study ofthe North Dump, Final Report, Prepared for Island Copper Mine. Van Zyl, D.J.A., Hutchison, I.P.G., and Kiel, J.E., ed., 1988. Introduction to Evaluation, Design and Operation of Precious Metal Heap Leaching Projects, Society of Mining Engineers, Inc., pp.61-67. Vos, R.J., and O’Heam, T.J., 1993. “Use of zeolite to treat acid rock drainage from Britannia minesite”, in Proceedings of] 7th Annual B. C. Mine Reclamation Symposium, Port Hardy B.C., pp.223-232.  164 APPENDIX I PETROGRAPRIC REPORT ON TEN POLISHED THIN SECTIONS FROM THE ISLAND COPPER DEPOSIT, B.C. FOR ACID ROCK DRAINAGE STUDIES Report for: Diane Lister, Department of Mining and Metallurgy, University of British Columbia, Vancouver, BC.  Invoice attached December 6, 1993  Samples submitted: Ten, from surface dumps (presumably most of the samples come from the open pit). SUMMARY:  The samples submitted form a suite of slightly oxidized, variably pyritic, altered porphyritic and volcanic rocks. The intent of the sampling was to obtain specimens as oxidized as possible by geologic weathering to determine, if possible, the The samples submitted, however, characteristics of oxidation. for a short time from a weathering to have only been subjected from inspection of the evident is This geological perspective. developed, and limonite of traces only show which hand samples, show specimens the section, thin In only. surfaces outside on mainly only a thin rind of limonite (± minor included erratic quartz and sericite grains), and only rarely does limonite penetrate the interior of the sample along fractures. Weathering alteration (as opposed to hypogene hydrothermal alteration) forms mainly very thin rinds up to 150 j.m thick except in one sample (4—8, from which two sections were cut) in which a 2—3 cm zone of “bleaching” (supergene ?clay alteration) is superposed on hydrothermal chloritic alteration. Most sulfide is pyrite, generally euhedral and disseminated and without differences in various parts of the section or in Minor veinlet-controlled chalcopyrite is various gangue hosts. present in two samples(4-8b and 3-7), one with inclusions of bornite (3—7), and traces of galena and ?sphalerite are included Magnetite, with minute inclusions of in pyrite in 8—1. chalcopyrite, is also found in this sample (8—1). Carbonate is found in only one sample (8—1), as rare crystals in a ?zeolite vein; zeolite is also present in 6—1 as Feldspar (plagioclase) occurs veins or matrix to massive pyrite. sericite is common and observed; not was gypsum 8-1; only in abundant, as are chlorite, clay, and quartz in these mainly phyllic, argillic, propylitic, or advanced argilic altered samples of intermediate volcanic rocks (andesite—?dacite). Rare minerals in the highly aluminous, advanced argillic alteration include blue dumortierite and clear ?diaspore.  Craig H.B. Leitch, Ph.D., P. Eng. (604) 666—4902 or 921—8780  _  :  Page 2  165  2-8: QUARTZ-PYROPHYLLITE-SERICITE-DUMORTIERITE-CLAY-PYRITE (ADVANCED ARGILLIC) ALTERED PORPHYRY On the fresh cut face, the hand sample is light creamy greywhite with spots of blue dumortierite and scattered pyrite, and is presumably from the hypogene high alumina alteration (pyrophyllite-dumortierite) zone that capped the Island Copper The specimen is not magnetic, and shows no reaction to deposit. The outside is It is scratched by steel. Rd. dilute cold coated by pale tan—coloured (?jarositic), soft pulverulent In limonite which does not appear to penetrate into the sample. polished thin section, the mineralogy is: 30% Quartz 20% Pyrophyllite 15% Sericite 15% Clay 10% Chlorite 5% Dumortierite 3% Pyrite 2% Rutile <1% Limonite The sample is composed of a generally fine-grained, tightly interlocking intergrowth of quartz and phyllosilicates, the Quartz is generally latter probably developed after feldspars. forms radiating pyrophyllite mm; 0.1 subhedral and less than sericite is finer, whereas diameter, mm to 0.1 masses of flakes Large patches of a size. in jm 20 about flakes subhedral forming LLm) low birefringence mineral look to be fine—grained (25—40 4 distinctly green in some places, with anomalous blue birefringence and length—slow character suggesting an Fe—rich chlorite; however, elsewhere a similar mineral but lacking the green colour may be clay, possibly kaolinitic. In places there are patches of coarser quartz (to 0.25 mm), Dumortierite (blue in hand pyrite, dumortierite and rutile. euhedral crystals up to 0.5 mm to sub— colourless specimen) forms as very fine—grained patches found is leucoxene) (or long; rutile pseudomorphing former Ti oxide crystals up to 0.3 mm long. Pyrite forms subhedral, scattered crystals up to 1.5 mm diameter that contain inclusions of quartz and phyllosilicate up to 0.3 mm Almost all the pyrite occurs in the coarse—grained across. patches, and is all similar in habit (disseminated; no veinlets). There does not appear to be any change in character of the pyrite from crystal edge to center, and there is no evidence of oxidation of individual crystals (no limonite rimming or The polish of pyrite crystals is penetrating along fractures). the section; grain boundaries edges of the at except excellent There are no other are generally smooth to rarely decussate. suif ides visible. Limonite is confined to narrow fractures less than 0.1 mm thick, which near the outside margin of the sample become In this sample, hydrothermal pervasive (affect 50% of the rock). (advanced argillic) hypogene alteration is easily distinguished from recent weathering, which appears to be restricted to No iron oxide rind on the specimen limonitic fracture coatings. section. thin in is visible  Page 3  166  2-9: QUARTZ-SERICITE-PYRITE (PHYLLIC) ALTERED FINE FELDSPAR ?OUARTZ PORPHYRY Fine—grained porphyry consisting of 20—30% 1 mm white feldspar crsytals and patches of disseminated pyrite, probably There are after former mafic sites, in a buff aphanitic matrix. 1.5 and scattered biotite, rare dark brown patches of ?secondary in general is rock The mineral. mm amygdules of a hard white softer than steel; it shows no reaction to HC1, and is not The outside is coated with soft, pulverulent, very magnetic. In polished thin pale tan-coloured (?jarositic) limonite. y: is approximatel mineralogy modal the section, 50% Sericite (after feldspars) 5% Muscovite (after mafic phenocrysts) 25% Quartz (mainly matrix) 5% (phenocrysts, veins) 10% Pyrite 5% Relict feldspar <1% Rutile This rock consists of highly altered feldspar and mafic mineral Feldspar relics in a very fine—grained siliceous groundmass. and have mm long, to 1 up outlines euhedral to subshow crystals been almost completely pseudotnorphed by fine—grained (<0.1 nun) Rarely, feathery anhedral feldspar remnants to 0.05 mm sericite. Mafic crystals also have long are seen in the feldspar sites. and have been diameter in euhedral outlines up to 0.5 mm muscovite as or sericite ined coarser—gra pseudomorphed by places by pyrite. and in diameter, mm 0.3 to euhedral flakes up to 1.7 mm across, up are ainygdules or ?phenocrysts Former quartz that are crystals euhedral to subhedral coarse of composed and highly fractured, somewhat strained and partly replaced by fine Similar grained secondary silica and lesser pyrite in places. quartz is found in rare quartz veins that are up to 0.3 mm thick. The groundinass is formed of extremely fine, rounded quartz crystals up to 25 j.m diameter. Pyrite forms generally subhedral crystals up to 1 mm diameter that frequently contain silicate inclusions up to 0.05 Pyrite grain nun diameter, with rare rutile inclusions. rough and highly to rounded and smooth from vary boundaries There is no apparent change in the character or decussate. association of the pyrite across the section; all pyrite grains Polish are disseminated rather than controlled along veinlets. is good except at the edges of the section, and there is no No obvious difference between the cores and rims of crystals. appear and gypsum zeolite other suif ides are visible; carbonate, to be absent. Limonite is essentially absent from this section (althogh seen on the outside of the hand sample, no rind can be seen or Rutile is common as fine—grairied measured in the section). replacements of former Ti02 minerals that were up to 0.15 nun long. Alteration to quartz-sericite-pyrite (phyllic assemblage) is intense and easily differentiable from supergerie oxidation and weathering, the effects of which are confined to the very outer surfaces of this sample.  Page 4  167  SERICITE-CLAY-CHLORITE ALTERED FINE FRAGMENTAL ?VOLCANIC Light grey-green, coarsely textured igneous rock containing grey-white patches to 0.5 cm diameter and dark green chioritic There is rio reaction to patches and minor disseminated pyrite. cold dilute MCi and no magnetism; the rock is about as hard as Limonite is confined to rare traces of orange—brown steel. In polished thin section, the modal goethite on fractures. y: approximatel mineralogy is 45% Sericite 20% Clay (?) 15% Chlorite 15% Quartz (?partly secondary) 3% Pyrite 1% Rutile tr Chalcopyrite (rare ?bornite) rock, composed of fragmental This appears to have been a fine to 0.5 up about rock ?volcanic of clasts subangular to subrounded finely comminuted more but similar of matrix cm size in a material, before thorough sericite-chiorite-clay alteration, Relict textures in the clasts are accompanied by minor pyrite. variable, from coarsely porphyritic (sericitized feldspar pseudomorphs to several mm long) to finely ?amygdular (0.1 mm However, the intense alteration has largely obscured quartz). the primary derivation of the fragments. Sericite forms abundant fine euhedral flakes and rosettes, Some patches are clearly rarely up to 50 um in diameter. probably feldspar, but others phenocrysts, former of s pseudomorph Other fragments are replaced appear to be after volcanic shards. by very fine-grained (10-25 m) flakes of a mineral that may be clay such as kaolinite; it appears to be intermixed with or Some feldspar crystals are gradational to the sericite. pseudomorphed by masses of a brownish (only partly translucent) mineral that may be clay-altered feldspar. Chlorite is found in discrete patches up to 1.2 mm across, Chlorite flakes are in places rimming cores of sericite. with optical characteristics diameter, jm to 25 subhedral and up birefringence) anomalous colour, green weak , (length—fast Quartz is found as fine suggestive of a Fe:Mg ratio about 1:1, um diameter in both fragments sub— to anhedral crystals up to 25 4 and matrix, and may be partly secondary. Pyrite in this sample forms fine, scattered disseminated crystals up to 1 mm diameter with sub- to euhedral outlines. Most have smooth boundaries, but some are highly decussate. I see no change in Inclusions are of gangue and are mainly rare. across the pyrite the of character disseminated the character or d clay— found fine—graine in most are host; gangue with section or Rare chalcopyrite is seen to 0.25 mm across, with sericite. highly anhedral outlines and very fine (5—10 LLm) inclusions of Rutile is common, as very fine crystals rarely to 0.1 ?bornite. There is no liinonite in the section. mm size. This is an intensely clay—sericite—chlorite±pyrite altered ?fragmental volcanic rock, showing no microscopic evidence of Carbonate, zeolite and recent (supergene weathering) alteration. gypsum appear to be absent. 3-7:  Page 5  168  4-8a: SERICITE-QUARTZ-CHLORITE-PYRITE ALTERED COARSE FRAGMENTAL VOLCANIC Coarsely fragmental volcanic rock containing large subrounded to subangular clasts up to 2 cm in size with dark green chioritic alteration (softer than steel) in a grey, The hand specimen shows siliceous matrix (harder than steel). react to cold dilute HC1. does not but magnetism, rare traces of 2.5 cm thick developed about rind weathering There is a clear the green colour is which in sample, the of outside from the The polished thin section covers only bleached to grey-white. Soft, pulverulent limonite (goethitic) this bleached portion. and ?clay coats the outside surface of the sample but does not Mineralogy in section is: penetrate the sample. 50% Sericite, clay 35% Quartz (largely secondary?) 10% Chlorite 5% Pyrite <1% Rutile This is a thoroughly quartz—sericite-chlorite—pyrite altered fragmental volcanic, in which the original textures are largely Most of the alteration is hypogene, destroyed by the alteration. but as noted from the hand specimen, in the area from which the section is cut, at least part of the clay-sericite alteration may This distinction be supergene (due to weathering processes). section. thin in cannot, however, be made Sericite forms very fine subhedral flakes, rarely up to 50 In places there jm in diameter, and in rosettes to similar size. may be minor clay (finer-grained, lower birefringence) intermixed Quartz is abundant, mainly secondary, in the with the sericite. form of anastamosing veinlets up to 0.5 mm thick, and as adjacent Crystals in or disseminated patches of finer replacement silica. the veins are up to 0.25 mm in diameter; in the fine patches, Chlorite is found as sub- to they average around 25-50 j.m. with bright blue anomalous diameter, mm euhedral flakes to 0.1 indicating moderately pleochroism green and interference colours intimately associated is It around 0.6). (Fe/Mg high Fe content of former alteration the results from probably and with rutile, 2 minerals) in the original volcanic rock. mafic minerals (± Ti0 Carbonate, zeolite, feldspars, and gypsum appear to be absent. Pyrite forms coarse, cubic, euhedral crystals up to 2 mm size, disseminated througout the section, generally away from any There is no clear association with any veins or fractures. particular gangue mineral, and therefore no change in association Most crystals have smooth, regular, straight across the section. with minor gangue inclusions being gangues, boundaries against poor polishing of the crystals. somewhat to leading and common No Cores of crystals are richer in inclusions than the rims. other sulf ides are present, but rutile crystals are present both as separate crystals and inclusions in pyrite to 0.1 mm. Limonite is very rare in the section, where it is restricted to the periphery of the rock, along a few thin fracture margins, It and as rare fine stains in chlorite and sericite near rutile. is not, however, present near pyrite.  Page 6  169  4-8b: INTENSELY CLAY-SERICITE-CHLORITE-QUARTZ (±HYDROBIOTITE, PYRITE) ALTERED VOLCANIC FRAGMENTAL The hand specimen for this sample is very small, and shows no traces of oxidation or limonite even on the outside surface; The fresh cut minor white material there may be only mud. altered volcanic chioritic green, ned, medium-grai a surface shows is slightly and HC1 dilute cold to react not does that rock It is much softer than steel, implying little quartz. magnetic. Modal mineralogy in polished thin section is: 50%% Clay—sericite 20% Chlorite (Fe-rich) 20% Quartz (mainly secondary) 5% Hydrobiotite 3% Pyrite 1% Rutile <1% Apatite tr Chalcopyrite This rock is similar to 4-Ba except that chlorite is more If it is from the green (less weathered) part of 4-8, abundant. then it clearly shows the difference between hydrothermal alteration (green, chloritic) and recent weathering alteration Large domains (subrounded to (bleached, clay-sericite rich). to cm size) consisting mainly of 1 subangular ?volcanic clasts, d (50-100 jnu) sericite coarse-graine chlorite, ruti].e, relatively hosted in a finer— are pyrite, some places in or muscovite, and are also domains There clay-sericite. of mesh jim) grained (10—25 of almost pure quartz (secondary, up to 0.5 mm subh- to anhedral Minor crystals, possibly fragments of altered plutonic rock. amounts of apatite, as minute crystals up to 0.1 mm long, are found associated with the coarse quartz. Chlorite has very strong anomalous blue and purple interference colours and green pleochroism, with length—slow character, indicating Fe-rich composition about 0.6 to 0.7 Subhedral flakes and radiating rosettes are up Fe/(Fe+Mg) ratio. Chlorite is not associated with pyrite, diameter. in to 0.25 mm domains (?former mafic phenocryst in found is frequently but relics) that contain small crystals of a brown, weakly pleochroic, flaky mineral of similar size to chlorite that may be This mineral might be associated with the hydrobiotite. weathering front in this rock (it could be the first stage in the breakdown of the chlorite, which is the mineral most noticeably lost during the weathering). Pyrite is found as scattered, euhedral crystals up to 0.75 The mm in size or as aggregates to 1 mm of finer cubic crystals. with or mineral gangue any with associated clearly not is pyrite the numerous quartz veins; the only exception to this is that the Traces of fine—grained aggregate occurs in a domain of chlorite. chalcopyrite are present as anhedral crystals to 0.05 in long, Most of clearly distributed only along a thin quartz veinlet. very few with surface, the pyrite polishes very well to a smooth at the crystals larger in inclusions showing, but concentrated Where s). ?overgrowth pyrite (suggesting cores, with clear rims quartz elongate shows pyrite the veinlet, quartz crossed by a inclusions, also suggesting possible late overgrowths.  Page 7  170  6-1: CLAY-?SERICITE-CHLORITE ALTERED. ?DACITIC VOLCANIC PORPHYRY Pale grey—green, fine feldspar-mafic mineral porphyry; softer than steel, slightly magnetic, no reaction to cold dilute Outside of sample is coated by thick coating of soft, HC1. pulverulent ?oxides and clay, generally creamy—coloured but tan Within the rock, the effects of weathering appear in places. Mineralogy in polished thin confined to hairline fractures. section is approximately: 25% Sericite 25% Clay (?kaolinitic) 25% Quartz (partly secondary) 10% Chlorite 10% Pyrite 5% Feldspar (plagioclase microlites, relict) <1% Rutile, sphene The section appears to be overly thin, making identifications on However, it is the basis of interference colours difficult. former which feldspar clearly a porphyritic volcanic rock in have been completely long) 1 to mm phenocrysts (30%, euhedral, up may be that mineral matted inn) (to 50 replaced by a fine-grained above) looks to be noted (as birefringence the sericite, although the if kaolinitic, possibly mineral, clay a could be It low. too Former mafic phenocrysts section is of true (30 J.Lm) thickness. (15%, euhedral, up to 0.5 mm long) have been pseudomorphed by chlorite, Fe—Ti oxides such as sphene and rutile, and pyrite. There were micrphenocrysts of Fe-Ti oxides up to 0.2 mm in size, The groundmass is now replaced by sphene and/or rutile. (25—30 inn) quartz, fine-grained very aphanitic, and consists of variably preserved which are microlites, clay (after plagioclase typical of ?sphene), and (rutile oxides Fe—Ti fine and in places) absent. are gypsum and zeolite Carbonate, volcanic. a dacitic Along a single fracture crossing the slide, all original minerals appear to be converted to clay; this could be due to supergene (weathering) alteration since it is along this fracture in the hand specimen that weathering is seen to be taking place. Hydrothermal alteration is clay-sericite-chiorite, or argillic in character. Pyrite forms euhedral cubic crystals up to 1 mm diameter, or in places aggregates of finer cubes, that are disseminated evenly In detail, it appears to show a preference through the specimen. with chlorite, but its size and habit sites, mafic altered for Pyrite are no different in those sites than in the groundmass. is not found along veins nor along the fracture crossing the slide, supporting its identication as a supergene rather than a In some areas, the pyrite crystals are hydrothermal feature. strongly fractured (notably at the edges of the section, but Elsewhere, contain only rare inclusions of silicate or rutiJ.e. cores) crystal at (mostly inclusions few a the crystals contain except smooth are of pyrite boundaries The unfractured. but are other are no irregular. There are which grains, aggregate for Limonite is not seen in thin section, either suif ides visible. around pyrite, along fractures, or on the outside of the sample, so the thickness of the weathering rind cannot be judged.  171 Page 8 6—2: MASSIVE PYRITE VEIN (MINOR SERICITE-OUARTZ-?ZEOLITEL WEATHERED TO CLAY-LIMONITE ON OUTSIDE SURFACES AND FRACTURES Mainly massive pyrite, presumably from a vein of at least 2 Some traces of a siliceous, altered walirock are ess. thickn cm There is a relatively thick (1 mm) weathering or attached. oxidation rind developed, in which buff-cream ?clays and brown This penetrates rarely along a goethitic limonite are abundant. specimen is not magnetic and The few fractures into the rock. In polished thin section, HC1. dilute shows no reaction to cold is: the modal mineralogy Hvpogene 75% Pyrite 5% Sericite 5% Zeolite (?) 5% Quartz (secondary, i.e. part of the vein) tr Rutile Supergene 5% Clay 5% te) ?jarosi ite ± Limonite (goeth Pyrite, the main component of this sample, is found as interlocking subhedra]. crystals up to 1.5 mm across, containing It is possible that some larger crystals up to 4 mm in length. the finer hosting crystals have been cataclasized (broken by deformation), since the larger crystals show prominent Inclusions (silicate and rare rutile) are not common fracturing. in the pyrite except in a few areas of the section, without any No other suif ides are seen obvious connection to other features. n pyrite crystals are betwee ices interst The in the section. ls. minera gangue of k networ filled with a Gangue minerals hosting the pyrite appear to be mainly Quartz forms anhedral, quartz, sericite and rare ?zeolite. interlocking crystals up to 0.1 mm diameter, and sericite forms The mineral tentatively subhedral flakes to 50 j.m diameter. identified as zeolite is found as sub- to euhedral crystals up to 0.5 mm long, with well—developed longitudinal cleavage and lesser cross—cleavage; birefringence is low (0.005—0.010), relief is lower than epoxy, and the crystals are length-slow with Optic angle is small and extinction angle of about 20°. a range of zeolites, such as s fit teristic charac these negative; h the large extinction althoug te; scoleci or ite chabaz , stilbite angle favours scolecite, zeolites are notably difficult to identify optically with certainty. Towards the outside of the section, a cloudy or semi—opaque alteration becomes prevalent; this is probably clay, a weathering In places, crystals or alteration, mixed with minor limoriite. igence could be birefrir e extrem aggregates to 0.25 mm with on of massive pyrite. oxidati the ed in expect l jarosite, a minera absent. be to appear gypsum and r, Carbonate, feldspa  Page 9  172  7-1: CLAY-SERICITE-CHLORITE-OUARTZ-EPIDOTE ALTERED VOLCANIC Fine-grained, grey-green, homogeneous finely porphyritic volcanic flow rock, strongly hydrothermally altered, with minor Cut by rare limonitic fractures; outside is disseminated pyrite. strongly coated by a weathering rind up to 0.1 mm thick of bright orange—brown (goethitic) limonite typical of oxidation of low The specimen is softer than steel due to the sulfide contents. a trace of magnetism, probably relict from shows and alteration, magnetite in the unaltered volcanic; it shows no reaction to cold dilute HC1. Mineralogy in polished thin section is approximately: 45% Clay—sericite 25% Quartz (partly secondary) 20% Chlorite 5% Epidote 2% Pyrite 2% leucoxene Rutile, 1% Limonite (goethitic) (propylitic lorite e-quartz—ch clay—sericit strongly a is This Former euhedral argillic) altered, finely porphyritic volcanic. to subhedral feldspar and mafic phenocrysts, respectively up to about 0.5 mm and 0.25 mm, are pseudomorphed by clay-sericite and They are set in a groundmass of quartz chlorite±epidote. (probaby partly secondary) as anhedral, highly interlocked crystals up to 50 ,.Lm size, pius minor sericite and rutile of Sericite replacing former feldspars finer grain size (15—20 ILm). It appears about 10—20 jnn size. of flakes scaly minute occurs as to be mixed with lesser amounts of a semi—opaque or lower birefringence material that could be clay and/or intermixed This material is even finer-grained, clay/epidote (saussurite). The clay could be the result of perhaps 5—10 jnn in size. weathering, rather than hydrothermal alteration. Chlorite replacing former inafic sites forms subhedral flakes up to 25 j.m diameter, with pale green colour, weak anomalous blue birefringence, and length-slow character suggesting an In places, fine intermediate Fe/(Fe+Mg) ratio about 0.5. euhedral crystals of epidote or clinozoisite (Fe-poor epidote) to Rarely, quartz, 25 jnn are found in the relict inafic sites. pyrite and ?sphalerite as 20-30 jim crystals are also found. A poorly preserved spheroidal texture to other chioritic patches suggests they may be remnants of ?volcanic glass rather than Rounded patches up to 0.5 mm diameter of mafic phenocrysts. fine—grained quartz and ?ciay may represent the sites of former axnygdules. Carbonate, zeolite, feldspar, and gypsum are not seen. Pyrite forms scattered, disseminated crystals of sub- to Inclusions of silicates euhedral habit up to 0.5 mm diameter. are not volumetrically but crystals most in found are rutile and abundant and show no obvious distribution patterns (cores vs. Apart from the occurrence of a few pyrite grains in rims). altered mafic sites, there is no clear control on pyrite No other sulf ides distribution (no veins, no concentrations). pyrite crystals. A on seen are are visible; no limonite coatings can be seen on the rind limonite thin (up to 100 jim) goethitic for 50 jim or penetrate stains limonite in places and specimen, are found along rare fractures.  Page 10  173  8-1: PLAGIOCLASE PORPHYRITIC, ANDESITIC VOLCANIC FLOW ALTERED TO CHLORITE-EPIDOTE-PYRITE (PROPYLITIC) AND LATE ZEOLITE-CALCITE Fine—grained, dark grey, siliceous (harder than steel) altered volcanic probably originally similar to 7-1 (before The rock contains strongly magnetic small dark alteration). spots and disseminated pyrite, but shows only traces of reaction to cold dilute HC1 in late fractures containing a white mineral. Fractures, along which alteration was concentrated, are common. The outside of the specimen is coated with a soft mixture of In polished thin section, orange—brown limonite and white ?clay. y: approximatel is the mineralogy 55% Plagioclase feldspar (?andesine) 20% Chlorite 10% Pyrite, trace galena, sphalerite inclusions 5% Epidote 5% Clay (?) 2% Ilmenite 1% Rutile, sphene, leucoxene 1% only) ?Zeolite (fractures <1% only) fractures (?calcite; Carbonate <1% Magnetite (trace chalcopyrite inclusions) The bulk of this rock is composed of plagioclase feldspar, as medium to large euhedral phenocrysts up to 2 mm long (3.5 mm They are twinned and show traces of where agglomerated). original compositional growth zonation, implying that the ), based on 35 presently observed composition of about andesine (An primary. is Y010, for 22° extinction angles of about Plagioclase also forms the bulk of the groundmass, as fine microlites of about 0.1 to 0.3 mm length, although there is a tendency to senate texture (all intermediary sizes between Most plagioclase crystals microlites and phenocrysts are seen). show only mild alteration, to clay—sericite and minor epidote (saussuritization), along cleavages and fractures. Former mafic sites are replaced by fine—grained (50-100 j.in) chlorite and lesser epidote, with minor ilmenite and ?sphene. Chlorite crystals show blue anomalous birefringence, green pleochroism, and length-slow character typical of Fe:Mg ratios about 0.6; epidote shows no pleochroism, indicating an Fe—poor character like that of clinozoisite. The groundmass to plagioclase microlites is so fine-grained that it is difficult to identify; it appears to be a mixture of 10—20 jm ?clay, chlorite, and Fe-Ti oxides (rutile, leucoxene). In the late fractures, feathery bladed short crystals of ?zeolite (low relief, birefringence about 0.010, length-fast) to about 0.1 mm are mixed with subhedral rhombs of carbonate up to 0.2 mm in Rare grains of size, presumably calcite by its reaction to MCi. chalcopyrite, of inclusions m 10—20 with mm, inagnetite to 0.1 are found in these fracture veins. Pyrite occurs as both disseminated cubic crystals up to 1 mm in diameter, as well as elongate patches of subhedral 0.25 mm The crystals distributed along microfractures and veinlets. disseminated euhedra characteristically display smooth boundaries and zre zoned with a clear core (occasionally containing a few large inclusions) and a rim full of inclusions, both silicate and  174 Page 11 Pyrite crystals along traces of galena and sphalerite (30 jnn). irregularly distributed larger, al, with euhedr less are es fractur inclusions, and have irregular boundaries against silicates hosting them. A thin rind, locally up to 100 j.m thick, of limonite is found around the outer rim of the specimen. There is no observable penetration of limonite into the sample. Grains of quartz and sericite caught up in the limonite are proabably accidental, agglomerated during weathering and erosion. As in other samples, a thin zone up to 150 j.m thick appears to rim the sample, in which ?clay is more prevalent; this may be alteration due to weathering as opposed to the pervasive propylitic (clay— chiorite-epidote—pyrite) hydrothermal alteration.  175 Page 12 8-2: INTENSELY OUARTZ-SERICITE-?CLAY-?DIASPORE (ADVANCED ARGILLIC) ALTERED ROCK; NO SULFIDES OR LIMONITE In polished thin section, the No hand sample provided. mineralogy is approximately: 45% Quartz (largely secondary) 35% Sericite 15% nitic) ?Clay (?kaoli 5% ?Diaspore 1% Rutile of large areas up made is It ides. no suif s contain This specimen up to 7 mm across of either coarse (to 1 mm) anhedral or fine (to 0.1 mm) subhedral quartz, locally with other minerals intermixed, In this matrix, quartz in a matrix of fine sericite and quartz. j.&m in diameter; 50 about to is anhedral to subhedral and up Both are same size. the about to flakes ral sericite forms subhed s to 20 j.tm crystal al euhedr to subFine ixed. intimately interm ions (Fe—rich reflect l interna red—brown with long of rutile, composition) are scattered throughout the matrix and in the quartz-rich patches. Other minerals in the quartz-rich patches include areas of a low—relief, low—birefringence mineral with brown “colour” (semi opaque character), length—fast, that could be a kaolinitic clay mineral such as halloysite (subhedral flakes up to 25 m size). There is also a second (?) clay mineral with similar optical properties but which is clear, and finer grained (10—15 nm). Also included in these areas are clusters of radiating, euhedral, clear crystals up to 0.25 nun long (high relief, high birefringence, length-fast) that may be diaspore (Al hydroxide). This mineral is found in some of the highly aluminous, advanced argillic alteration zones of Vancouver Island. There is no observable weathering rind or rim of limonite on this sample, and no traces of limonite are seen in the interior of the specimen; this is understandable, given the lack of any Carbonate, zeolite, feldspars and gypsum are not sulfide. All the alteration appears to belong to the advanced present. (quartz-sericite-caly-?diaspore); none can be assemblage argillic The protolith lithology of this ring. weathe to ted attribu intensely altered rock can no longer be distinguished.  176 APPENDIX 2  List of Sample Numbers and Description  Sample # I 2 3 4 5 6 7 8 9 10 11 12 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41  Site # I 1 1 2 2 2 3 3 3 4 4 4 2 4 5 6 7 8  1 1 1 2 2 2 3 3 3 4 4 4  Location!Descriptio Coil; pre-test Coil; pre-test Coil; pre-test Coi 2; pre-test Coi 2; pre-test Coi 2; pre-test Coi 3; pre-test Coi 3; pre-test Coi 3; pre-test Coi 4; pre-test Coi4;pre-test Col 4; pre-test HC 7; pre-test HC 8; pre-test HC 1; pre-test HC 6; pre-test HC 4&5; pre-test HC 2&3; pre-test Till; Upper Caps Till; L.N. Dump Coil; post-test Coil; post-test Coil; post-test Coi 2; post-test Coi 2; post-test Col 2; post-test Col 3; post-test Col 3; post-test Col 3; post-test Col 4; post-test Coi 4; post-test Col 4; post-test  <2 2 <2 <2 <2 <2 <2 <2 <2 2 <2 <2 <2 <2 <2  <2 <2 <2 <2 <2 <2 <2 <2 <2 <2 <2 <2 <2 <2 <2  5 <5 <5 <5 <5  38 40 34 26 25  3323 7.33 3268 7.62 3192 6.70 957 6.06 827 5.41  14  5219.0 7 17  16 41  3.91  3110463.98  7.1  67  <1 <1 <1 <1 1 1 <1 1 1 <1 9  .13 .15 .13 .07 .13 .14  .12 .10 .09 .09 .06 .05 .04 .04 .05 .05 .08 .08 .06 .08 .09 .06  2.30 3.80 3.95 3.84 2.43 2.25 2.15 1.34 1.57 1.62  6 9 11 10 9 9 8 12 15 65  9 2.23 7 2.19 7 2.08 121.75 7 2.21 331.90  .25 .16 .17 .16 .03 .03 .03 .04 .05 .05 .03 .02 .03 .05 .03 .08  34 81 135 81 45 34 48 19 20 29 56 58 44 36 58  1.07 1.64 1.63 1.68 1.21 33 43 40 44 61  48 1.18 31 1.10 24 .52 25 .63 20 .64 26 1.23 32 1.13 29 1.20  11.64 12 1.24  61.90189  4 8 7 7 3 4 4 3 3 3 6 6 6  4 6  .056 .070 .070 .068 .071 .075 .076 .086 .082 .086  2.00 2.14 1.88 2.22 .85 56 .67 51 .53 36 .73 401.00 42 .95  59 1.18 .078 54 1.08 .073 58 1.22 .076 44.93.092 55 1.25 .069  <2 <2 <2 <2 <2 <2 2 <2 <2 <2 16  <2 <2 <2 <2 <2  56  83 88 88 84 55  .50.087  36  DATE RECEIVED:  -  NOV 10 1993  AJW  SIGNED  SairLes beginning ‘RE’ are dupLicate samLes.  DATE REPORT MAILED:  SAMPLE TYPE: PULP  .JD.TOYE. C.LEONG, J.WANG; CERTIFIED B.C. ASSAYERS  .500 GRAM SAMPLE IS DIGESTED WITH 3ML 3-1-2 HCL-HNO3-H20 AT 95 DEC. C FOR ONE HOUR AND IS DILUTED TO 10 ML WITH WATER. ICP THIS LEACH IS PARTIAL FOR MN FE SR CA P LA CR MG BA TI B V AND LIMITED FOR NA K AND AL.  39  122  62  16  STANDARD C  36  71 <2 <2 64 <2 56 <2109 <2 94 <2 ‘2 <2 <2 <2 <5 <5 <5 <5 <5  25 13 18 18  11 685 4.72 10 640 4.72 11 573 4.71 8711 2.57  .5 .4 .4 .3 .4  17 140 21 147 15 102 51 227 20 131  24 844 29 831 24 660 38644 24 835  0039 0040 0041 0042 0043 585  2 <2 <2 2 <2  .6 .6 .3 1.0 .5  <2 <2 <2 <2 <2  <5 <5 <5 <5 <5  12 17 13 14 20  19 970 6.50 17 1051 6.27 7 526 2.44 8 640 2.37 8 648 2.43  38 29 15 14 14  .5 .5 .2 .6 .4  63 235 65 367 20 103 33 159 51 147  10  <2 <2 <2 <2 <2  .8 1.7 .4 .8 .7 50 75 60 68 83  <2 <2 <2 <2 <2  23 23 21 13 16  <2 2 3 <2 <2  .9 93 94 5.8 97 10.2 94 7.8 70 1.0  <2 2 <2 <2 <2  <2 <2 <2 <2 <2  <5 <5 <5 <5 <5  9 37 35 39 13  4.12 8.12 8.70 8.04 5.93  15 761 21 3228 22 3497 22 3286 18 958  27 24 22 27 30  7 170 .9 519 81 1068 1.7 593 2056 907 2.4 520 537 117 1444 1.5 .6 114 132 239  5 147 6 115 23 469 28 716 28 630  <1 .13 .14 .05 .05 .05  .04 .05 .01 .10 .08 2.10 4.13 .80 1.87 1.82 13 13 14 6 5 14 .08 36 .15 13 <.01 26 .27 25 .26 36 1.40 23 1.87 10 .24 36 .96 34 .93 3 4 3 5 5 62 .56 .035 87 .71 .071 15 .49 .090 80 2.39 .053 78 2.37 .050  1.13 1.22 .95 1.26 2.26  38 <2 3 <2 <2  <2 <2 2 <2 <2  <.2 1.0 1.6 .5 .3  47 138 25 66 64  <2 <2 <2 2 <2  <2 <2 <2 <2 <2  <5 <5 <5 <5 <5  90 101 28 13 11  23 2715 13.14 20 6709 13.87 7 401 1.56 12 583 3.92 12 572 3.66  32 16 24 25 24  21 779 781 215 9.6 12 325 316 420 2.8 .5 5 28 248 65 .1 4 191 9 83 .1 8 77 3 178  0025 0026 0027 0028 RE 0028  0034 0035 0036 0037 0038  1 1 <1 .06 .07 .07 .06 .13  .12 .09 .04 .08 .16  2.02 2.00 1.8 2.12 4.10 10 9 8 9 15 .03 .03 .03 .03 .07  49 54 36 56 32 13 12 31 20 106 5 6 3 6 4  .067 .068 .079 .072 .043  1.31 1.22 .61 1.48 3.48  50 53 46 52 83  <2 2 <2 <2 <2  <2 <2 3 <2 <2  .3 88 .4 92 62 1.5 .7 97 113 32.3  <5 <5 <5 <5 <5  18 19 20 17 39  591 3.62 572 3.60 931 5.83 653 3.67 4941 7.63  11 10 17 11 31  14 16 28 28 85  11 111 .6 20 128 .6 .5 77 317 .6 19 180 149 5225 2.0  19 637 23 847 5 118 25 657 6 288  <1  1 2 <1 2 1 .21 .06 .05 .05 .05  .06 .08 .04 .06 .06 11 1.93 14 1.59 15 1.34 131.56 10 1.78 .02 .06 .04 .05 .02  36 34 18 24 41  27 .94 12 .62 16 .60 12 .52 ii 1.04  4 4 3 3 5  .084 .089 .095 .078 .065  47 .69 62 .89 36 .88 371.04 43 1.14  <2 <2 <2 <2 3  0012 0022 0023 0024  11 7 7 6 6  1 1 1 <1 <1 .09 .10 .12 .12 .04  .08 .04 .03  12 3.46 11 1.81 9 2.02  .19 .02 .02  <2 <2 <2 <2 <2  0011  0029 0030 0031 0032 0033  1 <1 <1 1 1 .14 .06 .04 .04 .12  .11  15 3.85  1.3 .9 .4 .4 .4  62 105 70 100 60  <5 <5 <5 <5 <5  21 21 15 15 16  23 893 6.10 8 694 2.29 7 648 1.83 8 435 1.59 9 519 3.12  27 12 12 9 13  1.0 .5 .5 .3 .6  90 298 49 216 33 121 32 108 14 112  7 254 32 599 25 427 22 628 29 737  0009 0010  1 1 <1 1 <1 .09 .06 .05 .12 .12 .13  15 4.16  B  .18  Ti .19  1.48 1.58 1.46 .87 1.10  32 36 29 23 26  8 8 7 3 3  88 3.42 .065 84 2.49 .067 82 2.38 .064 45 .63 .080 47 .64 .075  <2 <2 <2 3 <2  P La Cr X ppm ppm  2 <2 <2 2 <2  Ca Z  113 3.6 103 4.3 96 6.1 61 2.0 49 1.9  ppw  Bi fl  Sb Pi  Cd 5  Sr  V 121 64 129 36 33  Ba Mg X PP  Page 1 V K K ppm  93—3258 Na K  #  PAZ(604)2531716  AL Z  File  Submitted by: Diane Lister  :...:E.pEo(6o4)253...3j58  (I1  0006 0007 0008  23 22 20 20 20  Th  Au PF’  U Fe As X ppii ppn  Mn p’an  1.8 1.8 1.7 .8 .6  36 31 22 26 35  Co ppII  Ni ppa  72 740 81 898 143 1125 110 433 64 413  Ag pçn  Zn pçii  5 480 3 490 5 466 6 142 10 133  Mo Cu Pb pç4I ppn pçn  Dept of Mining & MineraL, Vancouver BC V6T 1Z4  GEOCHEMICAL ANALYSIS CERTIFICATE  ST.. VANCOUVER s.c.  The University of British Columbia  •  0001 0002 0003 0004 0005  SAI4PLE#  1  •A  1  I  CD  CD  178  vv;v  a,  t’  -, I.  Z  0 OD000  — <  0.0—,0.0  .0 0  — w.0*ot-  0.  E  -00.  —  .,0”0.0n 0..frN  N  Z ‘.E  cc c  &  CE  .  N N N  p,’’O’.  0.  0  -  •a) 00000  0’ C  C.  W.0I-NN  0  —  —-  N  >E  c00.0N’0 0. .fl ‘0.0  ,..  .E  NNNN  n  &  tn&  vvvv  — -.  0E ‘.E  U’  N  U’  Ils.0’O’0  LE  NNNVJN VVVV  -  E  NNNNN V VVVV  I’.  —&  U’.  Pfl  a  fU’.U’  a  vvvvv  —  CE 0.  N.f’.$l’.SO  ,‘.  ZE  <a  C C C  ‘0 0  w1.  U.  .0N.tW  ‘  U  °‘“°‘  ‘0.U’U’.’0  0  OE  .—‘-‘000  —  —E  iP.Ot-I N N N  ‘E  ‘0..tN.30  0  CE  C  ‘C  &  ——  N  —  O.0U%  —  N&  .0 E  ‘0’0 ‘0 W W  —  DC a  ‘0O’’’  N  OE  !-,O’0NIts  ‘0  0.  0 .U’,0P’..0  ‘.  •..t0’0I•’.  NNN  ‘0  —  I.  z < (n  0000W  0000  V  179 APPENDIX 4  2 Calculation Method for APP and APPS  given that 1 weight percent equals 10 kg/tonne,  APPS=wt%Sx  10kg it  <  lOOgCaCO / 3 mol 32gS/mol  and  {  (kgCaCO / t) = wt%S _[wt%S0 2 APPS 3 [ 4  32gS / mol 4 /mol)jj 96gSO  10kg it  lOOgCaCO / mol 3 32gS/mol  4 6 8 12 16 20 pan 30 40 50 70 100 140 200 270 400 pan  Mesh # (Can.)  TOTAL  100.00  1.40 1.08 0.89 0.73 0.69 0.57 0.62 0.53 0.53 0.55 6.18 5.11 4.21 3.48 2.79 2.22 1.60 1.08 0.55  0.6 0.425 0.3 0.212 0.15 0.106 0.075 0.053 0.038  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  436.8  0.00 21.45 24.75 12.96 9.23 7.37 5.17 4.44 2.93 2.01 2.11  100.00 78.55 53.80 40.84 31.62 24.24 19.07 14.63 11.70 9.68 7.58  0.0 93.7 108.1 56.6 40.3 32.2 22.6 19.4 12.8 8.8 9.2 33.4 6.1 4.7 3.9 3.2 3.0 2.5 2.7 2.3 2.3 2.4  475.5 617.9 577.5 536.9 511.4 501.7 482.8 471.6 429.6 383.7 363.3 286.6 331 326.8 328.1 321,4 296.2 291.8 248.2 274.5 270 255.6  26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 0.85  1.06 0.75 0.53 0.318 0.265 0.187 0.132 0.0937 0.0661 0.0469 0.0331  475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9 354.1 253.2 324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2  Wt. % Retained  Wt. % Passing  Opening MassSve MassSve MassSamp (g) +Samp(g (g) (mm)  -  18-Oct-93 Date: Prolect: Island Copper ARD Study Done By: DL  Opening (in.)  Sieve Analysis 001-3 Sample: 434.9 Mass (g): col pre-test Desc.:  C  )4444(—  FF  FF  I  —  0.1  ,E 0.uI  10  l00:  II 111/ II III  —I-—  WI.  Opening Size (mm)  I liii  10  Li  11111 II III  I,,’’  II Ill  E1III1 mI1  Size Distribution 001-3  10  40  !  C  C  (0  10  15  20  -—  lfl  35  00  I—  pan ° 2 30 40 50 70 100 140 200 270 400 pan  4 6 8 12 16  Mesh # (Can.)  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  0,6 0.425 0.3 0.212 0.15 0.106 0.075 0.053 0.038  324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2  TOTAL  353.6 345.4 341.9 331.0 303.6 296.5 251.2 277.0 270.7 259.3  605.8  28.7 23.3 17.7 12.8 10.4 7.2 5.7 4.8 3.0 6.1 15.02 11.18 8.25 6.14 4.42 3.24 2.29 1.50 1.01  100.00  4.74 3.85 2.92 2.11 1.72 1.19 0.94 0.79 0.50 1.01  0.00 7.63 8.90 7.46 10.48 9.36 8.83 8.57 7.08 5.33 660  100.00 92.37 83.48 76.02 65.53 56.17 47.34 38.78 31.69 26.36 1976  475.5 570.4 523.3 525.5 534.6 526.2 513.7 504.1 459.7 407.2  0.0 46.2 53.9 45.2 63.5 56.7 53.5 51.9 42.9 32.3  475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9  1.06 0.75 0.53 0.318 0.265 0.187 0.132 0:0937 0.0661 0.0469 00331  26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 085  Wt. % Retained  Wt. % Passing  Opening MassSve MassSve MassSamp +Samp(g (g) (mm) (g)  -  Date: Project: Island Copper ARD Study Done By: DL  Opening (In.)  Sieve Analysis 004-3 Sample: 606.9 Mass (g): cot pre-test Desc.:  ?  0.1 0.01  t  —  —  -  V  .‘  ,.  I  0.1  —  V  V  % Passing  —I—  .  -  —  E:.  —  WI % Retained  10  .-_________  —  -  I Openingsize(mm)  :  .  —  0  25  100  V.  o  [I° i iti—ñ i iii j.N*TDII_—I I LI I III! —LI---i.11-I .L_I I I 111111 I I I IIILllzl I I 111111 I I I IIIIi1’°  100j  004-3  Size Distribution  00  4 6 8 12 16 20 pan 30 40 50 70 100 140 200 270 400 pan  Mesh # (Can.)  TOTAL  100.00  2.55 2.19 1.89 1.72 1.54 1.43 1.14 1.17 0.74 0.72 12.55 10.36 8.47 6.75 5.21 3.77 2.63 1.46 0.72  0.6 0.425 0.3 0.212 0.15 0.106 0.075 0.053 0.038  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  726.1  0.00 8.18 17.99 13.87 11,97 8.87 7.13 6.11 4.31 3.06 3.42  100.00 91.82 73.83 59.96 48.00 39.13 31.99 25.88 21.57 18.51 15.09  0.0 59.4 130.6 100.7 86.9 64.4 51.8 44.4 31.3 22.2 24.8 114.5 18.5 15.9 13.7 12.5 11.2 10.4 8.3 8.5 5.4 5.2  475.5 583.6 600.0 581.0 558.0 533.9 512.0 496.6 448.1 397.1 378.9 471.3 343.4 338.0 337.9 330.7 304.4 299.7 253.8 280.7 273.1 258.4  26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 0.85  1.06 0.75 0.53 0.318 0.265 0.187 0.132 0:0937 0.0661 0.0469 0.0331  475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9 354.1 356.8 324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2  Wt. % Retained  Wt. % Passing  Opening MassSve MassSve MasaSamp +Samp(g (g) (mm) (g)  20-Oct-93 Date: Project: Istand Copper- ARD Study Done By: DL  Opening (in.)  Sieve Analysis 007-3 Sample: 731.5 Mass (g): col pre-test Desc.:  .  •o  i  =  —  —  [j[1Jll  [Itfllf  F FR1fL :  :  U4IftlZ  1111111—  F flITE  -  [lulL  ttti[h  1ER  -  -  -  ——  F’l  -  -  Wi. % Passing  —4--  liTtlE H+I+I—  FUII1 T1= flillF  WI. % Retained  OpeningSize(mm)  L11iJ1I._. -14-141}lf I HI’+H+  tffl : FB141 :  :1 HIIE 11 El Ht1111 I 1 III E  jIj= IiII11E : JJ-’ :1 [fflj : HEITII FF1111— [Ufihl— -L 1111111—  -  :  E  007-3  Size Distribution  5 I1TIIF 144+H’o  riiiii  1ftII25 1-I-filL ll1.l5  35 flhII[  f45 hilMo  F!U1T°  00  4 6 8 12 16 20 pan 30 40 50 70 100 140 200 270 400 pan  Mesh # (Can.)  TOTAL  100.00  0.6 0.425 0.3 0212 0.15 0.106 0.075 0.053 0.038  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  674.2  2.74 2.31 1.84 1.48 1.31 0.98 0.85 0.71 0.58 1.25 11.30 8.99 7.15 5.67 4.36 3.38 2.54 1.82 1.25  26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 0.85  1.06 0.75 0.53 0.318 0.265 0.187 0.132 0.0937 0.0661 0.0469 0.0331  475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9 354.1 356.8 324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2  0.00 20.08 15.51 9.69 5.96 7.89 7.24 6.63 4.97 3.89 4.09  100.00 79.92 64.40 54.72 48.75 40.86 33.63 26.99 22.03 18.14 14.05  0.0 135,4 104.6 65.3 40.2 53.2 48.8 44.7 33.5 26.2 27.6 95.1 18.5 15.6 12.4 10.0 8.8 6.6 5.7 4.8 3.9 8.4  475.5 659.6 574.0 545.6 511.3 522.7 509.0 496.9 450.3 401.1 381.7 451.9 343.4 337.7 336.6 328.2 302.0 295.9 251.2 277.0 271.6 261.6  WI. % Retained  WI. % Passing  MassSve MassSve MassSamp (g) +Samp(g (g)  Opening (mm)  -  25-Oct-93 Date: Project: island Copper ARD Study Done By: DL  Opening (in.)  010-1 Sample: 675.6 Mass (g): coi pre-test Desc.:  Sieve Analysis  a  0  C  i  ‘I’ll  I liii  zI4nm—  r  I I III  ii Iii  ‘f-T-flTlW  TIll l’ TithE  IIIF ITllhIz  1{HIIIIE  141m11  -  IIIll  ——  WL % Passing  Opening  -  —i—  11111110  lli  Wt. % Retained  SIze (mm)  141111—’14t11I1E_ 01  n’Hj  r EllllE  =  0.01  Al  ‘U  °°  Size Distribution 010-1  5  100  1 tttiiit  ,n 11111  l  jjjjj-20  +1111  30  11111-40 11111 11111-35  11  ° 5 tflTllT  C  (0  c)  00  -  28-Oct-93 Date: Project: Island Copper ARD Study Done By: DL  30 40 50 70 100 140 200 270 400 pan  pan  4 6 8 12 16 20  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  1.06 0.75 0.53 0.318 0.265 0.187 0:132 0.0937 0.0661 0.0469 0.0331 0.6 0.425 0.3 0.212 0.15 0.106 0.075 0.053 0.038  26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 0.85  475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9 354.1 356.8 324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2 TOTAL  475.5 524.2 469.4 539.5 521.3 517.8 496.6 484.8 443.1 395.0 378.7 450.4 343.9 338.7 337.8 328.6 302.6 296.1 250.5 278.1 270.3 257.3 391.1  0.0 0.0 0.0 59.2 50.2 48.3 36.4 32.6 26.3 20.1 24.6 93.6 19.0 16.6 13.6 10.4 9.4 6.8 5.0 5.9 2.6 4.1 19.02 14.78 11.30 8.64 6.24 4.50 3.22 1.71 1.05  100.00 100.00 100.00 84.86 72.03 59.68 50.37 42.04 35.31 30.17 23.88  100.00  4.86 4.24 3.48 2.66 2.40 1.74 1.28 1.51 0.66 1.05  0.00 0.00 0.00 15.14 12.84 12.35 9.31 8.34 6.72 5.14 6.29  WI % Mesh # Opening Opening MassSve MassSve MassSamp Wt. % Retained Passing +Samp(g (g) (mm) (g) (in.) (Can.)  022-2 Sample: 391.7 Mass (g): HG pre-test Site #2 Desc.:  Sieve Analysis  It  a  1111111  I  I  ‘  I I  lillilil  = i_IIIii ii  Z jE  111111,  1,1.  ‘I”  lull  I  I  I  II  I  111111  ‘_L.L.LLLL’  IllIlul  Ii-lI--—-l—I-l-I—II-II4——i--i_I-Ii-IiI  ___SI  =  —,  i_iI_IIIIL,1_I  —1—rrrlr,n—  -  ,  I  ,  II  --  :1  I  I  I  Illill  IlIlill  P  I  LI_hulhIl_r_l._I ‘‘t’ l:huuII,,:—-L;,:-,—,III,  Ililillil  IM. % Passing  —.-.  M. % Retained  Opening Size (nan)  S  U)  1998_  0 -i;; TF iHF1fiiii  hIll?,,  11111111  IIIIIIIII i.t’t,._i  ulilu  LILI__LJ_IJLILIL_J_LLLILILI..._LJ_IJLILI  -—-l—t-t-I--IjA-r--—,  -—  —I-—I—pip-ut, —I—I-4tIIII-.——I—l-I-i--II-III_-.l—_i—I-I,-III 11llItl I-,,,,,ut———I—r,-I--, tin—-- r -—-1-t-t-l-I7In——t--l-I-.t-IlIl-—-P—t--rI—n-Iti-—r.i-iirIu  ri—I-I,- I,Ir——-l—rl-rI,Ir,  ..LLLLLLLLLI. :EIItt:E  hhiF liii, iii Iil?IIIiI  E,,j  —-t—II-I--Jt-Ili.-—I--4—I-4  P  F  —-------——________  1  I  10  100  Size Distribution 022-2  .  00  -  28-Oct-93 Date: Project: Island Copper ARD Study Done By: DL  30 40 50 70 100 140 200 270 400. pan  pan  4 6 8 12 16 20  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  1.06 0.75 0.53 0.318 0.265 0.187 0.132 0.0937 0.0661 0.0469 0.0331 0.6 0.425 0.3 0.212 0.15 0.106 0.075 0.053 0.038  26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 0.85 475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9 354.1 356.8 324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2 TOTAL  475.5 524.2 472.3 533.1 508.8 491.2 475.1 465.0 425.4 382.1 362.0 397.4 330.6 327.6 328.9 322.4 297.2 292.7 248.6 274.8 269.8 257.5 206.1  0.0 0.0 2.9 52.8 37.7 21.7 14.9 12.8 8.6 7.2 7.9 40.6 5.7 5.5 4.7 4.2 4.0 3.4 3.1 2.6 2.1 4.3 16.45 13.78 11.50 9.46 7.52 5.87 4.37 3.11 2.09  100.00 100.00 98.59 72.97 54.68 44.15 36.92 30.71 26.54 23.05 19.21  100.00  2.77 2.67 2.28 2.04 1.94 1.65 1.50 1.26 1.02 2.09  0.00 0.00 1.41 25.62 18.29 10.53 7.23 6.21 4.17 3.49 3.83  Wt. % Mesh # Opening Opening MassSve MassSve MassSamp Wt. % Passing Retained +Samp(g (g) (in.) (mm) (g) (Can.)  Sieve Analysis 024-1 Sample: 206.4 Mass (g): HC pre-test Site #5 Desc.:  rn  111111  —  11111111  rc  11111111  C  1T  -  r  iii,  I  11111111  I  IIJIIII  I  .i.%Re{ak1ej  10  III ILl  I 199B  100  rrnrIn—r1rIII  ——r——I—IIrIII 1 IIrryIIl  I  1 Op.ning Size (mm)  Hz55w19  0.1  -1-rrrIrIII——I,I,rIIIr—-  III IIIIl  —4-.  I  —  11111111 I 111111 SI -4I-4IISf-l-I.- CCIIIH ——f-.4-4-IHH 111111111 I I I I III I 4I Ill  .  I frI  —  I  r  --  IrIIIII:f  —_1  II  •II  0.1 0.01  i  r r  I 1111111 --4- I- -II-IIII 11111  —  ——I-.H4-IH—  024-1  Size Distribution  00 I-,’  -  Date: 28-Oct-93 Project: Island Copper ARD Study Done By: DL  30 40 50 70 100 140 200 270 400 pan  pan  4 6 8 12 16 20  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  1.06 0.75 0.53 0.318 0.265 0.187 0.132 0.0937 0.0661 0.0469 0.0331 0.6 0.425 0.3 0.212 0.15 0.106 0.075 0.053 0.038  26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 0.85  475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9 354.1 356.8 324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2 TOTAL  475.5 524.2 469.4 511.0 506.3 497.7 484.3 475.3 435.0 388.6 370.0 422.3 336.8 332.7 332.7 326.0 300.5 294.8 248.9 276.8 270.0 256.7 254.5  0.0 0.0 0.0 30.7 35.2 28.2 24.1 23.1 18.2 13.7 15.9 65.5 11.9 10.6 8.5 7.8 7.3 5.5 3.4 4.6 2.3 3.5 21.02 16.86 13.52 10.45 7.58 5.42 4.09 2.28 1.38  100.00 100.00 100.00 87.94 74.11 63.03 53.56 44.48 37.33 31.94 25.70  100.00  4.68 4.17 3.34 3.06 2.87 2.16 1.34 1.81 0.90 1.38  0.00 0.00 0.00 12.06 13.83 11.08 9.47 9.08 7.15 5.38 6.25  Wt. % Mesh # Opening Opening MassSve MassSve MassSamp Wt. % +Samp(g Passing Retained (g) (in.) (mm) (Can.) (g)  Sieve Analysis 025-f Sample: 255.1 Mass (g) HC pre-test Site #6 Desc.:  0.  Cl]  oa :  :1  ] II]]:  —  r rI] nIl I•IlllIll I I_Il II  —  1 I I ,-nI-,l—;.i -—I—I--I I 1W — 1 r ri—Irl I —1—1111111 i—i IIIl.4 —I 11111 I IlL I P. I I I I_I. I 1.1 I I j__Sl III  I r!t:tOtl :1:1 DC = ]: C CDEI l—I*I+Il_—--I—l—l-l—II-I  1,I  :41:U1tf  —  .1  -.  _4_I1Il__i:l_l1-Ui4_I1-:_i.1-l14  I1i1I]IiIilE*14I11i r 1114111 l—44+14141——-l-—I—I—I4-l4-lI-——I—I-l-I-4I-I —-1—1—1-1-11-114 I JIlL.: JLJ.1 _LL :: CUD  1 I  1  I 1  0.1  °  M. % Retaied Passrng  -.—  (mm)  IIIIIIII 10  Openg SIze  1  I  11111111 I IprIII\ l 1 1—r rrIIrIr\ 11 I 111111111111111 —  1—4———1———4—4———l—l—lI  1111111 I  I  11111111  I  I  I  11111  1111111  i998  0  c  3996  100  IIIIIIIII0  I  I  rri:CtrI;IIllIr:Ui1 rrlIlI:r [rInd ll1IC:nl:C lUll 1 —l—I——IrI,-lr-——l—-r—rI-—ItI —-1—I—rI—Id I I FI•ll4I L _I__I_J J. LLIJLIII __I_ I _IJ_I1III. IIL._J_L LIJLIL I *__ III I) I. I I 1 1111 1. LI I I I L, I I II II —  —  —  =  ——l—l—1-l-lI-IH ——I--I-I *I4l4I  =  ZiZttIibttZrlt41tliI1IZtII  0_i -———I——I 001  10  Il_lu  Size Distribution 025-1  -  Date: 28-Oct-93 Project: Island Copper ARD Study Done By: DL  pan  30 40 50 70 100 140 200 270 400  pan  6 8 12 16 20  06 0.425 03 0.212 0.15 0.106 0.075 0.053 0.038  460.2 452.2 416.8 3749 354.1 3568 3249 322.1 3242 318.2 293.2 289.3 245.5 272.2 267.7 253.2  3.35 2.36 1.7 118 0.85  0.132 0.0937 0.0661 00469 0.0331  00234 0.0165 00117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  475.5 524.2 469.4 480.3  26.5 19 13.2 9.5  1.06 0.75 0.53 0.318  197.6  18.0 19.5 14.7 123 12.5 462 86 7.3 58 4.8 4.5 3.4 2.8 2.1 1.9 3.6  478.2 471.7 431.5 3872 366.6 4030 3335 329.4 3300 323.0 297.7 292.7 248.3 274.3 269.6 256.8 TOTAL  0.0 0.0 0.0 31.3  475.5 524.2 469.4 511.6  100.00  435 3.69 294 2.43 2.28 1.72 1.42 1.06 0.96 1.82  9.11 9.87 7.44 622 6.33  52.53 42.66 35.22 2900 22.67 1832 14.63 1169 9.26 6.98 5.26 3.85 2.78 1.82  0.00 0.00 0.00 15.84  100.00 100.00 100.00 84.16  Wt. % Mesh # Opening Opening MassSve MassSve MassSamp Wt. % (g) Passing Retained +Samp(g (g) (in.) (mm) (Can.)  Sieve Analysis Sample: 027-2 200.2 Mass (g) HC pre-test Site #8 Desc.:  0.1  1  10  100  I  I  1  i  : i  1111  I  iir  :  I  I1I1rrr\-  I  --Wt.%RetainedI  OpeningSize(mm)  ‘“  1I1ITTI_fl  : :  Size Distribution 027-2  I  ii  n rI  1:  I  II-I  LILII  r  :  II  I998  3996  59.94  00  -  30 40 50 70 100 140 200 270 400 pan  pan  4 6 8 12 16 20  00234 0.0165 00117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  .  2.00 1.50 1.06 0.75 0.53 0.318 0265 0.187 0132 0.0937 0.0661 00469 0.0331 06 0.425 03 0.212 0.15 0.106 0.075 0.053 0.038  50.8 38.1 26.5 19 13.2 9.5 67 4.75 335 2.36 1.7 1.18 0.85  643.8 586.5 475.5 524.2 469.4 480.3 4711 469.5 4602 452.2 416.8 374.9 354.1 356.8 3249 322.1 3242 318.2 293.2 289.3 245.5 272.2 267.7 253.2 TOTAL  643.8 842.3 755.9 587.1 556.4 556.1 5514 541.7 5340 546.8 509.0 466.0 488.7 1293.5 4360 468.5 5247 520.6 450.5 367.6 262.4 283.8 274.0 257.3 2335.6  0.0 255.8 280.4 62.9 87.0 75.8 803 72.2 738 94.6 92.2 91.1 134.6 936.7 1111 146.4 2005 202.4 157.3 78.3 16.9 11.6 6.3 4.1 3527 29.00 2042 11.75 5.02 1.67 0.94 0.45 0.18  100.00 89.05 77.04 74.35 70.62 67.38 6394 60.85 5769 53.64 49.69 45.79 40.03 0  ‘  •ç  o  1  10  100  i  II  1  I 1111  —  I  TI  -  --  I  Ei  1 1 :,1u  111111  -  1  I  -:  —  rI  rE  TI  III  iF  1T  -  III  -  I  III  -  I  EL 1 -  OpenlngSlze(mm)  I  ‘I11  I  —  E  1998  3996  I IWO  ;Er:Ilic:ElcI!Ii  —  -  E  :-  i]&.1I4L  TIlT T1TITITr)r Ii ill rT IrrrITITI/ r --1-rrr1TTT7-1-r11rTrr11nnrrctT LIII L £ LIII I II I U UL 1 £ LIITII[ TIlT I I 111111 I I I III7 I I 111111  -  LICII  028  Size Distribution  100.00 Comment: blinding by fine particles on 40# to 140# sieves; therefore probable size misrepresentation  476 6.27 858 8.67 6.73 3.35 0.72 0.50 0.27 0.18  0.00 10.95 12.01 2.69 3.72 3.25 344 3.09 316 4.05 3.95 3.90 5.76  Wt. % Mesh # Opening Opening MassSve MassSve MassSamp Wt. % Passing Retained +Samp(g (g) (mm) (g) (in.) (Can.)  -  Sieve Analysis 05-Nov-93 Date: 028 Sample: Project: Island Copper ARD Study 2341.8 Mass (g) Desc.: Till Composite Upper Caps Done By: DL  00 00  -  30 40 50 70 100 140 200 270 400 pan  pan  4 6 8 12 16 20  0.0234 0.0165 0.0117 0.0083 0.0059 0.0041 0.0029 0.0021 0.0015  2.00 1.50 1.06 0.75 0.53 0.318 0.265 0.187 0.132 0.0937 0.0661 0.0469 0.0331  0.6 0.425 0.3 0.212 0.15 0.106 0.075 0.053 0.038  50.8 38.1 26.5 19 13.2 9.5 6.7 4.75 3.35 2.36 1.7 1.18 0.85  643.8 586.5 475.5 524.2 469.4 480.3 471.1 469.5 460.2 452.2 416.8 374.9 354.1 356.8 324.9 322.1 324.2 318.2 293.2 289.3 245.5 272.2 267.7 253.2 TOTAL  643.8 1105.5 634.1 628.5 640.6 619.7 601.8 573.7 556.1 559.9 513.4 462.1 487.3 1362.1 452.9 460.9 505.3 456.3 413.6 405.7 331.1 350.4 288.7 261.1 2863.5  0.0 519.0 158.6 104.3 171.2 139.4 130.7 104.2 95.9 107.7 96.6 87.2 133.2 1025.3 128.0 138.8 181.1 138.1 120.4 116.4 85.6 78.2 21.0 7.9 30.99 26.15 19.82 15.00 10.79 6.73 3.74 1.01 0.28  100.00 81.88 76.34 72.69 66.72 61.85 57.28 53.64 50.30 46.53 43.16 40.12 35.46 e  S  0.1 0.01  1  10  100  LISILl  —  :1:1:1] n  .2 _CCTStIIl  : r  —  —  —  —  —  : :  —  I_  .2 .4421 1.21..  I  rI  TI  IIIIIiI 0.1  F  —  I  ill  .:  Ill  h Passing  .2  M  .  —  —  C = 2.1  I  .1  —  11111  LLLL  1111111  I  I I_lAIr  : c [[[I]  —  _1_t.  ..2  uu__  .—l—..l—IIII I IIIIiiii 10 1 Opening Size (mm)  J-tr.q  :.::;:::  —  —  :  —  r I  II  0 C  3996  5994  iliiii:0 100  r I f  :nt  _L. I .1.1 11 UL -I LI- LIII. A. _I_LI4 Lila I_I. I- LISII7_ r V rIti r V rrtk#I ——t—I-r-It nit— —r 1111 nr- —-1 -1 mrrrl1TITIli1nrrmT,lrIrIr—vrrrlrl nnr r r CITA. 111111 I 1 111 I r T 111111 _i_ .111 IJLJI. ..J.. 1. L .1111 .1.. L I-LI/till... I. .1_I__IA I_Ill 11111111 11111 III 21111111 I I I 1111111 I  alL——  L1IIiI1I.._L .I.I..IJIlII •.1. 1.111 lIJI_ I .1 paIl 1111111111 I 1112111 11.1111111 I I 1111111 11111 II  [[31:  _J_LLLIIII_/_  I  :1: r  029  Size Distribution  100.00 Comment: blinding by fine particles on 50# to 140# sieves; therefore probable size misrepresentation  4.47 4.85 6.32 4.82 4.20 4.06 2.99 2.73 0.73 0.28  0.00 18.12 5.54 3.64 5.98 4.87 4.56 3.64 3.35 3.76 3.37 3.05 4.65  Wt. % Mesh # Opening Opening MassSve MassSve MassSamp Wt. % Passing Retained +Samp(g (g) (mm) (g) (in.) (Can.)  -  Sieve Analysis 02-Nov-93 Date: 029 Sample: Copper ARD Study Island Project: 2878.4 Mass (g) Till Composite L. N. Dump Done By: DL Desc.:  00  190 APPENDIX 6  Moisture Contents, Sites 1 through 4 Sample Pits  Sites I through 4 Sample Pit Profiles FROM(m)  TO(m)  MOISTURE%  1  0  0.1  1.5  1  1  0.1  0.5  2.6  1  1  0.5  1  2.8  2  1  0  0.3  6.9  2  1  0.3  0.3  9.3  3  1  0  0.08  9.8  3  1  0.08  0.2  2.2  3  1  0.2  0.5  3.9  3  1  0.5  1  2.5  3  2  0  0.08  7.9  3  2  0.08  0.2  2.7  3  2  0.2  0.5  3.2  3  2  0.5  1  4.4  4  1  0  0.1  4.1  4  1  0.1  0.3  3.3  4  1  0.3  0.5  1.8  4  1  0.5  1  2.7  SITE#  PROFILE#  1  191 APPENDIX 7  Leachate Quality Analytical Techniques  Dissolved Metals Appoximately 1.5 litres was filtered through O.45um cellulose nitrate for dissolved metals analyses. One twentieth of the total leachate volume was taken from the filtrate and added to a four week composite sample, with the remaining one litre of filtrate becoming the weekly dissolved metals sample. All samples were preserved with trace metal grade concentrated nitric acid (8 ml acid per 1000 ml of sample) before being sent for analysis.  pH pH was measured using a Corning combination electrode and Coming 150 pH/ion meter. The electrode was calibrated using 7.00 and 4.01 buffer solutions immediately prior to sample measurement, and checked using the 7.00 buffer every five samples measured. The Coming 150 pH/ion meter was equipped with an automatic sensor which outputs a pH value when the meter estimates that equilibrium was reached. For leachate pH measurement, a second equilibrium reading was taken after reaching initial equilibrium to ensure that the meter has recorded a true value. The pH measurement procedure was altered after week 6 when it was discovered that the electorde was getting coated with a precipitate from the leachate. Subsequently, the electrode was briefly immersed in 10% HC1 solution between sample measurements. This amendment produced more consistent results than previously.  Eh Eh was measured using a platinum and calomel reference electrode. As in the pH measurement procedure, both electrodes were briefly immersed in 10% HCI between sample measurements to prevent precipitate build-up. A standard +430 mV(cal.) reference solution was measured immediately prior to sample measurements. As in pH measurement, the meter was used in automatic equilibrium sensing mode, and two consecutive measurements were made for each sample to ensure that equilibrium had been attained. Results were recorded in millivolts with respect to the calomel electrode, and Eh with respect to the hydrogen electrode were calculated using: Eh (mV H°) = Eh (mV cal.) + (430-S) + 244, where S  =  measured Eh (mV cal.) of +430 mV standard  Conductivity Conductivity was measured using a Brinkmann electrode and Brinkmann 660 Conductometer, or a Radiometer conductivity meter. The Brinkmann meter was calibrated with a commercial standard solution immediately prior to sample measurement, while the Radiometer meter was calibrated internally. The electrodes were occasionally soaked in 10% HC1 to prevent excessive buildup of precipitate. Results were recorded in mS/cm.  192 Total Alkalinity Total alkalinity for samples with pH greater than 4.5 was measured using the 1CM Environmental Methods (1986) procedure. One hundred ml of leachate was titrated with 0.02 N HCI to pH 4.5 endpoint. The volume of 0.02 N HCI consumed was recorded and total alkalinity calculated using: l) = 50,000 N CaCO / Total Alkalinity (mg 3  *  V  vs where:  N= V= V=  normality of HCI (0.02) volume (ml) of HCI required to reach pH 4.5 sample volume (ml)  Acidity Acidity for samples with pH less than 6.0 was measured using the 1CM Environmental Methods Manual (1986). Fifty ml of leachate was boiled with five drops of concentrated hydrogen peroxide for several minutes to oxidize interfering ferrous ions and cooled to room temperature. The solution was titrated with 0.01 N NaOH until pH remains above 8.3 for at least 10 seconds. The volume of NaOH consumed was recorded and acidity calculated using: l) = V*N*5.000 CaCO / Acidity (mg 3 Vs where:  V= N= V=  volume NaOH (ml) normality NaOH volume of sample (ml) (50.00 ml)  Sulfate Sulfate was measured turbimetrically using a slightly amended procedure to the 1CM Environmental Methods Manual (1986) Leachate was diluted, based on its conductivity to give an estimated sulfate concentration between 20 and 50 ppm. The one hundred ml aliquot was added to 20 ml of acetate buffer solution and stirred briefly. Barium chloride was added and the solution stirred for exactly one minute. A portion of the solution was immediately poured into a 1 cm pathlength spectrophotometer cell which was then placed in a Perkin-Elmer Lambda 8 UV/VIS Spectrophotometer. Sample absorbence with respect to an acetate buffer-distilled water reference solution was recorded at 420 nm after four to five minutes. Absorbence was used to determine ppm sulfate from a calibration curve. Absorbence was recalibrated every five readings using the reference solution, and a known standard was usually included in each run.  193 Leachate Quality Analytical Replicates  APPENDIX 8  UNIVERSITY OF BRITISH COLUMBIA DEPT. OF MINING AND MINERAL PROCESS ENGINEERING PARAMETE pH pH units UNIT:  LEACHATE ANALYSIS REPLICATE TESTING Date 02-Oct-92 02-Oct-92 02-Oct-92 02-Oct-92 06-Nov-92 06-Nov-92 06-Nov-92 06-Nov-92 07-Dec-92 07-Dec-92 07-Dec-92 07-Dec-92 07-Dec-92 10-Dec-92 30-Jan-93 30-Jan-93 30-Jan-93 30-Jan-93 20-Feb-93 27-Feb-93 30-Mar-93 15-Apr-93 02-May-93 11-May-93 17-May-93 25-May-93 07-Jun-93 14-Jun-93 12-Aug-93 12-Aug-93 17-Aug-93 17-Aug-93 24-Aug-93 24-Aug-93 30-Aug-93 30-Aug-93 09-Sep-93 09-Sep-93 14-Sep-93 14-Sep-93  # Duplicates Std. Deviation 2X Std. Dev.  RepI.#1  Sample I.D. Week # 1 2 3 4 1 2 3 4 1 2 3 3 4 3 1 2 3 4 4 2 4 2 3 4 2 1 3 4 HC2 HC6 HC4 HC5 HC3 HC7 HCI HC8 HC5 HC6 HC2 HC4  14 20 20 20 20 24 25 29 31 34 35 36 37 39 40 1 1 2 2 3 3 4 4 5 5 6 6  6.919 2.822 7.278 7.421 8.366 2.579 8.081 7.968 8.283 2.391 8.280 7.911 8.484 8.266 7.572 2.305 6.979 7.572 7.762 2.396 7.875 2.277 8.078 7.865 2.236 8.506 8.195 7.884 4.125 2.469 2.918 2.843 4.390 3.345 7.609 7.077 2.826 2.326 4.535 2.672  RepL#2 7.034 2.799 7.089 7.563 8.350 2.563 8.180 7.943 7.794 2.414 7.957 7.849 8.397 8.443 7.898 2.378 7.631 7.628 7.82 2.395 7.91 2.279 7.663 7.834 2.141 8.498 8.170 7.901 4.129 2.471 2.918 2.816 4.408 3.346 7.576 7.134 2.822 2.321 4.557 2.671  IDELTAI  DELTA  %I  0.115 0.023 0.189 0.142 0.016 0.016 0.099 0.025 0.489 0.023 0.323 0.062 0.087 0.177 0.326 0.073 0.652 0.056 0.058 0.001 0.035 0.002 0.415 0.031 0.095 0.008 0.025 0.017 0.004 0.002 0 0.027 0.018 0.001 0.033 0.057 0.004 0.005 0.022 0.001  1.65 0.82 2.63 1.90 0.19 0.62 1.22 0.31 6.08 0.96 3.98 0.79 1.03 2.12 4.21 3.12 8.93 0.74 0.74 0.04 0.44 0.09 5.27 0.39 4.34 0.09 0.31 0.22 0.10 0.08 0.00 0.95 0.41 0.03 0.43 0.80 0.14 0.22 0.48 0.04  40 0.148 0.295  1.98 3.95  194  UNIVERSITY OF BRITISH COLUMBIA DEPT. OF MINING AND MINERAL PROCESS ENGINEERING PARAMETE Conductivity mS/cm UNIT:  LEACHATE ANALYSIS REPLICATE TESTING Date 02-Oct-92 02-Oct-92 02-Oct-92 02-Oct-92 08-Nov-92 08-Nov-92 08-Nov-92 08-Nov-92 09-Dec-92 09-Dec-92 09-Dec-92 09-Dec-92 09-Dec-92 30-Jan-93 30-Jan-93 30-Jan-93 12-Feb-93 19-Feb-93 27-Feb-93 05-Mar-93 25-Mar-93 01-Apr-93 15-Apr-93 10-May-93 11-May-93 17-May-93 25-May-93 07-Jun-93 14-Jun-93 12-Aug-93 12-Aug-93 17-Aug-93 17-Aug-93 24-Aug-93 24-Aug-93 30-Aug-93 30-Aug-93 09-Sep-93 09-Sep-93 14-Sep-93 14-Sep-93  # Duplicates Std. Deviation 2X Std. 0ev.  1 2 3 4 1 2 3 4 1 2 3 3 4 2 3 4 3 4 2 1 3 4 2 3 4 2 1 3 4 HC2 HC6 HC4 HC5 HC3 HC7 HCI HC8 HC5 HC6 HC2 HC4  RepI.#2  RepI.#1  Week #  Sample I.D.  20 20 20 23 24 25 26 28 29 31 34 35 36 37 39 40 1 1 2 2 3 3 4 4 5 5 6 6  0.551 4.350 0.482 2.390 0.218 2.640 0.292 1.640 0.257 3.940 0.210 2.110 0.235 3.93 0.18 2.18 0.23 2.09 3.34 0.3 0.225 1.308 4.49 0.179 0.792 4.330 0.271 0.184 0.539 1.470 3.610 2.180 2.440 0.241 0.755 0.151 0.354 1.890 3.930 0.199 2.520  IDELTAI 0.404 4.310 0.596 2.310 0.234 2.630 0.200 1.610 0.261 3.810 0.207 2.210 0.233 3.95 0.165 2.19 0.216 2.11 3.38 0.296 0.227 1.306 4.48 0.176 0.797 4.360 0.273 0.182 0.556 1.460 3.640 2.180 2.440 0.241 0.760 0.150 0.393 1.900 3.960 0.196 2.310  IDELTA %  0.147 0.040 0.114 0.080 0.016 0.010 0.092 0.030 0.004 0.130 0.003 0.100 0.002 0.020 0.015 0.010 0.014 0.020 0.040 0.004 0.002 0.002 0.010 0.003 0.005 0.030 0.002 0.002 0.017 0.010 0.030 0.000 0.000 0.000 0.005 0.001 0.039 0.010 0.030 0.003 0.210  30.79 0.92 21.15 3.40 7.08 0.38 37.40 1.85 1.54 3.35 1.44 4.63 0.85 0.51 8.70 0.46 6.28 0.95 1.19 1.34 0.88 0.15 0.22 1.69 0.63 0.69 0.74 1.09 3.11 0.68 0.83 0.00 0.00 0.00 0.66 0.66 10.44 0.53 0.76 1.52 8.70  41 0.047 0.095  7.96 15.92  195  UNIVERSITY OF BRITISH COLUMBIA DEPT. OF MINING AND MINERAL PROCESS ENGINEERING PARAMETE Eh mV UNIT:  LEACHATE ANALYSIS REPLICATE TESTING Date 33916 08-Nov-92 08-Nov-92 08-Nov-92 09-Dec-92 09-Dec-92 09-Dec-92 09-Dec-92 30-Jan-93 30-Jan-93 30-Jan-93 30-Jan-93 30-Jan-93 15-Feb-93 20-Feb-93 05-Mar-93 25-Mar-93 01-Apr-93 15-Apr-93 02-May-93 11-May-93 17-May-93 25-May-93 07-Jun-93 14-Jun-93 12-Aug-93 12-Aug-93 17-Aug-93 17-Aug-93 24-Aug-93 24-Aug-93 30-Aug-93 30-Aug-93 09-Sep-93 09-Sep-93 14-Sep-93 14-Sep-93  Sample l.D. Week # 1 2 3 4 1 2 3 4 1 2 3 4 4 3 4 1 3 4 2 3 4 2 1 3 4 HC2 HC6 HC4 HC5 HC3 HC7 HCI HC8 HC5 HC6 HC2 HC4  # Duplicates Std. Deviation 2X Std. Deviaton  Repl.#1 274.0 579.8 288.8 259.0 376.4  20 20 20 20 20 23 24 26 28 29 31 34 35 36 37 39 40 1 1 2 2 3 3 4 4 5 5 6 6  671.0 366.2 358.5 293.3 533.7 315.0 260.4 630.8 575.5 320.6 285.4 313.1 321.4 563.7 326.0 282.6 592.8 503 546 586 593 758 757 770 685 713 548 623 700 806 654 745  Repl.#2 279.6 574.0 275.9 287.4 391.0 613.2 327.0 369.2 306.9 530.4 286.9 274.4 615.3 553.6 319.2 287.6 324.6 278.8 558.3 345.2 287.8 591.5 527 582 584 610 758 762 769 653 720 569 682 701 806 638 745  IDELTA1 IDELTA %I 2.02 5.6 5.8 1.01 12.9 4.57 28.4 10.40 14.6 3.81 57.8 9.00 39.2 11.31 10.7 2.94 13.6 4.53 3.3 0.62 28.1 9.34 5.24 14.0 2.49 15.5 21.9 3.88 1.4 0.44 0.77 2.2 11.5 3.61 14.20 42.6 5.4 0.96 5.72 19.2 5.2 1.82 0.22 1.3 4.66 24.0 6.38 36.0 0.34 2.0 2.83 17.0 0.00 0.0 0.66 5.0 0.13 1.0 4.78 32.0 0.98 7.0 3.76 21.0 9.04 59.0 0.14 1.0 0.00 0.0 2.48 16.0 0.00 0.0  37 15.7 31.3  3.65 7.30  196  UNIVERSITY OF BRITISH COLUMBIA DEPT. OF MINING AND MINERAL PROCESS ENGINEERING PARAMETE Acidity UNIT: mg CaCO3/I  LEACHATE ANALYSIS REPLICATE TESTING Date 10-Oct-92 18-Nov-92 04-Jan-93 11-Jan-93 08-Feb-93 27-Feb-93 16-Apr-93 17-May-93 12-Aug-93 17-Aug-93 24-Aug-93 24-Aug-93 30-Aug-93 09-Sep-93 09-Sep-93 14-Sep-93 14-Sep-93  # Duplicates Std. Deviation 2X Std. Dev.  Sample I.D. Week # 2 2 2 2 2 2 2 2 HC6 HC4 HC3 HC7 HC8 HC5 HC6 HC2 HC4  17 18 22 25 31 36 1 2 3 3 4 5 5 6 6  Repl.#1 2058 1852 1840 2440 2285 1645 2680 2090 1465 835 10 154 0 645 2430 20 610  Repl.#2 2040 1813 1910 2510 2384 1570 2605 2110 1510 915 10 144 9 620 2430 10 1080  IDELTAI  18 39 70 70 99 75 75 20 45 80 0 10 9 25 0 10 470  17 109 218  IDELTA  %I  0.88 2.13 3.73 2.83 4.24 4.67 2.84 0.95 3.03 9.14 0.00 6.71  3.95 0.00 66.67 55.62  20.03 40.05  197  UNIVERSITY OF BRITISH COLUMBIA DEPT. OF MINING AND MINERAL PROCESS ENGINEERING PARAMETER Alkalinfty mg CaCO3/I UNIT:  LEACHATE ANALYSIS REPLICATE TESTING Date  Sample I.D.  10-Oct-92 10-Oct-92 10-Oct-92 18-Nov-92 18-Nov-92 18-Nov-92 04-Jan-93 04-Jan-93 04-Jan-93 08-Feb-93 08-Feb-93 08-Feb-93 08-Feb-93 05-Mar-93 25-Mar-93 05-Apr-93 01-May-93 11-May-93 25-May-93 07-Jun-93 14-Jun-93 12-Aug-93 HC8 17-Aug-93 HC8 30-Aug-93 HCI  # Duplicates Std. Deviation  RepI.#1  Week # 1 3 4 1 3 4 1 3 4 1 3 3 4 1 3 4 3 4 1 3 4  17 17 17 22 22 22 22 26 28 29 34 35 37 39 40 1 2 4  RepI.#2 105.5 99.5 193.0 98.0 76.0 150.5 88.4 55.5 105.0 87.5 49.0 108.5 128.4 89.5 53.5 91.0 42.0 93.0 85.5 42.5 108.5 12 9 21  IDELTA %I  IDELTAI 107.0 94.0 196.0 107.6 76.0 149.5 88.5 56.5 108.0 88.5 50.5 108.5 132.0 90.0 53.5 89.5 42.5 94.0 87 42 107.5 13 9 19.5  1.5 5.5 3 9.6 0 1 0.1 1 3 1 1.5 0 3.6 0.5 0 1.5 0.5 1 1.5 0.5 1 1 0 1.5  1.41 5.68 1.54 9.34 0.00 0.67 0.11 1.79 2.82 1.14 3.02 0.00 2.76 0.56 0.00 1.66 1.18 1.07 1.74 1.18 0.93 8.00 0.00 7.41  24 2.1 4.3  2.65 5.30  198  UNIVERSITY OF BRITISH COLUMBIA DEPT. OF MINING AND MINERAL PROCESS ENGINEERING PARAMETER Sulfate mg/I UNIT:  LEACHATE ANALYSIS REPLICATE TESTING  0 Date 29-Sep-92 29-Sep-92 16-Oct-92 16-Oct-92 16-Oct-92 16-Oct-92 25-Nov-92 25-Nov-92 25-Nov-92 25-Nov-92 15-Dec-92 15-Dec-92 11-Jan-93 11-Jan-93 11-Jan-93 21-Feb-93 27-Feb-93 09-Mar-93 17-May-93 05-Apr-93 20-Apr-93 10-May-93 11-May-93 17-May-93 25-May-93 15-Jun-93 15-Jun-93 17-Aug-93 17-Aug-93 17-Aug-93 17-Aug-93 24-Aug-93 24-Aug-93 30-Aug-93 30-Aug-93 14-Sep-93 14-Sep-93 17-Sep-93 17-Sep-93  # Duplicates Std. Deviation 2X Std. 0ev.  RepI.#1  Sample I.D. Week # 2 4 1 2 3 4 1 2 3 4 4 4 1 3 4 4 2 1 2 4 2 3 4 2 1 3 4 HC2 HC6 HC4 HC5 HC3 HC7 HC1 HC8 HC2 HC4 HC5 HC6  14 14 18 18 18 24 25 26 36 29 31 34 35 36 37 39 40 1 1 2 2 3 3 4 4 6 6 5 5  5750 1650 100 3000 142 1600 113 4180 96 1725 1600 70 3800 44 1550 1450 2550 64 62 750 3600 46 330 3050 65 50 195 870 2800 1450 1650 80 350 48 178 74 1450 1070 2900  RepI.#2 7000 1625 107 2250 146 1575 105 4120 106 1760 1625 79 4050 58 1575 1475 2450 64 59 750 3900 42 320 2800 60 48 160 965 2800 1600 1700 100 350 48 178 80 1425 1110 3200  IDELTAI 1250 25 7 750 4 25 8 60 10 35 25 9 250 14 25 25 100 0 3 0 300 4 10 250 5 2 35 95 0 150 50 20 0 0 0 6 25 40 300  IDELTA  %I  19.61 1.53 6.76 28.57 2.78 1.57 6.90 1.45 9.90 2.01 1.55 12.08 6.37 27.45 1.60 1.71 4.00 0.00 4.96 0.00 8.00 9.09 3.08 8.55 8.00 4.08 19.72 10.35 0.00 9.84 2.99 22.22 0.00 0.00 0.00 7.79 1.74 3.67 9.84  39 234.7006082 7.45678523 469.4012165 14.91357046  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  Week  Column i  .  65.5  65.3  65.3  65.0  64.4 64.9 64.6 85.0 64.9 64.9 64.8 64.9 65.0 64.7 64.9 65.3 65.0 65.0 64.9 65.1 65.0 64.9 65.0 65.0 65.1 64.9 64.9 64.7 65.0 64.9  Cal. Mass (kg) 5.10 5.95 5.53 6.17 6.04 6.04 5.85 6.04 6.23 5.67 6.04 6.79 6.23 6.23 5.97 6.40 6.19 5.97 6.19 6.19 6.40 5.95 5.95 5.74 6.17 5.95  0.57 0.53 0.34 1.11 0.58 0.55 0.34 0.92 0.56 0.56 0.36 0.95 0.54 0.59 0.35 1.01 0.49 0.57 0.38 0.90 0.60 0.52 0.35 1.00 0.46 0.50 0.21 0.76 0.54 0.69 0.37 0.84 0.50 0.58 0.37 1.25 0.48 0.51 0.40 1.00  2.098 4.975 3.370 10.785 5.765 5.380 3.415 8.940 5.450 5.420 3.570 8.990 5.570 5.400 3.424 8,164 4.840 5.270 3.323 8.920 5.920 4.770 3.180 9.340 3.770 5.750 3.094 7.776 4.570 6.420 3.640 9.350 5.070 5.180 3.220 10.540 4.020 4.920 4.030 9.400  5.625 5.325 3.475 11.175 5.850 5.500 3.450 9.300 5.625 5.700 3.700 9.325 5.550 5.750 3.550 8.750 5.750 5.825 3.800 9.125 6,250 5.075 3.500 10.200 4.575 5.075 2.100 7.800 4.650 7.000 3.725 9.600 5.050 5.925 3.750 11.100 4.650 5.125 4.050 10.225 7.22  6.81  6.81  6,16  Moisture Cont. (wt.%)  Average Flow (mI/mm)  Vol. Leachate (I)  Vol. Pumped (I) 7.64 7.46 7.05 6.92 6.91 8.04 8.02 8.13 8.37 7.90 8.26 7.82 7.79 8.49 8.64 8.64 8.67 8.05 7.98 7.90 8.15 7.99 8.02 8.07 7.89 8.00 8.09 8.16 8.12 8.15 7.86 7.87 8.21 8.33 8.06 8.36 8.51 7.97 7.98 8.17  pH  453 507 473 512 518 491 507 640 620 601 504 584619 556 511 537 544 586 580 544 529 507 545 593 489 489 492 560 541 513 552 504 527 413 420 510  Eh (mV H) 2.68 1.04 0.49 0.55 0.46 0.36 0.50 0.46 0.22 0.31 0.32 0.30 0.26 0.31 0.32 0.28 0.21 0.25 0.27 0.24 0.25 0.28 0.30 0.29 0.28 0.30 0.31 0.33 0.28 0.30 0.29 0.29 0.252 0.303 0.304 0.311 0.271 0.29 0.306 0.287  Cond. (mS/cm)  70 99 120 106 112 97 109 109 98 98 108 104 102 91 112 88 93 86 93 101 88 83 94 93 90 69 87 96 90 79 87 84 90 76 97 85.5 84 82 96  Acidity Alkalinity (mg CaCO3/l) (mg CaCO3/i)  General Parameters  1650 600 250 80 50 100 109 98 97 102 109 71 60 74 84 70 62 75 73 58 46 66 66 62 68 64 63 62 56 64 72 62 49 66 76 48 65 72 85 47  Sulphate (mg/I)  CD  E  a.  ()  0  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  Week  Column 2  60.8  61.0  61.5  61.0  60.3 60.3 60.3 61.0 60.7 60.9 60.0 60.9 60.8 60.8 60.8 61.2 60.9 61.0 60.8 61,0 60.8 60.8 60.8 60.9 61.0 60.8 60.7 60.8 60.8 61.0  Col. Mass (kg)  ,  6.025 5.725 3.250 11.000 5.650 4.875 3.250 9.350 5.370 5.825 3.900 9.675 5.550 5.850 3.750 8.825 5.900 5.525 3.600 11.400 6.725 4.850 3.550 12.525 5.700 4.275 3.400 8.500 4.600 6.800 3.575 10.200 6.250 5.950 3.800 9.525 5.925 5.700 4.000 10.875  Vol. Pumped (ml) 9.50 9.50 9.50 10.93 10.31 10.73 8.84 10.73 10.52 10.52 10.52 11.38 10.73 10.94 10.47 10.94 10.47 10.47 10.47 10.71 10.96 10.45 10.22 10.45 10.45 10.94  0.62 0.57 0.32 1.10 0.56 0.48 0.32 0.93 0.53 0.58 0.38 0.98 0.54 0.60 0.37 1.02 0.50 0.54 0.36 1.13 0.64 0.50 0.35 1.23 0.57 0.42 0.34 0.83 0.54 0.67 0.36 0.89 0.61 0.58 0.38 1.07 0.61 0.57 0.39 1.07  0.570 5.365 3.235 10.040 5.520 4.260 3.250 8.830 5.550 5.470 3.720 9.240 5.420 5.200 3.526 8.366 5.160 4.870 3.350 11.120 6.520 4.670 3.178 12.040 5.900 5.700 3.179 8.586 4.620 6.220 3.770 9.620 5.620 5.520 3.220 9.490 5.920 5.620 4.030 10.530 10.45  10.94  11.88  10.94  Moisture Cont. (wt.%)  Average Flow (mI/mm)  Vol. Leachate (ml) 2.20 2.53 2.68 2.82 2.70 2.81 2.59 2.49 2.58 2.54 2.35 2.48 2.41 2.38 2.26 2.24 2.20 2.37 2.37 2.38 2.37 2.37 2.30 2.39 2.40 2.36 2.14 2.15 2.25 2.29 2.28 2.32 2.23 2.22 2.26 2.24 2.22 2.27 2.13 2.35  pH  783 800 793 811 821 814 824 921 915 809 829 826 829 841 820 778 809 833 841 829 839 838 767 845 842 814 808 830 833 860 846 838 852 858 858 854  Eh (my H) 16.77 9.10 5.11 4.35 4.08 4.03 3.93 3.88 2.64 3.44 3.90 3.72 3.94 4.24 4.43 3.72 3.73 3.72 4.50 3.93 3.34 3.88 4.43 3.75 3.34 4.04 4.79 4.39 4.09 4.37 4.49 4.13 3.85 4.09 4.61 4.33 3.63 3.81 4.11 3.5  Cond, (mS/cm)  26000 5200 4900 2040 1980 2230 2020 1690 1850 2290 2090 1770 2000 2260 2230 1840 2470 2830 2380 1780 2280 3090 2290 1610 2500 3034 2975 2209 2400 2680 2160 1770 2180 2770 2090 1535 1906 2280 1700  Alkalinity Acidity (mg CaCO3/l) (mg CaCO3/l)  General Parameters  32000 11500 6400 3750 3200 3000 3425 3525 3600 3525 4150 3740 3550 3900 4350 3525 3275 3925 4450 3200 2550 3400 4500 3200 2550 3250 4200 4300 2950 3400 3750 3100 2700 2900 3650 3050 2300 2500 3350 2250  Sulphate (mg/I)  16-Nov-93  C  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  Week  Column 3  64.5  64.6  64.4  64.4  63.5 63.7 63.5 63.9 63.6 64.1 63.9 63.7 63.8 63.8 64.0 64.2 64.0 64.0 64.0 64.2 64.0 64.0 64.2 64.2 642 64.2 64,1 64.2 64.0 64.2  Col. Mass (kg)  .  Vol. Leachate (ml)  1.240 5.010 3.330 10.410 4.930 4.840 3.980 9.140 4.650 5.620 3.720 9.490 5.770 5.700 3.828 8.261 4.080 4.770 3.360 10.540 6.070 4.320 3.055 11.590 4.650 5.500 2.937 7.094 3.820 6.270 3.280 6.909 5.620 6.220 3.320 10.490 6.070 5.820 4.260 9.690  Vol. Pumped (ml)  5.800 5.400 3.375 10.900 5.100 5.050 4.050 9.425 4.800 5.800 3.750 9.475 5.925 6.050 3.800 8.700 5.450 5.100 3.650 10.575 6.125 4.625 3.450 12.400 4.600 4.775 3.085 7.525 3.950 6.750 3.325 8.025 6.100 6.300 3.875 10.050 6.175 6.100 4.300 10.150  0.59 0.54 0.33 1.09 0.51 0.50 0.40 0.93 0.47 0.57 0.36 0.96 0.57 0.62 0.38 1.00 0.46 0.50 0.36 1.04 0.58 0.48 0.34 1.22 0.46 0.47 0.31 0.73 0.46 0.66 0.33 0.70 0.60 0.62 0.39 1.13 0.64 0.61 0.42 1.00  (mi/mm)  Average Flow  15.06  15.28  14.83  14.83  12.97 13.44 12.97 13.90 13.20 14.22 13.81 13.40 13.60 13.60 14.01 14.42 14.01 14.01 13.93 14.38 13.93 13.93 14.38 14.38 14.38 14.36 14.14 14.36 13.91 14.36  Moisture Cont. (wt.%) 8.14 7.27 7.30 7.28 6.93 8.20 8.09 8.27 8.08 8.13 8.58 7.83 7.96 8.27 8.63 8.33 8.76 8.18 8.03 7.63 7.84 7.85 7.85 8.07 7.96 7.83 7.97 8.06 8.15 7.89 7.91 7.91 8.17 8.08 7.96 7.77 8.18 7.84 8.2 7.81  pH  453 528 497 525 533 484 526 609 610 598 552 566 553 553 536 559 586 556 615 581 639 545 518 557 576 507 496 558 567 571 545 545 563 562 546 594  Eh (my H) 2.56 2.12 0.65 0.48 0.63 0.43 0.37 0.34 0.29 0.21 0.26 0.26 0.21 0.20 0.31 0.08 0.19 0.18 0.20 0.18 0.15 0.20 0.23 0.17 0.18 0.24 0.24 0.23 0.21 0.22 0.22 0.20 0.191 0.179 0.204 0.172 0.156 0.17 0.184 0.158  Cond. (mS/cm)  48 80 100 100 98 87 87 88 76 86 68 59 59 60 65 56 57 59 48 45 49 57 48 47 48 55 54 54 50 50 47 47 40.5 44.3 37 35 38.5 42 39  Alkalinity Acidity (mg CaCO3/l) (mg CaCO3/l)  General Parameters  1800 1560 325 125 100 140 139 100 104 98 101 62 49 50 55 50 44 51 52 36 33 47 66 44 45 49 63 62 56 56 62 39 46 42 50 55 40 42 50 32  Sulphate (mg/I)  16-Nov-93  98 114.5 93 115.5 101 107  2.18 2.16 2.09 2.04 1.92 1.81 1.50 1.31 1.22 1.23 0.93 0.78 0.76 0.79 0.75 0.59 0.58 0.63 0.54 0.46 0.61 0.60  597 578 505 547 639 518 550 544 496 484  540 557 544 528 526 539 558  586 548 536 525  7.89 7.83 7.76 7.82 7.75  7.88 7.90 7.89 7.64 7.96  7.30 7.30 7.30 7.30  44  41 42 43  40  38 39  37  36  35  33 34  31 32  30  28 29  61.7  61.7  61.6  61.5  61.2  61.2  26 27  0.32 0.83 0.60 0.58 0.37 1.18  3.130 9.254  5.670 5.670  3.170  3.175  0.39  1.06 0.69 0.54 0.30 0.90  3.940  5.120 2.965  8.685  5.400 3.000  9.125  0.60  5.620  10.450 6.420  0.62  9.840 5.920  3.975 10.800 6.750  6.000 5.975  10.475  6.125 5.950 3.750  9.500  0.62  4.220 5.620  4.450 6.275  0.74 0.52  0.32  7.951  7.550  0.35 1.18 0.57 0.44  5.600 3.094  4.820 3.190 11.840 5.950  1.10 0.59 0.55  5.725 4.475 3.235  3.525 12.025  61.0 61.2 61.2 61.2  22 23 24 25  11.750 5.920  0.37  0.49 0.53  8.23  8.23  8.05  7.77  7.30  6.88 7.32  6.88  7.34 7.06  7.90  7.78  8.09 7.88 8.00 8.04  8.24 7.78  7.82  8.11 7.86  8.10 7.80  551  0.58  210 230  84 98  2.19 2.10  504 538  7.63  7.34 6.66  6.225 5.300  61.2  20 21  4.970 3.327  4.760  558  200  107  2.20 2.33  554 511  7.91 7.85  3.725 11.150  18 19  5.425  8.950 5.700  61.0 61.2  61.2  61.1 61.0  17  7.91  2.22  552  8.43  .  2.39 2.44 2.08  532 493 554  0.56 0.61 0.37 1.03  235 195 140 195 265 210  99.5 79 108  420 330 260 330  625 690  750  1150 900  1350 900  1450  1500  1500 1600  1500  1560 1650  1720 1760  1625 1650  1640 1600  1740  1625 1550  1525  1600  1600  88 108.5 99  111 91  104 72  108 98 129 96  107  lii 122  105 116  137 133 145  133  2.11  603  8.18 8.50 8.31  5.750 3.806 8.164  5.420 7.26 7.10  14 15 16  5.945 3.775  61.4 61.2 61.1  13  181  2.08  7.78 7.85  7.67 7.67  620  8.04  7.06  9.200 5.825  61.5  0.94  3.670 8.990  3.675  61.1  12  151 123  1.88 2.04  499 557  11  7.26  0.54 0.36  5.320  5.500  174  1.64  7.97 7.44  7.06  9 10  0.57  527 503  5.500  61.4  8  5.775  495  7.88  1600  61.1 61.2  61.4 61.0  6 7 7.83  61.2  4 5  7.67  154 183  2.20 2.19  529  1650 1575  1750 1900  (mg/I)  Sulphate  0.95  193  2.30  456  7.67 6.86  0.54 0.32  194 216 193  7.54 7.26  1.09 0.46  2.39 2.40  7.62  6.85  0.33  3.185  10.695  2.11  7.84  6.39  0.63 7.42 7.34 7.90  240  2.47  8.02  6.39  0.61  2.62  Acidity (mg CaCO3I1)  Alkalinity (mg CaCO3/l)  Cond. (ms/cm)  Eh (my H)  1.965  pH  5.210  (wt.%)  Cont,  (mi/mm)  3.230 9.340  61.0 61.3  3  6.000 6.325  (ml)  flow  3.250 9.600  60.8  (ml)  Moisture  Average  4.900 5.280  60.8  I  2  (kg)  Vol.  Leachate  Vol.  Pumped  Cd.  Mass  General Parameters  3.375 10.925 4.575 5.475  Week  Column 4  15-Nov-93  0 t’J  Week 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  0.053 0.028 0.033 0.042 0.030 0.030 0.027 0.027 0.036 0.021 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  0.0008 0.0007 0.0007 0.0006 0.0006 0.0006 0.0006 0.0005 0.0005 0.0005 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 0.044 0.049 0.045 0.045 0.044 0.046 0.049 0.039 0.039 0.049 0.053 0.044 0.05 0.042 0.058 0.05 0.046 0.046 0.046 0.044 0.045 0.043 0.05 0.039 0.036 0.05 0.044 0.038  <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  Ag (mg/I) Al (mg/I) As (mg/I) Ba (mg/I) Be (mg/I) 0.0013 0.26  Column 1  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  0.0059 0.0039 0.0029 0.0026 0.0029 0.0027 0.0042 0.0018 0.0015 0.0015 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  0.53 0.39 0.36 0.27 0.36 0.28 0.31 0.21 0.23 0.19 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 0.001 0.001 <0.0002 0.0002 <0.0002 <0.0002 0.0006 <0.0002 <0.0002 <0.0002 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  0.0074 0.0061 0.0052 0.0046 0.0046 0.0061 0.0046 0.0040 0.0033 0.0070 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 0.032 <0.030 <0.030 0.04 <0.030 <0.030 98 65 62 67 68 66 64 63 64 62 55.9 65.2 56.1 60.2 51.4 60.9 57.6 49.6 48.2 55.1 56.4 58.4 55.7 51.6 52.1 57.9 51.5 55.8 50 53 43.5 53.6 51.8 48.1 43.4 51.5 52.1 49.5  K (mg/I) 1.8  BI (mg/I) Ca (mg/I) Cd (mg/I) Co (mg/I) Cr (mg/I) Cu (mg/I) Fe (mg/I) 0.019 0.015 610 0.0022  Dissolved Metals  0.019 0.019 <0.015 0.017 0.015 0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  Li (mg/I)  2.8 1.8 1.8 2.0 2.1 2.0 1.8 1.9 2.0 1.8 1.7 1.91 1.65 1.77 1.61 1.75 1.76 1.45 1.43 1.66 1.77 1.6 1.74 1.53 1.96 1.73 1.62 1.66 1.6 1.52 1.49 1.57 1.66 1.34 1.26 1.64 1.64 1.38  0.020 0.017 0.0061 0.0037 0.0016 0.0018 0.0022 0.0014 0.0013 0.0014 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  0.045 0.043 0.042 0.041 0.056 0.048 0.045 0.049 0.049 0.052 0.052 0.048 0.053 0.048 0.058 0.047 0.049 0.049 0.049 0.051 0.036 0.036 0.034 0.036 0.036 0.037 0.036 0.037  27.0 19.0 17.0 15.0 14.0 13.0 11.0 10.0 10.0 8.1 7.9 8.2 6.9 7.3 6.6 6.5 6.8 5.2 5.3 6 6.4 5.3 6.2 5 7.3 5.5 5.5 5.5 5.3 4.8 5 4.7 5.1 4.1 5.7 5 4.9 3.9  Mg (mg/I) Mn (mg/I) Mo (mg/I) Na (mg/I) 0.060 15 83  C  16-Nov-93  Ni (mg/I)  <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020  Week  2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  Column 1  <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30  P (mg/I)  <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30  0.628 0.713 0.61 0.69 0.599 0.671 0.679 0.551 0.546 0.636 0.664 0.613 0.636 0.569 0.699 0.636 0.618 0.635 0.592 0.593 0.583 0.598 0.621 0.529 0.478 0.629 0.619 0.543  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  Pb (mg/I) Sb (mg/I) Se (mg/I) Sn (mg/I) Sr (mg/I) Th (mg/I)  <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  Ti (mg/I)  <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030  V (mg/I)  Dissolved Metals  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  0.076 0.061 0.015 0.015 0.016 0.015 0.04 1 0.0071 0.0030 0.0063 <0.005 0.016 0.011 0.016 <0.005 0.013 0.005 <0.005 <0.005 0.008 0.006 0.008 0.013 0.006 <0.005 <0.005 <0.005 0.009 <0.005 <0.005 <0.005 0.008 <0.005 0.03 0.045 <0.005 <0.005 0.038  W (mg/I) Zn (mg/I) 0.074  I’J  1 6-Nov-93  Week 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 0.018 <0.015 <0.015 0.015 0.025 <0015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  760 410 250 240 280 240 200 200 240 220 197 223 219 219 188 258 295 203 149 210 252 185 137 170 222 210 173 195 188 165 136 165 198 147 104 122 150 112  0.023 0.013 0.0012 0.0012 0.0019 0.0019 0.0016 0.0024 0.0035 0.0033 <0.20 0.22 0.27 0.23 0.21 0.26 0.29 0.22 <0.20 <0.20 0.23 0.24 <0.20 0.24 0.26 0.3 0.23 0.25 0.27 0.23 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  Ag (mg/I) Al (mgJI) As (mg/I) Ba (mg/I) Be (mg/I) 0.22 3800  Column 2  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 0.1 0.11 0.15 0.14 <0.10 <0.10 0.11 <0.10  430 510 540 550 560 520 550 560 570 510 484 470 400 379 377 379 387 295 261 299 312 263 231 233 253 238 217 221 209 197 155 174 191 155 133 142 149 130  0.25 0.11 0.15 0.14 0.16 0.14 0.11 0.13 0.15 0.13 0.108 0.117 0.115 0.106 0.103 0.131 0.153 0.114 0.093 0.111 0.14 0.106 0.079 0.096 0.12 0.124 0.104 0.108 0.107 0.094 0.086 0.096 0.111 0.093 0.065 0.065 0.077 0.07 0.543 0.589 0.58 0.571 0.499 0.658 0.747 0.55 0.413 0.573 0.669 0.531 0.419 0.508 0.668 0.643 0.569 0.627 0.596 0.572 0.452 0.545 0.627 0.511 0.394 0.435 0.513 0.409  0.022 0.024 0.032 0.028 0.019 0.032 0.03 0.029 0.022 0.034 0.045 0.035 0.028 0.035 0.049 0.049 0.035 0.046 0.046 0.042 0.034 0.039 0.051 0.031 0.027 0.035 0.045 0.032  14 7.6 3.5 3 3.5 3.2 2.7 2.9 3.3 2.8 2.64 2.88 2.68 2.63 2.27 2.97 3.38 2.3 1.81 2.48 2.87 2.19 1.82 2.06 2.69 2.52 2.23 2.42 2.3 2.12 1.79 2.06 2.41 1.89 1.51 1.74 2.03 1.59  110 77 63 70 99 110 100 130 180 180 175 218 243 249 222 322 404 291 223 374 507 311 234 336 506 475 376 453 453 400 306 374 480 352 236 297 372 263  Bi (mg/I) Ca (mg/I) Cd (mg/I) Co (mg/I) Cr (mg/I) Cu (mg/I) Fe (mg/I) 1700 67 82 3.1  Dissolved Metals  0.26 0.18 0.32 0.23 0.17 0.11 0.12 0.11 0.12 0.1 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0  K (mg/I) 1.4  0.051 0.057 0.054 0.054 0.046 0.059 0.068 0.049 0.037 0.062 0.077 0.065 0.046 0.056 0.077 0.081 0.066 0.075 0.074 0.067 0.059 0.068 0.082 0.074 0.055 0.069 0.084 0.069  153 66 34 31 37 35 28 29 33 29 26.5 27.7 25.8 26 23.1 28.7 32.2 24.2 19.3 26.6 33.2 31.8 26.7 29.3 38 40.2 38.8 41.3 41 39.7 344 38.9 46 39.3 33.5 37 43.3 38.8  33 13 7.6 7 8.3 7.7 6.1 6.5 7.2 6.5 5.64 5.93 5.78 5.77 5.11 6.41 7.2 5.32 4.17 5.7 7.2 6.53 5.28 5.89 7.8 8.21 7.77 8.23 8.18 7.9 6.69 7.54 8.94 7.78 6.52 7.2 8.41 7.57  <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030  15 11 10 10 11 10 9.3 9.3 10 8.1 8.4 8.5 7.7 7.6 7.4 8.1 9 6.7 6.2 7.7 8.4 7.4 7.2 6.9 8.8 8 8.2 8.2 8 7.7 7.2 7.1 8.1 7.2 7 7.2 8 7  Li (mg/I) Mg (mg/I) Mn (mg/I) Mo (mg/I) Na (mg/I) 690 200 46  c,I  1 6-Nov-93  Week 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 <0.050  <0.050 <0.050  0.74 0.795 0.671 0.678 0.639 0.711 0.766 0.558 0.479 0.602 0.638 0.537 0.486 0.52 0.6 0.563 0.555 0.584 0.55 0.552 0.507 0.537 0.614 0.506 0.441 0.487 0.53 0.468 <0.030 <0.030 <0.030 <0.030 <0.030  <0.030 <0030  <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030  <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30  <0.010 <0.010 0.011 0.014 0.011 0.013 0.014 0.016 0.012 0.015 0.019 0.014 0.011 0.016 0.02 0.019 0.015 0.02 0.021 0.017 0.015 0.018 0.026 0.019 0.017 0.019 0.023 0.019  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 0.61 0.66 1.03 0.85 0.53 0.62 0.75 0.49  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  0.42 0.49 0.67 0.74 0.51 0.82 1.16 0.77 0.35 1.01 1.69 1.22 0.53 1.13 1.93 2 1.29 1.72 1.79 1.45 0.83 1.19 1.93 1.3 0.48 0.73 1.06 0.77  0.617 0.687 0.673 0.655 0.569 0.73 0.846 0.593 0.452 0.63 0.785 0.576 0.481 0.558 0.762 0.724 0.641 0.712 0.684 0.638 0.53 0.612 0.707 0.575 0.451 0.486 0.577 0.447  <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050  V (mg/I)  Ti (mg/I)  Pb (mg/I) Sb (mg/I) Se (mg/I) Sn (mg/I) Sr (mg/I) Th (mg/I)  P (mg/I)  Dissolved Metals  Ni (mg/I)  Column 2  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  110 49 23 21 24 21 17 19 22 19 16.2 17.4 18.3 17.6 15.4 20.3 23.1 16.9 12.6 17.5 21.5 16.7 12.1 14.4 19 18.6 15.6 16.9 16.5 15.1 12 13.2 15.4 12.6 9.61 10.2 11.8 9.32  W (mg/I) Zn (mg/I) 610  1 6-Nov-93  Week 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  0.093 0.017 0.024 0.042 0.018 0.015 0.03 0.012 0.012 0.012 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  0.0001 0.0001 0.0003 0.0002 0.0002 0.0002 0.0002 0.0002 0.0002 0.0002 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 0.022 0.02 0.024 0.016 0.019 0.025 0.023 0.015 0.015 0.021 0.029 0.023 0.023 0.023 0.027 0.031 0.029 0.027 0.031 0.028 0.025 0.023 0.027 0.019 0.017 0.018 0.024 0.017  <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  Ag (mg/I) Al (mg/I) As (mg/I) Ba (mg/I) Be (mg/I) 0.0002 0.174  Column 3  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0,015 <0.015  0.51 0.33 0.29 0.31 0.38 0.21 0.32 0.28 0.37 0.23 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 0.034 0.021 0.015 0.013 0.012 0.012 0.0091 0.0082 0.0045 0.0034 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  0.0092 0.0061 0.0082 0.0061 0.0082 0.0052 0.015 0.0048 0.0024 0.039 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 0.048 <0,030 0.04 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 0.062 0.166 <0.030 <0.030 0.064 0.044 0.042 <0.030 0.0039 150 0.0015 86 0.0008 75 0.0008 85 83 0.0009 65 0.0005 0.0004 56 0.0005 54 55 <0.0002 0.0002 43 <0.010 38.3 <0.010 36.2 <0.010 38 <0.010 31.8 <0.010 34.7 39.3 <0.010 <0.010 38.2 <0.010 28.7 <0.010 25.5 <0.010 33.7 <0.010 40 31.2 <0010 31.7 <0.010 <0.010 31.7 <0.010 37.4 <0.010 37.8 <0.010 35.9 <0.010 36.3 <0.010 34.7 <0.010 34.1 <0.010 28.9 <0.010 29.2 <0.010 30.7 <0.010 23.8 <0.010 22.8 <0.010 26.1 <0.010 28.8 <0.010 24,3  K (mg/I) 2.4  Bi (mg/I) Ca (mg/I) Cd (mg/I) Co (mg/I) Cr (mg/I) Cu (mg/I) Fe (mg/I) 0.013 0.038 0.0055 660  Dissolved Metals  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  Li (mg/I)  3.8 2.1 1.9 2.1 2.1 1.5 1.4 1.5 1.8 1.2 1.06 0.972 1.05 0.784 0.945 1.07 1.06 0.674 0.646 0.948 1.27 0.859 0.918 0.891 1.16 1.15 1.04 1.04 1.1 0.934 0.829 0.788 0.886 0.605 0.62 0.738 0.872 0.686  0.25 0.049 0.015 0.0088 0.0058 0.0037 0.0041 0.0029 0.002 0.0017 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0,005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  0.094 0.093 0.102 0.085 0.106 0.104 0.109 0.088 0.068 0.093 0.099 0.085 0.087 0.085 0.092 0.088 0.09 0.086 0.088 0.083 0.067 0.065 0.07 0.037 0.037 0.038 0.051 0.038  15 8.9 8.6 8.3 8.1 5.9 5.9 5.9 7.1 4,9 5.2 4.8 5.2 3.8 5.1 5.1 5.2 3.3 3.5 4.9 6.5 4.1 5.3 4.5 6.7 5.6 5.8 5.5 6.1 4.9 4.6 4 5.1 3.3 3.8 4.2 4.9 3.6  Mg (mg/I) Mn (mg/I) Mo (mg/I) Na (mg/I) 0.89 84 20  16-Nov-93  Ni (mg/I)  <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0:020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020  Week  2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  Column 3  <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30  P (mg/I)  <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 ..  <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30  0.155 0.146 0.146 0.125 0.138 0.157 0.151 0.11 0.101 0.131 0.156 0.12 0.124 0.12 0.143 0.143 0.138 0.14 0.131 0.134 0.12 0.115 0.126 0.1 0.096 0.103 0.114 0.095  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  Pb (mg/I) Sb (mg/I) Se (mg/I) Sn (mg/I) Sr (mg/I) Th (mg/I)  <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  Ti (mg/I)  <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030  V (mg/I)  Dissolved Metals  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  0.32 0.11 0.066 0.072 0.067 0.062 0.04 0.035 0.032 0.035 0.037 0.029 0.024 0.029 0.021 0.03 0.024 0.032 0.024 0.024 0.02 1 0.034 0.027 0.03 0.025 0.041 0.037 0.045 0.037 0.048 0.047 0.051 0.042 0.051 0.045 0.05 0.043 0.047  W (mg/I) Zn (mg/I) 1.4  00  16-Nov-93  Week 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44  <0.015 <0.015 <0.015 <0.015 <0.015 0.029 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015. <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  0.49 0.5 0.042 0.006 0.015 0.086 0.039 0.048 0.009 0.018 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  0.0002 0.0002 0.0003 0.0003 0.0002 0.0002 0.0002 0.0002 0.0001 0.0001 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 0.029 0.026 0.035 0.028 0.038 0.035 0.039 0.025 0.039 0.047 0.06 0.037 0.044 0.046 0.062 0.048 0.045 0.052 0.061 0.049 0.058 0.059 0.075 0.059 0.069 0.086 0.086 0.062 0.062 0.095 0.112 0.088  <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  Ag (mg/I) Al (mg/I) As (mg/I) Ba (mg/I) Be (mg/I) 0.0003 0.36  Column 4  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  0.022 0.057 0.024 0.022 0.016 0.027 0.023 0.009 0.0071 0.012 0.014 0.014 0.012 0.018 <0.010 0.016 <0.010 <0.010 <0.010 <0.010 <0.010 0.013 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  1.4 0.99 0.77 0.7 0.72 0.62 0.65 0.56 0.61 0.5 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 <2.0 0.0007 0.0036 0.0006 0.0002 <00002 <0.0002 0.0003 <0.0002 <0.0002 <0.0002 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  0.015 0.015 0.01 0.01 0.0085 0.0061 0.003 0.0061 0.003 0.0061 <0.030 <0.030 <0.030 <0.030 <0.030 0.037 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 630 650 650 650 640 640 640 640 640 640 658 617 646 612 666 686 646 665 608 640 638 623 604 488 457 395 304 300 289 213 152 161 154 132 103 114 116 110 88.3 102 105 107  K (mg/I) 2.9  Bi (mg/I) Ca (mg/I) Cd (mg/I) Co (mg/I) Cr (mg/I) Cu (mg/I) Fe (mg/I) 0.017 0.095 0.0059 650  Dissolved Metals  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  32 32 31 31 30 26 21 18 17 14 12.2 9.81 8.91 7.08 6.31 5.39 4.83 3.61 2.89 2.75 2.83 1.92 1.5 1.35 1.81 1.38 1.15 1.19 1.45 0.875 0.949 0.953 1.36 0.807 0.91 1.16 1.36 0.827 0.845 1.33 1.63 1.08  1.3 1.3 1.2 0.84 0.48 0.34 0.15 0.067 0.037 0.045 0.023 0.018 0.015 0.014 0.009 0.012 0.006 0.006 0.007 0.005 0.005 0.005 <0.005 <0.005 <0.005 <0.005 0.005 0.005 <0.005 <0.005 <0.005 <0.005 0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  <0.030 <0.030 <0.030 <0.030 <0.030 0.036 0.033 0.038 <0.030 0.034 0.033 0.031 0.033 <0.030 0.032 0.031 0.032 0.032 0.037 0.037 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 0.035 0.035 0.031  45 24 19 16 15 12 10 9.5 9.8 7.8 7.8 7.1 7.5 6 6.7 5.8 6.4 4.7 5.3 6.1 7.9 4.7 5 5.1 8 5.6 5.2 5.8 7.2 4.6 5.1 5 7.5 4.6 5.4 6.7 7.8 4.6 4.9 7.4 11.2 5.8  LI (mg/I) Mg (mgII) Mn (mg/I) Mo (mg/I) Na (mg/I) 93 33 1.8  16-Nov-93  Week 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44  P (mg/I)  <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30  Ni (mg/I)  <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <Q.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020 <0.020  Column 4  <0.20 <0.20 <0.20 <020 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30 <0.30  4.46 4.04 3.96 3.48 3.49 3.06 2.96 2.6 2.38 2.33 2.33 2.03 1.82 1.51 1.46 1.26 1.01 1.03 1.01 0.743 0.649 0.627 0.658 0.536 0.489 0.514 0.534 0.478 0.423 0.473 0.518 0.481  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  <0.030 <0.030 <0.030 <0.030 <0.030 0.034 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030 <0.030  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010  <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050 <0.050  V (mg/I)  Pb (mg/I) Sb (mg/I) Se (mg/I) Sn (mg/I) Sr (mg/I) Th (mg/I) Ti (mg/I)  Dissolved Metals  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  0.07 0.26 0.13 0.12 0.054 0.14 0.12 0.065 0.035 0.094 0.06 0.075 0.064 0.095 0.03 0.062 0.021 0.058 0.032 0.043 0.031 0.067 0.03 0.019 0.006 0.012 0.006 0.007 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  W (mg/I) Zn (mg/I) 0.38  16-Nov-93  Cell # 0.860 0.526 0.480 0.614 0.428 0.517 0.450 0.389 0.469 0.776 0.564 0.474 0.521 0.491 0.492 0.448 0.428 0.508 0.832 0.615 0.481 0.082 0.258 0.473 0.399 0.507 0.522 0.850 0.515 0.505 0.627 0.458 0.488 0.354 0.364 0.411  1.000 0.500 0.500 0.500 0.500 0.500 0.500 0.500 0.500 1.000 0.500 0.500 0.500 0.500 0.500 0.500 0.500 0.500 1.000 0.500 0.500 0.500 0.500 0.500 0.500 0.500 0.500 1.000 0.500 0.500 0.500 0.500 0.500 0.500 0.500 0.500  0 0 0 1 2 3 4 5 6 0 0 0 1 2 3 4 5 6 0 0 0 1 2 3 4 5 6 0 0 0 1 2 3 4 5 6  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  2 2 2 2 2 2 2 2 2  3 3 3 3 3 3 3 3 3  4 4 4 4 4 4 4 4 4  Week  8 3 4 0 7  0 0 0  8 0 0 5 10  0 0 0  0 0 0 1 10  0 0 0  0 5 4 1 10  0 0 0  Pro-Leach Vol. Collected Wtr. Level (mm) (I)  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  Date  Vol. Added (I)  1 1 1 1 1 1 1 1 1  Humidity Cells  766 757 755 785 783 745  624 617 653 644 631 633  5.13 4.32 4.27 5.17 4.48 4.41 4.37 4.40 4.41 2.75 2.81 2.77 2.82 2.92 2.82 2.72 2.58 2.67  593 671 657 654 628 638  470 542 544 548 384 514  Eh (mV H)  4.24 4.25 4.25 4.12 4.36 4.37 4.42 4.48 4.54  7.85 7.67 7.70 7.98 7.68 7.42 7.56 7.71 7.76  pH  105 52 10 20 <10 5  25 21 10 20 10 20  2230 915 1205 1090 1270 1110  0.50 0.34 0.34 0.49 0.33 0.24 0.44 0.32 0.33 6.10 3.07 3.26 2.54 2.18 2.28 2.55 2.83 2.31  24.5 22.0 14.5 19.5 26.5 24.0  Alkalinity Acidity (mg CaCO3/l) (mg CaCO3/l)  0.62 0.61 0.59 1.47 0.45 0.22 0.21 0.32 0.20  0.40 0.20 0.19 0.16 0.18 0.11 0.15 0.20 0.13  Cond. (mS/cm)  General Parameters  52500 2550 2750 1975 1600 1790 1500 1880 1425  125 100 198 140 150  250 158 165  290 265 290 870 215 88 78 125 80  109 62 55 48 58 34 48 65 35  Sulphate (mg/I)  16-Nov-93  Cell # 0861 0.498 0.526 0.676 0.413 0.498 0.199 0.346 0.506 0.740 0.558 0.484 0.610 0.494 0.509 0511 0.377 0.494 0.831 0.520 0.500 0.501 0.616 0.484 0.412 0.406 0.420 0.890 0.561 0.500 0.691 0.593 0.501 0.591 0.448 0.416  1.000 0.500 0.500 0.500 0.500 0500 0.500 0.500 0.500 1.000 0.500 0.500 0.500 0.500 0.500 0.500 0.500 0.500 1.000 0.500 0.500 0.500 0.500 0.500 0.500 0.500 0.500 1.000 0.500 0.500 0.500 0.500 0.500 0.500 0.500 0.500  0 0 0 1 2 3 4 5 6 0 0 0 1 2 3 4 5 6 0 0 0 1 2 3 4 5 6 0 0 0 1 2 3 4 5 6  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  6 6 6 6 6 6 6 6 6  7 7 7 7 7 7 7 7 7  8 8 8 8 8 8 8 8 8  Week  12 1 14 7 1  0 0 0  16 0 2 6 5  0 0 0  10 0 8 2 8  0 0 0  2 0 7 1 10  0 0 0  Pre-Leach Vol. Collected Wir. Level (mm) (I)  03-Aug-93 04-Aug-93 04-Aug-93 09-Aug-93 16-Aug-93 23-Aug-93 30-Aug-93 06-Sep-93 13-Sep-93  Date  Vol. Added (I)  5 5 5 5 5 5 5 5 5  Humidity Cells  772 770 748 756 701 734  758 723 708 775 806 759  730 694 720 712 718 706  567 577 614 623 558 584  2.38 2.52 2.45 2.47 2.52 2.64 2.48 2.32 2.28 2.86 2.93 2.93 3.16 3.23 3.35 3.29 3.24 322 6.65 7.27 7.04 7.40 7.00 6.80 7.13 6.97 6.72  Eh (mV H)  2.89 2.80 2.84 2.81 2.84 2.84 2.94 2.82 2.84  pH  1440 1045 955 370 620 610  1510 1320 1310 1440 2430 3480  195 192 145 124 115 150  5 <10 <10 10  5.35 3.05 2.74 2.82 244 0.24 1.39 1.90 188 12.70 5.70 5.55 3.61 3.44 3.05 3.14 3.96 4.76 2.80 2.05 1.91 0.90 0.88 0.76 0.79 0.75 0.97 0.77 0.59 0.49 0.76 0.46 0.33 0.39 0.41 0.51  13.0 9.0 5.5 8.0 75 6.0  Alkalinity Acidity (mg CaCO3/l) (mg CaCO3/I)  Cond. (ms/cm)  General Parameters  350 272 245 422 220 158 178 140 250  1650 1350 1212 425 375 350 360 310 450  190000 6750 5750 2800 2450 2150 2075 3200 4925  48500 2675 2175 2150 1700 1475 700 1110 1125  Sulphate (mg/I)  16-Nov-93  L’3  Cell  2.17 2.19 2.06  <0.015 <0.015 <0.015  <2.0 <2.0 <2.0 <0.030 <0.030 <0.030  0.255 0.116 0.082 0.214 0.16 0.362 0.09 0.06 0.062 <0.015 <0.015 <0.015  0.135 0.056 0.043 0.037 0.017 0.032 0.03 0.021 0.021  196 160 212 328 211 305 99.3 80.8 141 139 71.2 130  0.12 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  <0.005 <0.005 <0.005 <0005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005 <0.005  <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 <0.010 0.04 0.025 0.025  <0.20 <0.20 <0.20 <0.20 <0.20 0.35 <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  195 102 79.4 125 78.4 131 21.3 14.9 14.3 <0.20 <0.20 <0.20  <0.015 <0.015 <0.015  <0.015 <0.015 0.032  <0.015 <0.015 <0.015  <0.015 <0.015 <0.015  1-2 3-4 5-6  1-2 34 5-6  1-2 3-4 5-6  1-2 34 5-8  1 2 3  1 2 3  1 2 3  1 2 3  5 5 5  6 6 6  7 7 7  8 8 8  <0.010 <0.010 <0.010  0.223 0.166 0.141  0.116 0.072 0.05  174 200 241  <0.10 <0.10 <0.10  <0.005 <0.005 <0.005  <0.010 <0.010 <0.010  <0.20 <0.20 <0.20  167 151 133  <0.015 <0.015 <0.015  1-2 3-4 5-6  1 2 3  4 4 4  <2.0 <2.0 <2.0  <0.030 <0.030 <0.030  0.624 0.452 0.481 <0.010 <0.010 <0.010  <0.015 <0.015 <0.015 <0.015 <0.015 <0.015  3.6 1.62 2.2  8.94 6.48 6.92  0.02 0.017 0.015 <2.0 <2.0 <2.0 1.84 1.17 1.22  3.53 2.47 3.76  <0.015 <0.015 <0.015  65.1 30.7 41.6 0.073 0.045 0.061 192 247 819  2.67 1.62 1.27 <0.015 <0.015 <0.015  <2.0 <2.0 <2.0  59.8 26.2 24.4 0.068 0.045 0.038 <2.0 <2.0 <2.0 26.4 27.9 21.3  2.13 2.22 1.8  <0.015 <0.015 <0.015  0.076 0.054 0.1  47.6 33.1 26.6 0.058 0.068 0.057 <2.0 <2.0 <2.0 24.9 46.7 80.6  0.383 0.407 0.397  9.48 1.02 0.814  <0.015 <0.015 <0.015  0.028 0.027 0.027  48.9 53.6 72.5  <0.10 <0.10 <0.10  <0.005 <0.005 <0.005  0.011 0.011 0.011  <0.20 <0.20 <0.20  0.4 0.38 0.36  <0.015 <0.015 <0.015  1-2 3-4 5-6  1 2 3  3 3 3  0.026 0.027 0.027  <0.015 <0.015 <0.015  2.9 <2.0 <2.0  0.427 <0.030 <0.030  1.75 0.228 0.193  <0.015 <0.015 <0.015  0.123 <0.015 <0.015  0.121 0.015 0.013  214 34.3 58.8  <0.10 <0.10 <0.10  0.006 <0.005 <0.005  0.026 0.011 0.011  <0.20 <0.20 <0.20  2.54 0.23 <0.20  <0.015 <0.015 <0.015  1-2 3-4 5-6  1 2 3  2 2 2  4.55 3.68 4.47 <0.015 <0.015 <0.015  <2.0 <2.0 <2.0  <0.030 <0.030 <0.030  <0.010 <0.010 <0.010  <0.015 <0.015 <0.015  <0.015 <0.015 <0.015  23.3 17.4 21.9  <0.10 <0.10 <0.10  <0005 <0.005 <0.005  0.04 0.013 0.029  <0.20 <0.20 <0.20  <0.20 <0.20 <0.20  <0015 <0.015 <0.015  1-2 3-4 5-6  <0.010 <0.010 <0.010  U (mg/I)  Mg (mg/I)  16-Nov-93 K (mg/I)  Fe (mg/I)  Cu (mg/I)  Cr (mg/I)  Ca (mg/I) Cd (mg/I) Co (mg/I)  1 2 3  BI (mg/I)  Be (mg/I)  Ba (mg/I)  As (mg/I)  Al (mg/I)  Ag (mg/I)  Week  Dissolved Metals  1 1 1  Comp#  Humidity Cells  Cell  0.007 0.005 0.007 <0.10 <0.10 <0.10 <0.30 <0.30 <0.30  <0.20 <0.20 <0.20 <0.20 <0.20 <0.20  <0.050 <0.050 <0.050  <0.30 <0.30 <0.30  <0,020 <0.020 <0.020  4.9 2.3 3.1  <0.030 <0.030 <0.030  0.113 0.046 0.047  1-2 3-4 5-6  1 2 3  8 8 8  0.842 0.435 0.543  <0.010 <0,010 <0.010 <0.010 <0.010 <0.010  <0.10 <0.10 <0.10 <0.10 <0.10 <0.10  <0.030 <0.030 <0.030  4.68 3.35 3.58  <0.10 <0.10 <0.10 <0.030 <0.030 <0.030 <0.30 <0.30 <0.30  <0.20 <0.20 <0.20  <0.20 <0.20 <0.20  <0.050 <0.050 <0.050  <0.30 <0.30 <0.30  0.34 0.266 0.297  <2.0 <2.0 <2.0  <0.030 <0.030 <0.030  1.99 1.43 1.51  1-2 3-4 5-6  1 2 3  7 7 7  0.073 0.058 0.064  6.42 3.76 6.13 <0,10 <0.10 <0.10 0.054 0.063 0.075 <0.10 <0.10 <0.10  0.152 0.086 0.069 <0.30 <0.30 <0.30  <0.20 <0.20 0.26  <0.20 <0.20 <0.20  <0.050 <0.050 <0.050  0.64 0.85 5.38  1.44 0.731 1.07  <2.0 <2.0 <2.0  <0.030 <0.030 <0.030  16.3 7.58 10.5  1-2 3-4 5-6  1 2 3  6 6 6  0.253 0.239 0.677  23.5 9 6.37 <0.10 <0.10 <0.10  0.065 0.034 0.03 0,033 0.035 0.022 <0.10 <0.10 <0.10 0.257 0.227 0.208  <0.30 <0.30 <0.30  <0.20 <0.20 <0,20  <0,20 <0.20 <0.20  <0.050 <0.050 <0.050  <0.30 <0.30 <0.30  1.15 0.476 0.422  <2.0 <2.0 <2,0  <0.030 <0.030 <0.030  41.1 17.2 15.6  1-2 3-4 5-6  ‘  1 2 3  5 5 5  18.7 10.9 7.29 <0.10 <0.10 <0.10 0.058 0.037 <0.030 0.031 0.05 0.054  <0.10 <0,10 <0.10  0.232 0.3 0.26  <0.30 <0.30 <0.30  <0.20 <0.20 <0.20  <0.20 <0.20 <0.20  <0.050 <0.050 <0.050  <0.30 <0.30 0.43  1.15 0.784 0.399  <2.0 2.5 2.4  <0.030 <0.030 <0.030  33.3 21.7 16.7  1-2 3-4 5-6  1 2 3  4 4 4  3.94 4.01 3.81 <0.10 <0.10 <0.10 <0.030 <0.030 <0.030 <0.010 <0.010 <0.010  <0.10 <0.10 <0.10  0.25 0.272 0.252  <0.30 <0.30 <0.30  <0.20 <0.20 <0.20  <0.20 <0.20 <0.20  <0.050 <0.050 <0.050  <0.30 <0.30 <0.30  0.591 0.641 0.694  <2.0 <2.0 <2.0  <0.030 <0.030 <0.030  0.999 1 0.946  1-2 3-4 5-6  1 2 3  3 3 3  17.5 1.97 1.6  <0.10 <0.10 <0.10 <0.030 <0.030 <0.030  <0.010 <0.010 <0.010  <0.10 <0.10 <0.10  1.04 0.153 0.175  <0.30 <0.30 <0.30  <0.20 <0.20 <0.20  <0.20 <0.20 <0.20  <0.050 <0.050 <0.050  <0.30 <0.30 <0.30  2.46 0.406 0.367  3,8 <2.0 <2.0  <0.030 <0.030 <0.030  4.55 0.491 0.387  1-2 3-4 5-6  1 2 3  0.008 0.007 0.006 <0.10 <0.10 <0.10  <0.030 <0.030 <0.030  <0.010 <0.010 <0.010  2 2 2  <0.10 <0.10 <0.10  0.103 0.079 0.097  <0.30 <0.30 <0.30  <0.20 <0.20 <0.20  <0.20 <0.20 <0.20  <0.050 <0.050 <0.050  <0.30 <0.30 <0.30  <0.020 <0.020 <0.020  <2.0 <2.0 2.5  <0.030 <0.030 <0.030  0.022 0.012 0.012  1-2 3-4 5-6  Zn (mg/I)  W (mg/I)  V (mg/I)  Ti (mg/I)  16-Nov-93  1 2 3  Th (mg/I)  Sr (mg/I)  Sn (mg/I)  Se (mg/I)  Sb (mg/I)  Pb (mg/I)  P (mg/I)  NI (mg/I)  Mn (mg/I) Mo (mg/I) Na (mg/I)  Week  Dissolved Metals  1 1 1  Comp#  Humidity Cells  -  ir-CMP  3 3 3  1 1 1 1 1 1  21.1  21.0  0.07  21.0  21.0  21.0  21  20.7 20 19.4 18.8 18.4 18.2  0.05  21.0  0  20.3 18.9 18.0 17.4 17.5 17.8  0.08 0.28 0.45 0.66 0.99 1.10  19.6 18.0 17.5 16.7 16.7 16.6  18.6 16.6 16.0 15.4 15.1 15.0  19.0 16.6 15.7 15.0 14.5 13.6  19.1 17.8 17.5 16.8 16.8 16.7  1. 2.14 1. 1. 1. 1.  19.0 16.3 15.4 14.3 13.5 13.4  0.2 0.5 0.8 1. 1.4 1.7  --  0.18 0.27 0.40 0.48 0.10  20. 20.1 I 1 9. 1 9•4 I 1 9i 20.  B  0.10 0.16 2.77 3. 7 1 2 0.  20. B 20. 4 19. 4 19. I 9. 7 21. 0  0.: 2 0. 0. B 1. 1 1. 4 1. 7  11. 18.:  0.16 0.39 0.37 0.39 0.34 0.10  0.12 0.19 0.14 0.13 0.10 0.23  11.: 9 0.1 D 1.: 2 3. 10. 9 18.’ 0  1 4  o.: 2 0J 5 0j 3  0.00 0.00 0.15 0.23 0.34 0.34  0.03 0.03 0.15 0.22 0.26 0.24  20.5 20.5 19.7 19.7 20.0 20.0  20. 20. 2 1 2 2  0.04  0.07 0.21 0.40 0.64 0.87 1.00  0.11 0.29 0.56 0.76 0.82 0.98 0.04  0.08 0.25 0.46 0.70 0.84 0.99 0.03  0.04  0.04 0.19 0.39 0.64 0.80 0.96  0.04  0.30 0.69 0.96 1.14 1.23 1.22  Mol% C02 week 39 IWeek 40 Week 41 Week 42 Week 43 Week 44 1July28  0.2 0.5 0.8 1.1 1.4 1.7  Mol% 02 I Column# Dist from top (rTlweek 39 IWeek 40 Week 41 Week 42 Week 43 IWeek 44 July 28  Is land Copper Waste Rock Column Tests Pore Gas Analysis Results 28-Mar-94 Date:  University of British Columbia Department of Mining and Mineral Process Engineering  1.  0  C-) 0  C  216 APPENDIX 11  Beach Dump Characterization  Proportions of rock in each ARD potential category (Figure 5.6) and the estimated current effluent conditions for the Beach dump (Table 5.11) used acid base accounting data from three holes drilled in 1988. An additional seven holes have since been completed in the Beach dump area.  The entire length of each of the three 1988 drill holes was used for estimating acidity and sulfate concentrations of insitu water in the Beach dump. However, only the portion of the dump above mean sea level (1000 feet elevation using Island Copper co-ordinate system) is considered to have potential for sulfide oxidation and release of metals and net acidity.  The remainder,  permanently submerged in Rupert Inlet, is not expected to oxidize. Approximately 13 million tons (11.8 million tonnes) of the total 595 million tons (540 million tonnes) are expected to be above the 1000 foot elevation level at time of mine closure (Ian Home, pers. comm.). Since 1988, Island Copper has routinely monitored water quality from both the Beach dump drill holes and from sea water directly off of the Beach dump face. Results are given in the minds annual environmental assessment reports, and will also be discussed in the mine closure plan to be published in 1994.  BHP (1990) calculated mean NNP values by drill hole using both: i) the entire hole, and ii) the interval above 1000 feet elevation. However, two sample paired t-test analysis of this data indicates that there is no significant difference in between mean NNP’s calculated using the two methods. The Beach dump was constructed using mainly push dumping (Section 2.3.3), and material from each truckload is vertically distributed in moderately steeply dipping two metre wide bands. Therefore, one would not expect a significant variance in acid base accounting results between upper and lower elevations of the dump.  217 15 >  20  -J  0  ci  Cl)  -20 0 .0  < 0  Z  -40 -60  0  -80  a) -100  -100  -80  -60  -40  -20  0  20  Entire Length of Hole • Mean NNP (kg CaCO3It)  Fig. 1 Comparison of Beach dump mean NNP values between the interval above sea level and the entire drill hole (n= 10 holes) In conclusion, the prediction of acidity and sulfate concentrations of insitu water in the Beach dump done as part of this study could be further refined by using the entire ten hole database. However, use of  only those  samples above 1000 feet elevation appears statistically unjustified.  Column leach testing of Beach dump material, to be initiated at Island Copper Mine in late April or early May 1994, will provide further information on the sulfate, acidity and metal loading from this dump area.  

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