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Investigating metal attenuation processes in mixed sulfide carbonate bearing waste rock Laurenzi, Laura 2016

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INVESTIGATING METAL ATTENUATION PROCESSES IN MIXED SULFIDE CARBONATE BEARING WASTE ROCK by  Laura Laurenzi  B.Sc., Simon Fraser University, 2004  A THESIS SUBMITTED IN PARTIAL FULFILLMENT OF THE REQUIREMENTS FOR THE DEGREE OF  MASTER OF SCIENCE in THE FACULTY OF GRADUATE AND POSTDOCTORAL STUDIES (Geological Sciences)  THE UNIVERSITY OF BRITISH COLUMBIA (Vancouver)   May 2016  © Laura Laurenzi, 2016 ii  Abstract  The objective of this study was to identify the trace metal/secondary mineral phase associations in a heterogeneous waste rock dump that contains carbonate bearing lithologies and a mix of metal sulfides. The identification of attenuation processes can be used to better predict the drainage chemistry from waste rock at this site and/or other sites with similar waste rock. This study also provides the opportunity to investigate metal attenuation at the largest scale of complexity and compare these observations to those made from the smaller scale tests conducted for this site and is useful for understanding scalability of the smaller scale tests. This study shows that in carbonate bearing waste rock the predominant processes that attenuate copper (Cu) and zinc (Zn) are precipitation of hydroxycarbonate and hydroxysulfate phases and sorption onto iron oxides. Arsenic (As) and molybdenum (Mo) are associated with iron oxides, although for Mo this association was observed in only a few samples. Lead (Pb) was observed in association with iron oxides. Wulfenite observed in a few samples provides an additional attenuation process for Mo and Pb. The stability of the phases and potential for remobilization of these metals can also be suggested from this study. The hydroxycarbonate/hydroxysulfate phases are the least stable phases identified and can dissolve at pH<5. Iron oxides are considered a stable phase, as such, the As, Cu, Pb and Zn associated with these phases may also be quite stable. Sorption of Mo is limited at neutral pH but wulfenite is a stable phase that is not expected to dissolve once formed.  Geochemical modelling of seepages from the dump show that iron oxides are supersaturated and wulfenite and gypsum are at equilibrium. Two mixed Cu:Zn hydroxycarbonate phases and hydrozincite were added to the geochemical database and are supersaturated, while malachite iii  and smithsonite were generally undersaturated. Brochantite and antlerite were also generally undersaturated, but the observations made in this study show that copper hydroxysulfates and mixed copper/zinc hydroxysulfates are precipitating. In mixed sulfide/carbonate bearing waste rock mixed of Cu:Zn hydroxycarbonate and hydroxysulfate phases may require consideration for adequate prediction of Cu and Zn concentrations in drainage.  iv  Preface  Chapter 2 and Chapter 3 are based on the results of my efforts to identify secondary phases and the metals associated with them. Under the supervision of Roger D. Beckie, I was responsible for the design of the experimental program, carrying out the tests and data reduction/interpretation for these chapters. A version of Chapter 3 has been published in the proceedings from the 10th International Conference on Acid Rock Drainage (ICARD) and International Mine Water Association (IMWA) Annual Conference. Laurenzi, L., Ulrich Mayer, K., and Beckie, R.D. (2015). A Metal Attenuation Study on Waste Rock Collected from the East Dump, Antamina Mine, Peru: A Combined Mineralogical and Geochemical Approach. In Proceedings from the 10th ICARD and IMWA Annual Meeting, (Santiago, Chile). v  Table of Contents  Abstract .......................................................................................................................................... ii Preface ........................................................................................................................................... iv Table of Contents ...........................................................................................................................v List of Tables ................................................................................................................................ ix List of Figures ............................................................................................................................... xi Acknowledgements ......................................................................................................................xx Dedication ................................................................................................................................... xxi Chapter 1 Introduction..................................................................................................................1 1.1 Problem Description ....................................................................................................... 1 1.2 Study Background ........................................................................................................... 3 1.3 Site Description ............................................................................................................... 5 1.4 Method ............................................................................................................................ 6 1.4.1 Sequential Extraction Procedure Used........................................................................ 7 1.5 Thesis Structure .............................................................................................................. 8 Chapter 2 Comparison of Two Sequential Extraction Procedures Using Waste Rock Material Collected from the East Dump, Antamina Mine, Peru ............................................12 2.1 Introduction ................................................................................................................... 12 2.2 Site and Sample Collection ........................................................................................... 15 2.3 Methods......................................................................................................................... 16 2.3.1 X-Ray Diffraction (XRD) Mineralogy of Samples .................................................. 16 2.3.2 Total Elemental Composition ................................................................................... 17 vi  2.3.3 Sequential Extraction Procedures and Comparison .................................................. 17 2.3.4 Method Quality and Inter-sample Variability and Sample-split Variability ............ 20 2.4 Results ........................................................................................................................... 21 2.4.1 XRD Mineralogy of Samples.................................................................................... 21 2.4.2 Bulk Elemental Analysis........................................................................................... 22 2.4.3 Sequential Extractions and XRD Mineralogy........................................................... 22 2.4.3.1 Extraction Results Step 1 – Water Soluble Phases ........................................... 23 2.4.3.2 Extraction Results Step 2 – Metals Bound by Weak Electrostatic Forces and Cation Exchange ............................................................................................................... 24 2.4.3.3 Extraction Results Step 3 – Weak Acid Dissolvable Phases ............................ 24 2.4.3.4 Extraction Results Step 4 – Amorphous Reducible Phases .............................. 25 2.4.3.5 Extraction Results Step 5 – Crystalline Reducible Phases ............................... 26 2.4.3.6 Cumulative Extraction Results for Steps 1 - 5 .................................................. 26 2.5 Discussion ..................................................................................................................... 27 2.6 Conclusions ................................................................................................................... 30 Chapter 3 Characterizing Trace Metal Attenuation and Secondary Phases in Carbonate Bearing Waste Rock Collected from the East Dump, Antamina Mine, Peru ........................54 3.1 Introduction ................................................................................................................... 54 3.1.1 Site Information ........................................................................................................ 56 3.1.2 Sample Collection and Selection .............................................................................. 56 3.2 Methods......................................................................................................................... 57 3.3 Results ........................................................................................................................... 60 3.3.1 Lithology and Mineralogy ........................................................................................ 60 vii  3.3.2 Total Elemental Composition ................................................................................... 61 3.3.3 SEP Results ............................................................................................................... 62 3.3.4 Optical Microscopy and SEM/BSE Imaging of Secondary Minerals ...................... 63 3.4 Discussion ..................................................................................................................... 66 3.4.1 Hydroxycarbonate and Hydroxysulfate Phases ........................................................ 67 3.4.2 Iron Oxides................................................................................................................ 68 3.4.3 Implications............................................................................................................... 70 3.5 Conclusions ................................................................................................................... 72 Chapter 4 Conclusions .................................................................................................................95 4.1 Summary of Key Findings ............................................................................................ 95 4.2 Future Work .................................................................................................................. 98 Bibliography ...............................................................................................................................100 Appendices ..................................................................................................................................106 Appendix A Reverse Circulation Drill Logs .......................................................................... 106 Appendix B Method Quality and Inter-sample Variability (Chapter 2) ................................. 115 B.1 Method .................................................................................................................... 115 B.2 Results ..................................................................................................................... 116 B.3 Conclusion .............................................................................................................. 117 Appendix C Comparison of Sequential Extraction Procedures Method 1 and Method 2 and Differential X-ray Diffraction (DXRD) diffractograms ......................................................... 123 C.1 SEP Results for Method 1 and Method 2 ............................................................... 124 C.2 XRD Results for Un-reacted Samples and Sample Residues After Application of Leaching Steps .................................................................................................................... 129 viii  Appendix D Geochemical Modelling of Oxalate Species and Saturation Indices for Metal Oxalates................................................................................................................................... 137 D.1 PhreeqC Input File .................................................................................................. 138 Appendix E Summary of XRD Results (Tabulated) and X-ray Diffractograms for East Dump Waste Rock Samples............................................................................................................... 142 E.1 Tabulated XRD Mineralogy ................................................................................... 143 E.2 XRD Patterns for Individual Samples..................................................................... 144 Appendix F Total Elemental Concentrations From 4-acid Digestions ................................... 160 Appendix G Sequential Extraction Procedure (SEP) Results for East Dump Waste Rock Samples ................................................................................................................................... 162 Appendix H SEM and EDS Investigation Images and Elemental Wt.% ................................ 181 H.1 EDS Un-normalized Wt.% of Elements from Semi-quantitative Analysis of EDS 182 Appendix I Drainage Chemistry Geochemical Modelling ..................................................... 196 I.1 Example PhreeqC Input File for Geochemical Modelling of East Dump Seepage Chemistry ............................................................................................................................ 197  ix  List of Tables Table 2.1 Samples selected for this study ..................................................................................... 43 Table 2.2 Sequential extraction procedures (SEPs) used in this study ......................................... 44 Table 2.3 Qualitative mineralogy from XRD ............................................................................... 45 Table 2.4 Elemental composition of via 4-acid digestion ............................................................. 46 Table 2.5 Acid base accounting (ABA) static test results ............................................................ 46 Table 2.6 Average concentrations and RPD values for Step 1 ..................................................... 47 Table 2.7 Average concentrations and RPD values for Step 2 ..................................................... 48 Table 2.8 Average concentrations and RPD values for Step 3 ..................................................... 49 Table 2.9 Average concentrations and RPD values for Step 4 ..................................................... 50 Table 2.10 Average concentrations and RPD values for Step 5 ................................................... 51 Table 2.11 RPD for cumulative concentrations of Step 1 through Step 5 .................................... 52 Table 2.12 Solubility product constants for metal oxalates .......................................................... 53 Table 3.1 Samples selected for metal attenuation study ............................................................... 74 Table 3.2 Sequential extraction procedure (SEP) (modified from Hall, Vaive, Beer, & Hoashi, 1996) ............................................................................................................................................. 76 Table 3.3 Summary of XRD identified mineralogy for East Dump waste rock samples ............. 77 Table 3.4 Trace element concentration by four-acid digestion and ICP-MS finish ..................... 79  Table B.1 Calculated RSD for triplicate analysis for Method 1 ................................................. 118 Table B.2 Calculated RSD for triplicate analysis for Method 2 ................................................. 119 Table B.3 Percent recovery (total elemental concentration vs. cumulative leached) ................. 120 Table B.4 Mass loss experiment ................................................................................................. 121 x  Table B.5 pH drift experiment .................................................................................................... 122  Table D.1 Tabulated results of SI and precipitation scenarios ................................................... 137  Table F.1 Tabulated elemental concentrations from 4-acid digestions ...................................... 161  Table H.1 EDS un-normalized wt.% of elements from semi-quantitative analysis of EDS ...... 182  xi  List of Figures Figure 1.1 Location of the Antamina Mine in Peru, South America (inset). From (Conlan, 2009).......................................................................................................................................................... 9 Figure 1.2 The Antamina Mine site, inset plan of the East Dump and pre-mining topography showing drill site locations and seeps downslope of drill sites .................................................... 10 Figure 1.3 Experimental framework ............................................................................................. 11 Figure 2.1 View of East Dump and drilling location (Site 1 and Site 3) ...................................... 33 Figure 2.2 Ca and Fe (% leached) SEP results for BH1-1, BH1-2, BH1-4 and BH3-3 ............... 34 Figure 2.3 As and Mo (% leached) SEP results for BH1-1, BH1-2, BH1-4 and BH3-3 .............. 35 Figure 2.4 Cu, Pb, and Zn (% leached) SEP results for BH1-1, BH1-2, BH1-4 and BH3-3 ....... 36 Figure 2.5 BH1-4 XRD results from Method 1 ............................................................................ 37 Figure 2.6 BH1-4 XRD results from Method 2 ............................................................................ 38 Figure 2.7 BH1-1 XRD results from Method 1 ............................................................................ 39 Figure 2.8 BH1-1 XRD results from Method 2 ............................................................................ 40 Figure 2.9 BH1-3 XRD results from Method 1 ............................................................................ 41 Figure 2.10 BH1-3 XRD results from Method 2 .......................................................................... 42 Figure 3.1 The Antamina Mine site, inset plan of the East Dump and pre-mining topography showing drill site locations and seeps downslope of drill sites .................................................... 80 Figure 3.2 Box and whisker plot of total metal concentration box per site .................................. 81 Figure 3.3 SEP results for Ca (%) with inset total concentration (ppm) ...................................... 82 Figure 3.4 SEP results for Fe (%) with inset total concentration (ppm) ....................................... 83 Figure 3.5 SEP results for As (%) with inset total concentration (ppm) ...................................... 84 Figure 3.6 SEP results for Cu (%) with inset total concentration (ppm) ...................................... 85 xii  Figure 3.7 SEP results for Mo (%) with inset total concentration (ppm) ..................................... 86 Figure 3.8 SEP results for Pb (%) with inset total concentration (ppm)....................................... 87 Figure 3.9 SEP results for Zn (%) with inset total concentration (ppm) ...................................... 88 Figure 3.10 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: Figure 10-A is from BH-1d (19.5 – 21.0); Figure 10-B and Figure 10-C are from BH-1d (91.5-92.4) ................................................................................................. 89 Figure 3.11 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: All images are from BH-3s (15.0 – 16.5) .................................... 90 Figure 3.12 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: Figure 3.12-A from BH-1s (1.5 – 3.0); Figure 3.12-B and -C from BH-1d (90.0 – 91.5) ...................................................................................................................... 91 Figure 3.13 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: Figure 13-A from BH-3d (24.0 – 25.5); Figure 13-B and Figure 13-C from BH-3d (48.0 – 49.5) .................................................................................................... 92 Figure 3.14 Equilibrium phases predicted from geochemical modelling of seeps downslope Site 1; CO-41 (top) and CO-57(bottom) .............................................................................................. 93 Figure 3.15 Equilibrium phases predicted from geochemical modelling of seeps downslope Site 3; CO-28 (top) and CO-56(bottom) .............................................................................................. 94  Figure A.1 BH-1s (0.0m - 23.50m)............................................................................................. 107 Figure A.2 BH-1d (0.0m - 49.50m) ............................................................................................ 108 Figure A.3 BH-1d (49.50m - 92.40m) ........................................................................................ 109 Figure A.4 BH-3s (0.0m – 22.50m) ............................................................................................ 110 xiii  Figure A.5 BH-3d (0.0m - 46.50m) ............................................................................................ 111 Figure A.6 BH-3d (46.50m - 91.50m) ........................................................................................ 112 Figure A.7 BH-3d (91.50m - 136.50m) ...................................................................................... 113 Figure A.8 BH-3d (136.50m - 145.00m) .................................................................................... 114  Figure C.1 Sequential extraction results Al (%) and Ca (%) ...................................................... 124 Figure C.2 Sequential extraction results Mn (%) and Mg (%) ................................................... 125 Figure C.3 Sequential extraction results Fe (%) and As (%) ...................................................... 126 Figure C.4 Sequential extraction results Cu (%) and Mo (%) .................................................... 127 Figure C.5 Sequential extraction results Pb (%) and Zn (%) ...................................................... 128 Figure C.6 BH1-1 XRD results from Method 1 ......................................................................... 129 Figure C.7 BH1-2 XRD results from Method 1 ......................................................................... 129 Figure C.8 BH1-3 XRD results from Method 1 ......................................................................... 130 Figure C.9 BH1-4 XRD results from Method 1 ......................................................................... 130 Figure C.10 BH3-1 XRD results from Method 1 ....................................................................... 131 Figure C.11 BH3-3 XRD results from Method 1 ....................................................................... 131 Figure C.12 BH3-2 XRD results from Method 1 ....................................................................... 132 Figure C.13 BH3-4 XRD results from Method 1 ....................................................................... 132 Figure C.14 BH1-1 XRD results from Method 2 ....................................................................... 133 Figure C.15 BH1-2 XRD results from Method 2 ....................................................................... 133 Figure C.16 BH1-3 XRD results from Method 2 ....................................................................... 134 Figure C.17 BH1-4 XRD results from Method 2 ....................................................................... 134 Figure C.18 BH3-1 XRD results from Method 2 ....................................................................... 135 xiv  Figure C.19 BH3-2 XRD results from Method 2 ....................................................................... 135 Figure C.20 BH3-3 XRD results from Method 2 ....................................................................... 136 Figure C.21 BH3-4 XRD results from Method 2 ....................................................................... 136  Figure E.1 BH-1s (1.5 - 3.0) ....................................................................................................... 144 Figure E.2 BH-1s (10.5 -12.0) .................................................................................................... 144 Figure E.3 BH-1s (19.5 - 21.0) ................................................................................................... 145 Figure E.4 BH-1d (1.5 - 3.0) ....................................................................................................... 145 Figure E.5 Rietveld analysis of BH-1d (19.5 – 21.0) ................................................................. 146 Figure E.6 BH-1d (25.5 - 27.0) ................................................................................................... 146 Figure E.7 BH-1d (39.0 - 40.5) ................................................................................................... 147 Figure E.8 BH-1d (42.0 - 43.5) ................................................................................................... 147 Figure E.9 BH-1d (49.5 - 51.0) ................................................................................................... 148 Figure E.10 BH-1d (54.0 - 55.5) ................................................................................................. 148 Figure E.11 BH-1d (64.5 - 66.0) ................................................................................................. 149 Figure E.12 BH-1d (70.5 - 72.0) ................................................................................................. 149 Figure E.13 BH-1d (76.5 - 78.0) ................................................................................................. 150 Figure E.14 BH-1d (81.0 - 82.5) ................................................................................................. 150 Figure E.15 BH-1d (90.0 - 91.5) ................................................................................................. 151 Figure E.16 BH-1d (91.5 - 92.4) ................................................................................................. 151 Figure E.17 BH-3s (1.5 - 3.0) ..................................................................................................... 152 Figure E.18 BH-3s (9.0 - 10.5) ................................................................................................... 152 Figure E.19 BH-3s (10.5 - 12.0) ................................................................................................. 153 xv  Figure E.20 Rietveld analysis of BH-3s (15.0 – 16.5) ................................................................ 153 Figure E.21 BH-3s (21.0 – 22.5)................................................................................................. 154 Figure E.22 BH-3d (10.5 - 12.0) ................................................................................................. 154 Figure E.23 BH-3d (24.0 - 25.5) ................................................................................................. 155 Figure E.24 BH-3d (37.5 - 39.0) ................................................................................................. 155 Figure E.25 BH-3d (39.0 - 40.5) ................................................................................................. 156 Figure E.26 BH-3d (48.0 - 49.5) ................................................................................................. 156 Figure E.27 BH-3d (72.0 - 73.5) ................................................................................................. 157 Figure E.28 BH-3d (88.5 - 90.5) ................................................................................................. 157 Figure E.29 BH-3d (91.5 - 93.0) ................................................................................................. 158 Figure E.30 BH-3d (94.5 - 96.0) ................................................................................................. 158 Figure E.31 BH-3d (112.5 - 114.0) ............................................................................................. 159 Figure E.32 BH-3d (121.5 - 123.0) ............................................................................................. 159  Figure G.1 SEP results for Al (ppm) inset graph shows total elemental from 4-acid digestions Al (ppm) ........................................................................................................................................... 163 Figure G.2 SEP results for Al (%) inset graph shows total elemental from 4-acid digestions Al (ppm) ........................................................................................................................................... 164 Figure G.3 SEP results for Ca (ppm) inset graph shows total elemental from 4-acid digestions Ca (ppm) ........................................................................................................................................... 165 Figure G.4 SEP results for Ca (%) inset graph shows total elemental from 4-acid digestions Ca (ppm) ........................................................................................................................................... 166 xvi  Figure G.5 SEP results for Fe (ppm) inset graph shows total elemental from 4-acid digestions Fe (ppm) ........................................................................................................................................... 167 Figure G.6 SEP results for Fe (%) inset graph shows total elemental from 4-acid digestions Fe (ppm) ........................................................................................................................................... 168 Figure G.7 SEP results for Mn (ppm) inset graph shows total elemental from 4-acid digestions Mn (ppm) .................................................................................................................................... 169 Figure G.8 SEP results for Mn (%) inset graph shows total elemental from 4-acid digestions Mn (ppm) ........................................................................................................................................... 170 Figure G.9 SEP results for As (ppm) inset graph shows total elemental from 4-acid digestions As (ppm) ........................................................................................................................................... 171 Figure G.10 SEP results for As (%) inset graph shows total elemental from 4-acid digestions As (ppm) ........................................................................................................................................... 172 Figure G.11 SEP results for Cu (ppm) inset graph shows total elemental from 4-acid digestions Cu (ppm) ..................................................................................................................................... 173 Figure G.12 SEP results for Cu (%) inset graph shows total elemental from 4-acid digestions Cu (ppm) ........................................................................................................................................... 174 Figure G.13 SEP results for Mo (ppm) inset graph shows total elemental from 4-acid digestions Mo (ppm) .................................................................................................................................... 175 Figure G.14 SEP results for Mo (%) inset graph shows total elemental from 4-acid digestions Mo (ppm) ........................................................................................................................................... 176 Figure G.15 SEP results for Pb (ppm) inset graph shows total elemental from 4-acid digestions Pb (ppm)...................................................................................................................................... 177 xvii  Figure G.16 SEP results for Pb (%) inset graph shows total elemental from 4-acid digestions Pb (ppm) ........................................................................................................................................... 178 Figure G.17 SEP results for Zn (ppm) inset graph shows total elemental from 4-acid digestions Zn (ppm) ..................................................................................................................................... 179 Figure G.18 SEP results for Zn (%) inset graph shows total elemental from 4-acid digestions Zn (ppm) ........................................................................................................................................... 180  Figure H.1 Images from BH-1d (19.5 – 21.0) A) Plane-polarized transmitted light photograph of silicate/calcite mineral with black and blue/green secondary mineral coating. B) SEM/BSE image ........................................................................................................................................... 184 Figure H.2 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of zinc silicate mineral (hemimorphite) with blue/green secondary mineral coating. B) SEM/BSE image ........................................................................................................................................... 184 Figure H.3 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image ....................................................................................................... 185 Figure H.4 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image ....................................................................................................... 185 Figure H.5 Images from BH-3s (15.0 – 16.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image ....................................................................................................... 186 Figure H.6 Images from BH-3s (15.0 – 16.5) A) Plane-polarized transmitted light photograph of blue/green secondary mineral associated with oxidized chalcopyrite grain (Cpy) B) SEM/BSE image ........................................................................................................................................... 186 xviii  Figure H.7 Images from BH-3s (15.0 – 16.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image ....................................................................................................... 187 Figure H.8 Images from BH-1s (1.5 – 3.0) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image ....................................................................................................... 187 Figure H.9 Images from BH-1d (19.5 – 21.0) A) Plane-polarized transmitted light photograph of mineral with coating B) SEM/BSE image .................................................................................. 188 Figure H.10 Images from BH-1s (1.5 – 3.0) A) Plane-polarized transmitted light photograph of pyrite mineral (Py) with iron oxide coatings in a calcite grain B) SEM/BSE image ................. 188 Figure H.11 Images from BH-1s (1.5 – 3.0) A) Plane-polarized transmitted light photograph of Chalcopyrite (Cpy) coated with iron oxide B) SEM/BSE image of same mineral showing zoning of iron oxide ................................................................................................................................ 189 Figure H.12 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of iron oxide grain. B) SEM/BSE image .................................................................................... 189 Figure H.13 Images from BH-1d (90.0 – 91.5) Plane-polarized transmitted light photograph of iron oxide coating hemimorphite (white mineral) B) SEM/BSE image ..................................... 190 Figure H.14 Image from BH-1d (90.0 - 91.5) A) Plane-polarized transmitted light photograph of iron oxide coating hemimorphite (white mineral) B) SEM/BSE image (300 micron scale) ...... 190 Figure H.15 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of weathered opaque mineral with iron oxide coating B) SEM/BSE image .............................. 191 Figure H.16 Images from BH-1d (19.5 – 21.0) A) Plane-polarized transmitted light photograph of pyrite (with Zn) mineral with iron oxide coatings B) SEM/BSE image ................................ 191 Figure H.17 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of pyrite (with 0.2 wt.% Zn) mineral with iron oxide coating B) SEM/BSE image .................. 192 xix  Figure H.18 Images from BH-3d (24.0 – 25.5) A) Plane-polarized transmitted light photograph of calcite mineral with iron oxide coatings B) SEM/BSE image ............................................... 192 Figure H.19 Image from BH-3d (24.0 – 25.5) A) Plane-polarized transmitted light photograph of garnet (?) mineral with iron oxide coatings B) SEM/BSE image ............................................... 193 Figure H.20 Images from BH-3d (24.0 – 25.5) A) Plane-polarized transmitted light photograph of garnet and other mineral (calcite?) with iron oxide coating B) SEM/BSE image magnified into iron oxide coating (50 micron scale) .......................................................................................... 193 Figure H.21 Images from BH-3d (48.0 – 49.5) A) Plane-polarized transmitted light photograph of pyrite (with 0.23 wt.% Zn) with iron oxide oxidation rim B) SEM/BSE image .................... 194 Figure H.22 Image from BH-3d (48.0 -49.5) A) Plane-polarized transmitted light photograph of pyrite (Py) (with 0.22 wt.% Zn) and chalcopyrite (Cpy) with iron oxide coating B) reflected light microscope image C) SEM/BSE image ...................................................................................... 194 Figure H.23 Image from BH-3d (91.5 – 93.0) A) Plane-polarized transmitted light photograph of silicate minerals with iron oxide coatings. B) SEM/BSE image ................................................ 195 Figure H.24 Image form BH-3d (91.5 – 93.0) A) Plane-polarized transmitted light photograph of pyrite (Py) (with Zn) with iron oxide coating B) SEM/BSE image ........................................... 195    xx  Acknowledgements  Natural Sciences and Engineering Research Council (NSERC) and Compañia Minera Antamina (CMA) provided the funding that made this study possible.  I thank my supervisors Roger Beckie and Uli Mayer for the opportunity to be part of a large-scale, multi-faceted project such as this. Throughout my appointment they were supportive, patient, and fun to talk to about the complexities of this study. I could not have asked for better supervisors.  Bevin Harrison, Roberto Manrique, Celedonio Aranda, Micheal Edsael Sanchez, Bartolomeo Vargas at Antamina provided excellent technical support and expert site-specific knowledge.  UBC is an amazing place and I am forever indebted to the EOAS staff that trained me and helped me throughout my journey, namely; Maureen Soon, Mati Raudsepp, Jenny Lai, Lan Kato, Elisabetta Pani, Edith Czech. I am also grateful to Les Lavkulich for some very interesting discussions on sequential extractions.   I owe many thanks to my cohort, office mates, colleagues, and predecessors. I am humbled by how much you know, how willingly you shared what you know, and how much coffee we drank in the process. xxi  Dedication  This study is dedicated to my family and friends who offered encouraging and motivating words when the light at the end of the tunnel seemed to get further away and, in the final stages, when it got closer and closer.  A special dedication goes to my partner Benjamin Woolsey for every silver lining and every golden moment I would have missed without him.1  Chapter 1 Introduction 1.1 Problem Description Non-economic waste rock typically constitutes the largest volume of waste produced during mining, especially in open-pit operations. After being exhumed, waste rock is stockpiled onsite in dumps and left exposed to ambient conditions. The primary environmental concern with respect to waste rock is drainage quality, and at all stages of mining significant effort is put forth to understand the hydrologic, geochemical, and microbiological influences on the waste rock with the goal of producing accurate predictions of water quality. Initial geochemical investigations of waste rock involve assessing primary mineralogy, trace element composition and potential to generate/buffer acidity of the waste rock. However, researchers have concluded that metals associated with secondary mineral phases are more important to identify and quantify because they exert a much stronger control on water quality (Al et al., 2000; Sloot and Zomeren, 2012). The chemistry of drainage from a mine site is the result of the competing processes of acid generation, acid neutralization, and secondary mineral production resulting in metal attenuation. The oxidation of sulfide minerals proceeds either via dissolved oxygen (O2) or dissolved ferric iron (Fe3+) acting as the oxidizing agents (Nordstrom and Alpers, 1999). The relative importance of oxidation by dissolved oxygen or ferric iron depends on the pH of the water and microbiology. Oxygen dominates oxidation at near-neutral conditions (Nordstrom and Alpers, 1999). Dissolved ferric iron is more effective than dissolved oxygen in oxidizing sulfide minerals, but has limited solubility in neutral pH conditions and is generally a more effective oxidizing agent at low-pH conditions (though both are active at low pH) (Nordstrom and Alpers, 1999). Typically pyrite is the most abundant sulfide mineral that oxidizes to generate acidity:  2  𝐅𝐞𝐒𝟐  + 𝟕𝟐𝐎𝟐  +  𝐇𝟐𝐎 = 𝐅𝐞𝟐+ +  + 𝟐𝐒𝐎𝟒𝟐−  +  𝟐𝐇+     1-1 𝐅𝐞𝐒𝟐  +  𝟏𝟒𝐅𝐞𝟑+  +  𝐇𝟐𝐎 = 𝟏𝟓𝐅𝐞𝟐+ +  + 𝟐𝐒𝐎𝟒𝟐−  +  𝟏𝟔𝐇+    1-2 Other sulfide minerals such as pyrrhotite, sphalerite, galena, and chalcopyrite can also oxidize to produce acidity depending on the oxidizing agent. Dissolution of some carbonate minerals, such as calcite, consumes the acidity generated by sulfide oxidation:  𝐂𝐚𝐂𝐎𝟑 +  𝐇+ = 𝐂𝐚𝟐+ + 𝐇𝐂𝐎𝟑−        1-3 While other carbonate minerals, such as siderite and rhodochrosite, will cause the release of acidity due to hydrolysis/precipitation of the released Fe2+ and Mn2+. Aluminosilicate minerals can also consume acid, however, their reactivity is slower and less effective than carbonates with the exception of wollastonite and olivine (Jambor et al., 2002). It is the balance between the acidity generated by oxidation versus the dissolution of buffering minerals that determines if acid rock drainage (ARD) or neutral rock drainage (NRD) will dominate the overall drainage of the waste material. However, in a heterogeneous waste rock dump with both potentially acid-generating (PAG) and non-acid generating (NAG) waste rock, it is possible to have acidic zones producing acidic drainage and neutral zones producing neutral drainage. Although acidity in mine drainage defines the problem, the issues are related to dissolved metals and metalloids (referred herein as metals) in the drainage for which many countries will have regulated concentrations for release. Dissolved metal concentrations differ in ARD and NRD because metal mobility is highly dependent on pH. In acidic conditions metals such as Al, Fe, Cu, Pb, Zn, Cd, and Mn are mobile, while in neutral conditions elements which are either weakly hydrolyzing, like Zn, or oxyanion forming like Mo and As are mobile (Price, 2009; Stumm and Morgan, 1995).  3  Studies have shown that significant metal sequestration occurs on secondary mineral surfaces due to sorption onto metal oxides (Nordstrom, 2011). The most prolific metal oxides in sulfidic mining wastes are iron oxides. In addition to sorption of metals, precipitation of secondary minerals due to solubility limitations in highly concentrated waters, or by evapo-concentration, also sequesters metals and stores them in the solid phases. Secondary phases that form as a result of precipitation from concentrated solutions are often initially amorphous and are low abundance compared to primary minerals (Bigham et al., 1996; Jang et al., 2003). The potential for long-term storage of trace metals depends on the stability of the phase that is attenuating metals and the stability of the environment it was sequestered in. Consequently, understanding the process of attenuation gives insight into the local environment that attenuation occurred and the potential for remobilization.  1.2 Study Background Since 2006, UBC, Compañia Minera Antamina (CMA) and Teck-ART have been collaborating on a multi-scale study of the hydrology, geochemistry and microbiology of waste rock at the Antamina mine site. The field studies include five 36m x 36m x 10m tall experimental waste rock piles (Bay, 2009). Three of these piles are composed of a single waste rock class (Peterson, 2014) and two are composed of a combination of waste rock classes (Blackmore, 2015). In addition to the experimental piles, smaller scale field studies involving field barrels of the single lithology and of mixed lithologies were constructed. The mixed lithology field barrels were designed to identify attenuation of Mo and Zn (Hirsche et al., 2012). Laboratory tests were also conducted which include batch tests, humidity cell testing and mixed lithology column tests (Blackmore, 2015; Conlan et al., 2012; Dockrey et al., 2014; Hirsche et 4  al., 2012). This study was motivated, in part, by the conclusions of previous researchers using smaller scale studies at Antamina to identify attenuation mechanism and secondary phases. These conclusions are: 1) Mo studies showed that wulfenite precipitation is favoured over kinetically limited powellite, however in the absence of Pb, precipitation of powellite should be the only significant attenuation mechanism for Mo in a carbonate buffered system (Conlan et al., 2012).  2) Acidic micro-environments in waste rock could lead to the precipitation of iron oxides where Mo sorption is possible (Dockrey et al., 2014). 3) Attenuation of Mo was observed when Mo leaching waste rock was placed over Pb bearing waste rock (Hirsche et al., 2012) 4) Zn attenuation was suggested to occur via precipitation of a carbonate or hydroxide phase or may be incorporated into crystalline structure of phyllosilicate clay minerals such as clinochlore (Hirsche et al., 2012). Hirsche et al. (2012) also suggested from modelling that the precipitation of smithsonite could not account for all of the Zn attenuated in his studies. 5) In one of the experimental waste rock piles a blue precipitate consisting of gypsum, malachite and mostly amorphous phases in which X-ray absorption spectroscopy (XAS) showed that Cu in the amorphous material was bound primarily to sulfates in the form of brochantite and to some carbonates in the form of malachite (Peterson, 2014). 5  The primary objective of this study is to identify metal attenuation processes in the full scale waste rock dump and hypothesizes that the processes identified at this scale will be similar to what was identified at the smaller scales of investigation.  1.3 Site Description The Antamina deposit is a large copper-zinc-molybdenum skarn deposit with smaller quantities of silver, bismuth and lead that formed by the intrusion of a quartz monzonite body into limestones (Lipten & Smith, 2004; Love, Clark, & Glover, 2004; Redwood, 1999). The Antamina mine is located approximately 270km NE of Lima, Peru in the department of Ancash (Figure 1.1). It is situated in the Andes, at an elevation ranging between 4200 and 4700masl. Antamina receives approximately 1200-1300mm precipitation per year, 80% of which falls as rain during the region’s wet season (October – April). The mean annual temperature at the mine site is ~5.0°C measured at the meteorological station at Punto B.  The East Dump receives both potentially acid-generating (PAG) waste rock and non-acid generating/acid-buffering (NAG) waste rock from a range of lithologies, i.e., limestone, marble, hornfels, exo- endo- skarn and intrusive, and is thus considered geochemically heterogeneous. At the time of drilling (November 2012 – February 2013), the deepest material in the East Dump had been in place for over 10 years, while the shallowest material for more than 5 years. Two holes were drilled at two sites, Site 1 and Site 3, on the East Dump (Figure 1.2). The holes were drilled using air-driven reverse circulation (RC) using a Casagrande C8 drill. Drilling fluids were not used to minimize the alteration of secondary phases. As part of the drilling program drill cuttings were logged by the Antamina Geology Department; these logs are presented in Appendix A. The size fraction selected for testing was <2mm, based on work conducted by 6  Strömberg & Banwart (1999) which identified that particles with diameters smaller than 0.25mm contribute to approximately 80% of the sulfide and silicate dissolution, and carbonate minerals larger than 5–10mm react too slowly to neutralize the acid produced from sulfides. Thus, the <2mm material was selected for this study with the expectation that these size fractions would host the secondary minerals that form coatings on grains as a result of sulfide oxidation and carbonate neutralization.  1.4 Method The observations made in this study are from a suite of geochemical and mineralogical tests. Figure 1.3 presents the experimental framework used for this study. X-ray diffraction (XRD) and total digests were used to characterize the mineralogy and total elemental content of the samples, respectively. Sequential extractions (SEP) were used to leach metals out of operationally defined “pools” representative of secondary minerals/phases. The results of these tests were then used to select a smaller number of samples to be prepared into thin-sections and undergo a more detailed mineralogical investigation. The detailed mineralogical investigation involved using transmitted and reflected light microscopy and scanning electron microscopy to identify secondary coatings on primary mineral surfaces. Energy dispersion x-ray spectroscopy (EDS) was used to obtain a semi-quantitative analysis of metal content of the secondary phases by focusing the beam on spot locations. Aqueous chemistry of seepages located downslope of the drill sites (also shown in Figure 1.2) was analysed using PhreeqC (Parkhurst and Appelo, 2013) to obtain saturation indices (SIs) of potential controlling phases. 7  1.4.1 Sequential Extraction Procedure Used It has only been in the last few decades that SEPs have been used to investigate metal attenuation (metal mobility and metal partitioning) in mine waste materials. Most of these studies have been focused on tailings (Carlsson et al., 2002; Dold, 2003; Fanfani et al., 1997, 1997; Hall et al., 1996); and only a few studies have conducted SEPs on waste rock (Jeong and Lee, 2003; Singh and Hendry, 2012; Stockwell et al., 2006). In this study, sequential extractions were used to suggest what phases metals were associated with and to identify samples in which high concentrations of elements might allow for visual identification of these phases in thin-section. Sequential extraction procedures also provide a way of obtaining useful information concerning the stability of the metals and associated phases. SEPs are “operationally” defined meaning that different extraction reagents can be used depending on the metals of interest and the expected secondary phases. Accordingly, multiple SEPs have been proposed. The main criticisms of SEPs are the lack of selectivity due to the wide range of secondary minerals possible (especially in mining wastes) and the potential to dissolve non-targeted phases causing difficulty in the interpretation of results. Two SEP methods were identified in the literature and were both suitable for this study because they accounted for the secondary minerals expected in this waste rock environment (i.e., water soluble phases, carbonate phases, and iron oxides). These two methods were investigated in Chapter 2 of this thesis to determine if there was a difference in the results obtained between the two methods and, if so, to what those difference could be attributed. A sub-set (N=8) of East Dump waste rock samples were selected for this study. SEP leachate results were paired with XRD scans of the unreacted material and of the residues after each extraction step, in order to compare the concentrations of metals leached with the minerals dissolved or precipitated during each step. 8  1.5 Thesis Structure This thesis is written in a paper-based format. It consists of four chapters: an introduction, two papers and a conclusion. The two papers are self-contained – each with an introduction, methods, results, discussion sections. The first paper (Chapter 2) presents the investigation of the two sequential extraction procedures conducted on samples from the East Dump. The second paper (Chapter 3) uses the results of the experimental framework described above to discuss the attenuation of As, Cu, Pb, Mo, and Zn in mixed sulfide/carbonate bearing waste rock collected from East Dump, Antamina Mine, Peru.   9  Figures  Figure 1.1 Location of the Antamina Mine in Peru, South America (inset). From (Conlan, 2009). 10    Figure 1.2 The Antamina Mine site, inset plan of the East Dump and pre-mining topography showing drill site locations and seeps downslope of drill sites  11   Figure 1.3 Experimental framework   12  Chapter 2 Comparison of Two Sequential Extraction Procedures Using Waste Rock Material Collected from the East Dump, Antamina Mine, Peru 2.1 Introduction Sequential extraction procedures (SEPs) were first proposed in mineral exploration because they allowed for the distinction between primary and secondary metal bearing minerals, the latter of which was used to infer “signals” of hidden deposits (Gatehouse et al., 1977; Sondag, 1981; Tessier et al., 1979). SEPs were also developed to understand metal retention and mobility in contaminated environments (Chao and Zhou, 1983; Wenzel et al., 2001). In the last few decades SEP methods have been proposed and used on mine wastes, mainly tailings (Carlsson et al., 2002; Dold, 2003; Fanfani et al., 1997, 1997; Hall et al., 1996) while only a few studies have focused on waste rock (Jeong and Lee, 2003; Singh and Hendry, 2012; Stockwell et al., 2006). SEPs are ideal for determining metal retention in secondary phases that cannot be identified via instrumental methods because they are either too low in concentration (<1%), or nano-crystalline/amorphous or sorbed onto other mineral phases. SEPs consist of a sequence of leaching steps each of which targets specific phases. All SEPs are operationally defined, meaning that the reagent used in each step is optimized to target, or select, metals retained by specific phases. The main criticism of SEPs is that they are not selective essentially dissolving more than the intended phase, particularly in mine waste which contains a wide range of primary and secondary minerals. Another criticism is that there is potential for redistribution of metals during the procedure such that metals are incorrectly associated with phases and consequently the contributions of specific phases to metal attenuation can be over- or underestimated. Many of the SEP methods recommend the use of other mineralogical techniques in parallel to improve 13  phase identification (Caraballo et al., 2009; Dold, 2003; Hall et al., 1996; Ryan et al., 2008; Tessier et al., 1979). Several studies on a variety of natural and synthetic samples have shown that SEPs are never ideally selective, particularly for steps that target amorphous and crystalline iron oxides (Caraballo et al., 2009; Chao and Zhou, 1983; Dold, 2003; Hall and Pelchat, 1999; Larios et al., 2012). For example, Chao and Zhou (1983) found that the presence of magnetite in a sample catalyzed the dissolution of crystalline iron oxides during the amorphous iron oxide step when using oxalate as a reagent. The study concluded that an acidified hydroxylamine hydrochloride solution was the most desirable reagent for use on amorphous iron oxides due the close agreement with the results of extraction using oxalate (Chao and Zhou, 1983). The study also showed that Mn oxides are dissolved during the Fe oxide steps (Chao and Zhou, 1983). Caraballo et al. (2009) showed that Al hydroxides were also dissolved together with iron oxides when using oxalate as a reagent (Caraballo et al., 2009). Hall et al. (1996) proposed using hydroxylamine hydrochloride for the dissolution of both amorphous and crystalline iron oxides instead of oxalate because of the lack of selectivity of oxalate in the presence of magnetite when dissolving amorphous iron oxides, as described by Chao and Zhou (1983), and because the UV light set-up required for oxalate dissolution of crystalline iron oxides was considered to be too cumbersome. Dold (2003) compared hydroxylamine hydrochloride and oxalate in darkness for the dissolution of amorphous Fe oxides and Mn oxides concluding that both mineral phases were dissolved by these two reagents, but proposed oxalate for oxides in Cu-bearing sulfide tailings. Broadhurst et al. (2009) qualitatively compared the SEP proposed by Hall et al. (1996) to the method proposed by Dold (2003) using tailings material. Their results showed that there was relative consistency between the leachate results of the two SEPs; however, the study did not attempt to identify the minerals dissolved during the different steps or identify the phases from 14  which the elements leached and concluded that more research was required to identify the phases and mechanisms. While there is no doubt that oxides are excellent scavengers for trace metals in mine impacted sediments and tailings in the presence or absence of carbonates, studies involving carbonate waste rock where there is a focus on determining the potential for metal attenuation via precipitation of metal carbonates or sorption onto carbonate mineral phases have not been carried out. This is partly because SEPs are typically designed such that metals extracted via carbonates and cation exchange/weak sorption are assessed in a single step, although these are two very different attenuation mechanisms. The objective of this study was to identify selectivity issues in two SEPs in weathered waste rock containing As, Fe, Cu, Pb, Zn, and Mo bearing sulfides and carbonates. The two methods used, Method 1 and Method 2, were modified from Hall et al. (1996) and Dold (2003), respectively. These two published SEPs were selected for this study as they targeted the same phases, in the same order, using different reagents, and because the methods could be modified to include supplementary steps to address additional expected attenuation mechanism for carbonate bearing waste rock material. The criteria used to evaluate the two SEPs were: 1) Selectivity (i.e., the ability of a reagent to dissolve only the intended phases); and, 2) Retention of metals in the extraction solution (i.e., minimal confounding effects due to secondary mineral formation with the extraction reagents). Our study used heterogeneous waste rock material with a range of acid generating and acid neutralizing potentials and a range of leaching potentials for As, Cu, Pb, Mo, and Zn to compare two SEP methods modified from published SEPs. These SEPs were designed to separate metals attenuated by secondary water soluble phases, cation exchange/weak sorption, weak-acid soluble 15  phases (targeting carbonates), metals sorbed to amorphous iron oxides and metals sorbed to crystalline iron oxides from metals associated with primary silicate and sulfide minerals. Using observed leachate chemistry in tandem with mineralogical analyses of the residuals after each step, our study identified not only selectivity issues with respect to oxide phases but also with respect to phases that are weak-acid soluble.  2.2 Site and Sample Collection The Antamina Mine is located approximately 270km North of Lima, Peru. The climate at Antamina is bimodal with two distinct annual seasons; a wet season and a dry season. During the wet season approximately 80 - 90% of the total annual precipitation (~1200-1300mm) occurs. During the dry season, rainfall is limited and evaporation is high (Peterson, 2014). The mine exploits Cu-Zn-Mo (Bi, Pb and Ag) from a skarn deposit hosted in limestone (Lipten and Smith, 2004; Love et al., 2004; Redwood, 1999) and exhumes limestone, marble, hornfels, skarn, and intrusive waste rock with a range of neutralization/acid-generating potentials and metal contents during the mining process. The waste rock samples used in this study were collected from boreholes drilled in the East Dump, one of the operating waste rock dumps at the mine. Although Antamina segregates waste rock into specific dumps based on reactivity and metal content the East Dump is designed to accept all types of waste rock, and diverts seepage from the dump into the tailings pond where it is treated as part of the water quality management program. Figure 2.1 shows the drilling locations that provided the samples for this study (Site 1 and Site 3). At the time of drilling, the deepest material had been in place for over ten years, and the shallowest material for more than 5 years. A total of four boreholes (2 per site) were drilled using air-driven reverse circulation drilling, in which no drilling fluids were used to minimize the potential for alteration/dissolution of secondary phases. At each site, waste rock was collected from a deep 16  hole (>100m) and a shallow hole (~20m). To obtain the waste rock material, the drill casing was advanced in 1.5m intervals and waste rock drill cuttings were blown up the hole by compressed air and collected in trays. The drill cuttings from each 1.5m interval were sieved using a 2mm (#10) mesh and 500g of the passing material was saved for this study. Table 2.1 lists the eight samples selected for this study along with lithology, visible sulfide mineralization, and visible secondary minerals. Prior to testing, the samples were air dried at room temperature in a fume hood for several days. The dried sample was then ground to a fine powder in a swing mill for ~30s to further homogenize the samples before application of the SEPs. Rao et al. (2008) presented a review on the effects that sample pre-treatment had on extraction results, summarizing that air-drying at low temperature had a minimal impact on the results of extractions. Grinding may have an effect on metal extractability by increasing the availability of phases; however, in this study this was not investigated. 2.3 Methods 2.3.1 X-Ray Diffraction (XRD) Mineralogy of Samples X-ray diffraction (XRD) was used to identify the most abundant crystalline minerals in the samples prior to the SEPs and in residues collected after the application of each step of the SEPs. To prepare the samples for XRD analysis, the samples were ground into a powder-slurry using a mortar and pestle and ethanol. The slurry was then smeared onto a glass slide. XRD data were collected using a Bruker D8 Focus Diffractometer with a scanning step of 0.029 2θ and counting time of 100.1s over a range of 3-80 2θ. Mineral phases in the X-ray diffractograms were matched to mineral phases using the International Centre for Diffraction Database PDF-4 and Search Match software by Bruker. 17  2.3.2 Total Elemental Composition Total elemental composition of each sample was determined by 4-acid digestion of an approximately 0.26g of sample (after homogenization and weighing) and analysis by inductively-coupled plasma – optical emission spectroscopy (ICP-OES) and inductively-coupled plasma – mass spectroscopy (ICP-MS) at SGS, Burnaby, Canada. Sulfur speciation and total inorganic carbon were analysed at ALS, Peru. 2.3.3 Sequential Extraction Procedures and Comparison For both SEP methods, each sample was prepared in the same manner: Eight splits of the sample were taken; 3 splits to be leached using the full SEP to calculate an average leachate chemistry for each step and 5 splits to be leached using a “parallel” extraction method to obtain post-step residues to be analysed using XRD. The “parallel” extractions were designed such that the first of the five splits would undergo only the first step and its residue analysed using XRD, the second of the five splits would only undergo the first and second step and its residue analysed using XRD, and so on for the rest of the splits/steps. The residues were air dried and prepared for XRD analysis as described above. The two methods, referred herein as Method 1 (modified from Hall et al., 1996) and Method 2 (modified from Dold, 2003), are presented in Table 2.2. The main differences between the two methods are the reagents used to dissolve weak-acid soluble phases (Step 3) and amorphous and crystalline iron oxides (Step 4 and Step 5, respectively), as well as the reaction time and temperature used for these steps. The two SEPs were modified by including additional steps, but no modifications were made to the specific steps prescribed by each method. The SEPs were modified based upon previous metal attenuation studies at Antamina (Hirsche, 2012; Peterson, 2014): 1) to include a water-soluble phase extraction step in Method 1; and, 2) to use 18  two steps instead of one to separate metals attenuated by weak sorption/cation exchange and weak-acid soluble phases, which are typically extracted together, in both SEPs. The reasons for the modifications are based on anticipated attenuation mechanisms, which were: 1) Precipitation of water soluble sulfates – Equilibrium geochemical modelling of seep water from locations downslope of the drilling sites indicate that gypsum is at equilibrium (Laurenzi, Chapter 3). Studies of waste rock in semi-arid to arid climates where evaporation is high show that the formation of efflorescent water soluble metal salts is possible (Carbone et al., 2013a; Smuda et al., 2007). Accordingly, a water soluble step would provide support for the model-predicted gypsum equilibrium and identify if additional water-soluble metal salts precipitated. 2) Attenuation via weak sorption/cation exchange or carbonates – Cation exchange onto clay minerals was proposed as a potential attenuation mechanism for Zn in Antamina waste rock (Hirsche, 2012). Smithsonite (ZnCO3) was identified using XRD with Rietveld refinement on waste rock collected from a field barrel after 1 year of weathering (Dockrey, 2010). Equilibrium geochemical modelling of seep water from locations downslope of the drilling sites indicate that copper-sulfates such as brochantite (CuSO4·3Cu(OH)2) and antlerite (Cu3(SO4)(OH)4) are generally undersaturated and supersaturated at times when the pH is between 7 and 5. The geochemical modelling also shows that malachite and smithsonite are undersaturated. Furthermore, a blue, mostly amorphous, precipitate containing some malachite and gypsum was observed associated with an experimental test pile composed of intrusive material (Peterson, 2014). Using synchrotron-based X-ray adsorption near-edge structure analysis (XANES), copper in the amorphous precipitate was determined to be bonded predominantly with sulfate in the 19  form of brochantite and some carbonate in the form of malachite (Peterson, 2014, personal communication M. Lindsay, Univ. Saskatchewan). Accordingly, a separate extraction step that removed metals bound by weak electrostatic forces and cation exchange (Step 2) was added to distinguish between this attenuation mechanism and metals associated with carbonates (Step 3). The extraction reagent (MgCl2) used for Step 2 was based on established methods documented in the literature (Tessier et al., 1979). Step 6 of each SEP was conducted by a commercial lab (SGS, Burnaby) using the four-acid digest method as described previously for total elemental composition. All reagents were prepared the day of, or one day prior to, testing. Approximately 1g of solid was weighed in a 50ml Falcon ™ tube and reacted in the sequence prescribed by the methods, see Table 2.2. After the reaction time was complete, the tube was centrifuged at ~2500RPM and the leachate was decanted into a syringe and filtered using a 0.45µm filter. The residue was rinsed with 5ml of de-ionized water (DI) and centrifuged again; the rinse water was then added to the leachate. The rinse step was then repeated. The leachates were analysed for Al, As, Ca, Cu, Fe, Mn, Mo, Pb, S, Si, and Zn by ICP-OES. Since the ICP-OES reports chemistry in mg/L of liquid and because the steps of the SEPs use different liquid:solid ratios, leachate concentrations were converted to ppm. The detection limits of the ICP-OES are generally 0.2mg/L; however, based on the volume of reagent used in each extraction step this resulted in varying detection limits in ppm (presented in Table 2.2). Element concentrations below the detection limits are reported herein as <DL. The average leachate chemistry for each step was calculated from the leachate concentrations (ppm) of the three splits of each sample that underwent the full extraction sequence. For each average leachate chemistry calculation, when 20  the detection limit was encountered a zero value was used. Relative percent difference (RPD) (USEPA, 2010) was used to compare the average leachate chemistries between the two SEPs, using the equation:  𝐑𝐏𝐃 =  |[𝐱𝟏−𝐱𝟐]|?̅?× 𝟏𝟎𝟎          2-1  Where x1 and x2 are the average leachate chemistry for a particular step in Method 1 and Method 2, respectively, and ?̅? is the average of the two. RPD is typically used in two ways; to calculate the precision from duplicate measurements and to compare two measured values when an exact (true) value is not known (USEPA, 2010). In this study it was used based on the assumption that if the two methods are targeting the same attenuation mechanism/secondary phases per step then the leachate chemistries between the two methods should be similar. If they are not, then either additional minerals are being dissolved or precipitated or there is carry-over of metals between steps due to the reagent used or incomplete dissolution of target phases. In this study an RPD of greater than 30% was used as an indication that there was a significant difference in the leachate chemistry of the steps being compared. The first two steps of both methods were identical in reagent, and leach time, thus comparison of the leachate results was also used as an indication that sampling biases were minimal in the splitting of samples to make replicates for the investigation; this is discussed further in the results section. 2.3.4 Method Quality and Inter-sample Variability and Sample-split Variability Appendix B presents measurements that were taken during the extraction steps to ensure solids were not lost during the manipulations of the SEPs (mass loss) and that the pH of the 21  carbonate extraction step (Step 3) remained at the targeted pH (pH drift). Calculation of cumulative leached versus total metal concentration were made to also determine if mass was conserved during the SEPs. Relative standard deviation (RSD) for the triplicate analyses for Al, As, Cu, Fe, Mg, Mn, Mo, Pb, Si and Zn were made to determine sample/split variability.   2.4 Results 2.4.1 XRD Mineralogy of Samples Table 2.3 lists the initial, pre-SEP sample mineralogy identified from the XRD diffractograms. The minerals found in the samples are consistent with the lithologies noted in the drill logs (Appendix A) and the mineralogies reported previously in “fresh” waste rock (Peterson, 2014). Most samples were mineral mixtures typical of the dominant rock type noted in the drill logs and the other lithologies also noted in the drill logs. The primary minerals in marble samples were predominantly calcite, but they also contained quartz, orthoclase, albite and biotite from igneous intrusive rock. The igneous intrusive samples contained quartz, orthoclase, albite, biotite, muscovite, pyrite, chalcopyrite, and molybdenite as well as calcite from marble and limestone. The primary minerals of the exoskarn waste rock were the garnet minerals hibschite and andradite, vesuvianite and wollastonite along with quartz, orthoclase, albite, biotite and calcite from igneous intrusive and limestone/marble waste rock, respectively. The secondary minerals that were identified in the samples were gypsum, hemimorphite, smithsonite, and wulfenite. It should be noted that while hemimorphite and smithsonite are possible secondary minerals that can form from the oxidation of zinc-bearing sulfide minerals, at Antamina these minerals are also associated with the supergene mineralization of the deposit (Personal communication 2013, L. Plascencia) and therefore may have been placed in the dump during 22  construction rather than have been formed in the dump as a secondary phase. Hemimorphite was also noted in “fresh” skarn material (Peterson, 2014), further suggesting that it may have been present pre-mining.  2.4.2 Bulk Elemental Analysis The results of the bulk chemistry from the four-acid digestions of the samples are presented in Table 2.4. The samples contained total Cu ranging between 1020 ppm – 18600 ppm. Total Zn contents in the samples ranged between 473 ppm – 8290 ppm. Arsenic ranged between 45 ppm – 193 ppm and Mo between 24.5 ppm – 331 ppm. Lead concentrations in the samples ranged between 50.1 ppm – 874 ppm. Sulfur speciation and total inorganic carbon (TIC) content are presented in Table 2.5 along with the calculated acid-potential and carbonate neutralization potential of each sample. Samples that were composed of primarily igneous intrusive waste rock contained between 2.9 and 14.9 % total sulfur and between 1.3 and 1.8% TIC. The samples that were composed of primarily marble contained between 0.3 – 0.6% total sulfur and between 3.6 – 9.6% TIC. The sample that contained primarily marble diopside had 1.5% total sulfur and 2.9% TIC and the green garnet exoskarn sample had 0.8% total sulfur and 4.9% TIC. For all of the samples, greater than 86% of the total sulfur was in the form of sulfide. Of the eight samples, five would be characterized as non-acid generating, one (BH3-2) would be characterized as uncertain and two (BH1-2 and BH3-4) would be characterized as potentially acid generating (Price, 2009). 2.4.3 Sequential Extractions and XRD Mineralogy In this section, the average Ca, Fe, Cu, Pb, Zn, As and Mo leachate results for each step of Method 1 and Method 2 for select samples are compared qualitatively using % leached figures, quantitatively using calculated RPD for all samples, and XRD analysis of post-step 23  residues on select samples. In each XRD analysis figure, the bottom diffraction pattern is the un-treated sample and the above diffraction patterns correspond to Step 1 through Step 5, in order. Appendix C presents the % leached Al, Ca, Fe, Mn, As, Cu, Mo, Pb, and Zn for all samples as well as the XRD analyses for both methods for all samples.  Generally, from the figures indicating the % leached it appears that there is relative consistency between Methods 1 and 2 with respect to the % leached Fe (Figure 2.2), and, As and Mo (Figure 2.3). Figure 2.2, shows that there are notable inconsistencies with % leached Ca specifically during Step 4 and Step 5, where Ca is found in the leachates from Method 1 and not Method 2. It also appears that there are inconsistencies with % leached Cu, Pb, and Zn (Figure 2.4) during Step 3 and Step 4, where more metals are leached from Method 1. In-line with the conclusions of Broadhurst et al. (2009), from this qualitative assessment it is not possible to attribute the inconsistencies in Ca, Cu, Pb, and Zn to selectivity issues. Thus, these inconsistencies are partially addressed in the sections below using direct comparison of the leachate chemistries aided by XRD mineralogy of the residuals. The focus of the comparison is on Step 3 through Step 5 in which the reagents, reaction time, and temperature differ between methods.  2.4.3.1 Extraction Results Step 1 – Water Soluble Phases The water-soluble extraction leachate concentrations, shown in Table 2.6, were similar between both methods, with all RPD values less than 15%. The concentrations of most elements were low or below detection limit. There was measurable Ca, Si and S (S not presented) in most samples; however, Si was very close to the detection limit in most samples. Ca and S made the bulk of the leached elements and were calculated to have an approximate 1:1 molar ratio, which is consistent with the dissolution of gypsum.  24  Samples BH1-1, BH1-3, and BH3-1 leached the lowest concentration of Ca during this step and gypsum was not detectable in the XRD patterns for these samples. The rest of the samples leached higher Ca in this step and had detectable gypsum in their XRD patterns. It can be seen for sample BH1-4 in Figure 2.5 and Figure 2.6 for Method 1 and Method 2, respectively, that after Step 1, gypsum was no longer detected in the residues suggesting that gypsum was successfully removed during this step.  2.4.3.2 Extraction Results Step 2 – Metals Bound by Weak Electrostatic Forces and Cation Exchange The MgCl2 reagent used in this step is meant to remove only those elements that are bound by weak electrostatic forces and cation exchange, and is not intended to dissolve a solid phase. Table 2.7 shows the average leachate concentrations for both SEPs and calculated RPDs for Step 2. RPD values were generally less than 30%. RPD values greater than 30% were due to values at or near the detection limits of the ICP-OES where error is high. The concentrations of elements in the leachates from this step were generally near or below detection limit, with the exception of Ca, Mn, Si, Cu, and Zn in most samples, and Mo, and Pb in a few samples. 2.4.3.3 Extraction Results Step 3 – Weak Acid Dissolvable Phases From Table 2.2, the reagents used for Method 1, Na-CH3COO – sodium acetate at pH 5.0, and Method 2, NH3-CH3COO - ammonium acetate at pH 4.5, differ in pH, counter ion and leach time. Table 2.8 presents the average leachate concentrations for the two SEPs and the calculated RPDs. The RPDs for calcium, manganese and copper were good, below 30%, which is an indication that both methods dissolved similar weak-acid soluble Ca, Mn, and Cu bearing phases. However, the RPDs for Fe, Si, Pb, and Zn were greater than 30% for many samples, while Mo and As were below detection.  25  For both methods, XRD analysis of the residuals of the step showed that calcite was successfully dissolved from each sample. Also noted in the XRD pattern was the dissolution of hemimorphite (Zn4Si2O7(OH)2·H2O) during this step during application of both methods, Figure 2.5 and Figure 2.6, respectively. Also seen in Figure 2.5 and Figure 2.6, after the dissolution of the carbonate fraction in Step 3, the relative proportion of the remaining mineral phases increased, allowing for the identification of the scapolite mineral meionite (Ca4Al6Si6O2*4CO3) in a sample where scapolite was noted in the drill logs, merwinite (Ca3Mg(SiO4)2), and wollastonite (CaSiO3), which is part of the Antamina porphyry-skarn assemblage (Lipten and Smith, 2004). It is unlikely that these minerals were precipitated during the extractions. 2.4.3.4 Extraction Results Step 4 – Amorphous Reducible Phases Generally, the concentrations of elements extracted by the two methods at this step were not similar and the RPDs were >30%, with the exception of Fe (See Table 2.9). However, the concentrations of Fe in the leachates from this step were generally higher in Method 1 than in Method 2, which could be due to the enhanced dissolution of Fe oxides during Step 3, when using Method 2, as discussed above.  XRD diffractograms for residues from both methods showed that wollastonite was dissolved or partially dissolved in this step, Figure 2.7 and Figure 2.9 for Method 1 and Figure 2.8 and Figure 2.10 for Method 2. After Step 4 of Method 2, the post-step residue XRD patterns for BH1-1 (Figure 2.8), BH1-2, BH1-3 (Figure 2.10), BH1-4 (Figure 2.6), and BH3-3 showed three peaks at 17.347, 17.746 and 29.263 (2θ), that were not identified in the Method 1 post-step residues. The mineral was identified as whewellite (CaC2O4·H2O – calcium oxalate). The identification of a Ca-oxalate precipitate suggested the potential for precipitation of other metal oxalates but at concentrations too low to be identifiable with XRD, the possibility of which is 26  examined in the discussion section. Since both methods dissolved calcite in the previous step, dissolution of wollastonite is considered the main source of Ca in the leachates at this step. A nearly 1:1 Ca:Si molar ratio calculated in the Step 4 leachates from Method 1 is used as a proxy to support this assumption, with the exception of BH1-1 and BH1-3 which had Ca:Si molar ratios of 16 and 13, respectively, suggesting that other processes may have played a role in the case of these samples. Close examination of the XRD results did not provide conclusive information on the dissolution or formation of other mineral phases. 2.4.3.5 Extraction Results Step 5 – Crystalline Reducible Phases Table 2.10 presents the average leachate concentrations for both methods and the calculated RPDs for the comparison of this step. Similar to the results of Step 4, the leachate concentrations were not comparable and most RPDs were >30%.  The only observed changes in the mineralogy of the post-step residues from this step were found in the samples that underwent Method 2. The whewellite peaks became more prominent in the samples previously noted to have precipitated whewellite and whewellite was detected in BH3-4 (shown in Appendix C) indicating that additional precipitation of this phase occurred during Step 5. A nearly 1:1 Ca:Si molar ratio calculated in the Step 5 leachates from Method 1 is used as a proxy to support the assumption that dissolution of wollastonite during this step supplies the Ca for the precipitation of whewellite as was observed in samples treated by Method 2. 2.4.3.6 Cumulative Extraction Results for Steps 1 - 5 The sum of concentrations for Step 1 through Step 5, for each method, and the calculated RPDs for these sums are presented in Table 2.11. The RPDs for Al, Ca, Fe, Mg, Mn, Pb, Si, and Zn were generally below 30%. The elemental releases of Cu, Pb, Zn were generally higher for 27  Method 1 and the elemental releases for As and Mo were higher for Method 2.The low-moderate RPDs in the cumulative analysis indicated that each method dissolved similar total amounts of minerals, and suggested that the differences in the comparison of leachates is due to selectivity rather than sample variability. For example, cumulative Fe leached for both methods were similar suggesting that similar phases were dissolved but the Fe phases dissolved were distributed over Steps 3 - 5 differently for each method.  2.5 Discussion The reagents and reaction times for Step 1 and Step 2 of both methods were the same, the average leachate chemistries were characterized by low RPDs and XRD results for both extraction methods were similar, suggesting sample-split variability was low. The leachate chemistry of the extractions and XRD mineralogy showed that gypsum was the only mineral dissolved in Step 1. Based on these results, a water soluble extraction step should always be included in an SEP for carbonate bearing waste rock as gypsum is a typical mineral phase that controls the concentration of Ca and S in pore water. The leachate chemistries from Step 2 of both methods had measureable Cu and Zn in some samples but these concentrations were low compared to the concentrations leached from subsequent extraction steps. The two methods documented in the literature (Dold, 2003; Hall et al., 1996) include cation exchange/weak sorption and carbonate phases in a single step; however, our study showed through modification of these methods that in carbonate bearing waste rock, metals are more likely to be associated with weak-acid soluble phases rather than cation exchange/weak sorption. Thus, modifying each method to include a weak sorption/cation exchange step (Step 2) allowed metals bound by this attenuation mechanism to be discriminated from metals bound in weak-acid soluble phases (Step 3). 28  Step 3 of both methods did not selectively extract carbonates. While the RPDs for Ca, Mn and Cu were acceptable, the RPDs for Fe, Pb and Zn were >30% for most samples. Two possible reasons for Fe in the leaches are dissolution of siderite (FeCO3) and reductive dissolution of low-order amorphous iron oxides in the presence of acetate. The concentrations of Fe ranged between 75ppm and 1500ppm for Method 1 and between 220ppm and 1800ppm for Method 2, thus higher concentrations of Fe were reported in the leachates from Method 2 as compared to Method 1. Heron et al. (1994) tested whether sodium acetate at pH 5 (same as Method 1) would selectively dissolve siderite and found that siderite was not sufficiently soluble in the presence of this reagent. Caraballo et al. (2009) used ammonium acetate at pH 4.5 (same as Method 2) in an SEP to dissolve poorly crystalline Fe - phases and found that this reagent dissolved schwertmannite; however, this study did not specify if siderite was also present in the samples and if so, whether it dissolved. Based on this information, it can be concluded that the reagent used in Method 1 is unlikely to dissolve siderite, if present, and that iron oxides are more probably sources for Fe released. Method 2 has been demonstrated to dissolve iron oxides, and their enhanced dissolution may also explain the enhanced release of Fe and potentially sorbed metals. The lower pH value of this reagent may also be more favorable to promote the dissolution of siderite, if present. While siderite was not noted in the XRD patterns in the samples collected for this study, fresh waste rock samples of igneous and skarn lithologies at Antamina had between 0.1 – 0.3% siderite (Peterson, 2014). BH1-1 and BH1-3 from the current study were composed of predominantly marble and marble diopside lithologies, thus are not expected have siderite, in both cases 3x and 4x, respectively, more Fe was noted in the leachates from Method 2 than Method 1 suggesting that iron oxides or additional Fe-bearing phases are being dissolved by both reagents but more so in Method 2. 29  Step 4 and Step 5 of both methods did not selectively dissolve iron oxides. The % leached Ca and leachate comparison of Ca for both steps showed that more Ca was found in the leachates from Method 1. The XRD results showed that for both methods wollastonite dissolved during these steps. A nearly 1:1 molar ratio of Ca:Si was calculated for the leachates from Step 4 and Step 5 of Method 1 for most samples, supporting the dissolution of wollastonite. However, other processes are suggested but not confirmed from samples that leached Ca:Si ratios greater than 1. In contrast to Method 1, the calcium concentrations in leachates from Step 4 of Method 2 were low as an artifact from the precipitation of whewellite (Ca-oxalate). To gain insight into the potential for precipitation of other oxalate minerals during this step, a PhreeqC (Parkhurst and Appelo, 2013) simulation was developed to mimic Step 4 of Method 2. Mineral saturation indices (SI) were investigated for metal oxalates using the solubility product constants (Ksp; at 25 °C) of calcium oxalate and other metal oxalates presented in Table 2.12. The PhreeqC input file and results are presented in Appendix D. The simulation results indicated that precipitation of Ca-oxalate and metal oxalates was possible (SI > 1). When these minerals are allowed to precipitate, in all scenarios modelled, Ca-oxalate and Zn-oxalate precipitated and in one case Cu-oxalate precipitated as well. Although Ca-oxalate was the only oxalate phase noted in the XRD patterns of the residues from Step 4 and Step 5 of Method 2, the leachate concentrations and % leached Cu, Pb and Zn were generally lower in Method 2 than in Method 1, which is consistent with the geochemical modelling predictions of precipitation of trace metal oxalates during this extraction step.  While previous studies show that oxalate in the presence of Fe2+ can catalyze the dissolution of iron oxides, this study shows that both oxalate and hydroxylamine hydrochloride will also dissolve wollastonite, which has not been previously reported, and in the case of an 30  oxalate-based reagent can cause the precipitation of Ca-oxalate minerals and potentially other metal oxalates. Our study also shows that while there are relative consistencies in the leachate chemistries from each step, as was concluded by Broadhurst et al. (2009), there are instances where a non-targeted phase was being dissolved as shown by the presence of an element in the leachate of one method (i.e., dissolved Ca in Step 4 of Method 1) but masked by the precipitation of a mineral phase in the other method (i.e., precipitation of whewellite, Method 2). The use of XRD on post-step residues allowed for this determination. Sampling biases were partially addressed above, where the leachate comparison between Step 1 and Step 2 of the two methods show low variation (high precision) suggesting that sampling biases were minimal. Thus the differences between the two methods in Step 3 are due to non-specific dissolution of iron oxides and the differences in Steps 4 – 5 are due to precipitation of metal oxalates in Method 2. Sampling biases were also addressed using a calculation of relative standard deviation (RSD) of triplicate analyses for both methods (in Appendix B). The RSDs for triplicate analyses were better for Method 1 than Method 2 but most were generally less than 10% with some between 10 % and 30%; higher RSDs (>30%) were typically elements that were low concentration and near the detection limit of the ICP-OES and in Steps 4 and 5 of Method 2 where precipitation of metal oxalates will have affected the overall chemistry of the leach solution. 2.6 Conclusions Using heterogeneous samples of waste rock collected from the East Dump at the Antamina Mine, two sequential extraction procedures were compared to identify selectivity issues, if present in the two methods, and to ultimately choose one procedure that would be appropriate for a larger study on metal attenuation. To investigate the two methods the leachate 31  chemistries of equivalent steps of each procedure were compared using RPD to quantify if the chemistries were similar and XRD was used to identify minerals dissolved or precipitated in residues collected after each step in “parallel” designed extractions. This study shows that using both qualitative and quantitative assessments of leachate chemistry along with mineralogy of residues from extraction steps can produce a meaningful assessment of selectivity in heterogeneous waste rock samples. The results show that both methods have selectivity issues with Step3, Step 4 and Step 5. Step 3 of both methods is designed to dissolve carbonate minerals and other weak-acid soluble minerals. Both methods appear to successfully dissolve all of the calcite and similar Cu bearing phases but both methods have unresolved selectivity issues with respect to the dissolution of Fe bearing phases, specifically siderite and amorphous iron oxides. The reagent used in Step 3 of Method 1 has been shown to be insufficient to dissolve siderite which would carry over to Step 4. The reagents used in both methods may also dissolve some amorphous iron oxides in this step but from the data the reagent used in Method 2 may cause the dissolution of more iron oxides than Method 1. Step 4 and Step 5 of both methods is designed to dissolve only reducible oxide phases such as amorphous iron oxides and crystalline iron oxides, however, both reagents used also dissolved wollastonite at these steps. Although both methods have selectivity issues with the reagents used for Step 4 and Step 5, Method 1 had the following advantage over Method 2: The hydroxylamine hydrochloride used in Step 4 and Step 5 of Method 1 did not further react with the metals leached into solution as did the oxalate used in Method 2. The precipitation of whewellite in Step 4 and Step 5 of Method 2 due to the dissolution of wollastonite and the use of oxalate as the reagent is an indication that oxalate is not an appropriate reagent for samples containing wollastonite. The identification of whewellite and lower concentrations of Cu, Pb and Zn in the Step 4 and Step 5 leachates of Method 2 are an 32  indication that oxalates of these metals may have also precipitated, although none were noted in XRD analysis. Geochemical modelling of this system supports this hypothesis.   33  Figures  Figure 2.1 View of East Dump and drilling location (Site 1 and Site 3)   (Image by Lorca-Ugalde, 2013) 34   Figure 2.2 Ca and Fe (% leached) SEP results for BH1-1, BH1-2, BH1-4 and BH3-3   Ca (%) Fe (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-1Step 1 Step 2 Step 3 Step 4 Step 5 residual35   Figure 2.3 As and Mo (% leached) SEP results for BH1-1, BH1-2, BH1-4 and BH3-3 As (%) Mo (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-1Step 1 Step 2 Step 3 Step 4 Step 5 residual36   Figure 2.4 Cu, Pb, and Zn (% leached) SEP results for BH1-1, BH1-2, BH1-4 and BH3-3  Cu (%) Pb (%) Zn (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-1Step 1 Step 2 Step 3 Step 4 Step 5 residual37   Figure 2.5 BH1-4 XRD results from Method 1    Gypsum Hemimorphite Calcite 38   Figure 2.6 BH1-4 XRD results from Method 2   Gypsum Hemimorphite Calcite Whewellite 39   Figure 2.7 BH1-1 XRD results from Method 1    Wollastonite 40   Figure 2.8 BH1-1 XRD results from Method 2    Whewellite Wollastonite 41   Figure 2.9 BH1-3 XRD results from Method 1    Wollastonite Wollastonite Wollastonite 42   Figure 2.10 BH1-3 XRD results from Method 2   Wollastonite Wollastonite Wollastonite Whewellite 43  Tables Table 2.1 Samples selected for this study Sample ID Borehole From To Major Lithology (>50%) Minor Lithology (<50%) Comments Sulfides Secondary minerals   (m) (m)      BH1-1 BH1-1 10.5 12 M  scapolites, some XV Cp, Bn, Py, Po  BH1-2 19.5 21 IQM M  Cp, Mo, Py FeOx BH1-3 BH-1d 64.5 66 M MDP some XV, XW, IQM Cp, Py, Po, Bi FeOx BH1-4 91.5 92.4 MDP XV some IQM Cp, Sp, Py, Po FeOx, Malachite BH3-1 BH-3s 1.5 3 M IQM  Cp, Py, Po FeOx BH3-2 15 16.5 IQM M  Cp, Py, Po FeOx, Malachite BH3-3 BH-3d 24 25.5 XV C  Cp, Py FeOx, Malachite BH3-4 37.5 39 IQM C   Cp, Py FeOx, Malachite NOTES:  C - limestone; IQM – Igneous intrusive; M – Marble; MDP – Marble Diopside; XV – Green Garnet Exoskarn; XW – Wollastonite Exoskarn; FeOX – visible iron oxide staining; Cp – Chalcopyrite; Bn – Bornite; Py – Pyrite; Po – Pyrrhotite 44  Table 2.2 Sequential extraction procedures (SEPs) used in this study Step Phases Method 1(1,4) Method 2(4) ICP-OES Detection Limit ppm of solid 1 Water Soluble 50mL deionized water,  shake for 1h 50mL deionized water, shake for 1h <10 ppm 2 Weakly sorbed /Exchangeable(2) 40 mL 1M MgCl2, 40 mL 1M MgCl2, <10 ppm Shake for 1 hour at room temperature Shake for 1 hour at room temperature 3 Carbonates 20mL 1M CH3COONa (sodium acetate) at pH 5,  shake for 6h, 20mL 1M CH3COONH4 (ammonium acetate) at pH4.5, shake for 2h, at room temperature   < 6 ppm repeat Step 4 Amorphous Iron Oxides 20mL 0.25M NH2OH*HCl (hydroxylamine hydrochloride) in 0.25 HCl placed in 60⁰C water bath for 2h, every 30min vortex contents, 20mL 0.2 M NH4-C2O4 (ammonium-oxalate) at pH3.0, shake for 1h in darkness, at room temperature  <6 ppm repeat Step, but heat for only 30 min 5 Crystalline Iron Oxides 30mL of 1 M NH2OH*HCl (hydroxylamine hydrochloride) in 25% CH3COOH (acetic acid), place in 90⁰C water bath for 3h, vortex every 20 min, 30mL 0.2 M NH4-C2O4 (ammonium-oxalate) at pH3.0, heat in water bath 80 °C for 2h <8 ppm repeat Step, but heat for only 1.5 hours 6 Residual Four Acid Digest(3) Four Acid Digest(3)  NOTES:  1) Steps 3 – 5 are repeated using the same liquid solid ratio (LSR) but extraction time was shortened. Leachates are analysed separately and the concentrations are summed. 2) Weakly sorbed/exchangeable step taken from (Tessier et al., 1979) 3) Residual fraction was determined at SGS, Burnaby (BC, Canada) 4) Method 1 modified from Hall et al. (1996) Method 2 Modified from Dold (2003) 45   Table 2.3 Qualitative mineralogy from XRD Mineral Mineral Formula BH1-1 BH1-2 BH1-3 BH1-4 BH3-1 BH3-2 BH3-3 BH3-4 Quartz SiO2 x x x x x x x x Orthoclase KAlSi3O8 x x x x x x x x Albite NaAlSi3O8  x x x x x x x Biotite K(Mg,Fe)3(AlSi3O10)(F,OH)2 x x x x x x x x Muscovite KAl2(AlSi3O10)(F,OH)2        x Calcite CaCO3 x x x x x x x x Hibschite Ca3Al2(SiO4)2(OH)4  x  x   x  Andradite Ca3Fe2(SiO4)3    x     Vesuvianite Ca10Mg2Al4(Si2O7)2(SiO4)5(OH)4  x  x   x  Wollastonite CaSiO3 x  x    x  Actinolite Ca2(Mg,Fe)5Si8O22(OH)2 x x    x x x Tremolite Ca2Mg5Si8O22(OH)2    x   x  Magnetite Fe3+2Fe2+O4      x   Pyrite FeS2  x  x x x x x Molybdenite MoS2     x x x x Chalcopyrite CuFeS2    x x x x  Sphalerite (Zn,Fe)S       x  Chlorite    x x  x  x Gypsum CaSO4  x  x  x x x Smithsonite ZnCO3    x     Wulfenite PbMoO4       x x Hemimorphite Zn4Si2O7(OH)2•(H2O)       x          46  Table 2.4 Elemental composition of via 4-acid digestion   units BH1-1 BH1-2 BH1-3 BH1-4 BH3-1 BH3-2 BH3-3 BH3-4 Al % 0.55 1.92 0.94 0.58 4.95 3.11 3.06 3.23 As ppm 46 45 65 94 98 86 193 55 Ca % >15 3.83 >15 10.1 10.3 8.9 24.4 5.76 Cu ppm 1170 6360 1020 4030 1250 18600 6320 3760 Fe % 0.63 6.74 1.07 6.15 2.05 7.78 4.88 23.4 Mn ppm 258 744 505 1090 569 610 1850 360 Mo ppm 24.5 50.2 42.1 107 213 124 331 154 Pb ppm 138 293 570 874 367 50.1 700 117 Zn ppm 1300 4600 1300 7310 1590 1090 8290 473  Table 2.5 Acid base accounting (ABA) static test results   BH1-1 BH1-2 BH1-3 BH1-4 BH3-1 BH3-2 BH3-3 BH3-4 Major Lithology M IQM M MDP M IQM XV IQM Minor Lithology  M MDP XV IQM M C C S(T) (%) 0.35 10 0.5 1.52 0.6 2.85 0.75 14.85 S(SO4) (%) 0.03 0.3 0.01 0.15 0.01 0.39 0.04 0.27 S(SO4) (%) 0.04 0.3 0.01 0.1 0.01 0.39 0.04 0.27 S(2-) (%) 0.31 9.7 0.49 1.37 0.59 2.46 0.71 14.6 TIC (%) 9.6 1.1 8.06 2.86 3.63 1.33 4.91 1.82 AP kg CaCO3/t 9.7 303 15.3 42.8 18.4 76.9 22.2 456 Carb-NP kg CaCO3/t 801 93 671 238 302 111 409 152 47  Table 2.6 Average concentrations and RPD values for Step 1 sample   Al_avg Ca_avg Fe_avg Mn_avg Si_avg As_avg Cu_avg Mo_avg Pb_avg Zn_avg ID Units ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm BH1-1 Method 1 <DL 406.1 <DL <DL 13.0 <DL <DL <DL <DL <DL Method 2 <DL 440.9 <DL <DL 14.3 <DL <DL <DL <DL <DL RPD - 8% - - 10% - - - - - BH1-2 Method 1 <DL 7726.9 <DL <DL 15.2 <DL <DL <DL <DL <DL Method 2 <DL 8651.4 <DL <DL 16.8 <DL <DL <DL <DL <DL RPD - 11% - - 10% - - - - - BH1-3 Method 1 <DL 574.7 <DL <DL 14.6 <DL <DL <DL <DL <DL Method 2 <DL 575.8 <DL <DL 16.5 <DL <DL <DL <DL <DL RPD - 0% - - 12% - - - - - BH1-4 Method 1 <DL 2196.3 <DL <DL 38.9 <DL <DL <DL <DL <DL Method 2 <DL 2232.2 <DL <DL 37.3 <DL <DL <DL <DL <DL RPD - 2% - - 4% - - - - - BH3-1 Method 1 <DL 338.3 <DL <DL <DL <DL <DL <DL <DL <DL Method 2 <DL 346.5 <DL <DL <DL <DL <DL <DL <DL <DL RPD - 2% - - - - - - - - BH3-2 Method 1 <DL 5155.2 <DL <DL 33.8 <DL <DL <DL <DL <DL Method 2 <DL 5263.6 <DL <DL 32.7 <DL <DL <DL <DL <DL RPD - 2% - - 3% - - - - - BH3-3 Method 1 <DL 936.7 <DL <DL 21.8 <DL <DL <DL <DL <DL Method 2 <DL 975.6 <DL <DL 25.4 <DL <DL <DL <DL <DL RPD - 4% - - 15% - - - - - BH3-4 Method 1 <DL 3483.6 <DL <DL <DL <DL <DL <DL <DL <DL Method 2 <DL 3562.6 <DL <DL <DL <DL <DL <DL <DL <DL RPD - 2% - - - - - - - - NOTES:   Indicates RPD >30% <DL denotes where the minimum detection limit of the OES was reported    48  Table 2.7 Average concentrations and RPD values for Step 2 sample   Al_avg Ca_avg Fe_avg Mn_avg Si_avg As_avg Cu_avg Mo_avg Pb_avg Zn_avg ID Units ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm BH1-1 Method 1 <DL 3901.8 <DL <DL 10.4 <DL <DL <DL <DL 14.5 Method 2 <DL 3683.6 <DL <DL <DL <DL <DL <DL <DL 13.1 RPD - 6% - - - - - - - 10% BH1-2 Method 1 <DL 3878.8 <DL 41.4 11.7 <DL 29.0 <DL <DL 67.0 Method 2 <DL 3502.7 <DL 38.7 <DL <DL 31.4 <DL <DL 79.7 RPD - 10% - 7% - - 8% - - 17% BH1-3 Method 1 <DL 3855.4 <DL <DL 15.3 <DL <DL <DL 13.6 13.6 Method 2 <DL 3884.6 <DL <DL 12.4 <DL <DL <DL 14.4 15.1 RPD - 1% - - 21% - - - 6% 10% BH1-4 Method 1 <DL 4622.5 <DL 16.0 39.2 <DL 11.6 17.2 <DL 29.4 Method 2 <DL 4165.9 <DL 14.5 33.4 <DL 10.9 14.5 <DL 32.7 RPD - 10% - 10% 16% - 6% 18% - 11% BH3-1 Method 1 <DL 4118.9 <DL <DL 32.7 <DL <DL <DL <DL <DL Method 2 <DL 3749.1 <DL 16.0 30.6 <DL <DL <DL <DL 16.8 RPD - 9% - - 6% - - - - - BH3-2 Method 1 <DL 5459.2 <DL 10.5 76.3 <DL 20.1 <DL <DL <DL Method 2 <DL 5278.9 <DL 22.8 70.8 <DL 27.4 <DL <DL 12.7 RPD - 3% - 74% 7% - 31% - - - BH3-3 Method 1 <DL 4948.9 <DL <DL 68.9 <DL 11.9 16.1 <DL 12.4 Method 2 <DL 5166.7 <DL 20.9 70.4 <DL 18.2 22.9 5.6 26.1 RPD - 4% - - 2% - 41% 35% - 72% BH3-4 Method 1 10.5 4542.0 <DL <DL 39.1 <DL <DL <DL <DL <DL Method 2 <DL 4164.2 <DL 13.1 29.9 <DL <DL <DL <DL 15.5 RPD - 9% - - 27% - - - - - NOTES:   Indicates RPD >30% <DL denotes where the minimum detection limit of the OES was reported   49  Table 2.8 Average concentrations and RPD values for Step 3 sample   Al_avg Ca_avg Fe_avg Mn_avg Si_avg As_avg Cu_avg Mo_avg Pb_avg Zn_avg ID Units ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm BH1-1 Method 1 16.3 315779.9 75.1 113.7 137.8 <DL 37.6 <DL 64.6 48.3 Method 2 19.4 310419.6 220.5 148.1 60.1 8.0 29.5 <DL 116.1 19.9 RPD 17% 2% 98% 26% 79% - 24% - 57% 83% BH1-2 Method 1 326.0 17355.9 1503.3 57.9 560.0 <DL 1747.3 <DL 308.0 1916.4 Method 2 364.5 15017.4 1814.0 75.9 344.7 <DL 1312.9 <DL 307.4 1776.3 RPD 11% 14% 19% 27% 48% - 28% - 0% 8% BH1-3 Method 1 50.9 277960.6 106.2 218.5 194.8 <DL 94.0 <DL 456.5 207.6 Method 2 83.6 272885.5 409.1 300.0 144.8 <DL 91.7 <DL 715.9 240.9 RPD 49% 2% 118% 31% 29% - 3% - 44% 15% BH1-4 Method 1 116.2 64057.9 263.6 227.9 1621.5 <DL 2400.8 <DL 624.2 7204.8 Method 2 158.9 72539.5 448.4 212.3 1419.3 <DL 2148.9 <DL 732.8 8664.1 RPD 31% 12% 52% 7% 13% - 11% - 16% 18% BH3-1 Method 1 108.7 81528.0 669.7 131.8 376.3 <DL 78.8 <DL 179.2 99.6 Method 2 156.4 100577.8 1221.1 172.3 188.4 <DL 56.0 <DL 282.6 54.0 RPD 36% 21% 58% 27% 67% - 34% - 45% 59% BH3-2 Method 1 147.6 71440.2 496.7 132.4 658.9 <DL 1707.1 <DL <DL 91.5 Method 2 217.5 79562.0 824.6 140.5 414.3 <DL 1695.8 <DL 25.0 128.5 RPD 38% 11% 50% 6% 46% - 1% - - 34% BH3-3 Method 1 86.0 120831.9 223.9 503.1 815.6 25.3 1578.6 <DL 617.1 1642.3 Method 2 150.8 123985.2 500.5 543.7 534.1 26.5 1684.3 <DL 813.3 2151.2 RPD 55% 3% 76% 8% 42% 5% 6% - 27% 27% BH3-4 Method 1 415.1 22644.9 629.2 49.0 570.4 <DL 305.8 <DL 23.8 64.9 Method 2 506.2 17852.9 818.0 72.6 342.9 <DL 242.9 <DL 59.0 26.1 RPD 20% 24% 26% 39% 50% - 23% - 85% 85% NOTES:   Indicates RPD >30% <DL denotes where the minimum detection limit of the OES was reported   50  Table 2.9 Average concentrations and RPD values for Step 4 sample   Al_avg Ca_avg Fe_avg Mn_avg Si_avg As_avg Cu_avg Mo_avg Pb_avg Zn_avg ID Units ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm BH1-1 Method 1 649.1 54078.2 1247.6 52.9 2356.9 15.9 117.4 <DL 96.1 126.9 Method 2 60.0 14.8 673.9 7.7 305.7 10.5 17.6 <DL 6.0 21.4 RPD 166% 200% 60% 149% 154% 41% 148% - 176% 142% BH1-2 Method 1 2027.3 2557.1 17802.5 110.4 2697.1 12.3 1217.0 <DL 395.8 2119.7 Method 2 839.8 60.9 13567.9 50.7 786.1 14.5 437.0 <DL 85.9 411.5 RPD 83% 191% 27% 74% 110% 17% 94% - 129% 135% BH1-3 Method 1 1578.3 48327.0 1754.2 69.0 2526.5 10.1 119.5 <DL 355.5 268.2 Method 2 172.2 21.3 1924.6 14.9 824.8 14.4 43.5 <DL 16.7 165.0 RPD 161% 200% 9% 129% 102% 35% 93% - 182% 48% BH1-4 Method 1 1639.6 3206.0 19277.9 88.1 2506.5 8.8 1307.8 <DL 835.1 3095.2 Method 2 611.2 67.3 16098.6 67.2 1484.6 34.8 1053.7 34.4 162.0 387.5 RPD 91% 192% 18% 27% 51% 120% 22% - 135% 155% BH3-1 Method 1 2321.4 2305.7 4832.8 175.8 1789.8 9.1 90.9 <DL 224.0 307.8 Method 2 548.2 47.1 7419.9 198.5 655.5 18.6 27.3 8.1 31.8 142.6 RPD 124% 192% 42% 12% 93% 68% 108% - 150% 73% BH3-2 Method 1 1533.7 1558.4 22572.5 150.0 1972.5 33.0 1512.7 <DL 58.9 136.1 Method 2 770.0 54.3 20961.4 138.7 1527.2 56.0 518.4 9.7 25.4 89.4 RPD 66% 187% 7% 8% 25% 51% 98% - 79% 41% BH3-3 Method 1 1636.3 12749.1 4901.6 158.0 6814.8 51.0 1245.7 7.9 456.6 1404.5 Method 2 427.2 36.4 4852.5 51.1 2762.5 70.2 340.1 14.8 25.6 929.7 RPD 117% 199% 1% 102% 85% 32% 114% 61% 179% 41% BH3-4 Method 1 5592.0 5217.3 5403.9 25.5 4856.6 15.9 494.4 <DL 93.9 101.3 Method 2 1458.7 38.5 5817.7 25.5 1356.3 25.2 141.3 6.2 17.4 100.5 RPD 117% 197% 7% 0% 113% 45% 111% - 137% 1% NOTES:   Indicates RPD >30% <DL denotes where the minimum detection limit of the OES was reported   51  Table 2.10 Average concentrations and RPD values for Step 5 sample   Al_avg Ca_avg Fe_avg Mn_avg Si_avg As_avg Cu_avg Mo_avg Pb_avg Zn_avg ID Units ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm BH1-1 Method 1 194.6 611.1 659.1 3.6 256.4 <DL 51.2 <DL 35.7 49.2 Method 2 438.0 25.6 1011.1 19.8 1033.8 12.2 27.5 <DL 26.7 38.9 RPD 77% 184% 42% 138% 121% - 60% - 29% 23% BH1-2 Method 1 1203.9 5075.4 5132.1 64.7 2661.9 <DL 546.9 <DL 73.6 591.6 Method 2 1870.2 36.4 9428.0 93.3 3299.7 <DL 544.1 <DL 131.4 1381.3 RPD 43% 197% 59% 36% 21% - 1% - 56% 80% BH1-3 Method 1 659.7 1208.2 1537.8 12.3 774.8 <DL 45.9 <DL 184.1 75.7 Method 2 1219.1 27.3 1519.3 24.8 1721.0 3.1 47.5 <DL 55.8 76.1 RPD 60% 191% 1% 67% 76% - 3% - 107% 1% BH1-4 Method 1 1011.9 7370.5 20792.5 58.2 3622.0 24.8 243.5 51.2 170.9 301.5 Method 2 1784.3 50.1 49717.3 65.2 5239.2 53.5 699.4 52.3 379.6 443.0 RPD 55% 197% 82% 11% 36% 73% 97% 2% 76% 38% BH3-1 Method 1 628.2 1129.2 2346.4 23.1 879.6 <DL 37.2 <DL 44.1 35.7 Method 2 2038.1 51.5 1762.6 56.0 3141.2 <DL 19.9 <DL 40.7 39.8 RPD 106% 183% 28% 83% 112% - 61% - 8% 11% BH3-2 Method 1 376.1 1080.2 5063.9 18.9 908.4 3.2 591.3 <DL <DL 66.1 Method 2 800.7 54.9 16258.6 81.6 3578.4 <DL 10.8 <DL 28.7 15.8 RPD 72% 181% 105% 125% 119% - 193% - - 123% BH3-3 Method 1 1184.7 6310.7 4031.9 64.8 2943.0 11.7 445.5 <DL 71.4 646.6 Method 2 2709.9 47.5 7094.6 112.5 4739.3 18.8 222.8 <DL 57.6 421.8 RPD 78% 197% 55% 54% 47% 46% 67% - 21% 42% BH3-4 Method 1 1190.1 1522.7 1237.2 10.6 1392.2 <DL 445.1 <DL 18.3 28.2 Method 2 5610.0 47.0 1894.9 10.1 4493.6 <DL 245.6 <DL 37.3 30.9 RPD 130% 188% 42% 5% 105% - 58% - 68% 9% NOTES:   Indicates RPD >30% <DL denotes where the minimum detection limit of the OES was reported   52  Table 2.11 RPD for cumulative concentrations of Step 1 through Step 5 sample   Al_avg Ca_avg Fe_avg Mn_avg Si_avg As_avg Cu_avg Mo_avg Pb_avg Zn_avg ID Units ppm ppm Ppm ppm ppm ppm ppm ppm ppm ppm BH1-1 Method 1 860.0 374777.2 1981.7 170.2 2774.5 15.9 206.2 <DL 196.4 239.0 Method 2 517.4 314584.5 1905.6 175.6 1413.8 30.7 74.6 <DL 148.8 93.3 RPD 50% 17% 4% 3% 65% 64% 94% - 28% 88% BH1-2 Method 1 3557.2 36594.1 24437.9 274.4 5945.9 12.3 3540.3 <DL 777.4 4694.7 Method 2 3074.5 27268.7 24809.9 258.6 4447.3 14.5 2325.4 <DL 524.7 3648.8 RPD 15% 29% 2% 6% 29% 17% 41% - 39% 25% BH1-3 Method 1 2288.9 331925.8 3398.2 299.8 3526.1 10.1 259.5 <DL 1009.6 565.0 Method 2 1475.0 277394.6 3852.9 339.7 2719.6 17.5 182.7 <DL 802.8 497.1 RPD 43% 18% 13% 12% 26% 54% 35% - 23% 13% BH1-4 Method 1 2767.8 81453.2 40334.0 390.2 7828.1 33.6 3963.7 68.5 1630.2 10631.0 Method 2 2554.5 79055.0 66264.4 359.1 8213.7 88.3 3912.8 101.2 1274.3 9527.3 RPD 8% 3% 49% 8% 5% 90% 1% 39% 25% 11% BH3-1 Method 1 3058.3 89420.1 7848.9 330.7 3078.4 9.1 207.0 <DL 447.3 443.0 Method 2 2742.7 104772.1 10403.5 442.8 4015.6 18.6 103.3 8.1 355.1 253.1 RPD 11% 16% 28% 29% 26% 68% 67% - 23% 55% BH3-2 Method 1 2057.4 84693.2 28133.0 311.8 3649.8 36.2 3831.2 <DL 58.9 293.7 Method 2 1788.2 90213.7 38044.7 383.6 5623.4 56.0 2252.4 9.7 79.2 246.5 RPD 14% 6% 30% 21% 43% 43% 52% - 29% 17% BH3-3 Method 1 2907.0 145777.2 9157.4 725.9 10664.1 88.0 3281.7 24.0 1145.0 3705.8 Method 2 3287.8 130211.5 12447.6 728.1 8131.7 115.4 2265.3 37.7 902.2 3528.8 RPD 12% 11% 30% 0% 27% 27% 37% 44% 24% 5% BH3-4 Method 1 7207.8 37410.5 7270.3 85.0 6858.2 15.9 1245.3 <DL 136.1 194.4 Method 2 7575.0 25665.3 8530.6 121.3 6222.6 25.2 629.8 6.2 113.8 173.0 RPD 5% 37% 16% 35% 10% 45% 66% - 18% 12% NOTES:   Indicates RPD >30% <DL denotes where the minimum detection limit of the OES was reported   53  Table 2.12 Solubility product constants for metal oxalates Mineral Formula Ksp (at 25°C) Calcium oxalate monohydrate CaC2O4× H2O 2.32×10-9 Copper(II) oxalate CuC2O4 4.43×10-10 Lead(II) oxalate PbC2O4 8.5×10-9 Magnesium oxalate dihydrate MgC2O4× 2H2O 4.83×10-6 Manganese(II) oxalate dihydrate MnC2O4× 2H2O 1.70×10-7 Zinc oxalate dihydrate ZnC2O4× 2H2O 1.38×10-9 NOTES: Ksp’s from (Olmsted and Williams, 2007)54  Chapter 3 Characterizing Trace Metal Attenuation and Secondary Phases in Carbonate Bearing Waste Rock Collected from the East Dump, Antamina Mine, Peru  3.1 Introduction Understanding the processes that attenuate metals is essential for reliable predictions of the quality of water emanating from waste materials (Al et al., 2000; Mayer et al., 2003). However, characterization of metal attenuation mechanisms and the identification of secondary minerals are difficult because secondary phases are often amorphous or nano-crystalline and in low abundance compared to primary minerals. In addition, attenuation mechanisms such as sorption require methods with low detection limits and smaller spatial resolution which can require time-consuming and expensive analyses.  The objective of this study is to identify processes that attenuate metals in waste rock collected from the East Dump, Antamina Mine, Peru. The East Dump at the Antamina Mine is a unique environment to study these processes because it hosts both acid-buffering and acid-producing waste rock (Dockrey, 2010; Peterson, 2014). The elements of interest in this study are arsenic (As), copper (Cu), molybdenum (Mo), lead (Pb), and zinc (Zn). Metal attenuation studies on environmentally impacted sediments and mine tailings are well documented; however, there are few studies of these processes in full-scale waste rock dumps (Carbone et al., 2013b; Smuda et al., 2007). Metal attenuation studies using waste rock from smaller-scale studies, such as humidity cells, field barrels and experimental piles are more common (Andrina, 2009; Conlan, 2009; Dockrey, 2010; Hannam, 2012; Hirsche, 2012; Peterson, 2014; Stockwell et al., 2006). 55  However, smaller-scale studies may not reproduce all the processes and overall complexity from full-scale heterogeneous systems. Secondary phases and metal attenuation in waste rock is commonly predicted from equilibrium modelling of leachate emanating from waste rock dumps, however, it can be challenging to compare the results from small-scale solid phase samples sourced from specific locations with results derived from leachate composition (Hochella et al., 1999; Petrunic et al., 2006). This is further complicated by the fact that thermodynamic data for complex phases that precipitate from concentrated solutions are often not well-established and as such are not included in common geochemical databases (Younger, 2000); however, identification of such phases by mineralogical analysis provides motivation for inclusion of these phases in the model database. In this study, metal attenuation is investigated at the scale of an operational dump where both geochemical and physical heterogeneity can affect the geochemical processes that attenuate metals. Processes investigated at the full scale are compared to those operating in smaller-scale experiments containing waste rock from the same site. The chemistry of leachate from seeps emanating from the dump associated with the waste rock sampling locations is set in context with the observed mineralogy from the samples collected from within the dump.  The first section of this chapter introduces the site, the sample collection and selection, and the methods used in this study. The second section describes the results from solid phase analyses and geochemical modelling. The third section provides a discussion of the results, including their significance for the study of drainage from waste rock and their relationship to previous studies. The final section concludes on the results and discussion of this chapter. 56  3.1.1 Site Information The Antamina mine, located approximately 270km NE of Lima, Peru in the department of Ancash, is considered one of the ten largest mines in the world (Antamina, 2015). The skarn deposit was formed by the intrusion of a quartz monzonite body into limestones and hosts copper, zinc, molybdenum, lead, silver and bismuth bearing ore minerals (Lipten and Smith, 2004; Love et al., 2004; Redwood, 1999). The deposit is situated in the Andes, at an elevation ranging between 4200 and 4700m.a.s.l. Antamina receives approximately 1200-1300mm precipitation per year, almost all as rain. 80% of the precipitation falls during the region’s six to seven month wet season (October – April) while the remaining 20% falls during the months of May to September, the dry season. The mean annual temperature at the mine site is approximately 5.0°C.  3.1.2 Sample Collection and Selection The waste rock used for this study was collected by taking advantage of boreholes drilled to monitor gases within an operating waste dump. A total of four boreholes were drilled: one shallow and one deep, drilled at each of two sites (Site 1 and Site 3, Figure 3.1) on the East Dump. The waste rock collected from the boreholes is broadly representative of the different lithologies at the site, and increases in age with depth where the deepest samples would have been in the dump for approximately 10 years at the time of drilling. The deepest borehole is at Site 3 and penetrates the entire depth of the waste rock dump (145m) into the underlying bedrock.  The East Dump receives both potentially acid-generating (PAG) waste rock and non-acid generating/acid-buffering (NAG) waste rock from a range of lithologies, i.e., limestone, marble, hornfels, exo- endo- skarn and intrusive, and is thus considered geochemically heterogeneous. 57  All water seeping from the East Dump is diverted and captured in the tailings pond where it is treated as part of Antamina’s water management plan. Air-driven reverse circulation (RC) drilling was used for all boreholes, minimizing the alteration and dissolution of secondary mineral phases that would be caused by drilling fluids. The drill casing was advanced in 1.5m intervals and the waste rock cuttings for each interval were blown out of the borehole pneumatically and collected in trays. The trays of drill cuttings from each 1.5m interval were homogenized and split for experimental testing, drill logging and acid – base accounting (ABA) tests. The samples used for the experiments in this study consisted of 500g of waste rock cuttings that had been sieved, passing a 2mm (#10) mesh (the -2mm fraction). The drill logs were produced by Antamina’s Geology Department from 2kg splits of each 1.5m interval of drill cuttings (see Appendix A).  In this study, the sample naming convention is: “BH” for borehole, number of site drilled (“1” or “3”), the depth of the borehole (“s” for shallow; “d” for deep), and the interval (in meters below ground surface) from which they were collected (example ID: BH-1s (0.0 – 1.5)). Based on visual inspection of oxidation products on the cuttings and the drill logs, 32 samples (from >150 intervals) were selected for this metal attenuation study. Table 3.1 lists the samples selected for testing along with the drill-log information of major lithology (lithology making up >50% of the sample), minor lithology (lithology making up <50% of the sample), other lithologies noted in the sample, sulfides, weathering products as well as the results of ABA testing and acid-rock drainage (ARD) classification (Price, 2009) of each sample.  3.2 Methods  Metal attenuation and secondary phases were investigated experimentally using both geochemical and mineralogical tests. The total elemental composition of the -2mm material 58  fraction was determined by inductively coupled plasma – mass spectrometry and -optical emission spectrometry (ICP-MS/-OES) analyses after sample digestion using 4-acids (i.e., HCl (hydrochloric acid), HNO3 (nitric acid), HF (hydrofluoric acid) and HClO4 (perchloric acid)) at SGS in Vancouver, Canada. X-ray Diffraction (XRD) was used to identify the crystalline minerals in the samples. Qualitative XRD data was collected for all samples and quantitative XRD using Rietveld refinement was collected for only two samples. Samples prepared for qualitative XRD were ground in mortar and pestle under ethanol. Samples prepared for quantitative XRD were reduced to <10μm by grinding under ethanol in a vibratory micronizing mill. XRD data were collected using a Bruker D8 Focus Diffractometer with a scanning step of 0.029 2θ and counting time of 100.1 s over a range of 3-80 2θ. Mineral phases in the X-ray diffractograms were matched to mineral phases using the International Centre for Diffraction Database PDF-4 and Search Match software by Bruker. Quantitative mineralogical data was obtained using the Rietveld method with TOPAS 4.2 software package (Bruker AXS). While XRD is useful for high abundance crystalline mineral phases, it has limited ability to detect minerals below ~0.5 - 1 wt.% abundance or those that are amorphous or nano-crystalline. Accordingly, a modified sequential extraction procedure (SEP) proposed by Hall et al. (1996) (Table 3.2) was used to selectively dissolve the secondary phases from a 1g sample. The SEPs were performed on duplicates of each sample and the average leachate concentration for each step is reported. The extraction solutions were analyzed by a Varian 725-ES ICP-OES. The SEP results are reported as % leached. Values below the detection limits of the ICP-OES, generally 0.2mg/L, are reported as <DL (i.e., less than the detection limit). Table 3.2 presents the detection limits for each extraction step of the SEP in units of ppm of solid. Seven samples were examined in thin section using transmitted and reflected light microscopy to identify secondary phases. 59  These phases were then examined using scanning electron microscopy (SEM) and energy dispersion spectroscopy (EDS). Thin sections for the SEM were coated with evaporated carbon and examined on a Philips XL30 SEM equipped with a Bruker Quantax 200 Microanalysis system and a light element XFLASH 4010 Silicon Drift detector. Secondary phases were assumed to occur as coatings on mineral surfaces and/or rock particles, and as discrete mineral grains of phases that were not part of the primary mineral assemblage noted in “fresh” waste rock. A semi-quantitative analysis of the wt.% abundance of elements associated with these phases was determined using the EDS on spot locations.  Figure 3.1 presents the location of the drill sites on the dump, the locations of monitored seeps found at the toe of the slopes below the drill sites, and the pre-dump topography in which natural catchments are inferred for drainage from waste rock dumped at Site 1 and Site 3. These seeps drain into the tailings ponds where they are managed as part of the water quality management program. Site 1 is in the catchment in which seeps CO-41 and CO-57 were located and Site 3 is in the catchment in which seeps CO-28 and CO-56 were located. Seepage chemistry for each of these monitored locations was provided from December 2004-May 2012 for CO-28 and CO-41 and November 2011-May 2012 for CO-56 and CO-57. The seep CO-41 had a data gap between August 2008-April 2011. The saturation indices (SIs) of phases that control the seepage chemistry were determined with PHREEQC (Parkhurst and Appelo, 2013) using the WATEQ4F database modified to include Mo species data from the MINTEQ.V4 database. The database used was further modified to include the phases hydrozincite, Zn5(OH)6(CO3)2 (Preis and Gamsjäger, 2001), aurichalcite, (Cu,Zn)5(OH)6(CO3)2 and rosasite, (Cu, Zn)2(OH)2CO3 (Alwan et al., 1980), taking into consideration that the mineralogical studies indicated the presence of mixed Cu and Zn phases (discussed in the sections below). 60  3.3 Results 3.3.1 Lithology and Mineralogy Table 3.1 presents the logged lithologies, visible sulfides and secondary phases, and ABA results for the 32 samples investigated in this study. These data show that most samples were composed of a mixture of lithologies which is partly due to the drilling method used. The major lithology of most of the Site 1 samples was marble, with some samples composed mostly of marble diopside and igneous intrusive waste rock. The major lithology of most of the Site 3 samples was igneous intrusive, with a few samples that were marble, hornfels and exoskarn. Iron oxides were visible in most of the samples and a blueish mineral, logged as malachite, was visible in many of the samples (N=16). The degree of iron oxidation was qualified using a scale of 0 (none) - 4 (strong) and many of the samples had visible oxidation products. Using the ARD criteria from Price (2009), the ABA results show that most of the samples from each site were non-acid generating (NAG) (N=20); however, three samples from Site 1 were classified as potentially acid-generating (PAG) and one as uncertain while four samples from Site 3 were PAG and three uncertain. Table 3.3 summarizes the sulfide minerals and secondary minerals identified in the waste rock samples using XRD analysis. A detailed tabulation of mineralogy and XRD patterns for all the samples is presented in Appendix E. The minerals identified in the samples were consistent with minerals previously identified using XRD in homogenous “fresh” waste rock samples by Peterson (2014). The sulfide minerals pyrite (N=29), chalcopyrite (N=18), and molybdenite (N=17) were identified in many of the samples. Sphalerite was also identified, but in fewer samples (N=3). All of the samples contained some calcite; gypsum (CaSO4·2H2O) (N=23), hemimorphite (Zn4Si2O7(OH)2·H2O) (N=2), smithsonite (ZnCO3) (N=4), and wulfenite 61  (PbMoO4) (N=8) were also identified. Aside from hemimorphite and smithsonite, these minerals were not previously reported in “fresh” waste rock samples and were considered secondary. While hemimorphite and smithsonite may be secondary minerals that formed due to the weathering of Zn-bearing sulfides, they may also have been present prior to construction of the waste rock pile considering that they were found as part of the supergene mineralization of the Antamina deposit (Personal communication 2013, L. Plascencia). As such, hemimorphite and smithsonite cannot be unambiguously considered secondary minerals that formed in the dump. Although many of the samples had a “rusty” appearance, and iron oxides were noted in the drill logs, crystalline iron oxides were not identified in the XRD patterns of any of the samples, potentially due to low abundance or because these oxide phases are actually amorphous. 3.3.2 Total Elemental Composition Table 3.4 presents the total solid phase concentrations of As, Cu, Mo, Pb and Zn in the samples. The total elemental concentrations for all elements are provided in Appendix F. Figure 3.2 presents box plots of the total metal concentrations for As, Cu, Mo, Pb, and Zn for the 16 samples from Site 1 and 16 samples from Site 3. The box plots show the median value (red line), the 25th – 75th percentile (blue box), the 9th-91st percentile (whiskers) and outliers (red crosses). The data shows that there were higher concentrations of As and Mo in the Site 3 samples and higher concentrations of Pb and Zn in the Site 1 samples. Some of the differences in total elemental composition between the sites can be explained by the distinct lithologies of the samples. The Site 1 samples contained more marble bearing waste rock which is expected to have higher Pb contents than other waste rock types (Hirsche, 2012; Lipten and Smith, 2004; Love et al., 2004). There were more samples from Site 3 that contained igneous intrusive waste rock in which Mo bearing sulfides are predominantly found (Hirsche, 2012; Lipten and Smith, 62  2004; Love et al., 2004). Site 1 samples contain more skarn and hornfels waste rock, known to contain higher Zn. While total Zn was generally lower in the Site 3 samples, samples from the base of the Site 3 borehole were composed mostly of hornfels and skarn and thus contained relatively high Zn. 3.3.3 SEP Results Figure 3.3 through Figure 3.9 present the SEP results as percent leached (%) Ca, Fe, As, Cu, Mo, Pb, and Zn, respectively, each figure contains an inset graph of the total metal concentration of each sample (reproduced from Table 3.4). Appendix G presents similar SEP plots for both ppm leached and % leached Al, Ca, Fe, Mn, As, Cu, Mo, Pb, Zn. The results of the SEPs indicated that weak-acid soluble phases and amorphous reducible phases contained the largest fraction of the metals of interest in this study. The highest proportion of Ca was leached during the weak-acid soluble step, Step 3 (see Figure 3.3), as was expected due to the high carbonate content of the waste rock. Ca was also leached during the water soluble extraction step, which can be attributed to gypsum dissolution. In Steps 4 and 5 Ca is attributed to the dissolution of wollastonite (see Chapter 2). Wollastonite was noted in the X-ray diffractograms of 7 samples (Appendix E) and makes up part of the exoskarn lithology (Lipten and Smith, 2004; Love et al., 2004). Jambor et al. (2002) noted that the neutralization potential (NP) of wollastonite was comparable to the NP provided by calcite, and due to its significant abundance could be an additional source of neutralization in this waste-rock pile. The highest proportion of Fe was leached during the two reductive dissolution steps that target iron oxides, Step 4 and Step 5 (Figure 3.4). The samples that leached the highest amounts of Fe in these extraction steps were generally obtained from material that was noted in the drill logs to have moderate to strong visible iron oxide alteration. 63  The highest proportions of extracted Cu, Pb and Zn (Figure 3.6, Figure 3.8, and Figure 3.9, respectively) were found in the weak-acid soluble and the amorphous reducible phases extraction steps. The highest proportion of extracted As (Figure 3.5) was found in the amorphous reducible phase extraction Step 4. Most Mo remained in the residual step and only a few samples (N=8) leached Mo from the secondary phases targeted by the SEP method (Figure 3.7). Mo was leached in varying concentrations during Steps 1, 2, 4 and 5 in only a few samples (N = 8) at low concentrations generally representing <10% of the total Mo with the exception of samples BH-1d (90.0 – 91.5) and BH-1d (91.5 – 92.4). These two samples leached 60 – 80% of the total Mo during Step 5 (crystalline iron oxides) but had low total Mo (~100ppm) initially. 3.3.4 Optical Microscopy and SEM/BSE Imaging of Secondary Minerals Plane-polarized microscope and SEM images were used to investigate the secondary phases that could be observed in thin section. Semi-quantitative analysis of EDS spectra from spot locations was used to determine wt.% of elements associated with the secondary phases found. Appendix H presents the full results of the microscopy and SEM/EDS imaging investigation. From the mineralogical investigation two distinct coatings on silicate, carbonate and sulfide minerals were found: a blue coating and rusty-brown coating. Images of the “blue precipitate” are presented in Figure 3.10 and Figure 3.11 for Site 1 and Site 3 samples, respectively. Figure 3.10-A shows a calcite grain from BH-1d (19.5 – 21.0) that is coated by a blue precipitate, with high Cu, Zn, O, C and S (with Cd, Fe, Mn, Al, Si and Ca making up <1% of the phase). Figure 3.10-B and Figure 3.10-C present images and results from BH-1d (91.5-92.4), in these examples analysis of the EDS spectra showed that the mineral phase contained Cu, Zn, O and C (with Si making up <0.5% of the phase). Also in Figure 3.10-B the mineral associated with the “blue precipitate” was composed predominantly of Zn, Si and O and 64  was inferred to be the hemimorphite that was identified from the XRD investigation of this sample. All images in Figure 3.11 are from BH-3s (15.0 – 16.5); the analysis of these EDS spectra showed high wt.% Cu, O, C and S, with only one specimen containing 0.8 wt% Zn. Generally, other elements made up <1% of the EDS spectra except in Figure 3.11-A where volume effects of the SEM may have excited Ca in the calcite that the “blue precipitate” was coating, causing 1.8% Ca in the precipitate. Figure 3.12 and Figure 3.13 present the plane polarized and SEM images of iron oxides found in samples from Site 1 and Site 3, respectively. Iron oxides formed as coatings around silicate minerals (Figure 3.12-B and Figure 3.13-A), and as coatings and rims around weathered sulfide minerals such as pyrite (+Zn) (Figure 3.12-A, Figure 3.12-C and Figure 3.13-C) and chalcopyrite (Figure 3.13-B). Analysis of the EDS spectra for these iron oxides showed that Cu, Zn, Pb (with 0.01% Mo in Figure 3.12-B) were associated with the samples examined from Site 1 and Cu, Zn, Pb, As and Mo were associated with the samples examined from Site 3.  Seepage Geochemistry Taken together, the results of the SEPs and SEM/EDS investigations suggested that in addition to sorption onto iron oxides, Cu and Zn were precipitating as hydroxycarbonate and hydroxysulfate phases. In some samples, precipitated phases appear to contain both Cu and Zn. Prior to conducting speciation calculations on the seepage samples, thermodynamic data for aurichalcite (Cu, Zn)5(CO3)2(OH)6 and rosasite (Cu,Zn)2(CO)3(OH)2 (Alwan et al., 1980) were integrated into the model database to represent the Cu and Zn-bearing blue precipitate observed in the East Dump samples. Both aurichalcite and rosasite are copper-zinc hydroxycarbonates that are found as secondary minerals in copper/zinc deposits (Frost et al., 2007a, 2007b) and are light blue to green in color. Both minerals are noted to have varying Cu:Zn ratios in natural samples 65  and have been synthesized with varying Cu:Zn ratios (Frost et al., 2007a, 2007b). Because hydrozincite was previously observed as a secondary phase in neutral drainage sites with high Zn contamination (Jacquat et al., 2008; Younger, 2000), it was also added to the database, even though this phase was not observed in our mineralogical investigations. Interestingly, aurichalcite is stoichiometrically similar to hydrozincite and rosasite is stoichiometrically similar to malachite, and solid solutions between each have been suggested (Yoder et al., 2011). Thermodynamic data for Cu/Zn hydroxysulfates were not available in the literature and could not be added to the database. Namuwite, (Zn,Cu)4SO4(OH)6·4H2O, which is stoichiometrically similar to brochantite (Yoder et al., 2011), has been studied for its crystalline structure and is noted to occur jointly with hydrozincite (Groat, 1996). The only documented Cu/Zn-carbonate-sulfate secondary mineral to naturally occur is schulenbergite, (Cu,Zn)7(SO4,CO3)2(OH)10·3(H2O), which was discovered at the Glücksrad mine, Germany and at the Hirao mine, Japan (Ohnishi et al., 2007), however, as with namuwite thermodynamic data for this phase are not yet available. Figure 3.14 and Figure 3.15 present the measured pH and saturation indices predicted from the chemistries of two seeps monitored at the toe of the slope below Site 1 and Site 3, respectively. An example PhreeqC input file is provided in Appendix I. The modelled geochemistry for the neutral seep associated with Site 1, CO-41, showed that goethite, Fe(OH)3(a), rosasite, aurichalcite and calcite were supersaturated throughout the simulation period. The SI of gypsum oscillated between -1 and 0 between the wet season and dry season, respectively until the end of the wet season in 2011 after which it remained at equilibrium (SI=0). At the end of the 2011 wet season, the pH declined from 7.5 to 6.5 and the SIs of gypsum, brochantite, hydrozincite, and wulfenite increased from undersaturated values close to 66  equilibrium (between -0.5 and 0.5), while the SI of calcite decreased from supersaturated values of ~1 to undersaturated (SI <-0.5). Drainage water from the neutral seep associated with Site 3, CO-28, is similar to that from the neutral seep associated with Site 1, in that goethite, Fe(OH)3(a), rosasite, aurichalcite, and calcite were supersaturated throughout the simulation period; however, hydrozincite was additionally supersaturated throughout the simulation period. Gypsum was at equilibrium. Over the time period modelled, the SI of smithsonite was always between 0 and -0.5, while the SI of wulfenite varied seasonally between -1 and 1, generally trending between supersaturated conditions in the wet season and undersaturated conditions in the dry season until after the end of the wet season in 2008 when the SI remained at equilibrium. During the wet season of 2006, 2007 and 2008 the pH at CO-28 declined to a minimum of 6.7, 5.9 and 3.8, respectively, but rebounded back to neutral at the end of each wet season. During these events the SIs of brochantite, antlerite, malachite and jarosite increased from undersaturated to supersaturated conditions with SIs > 3. In both neutral seepages, supersaturation of calcite is likely an artifact due to the degassing of CO2 from the sample, prior to or at the time of the pH measurement, which tends to cause an increase in pH. If CO2 degassing had taken place, it would also cause an over-prediction of the SIs for the hydroxide, hydroxycarbonate and hydroxysulfate phases. The modelled geochemistry of the acidic seeps associated with Site 1 (CO-57) and Site 3 (CO-56) are similar in that jarosite and goethite are supersaturated with SIs > 1 and gypsum is at equilibrium over the simulation period. 3.4 Discussion From the mineralogical and geochemical investigations used in this study it is possible to distinguish two predominant attenuation processes active in the waste-rock dump: precipitation of hydroxycarbonate and hydroxysulfate phases that attenuate Cu and Zn and sorption or co-67  precipitation onto precipitating iron oxides that attenuate As, Cu, Pb, and Zn. The presence of a weak-acid soluble Pb-bearing phase (such as cerrusite, PbCO3) is inferred from the SEP results but is neither supported by the geochemical modelling of seepage chemistry, nor is it identified in the mineralogical examinations. From the data, the predominant attenuation mechanism for Mo cannot be conclusively identified; however, the data suggest that Mo is attenuated by sorption onto iron oxides in only a few samples at low concentrations and by the precipitation of wulfenite.  3.4.1 Hydroxycarbonate and Hydroxysulfate Phases Blue colored coatings on primary minerals identified in the thin sections containing variable Cu, Zn, C and S, likely represent the phases that leached Cu and Zn in the weak-acid soluble step of the SEP. The abundance of Cu, Zn, C and S associated with these phases appear to be influenced by the rock types present in the sample and the ARD classification of the material. The “blue precipitate” found in the Site 1 samples contained both Cu and Zn and is likely due to the higher total Zn concentrations present in Site 1 samples (from marble waste rock) as compared to Site 3 samples. The presence or absence of S in these phases appears to be related to the ARD character of the samples where S is present in samples that were characterized as PAG and uncertain, but not in samples that were identified as NAG. From the mineralogical investigation it is clear that a complex assortment of secondary hydroxycarbonate and hydroxysulfate mineral phases is present in the dump, potentially owing to the heterogeneous geochemical/lithological nature of the waste rock. In an experimental pile study of PAG waste rock at Antamina (Peterson, 2014) a “blue precipitate” that had formed from drainage emanating from the pile was collected and using XRD determined that the precipitate was mostly amorphous in nature containing some gypsum and malachite. X-Ray absorption 68  spectroscopy (XAS)–fine structure (XAF) and near edge structure (XANES) was then used to show that much of the copper in the sample was bonded to sulfate. Using EDS, Peterson (2014) also showed that the precipitate contained Zn, Cu, Si, Al, Mn, Mg, S, and C, but did not quantify the relative abundance of these elements. Without the addition of aurichalcite, rosasite, and hydrozincite to the geochemical database, speciation modeling would fail to confirm hydroxycarbonates as sinks for Cu and Zn, as malachite and smithsonite were, for the most part, undersaturated in the drainage chemistries of the seeps. Aurichalcite and rosasite were consistently supersaturated in neutral drainage from Site 1 and Site 3, but undersaturated in the acidic seeps. Brochantite and malachite were supersaturated only when pH remained in the range between 5 and 7. Hydrozincite was shown to be supersaturated in the drainage from Site 3, but was undersaturated in the drainage from Site 1 (until the pH declined below 7, at which point it became supersaturated). 3.4.2 Iron Oxides Iron oxides are widely documented to be one of the most important sinks for both cationic and anionic metals in many mining waste and environmentally impacted sediment studies in a host of different acid-rock drainage (ARD) and neutral-rock drainage (NRD) environments (Caraballo et al., 2009; Carlsson et al., 2002; Dold and Fontboté, 2002, 2002; Segura et al., 2006; Smuda et al., 2007). During the SEPs the highest concentrations of Fe were leached during the amorphous reducible phase step (Step 4), implying that the Fe-oxides are likely amorphous or poorly crystalline, consistent with the fact that Fe-oxides were not identified in the XRD investigation. The geochemical modelling showed that goethite was supersaturated in both the acidic and neutral seeps and FeOH3(a) was supersaturated in the neutral seeps. From Figure 3.4, the amount of Fe leached from the Site 1 samples was similar regardless of total iron 69  content, ARD classification, depth or material type, while there appears to be slight differences in Fe leached from the Site 3 samples. Also leached during Step 4 of the SEP were As, Cu, Pb and Zn and these associations were confirmed by the SEM/EDS investigations. During the SEPs Mo remained until the residual phase step of most samples. The residual phases include stable silicate phases, sulfide phases, and potentially wulfenite which is considered a very stable phase (Vlek and Lindsay, 1977). Wulfenite was identified in the XRD patterns of 8 of the 32 samples. Mo was leached from 8 non-wulfenite bearing samples during Step 1, Step 2, Step 4 and Step 5 of the SEP but at very low concentrations compared to the total Mo of the sample. This SEP result appears to suggest multiple attenuation processes for Mo; however, it is also possible that Mo is released by desorption from iron oxides which is expected to occur in pH >5 conditions (Conlan, 2009; Geng et al., 2013), as present during Step 1 and Step 2, or desorption due to the complete dissolution of iron oxides, during Step 4 and Step 5. The samples that leached Mo during these steps, when investigated using SEM/EDS, were found to have Mo associated with iron oxides. The association of Mo with iron oxides suggests that the pore water pH in these zones may be acidic as sorption of the oxyanion Mo increases with decreasing pH. These observations are consistent with previous studies at Antamina that indicated there are two likely attenuation mechanisms for Mo: sorption onto iron oxides in acidic micro-environments (Dockrey et al., 2014) and the precipitation of wulfenite when Mo bearing pore water is in contact with Pb-bearing waste rock (Conlan, 2009; Conlan et al., 2012; Hirsche, 2012). Finally, during the weakly sorbed/cation exchange extraction step (Step 2), Cu, Pb, and Zn were detected in the leachate but at low concentrations (<3% of the total leached). Cation exchange was suggested as a potential attenuation mechanism for Zn in a stacked barrel study by Hirsche et al. 70  (2012), however, based on the SEP results this mechanism does not appear to be the dominant attenuation mechanism for Zn. The sequential extractions used in this study are useful to indirectly assess the potential mobility of the metals studied in the mine waste. The hydroxysulfate and hydroxycarbonate Cu and Zn phases will be stable under neutral drainage conditions, but will tend to dissolve and release metals if acidic conditions (pH<5) arise, for example in locations of low carbonate mineral content in which carbonate minerals become depleted or passivated. Hemimorphite and smithsonite, although not conclusively determined to be secondary minerals that formed in the East Dump, should also be considered phases from which Zn could be released, if the pH decreased below 5 (see Chapter 2). The metals associated with the iron oxides are expected to be stable in neutral to slightly acidic conditions and remobilization would require the development and persistence of acidic pH conditions (pH<3) or reducing conditions to cause dissolution of these oxide phases. Mo would be the exception to enhanced metal release under low pH conditions, since sorbed Mo would be expected to become more stable, provided that dissolution of iron oxides does not occur. Mo associated with iron oxides was identified in only a few samples (N = 8) suggesting that neutral conditions currently prevail in the dump. Wulfenite observed in samples (N = 8) during the XRD investigation is considered to be a stable mineral phase in neutral conditions and may provide a very stable phase for the attenuation of Mo (Vlek and Lindsay, 1977). 3.4.3 Implications The observations made in this study support the occurrence of sorption on iron oxides as a process of metal attenuation for As, Cu, Pb and Zn (and Mo in zones that are acidic) and also suggest that Cu and Zn attenuation in carbonate-bearing waste rock needs to be accounted for 71  using mixed Cu:Zn hydroxycarbonate and hydroxysulfate phases, although thermodynamic data for such phases are limited. Thermodynamic data for rosasite and aurichalcite were added to the geochemical model to represent the observed of Cu-Zn-C-O phases in the East Dump samples. While the actual phases precipitating in the dump may have different Cu:Zn ratios than those of rosasite and aurichalcite, the simulations show that these phases were supersaturated, while phases like smithsonite and malachite were generally undersaturated in the neutral drainage. Hydrozincite was not observed in the mineralogical investigations, but it was also predicted to be supersaturated by the geochemical model in one of the neutral seeps; thus, should be considered a possible attenuating phase for Zn until it can be conclusively ruled out. Thermodynamic data for mixed Cu:Zn hydroxysulfate phases were not available; however, the mineralogical investigation shows phases that contain Cu-Zn-S-O-C to be present in samples that were PAG and uncertain. The SEPs also show that a significant proportion of Cu and Zn are stored in weak acid soluble phases and that the onset of ARD may cause the release of Cu and Zn via dissolution of these phases. Although not the focus of this study, two primary minerals have been shown to be of potential interest with respect to ARD and metal leaching. The EDS examination showed that some of the pyrite contained trace amounts of Zn (0.20 – 0.23 wt.%; Appendix H Figure H.17, Figure H.21, and Figure H.22), oxidation of pyrite with Zn impurities may be another source of Zn in solution, in addition to the oxidation of sphalerite. The dissolution of wollastonite may contribute to the neutralization potential of the waste rock (Jambor et al., 2002); however, up until now calcite has been considered the only phase to contribute to acid neutralization.  72  3.5 Conclusions The analysis of waste rock and seepage geochemistry from Antamina’s East Dump show that precipitation of weak-acid soluble hydroxycarbonate and hydroxysulfate phases and sorption/co-precipitation onto iron oxides are the predominant attenuation processes for Cu and Zn. From the SEP data these hydroxycarbonate and hydroxysulfate phases appear to be as important as iron oxides in attenuating Cu and Zn. From the mineralogical investigation these phases can contain both Cu/Zn which may vary compositionally as a function of the relative abundances of Cu and Zn. The results of the sequential extraction procedure used in this study also suggest that weak-acid soluble and iron-oxide phases are attenuating Pb, although the SEM/EDS investigations only found Pb associated with iron oxides. The SEP results and SEM/EDS investigations show that As is associated with amorphous iron oxides. In most of the samples investigated, Mo was released predominantly during the residual step of the SEP suggesting that Mo is associated with sulfides, and stable phases such a wulfenite. Only a few samples leached Mo and the SEP and SEM/EDS investigations suggest that Mo is being released mainly by desorption from iron oxide surfaces. Since Mo associated with iron oxides was identified in only a few samples it can be suggested that neutral conditions currently prevail in the dump. Geochemical modelling of the seepage chemistry provides evidence for the type of mineral phases that might be controlling metal solubility; however, the database used required modification to capture the complexity of the secondary phases that were observed in the samples. In high carbonate waste rock with both Cu and Zn sulfides, precipitation of mixed Cu:Zn hydroxycarbonate and hydroxysulfate phases in addition to pure end-member Cu or Zn hydroxycarbonate or hydroxysulfate phases may exert significant control on the aqueous 73  concentrations of Cu and Zn. Thermodynamic data for these phases is limited, highlighting the need for more geochemical studies on these phases. 74  Tables Table 3.1 Samples selected for metal attenuation study  Sample ID Major Lithology (>50%) Minor lithology (<50%)  Other Lithology  Primary Sulfides  Secondary minerals S(T) S(2-) TIC AP (S2-) Carb-NP NPR ARD Classification (Price, 2009) (%) (%) (%) kg CaCO3/t kg CaCO3/t - BH-1s (1.5 - 3.0)* M MDP XV Cp, Mo, Py, Bi FeOx (2) 0.6 0.5 4.2 16 351 22.4 NAG BH-1s (10.5 - 12.0) M  XV Cp, Bn, Py, Po  0.4 0.3 9.6 10 801 82.7 NAG BH-1s (19.5 - 21.0) IQM M  Cp, Mo, Py FeOx (3) 10.0 9.7 1.1 303 93 0.3 PAG BH-1d (1.5 - 3.0) MDP M XCVC, IQM Cp, Py, Po FeOx (1) - - - - - - - BH-1d (19.5 - 21.0)* IQM  M, BN sec Cp, Mo, Py, FeOx (3), Malachite 8.6 8.0 1.4 251 117 0.5 PAG BH-1d (25.5 - 27.0) XCVC MDP M, IQM, BN Cp, Sp, Py, Po FeOx (4), Malachite 12.9 12.4 1.7 388 143 0.4 PAG BH-1d (39.0 - 40.5) M XCVC MDP, IQM, BN Cp, Py, Po FeOx (2), Malachite 6.6 6.4 3.5 198 287 1.5 uncertain BH-1d (42.0 - 43.5) MDP M XC Cp, Sp, Py, Po FeOx (1), Malachite 1.7 1.7 4.5 53 374 7.1 NAG BH-1d (49.5 - 51.0) MDP XV IQM, M Cp, Py FeOx (1) 2.5 2.4 2.6 76.2 213 2.8 NAG BH-1d (54.0 - 55.5) M IQM MDP, XV Cp, Mo, Py, Po FeOx (1) 1.3 1.3 2.1 40 178 4.5 NAG BH-1d (64.5 - 66.0) M MDP IQM, XCVC, XW Cp, Py, Po, Bi FeOx (1) 0.5 0.5 8.1 15.3 671 43.9 NAG BH-1d (70.5 - 72.0) IQM NC M, XW, MDP, BN Cp, Bn, Mo, Py, FeOx (1), Malachite 1.7 1.7 2.5 51.6 207 4.0 NAG BH-1d (76.5 - 78.0) M MDP IQM, XCVC, BN Cp, Sp, Py,Po FeOx (1), Malachite 1.5 1.5 5.9 46.6 491 10.5 NAG BH-1d (81.0 - 82.5) IQM M M, XCVC Cp, Py, Po, FeOx (3), Malachite 1.0 1.0 3.2 31 268 8.8 NAG BH-1d (90.0 - 91.5)* M MDP IQM, XV Cp, Sp, Py, Po FeOx (4), Malachite 1.2 1.1 4.7 34 391 11.5 NAG BH-1d (91.5 - 92.4) MDP XV M, IQM Cp, Sp, Py, Po FeOx (4), Malachite 1.5 1.4 2.9 43 238 5.6 NAG BH-3s (1.5 - 3.0) M IQM  Cp, Py, Po FeOx(1) 0.6 0.6 3.6 18 302 16.4 NAG BH-3s (9.0 - 10.5) M MDP IQM Py, Po FeOx (1), 1.6 1.5 6.2 48 517 10.8 NAG 75   Sample ID Major Lithology (>50%) Minor lithology (<50%)  Other Lithology  Primary Sulfides  Secondary minerals S(T) S(2-) TIC AP (S2-) Carb-NP NPR ARD Classification (Price, 2009) (%) (%) (%) kg CaCO3/t kg CaCO3/t - malachite BH-3s (10.5 - 12.0) M  IQM Cp, Py, Po FeOx (2) 1.2 1.2 5.5 37 458 12.5 NAG BH-3s (15.0 - 16.5)* IQM M  Cp, Py, Po FeOx (4), malachite 2.9 2.5 1.3 77 111 1.4 uncertain BH-3s (21.0 - 22.5) M XV  Py, Po FeOx (2), malachite 0.4 0.4 7.1 13 591 47.2 NAG BH-3d (10.5 - 12.0) C   Py FeOx (1) 1.3 1.3 6.9 40 576 14.5 NAG BH-3d (24.0 - 25.5)* XV C  Cp, Py FeOx (1), Malachite 0.8 0.7 4.9 22 409 18.4 NAG BH-3d (37.5 - 39.0) IQM C  Cp, Py FeOx (2), Malachite 14.9 14.6 1.8 456 152 0.3 PAG BH-3d (39.0 - 40.5) IQM C  Cp, Py FeOx (2) 19.3 19.0 1.3 593 107 0.2 PAG BH-3d (48.0 - 49.5)* IQM XCVC  Cp, Py FeOx (1) 1.7 1.7 1.4 52 117 2.3 NAG BH-3d (72.0 - 73.5) XV HG  Cp, Py  12.5 12.2 2.7 383 227 0.6 PAG BH-3d (88.5 - 90.0) IQM   Py FeOx (1) 2.9 2.8 1.2 89 102 1.2 uncertain BH-3d (91.5 - 93.0)* IQM XV  Cp, Py FeOx (1), Malachite 4.6 4.5 1.7 141 140 1.0 PAG BH-3d (94.5 - 96.0) IQM NC  Cp, Sp, Mo, Py  4.7 4.6 2.8 144 236 1.6 uncertain BH-3d (112.5 - 114.0) IQM M  Cp, Py FeOx (1), Malachite 0.5 0.5 3.1 17 259 15.6 NAG BH-3d (121.5 - 123.0) HG IQM  Py, Po FeOx (1) 0.5 0.4 5.4 13.8 450 32.6 NAG NOTES:  * indicates samples that were selected for SEM/BSE investigations Acronyms used in the table are defined as such:IQM – Igneous Intrusive;M – Marble; MDP – Marble Diopside; XCVC – Brown and Green Garnet Exoskarn; XV – Green Garnet Exoskarn; C – Limestone; NC – Brown Garnet Endoskarn; Cp –Chalcopyrite; Sp – Sphalerite; Mo – Molybdenite; Py – Pyrite; Po – Pyrrhotite; Mt – Magnetite; Bi – Bismuthite;Bn – Bornite; FeOx – Iron Oxides (1-4 denotes intensity of oxidation; where 1 = trace, 2 = weak, 3 = moderate and 4 = strong)76  Table 3.2 Sequential extraction procedure (SEP) (modified from Hall, Vaive, Beer, & Hoashi, 1996) Step Phases Method(1) ICP-OES Detection Limit  ppm of solid(4) Weight out 1g of sample and react in the prescribed order:  1 Water Soluble 50 mL deionized water shake for 1h <10 ppm 2 Weakly sorbed /Exchangeable(2) 40 mL MgCl2 shake for 1h <10 ppm 3 Weak acid soluble 20 mL CH3COONa (sodium acetate) at pH 5 shake for 6h centrifuge for 10min < 6 ppm Repeat Step 4 Amorphous reducible phases 20 mL 0.25M NH2OH*HCl (hydroxylamine hydrochloride) in 0.25 HCl place in 60⁰C water bath for 2h every 30min vortex contents <6 ppm Repeat Step but heat for only 30 min 5 Crystalline reducible phases 30 mL of 1 M NH2OH*HCl (hydroxylamine hydrochloride) in 25% CH3COOH (acetic acid) place in 90⁰C water bath for 3h, vortex every 20 min <8 ppm Repeat Step but heat for only 1.5 hours 6 Residual Phases (silicates, sulfides) 4 - Acid Digest(3)   NOTES:   1)  Steps 3 – 5 are repeated using the same liquid solid ratio (LSR) but extraction time was shortened. Leachates are analysed separately and the concentrations are summed.  2)  Weakly sorbed/exchangeable step taken from (Tessier et al., 1979)  3)  Residual fraction was determined at SGS, Burnaby (BC, Canada)  4)  Detection limit in ppm of solid based upon detection limit of the ICP-OES (0.2mg/L) and volume of regent  77  Table 3.3 Summary of XRD identified mineralogy for East Dump waste rock samples Sample ID Pyrite Molybdenite Chalcopyrite Sphalerite Gypsum Smithsonite Wulfenite Hemimorphite Mineral Formula FeS2 MoS2 CuFeS2 ZnS CaSO4 ZnCO3 PbMoO4 Zn4Si2O7(OH)2•(H2O) BH-1s (1.5 - 3.0)* x x     x       BH-1s (10.5 - 12.0)                 BH-1s (19.5 - 21.0) x       x       BH-1d (1.5 - 3.0) x           x   BH-1d (19.5 - 21.0)*(1) x   x   x   x   BH-1d (25.5 - 27.0) x   x   x       BH-1d (39.0 - 40.5) x x     x   x   BH-1d (42.0 - 43.5) x     x x       BH-1d (49.5 - 51.0) x x x   x x     BH-1d (54.0 - 55.5) x x             BH-1d (64.5 - 66.0)                 BH-1d (70.5 - 72.0) x x x   x       BH-1d (76.5 - 78.0) x   x   x       BH-1d (81.0 - 82.5) x   x   x       BH-1d (90.0 - 91.5)* x   x   x     x BH-1d (91.5 - 92.4) x   x   x x   x BH-3s (1.5 - 3.0) x x x           BH-3s (9.0 - 10.5) x x     x x     BH-3s (10.5 - 12.0) x   x   x       BH-3s (15.0 - 16.5)*(1) x x x   x   x   BH-3s (21.0 - 22.5)       x x       BH-3d (10.5 - 12.0) x       x       BH-3d (24.0 - 25.5)* x x x x x   x   BH-3d (37.5 - 39.0) x x     x   x   BH-3d (39.0 - 40.5) x x     x       78  Sample ID Pyrite Molybdenite Chalcopyrite Sphalerite Gypsum Smithsonite Wulfenite Hemimorphite Mineral Formula FeS2 MoS2 CuFeS2 ZnS CaSO4 ZnCO3 PbMoO4 Zn4Si2O7(OH)2•(H2O) BH-3d (48.0 - 49.5)* x x x   x       BH-3d (72.0 - 73.5) x x x   x   x   BH-3d (88.5 - 90.0) x x x   x x     BH-3d (91.5 - 93.0)* x x x   x       BH-3d (94.5 - 96.0) x x x   x   x   BH-3d (112.5 - 114.0) x x x           BH-3d (121.5 - 123.0) x               NOTES:  * Indicates samples that were selected for SEM/BSE investigations (1) Indicates samples in which the Rietveld method was used to determine quantitative mineralogy 79  Table 3.4 Trace element concentration by four-acid digestion and ICP-MS finish Sample ID As Cu Mo Pb Zn   ppm ppm ppm ppm ppm BH-1s (1.5 - 3.0)* 104 1340 468 2080 674 BH-1s (10.5 - 12.0) 46 1170 24.5 138 1300 BH-1s (19.5 - 21.0) 45 6360 50.2 293 4600 BH-1d (1.5 - 3.0) 151 1030 27.2 800 1550 BH-1d (19.5 - 21.0)* 49 5740 104 195 4900 BH-1d (25.5 - 27.0) 68 7070 71.5 646 7660 BH-1d (39.0 - 40.5) 92 2410 169 563 2460 BH-1d (42.0 - 43.5) 48 1390 62.7 540 6550 BH-1d (49.5 - 51.0) 68 1720 103 462 3990 BH-1d (54.0 - 55.5) 28 1440 103 108 668 BH-1d (64.5 - 66.0) 65 1020 42.1 570 1300 BH-1d (70.5 - 72.0) 58 >10000 285 160 991 BH-1d (76.5 - 78.0) 67 6650 80.2 970 4190 BH-1d (81.0 - 82.5) 111 3880 215 295 857 BH-1d (90.0 - 91.5)* 118 5530 73.3 593 >10000 BH-1d (91.5 - 92.4) 94 4030 107 874 7310 BH-3s (1.5 - 3.0) 98 1250 213 367 1590 BH-3s (9.0 - 10.5) 79 3540 94.4 1800 3750 BH-3s (10.5 - 12.0) 90 4400 84.7 412 951 BH-3s (15.0 - 16.5)* 86 18600 124 50.1 1090 BH-3s (21.0 - 22.5) 152 4750 77 631 4000 BH-3d (10.5 - 12.0) 69 5510 41.8 168 850 BH-3d (24.0 - 25.5)* 193 6320 331 700 8290 BH-3d (37.5 - 39.0) 55 3760 154 117 473 BH-3d (39.0 - 40.5) 47 3180 150 74 309 BH-3d (48.0 - 49.5)* 78 6120 309 258 519 BH-3d (72.0 - 73.5) 47 2650 129 156 997 BH-3d (88.5 - 90.0) 106 5350 310 208 4370 BH-3d (91.5 - 93.0)* 151 4500 195 276 3720 BH-3d (94.5 - 96.0) 65 2900 242 294 2450 BH-3d (112.5 - 114.0) 123 2050 145 105 1950 BH-3d (121.5 - 123.0) 90 542 16 289 1010 NOTES: * indicates samples that were selected for SEM/BSE investigations   80  Figures  Figure 3.1 The Antamina Mine site, inset plan of the East Dump and pre-mining topography showing drill site locations and seeps downslope of drill sites 81   Figure 3.2 Box and whisker plot of total metal concentration box per site82   Figure 3.3 SEP results for Ca (%) with inset total concentration (ppm)  83   Figure 3.4 SEP results for Fe (%) with inset total concentration (ppm) 84   Figure 3.5 SEP results for As (%) with inset total concentration (ppm)  85   Figure 3.6 SEP results for Cu (%) with inset total concentration (ppm)  86   Figure 3.7 SEP results for Mo (%) with inset total concentration (ppm)  87   Figure 3.8 SEP results for Pb (%) with inset total concentration (ppm)  88   Figure 3.9 SEP results for Zn (%) with inset total concentration (ppm)89   Figure 3.10 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: Figure 10-A is from BH-1d (19.5 – 21.0); Figure 10-B and Figure 10-C are from BH-1d (91.5-92.4)  90   Figure 3.11 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: All images are from BH-3s (15.0 – 16.5)  91   Figure 3.12 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: Figure 3.12-A from BH-1s (1.5 – 3.0); Figure 3.12-B and -C from BH-1d (90.0 – 91.5)  92   Figure 3.13 Plane polarized (left)/SEM (right) images of secondary phases with semi-quantitative wt% of elements: Figure 13-A from BH-3d (24.0 – 25.5); Figure 13-B and Figure 13-C from BH-3d (48.0 – 49.5)93   Figure 3.14 Equilibrium phases predicted from geochemical modelling of seeps downslope Site 1; CO-41 (top) and CO-57(bottom) 94   Figure 3.15 Equilibrium phases predicted from geochemical modelling of seeps downslope Site 3; CO-28 (top) and CO-56(bottom)95  Chapter 4 Conclusions  This research is one part of a large scale study that is aimed at understanding the hydrologic and geochemical influences on water in a neutral drainage waste rock dump that contains a mix of metal sulfides and variable carbonate content. The principal objective of this research was to identify secondary phases and characterize metal attenuation. To this end, waste rock samples were collected from a full-scale operational waste rock dump and investigated using total digestions, acid-base-accounting (ABA) tests, X-ray diffraction (XRD), sequential extraction procedures (SEPs), reflected and transmitted light microscopy, and scanning electron microscopy with electron dispersion spectroscopy (SEM/EDS). The metals of interest were arsenic (As), copper (Cu), molybdenum (Mo), lead (Pb), and zinc (Zn). This chapter summarizes the findings presented in the two main chapters of this thesis, discuss the uncertainties in the data and future research that could reduce uncertainties in the data. 4.1 Summary of Key Findings In Chapter 2 two sequential extraction procedures (SEPs) were investigated to identify selectivity and carry-over associated with geochemically heterogeneous, naturally weathered materials, the key findings of this chapter may be of interest to those investigating skarn waste rock or any other type of material consisting of silicate minerals mixed with sulfides and carbonates. The conclusions from this chapter are:  For mixed sulfide, carbonate bearing waste rock SEPs should be designed to: 1) remove water soluble phases, such as gypsum; 2) distinguish between metals bound by cation exchange/weak sorption and metals associated with precipitation of weak acid soluble 96  phases, such as metal carbonate/hydroxycarbonate/hydroxysulfate phases; and, 3) distinguish metal bound by sorption onto reducible phases.  Oxalate should not be used as a reagent to dissolve iron oxides when wollastonite is present in the samples. Wollastonite dissolved in reagents that were acidified to pH < 3. The dissolution of wollastonite released Ca2+ into solution which reacted with oxalate to precipitate whewellite (CaC2O4·H2O). It is postulated that the dissolution of wollastonite may have neutralized the acidity of the oxalate reagent used (Jambor et al., 2002), allowing for precipitation of whewellite. While whewellite was the only phase identified, geochemical modelling suggests that other insoluble metal oxalates (Cu-, Pb- and Zn-Oxalates) could have also formed affecting predictions of Cu, Pb and Zn release. In Chapter 3 mixed sulfide/carbonate bearing waste rock samples collected from the East Dump were investigated to identify the phases that metals where attenuated by. The conclusions from this chapter are:  Precipitation of blue colored weak-acid soluble hydroxycarbonate and hydroxysulfate phases are predominant attenuation processes for Cu and Zn. From the SEP data these hydroxycarbonate and hydroxysulfate phases appear to be as important as iron oxides in attenuating Cu and Zn, but are relatively less stable than iron oxides. From the mineralogical investigation these phases can contain both Cu/Zn and C/S and the variation in elemental composition may be a function of the total metal content of the waste rock as well as the total Sulfur (Stot) and Total Inorganic Carbon (TIC). The blue colored Cu/Zn hydroxycarbonate and hydroxysulfate phases observed in this study are 97  consistent with blue precipitates observed from an experimental pile of similar waste rock (Peterson, 2014).  Modelling of seepage chemistry showed that most end-member Cu or Zn hydroxycarbonate and hydroxysulfate minerals were undersaturated. The observations as discussed in the bulleted point above allowed for the inclusion of mixed Cu/Zn hydroxycarbonate phases (aurichalcite and rosasite, (Alwan et al., 1980) and hydrozincite (Preis and Gamsjäger, 2002)) in the database. These phases were generally supersaturated. For future geochemical and reactive transport modelling these phases may be adequate analogues to the secondary phases observed in the East Dump samples.  As, Cu, Pb, and Zn were leached from amorphous iron oxide dissolution steps and found associated with iron oxides in the SEM/EDS investigation. Iron oxides are well known for their ability to scavenge metals out of solution many mining waste studies (Caraballo et al., 2009; Carlsson et al., 2002; Dold and Fontboté, 2002, 2002; Segura et al., 2006; Smuda et al., 2007). From these associations sorption/co-precipitation of metals onto iron oxides is also considered a predominant attenuation process in the dump. Conceptually, these iron oxides are considered stable phases and dissolution and release of the metals associated with iron oxides would require the onset and persistence of pH<3 conditions or reducing conditions.  A Pb-bearing weak-acid soluble phase is suggested from the SEPs, although the mineralogical and SEM/EDS investigations only found Pb associated with iron oxides and the XRD investigations found wulfenite. Conceptually, this suggests that there is an “available” pool of Pb from an unknown phase from which release would require pH<5 conditions to arise. 98   Mo was released predominantly during the residual step of the SEP suggesting that Mo is associated with sulfides and stable phases. Only a few samples leached Mo during the SEP and the SEP and SEM/EDS investigations suggest that Mo is being released mainly by desorption from iron oxide surfaces. Mo released from the weak-sorption sites (Geng et al., 2013) and the dissolution of iron oxides in the SEP is indicative of locations in the dump in which pH<5 (Dockrey et al., 2014). The XRD investigation identified wulfenite in a few samples and Rietveld analysis of two samples show that wulfenite is present locally in low quantity (0.15 – 0.34 %; Appendix E Figure E.5, and Figure E.20). Wulfenite is considered a stable phase (Vlek and Lindsay, 1977) that forms quickly when Mo releasing leachate is in contact with Pb-bearing waste rock (Conlan, 2009; Conlan et al., 2012; Hirsche, 2012).  Pyrite investigated in the samples contained some Zn (0.20 – 0.23 wt.%; Appendix H Figure H.17, Figure H.21, and Figure H.22) oxidation of these pyrite grains would release Zn into solution. This should be considered a source of Zn, in addition to sphalerite.  Wollastonite makes up part of the skarn lithology of the Antamina ore deposit and in the waste rock. Other researchers have shown that wollastonite may provide neutralization of acidity generated similar to that of calcite (Jambor et al., 2002). Up until now this silicate mineral has not been considered for its neutralization potential in reactive transport models at the Antamina Site. 4.2 Future Work While the SEM/study shows that Cu-Zn-C-S-O are present in varying ratios in samples, it is not able to identify the bonds between elements. These phases were potentially amorphous 99  and/or crystalline but low in content as to not be identifiable using XRD. Micro X-ray diffraction (µ-XRD) or accelerated light techniques such as XAFS/XANES can be used on these phases to assess mineralogy (if possible) and bonding relationships, providing sufficient amount of this material can be collected for these techniques. Thermodynamic data for mixed Cu/Zn hydroxycarbonate phases are limited and thermodynamic data for mixed Cu/Zn hydroxysulfate phases are non-existent. 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December 1, 2015.106  Appendices Appendix A Reverse Circulation Drill Logs This Appendix presents the Drill Logs created during drilling of the East Dump.107   Figure A.1 BH-1s (0.0m - 23.50m) 108   Figure A.2 BH-1d (0.0m - 49.50m) 109   Figure A.3 BH-1d (49.50m - 92.40m) 110   Figure A.4 BH-3s (0.0m – 22.50m) 111   Figure A.5 BH-3d (0.0m - 46.50m) 112   Figure A.6 BH-3d (46.50m - 91.50m) 113   Figure A.7 BH-3d (91.50m - 136.50m) 114   Figure A.8 BH-3d (136.50m - 145.00m)115   Appendix B Method Quality and Inter-sample Variability (Chapter 2) B.1 Method Sub-sample Precision To identify if there was variability in the sub-sampling of each sample, precision in the leachates from the triplicate samples tested was calculated using the relative standard deviation (RSD), as follows:  %𝑹𝑺𝑫 =  𝒔?̅?× 𝟏𝟎𝟎        0-1  where s is the standard deviation and ?̅? is the mean. RSD is widely used in analytical chemistry to express the precision and repeatability of an assay. A low RSD was indicative of low variability in sub-samples. Percent Recovery Percent recovery was used as a proxy for complete dissolution of all mineral phases in the samples tested. Percent recovery was calculated for each element by comparing the sum of mas extracted during all steps of the sequential extractions, including the residual step, to the total bulk elemental content determined via 4-acid digest. Mass Loss  The mass of solids lost to the centrifugation, decanting and filtration of the leachates at each step was determined by weighing the syringe filter before and after use and drying. This determine that minimal solids were lost during the decanting step and that the desired liquid:solid ratio was maintained.  116  pH Drift  To ensure that the liquid:solid ratio was sufficient to dissolve the carbonates present in the samples, the pH was monitored during the carbonate step of the SEPs (step 3). If the initial and final pH were sufficiently close, then complete dissolution of the carbonates was suggested. B.2 Results Triplicate Analyses Calculated RSD’s for triplicate results are presented in Table B.1 and Table B.2 for Method 1 and Method 2, respectively. The RSDs for the triplicate analyses were generally better than 10%, but some were between 10 % and 30% depending on the element and the detection limit. Only a few elements had RSDs higher than 30% and were typically elements that were at low concentration or near the detection limit of the ICP-OES.  Percent Recovery The calculated % recovered for Al, As, Ca, Cu, Fe, Mn, Mo, Pb, and Zn leached from Method 1 and Method 2 are presented in Table B.3. For both methods the % recovered were generally between 70% and 130%, and between methods were similar. In only a few instances was the % recovery less than 70% and in more instances was the % recovery greater than 130%. The greater amount leached versus total elemental composition is most likely due to comparing results from duplicate samples and in some of the cases for Ca is due to the results of the total digestion reported greater than the maximum detection limits of the analysis. The similarity of % recovery between methods is a good indication that both methods will dissolve the same minerals over the entirety of the sequential extraction procedure. 117  Mass Loss  Table B.4 presents the mass loss results. The maximum recorded loss was 4.18% and the minimum recorded loss was 0%, the average calculated loss was 1.52%. The mass losses were higher for the Hall method than the Dold Method. This is most likely due to the repeat analysis for steps 3 – 5 in the Hall method which required additional rinsing, centrifuging, and decanting steps.  pH Drift  Table B.5 presents the initial and final pH values for step 3 of each method. Based on the initial and final pH, both methods remained at the target pH implying that there was sufficient acid for the carbonate step.  B.3 Conclusion Each method was individually assessed by testing samples in triplicate, monitoring mass loss, pH drift and by calculating % recovery for specific elements. The calculated RSDs for both methods were low, indicating good precision in the triplicate samples. Higher mass losses were identified in Method 1 as compared to Method 2. This is expected due the higher number of manipulations of the sample prescribed by the method. A higher RPM during centrifugation of the samples would likely reduce the mass loss during decanting the leachates. The pH for Step 3 of each method remained close to the target value and thus no changes to the reagent pH are necessary. The calculated percent recoveries for specific elements were generally good for both methods and between methods were similar. 118  Table B.1 Calculated RSD for triplicate analysis for Method 1   119  Table B.2 Calculated RSD for triplicate analysis for Method 2  120   Table B.3 Percent recovery (total elemental concentration vs. cumulative leached) Al As Ca Cu Fe Mn Mo Pb Zn Al As Ca Cu Fe Mn Mo Pb Znppm ppm ppm ppm ppm ppm ppm ppm ppm %recovery % % % % % % % % %Cummulative Method 1 5849.145 29.40005 383592.8 734.6647 5257.954 244.4205 14.00136 215.2124 1205.087 Method 1 106% 64% 256% 63% 83% 95% 57% 156% 93%Cummulative Method 2 6251.744 45.33075 332155.3 677.8239 5522.783 271.7778 24.09525 223.5705 990.7009 Method 2 114% 99% 221% 58% 88% 105% 98% 162% 76%BH-1s (10.5 - 12.0) 5500 46 >150000 1170 6300 258 24.5 138 1300Cummulative Method 1 25132.69 40.00991 72198.11 6610.484 90223.77 865.1267 36.79716 783.3079 6618.783 Method 1 131% 89% 189% 104% 134% 116% 73% 267% 144%Cummulative Method 2 26099.02 46.14248 63710.69 6675.823 91732.82 889.6849 41.49972 676.6827 6470.909 Method 2 136% 103% 166% 105% 136% 120% 83% 231% 141%BH-1s (19.5 - 21.0) 19200 45 38300 6360 67400 744 50.2 293 4600Cummulative Method 1 12919.6 40.50697 349859.8 710.9772 10535.64 495.8599 32.01614 1054.311 1293.143 Method 1 137% 62% 233% 70% 98% 98% 76% 185% 99%Cummulative Method 2 12630.85 50.85265 305358.4 720.1278 11125.48 553.3211 33.06944 1072.579 1323.441 Method 2 134% 78% 204% 71% 104% 110% 79% 188% 102%BH-1d (64.5 - 66.0) 9400 65 >150000 1020 10700 505 42.1 570 1300Cummulative Method 1 18955.24 85.23529 161689 4424.11 91714.87 1125.416 91.2796 1650.278 11127.71 Method 1 327% 91% 160% 110% 149% 103% 85% 189% 152%Cummulative Method 2 19322.33 131.2723 168495.7 4596.769 117983.2 1177.997 129.0266 1707.352 10201.88 Method 2 333% 140% 167% 114% 192% 108% 121% 195% 140%BH-1d (91.5 - 92.4) 5800 94 101000 4030 61500 1090 107 874 7310Cummulative Method 1 41781.12 50.10563 104623.7 802.7186 17767.24 477.4498 145.7708 459.4543 1172.295 Method 1 84% 51% 102% 64% 87% 84% 68% 125% 74%Cummulative Method 2 41933.51 73.04047 122439.2 980.3522 21694.4 625.2274 187.5032 496.5238 1301.558 Method 2 85% 75% 119% 78% 106% 110% 88% 135% 82%BH-3s (1.5 - 3.0) 49500 98 103000 1250 20500 569 213 367 1590Cummulative Method 1 29204.25 61.56246 100455.9 14938.3 85651.1 650.578 93.78637 61.65474 952.0413 Method 1 94% 72% 113% 80% 110% 107% 76% 123% 87%Cummulative Method 2 29869.8 99.53734 112139.3 17454.36 91933.79 750.6287 111.5939 106.644 1122.745 Method 2 96% 116% 126% 94% 118% 123% 90% 213% 103%BH-3s (15.0 - 16.5) 31100 86 89000 18600 77800 610 124 50.1 1090Cummulative Method 1 29819.59 154.4257 256187.9 5701.961 54815.54 1930.201 311.9222 1157.007 6897.801 Method 1 97% 80% 105% 90% 112% 104% 94% 165% 83%Cummulative Method 2 31585.86 214.5543 262095.9 6961.461 59670.54 2083.118 324.9682 1217.201 9152.685 Method 2 103% 111% 107% 110% 122% 113% 98% 174% 110%BH-3d (24.0 - 25.5) 30600 193 244000 6320 48800 1850 331 700 8290Cummulative Method 1 30326.14 31.20714 47943.76 2933.972 166736.7 287.3294 108.8965 136.0508 350.6762 Method 1 94% 57% 83% 78% 71% 80% 71% 116% 74%Cummulative Method 2 31760.79 42.52382 34603.55 3024.954 185737.2 322.5714 110.3972 151.9514 400.0459 Method 2 98% 77% 60% 80% 79% 90% 72% 130% 85%BH-3d (37.5 - 39.0) 32300 55 57600 3760 234000 360 154 117 473Notes:For the % recovery calculation of Ca, the maximum detection value was used when greater the detection limit was encountered57% green highlights indicate cumulative values that are 70%<x>130% of total value121  Table B.4 Mass loss experiment Sample Initial mass (g) Loss (g) Loss (%) 3Ha 1.0039 0.0324 3.23% 16Ha 1.0012 0.0209 2.09% 17Ha 1.0027 0.0099 0.99% 20Ha 1.0003 0.0152 1.52% 23Ha 1.0026 0.0121 1.21% 24Ha 1.0021 0.04188 4.18% 3Dc 1.0002 0.0243 2.43% 16Dc 1.0029 0.0221 2.20% 17Da 1.0018 0.0008 0.08% 20Da 1.0034 0 0.00% 23Da 1.0014 0.0004 0.04% 24Da 1.0022 0.0033 0.33% 122   Table B.5 pH drift experiment  BH-1s (10.5 - 12.0) BH-1s (19.5 - 21.0) BH-1d (64.5 - 66.0) BH-1d (91.5 - 92.4) BH-3s (1.5 - 3.0) BH-3s (15.0 - 16.5) BH-3d (24.0 - 25.5) BH-3d (37.5 - 39.0) pH initial Method 2 4.5 4.5 4.5 4.5 4.5 4.5 4.5 4.5 pH final Method 2 n/a 4.6 n/a 4.68 4.69 4.67 4.73 4.6 pH initial Method 1 5 5 5 5 5 5 5 5 pH final Method 1 5.59 5.06 5.42 5.22 5.3 5.24 5.46 5.1 123   Appendix C Comparison of Sequential Extraction Procedures Method 1 and Method 2 and Differential X-ray Diffraction (DXRD) diffractograms  124  C.1 SEP Results for Method 1 and Method 2  Figure C.1 Sequential extraction results Al (%) and Ca (%) Al (%) Ca (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-2 0% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-20% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-40Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual0Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual125   Figure C.2 Sequential extraction results Mn (%) and Mg (%) Mg (%) Mn (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-20% 20% 40% 60% 80% 100%Method 2Method 1BH1-20Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual0Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual126   Figure C.3 Sequential extraction results Fe (%) and As (%) Fe (%) As (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-30Method 2Method 1LegendStep 1 Step 2 Step 3 Step 4Step 5 residual Step 1 Step 2Step 3 Step 4 Step 5 Residual0% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-20% 20% 40% 60% 80% 100%Method 2Method 1BH1-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-40Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual0Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual127   Figure C.4 Sequential extraction results Cu (%) and Mo (%) Cu (%) Mo (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-20% 20% 40% 60% 80% 100%Method 2Method 1BH1-20Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual0Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual128   Figure C.5 Sequential extraction results Pb (%) and Zn (%)   Pb (%) Zn (%)0% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-10% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-30% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-10% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-20% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-30% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH3-40% 20% 40% 60% 80% 100%Method 2Method 1BH1-20% 20% 40% 60% 80% 100%Method 2Method 1BH1-20Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual0Method 2LegendStep 1 Step 2 St p 3 Step 4 Step 5 Residual129  C.2 XRD Results for Un-reacted Samples and Sample Residues After Application of Leaching Steps   Figure C.6 BH1-1 XRD results from Method 1  Figure C.7 BH1-2 XRD results from Method 1 130   Figure C.8 BH1-3 XRD results from Method 1  Figure C.9 BH1-4 XRD results from Method 1 131   Figure C.10 BH3-1 XRD results from Method 1  Figure C.11 BH3-3 XRD results from Method 1 132   Figure C.12 BH3-2 XRD results from Method 1  Figure C.13 BH3-4 XRD results from Method 1 133   Figure C.14 BH1-1 XRD results from Method 2  Figure C.15 BH1-2 XRD results from Method 2 134   Figure C.16 BH1-3 XRD results from Method 2  Figure C.17 BH1-4 XRD results from Method 2 135   Figure C.18 BH3-1 XRD results from Method 2  Figure C.19 BH3-2 XRD results from Method 2 136   Figure C.20 BH3-3 XRD results from Method 2  Figure C.21 BH3-4 XRD results from Method 2   137   Appendix D Geochemical Modelling of Oxalate Species and Saturation Indices for Metal Oxalates Table D.1 Tabulated results of SI and precipitation scenarios sim si_CaOxalate si_CuOxalate si_MgOxalate si_MnOxalate si_PbOxalate si_ZnOxalate 1  6.427 6.2691 3.0977 0.9617 3.752 6.6417 0 0 -3.1274 -4.6947 -0.8176 0 2  6.5989 6.2782 3.1722 0.9085 3.6911 6.7162 0 -0.1097 -3.2188 -4.9295 -1.0193 0 3  6.7544 6.2864 3.2481 0.8562 3.6311 6.792 0 -0.2547 -3.2982 -5.1499 -1.2151 0 4  6.8989 6.2939 3.3252 0.8046 3.5719 6.8691 0 -0.3888 -3.3654 -5.3563 -1.4031 0 5  7.0356 6.3008 3.4035 0.7539 3.5136 6.9475 0 -0.5151 -3.4237 -5.5523 -1.5853 0 6  -999.999 6.2221 2.8133 1.1839 4.0035 6.3573 -999.999 0 -2.7842 -3.763 -0.1423 0    138  D.1 PhreeqC Input File  SELECTED_OUTPUT     -file                 SIs_T_UM.xls     -reset                false     -simulation           true     -step                 true     -ph                   true     -pe                   true     -molalities           Oxalate-2  H(Oxalate)-  H2(Oxalate)     -saturation_indices   CaOxalate  CuOxalate  MgOxalate  MnOxalate                           PbOxalate  ZnOxalate SOLUTION_MASTER_SPECIES     Oxalate       Oxalate-2        2     88.06           88.06  SOLUTION_SPECIES Oxalate-2 = Oxalate-2     log_k     0 H+ + Oxalate-2 = H(Oxalate)-1  log_k 4.2798 H+ + H(Oxalate)-1 = H2(Oxalate)  log_k 1.27             PHASES CaOxalate     CaOxalate:H2O = Ca+2 + H2O + Oxalate-2     log_k     -8.6345 CuOxalate     CuOxalate = Cu+2 + Oxalate-2     log_k     -9.3536 MgOxalate     MgOxalate:2H2O = 2H2O + Mg+2 + Oxalate-2     log_k     -5.3161 MnOxalate     MnOxalate:2H2O = 2H2O + Mn+2 + Oxalate-2     log_k     -6.7696 PbOxalate     PbOxalate = Oxalate-2 + Pb+2     log_k     -8.0706 ZnOxalate     ZnOxalate:2H2O = 2H2O + Oxalate-2 + Zn+2     log_k     -8.8601  SOLUTION 1     temp      25     pH        3     pe        3     redox     pe     units     mol/l     density   1     Oxalate   0.2     Ca        0.2     Cu        0.2     Mg        0.2 139      Mn        0.2     Pb        0.2     Zn        0.2     -water    1 # kg  EQUILIBRIUM_PHASES 1     CaOxalate 0 0     CuOxalate 0 0     MgOxalate 0 0     MnOxalate 0 0     PbOxalate 0 0     ZnOxalate 0 0 END   SOLUTION 2     temp      25     pH        3     pe        3     redox     pe     units     mol/l     density   1     Oxalate   0.2     Ca        0.25     Cu        0.2     Mg        0.2     Mn        0.2     Pb        0.2     Zn        0.2     -water    1 # kg  EQUILIBRIUM_PHASES 1     CaOxalate 0 0     CuOxalate 0 0     MgOxalate 0 0     MnOxalate 0 0     PbOxalate 0 0     ZnOxalate 0 0 END  SOLUTION 3     temp      25     pH        3     pe        3     redox     pe     units     mol/l     density   1     Oxalate   0.2     Ca        0.3     Cu        0.2     Mg        0.2     Mn        0.2     Pb        0.2     Zn        0.2     -water    1 # kg 140   EQUILIBRIUM_PHASES 1     CaOxalate 0 0     CuOxalate 0 0     MgOxalate 0 0     MnOxalate 0 0     PbOxalate 0 0     ZnOxalate 0 0 END  SOLUTION 4     temp      25     pH        3     pe        3     redox     pe     units     mol/l     density   1     Oxalate   0.2     Ca        0.35     Cu        0.2     Mg        0.2     Mn        0.2     Pb        0.2     Zn        0.2     -water    1 # kg  EQUILIBRIUM_PHASES 1     CaOxalate 0 0     CuOxalate 0 0     MgOxalate 0 0     MnOxalate 0 0     PbOxalate 0 0     ZnOxalate 0 0 END  SOLUTION 5     temp      25     pH        3     pe        3     redox     pe     units     mol/l     density   1     Oxalate   0.2     Ca        0.4     Cu        0.2     Mg        0.2     Mn        0.2     Pb        0.2     Zn        0.2     -water    1 # kg  EQUILIBRIUM_PHASES 1     CaOxalate 0 0     CuOxalate 0 0     MgOxalate 0 0 141      MnOxalate 0 0     PbOxalate 0 0     ZnOxalate 0 0 END  SOLUTION 5     temp      25     pH        3     pe        3     redox     pe     units     mol/l     density   1     Oxalate   0.2     Ca        0.0     Cu        0.2     Mg        0.2     Mn        0.2     Pb        0.2     Zn        0.2     -water    1 # kg  EQUILIBRIUM_PHASES 1     CaOxalate 0 0     CuOxalate 0 0     MgOxalate 0 0     MnOxalate 0 0     PbOxalate 0 0     ZnOxalate 0 0 END   142   Appendix E Summary of XRD Results (Tabulated) and X-ray Diffractograms for East Dump Waste Rock Samples 143  E.1 Tabulated XRD Mineralogy   Sample ID Quartz OrthoclaseAlbite Biotite/Phlogopite Muscovite Calcite Diopside Hibschite Grossular Garnet Andradite Vesuvianite WollastoniteActinolite Tremolite MagnetitePyrite MolybdeniteChalcopyriteSphaleriteChlorite Gypsum Smithsonite WulfeniteHemimorphite Al2O3Mineral Formula SiO2 KAlSi3O8 NaAlSi3O8 K(Mg,Fe)3(AlSi3O10)(F,OH)2 KAl2(AlSi3O10)(F,OH)2 CaCO3 CaMg(Si2O6)Ca3Al2(SiO4)2(OH)4 Ca3Al2(SiO4)3 Ca3Fe2(SiO4)3Ca10Mg2Al4(Si2O7)2(SiO4)5(OH)4 CaSiO3 Ca2(Mg,Fe)5Si8O22(OH)2. Ca2Mg5Si8O22(OH)2 Fe3+2Fe2+O4FeS2 MoS2 CuFeS2 (Zn,Fe)S CaSO4 ZnCO3 PbMoO4 Zn4Si2O7(OH)2•(H2O)BH-1s (1.5 - 3.0) x x x x x x x x x x x x x xBH-1s (10.5 - 12.0) x x x x x xBH-1s (19.5 - 21.0) x x x x x x x x x xBH-1d (1.5 - 3.0) x x x x x x x x x x xBH-1d (19.5 - 21.0) x x x x x x x x x x x x xBH-1d (25.5 - 27.0) x x x x x x x x x x x x x xBH-1d (39.0 - 40.5) x x x x x x x x x x x x x x x xBH-1d (42.0 - 43.5) x x x x x x x x x x x x x xBH-1d (49.5 - 51.0) x x x x x x x x x x x x x xBH-1d (54.0 - 55.5) x x x x x x x x x xBH-1d (64.5 - 66.0) x x x x x x xBH-1d (70.5 - 72.0) x x x x x x x x x x x x xBH-1d (76.5 - 78.0) x x x x x x x x x x x x xBH-1d (81.0 - 82.5) x x x x x x x x x xBH-1d (90.0 - 91.5) x x x x x x x x x x x x x xBH-1d (91.5 - 92.4) x x x x x x x x x x x x x x xBH-3s (1.5 - 3.0) x x x x x x x xBH-3s (9.0 - 10.5) x x x x x x x x x x xBH-3s (10.5 - 12.0) x x x x x x x x x xBH-3s (15.0 - 16.5) x x x x x x x x x x x x x x xBH-3s (21.0 - 22.5) x x x x x x x x x x xBH-3d (10.5 - 12.0) x x x x x x x x xBH-3d (24.0 - 25.5) x x x x x x x x x x x x x x x xBH-3d (37.5 - 39.0) x x x x x x x x x x x xBH-3d (39.0 - 40.5) x x x x x x x x x x xBH-3d (48.0 - 49.5) x x x x x x x x xBH-3d (72.0 - 73.5) x x x x x x x x x x x x x xBH-3d (88.5 - 90.0) x x x x x x x x x x x x x xBH-3d (91.5 - 93.0) x x x x x x x x x x x x x xBH-3d (94.5 - 96.0) x x x x x x x x x x x x x x xBH-3d (112.5 - 114.0) x x x x x x x x x xBH-3d (121.5 - 123.0) x x x x x x x x144  E.2 XRD Patterns for Individual Samples  Figure E.1 BH-1s (1.5 - 3.0)  Figure E.2 BH-1s (10.5 -12.0) 145   Figure E.3 BH-1s (19.5 - 21.0)  Figure E.4 BH-1d (1.5 - 3.0) 146   Figure E.5 Rietveld analysis of BH-1d (19.5 – 21.0)  Figure E.6 BH-1d (25.5 - 27.0) 2Th Degrees75706560555045403530252015105Counts15,00010,0005,0000S5_laurenzi_XRDR.raw Quartz low 18.11 %Orthoclase 21.44 %Calcite 12.87 %Gypsum 4.00 %Pyrite 14.96 %Chalcopyrite 0.53 %Albite low 5.83 %Wulfenite 0.15 %Diopside 8.28 %Vesuvianite 2.46 %Actinolite 1.02 %Phlogopite 1M 3.50 %Grossular 6.50 %147   Figure E.7 BH-1d (39.0 - 40.5)  Figure E.8 BH-1d (42.0 - 43.5) 148   Figure E.9 BH-1d (49.5 - 51.0)  Figure E.10 BH-1d (54.0 - 55.5) 149   Figure E.11 BH-1d (64.5 - 66.0)  Figure E.12 BH-1d (70.5 - 72.0) 150   Figure E.13 BH-1d (76.5 - 78.0)  Figure E.14 BH-1d (81.0 - 82.5) 151   Figure E.15 BH-1d (90.0 - 91.5)  Figure E.16 BH-1d (91.5 - 92.4) 152   Figure E.17 BH-3s (1.5 - 3.0)  Figure E.18 BH-3s (9.0 - 10.5) 153   Figure E.19 BH-3s (10.5 - 12.0)  Figure E.20 Rietveld analysis of BH-3s (15.0 – 16.5) 2Th Degrees75706560555045403530252015105Counts30,00025,00020,00015,00010,0005,0000-5,000S20_laurenzi_XRDR_redo.raw_1 Quartz low 28.63 %Orthoclase 18.80 %Calcite 14.65 %Gypsum 4.93 %Pyrite 2.95 %Chalcopyrite 3.57 %Albite low 8.41 %Wulfenite 0.34 %Diopside 3.51 %Phlogopite 1M 3.22 %Actinolite 1.32 %Molybdenite 2H 0.10 %Magnetite 5.79 %Andradite 3.79 %154   Figure E.21 BH-3s (21.0 – 22.5)  Figure E.22 BH-3d (10.5 - 12.0) 155   Figure E.23 BH-3d (24.0 - 25.5)  Figure E.24 BH-3d (37.5 - 39.0) 156   Figure E.25 BH-3d (39.0 - 40.5)  Figure E.26 BH-3d (48.0 - 49.5) 157   Figure E.27 BH-3d (72.0 - 73.5)  Figure E.28 BH-3d (88.5 - 90.5) 158   Figure E.29 BH-3d (91.5 - 93.0)  Figure E.30 BH-3d (94.5 - 96.0) 159   Figure E.31 BH-3d (112.5 - 114.0)  Figure E.32 BH-3d (121.5 - 123.0)   160   Appendix F Total Elemental Concentrations From 4-acid Digestions 161   Table F.1 Tabulated elemental concentrations from 4-acid digestions Date : Sept 26 2013Sample ID Ag Al Ba Ca Cr Cu Fe K Li Mg Mn Na Ni P S Sr Ti V Zn Zr As Be Bi Cd Ce Co Cs Ga Hf In La Lu Mo Nb Pb Rb Sb Sc Se Sn Ta Tb Te Th Tl U W Y Ybppm % ppm % ppm ppm % % ppm % ppm % ppm ppm % ppm % ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppmMethod Code IC40M IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40MLOD 0.02 0.01 1 0.01 1 0.5 0.01 0.01 1 0.01 2 0.01 0.5 50 0.01 0.5 0.01 2 1 0.5 1 0.1 0.04 0.02 0.05 0.1 5 0.1 0.02 0.02 0.1 0.01 0.05 0.1 0.5 0.2 0.05 0.1 2 0.3 0.05 0.05 0.05 0.2 0.02 0.05 0.1 0.1 0.1BH-1s (10.5 -12) 1.87 0.55 42 >15 18 1170 0.63 0.43 10 0.53 258 0.11 4.3 160 0.45 2750 0.03 9 1300 1.1 46 0.2 8.23 3.34 5.83 3.6 6 1.9 0.04 0.42 3.2 0.02 24.5 2.2 138 16.3 2.84 0.5 3 0.9 0.23 0.06 0.11 0.2 0.22 0.89 23.3 1.2 0.1BH-1s (19.5 -21) 6.69 1.92 43 3.83 58 6360 6.74 0.95 16 0.56 744 0.11 7.9 350 >5 48.1 0.09 41 4600 11.7 45 0.8 12.9 10.4 27.5 11.6 <5 15.3 0.62 6.84 17.6 0.15 50.2 5.8 293 35.9 16.2 3 17 42.9 0.4 0.3 1.27 2.1 0.74 4.67 30.2 8.9 0.9BH-1d (64.5 -66) 3.57 0.94 48 >15 27 1020 1.07 0.65 12 0.54 505 0.09 5 370 0.53 774 0.05 19 1300 7.3 65 0.4 16.6 3.76 18 3.4 <5 4 0.25 0.32 9.3 0.07 42.1 3.3 570 20.1 10.7 1.1 5 1.7 0.33 0.19 0.34 0.9 0.31 1.66 19.4 4.8 0.5BH-1d (91.5 -92.4) >10 0.58 34 10.1 53 4030 6.15 0.46 17 0.71 1090 0.05 7.2 720 1.52 165 0.09 41 7310 12 94 0.9 84.3 21.7 23.6 12.3 <5 15.4 0.48 10.9 16.5 0.09 107 5.3 874 32.2 13.1 1 30 52.3 0.51 0.23 1.27 0.6 0.5 4.44 177 4.5 0.6Date : July 15, 2013Sample ID Ag Al Ba Ca Cr Cu Fe K Li Mg Mn Na Ni P S Sr Ti V Zn Zr As Be Bi Cd Ce Co Cs Ga Hf In La Lu Mo Nb Pb Rb Sb Sc Se Sn Ta Tb Te Th Tl U W Y Ybppm % ppm % ppm ppm % % ppm % ppm % ppm ppm % ppm % ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppmMethod Code IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40A IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40M IC40MLOD 0.02 0.01 1 0.01 1 0.5 0.01 0.01 1 0.01 2 0.01 0.5 50 0.01 0.5 0.01 2 1 0.5 1 0.1 0.04 0.02 0.05 0.1 5 0.1 0.02 0.02 0.1 0.01 0.05 0.1 0.5 0.2 0.05 0.1 2 0.3 0.05 0.05 0.05 0.2 0.02 0.05 0.1 0.1 0.1BH-3s ( 1.5-3.0 ) 4.26 4.95 410 10.3 69 1250 2.05 3.04 36 0.7 569 0.52 4.4 620 0.83 511 0.11 37 1590 9.3 98 1.2 23.2 3.54 28.9 6.9 6 14.4 0.25 0.42 15.9 0.09 213 5.2 367 114 7.03 4.1 3 3.1 0.15 0.29 0.5 6.3 1.6 1.47 18.9 7.8 0.6BH-3s (15.0 - 16.5) 11 3.11 259 8.9 71 18600 7.78 2.24 24 0.66 610 0.53 4.2 570 4.07 326 0.09 71 1090 7.4 86 0.9 4.46 2.12 27.8 22 <5 17 0.2 1.65 14.7 0.07 124 2.6 50.1 73.9 15.3 2.8 4 22.2 0.1 0.24 0.2 4.3 2.99 2.28 8.1 5.2 0.5BH-3d (24.0 - 25.5) 21.3 3.06 37 24.4 83 6320 4.88 0.43 16 1.27 1850 0.06 3.4 680 1.06 353 0.13 45 8290 44 193 1.3 145 18.8 33.8 16.3 6 12.1 1.18 3.23 20.9 0.15 331 5.2 700 25.1 29.6 4.2 6 13.7 0.06 0.35 0.75 5.6 0.37 3.8 48.5 10.9 1BH-3d (37.5 - 39.0) 4.04 3.23 84 5.76 71 3760 23.4 1.24 16 0.64 360 0.09 4.9 450 26.8 89.9 0.1 61 473 14.1 55 1.2 175 0.85 29.6 43.5 5 11.5 0.39 0.64 18.3 0.09 154 5.3 117 54.5 5.96 3.8 22 12.8 0.31 0.27 1.08 4.8 0.45 2.4 162 7.2 0.6Date : January 22, 2014Sample ID Ag Al Ba Ca Cr Cu Fe K Li Mg Mn Na Ni P S Sr Ti V Zn Zr As Be Bi Cd Ce Co Cs Ga Hf In La Lu Mo Nb Pb Rb Sb Sc Se Sn Ta Tb Te Th Tl U W Y Ybppm % ppm % ppm ppm % % ppm % ppm % ppm % % ppm % ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppm ppmMethod Code ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40B ICM40BLOD 0.02 0.01 1 0.01 1 0.5 0.01 0.01 1 0.01 2 0.01 0.5 0.005 0.01 0.5 0.01 2 1 0.5 1 0.1 0.04 0.02 0.05 0.1 1 0.1 0.02 0.02 0.1 0.01 0.05 0.1 0.5 0.2 0.05 0.5 2 0.3 0.05 0.05 0.05 0.2 0.02 0.05 0.1 0.1 0.1BH-1s (1.5-3.0) 5.37 4.87 144 >15 75 1340 4.21 1.82 17 1.37 1520 0.32 9.9 0.074 0.73 371 0.21 66 674 48.4 104 2.2 34 2.11 45.5 7.7 9 13.4 1.59 1.03 25.8 0.22 468 11.1 2080 79.6 14.1 6.4 7 9.7 1.15 0.5 1.36 8.1 0.61 3.2 53 14.2 1.3BH-1d (1.5-3.0) 4.08 2.3 92 >15 35 1030 2.35 1.12 12 1 590 0.14 8.5 0.055 1.36 809 0.1 51 1550 17.9 151 1 20.4 5.02 22 6 6 5.6 0.39 0.62 11.7 0.12 27.2 4.7 800 58.6 8.18 3.6 4 4 0.32 0.26 0.78 3.6 1.01 1.88 29.5 8.2 0.7BH-1d (19.5-21.0) 6.74 3.72 251 7.65 81 5740 8.55 2.66 21 0.88 518 0.37 2.6 0.086 >5 260 0.13 54 4900 15.9 49 1.4 23.2 14.9 24.6 8.4 5 16.7 0.5 5.83 13.2 0.12 104 4.8 195 90.4 7.21 4.5 12 37.3 0.38 0.28 1.25 5.2 0.94 3.17 26 8.1 0.7BH-1d (25.5-27.0) 9.48 1.95 34 8.06 97 7070 >15 0.78 13 0.73 653 0.07 9.5 0.056 >5 103 0.09 51 7660 18.8 68 1.2 426 23.9 21.6 27.6 18 10.7 0.56 5.39 12 0.12 71.5 5 646 28.9 10.7 3.3 27 27 0.36 0.27 3.18 3.6 0.55 2.6 586 7.9 0.7BH-1d (39.0-40.5) 5.01 3.73 147 13 80 2410 8.42 1.92 15 0.82 563 0.18 5.4 0.055 >5 382 0.12 59 2460 20.8 92 1.4 59 8.12 27.9 28.8 29 10.8 0.57 1.31 15 0.12 169 5 563 64.7 8.51 4.1 9 12.8 0.38 0.3 1.12 5 1.28 2.01 113 8.2 0.7BH-1d (42.0-43.5) 14.3 3.22 61 >15 87 1390 6.79 0.98 13 1.11 1620 0.1 19.1 0.071 2.21 403 0.16 71 6550 42.4 48 1.2 182 17.6 25.5 9.1 4 10.4 1.14 2.35 13.7 0.18 62.7 7.6 540 39.3 8.04 4.9 10 10.4 0.51 0.36 2.79 4.8 0.39 2.22 104 11.8 1BH-1d (49.5-51.0) 5.8 4.11 286 12.3 106 1720 5.59 3.13 22 0.8 774 0.44 23.7 0.065 3.58 435 0.15 54 3990 20.2 68 1.4 92 11.4 27.1 12.2 12 12.7 0.63 1.67 13.6 0.12 103 5.5 462 93.9 15 4.4 7 9.3 0.4 0.31 1.57 5.1 0.9 1.71 82.3 8.9 0.7BH-1d (54.0-55.5) 2.26 5.67 482 9.02 76 1440 2.7 4.1 25 0.77 432 1.06 5.4 0.069 1.46 525 0.18 59 668 13.3 28 2.1 20.3 1.94 32.6 7.6 8 16.1 0.42 0.44 15.8 0.13 103 5.9 108 131 4.34 5.6 2 5.3 0.48 0.37 0.32 6.3 0.79 1.53 39.4 9.7 0.8BH-1d (70.5-72.0) 8.96 3.3 274 12.5 98 >10000 5.61 2.59 19 0.63 1260 0.55 4.4 0.039 2.11 474 0.1 39 991 14.8 58 1.3 27.6 2.56 26 12.7 5 12.5 0.46 2.52 11.6 0.09 285 3.3 160 83.2 13.3 2.7 5 24.1 0.24 0.25 0.48 4 1.46 2.84 9.7 6.1 0.5BH-1d (76.5-78.0) 8.81 2.51 105 >15 66 6650 5.96 1.14 14 1.03 1200 0.19 9.3 0.067 2.98 549 0.11 81 4190 22.8 67 1 50.5 12.5 23.8 15.6 4 10.7 0.58 1.95 12.8 0.13 80.2 4.4 970 47.6 15.7 4.1 8 9 0.29 0.3 1.21 3.8 1.12 2.22 24.8 9.2 0.8BH-1d (81.0-82.5) 2.82 4.11 384 10.4 85 3880 2.28 3.39 27 0.6 517 0.63 5.8 0.057 1.09 483 0.09 39 857 6.2 111 1.2 7.99 2.1 22.4 6.8 5 12.1 0.17 0.62 10.8 0.07 215 2.3 295 106 13 3.4 3 4.4 0.18 0.23 0.23 4.6 1.52 1.3 6.4 6 0.4BH-1d (90.0-91.5) 22.9 2.18 26 >15 70 5530 15 0.43 11 0.71 1360 0.06 2.1 0.068 1.86 166 0.1 44 >10000 33 118 1.6 68.8 25.1 23 22.5 3 16 0.86 12.5 15.7 0.16 73.3 4.8 593 27.6 12.4 3.4 27 55.4 0.3 0.28 1.24 3.1 0.48 5.38 176 10.3 0.9BH-3s (9.0-10.5) 8.96 3.04 215 >15 61 3540 3.37 2.29 20 0.93 820 0.28 16.5 0.055 1.89 950 0.12 47 3750 21.1 79 1.1 41.8 11 24.9 8.8 5 9.9 0.45 1.26 13.1 0.11 94.4 5.5 1800 79 11.4 3.7 5 6.3 0.38 0.28 1.73 4.7 1.09 1.93 29.2 8.3 0.7BH-3s (10.5-12.0) 2.73 4.21 298 11.6 89 4400 3.85 2.95 24 0.9 721 0.4 10.1 0.059 1.92 458 0.14 56 951 17.1 90 1.4 5.86 3.38 29.7 10.5 7 12.6 0.34 0.76 14.8 0.13 84.7 6.2 412 103 6.57 4.9 2 7.1 0.4 0.35 0.33 5.5 1.32 1.38 29.6 9.7 0.8BH-3s (21.0-22.5) 8.5 2.45 49 >15 67 4750 4.52 0.54 11 1.12 1270 0.08 6.7 0.06 0.64 665 0.11 42 4000 32.5 152 1.1 65 9.85 22.6 10.7 5 9 0.87 2.85 12.5 0.13 77 4.7 631 27 21.4 3.5 4 13.7 0.29 0.27 0.73 3.7 0.37 2.62 47.2 9 0.8BH-3d (10.5-12.0) 3.23 1.72 127 >15 59 5510 2.92 1.25 15 0.68 503 0.22 3.5 0.038 1.27 521 0.06 32 850 5.5 69 0.8 5.68 2.31 14.1 8.9 5 6.9 0.17 0.5 6.8 0.06 41.8 2 168 46.6 12.4 2.2 2 5.4 0.12 0.16 0.14 2.4 0.85 1.77 3.6 4.5 0.4BH-3d (39.0-40.5) 2.51 3.04 61 4.79 142 3180 >15 1.31 15 0.64 284 0.1 0.7 0.05 >5 123 0.09 57 309 17.4 47 1.3 107 0.81 26.1 39.4 6 9.6 0.34 0.61 14.7 0.09 150 4.3 74 51 4.47 3.8 20 12.5 0.3 0.26 1.11 4.7 0.42 2.02 148 7.2 0.6BH-3d (48.0-49.5) 4.71 4.76 384 6.2 135 6120 3.69 3.74 28 0.59 830 0.66 5 0.058 1.84 297 0.13 44 519 14 78 2.2 14 1.34 35.8 12.8 8 14.1 0.46 0.66 20.8 0.1 309 4.2 258 126 10.9 4.2 3 7.4 0.36 0.3 0.46 6.7 1.44 1.99 14.3 7.7 0.6BH-3d (72.0-73.5) 2.92 3.23 101 8.76 97 2650 14.7 1.59 16 0.69 407 0.19 1.3 0.049 >5 237 0.1 52 997 16.3 47 1.4 94 2.56 26.5 32.7 6 9.8 0.44 0.86 14.8 0.09 129 4.6 156 57.1 6.56 3.7 16 11.3 0.34 0.27 1.47 4.6 0.48 1.94 144 7.4 0.6BH-3d (88.5-90.0) 16.1 5.16 326 10.1 111 5350 6.01 3.58 27 0.96 1150 0.58 9.3 0.064 3.62 245 0.16 55 4370 27 106 2.3 268 12.2 34 12.8 7 16.9 0.84 3.88 17.5 0.13 310 5.9 208 120 52.2 4.9 8 15.9 0.46 0.35 3.97 6.8 0.81 2.69 532 9.3 0.8BH-3d (91.5-93.0) 9.5 4.06 189 9.09 109 4500 11.2 2.62 25 0.82 918 0.26 4.2 0.055 >5 176 0.13 56 3720 22 151 1.9 289 10.8 30.5 29.6 7 14.6 0.62 3.59 16.1 0.12 195 5.9 276 88.9 40.6 4.4 14 19.4 0.44 0.32 4.73 6 0.7 2.56 457 8.5 0.7BH-3d (94.5-96.0) 6.04 3.51 144 11.2 108 2900 12 1.8 16 0.84 869 0.26 4.2 0.053 >5 231 0.12 64 2450 24.2 65 1.7 108 6.58 27.5 17.6 5 12.7 0.62 2.1 14.8 0.11 242 5.8 294 65.3 15.7 4.4 11 14.2 0.44 0.29 2.07 5.4 0.49 2.38 299 8.2 0.7BH-3d (112.5-114.0) 2.33 5.04 413 8.24 109 2050 1.99 3.81 35 0.74 587 0.61 5.8 0.051 0.71 373 0.12 44 1950 11.3 123 1.7 16.3 4.72 27.8 6.2 5 13.5 0.3 0.31 14.8 0.09 145 3.4 105 122 3.59 4.7 <2 2.8 0.27 0.25 0.44 6.2 1.22 1.01 12.7 7.2 0.6BH-3d (121.5-123.0) 1.98 4.55 135 >15 72 542 2.81 1.6 31 1.24 664 0.3 15.7 0.084 0.77 600 0.24 85 1010 49.8 90 1.8 24 2.94 45.9 8.4 23 11.5 1.03 0.53 23.5 0.26 16 11.8 289 85.5 7.13 8.5 <2 4.6 0.8 0.62 0.67 8 0.78 2.1 55.8 18.6 1.6162  Appendix G Sequential Extraction Procedure (SEP) Results for East Dump Waste Rock Samples 163   Figure G.1 SEP results for Al (ppm) inset graph shows total elemental from 4-acid digestions Al (ppm)   164   Figure G.2 SEP results for Al (%) inset graph shows total elemental from 4-acid digestions Al (ppm)   165   Figure G.3 SEP results for Ca (ppm) inset graph shows total elemental from 4-acid digestions Ca (ppm)   166   Figure G.4 SEP results for Ca (%) inset graph shows total elemental from 4-acid digestions Ca (ppm)   167   Figure G.5 SEP results for Fe (ppm) inset graph shows total elemental from 4-acid digestions Fe (ppm)   168   Figure G.6 SEP results for Fe (%) inset graph shows total elemental from 4-acid digestions Fe (ppm)   169   Figure G.7 SEP results for Mn (ppm) inset graph shows total elemental from 4-acid digestions Mn (ppm)   170   Figure G.8 SEP results for Mn (%) inset graph shows total elemental from 4-acid digestions Mn (ppm)   171   Figure G.9 SEP results for As (ppm) inset graph shows total elemental from 4-acid digestions As (ppm)   172   Figure G.10 SEP results for As (%) inset graph shows total elemental from 4-acid digestions As (ppm)   173   Figure G.11 SEP results for Cu (ppm) inset graph shows total elemental from 4-acid digestions Cu (ppm)   174   Figure G.12 SEP results for Cu (%) inset graph shows total elemental from 4-acid digestions Cu (ppm)   175   Figure G.13 SEP results for Mo (ppm) inset graph shows total elemental from 4-acid digestions Mo (ppm)   176   Figure G.14 SEP results for Mo (%) inset graph shows total elemental from 4-acid digestions Mo (ppm)   177   Figure G.15 SEP results for Pb (ppm) inset graph shows total elemental from 4-acid digestions Pb (ppm)   178   Figure G.16 SEP results for Pb (%) inset graph shows total elemental from 4-acid digestions Pb (ppm)   179   Figure G.17 SEP results for Zn (ppm) inset graph shows total elemental from 4-acid digestions Zn (ppm)   180   Figure G.18 SEP results for Zn (%) inset graph shows total elemental from 4-acid digestions Zn (ppm) 181   Appendix H SEM and EDS Investigation Images and Elemental Wt.% 182  H.1 EDS Un-normalized Wt.% of Elements from Semi-quantitative Analysis of EDS Table H.1 EDS un-normalized wt.% of elements from semi-quantitative analysis of EDS     As Mo Cu Pb Zn C S O Fe K Al Ca Cd Mn Si Ti Bi Total Figure Sample ID % % % % % % % % % % % % % % % % % % Figure H.1-1 BH-1d (19.5 – 21.0)   8.6 0 28.5 7.8 0.2 21.8 0.1  0.1 0.4 0.3 0.4 0.1   68.3 Figure H.1-2 BH-1d (19.5 – 21.0)   46.1 0 1.5 5.5 6 15.3 0.4  0.2 0.1   0.2   75.3 Figure H.2-1 BH-1d (90.0 – 91.5)     49.5 2.7  16.1       9   77.3 Figure H.2-2 BH-1d (90.0 – 91.5)   41.6 0 12.7 9.4  22.8          86.5 Figure H.3 BH-1d (90.0 – 91.5)   18.2 0 32.1 9.6  18.9       0.2   79 Figure H.4 BH-1d (90.0 – 91.5)   45.8 0 8.9 6.6  21 0.5      0.4   83.2 Figure H.5 BH-3s (15.0 – 16.5)   53.5 0 0 6.5 7.8 13.9 0.2  0.7 1.8   0.3   84.7 Figure H.6-1 BH-3s (15.0 – 16.5)   64.4 0 0 3.7 7.2 8   0.2    0.1   83.6 Figure H.6-2 BH-3s (15.0 – 16.5)   8.5    3.2 28.6 25.3 1.14 4.9 1.5   5.1   78.24 Figure H.7 BH-3s (15.0 – 16.5)   62.3 0 0.8 6.2 6.2 12.5 0.3  0.3 0.2  0.1 0.3   89.2 Figure H.8 BH-1s (1.5 – 3.0)    3  5.2 9.6 30.3 31.4 6.4      0.2  86.1 Figure H.9 BH-1d (19.5 – 21.0)   0.4   6.7 9.4 31 26.1 5.5      0.34  79.44 Figure H.10 BH-1s (1.5 – 3.0)   0.5 4.3 1 19.9  21.4 27.7   1.2   2.7   78.7 Figure H.11-1 BH-1s (1.5 – 3.0)   0.8 2.4 1.1 5.9  31.7 39.4  0.6 2.4   6.5   90.8 Figure H.11-2 BH-1s (1.5 – 3.0)   1.1 1.3 1 6.5  28.5 39.1  0.3 2.1   5.3   85.2 Figure H.12 BH-1d (90.0 – 91.5)   4.1 1.2 4.5  0.1 23 43  0.1 0.6   4   80.6 Figure H.13 BH-1d (90.0 – 91.5)  0.01 1.9 2 7.8   26 40.8  1.4 0.4   4.3   84.61 Figure H.14 BH-1d (90.0 - 91.5)   2.1 1.4 5.2   10.4 45.4  0.3 0.4   3.1   68.3 Figure H.15-1 BH-1d (90.0 – 91.5)   0.4 0.7 1.8 6.9  30.8 48.6   0.3   3   92.5 Figure H.15-2 BH-1d (90.0 – 91.5)   0.4 0.1 1.5 9.6  30.8 30  0.2 0.3   13.3   86.2 Figure H.16 BH-1d (19.5 – 21.0)   0.3  1.1 6 0.6 32.6 45.7  0.4 0.3   2.7   89.7 Figure H.17 BH-1d (90.0 – 91.5)   2.8  1 8 0.39 30.4 35.1  0.6 0.6   8.1   86.99 Figure H.18 BH-3d (24.0 – 25.5) 0.6  4.5  3.3 10.2  23.3 30.8  1 2.1  0.1 6   81.9 Figure H.19 BH-3d (24.0 – 25.5) 1.5 0.04 4.6 2.1 6 5  22.5 33   1.1  0.1 3.8  4.4 84.14 183      As Mo Cu Pb Zn C S O Fe K Al Ca Cd Mn Si Ti Bi Total Figure Sample ID % % % % % % % % % % % % % % % % % % Figure H.20 BH-3d (24.0 – 25.5) 0.3 0.11 6.6 0.7 2.5 11.6  19.7 31.5   0.5   3.4  0.3 77.21 Figure H.21-1 BH-3d (48.0 – 49.5) 0.2  0.05 1.5 0.48   23.5 53  0.2 0.4   1.2   80.53 Figure H.21-2 BH-3d (48.0 – 49.5) 0.4 0.01 0.5 0.6 2.3   20.6 47  0.9 0.5   1.9   74.71 Figure H.22 BH-3d (48.0 -49.5) 0.2 0.5 13.3  0.7  0.9 22.3 30.4 0.2 0.7 2.3   5.9   77.4 Figure H.23 BH-3d (91.5 – 93.0) 0.3  1  0.4  2.1 23.7 42.1 0.6 3.8 2.6  0.2 3   79.8 Figure H.24 BH-3d (91.5 – 93.0) 0.2  2.05  4.15 7.02 1.4 29.1 38.7  0.7 1.17   2.7   87.19 184   Figure H.1 Images from BH-1d (19.5 – 21.0) A) Plane-polarized transmitted light photograph of silicate/calcite mineral with black and blue/green secondary mineral coating. B) SEM/BSE image   Figure H.2 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of zinc silicate mineral (hemimorphite) with blue/green secondary mineral coating. B) SEM/BSE image  185   Figure H.3 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image   Figure H.4 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image  186   Figure H.5 Images from BH-3s (15.0 – 16.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image   Figure H.6 Images from BH-3s (15.0 – 16.5) A) Plane-polarized transmitted light photograph of blue/green secondary mineral associated with oxidized chalcopyrite grain (Cpy) B) SEM/BSE image  187   Figure H.7 Images from BH-3s (15.0 – 16.5) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image   Figure H.8 Images from BH-1s (1.5 – 3.0) A) Plane-polarized transmitted light photograph of mineral B) SEM/BSE image  188   Figure H.9 Images from BH-1d (19.5 – 21.0) A) Plane-polarized transmitted light photograph of mineral with coating B) SEM/BSE image    Figure H.10 Images from BH-1s (1.5 – 3.0) A) Plane-polarized transmitted light photograph of pyrite mineral (Py) with iron oxide coatings in a calcite grain B) SEM/BSE image  189   Figure H.11 Images from BH-1s (1.5 – 3.0) A) Plane-polarized transmitted light photograph of Chalcopyrite (Cpy) coated with iron oxide B) SEM/BSE image of same mineral showing zoning of iron oxide   Figure H.12 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of iron oxide grain. B) SEM/BSE image  190   Figure H.13 Images from BH-1d (90.0 – 91.5) Plane-polarized transmitted light photograph of iron oxide coating hemimorphite (white mineral) B) SEM/BSE image   Figure H.14 Image from BH-1d (90.0 - 91.5) A) Plane-polarized transmitted light photograph of iron oxide coating hemimorphite (white mineral) B) SEM/BSE image (300 micron scale)   191   Figure H.15 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of weathered opaque mineral with iron oxide coating B) SEM/BSE image   Figure H.16 Images from BH-1d (19.5 – 21.0) A) Plane-polarized transmitted light photograph of pyrite (with Zn) mineral with iron oxide coatings B) SEM/BSE image  192   Figure H.17 Images from BH-1d (90.0 – 91.5) A) Plane-polarized transmitted light photograph of pyrite (with 0.2 wt.% Zn) mineral with iron oxide coating B) SEM/BSE image   Figure H.18 Images from BH-3d (24.0 – 25.5) A) Plane-polarized transmitted light photograph of calcite mineral with iron oxide coatings B) SEM/BSE image  193   Figure H.19 Image from BH-3d (24.0 – 25.5) A) Plane-polarized transmitted light photograph of garnet (?) mineral with iron oxide coatings B) SEM/BSE image   Figure H.20 Images from BH-3d (24.0 – 25.5) A) Plane-polarized transmitted light photograph of garnet and other mineral (calcite?) with iron oxide coating B) SEM/BSE image magnified into iron oxide coating (50 micron scale)   194   Figure H.21 Images from BH-3d (48.0 – 49.5) A) Plane-polarized transmitted light photograph of pyrite (with 0.23 wt.% Zn) with iron oxide oxidation rim B) SEM/BSE image  Figure H.22 Image from BH-3d (48.0 -49.5) A) Plane-polarized transmitted light photograph of pyrite (Py) (with 0.22 wt.% Zn) and chalcopyrite (Cpy) with iron oxide coating B) reflected light microscope image C) SEM/BSE image  195   Figure H.23 Image from BH-3d (91.5 – 93.0) A) Plane-polarized transmitted light photograph of silicate minerals with iron oxide coatings. B) SEM/BSE image   Figure H.24 Image form BH-3d (91.5 – 93.0) A) Plane-polarized transmitted light photograph of pyrite (Py) (with Zn) with iron oxide coating B) SEM/BSE image    196   Appendix I Drainage Chemistry Geochemical Modelling 197  I.1 Example PhreeqC Input File for Geochemical Modelling of East Dump Seepage Chemistry   SELECTED_OUTPUT    -file                 CO_28.xls    -reset                false    -simulation           true    -solution             true    -ph                   true    -pe                   true    -alkalinity           true    -charge_balance       true    -percent_error        true    -saturation_indices   Calcite  Gypsum  Malachite  Antlerite                          Brochantite  Fe(OH)3(a)  Jarosite-K  Goethite                          Goslarite  Rhodochrosite  Cerrusite  Rosasite Aurichalcite                          Smithsonite  Hydrozincite_Preis    Wulfenite                          PowelliteSOLUTION_SPREAD    -pe       12    -units    mg/l number  Alkalinity    pH  Temp     S(6)      Cl    N(5)    P      Ag     Al     As      Ba     B     Si       Ca      Cd     Cu     Fe      K      Li      Mg      Mn       Mo     Na     Ni      Pb      Se      Sr      Zn      F             pe          as HCO3                                                                                                                                                                                      O2(g) -0.8864508 155.6 8      457 19 1.02 0.3 0.01 0.18 0.001 0.037 0.03 8.6 248.1 0.003 0.045 0.001 3.48 0.02 15.26 0.081 0.01 3.02 0.01 0.015 0.0002 1.688 0.403 0.28 12

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