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A study of pyrite reactivity and the chemical stability of cemented paste backfill Bertrand, Valérie J. 1998

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A STUDY OF PYRITE REACTIVITY AND THE CHEMICAL STABILITY OF CEMENTED PASTE BACKFILL  by V A L E R I E J. B E R T R A N D B . S c , University o f Ottawa, 1991  A THESIS S U B M I T T E D I N P A R T I A L F U L F U L L M E N T O F T H E REQUIREMENTS FOR T H E D E G R E E OF M A S T E R » OF A P P L I E D S C I E N C E in T H E F A C U L T Y OF G R A D U A T E STUDIES (Department o f Mining and Mineral Process Engineering)  We accept this thesis as conforming to the required standard  T H E UNIVERSITY OF BRITISH C O L U M B I A December 1998 © V a l e r i e J. Bertrand, 1998  In  presenting  degree  this thesis  in partial fulfilment of  the  requirements  for  an  at the University of British Columbia, 1 agree that the Library shall make it  freely available for reference and study. I further agree that permission for copying  of  department  this thesis for scholarly purposes or  by  his  or  her  representatives.  may be granted It  is  permission.  Department The University of British Columbia Vancouver, Canada  DE-6 (2/88)  - ~ / £ ~~  extensive  by the head of my  understood  that  publication of this thesis for financial gain shall not be allowed without  Date  advanced  copying  or  my written  ABSTRACT  A two-fold study was carried out to 1) characterize the evolution o f the reactivity o f pyrite i n the early cycles o f kinetic test leaching, using cyclic voltamperometry, and 2) document the weathering characteristics o f various paste backfill mixtures that contain pyritic tailings, when exposed to leaching environments similar to those encountered i n mine settings.  Pyrite leaching experiments were carried out on 6 different pyrite samples from existing mines.  C y c l i c voltamperometry was performed on carbon paste electrodes ( C P E )  containing fine grained pyrite samples on the unleached samples and after leaching periods o f 4, 10 and 20 weeks. Pyrite reactivity profiles, supported by scanning electron microscope ( S E M ) observations and leachate chemistry data showed that minor phases o f sphalerite and galena present in the pyrite samples were the most important parameters affecting pyrite reactivity in the initial leaching cycles. Sphalerite and galena were found to effectively retard the oxidation o f pyrite in the early leaching cycles. A s sphalerite and galena were leached out, an increase i n the reactivity o f pyrite was observed, followed by a gradual loss o f reactivity from precipitate coatings.  A t a fundamental level, mineral  surface characterization by cyclic voltamperometry was found useful i n the interpretation of kinetic test data for the prediction o f acid rock drainage ( A R D ) generation.  For the backfill weathering study, paste backfill samples o f 4 different mines were leached i n deionized water (pH 5.5) in flooded and alternating air-flooded environments ii  and i n a simulated A R D solution (Fe=500 mg/1, S04=1.5 g/1 and p H 2.5) for 20 weeks. S E M , solid phase chemistry, paste p H , acid-base accounting measurements and leachate chemistry were also used to document the weathering characteristics o f cemented paste backfill ( C P B ) .  This study revealed that hydrated portland cement minerals are p H  sensitive and highly soluble.  Short-term exposures o f portland-CPB to circum-neutral  water or to A R D solution promoted the dissolution o f the binder material, increasing the porosity o f the backfill and further infiltration o f aqueous solution. Long-term exposure or flooding o f C P B was found to promote the precipitation o f secondary, expansive minerals such as gypsum i n addition to solubilizing primary cement minerals. Detailed chemical analyses and acid-base accounting indicated that the neutralizing potential added to the material by the cement phase is short-lived and the small volumes added are insufficient to neutralize the acid generating potential o f the mixture. A l l A R D solutionleached C P B samples formed an increasingly thick crust o f precipitates that, with time, reduced the ability o f the C P B to neutralize the A R D solution  111  TABLE OF CONTENTS  ABSTRACT  JI  TABLE OF CONTENTS  IV  LIST OF TABLES  IX  LIST OF FIGURES  XI  LIST OF PLATES  XVII  ACKNOWLEDGEMENT 1  INTRODUCTION 1.1  1.2  1  1.1.1  Pyrite Reactivity Study  1.1.2  Chemical Stability of Cement Paste Backfill  1  BACKGROUND  Pyrite Reactivity Study  7  Pyrite Oxidation Controls  9  \.2.\2  Electrochemistry Applied to the Study of Pyrite Reactivity  11  Cemented Paste Backfill in the Mining Industry  12  1.2.2.1  Cement Chemistry, Hydration and Chemical Stability  14  1.2.2.2  CPB in the Mine Environment  17  MINE SITE DESCRIPTIONS 2.1  3 7  1.2.1.1  1.2.2  2  1  P R O B L E M DEFINITION A N D OBJECTIVES  1.2.1  3  XXIII  26  C O M P A R I S O N OF M I N E SITE A N D E X P E R I M E N T A L CONDITIONS  EXPERIMENTAL METHODOLOGY  28  29  iv  3.1  3.2  PYRITE EXPERIMENTS  3.1.1  Sample Preparation and Analyses  3.1.2  Leaching Apparatus  30  3.1.3  Chemical analyses  32  3.1.4  Mineralogical and Electrochemical Characterisation of Pyrite  29  PASTE EXPERIMENT  3.2.1  33 36  Paste Sample Preparation  36  3.2.1.1  Paste Components  36  3.2.1.2  Paste Mixing  37  3.2.1.3  Pre-Leaching Sample Preparation  42  3.2.2  Leaching Apparatus and Leaching Cycles  43  3.2.3  Post-Leaching Sample Preparation  44  3.2.4  Chemical Analyses of the Paste Experiment  3.2.4.1  4  29  Paste pH and Acid-Base Accounting Measurements  RESULTS - PYRITE REACTIVITY STUDY  46 48  49  4.1  M I N E R A L O G Y , C H E M I S T R Y A N D STOICHIOMETRY OF U N L E A C H E D PYRITES  49  4.2  ELECTROCHEMISTRY OF T H E U N L E A C H E D PYRITE SAMPLES  56  4.3  E V O L U T I O N OF L E A C H A T E CHEMISTRY  62  4.4  E V O L U T I O N OF PRECIPITATE C O A T I N G S D U R I N G L E A C H I N G  67  4.5  4.4.1  Huckleberry pyrite  67  4.4.2  Louvicourt Pyrites  68  4.4.3  Tizapa Pyrite  68  4.4.4  Zimapan Pyrite  69  4.4.5  Brunswick Pyrite  70  E V O L U T I O N OF PYRITE R E A C T I V I T Y  77  4.5.1  Huckleberry pyrites  77  4.5.2  Louvicourt Pyrites  78  V  5  4.5.3  Tizapa pyrite  79  4.5.4  Zimapan pyrite  79  4.5.5  Brunswick  4.5.6  Evolution ofPyrite Reactivity  80 88  RESULTS - CEMENTED PASTE B A C K F I L L STUDY  88  5.1  C H E M I C A L A N A L Y S E S OF TAILINGS, BINDER A N D P A S T E  88  5.2  A S P E C T OF C U R E D PASTE SAMPLES  91  5.3  W A T E R ABSORPTION BY T H E PASTE  92  5.4  CHEMISTRY OF P A S T E L E A C H A T E S  95  5.5  5.6  5.4.1  Flooded and Cycled Water-Leached Environments  5.4.2  Fe (S0 ) Solution-Leached Environment  5.4.3  Summary of Observations from Leachate Chemistry  5.8  4  108  3  SOLID P H A S E CHEMISTRY OF L E A C H E D P A S T E  117 120  5.5.1  Tizapa  5.5.2  Brunswick  5.5.3  Louvicourt  5.5.4  Francisco I. Madero  5.5.5  Summary of Observations from the Solid Phase Chemistry  120 '.  122 123 124  BUFFERING CAPACITY AND ACID-BASE ACCOUNTING  5.6.1 5.7  2  95  129 131  Evolution ofAcid Producing and Acid Neutralizing Potentials:  SCANNING E L E C T R O N MICROSCOPE ( S E M ) OBSERVATIONS  131 137  5.7.1  Tizapa  137  5.7.2  Brunswick  141  5.7.3  Louvicourt  143  5.7.4  Francisco I. Madero  5.7.5  Summary of Observations from Scanning Electron Microscopy  146  COMPRESSIVE STRENGTH MEASUREMENTS  164 167  vi  6  DISCUSSION 6.1  172  PYRITE REACTIVITY STUDY  6.1.1  172  Effect of Precipitate Coatings on the Passivation of Pyrite: Huckleberry and Louvicourt-1  Samples 172 6.1.2  Effect of Stoichiometry on Pyrite Reactivity: Louvicourt-1 andLouvicourt-2 Samples.... 175  6.1.3  Effect of Mineral Impurities on Pyrite Reactivity  6.1.3.1  Galvanic Protection by Sphalerite: Louvicourt-2 and Tizapa Samples  176  6.1.3.2  Effect of the Presence of Galena: Zimapan and Brunswick Samples  178  6.1.4 6.2  The Huckleberry Problem  179  C H E M I C A L STABILITY OF PASTE MIXTURES  180  6.2.1  Effect of Cementing Tailings Waste  6.2.2  Cemented Backfill Alteration with Leaching  6.2.2.1  6.2.3 7  Water-Leached Environments  7.2  8  180 182 182  Ferric Sulfate or Artificial ARD Environments  CONCLUSIONS 7.1  176  189 193  PYRITE REACTIVITY STUDY  193  7.1.1  Huckleberry Pyrite  195  7.1.2  Application of this Study to Field Investigations  CEMENTED PASTE BACKFILLLEACHING STUDY  195 197  202  RECOMMENDATIONS 8.1  PYRITE REACTIVITY STUDY  202  8.2  PASTE BACKFILL STUDY  203  REFERENCES  206  APPENDIX I - ANALYTICAL RESULTS: PYRITE LEACHATE  214  APPENDIX II - ANALYTICAL RESULTS: PASTE BACKFILL LEACHATE  219  vii  APPENDIX III - ANALYTICAL RESULTS: PASTE BACKFILL SOLID PHASE  viii  LIST OF TABLES  Table 1.1: Chemical Composition o f Normal Portland Cement (Canada type 10 or M e x i c o no. 1)  15  Table 1.2 Concrete Subjected to Sulfate Attack  Table  1.3:  Replacement  Components  o f Portland Cement  Resistance o f Concrete  20  to Improve  Sulfate 23  Table 2.1: Summary o f Sample Geological Setting and Mineralogy  26  Table 2.2 Hydrogeologic Conditions at M i n e Sites where Backfill is or w i l l be Used... 28  Table 3.1 Leachate Analyses Carried Out at U A S L P  32  Table 3.2 Leachate and Solids Analyses Carried Out at C I D T  33  Table 3.3 Paste Sample Preparation  40  Table 3.4 Paste Leaching Scenarios  43  Table 3.5 Solids Phase Analyses for the Paste Experiments  48  Table 4.1 Solid Phase chemistry o f Pyrite Samples  51  Table 4.2 Normative Mineralogy o f Pyrite Samples  52  Table 4.3 Average Stoichiometry o f Pyrites  53  ix  Table 4.4 Open Current Potentials o f Pyrite Samples (Volts, S H E )  56  Table 4.5 Relative Reactivities o f Pyrites  90  Table 5.1 Chemical Analyses o f Tailings and Binders Used in Paste Mixtures  90  Table 5.2 Final Composition o f Paste Samples  90  Table 5.3a, b Paste Water Content: 5 weeks and 20 weeks  93  Table 5.4 Tizapa - Description o f Leached Paste Sections  125  Table 5.5 Brunswick - Description o f Leached Paste Sections  125  Table 5.6 Louvicourt - Description o f Leached Paste Sections  126  Table 5.7 Francisco I. Madero - Description o f Leached Paste Sections  127  Table 5.8 a, b, c, d Paste p H and Acid-Base Accounting Test Results  133  Table 5.9 Unconfined Compressive Strength o f Paste Samples  169  Table 6.1 Dissociation Constants for Secondary Minerals o f Z n and Pb  179  Table 6.2  Solubility Products and Dissociation Constants o f Secondary Minerals  Derived from Hydrated Portland Cement and Other Reference Minerals  185  LIST OF FIGURES  Figure 1.1 Paste Backfill Leaching Environments  20  Figure 2.1 Sites Location o f Huckleberry, Louvicourt and Brunswick mines, Canada... 27  Figure 2.2  Site Location o f Tizapa, Zimapan and Francisco I. Madero ( F I M ) mines,  Mexico  27  Figure 3.1 Pyrite Leaching Apparatus  32  Figure 3.2 Preparation o f Carbon Paste Electrode ( C P E )  35  Figure 3.3 Electrochemical C e l l (front clip attached to C P E )  35  Figure 3.4: a. M i x i n g Paste Sample b. Standard Slump Test  42  Figure 3.5 Rodding Paste into M o u l d Figure 3.6 Curing o f Paste Samples  42  Figure 3.7 Leaching Cells Figure 3.8 Lab Set up, Paste Leaching Study  44  Figures 3.9 a, b, c: Paste Sample Preparation Steps for Mineralogical Observations  45  Figure 4.1 A l l unleached pyrite samples Figure 4.2 Close up view  60  xi  Figure 4.3 Galena Figure 4.4 Sphalerite  61  Figure 4.5 Pyrite Leachate p H  63  Figure 4.6 Pyrite Leachate Redox Conditions  64  Figure 4.7 Pyrite Leachate Conductivities  64  Figure 4.8 Pyrite Leachate Sulfate Concentrations  65  Figure 4.9 Pyrite Leachate Total Iron Concentrations  65  Figure 4.10 Pyrite Leachate Zinc Concentrations  65  Figure 4.11 Pyrite Leachate Lead Concentrations  66  Figure 4.12 Pyrite Leachate Copper Concentrations  66  Figure 4.13 Pyrite Leachate Arsenic Concentration  66  Figure 4.14 Huckleberry Water-Leached Pyrites Figure 4.15 Huckleberry Solution-Leached Pyrites  82  Figure 4.16 Louvicourt-1 Pyrite Figure 4.17 Louvicourt-2 Pyrite  83  Figure 4.18 Tizapa Pyrite Figure 4.19 Zimapan Pyrite  85  xii  Figure 4.20 Z o o m on Initialization points, Zimapan Pyrites  86  Figure 4.21 Brunswick pyrites Figure 4.22 Z o o m o f Initialization Points  87  Figure 4.23 A l l Pyrites, 4-week Leached  88  Figure 4.24 A l l Pyrites, 10-week Leached  Figure 4.25 A l l Pyrites, 20-week Leached  89  Figure 5.1 Superficial Oxidation of Freshly Cured Tizapa Paste Samples  91  Figure 5.2 Water Absorption i n Cyclic-Leaching (water) Environment  94  Figure 5.3 Water Absorption i n Flooded-Leaching (water) Environment  94  Figure 5.4 Water Absorption in F e ^ S O ^ Solution  94  Figure 5.5 Flooded Cells - Leachate p H Figure 5.6 Cycled Cells - Leachate p H  100  Figure 5.7 Flooded Cells - Conductivity Figure 5.8 Cycled Cells - Conductivity  100  Figure 5.9 Flooded Cells - Redox Potential Figure 5.10 Cycled Cells - Redox Potential  101  Figure 5.11 Flooded Cells - Dissolved Iron Figure 5.12 Cycled Cells - Dissolved Iron  102  xiii  Figure 5.13 Flooded Cells - Dissolved Zinc Figure 5.14 C y c l e d Cells - Dissolved Z i n c  102  Figure 5.15 Flooded Cells - Dissolved Lead Figure 5.16 Cycled Cells - Dissolved Lead  103  Figure 5.17 Flooded Cells - Sulfate Concentration Figure 5.18 Cycled Cells - Sulfate Concentration  103  Figure 5.19 Flooded Cells - Calcium Concentration Figure 5.20 Cycled Cells - Calcium Concentration  104  Figure 5.21 Flooded Cells - Calcium depletion rate Figure 5.22 Cycled Cells - Calcium depletion rate  104  Figure 5.23 Flooded Cells - Magnesium Concentration Figure 5.24 Cycled Cells - Magnesium Concentration  105  Figure 5.25 Flooded Cells - Magnesium depletion rate Figure 5.26 Cycled Cells - Magnesium depletion rate  105  Figure 5.27 Flooded Cells - Potassium Concentration Figure 5.28 C y c l e d Cells - Potassium Concentration  106  Figure 5.29 Flooded Cells - Potassium Depletion Rate Figure 5.30 Cycled Cells - Potassium Depletion Rate  xiv  106  Figure 5.31 Flooded Cells - Silicon Concentration Figure 5.32 C y c l e d Cells - Silicon Concentration  107  Figure 5.33 F e ( S 0 ) Cells - p H  112  Figure 5.34 F e ( S 0 ) 3 Cells - Conductivity  112  Figure 5.35 F e ( S 0 ) 3 Cells - Redox Potential  112  Figure 5.36 Ferric Hydroxide Coating on Paste Sample  113  Figure 5.37 F e ( S 0 ) Cells - Dissolved Iron  113  Figure 5.38 F e ( S 0 ) Cells - Dissolved Zinc  113  Figure 5.39 F e ( S 0 ) Cells - Dissolved Lead  114  Figure 5.40 F e ( S 0 ) Cells - Sulfate Concentration  114  Figure 5.41 F e ( S 0 ) Cells - Calcium Concentration  114  Figure 5.42 F e ( S 0 ) Cells - Calcium Depletion Rate  115  Figure 5.43 F e ( S 0 ) Cells - Magnesium Concentration  115  Figure 5.44 F e ( S 0 ) Cells - Magnesium Depletion Rate  115  Figure 5.45 F e ( S 0 ) Leach Cells - Potassium Concentration  116  Figure 5.46 F e ( S 0 ) Leach Cells - Potassium Depletion Rate  116  2  2  4  4  2  4  2  4  2  2  2  3  4  2  2  3  4  2  2  3  4  2  2  3  3  4  4  4  4  4  4  3  3  3  3  3  3  xv  Figure 5.47 F 62(804)3 Leach Cells - Silicon Concentration  116  Figure 5.48 Tizapa Paste - U C S with Leaching Time  170  Figure 5.49 Brunswick Paste - U C S with Leaching Time  170  Figure 5.50 Louvicourt Paste - U C S with Leaching Time  170  Figure 5.51 Francisco I. Madero Paste - U C S with Leaching Time  171  xvi  LIST O F P L A T E S  Plate 4.1 Zimapan Pyrite Plate 4.2 Brunswick Pyrite Plate 4.3 Louvicourt-2 Pyrite Plate 4.4 Tizapa Pyrite Plate 4.5 Huckleberry Pyrite Plate 4.6 Louvicourt-1 Pyrite  HUCKLEBERRY PYRITES Plate 4.7 Water-Leached, 20 weeks Plate 4.8 Water-Leached, 20 weeks Plate 4.9 Solution-Leached, 20 weeks Plate 4.10 Solution-Leached, 20 weeks  LOUVICOURT-1 PYRITES Plate 4.11 4-week leached Plate 4.12 20-week leached Plate 4.13 20-week leached  LOUVICOURT-2 PYRITES Plate 4.14 Pyrite surface at 10 weeks Plate 4.15 Pyrite surface at 20 weeks Plate 4.16 Pyrite at 10 weeks Plate 4.17  20-week leached  xvii  TIZAPA PYRITES Plate 4.18 Pyrite surface at 20 weeks Plate 4.19 Oxidized sphalerite, 20 weeks Plate 4.20 Precipitate cover, 20 weeks Plate 4.21 Oxidized galena, 20 weeks  72  Zimapan Pyrites Plate 4.22 Pyrite surface at 20 weeks Plate 4.23 Corrosion pits at 4 weeks Plate 4.24 Anglesite precipitates, 20 weeks  73  BRUNSWICK PYRITES Plate 4.25 Pyrite surface at 20 weeks Plate 4.26 Pyrite corrosion pits, 20 weeks Plate 4.27 Anglesite and galena, 4 weeks  74  PASTE BACKFILL PLATES Plate 5.1 Portlandite crystals Plate 5.2 Typical E D X o f portlandite Plate 5.3 Secondary gypsum crystals Plate 5.4 Typical E D X o f gypsum  149  Plate 5.5 Acicular ettringite Plate 5.6 Typical E D X o f ettringite  150  xviii  Plate 5.7 Amorphous tobermorite Plate 5.8 Typical E D X o f amorphous tobermorite  150  TIZAPA BACKFILL Plate 5.9 Portlandite cover on sample surface Plate 5.10 Bottom o f layer 1 Plate 5.11 Bottom o f layer 1 Plate 5.12 Poorly developped Tobermorite, below layer 1 Plate 5.13 Core o f sample Plate 5.14 Primary ettringite on incomp-letely hydrated cement grain  151  Plate 5.15 Poorly developped Tb, layer 1 Plate 5.16 Ettringite, upper layer 1 Plate 5.17 Secondary gypsum, upper layer 1 Plate 5.18 General paste aspect, layer 1 Plate 5.19 Ettringite i n core o f sample Plate 5.20 Good tobermorite development at core o f sample  ..152  Plate 5.21 Porous layer 1 Plate 5.22 Depleted tobermorite, layer 1 Plate 5.23 Area o f good tobermorite, layer 1 Plate 5.24 Layer 1, iron sulfate precipitate Plate 5.25 E D X o f iron sulfate precipitate  Plate 5.26 Good tobermorite cover, layer 2  154  xix  BRUNSWICK BACKFILL Plate 5.27 Layer 1 Plate 5 28 Layer 1 Plate 5 29 Less resistant tobermorite, layer 1 154  Plate 5 30 Abundant tobermorite, layer 2  Plate 5.31 G o o d tobermorite cover, layer 1 Plate 5.32 General aspect o f paste, layer 1 Plate 5.33 Paste surface Plate 5.34 General aspect o f paste, layer 1 Plate 5.35 Poor tobermorite cover, layer 1 Plate 5.36 Possible 2ry ettringite, layer 1  155  Plate 5.37 General aspect o f paste, layer 2 Plate 5.38 Fe-sulfate precipitate, layer 2 Plate 5.39 Inside large pore, layer 2 Plate 5.40 General aspect o f paste, layer 3 Plate 5.41 Layer 3  156  LOUVICOURT BACKFILL Plate 5.42 Massive tobermorite Plate 5.43 Dendritic tobermorite Plate 5.44 E D X o f tobermorite developed from slag Plate 5.45 General aspect o f paste mixture  157  xx  Plate 5.46 Less tobermorite in upper layer 1 Plate 5.47 M o r e abundant tobermorite i n lower layer 1  Plate 5.48 Poorly developed Tb, upper layer 1 Plate 5.49 Upper layer 1 Plate 5.50 More developed Tb, lower layer 1 Plate 5.51 Secondary gypsum, lower layer 1 Plate 5.52 Tobermorite-lined fracture in paste Plate 5.53 Close-up view o f tobermorite  Plate 5.54 Depleted tobermorite, layer 1 Plate 5.55 Masses o f tobermorite at upper layer 1 Plate 5.56 Increased Tb content, below layer 1 Plate 5.57 Fracture developed possibly along existing plane o f weakness  Plate 5.58 Botryoidal mass o f iron sulfate precipitate Plate 5.59 E D X o f iron sulfate precipitate (type o f jarosite?) Plate 5.60 Possibly ettringite, layer 2 Plate 5.61 W e l l developed tobermorite, lower layer 2  FRANCISCO I. MADERO BACKFILL Plate 5.62 Layer 1, flooded sample Plate 5.63 Core o f flooded sample Plate 5.64 Layer 1, hydrated cement phase  xxi  Plate 5.65 General aspect o f paste Plate 5.66 Good development o f tobermorite, layer 1 Plate 5.67 Layer 1 Plate 5.68 Monosulfoaluminate (?), below layer 1 Plate 5.69 Core o f sample  xxn  ACKNOWLEDGEMENT  This project was the first o f an international collaboration and technology exchange between the University o f British Columbia and the Universidad Autonoma de San Luis Potosi. Although the logistics for this project were sometimes challenging, the project was carried out with great success thanks to the efforts o f many determined people. In particular I would like to sincerely thank my supervisor and project co-ordinator D r Richard Lawrence for his time and attention, D r Marcos Monroy, D r George^ Poling and D r Ignacio Gonzalez for their supervision, valuable comments and discussions throughout the study and for their support and friendship. I would also like to thank D r Marcello Veiga, and D r Ralph Hackle for their comments and revisions. Funding o f this research was provided by a grant from the National Science and Engineering Research Council o f Canada ( N S E R C ) , by Noranda Technology Centre ( N T C ) , Centra de Investigation y Desarollo Tecnologico (CIDT) o f Pefioles, the University o f British Columbia, the Universidad Autonoma de San Luis Potosi, Universidad Autonoma Metropolitana-Ixtapalapa o f M e x i c o City. In particular, I would like to thank L u c St-Arnaud and Michael L i o f N T C , Louis Racine o f Louvicourt mine and A r i e Moerman o f Brunswick mining division, Noranda. Thanks also goes out to Benjamin R u i z and Carlos Lara o f C I D T , and Peter Campbell formerly o f Huckleberry mines. I thank Sally, Frank, Marina, Terry, Shannon, Pary, Dr. Dunbar and D r . Scoble at the department o f M i n i n g and Mineral Process Engineering o f U B C for their support. M e gustaria agradecer a Irene Chavira y Roel Cruz-Gaona por su invaluable ayuda con los experimentos que lleve a cabo en Mexico, y tambien por su paciencia y amistad. Tambien agradezco a Ulices, Pancho, Israel, Lorena, Teresita y Eliceo por sus atenciones y, claro, su amistad. Finalement, j'aimerais remercier du fond du coeur Stephane D ' A o u s t et Therese Bertrand ainsi que Kathy et Tracy pour leur soutien moral, encouragement et amitie.  xxiii  1  1.1  INTRODUCTION  P R O B L E M DEFINITION AND OBJECTIVES  The oxidation o f sulfide-rich mining wastes produces drainage water o f poor quality, contaminated by dissolved heavy metals and which is often o f l o w p H .  This  contaminated drainage is termed acid rock drainage ( A R D ) and is the most costly environmental problem facing the mining industry today.  In many cases, the only  solution is long term treatment once the process o f oxidation is under way.  A key  mineral in A R D generation is pyrite ( F e S 2 ) which, although not the most reactive sulfide mineral, is by far the most common, frequently present in large quantities i n mining wastes.  This study has been carried out to observe the electrochemical behaviour o f  pyrite oxidation to determine i f such fundamental observations might be used to refine the prediction o f the behaviour o f pyritic mine wastes.  Building on the knowledge o f  pyrite reactivity, experiments have been carried out to observe the chemical stability o f paste backfill  containing pyritic tailings, exposed to various leaching conditions  encountered in mine settings.  1.1.1  Pyrite Reactivity Study  A considerable amount o f research has been carried out in the last two decades to improve the understanding o f mineral waste oxidation and provide methods to control the generation o f A R D .  Fundamental studies o f pyrite oxidation such as the processes  involved, oxidation rates and reaction products, abound i n the literature (Singer and  1  Stumm, 1970; Lowson, 1982; Moses et al., 1987; Brown and Jurinak, 1989; Nicholson et al, 1989; Moses and Herman, 1991; Nicholson, 1994; Blowes et al, 1995; Eidsa et al, 1997).  The application o f this fundamental knowledge to field conditions, to predict  drainage water quality, remains difficult because o f the heterogeneous nature o f mineral waste piles and the multitude o f chemical reactions and physical factors that can affect the generation o f A R D and subsequent water quality.  The mining industry has typically relied on the use o f simple and often short-term laboratory weathering tests such as humidity cells to attempt to predict the behaviour o f mining wastes exposed to the environment and quality o f the drainage water.  An  increasing number or government agencies require that specific static and kinetic tests be carried out on mineral wastes, prior to permitting a new mining operation to determine the potential o f waste to generate acid. Mineral waste characterization programs required by the government o f British Columbia, Canada, are presented i n the B C A M D Draft A c i d Rock Drainage Technical Guide (Steffen, Robertson and Kirsten L t d . , 1989), with revisions by Price and others (1997). The limitation to kinetic tests is that only products are measured without detailed information on sulfide oxidation kinetics or the possible change in oxidation rate, for example, with the evolution o f precipitation products.  In  addition, the data generated by some o f these tests has been shown to vary depending on the procedures followed for a given tests (Lawrence and Wang, 1997), and to be difficult to extrapolate to predict the.future chemical behaviour o f mine wastes and the quality o f the leachates derived from them (Bethune et al., 1997; Otwinowski, 1997).  Inaccurate  predictions on A R D generation can have costly consequences such as over-design o f treatment facility or worse yet, unplanned environmental restoration costs due to design  2  failures. Reduction i n risks and costs associated with the management o f mining wastes is the objective o f the great amount o f research actively being carried out to improve prediction techniques.  In the first part o f this thesis, cyclic voltamperometry was used to study the effects o f oxidative leaching on the reactivity o f pyrite. Cyclic voltamperometry is an established investigative tool used in electrochemical studies to characterize  surface  and/or  semiconducting properties o f metals and minerals. Electrochemical techniques o f various kinds have been used extensively i n the mineral processing industry (review by Peters, 1984). C y c l i c voltamperometry was used in conjunction with mineralogical investigation techniques to characterize the relative reactivity o f various pyrite samples, and the evolution o f reactivities as leaching progressed. The study was carried out with the aim to document the influence o f intrinsic and extrinsic factors on the reactivity o f pyrite. The objectives o f this first part o f the study are:  1) Investigate the usefulness o f this relatively rapid and simple technique as an acid rock drainage predictive tool,  2) Measure the initial effects  o f leaching on pyrite to attempt to improve the  understanding and subsequent interpretation o f humidity cell data i n the early stages of leaching.  1.1.2 Chemical Stability of Cement Paste Backfill Due to the uncertainty associated with oxidation o f mining wastes and the lack o f failproof, effective protection against A R D , especially for sulfide-rich wastes, the mining  3  industry is increasingly pushing to minimize the surface disposal o f waste to try to avoid the possible environmental problems associated with that practice.  A n option that is  gaining popularity in the mining industry is the use o f total tailings, including sulfide-rich tailings, as backfill material within the mine. Backfill is commonly used i n mining operations to provide underground support and improve ore recovery, and as a method to dispose o f some o f the waste generated. Recent technological developments allow the use o f total tailings, including the fine portion, which has traditionally been avoided because o f excessive water retention and associated backfill stability problems.  Cement  can be added in small proportion to the mixture to increase the short and/or long term strength o f the backfill (Landriault et al., 1998). This material is referred to as cemented paste backfill ( C P B ) .  Advantages to using cemented, total tailings backfill include the decrease i n volume o f waste to be disposed o f on the surface, thereby reducing the liability associated with the long-term care o f a tailings disposal facility and the possible environmental problems associated with the oxidation o f that waste in the case o f reactive tailings.  Another  reported advantage is the added neutralization capacity provided by the cement to reactive tailings.  The neutralized tailings could potentially serve a dual purpose:  preventing the oxidation o f the waste and the subsequent generation o f A R D , as well as neutralizing existing A R D actively produced within the mine when coming i n contact with the backfill material.  Backfill stability studies generally focus on the rheological or geotechnical properties o f the backfill material. M a n y investigations are carried out to evaluate lower cost mineral additives to replace portions o f Ordinary Portland Cement ( O P C ) while preserving the  4  strength o f the cured material and into improvements in the short and/or long term strength o f the mixtures (Hopkins and Beaudry, 1989; L o r d and L i u , 1998; Gay and Constantiner, 1998; Noranda Technology Centre, 1998a,b,c,d; Archibald and Chew, 1998). Very few studies have been carried out on the reactivity or chemical stability o f cured paste mixtures exposed to various environmental conditions. It is well known i n the concrete industry that fine, sulfidic aggregate has deleterious effects on the setting ability o f hydrated cement and its long-term durability. Furthermore, high concentrations o f sulfate such as those commonly present in the tailings water used to make the mixtures or present i n the mine waters (or A R D ) coming i n contact with the backfill are recognized in the concrete industry as aggressive solutions. Sulfates react with the hydrated cement to produce expansive minerals that cause the material to crack and lose its strength.  The  resulting physical weakening o f the backfill concrete is highly undesirable, especially when the backfill is used to provide physical support o f the underground structures.  The  physical breakdown o f the backfill can also result i n the exposure o f reactive sulfide particles to the environment, thereby increasing the risk o f oxidation.  In the second part o f this study, the chemical stability o f cemented paste mixtures that contain reactive waste was studied following 20-week leaching periods i n environments similar to those encountered in a mine setting. The principal objective o f the paste backfill stability study was to document the changes in properties occurring within various pastes when exposed to these leaching environments.  The objectives o f this second part o f the study were to provide insights on questions and actual problems reported by mines currently using or planning to use C P B , namely:  5  1) Decreasing backfill strength with time resulting i n higher than expected dilution o f ore when blasting next to backfilled areas,  2) Effectiveness o f C P B to neutralize reactive tailings used in the backfill mixture and to buffer A R D produced within the mine.  3) The feasibility o f using reactive tailings i n above-ground applications o f C P B  In the first part o f the study, 6 samples o f pyrite from mines i n Canada and M e x i c o (Huckleberry, Louvicourt samples 1 and 2, Brunswick and Zimapan, Tizapa respectively) were studied using cyclic voltamperometry, scanning electron microscope ( S E M ) , energy dispersive x-ray analysis ( E D X ) , chemical analyses o f solids and o f the leachate to obtain information on the reactivity o f each pyrite i n the early cycles o f leaching. In the second part, 4 paste backfill samples were prepared using the formulation specified by each mine (Louvicourt, Brunswick, Tizapa and Francisco I. Madero) to study their weathering characteristics when subjected to various leaching environments. S E M , E D X , acid-base accounting analyses as well as o f solid phase and leachate chemistry was used for this investigation.  In the remaining section o f this chapter, the reactivity o f pyrite w i l l be discussed followed by a detailed review o f the use o f paste backfill in the mining industry and the potential problems associated with exposure to environments encountered in a mine setting. Chapter 3 presents the methodology and sample preparation procedures developed for both the pyrite and backfill studies.  Chapter 4 presents the results, analyses o f data,  6  together with brief summaries o f observations. A discussion o f the results is presented i n Chapter  5.  Chapter  6  presents  the  conclusions o f the  study  together  with  recommendations for further investigations and which could benefit similar studies.  1.2  1.2.1  BACKGROUND  P y r i t e Reactivity Study  Pyrite oxidation reactions and their mechanisms are generally well understood.  The  overall reactions can be summarized as follows:  Initial circum-neutral oxidation o f pyrite:  FeS  + H 0 + / 0 7  2  2  2  Slow, rate limiting Fe  2Fe  2 +  + V 0 2  -> F e + 2 S 0 ~ + 2 H 2 +  2  2  (1.1)  +  4  oxidation (more rapid, bacterially mediated at p H < 3.5):  2  + 2H  +  -> 2 F e  3 +  + H 0  (1.2)  2  Ferric iron oxidation o f pyrite at p H < 3.5:  FeS  2  + 14 F e  3 +  + 8 H 0 -> 15 F e  + 2 S 0 " + 16 H  2 +  2  2  4  +  (1.3)  Although pyrite may not be the most reactive sulfide mineral, the complete oxidation o f one mole o f pyrite i n a bacterially mediated environment p H < 3.5 can liberate 16 protons and a considerable amount o f dissolved metals.  B y comparison, the oxidation o f one  mole o f a mono sulfide mineral such as sphalerite (ZnS), galena (PbS) and covellite (CuS) w i l l liberate one mole o f dissolved metal ( M e ) , one mole o f sulfate but no protons: 2+  7  MeS + 2 0  2  ^  Me  z+  + S 0 2-  (1.4)  4  Under circum-neutral p H conditions, dissolved metal w i l l combine with hydroxide to precipitate sparingly soluble metal hydroxide according to equation 5, slowly decreasing the activity o f O H - i n solution.  Me  / +  + 2 OH  > Me(OH)  (1.5)  2  Reactions (1.1) to (1.5) show that, once started, the oxidation o f sulfide minerals becomes difficult to stop or slow down without costly control measures. For this reason, the key to successful management o f potentially acid-generating waste is to prevent the initiation o f oxidation reactions.  Static and kinetic tests have been devised to provide information on the potential o f mineral wastes to oxidize, the time o f initiation o f acid generation and the expected loading o f metals and low p H water to the receiving environment.  This information is  obtained from the interpretation o f the leachate chemistry data, from the comparison o f sulfide oxidation rates and carbonate mineral depletion rates measured during leaching. Extrapolation o f the calculated rates and ion loading suggests a time frame for acid generation: onset, duration and metal loading to receiving waters.  In many cases, however, leachate data yields data that can be difficult to interpret and, therefore, make predictions with any degree o f certainty.  Such is the case for the low-  grade ore waste o f the new Huckleberry mine in British Columbia, one o f the pyrites studied i n this project.  Kinetic test results for Huckleberry low-grade ore  was  inconclusive as to the onset o f oxidation o f sulfide minerals contained i n the rock. Static  8  tests c a r r i e d out o n this material s h o w e d a n e u t r a l i z a t i o n potential ratio ( N P R ) o f 0.78, characterizing  it as likely  acid  according  generating  to current  British  Columbia  G u i d e l i n e s ( P r i c e et al., 1997). D u r i n g 3 years o f k i n e t i c ( c o l u m n leach) tests, h o w e v e r , p H , sulfate a n d c a l c i u m levels r e m a i n e d elevated ( L a w r e n c e , 1997). T h e h i g h sulfate a n d c a l c i u m concentrations w e r e attributed to the d i s s o l u t i o n o f the h i g h g y p s u m content o f the m a t e r i a l .  C o n s e q u e n t l y , n o sulfide o x i d a t i o n rates o r n e u t r a l i z a t i o n d e p l e t i o n rates  w e r e extractable f r o m the leachate data, s u c h that n o p r e d i c t i o n s w e r e p o s s i b l e as to the onset o f A R D o r e x p e c t e d m e t a l l o a d i n g .  A s a result, the m i n e h a d to assume the  m a t e r i a l to be reactive a n d dispose o f the material into the t a i l i n g s p o n d rather t h a n use it as c o n s t r u c t i o n m a t e r i a l o f the inner p o n d w a l l where it w o u l d b e c o m e f l o o d e d w i t h i n 8 years.  In l i g h t o f the costly d i s p o s a l alternative f o r this material o f u n c e r t a i n A R D p o t e n t i a l , the first part o f this thesis studied the reactivity o f pyrite (the p r i n c i p a l sulfide m i n e r a l i n the waste) i n b o t h gypsum-saturated  a n d gypsum-free  environments,  e v o l u t i o n o f the r e a c t i v i t y o f pyrite as l e a c h i n g progressed.  to d o c u m e n t  the  T h e results w e r e e x p e c t e d to  h e l p evaluate the p r o b a b i l i t y o f the waste to o x i d i z e w i t h i n the t i m e frame o f 8 years e x p o s e d to air a n d water.  F i v e other pyrites w e r e studied a l o n g w i t h H u c k l e b e r r y to  evaluate the effects o f m i n e r a l o g i c a l characteristics o n the r e a c t i v i t y o f p y r i t e i n general.  1.2.1.1 Pyrite Oxidation Controls S u l f i d e m i n e r a l o x i d a t i o n reactions are d o c u m e n t e d as surface c o n t r o l l e d , based o n the a v a i l a b i l i t y o f reactive sites to participate i n the exchange o f charges w i t h a n o x i d i z i n g agent s u c h as o x y g e n o r ferric i r o n i n the case o f pyrite ( L o w s e n , 1 9 8 2 ; M o s e s et al.,  9  1987; B r o w n and Jurinak, 1989; Nicholson et al,  1988; Moses and Herman, 1991;  Nicholson, 1994). Bacterial oxidation o f pyrite, for example by T. ferrooxidans, is also documented to be surface controlled. The oxidation o f ferrous iron and/or sulfide was found to be facilitated by the attachment o f the bacteria to the mineral surface (Herrera et al, 1989; Free et al., 1993).  When pyrite is exposed to air and water, the oxidation  products o f pyrite produce coatings that decrease the surface area o f the grain available for oxidation, effectively decreasing the rate o f pyrite oxidation.  The occurrence o f  precipitate coatings on pyrite and its effect o f decreasing reaction rates are documented by Nicholson and others (1990).  Thick precipitate coatings o f iron hydroxide, iron  oxyhydroxides and jarosite on oxidized pyrite grains are also documented by Jambor (1994), Bigham (1994) and Alpers and others (1994).  Oxidation being a surface controlled reaction, the rate o f oxidation o f a pyrite is documented to be influenced by the mineralogy, stoichiometry, crystal morphology and defects o f pyrite crystals. M c K i b b e n and Barnes (1986), K w o n g (1993) and K w o n g and Lawrence (1994) have indicated that the relative abundance o f physical or chemical defects and the occurrence o f mineralogical impurities associated with pyrite, influence the distribution o f surface free energy on pyrite grains.  Locations o f higher energy  created by defects are more likely to oxidize than grains or parts o f grains having lower surface energy. These researchers conclude that oxidation rate o f pyrite depends heavily i  on the mineralogical characteristics o f the mineral. Little is known, however, on h o w these parameters interact to influence the reactivity o f pyrite. K w o n g (1994) suggested a theoretical order o f importance o f mineralogical factors influencing the rate o f sulfide oxidation based on laboratory weathering tests. Measurements o f local redox potentials  10  gave indications o f local reactivity but did not provide a bulk measurement o f the reactivity o f a pyrite sample, which could be compared to that o f another sample.  Bulk  reactivity measurements on a bulk sample o f pyrite (averaging all defects) would provide information on the acid generation potential o f pyrite that may be closer to field conditions.  1.2.1.2 Electrochemistry Applied to the Study of Pyrite Reactivity Electrochemical oxidation o f pyrite combined with surface spectroscopy studies has identified kinetic processes o f oxidation in various ionic solutions and p H environments. The application o f cyclic voltamperometry to hydrometallurgical studies is reviewed by L i and others (1992). Buckley and others (1988), Wadsworth and others (1993), and L i and Wadsworth (1993) used cyclic voltamperometry to document the formation o f sulfur, or polysulfide layers on the surface o f pyrite following electrode oxidation i n slightly acidic or neutral p H solutions.  These coatings were found to effectively decrease the  leachability or reactivity o f the underlying pyrite.  Doyle and M i r z a (1996) used cyclic voltamperometry to characterize pyrites from different sources to evaluate the effects o f pyrite composition and electric properties (rest potential, resistivity, net concentration o f donors and charge carrier concentration) on the oxidation behaviour o f pyrite. They found poor correlation between chemical or electric characteristics and pyrite reactivity using their methodology.  Their working electrodes  consisted o f polished pyrite grain, documented in later studies to be largely affected by polishing o f the grain, with responses often dominated by fractures or irregularities i n the surface o f the grain-electrode.  Lazaro and others (1995) indicated that electrochemical  11  responses obtained from carbon paste electrodes ( C P E ) - a mixture o f pyrite, graphite and non-conducting silicon o i l - are average responses o f the many grains exposed at the electrode surface, including those given by grain-surface impurities and crystal defects. The averaging effect o f C P E was found to produce highly repeatable results.  C y c l i c voltamperometry using C P E was used in this study to document the reactivity o f six different pyrite samples and try to correlate the mineralogical characteristics o f each sample with its measured reactivity. The evolution o f pyrite reactivity with increasing leaching time was also studied by that method.  1.2.2  Cemented Paste Backfill in the Mining Industry  Backfill can consist o f any mining residue depending on the purpose o f the backfill; from large waste rock blocks to the coarser portion o f mineral processing residues (tailings) or a mixture o f all rock sizes. The very fine size portion o f tailings has traditionally been avoided as the water it contains drains out more slowly, resulting i n backfill stability problems.  Mineral processing fines are typically discarded to the tailings containment  areas. Decreasing grades o f mineral deposits now being exploited has resulted, however, in increasing volumes o f mineral processing waste to be disposed of. Shortage o f land or resources to build additional tailings impoundments is forcing the mining industry to evaluate ways to reuse the additional waste being generated. Recent technological advancements i n tailings dewatering have permitted the inclusion o f larger proportions o f fines i n backfill without the problem o f excessive bleeding (drainage o f excess water from the mixture)  (Cincilla et al, 1997; Dahlstrom, 1997; Williams, 1997).  Backfill  containing a high proportion o f fines is a popular solution to the problem o f increasing  12  mineral processing wastes. Paste backfill refers to the thick yet fluid consistency o f the fresh mixture, which is specially designed for pipe flow and easy emplacement i n stopes of various sizes and shapes. Contrary to plain backfill, paste fill can contain total tailings, which include a considerable proportion o f fines. To facilitate pumping and placement o f the mixture and minimize erosion o f the pipe used for the transport o f the paste, it is recommended that the backfill mixture contain a minimum o f 15 to 20 % solids finer than 20 urn (Landriault et al, 1998).  Cement can be added to paste mixtures to decrease bleeding o f the backfill once i n place and/or to increase the strength o f the backfill upon curing (Landriault et al, 1998). One of the first uses o f cemented backfill in the mining industry occurred at the B H P Mount Isa mine i n Australia where, since the early 1930's, large blocks o f waste rock were thrown into a vertical shaft along with hydrolysed cement to fill open stopes and accommodate their particular mining sequence. A n overview o f the Canadian experience with the various types o f backfill is given by U d d (1989).  Another advantage o f adding cement to paste backfill is the highly alkaline composition of cement material, which provides additional neutralizing capacity to tailings that are potentially acid generating, such as pyritic tailings (Levens and Boldt, 1994). Ideally, this waste management practice could alleviate the need to dispose o f reactive waste i n an engineered impoundment where, apart from the construction, maintenance and land costs, the reactive waste could oxidize and become an expensive environmental liability.  13  1.2.2.1 Cement Chemistry, Hydration and Chemical Stability Cement is one o f the most commonly used materials in the construction industry and much is already known about its chemistry, its strengths and weaknesses and its interaction with a variety o f aggregates (Taylor, 1997). Under normal circumstances, the inherent chemical stability, physical strength and workability o f concrete allows its use i n a variety o f settings. Table 1.1 describes the various elements that make up Ordinary Portland Cement ( O P C ) used i n C P B , along with the principal hydration reactions involved i n the curing o f cement mixtures.  The cement industry uses abbreviated  nomenclature for the unhydrated cement components such that the oxide in the mineral phases is referred to by one letter:  C=CaO, S = S i 0 , 2  A=A1 0 , 2  3  F=Fe 0 , S=S0 , 2  3  4  H = H 0 . In the cement literature, the alite phase o f cement is written C S , referring to the 2  3  composition: 3 C a O - S i 0 , or, in analytical chemistry, CasSiOs. This terminology is used 2  in parts o f Table 1.1 and in subsequent text to permit the correlation with the cement literature.  Some agents are known to interact destructively with the components o f the concrete, undermining its integrity. O f these, reactive mineral aggregates (i.e. sulfidic aggregates), high concentrations o f sulfate i n the mixing water or excess atmospheric C 0  2  upon  drying make up the principal destabilizing agents o f concrete (Kosmatka et al. 1995). These processes are briefly described as:  14  Table 1.1: Chemical Composition of Normal Portland Cement (Canada type 10 or Mexico no. 1) Principal components of dr\ ci-im-nl tricalcium silicate: C S C a S i 0 (alite)  T\pir:il portion  Principal hydration pioducts  -50 %  1) 2 C S + 6 H 0 ^ C a S i 0 - 3 H 0 + 3Ca(OH) >  3  3  5  3  2  (a)  Rapid hardening, early strength development  3  2  7  (b  2  2  70% reacted in 28 days  dicalcium silicate: C S C a S i 0 (belite)  Function  "Tobermorite gel: principal binding agent of cement Portlandite: no cementing properties  b)  2  2  -25 %  4  2  6  2  3  2  2  3) C A + 12 H 0 + Ca(OH) -> Ca Al 0 (OH) 12H 0 3  -10%  2  4  2  6  2  2  4  2  4  2  4  4  Monosulfoaluminate and Ettringite: expansive minerals produced from the reaction of dissolved gypsum with C A .  e)  2  3  2  4  2  (e)  2  4  2  2  6) C A F + 10 H 0 + 2 Ca(OH) -> Ca Al Fe 0 12H 0 4  -8%  3  2  2  ( 0  6  2  2  14  2  1  Gypsum: C S H CaS0 -2H 0  d)  5) C A + 26H 0 + 3 CaS0 -2H 0-> Ca Al (S0 ) (OH), -26H 0 6  2  2  (d)  2  Tetracalcium aluminate hydrate: some early strength development.  2  4) C A + 10 H 0 + CaSCv2H 0 -> Ca Al (S0 )12H 0  Consumes Ca(OH) , produces a high heat of hydration  c)  2  (c)  3  tetracalcium aluminoferrite: C A F Ca4Al Fe 0 o Manufacturing purpose to reduces clinkering T°, responsible for colour effects in cement  Same as above  2  30% reacted in 28 days, 90% in 1 year  3  3  2  (b  Slow hardening, later stage strength development (after 1 week) tricalcium aluminate: C A Ca Al 0  2) 2 C S + 4 H 0 -> C a S i C y 3 H 0 « + Ca(OH) >  3  °Calcium aluminoferrite hydrate: rapid hydration but little strength contribution.  7) C A F + 50 H 0 + 6 CaS0 -2H 0 + Ca(OH) -> 2 Ca«Al (S0 ) (OH) -26H 0 4  2  4  2  2  (e)  2  4  3  l2  2  2  ~5 %  2  Very rapid dissolution, slows the rate of C A hydration to avoid flash setting 3  Fe, K , M g  Dissolved gypsum may participate in reactions (4), (5), (6) and (7), depending on the local pore water chemistry.  Too much gypsum may favour the formation of ettringite over portlandite.  May be included in any cement phase in solid solution.  Few %  Present in clays used to make Portland cement. Notes:  Slow hydration reaction to produce ettringite  This table adapted from Kosmatka 1995 and Taylor 1997 tricalcium silicate hydrate or "tobermorite gel": composition may vary and may include trace concentrations of Fe, Mg, K.  (1)  15  Carbonation: Carbonation refers to the excessive shrinkage o f concrete upon drying caused by the penetration o f atmospheric C O 2 , transforming hydroxides to carbonates. These reactions lower the alkalinity o f concrete, destabilizing the curing process.  H i g h waterxement  ratios, l o w cement content and/or short curing period enhance the potential for carbonation to occur. This phenomenon is normally restricted to shallow depth or at the surface o f the concrete.  Alkali-aggregate reactions: The reaction between reactive mineral aggregates used in the concrete mixture and the sodium and potassium alkalis present in the cement cause expansive secondary mineral growth. Growths o f the secondary minerals create internal stresses within the concrete causing it to crack and lose its strength.  Sulfate attack: The interaction o f the sulfate ion present in the pore water with the hydrated compounds o f the cement also results in expansive secondary mineral growth (gypsum and/or ettringite) combined with disintegration o f the primary binding material (tobermorite gel).  Water dissolution: Another deleterious agent o f particular importance i n the backfill environment is the interaction o f neutral p H water with concrete, dissolving and leaching out some o f the p H sensitive or water-soluble components o f hydrated cement such as portlandite (Ca(OH)2)  16  or tobermorite gel (Ca3Si207"3H20). This can be particularly deleterious when contact occurs during the curing period o f the freshly mixed cement (Adenot and B u i l 1992; Carde and Francois 1997).  1.2.2.2 CPB in the Mine Environment The utilization o f cement in the mining industry as a binder o f tailings i n paste backfill is a unique application as a very small proportion of cement is normally used to bind tailings, commonly less than 10 or even 5 % o f dry weight, when common concentrations o f cement i n concrete range from 30 to 40%. In addition, water to cement (w/c) ratios, important i n the hydration and subsequent curing o f cement, are also considerably higher in C P B applications compared to normal concrete. This ratio is expressed as the mass o f water divided by the mass o f cementing materials. Kosmatka and others (1995) explain that lower w/c ratios provide the greatest unconfined compressive strength (ucs) i n normal concrete. Lamos and Clark (1989) come to the same conclusion with respect to tailings backfill.  Normal concrete w/c ratios, typically around 0.5, yield 28-day ucs  values ranging between 25 and 35 M P a . Typical C P B applications, such as the four mines studied i n this work, have w/c ratios ranging from 5 to 10. C P B mixtures can, therefore, be expected to develop poorer ucs values upon curing than concrete containing the same amount o f cement.  In addition, the tailings or aggregate mixed with the cement to form paste backfill is regarded i n the cement industry as undesirable aggregate because o f its very fine grain size and most importantly, because o f its composition, i n the case o f high sulfide tailings. The Canadian Standards Association (Standard A23.1) specifies the lower and upper  17  limits o f grain size that should make up the fine size portion o f aggregate. Standard A23.1 specifies a lower limit o f 5 to 10% passing 160um and upper limit o f 80 to 100% passing 2.5 mm. Furthermore, the Canadian Portland Cement Association ( C P C A ) (Kosmatka et al., 1995) indicates that the fine aggregate content should be no larger than 4 5 % by mass or volume o f the total aggregate content, above which the cement cannot efficiently coat all aggregate particles.  When a large proportion o f small size particles are used as  aggregate, the mixture w i l l require a larger cement content to effectively coat all the particles and meet specified strength requirements. In the case o f C P B , the tailings used as aggregate are often finer than 150u.m. Consequently, the cement added to the paste mixture can be expected to underperform in terms o f strength development upon curing compared to a similar cement proportion used i n conjunction with standard aggregate.  Pyritic aggregate is normally avoided i n cement mixtures because o f its reactivity and the consequent production o f sulfates.  Sulfate is a documented aggressive agent that  participates i n expansive secondary mineral growths within the concrete (Shayan, 1988; DeCeukelaire, 1991; Idorn, 1992; Casanova et al., 1996). These reactions create internal stresses that lead to cracking and disintegration o f the cured concrete.  The details o f  these reactions are discussed in the later paragraphs.  In summary, the constituents o f paste backfill are far from ideal to provide a mixture o f optimum compressive strength. The strength requirement for paste backfill is, however, commonly much lower than for a building material. In the four cases studied, the ucs requirement range between 0.5 and 3.5 M P a , depending on the purpose o f the backfill. Indeed, backfill used to fill empty voids surrounded by rock mass requires minimal  18  strength, enough to prevent liquefaction when blasting other areas o f the mine or i n the event o f seismic activity.  Higher backfill strengths are required when the material is used to hold rock faces against which active mining is planned. In these cases, the strength o f the backfill material must not be decreased by cement-altering reactions.  In a mine setting, the backfill is  unfortunately exposed to various conditions that can be detrimental to the chemical stability o f cemented tailings backfill such as sulfate attack from sulfate rich A R D generated within the mine or present i n the tailings water used to make up the backfill mixture and water dilution o f the cement phase o f incompletely cured backfill.  Figure  1.1 shows different leaching environments to which backfill can be exposed. A l l stopes are backfilled in this figure, the red colour backfill indicates exposure to A R D generated within the host rock whereas the grey colour indicates exposure to neutral p H , infiltrating water.  The cyclic-leached environment (top) represents cases where backfill  is  temporarily exposed to either infiltrating rain water underground but above the water table or exposed to meteoritic water i n above ground applications. environment  (bottom  left)  represents  backfill  submerged  in  Flooded-leached  circum-neutral p H  groundwater, and the A R D environment (bottom right) represents backfill submerged i n l o w p H A R D water containing high concentrations o f sulfate and metals. potential leaching environments were studied i n this project.  19  These three  F i g u r e 1.1 Paste B a c k f i l l L e a c h i n g E n v i r o n m e n t s  1.2.2.2.1  Sulfate attack:  Sulfate i s o m n i p r e s e n t i n m o s t m e t a l m i n e s , either i n the groundwater i n contact w i t h the deposit o r i n the wastewater generated b y ore p r o c e s s i n g . m i n e r a l p r o c e s s i n g water are often greater  Sulfate concentrations i n  than 1.5 g/1, c l a s s i f i e d as aggressive water i n  the cement i n d u s t r y ( T a b l e 1.2).  T a b l e 1.2 Concrete Subjected to Sulfate A t t a c k D i s s o l v e d sulfate i n contact w i t h concrete (g/1)  P o r t l a n d cement type to be used  >2.0  > 10.0  50  Severe  0.2-2.0  1.5-10.0  50  Moderate  0.1-0.2  0.15 - 1.5  20, 40 or 50  Degree o f exposure  W a t e r s o l u b l e sulfate in soil sample (%)  V e r y severe  Adapted from Kosmatka et al., 1995  Cemented  backfill  paste  i s n o r m a l l y made  f r o m tailings p i p e d d i r e c t l y f r o m the  p r o c e s s i n g plant to w h i c h cement is added i n a secondary m i x i n g step.  20  The hydration  water o f the mixture consists i n large part o f this sulfate-rich residue water. The only method o f reducing the sulfate concentration from the tailings water would be to thoroughly wash the tailings or precipitate out the sulfate as a stable mineral phase, both uneconomical and inefficient (Noranda Technology Centre, 1998a). A second source o f sulfate is acid rock drainage ( A R D ) which may not necessarily be acidic but may contain high (aggressive) concentrations o f sulfate generated by the oxidation o f sulfide minerals. In the first case, where sulfate is found within the tailings water, the chemical stress is generated within the paste.  In the case o f A R D , the chemical stress or aggressivity is  external.  The chemical process o f sulfate attack on hydrated concrete can be summarised as follows:  Free sulfate ions present in solution (from A R D or from sulfate present i n the pore solution o f the backfill) can combine with calcium of dissolved portlandite (Ca(OH)2) to form gypsum according to reactions 1.6:  S0 " + 2  4  Ca  2 +  + 2H 0 ^  CaS0 -2H 0  2  4  (1.6)  2  Gypsum growth i n pore spaces creates pressure that, when occurring at a large scale, can induce cracking o f the backfill. (Ouellet et al.,  Crystallisation pressures can reach 70 to 2000 M P a  1998). Another expansive reaction is the formation o f ettringite  (Ca6Al2(S0 )3(OH)i2"26 H 0 ) from the reaction o f aqueous sulfate and calcium, (from 4  2  dissolved portlandite and free sulfate or from dissolved gypsum) and the monosulfate phase o f cement ( C a 4 A l ( S 0 ) - 1 2 H 2 0 ) according to reaction 1.7: 2  4  21  Ca4Al (S0 )- 12H 0 + 2S0 " + Ca 2  2  The  4  2  4  2 +  + 20 H 0 -> C a 6 A l ( S 0 ) ( O H ) - 2 6 H 0 2  2  4  3  12  2  (1.7)  higher molecular weight and larger crystal lattice o f ettringite compared to  monosulfate phase w i l l also induce expansive forces or crystallisation pressures within the backfill that can result in cracking and disintegration (Fu et al., 1995; Taylor, 1997).  Mineral additives such as ground blast furnace slag, silica fumes or fly ash can be added or partially replace O P C to modify fill mixture properties, such as workability o f the mixture, to increase the strength o f the fill at a particular stage o f curing or resistance to chemical attack (Mangat and Khatib, 1995; Gifford and Gillott, 1997; Taylor, 1997). Mineral additives such as those presented i n Table 1.3 are used to reduce the cost o f the binding agent without decreasing the strength o f the fill. In order to improve the sulfate resistance o f concrete, the amount o f calcium hydroxide and calcium aluminate hydrate must be minimized.  22  Table 1.3: Replacement Components of Portland Cement to Improve Sulfate Resistance of Concrete I", licet  \dditi\c F l y ash: coal combustion residue; l o w calcium ( T o r i i e t a / . 1995) (Djuric eta/. 1996)  Lower air and pore volume, reduced permeability, increased resistance to sulfate absorption into concrete.  Silica fumes: Silicon, silicon alloy smelting ash residue; minimum 75% silicon, very l o w calcium and aluminium oxides ( A k o z e t a / . 1995)  Lower air and pore volume, very reduced permeability, decreases gypsum and ettringite formation, increased electrical resistivity (corrosion protection)  Blast furnace slag: glassy iron smelting residue; calcium silicates and aluminosilicates may be high (13-15%) A 1 0 slag or l o w (3-5% ) AI2O3 slag (Irassar et al. 1996) 2  Lower air and pore volume, reduced permeability, may increase mixture strength in the end.  3  1.2.2.2.2 Dissolution of the Cement Phase of Backfill The dissolution o f cement phases i n contact with meteoritic and groundwater has been studied by the French Commission on Atomic Energy with respect to degradation o f cemented containers o f nuclear waste. Carde and Francois (1997) who used ammonium nitrate leaching solution previously determined to leach cements i n a similar but more rapid way than water, point out that a zone o f lesser strength is formed i n the leached area and that the decreased strength is due to increased porosity resulting from the complete leaching o f the portlandite phase o f cement (Ca(OH) ). 2  The increase i n porosity was  calculated to be equal to the proportion o f portlandite in their concrete mixtures.  23  A  progressive decalcification o f the tobermorite phase was also observed.  Leaching o f  these phases were the only deterioration found to occur i n concrete samples kept immersed i n the leaching solution. In wet-dry cycled experiments, an increase i n pore solution ion concentration was found to promote the precipitation o f secondary expansive minerals such as ettringite causing internal stresses and microcracking o f the concrete. Adenot and B u i l (1992) used deionized water i n similar leaching experiments and observed similar alteration: thin leached zones characterized by partial or complete dissolution  of  portlandite  but  preceded  by  the  dissolution o f  ettringite  and  monosulfoaluminate phases. Calcium to silicon ratios o f the tobermorite phase were also found to decrease from the core o f the specimens towards the surfaces, reflecting a decreasing calcium concentration in the pore solution o f the edges o f the samples. Advancement o f the dissolution front was calculated to be proportional to the diffusion rate o f deionized water i n the concrete. The cores o f their flooded samples were observed to possess a similar composition to the cores o f the unleached samples.  In similar  experiments to those o f Adenot and B u i l , Revertegat and co-workers (1992) determined that dissolution o f portlandite started to occur at p H 12.5 and became more severe as the p H o f the pore solution dropped. Using additional samples made o f cement and fly ash, they documented the increased resistance o f this binder mixture to leaching.  They  determined that this binder mixture uses portlandite in its hydration process, effectively decreasing the amount o f portlandite available to be leached, resulting in a lower loss o f porosity.  24  That research indicates that deionized water is indeed an effective leaching agent o f concrete, dissolving away portlandite thereby increasing the porosity o f the material, as well as decalcifying or degrading tobermorite gel, the principal binding agent o f cement. The considerably higher porosity o f backfill material compared to concrete suggests that leaching solution w i l l more effectively penetrate the backfill and alter the pore solution chemistry, creating disequilibrium conditions between the solution and the solid phase and accelerating the dissolution o f the cement phases o f the backfill.  25  2  MINE SITE DESCRIPTIONS  Figure 2.1 shows the location o f the Canadian mines sites and Figure 2.2, the location o f M e x i c a n mine sites studied in this project. Table 2.1 below identifies the mine sites from which samples were obtained for both the pyrite reactivity and cemented paste chemical stability studies. A summary description o f the geological setting, deposit type and mode o f pyrite occurrence is also presented.  Table 2.1: Summary of Sample Geological Setting and Mineralogy Mine  Location  Deposit type  Pyrite occurrence  Huckleberry  British Columbia, Canada  Skarn deposit  V e i n fill, coarse granular pyritohedrons  1997 - present  Louvicourt  Quebec, Canada  1992 - present  Cu, M o Volcanogenic massive sulfide in felsic tuff, chert and mudstone host rock . Cu, Zn, A g , A u Volcanogenic massive sulfide in felsic volcanic and volcaniclastic host rock Pb, Z n , C u , A u Stratiform volcanogenic massive sulfide i n andesitic metamorphised volcaniclastic host rock Z n , Pb, C u , A u , A g Stratiform volcanogenic sulfides i n metamorphised andesitic volcaniclastic host rock Pb-Zn, C u (Ag) Skarn Pb-Zn deposit 2  Brunswick  N e w Brunswick, Canada  1950'spresent  Tizapa  M e x i c o state, Mexico  1994 - present  Francisco I. Madero  Zacatecas, Mexico  (FIM) Project  Zimapan 1920's(?)present  Hidalgo, Mexico  Pb, Z n , A g , C u , (Au)  (1) Pyrite mode of occurrence in samples used for tests (2) Tourigny et al., 1994; (3) Luff et al, 1992  26  Fine-grained massive pyrite, cubic Some chalcopyrite and sphalerite Fine-grained massive pyrite, cubic  Fine-grained massive pyrite, cubic  Fine-grained massive pyrite, cubic  M e d i u m to large grained cubic pyrite crystals  Mine Site Locations  Figure 2.1 Sites Location of Huckleberry, Louvicourt and Brunswick mines, Canada  27  2.1  COMPARISON OF MINE SITE AND EXPERIMENTAL CONDITIONS  Table 2.2  Describes the site conditions encountered at the different mine sites where  paste backfill w i l l be used.  These conditions can be compared to the experimental  conditions carried out in this study, presented in Chapter 3, Section 3.2.2, Table 3.4.  Table 2.2 Hydrogeologic Conditions at Mine Sites where Backfill is or will be Used Site Louvicourt  Mine Site Water Quality  Backfill Mix Water Quality  p H : 8.3 S 0 : 620 mg/1 F e : 0.05 mg/1 C u : 0.03 mg/1 Z n : 0.10 mg/1  p H : 7.55 S 0 : 700 mg/1  Tizapa  S C V " : 2.4 mg/1  n.a.  Brunswick  p H : 2.9 > S 0 : 6900 mg/1  n.a.  4  tot  ([  4  Fetot: 0.64  mg/1  C u : 0.03 mg/1 Z n : 0.40 mg/1  mg/1  C u : 8.4 mg/1 Pb: 2.7 mg/1 Z n : 1650 mg/1  Francisco I. Madero  Pyritic tailings, Underground use, (since 1992), future surficial use possible. Planned underground use. Underground use. June 1998 startup o f backfill plant.  4  Fetot: 1100  Use of Backfill (or projected use)  Early development stage, no data available.  Planned underground use.  (1) Average water chemistry from 0-mile brook, from January 13, 1989 to July 3, 1997 (Moerman, 1997). n.a.: not available  28  3  EXPERIMENTAL METHODOLOGY  A l l pyrite and paste experiments, as well as parts o f the analyses, were carried out i n Mexico.  The laboratory equipment for both studies was constructed and previously  tested at the University o f British Columbia ( U B C ) .  The laboratory set up was then  dismounted and rebuilt at the Universidad Autonoma de San Luis Potosi ( U A S L P ) , i n M e x i c o . Pyrite, tailings and slag samples were sent directly from the mines to U A S L P where they were prepared and the paste samples were mixed and cured prior to leaching. The 20-week leach cycles for both experiments, as well as most mineralogical analyses, were carried out at U A S L P . carried out at U B C .  Mineralogical analyses o f the leached paste mixtures were  X-ray diffraction analyses were carried out at U A S L P .  Cyclic  voltamperometry was carried out at the Universidad Autonoma Metropolitana - Ixtapala ( U A M I ) i n M e x i c o City. experiments  were  Leachate samples from both the pyrite and the paste  sent for analyses to the  Centro de Investigation y Desarollo  Tecnologico ( C I D T ) o f Penoles in Monterrey, M e x i c o .  3.1  3.1.1  PYRITE EXPERIMENTS  Sample Preparation and Analyses  Pyrite was obtained from composite hand samples chosen by the geologist o f each mine to represent the most common mode o f occurrence at the site. The samples containing coarse pyrite crystals (Huckleberry and Zimapan) were put in a double bag o f 6 m i l polyethylene and fragmented  with a steel head hammer to release pyrite crystals.  29  Individual pyrite crystals containing no visible oxidation coating or inclusions o f other minerals were hand-selected for use i n the tests.  Hand samples o f pyrite from the Louvicourt and Tizapa mines consisted o f massive fine grained pyrite (75-90% pyrite content), making the segregation o f individual crystals by hand practically impossible. The whole samples were therefore prepared directly without the separation o f gangue or minor impurities. Samples were reduced to -6.25 m m using a j a w crusher previously cleaned by repeated passings o f clear glass. N o pure silica sand was available for this purpose.  Pyrite crystals and crushed particles were soaked in an acid bath o f 2 N HC1 for 1 hour to remove oxidation coatings, followed by six repeated rinsing i n deionized water. The pyrite samples were oven dried at 35°C for 24 hours. Dry grinding was performed i n two steps, first using a glass-cleaned ring pulverizer to reduce the grain size and obtain a first batch o f the required particle size and second, with an agate mortar to grind down the coarse material leftover from the ring pulverizer. The ground pyrite was dry sieved using standard 105 and 150 urn Tyler sieves. The +105 -150 um fraction was retained for the leaching tests. Sieves were washed in an ultrasound bath and completely dried between the different samples.  Pyrite samples were kept i n a glass dessicator under vacuum  between the various preparation steps and when not in use to avoid oxidation o f pyrite surfaces.  30  3.1.2  Leaching Apparatus  The leaching apparatus was designed to promote the oxidation o f the different pyrite samples and facilitate the measurement and comparison o f the early effects o f leaching on pyrite grains. Mineralogical observations, as well as chemical and electrochemical analyses, were performed on the unleached pyrite and after 4, 10 and 20 weeks o f leaching.  The leaching apparatus consisted o f 5 cm diameter Biichner Funnels™: a  polyethylene container with a flat perforated bottom and open top, attached at the bottom to a funnel (Figure 3.1). A 45 um filter paper was placed at the bottom o f the container to retain the pyrite particles.  20 gm o f pyrite was placed i n each funnel.  The leaching  solution was devised to simulate rainwater: distilled water with a p H adjusted to 5.5 by the addition o f C O 2 , prepared immediately before use.  15 m l o f leaching solution,  enough to cover the sample, was added to each sample twice weekly. The pyrite samples were left inundated for 3 hours, after which the solution was extracted by vacuum suction, re-filtered and analyzed. The pyrite samples were left exposed to ambient air between each leaching cycle.  Three (3) samples o f each pyrite were prepared and  leached simultaneously. One (1) sample was removed at the end o f each period for the various analyses.  Additional samples o f Huckleberry pyrite were leached with Huckleberry waste rock leach solution. The solution was generated in two leaching columns, each containing 2.5 kg o f crushed low-grade ore (-3.33mm +1.68 mm).  The crushed ore was previously  rinsed with 5 litres o f deionized water to wash out all the fines (finer than 1.68 mm). T o each column was added 250 m l o f deionized water with the p H adjusted to 5.5 with C O 2 . The water was retained i n the columns for 1 hour after which the bottom valve was  31  opened to let the solution freely drain out.  The leachate was then filtered and used to  leach samples o f Huckleberry pyrite prepared in the same way as the  water-leached  pyrite. The remaining solution was acidified and submitted for chemical analyses.  Figure 3.1 Pyrite Leaching Apparatus 3.1.3  Chemical analyses  Routine chemical analyses were performed on leachate solutions obtained by the pyrite experiment and by the Huckleberry waste rock columns for the parameters listed in Tables 3.1 and 3.2. The solution p H , redox potential, conductivity and ferrous iron were analysed upon collection at U A S L P , the other leachate parameters, as well as solids analyses o f the unleached pyrite samples were analyzed at the chemical laboratory o f the Centro de Investigation y Desarollo Tecnologico (CIDT) o f Penoles in Monterrey,  32  M e x i c o . The pyrite leachate solutions were acidified with concentrated HC1 to a p H < 2 and stored at 4°C between monthly shipments to Pefioles v i a courier.  Table 3.1 Leachate Analyses Carried Out at UASLP Analysis  Instrument  p H , redox potential  Beckman (t>320 combined p H and redox meter  Conductivity  Cole Parmer portable conductivity meter M o d e l 19820-00 0.1N KC1 (Calomel) probe, all values adjusted to the standard hydrogen electrode (SHE).  Ferrous iron  U V Spectrophotometer Beckman D U 650, using O-phenanthroline indicator  Table 3.2 Leachate and Solids Analyses Carried Out at CIDT Sample; type Leachate Solutions (both pyrite and paste backfill experiments) Solids (unleached pyrite samples)  Analyses  Methodology and Instrument  F e a i , C u , Z n , Pb, tot  As,  A l , Sb, S, Si, C a , M g , K S0 " z  4  C 0 , S 0 , CaO, K 0 , MgO, MnO, Na 0, Si0 , A I 2 O 3 A g , As, B i , C, Cd, Cu, Cr, Fe, N i , Pb, S, Se, Te, Z n . 3  4  2  2  Atomic absorption spectrometer ( A A ) determination (Perkin-Elmer model 5000) Gravimetric determination Aqua-regia digestion and A A analysis (Perkin-Elmer 5000)  2  3.1.4 Mineralogical and Electrochemical Characterisation of Pyrite A l l pyrite samples were characterized mineralogically and chemically prior to leaching to determine stoichiometric compositions, as well as crystal habit and form. Polished  33  sections o f the hand samples containing the pyrite from each mine were prepared for observation under both optical (reflected light) microscope and scanning electron microscope ( S E M ) . Reflected light microscopy was carried out using a Versamet U n i o n microscope, whereas S E M observations were carried out using a Philips X L 3 0 Scanning Electron Microscope.  Quantitative chemical microanalyses were carried out on the  polished sections using the energy dispersive x-ray ( E D A X ) o f the S E M .  Loose grain  mounts o f the leached pyrites were prepared after each leaching period (4, 10 and 20 weeks).  The grain mounts consisted a dusting o f the pyrite grains over double-sided  carbon tape fixed to a metal holder. observations.  These samples were carbon-coated for S E M  Qualitative chemical microanalyses were carried out using E D A X to  identify the precipitation products formed after each leaching period.  Electrochemical analyses were carried out on unleached pyrites and after 4, 10 and 20 weeks o f leaching to characterise pyrite surfaces and verify the presence o f precipitates imperceptible with the S E M . A t the end o f each leaching cycle, the pyrite samples were oven dried at 35°C and transported to the electrochemistry laboratory o f the U A M I i n air tight polyethylene containers, for voltamperometric studies.  A n E G & G P A R M273  Potentiostat coupled to a P C with the M 2 7 0 software were used to generate a voltage cycle o f 20 mV/sec and record the voltamperometric response respectively.  A 3-  electrode system was used, o f which a carbon paste (graphite-pyrite, 50% wt) electrode (CPE) was used as the working electrode. The carbon and pyrite were carefully blended together with laboratory grade silicon o i l i n an agate mortar (Figure 3.2). The resulting paste was placed i n a 0.5 m l plastic syringe into which was inserted a platinum electrode welded with silver to a copper wire end, in turn connected to the current source.  34  The  counter electrode consisted o f a graphite rod, the reference electrode a H g / H g S G y K ^ S C M (sat),  S S E (Eh=0.615V, S H E ) .  The electrode system was placed i n a Pyrex™ glass cell  containing an electrolyte solution o f 0.1 M N a N 0  3  with a p H o f 6.5, through which was  bubbled purified nitrogen gas for a minimum o f 45 minutes before the start o f the experiments. Nitrogen was blown on top o f the solution throughout the tests to provide an inert atmosphere within the cell. The set-up is shown on Figure 3.3.  Figure 3.2 Preparation of Carbon Paste Electrode (CPE)  35  Figure 3.3 Electrochemical Cell (front clip attached to CPE)  T h e z e r o current p o t e n t i a l (rest potential) o f the C P E w a s established before c o n d u c t i n g the p o t e n t i a l sweep b y letting the electrode system rest f o r a m i n i m u m o f 5 m i n u t e s u n t i l a stable r e a d i n g w a s obtained. anodic  and cathodic  S i n g l e - c y c l e potential sweeps were initiated i n b o t h the  directions (to induce  pyrite  surface  oxidation and reduction  r e s p e c t i v e l y ) w i t h i n the range o f - 1 . 0 to +0.7 V . T h i s range w a s established f r o m the " n o current r e s p o n s e " o f a C P E that d i d not c o n t a i n pyrite, p l a c e d i n the 0.1 M N a N 0 3 electrolyte s o l u t i o n .  A n e w w o r k i n g electrode ( C P E ) surface w a s o b t a i n e d after e a c h  s w e e p b y s q u e e z i n g out a s m a l l a m o u n t o f paste f r o m the syringe a n d l e v e l l i n g the C P E tip o n a 6 0 0 grit s i l i c a carbide sand paper.  M u l t i p l e analyses were c a r r i e d out o f e a c h  sweep range i n order to o b t a i n a repeatable c o m p a r a t i v e study.  36  (representative)  response  usable f o r a  3.2  PASTE EXPERIMENT  3.2.1 Paste Sample Preparation  3.2.1.1 Paste Components Tailings were sent by courier from the mine site to U A S L P in airtight containers with enough residual water to cover the tailings and prevent oxidation during transport. Wet tailings were oven dried at 40°C for 72 hours or until constant weight. F I M tailings were shipped dry and were used without preliminary treatment.  Dry tailings were used i n the paste mixture i n order to standardize the mixing procedure and to have control o f the water content o f each mixture. City tap water was used to make up the paste samples since not enough tailings water was available to make the required volumes o f samples, and some tailings' water chemistry ( F I M samples) were unavailable.  The tailings were not washed prior to drying, therefore the sulfate present  in the tailings water and pore water likely precipitated then redissolved upon hydration o f the paste mixture. Drying the tailings prior to use is not believed to have caused major changes i n the chemistry o f the tailings.  Ordinary Portland Cement ( O P C ) Type 1, the Mexican equivalent o f Canadian Portland Cement N o . 10, was used as binding agent as prescribed in all paste recipes submitted by the mines. The cement was analyzed chemically before forming the paste samples to make sure that it met specifications. The chemical composition o f the cement is shown i n Table 5.5 o f Chapter 5.  37  3.2.1.2 Paste Mixing The paste backfill formulation o f each mine was /followed as closely as possible i n the preparation o f paste samples.  Since the F I M project is still i n the feasibility stage, the  project consultant's suggested formulation (based on general guidelines) was followed for that mine. The formulation to be used at F I M in the future may differ from that made in this study. Table 3.3 shows the proportions o f tailings, water and binder required for each paste recipe (shaded) and actual proportions used to make up the paste for each mine for this investigation (un-shaded).  Standard slump tests ( C S A test method A23.2-5C)  were carried out to determine paste consistency using a standard slump cone. Slump test results are expressed as cm o f slump o f the material once the cone had been removed. A large slump corresponds to a more liquid, less consistent paste.  Initially, a single tailings sample was used to determine the relative proportions o f cement, water and aggregate necessary to achieve the right slump since not enough tailings samples were available to carry out repeated tests. The test indicated that only the site-specific tailings could give the required slump within the given range o f component proportions.  Consequently paste mixtures were made by using a fixed  amount o f binder and tailings from each mine to which was added the minimum amount o f water suggested i n each recipe. The amount o f water was then slowly increased i f necessary until the mixture reached the specified slump range.  In some cases, the  maximum water content specified by the recipe would not yield an adequate slump and the mixture would be too stiff. In these instances, a compromise between water content and acceptable slump range was made i n order to obtain an adequate mixture.  38  Paste mixing and moulding was carried out according to A S T M specifications C192-90a and C470 for concrete mixtures. K e y elements o f these specifications are as follows:  1) M i x i n g o f paste (Figure 3.4 a, b):  •  A plastic, non reactive container was used to m i x the paste components,  •  The batch size o f the backfill mixture was larger than the volume o f the slump cone,  •  A l l constituents were dry and well mixed prior to water addition,  •  The batch was hand mixed for 20 minutes until homogeneous, following which the slump test was carried out,  •  The water content was adjusted i f necessary and the batch was re-mixed for 10-15 minutes following which a second slump measurement was taken,  •  Slump measurements were not repeated more than twice per batch,  •  The paste material used for the slump test was put back i n the mixing container and used to fill the moulds.  39  T3  be  t?; >•  3  SI) 00  u  00  T3  oo  00  00  T3  T3  T3  m  o  E  'do  o  00  33  e r i  o  CN  00  o  CN  -f  a o  00  a SS o  r i CN  U  CN  s-  2  a>  ri  O CN  CN CN  CN  ri  in  cs  O-, T3 O  T3  I* ••"  z.• u  -4->  ON  s  CN CN  in  o  o +o-»  MS r«".  U  n  o  cj  <5  c o V* ft ses  T3  #  a  i-  PH a> "a  S «  oo  s  s. IT)  5 "53  o 00 00  C/3  oo  oc  00 ^iiiiiii 'I, ^ fN 00  C3  «  ss o o  t--  <+H O  O fl  H  a 00  H) "S ^ 5 J ^<U . & St  - !  g  -P <D O  00  o 3  oo f-  (N  3  jS p<^ <  O  a  « .9  o  >>  T3  to  a ss M  o  op 131 •S £ o  •S t;  so  PM  -y. O  «  m  2) M o u l d material and mould filling (Figure 3.5):  •  N o n reactive, non-absorptive, watertight, single use cylindrical polyethylene moulds were used having an internal diameter equal to Vi the vertical height (5 c m diameter x 10 c m height),  •  The filling sequence was dictated by the size and shape o f the mould: •  filling was carried out in 2 layers o f 5 c m each,  •  each layer was given 25 strokes with a rounded end glass rod to evacuate air bubbles,  •  a slight tap on bottom o f mould was given after rodding.  3) Curing o f samples (Figure 3.6): •  Samples were covered with a plastic sheet to prevent rapid drying o f the surface. Wet sponges were kept under the plastic sheet as sources o f humidity,  •  Curing was carried out in a temperature and humidity monitored environment with minimised air currents. The average curing temperature was 20°C ± 2 °C and average humidity was 80% ± 1 0 % .  Concrete  samples,  for  which  the  above  specifications were  designed,  possess  considerably more binder or cement (commonly around 30 % binder) than paste material made by the mining industry (commonly less than 8% binder). A s a consequence, a deviation to specification C I 9 2 had to be made where the paste samples were not demoulded after 24 hours because o f the lack o f strength of some samples. A l l samples were de-moulded after a humid cure period o f 14 days.  41  Figure 3.4: a. Mixing Paste Sample  Figure 3.5 Rodding Paste into Mould  b. Standard Slump Test  Figure 3.6 Curing of Paste Samples  3.2.1.3 Pre-Leaching Sample Preparation Cured samples were wet-cut into two pieces using a thoroughly cleaned and degreased circular saw (diamond saw).  The pieces consisted o f a cylinder o f 10 c m length for  compressive strength test and a separate piece o f 2 to 3 cm thick (termed a puck) polished on one side intended for mineralogical observations. The sample surfaces in contact with the mould were lightly sanded to remove the thin oxidation coating that had started to  42  form upon curing for some samples. Both pieces o f each sample were oven dried at 40°C for 48 hours, weighed and measured before placement in the leaching cells.  3.2.2  Leaching Apparatus and Leaching Cycles  The leach solutions and leach cycles were devised to simulate the different modes o f exposure o f a cemented paste backfill, as shown in Figure 1.1 o f Chapter 1. Table 3.4 explains how these environments were simulated in the laboratory to recreate the various leaching scenarios.  Table 3.4 Paste Leaching Scenarios Mine Setting Sub-arial backfill or in underground mine above the water table  Mode of Leaching Alternating dry (ambient air) and flooded cycles  Leaching Solution Simulated rain water: distilled water, p H adjusted to 5.5 with C 0 2  Flooded backfill (below the water table), absence o f A R D i n mine water  Constantly flooded cycles, 2 week water circulation period  Flooded backfill i n contact with acidic groundwater produced i n the mine  Constantly flooded cycles, bi-weekly solution replacement  Simulated rain water: distilled water, p H adjusted to 5.5 with C 0 2  Fe (S04)3 solution, 0.005 M or (1.5g/lS0 -) p H 2.4 - 2.6 2  2  4  Leaching Frequency 1 week cycle: 24 hours flooded then drained and let open to ambient air for 6 days 1 week cycle: Vi cell volume taken out for analyses and replaced with fresh solution Vi week cycle: replacement o f entire cell solution  The leaching cells consisted o f 8.5 cm diameter, 15 c m tall, clear acrylic cylinders equipped with an evacuation valve at the bottom and removable cap and valve at the top (Figure 3.7).  Some additional cells were made at U A S L P having slightly different  dimensions (5 c m diameter, 20 c m tall) for the lack o f similar size material but with an  43  equal volume as the cells made at U B C . cemented paste fill leaching study.  Figure 3.8 shows the laboratory set-up for the  Three cells were prepared for each sample and  leaching environment so that one sample could be removed after each designated leaching period for destructive analyses. Three (3) samples were prepared and leached simultaneously i n a ferric sulfate solution, 3 in a flooded water environment and 3 in the alternating air-water environment for each mine. After every leaching period (5, 10 and 20 weeks) 1 leaching cell was dismantled for analysis o f the solid phase. In each o f the first five weeks, leachate solutions o f the three leach cells characterizing the same environment were combined to form one large sample for analysis. Between the 5 10  th  th  and  week, leachate samples were a composite o f the two remaining cells o f the same  environment and leachate samples between the 10  th  and 2 0  th  week originated from a  single leaching cell.  Figure 3.7 Leaching Cells 3.2.3  Figure 3.8 Lab Set up, Paste Leaching Study  Post-Leaching Sample Preparation  After each leaching period, the retrieved samples - both cylinder and puck - were measured, weighed and oven dried at 50°C for approximately 72 hours or to a constant weight. The difference between the wet and dry weights was noted for later analysis o f  44  water absorption by the paste sample in the different leaching environments. The dry samples were placed into individual bags o f 2-mil polyethylene, properly closed to exclude contact with external moisture.  The cylinders were brought to the Soil  Mechanics laboratory o f the Instituto Politecnico National (IPN) o f M e x i c o City for unconfined compressive strength (ucs) analyses using a Losenhausenwerk hydraulic compressor.  The leached paste pucks were very friable and fragile.  Consequently, a sample  preparation procedure was devised to minimize handling o f the sample and maximize the subsequent quality o f S E M imaging. The pucks were cut transversally with a fine tooth metal saw and sanded down using 150, 220 then 1500 grit silica sand paper to obtain a flat surface.  In the case where pucks were too thin or inappropriate for mineralogical  observations, slices o f the cylinders were cut out and prepared i n the same way. Excess dust and loose particles were lifted from the surface o f the cut surface with adhesive tape. The sample was then fixed to aluminium foil with double-sided carbon tape exposing only the flattened, particle-free surface (Figure 3.9 a, b, c). A l l prepared puck surfaces were photo scanned then carbon coated for S E M observations. Some samples required 2 or 3 coatings o f carbon because o f their porous nature. Samples were kept i n a dessicator between preparation and observation steps.  45  Figures 3.9 a, b, c: Paste Sample Preparation Steps for Mineralogical Observations  T h e different o x i d a t i o n zones o f the leached samples were i d e n t i f i e d b y c o l o u r and/or textural changes v i s i b l e o n the p u c k sections and/or c y l i n d e r s .  A l l the v i s i b l y different  layers, as w e l l as the core o f each sample, were carefully scraped o f f the 2 0 w e e k - l e a c h e d samples ( c y l i n d e r s ) a n d c o l l e c t e d for analyses. T h e 5 a n d 10 w e e k - l e a c h e d samples w e r e not u s e d for this purpose, as the o x i d a t i o n layers were  i n general too t h i n to  be  s u c c e s s f u l l y separated. I n cases where no distinct o x i d a t i o n layers were apparent, a 1 m m t h i c k s u r f i c i a l l a y e r w a s scraped o f f for analyses.  M i n e r a l o g i c a l observations were carried out o n the prepared p u c k s u s i n g a P h i l i p s X L 3 0 scanning electron microscope ( S E M ) .  C h e m i c a l m i c r o a n a l y s e s were p e r f o r m e d u s i n g a n  I M I X energy d i s p e r s i v e x - r a y ( E D X ) instrument at U B C .  E l e m e n t a l spectra o f the  cement phases a n d precipitates were taken, thereby p r o v i d i n g qualitative i n f o r m a t i o n about the general c o m p o s i t i o n o f these phases.  N o quantitative i n f o r m a t i o n c o u l d be  extracted f r o m the E D X spectra as the phases o f interest were too s m a l l a n d scarce to p r o v i d e accurate quantitative data w i t h the a n a l y t i c a l procedures used.  46  3.2.4  Chemical Analyses of the Paste Experiment  Routine chemical analyses were performed on leaching solutions (water and simulated A R D or ferric sulfate solutions) and on the leachate produced by the paste backfill leaching cells. Parameters analyzed for the paste backfill study are listed i n Tables 3.1 and 3.2 i n section 3.1.3. Leachate p H , redox potential, conductivity and ferrous iron were analysed upon collection at U A S L P .  Other leachate parameters and the hole-rock  analyses o f tailings, cement and slag were carried out at the analytical laboratory o f the Centro de Investigation y Desarollo Tecnologico (CIDT) o f Pefioles in Monterrey, M e x i c o . The leachate solutions were acidified with concentrated HC1 to a p H <2 and stored at 4°C between monthly shipments to Pefioles v i a courier. Solid phase analyses o f the weathered backfill samples were carried out according to the methods listed i n Table 3.5. The ferric sulfate-leached Brunswick samples were sent to Chem-Met laboratory o f Vancouver for sulfur species analyses.  This lab was not equipped to analyse the  requested suite o f metals and major ions, hence the unused portions o f these samples were sent to Chemex laboratory o f North Vancouver for analyses o f metals and major ions listed i n Table 3.5.  A l l the other weathered paste samples were sent directly to  Chemex laboratory for sulfur species determination as well as metals and major ions analyses (when sufficient sample mass was available for analyses).  47  Table 3.5 Solids Phase Analyses for the Paste Experiments Sample  Analvses  Unleached solids (tailings, O P C and slag)  FcO. C O i . S 0 , CaO, K 0 , MgO, MnO, Na 0, Si0 , A1 0 Ag, As, Bi, C, Cd, Cu, Cr, Fe, N i , Pb, S, Se, Te, Z n . Total Sulfur  Pefioles C I D T : Aqua-regia digestion A A analysis (Perkin-Elmer 5000)  Sulfate  Chem M e t and Chemex: Dilute (10%) HC1 leach and gravimetric determination  A l , C a , M g , Fe, C u , Z n  Chemex: H F acid digestion and I C P A E S analysis (Perkin-Elmer Optima)  Leached paste  Vleihodoloi'.v .md Instrument 4  2  2  2  2  3  Chem Met: A c i d leach (oxidation) and gravimetric determination Chemex: Leco 420 combustion furnace with Infra R e d detector  1  Chemex: H F acid digestion and A A analysis (Varian 220)  Pb  ICP-AES: Inductively coupled plasma atomic emission spectrocopy  3.2.4.1 Paste pH and Acid-Base Accounting Measurements Paste p H measurements were taken on dry tailings and unleached paste backfill for the purpose o f comparison and to verify the presence o f acidity or available buffering minerals o f the samples. Acid-base accounting ( A B A ) tests were carried out at U A S L P to monitor the evolution o f the paste buffering capacity as leaching proceeded.  The  Modified Sobek Method o f A B A analysis was followed on all tailings and unleached paste samples as well as on the different oxidation layers o f all 20-week leached paste samples.  48  4  RESULTS - PYRITE REACTIVITY STUDY  4.1  MINERALOGY, CHEMISTRY AND STOICHIOMETRY OF UNLEACHED PYRITES  A l l pyrite samples studied possessed different chemical compositions, stoichiometries and mineralogical associations.  Solid phase chemistry o f the pyrite samples, normative  mineralogy o f the samples and average stoichiometry o f pyrite crystals i n the samples are summarized i n Tables 4.1, 4.2 and 4.3 respectively. S E M micrographs showing sample mineralogy are shown i n Plates 4.1 to 4.6.  Mineralogical observations o f unleached polished sections showed that the Brunswick and Zimapan samples  possessed  the  greatest amount  o f sphalerite,  galena  and  arsenopyrite impurities, mainly occurring as small inclusions within pyrite (Plates 4.1 and 4.2). The Brunswick pyrite sample consisted o f pyrite concentrate whose particles were ground small enough to expose the separate mineral phases. Louvicourt-2 and Tizapa pyrites contained sphalerite, chalcopyrite, galena as well as some tetrahedrite and arsenopyrite impurities occurring mainly as separate phases in the interstices o f pyrite grains (Plates 4.3 and 4.4). Some o f these mineral impurities also occurred as inclusions 1  within pyrite crystals. Huckleberry and Louvicourt-1 samples contained the least amount of mineral impurities.  1  Huckleberry pyrite contained no visible inclusions whereas  Mineral impurities refers to the presence of non-pyrite sulfide minerals in the pyrite samples. 49  Louvicourt-1 contained very few and dispersed  inclusions o f sphalerite,  galena,  chalcopyrite and arsenopyrite (Plates 4.5 and 4.6).  Stoichiometric microanalyses o f uncrushed pyrite crystals (polished sections) indicated that pyrites i n all samples possessed some nickel (Ni), copper (Cu) and arsenic (As) i n their crystal lattice. The amount o f lattice impurities varied considerably within a single grain and also between the various grains o f one sample, as reflected by the high standard deviations o f the measurements (Table 4.3). Cross-sectional microanalyses o f individual pyrite grains did not reveal any texture differences associated with chemical variations i n any o f the pyrites.  50  "S3 .s  g(0 Br  c  5 5 *•  1i si 1 £  c >  CD 3 O)  CD  o  £  CO  O CD  CL  co  Q.  O  c  CO  T—  V CD — C CD CO w  o '  CD _ C CO  «= "O o ©  -s OO  n  >  N  —  CO  w  CD  O  w  "O <D  —  CO  w  !  Q_ CO  E _  Q. .o e£ CD CO  CM  CD _ C CO  3  s  T~  I  o <  CD 3  O  TO O)  CL CO  ~ s CO  V  T  co"  Q_  •3 •o  3  8  Oc w  E  O  f  5? .c  5  <o  (0  to  co  co  o  , _  CM  VP  c  CO TO  CL  O  (0  CL  £ o  CO  CO  CO  CO  Q. O CL <  >; c£ a. co O)  >>  3 £  a.  < o co  CD  j  CO  in  a> c CO TO  00 I  CO CD  CD T3 JZ CD 3 JZ CD o CD —  I  a.  <D  "  o  Q. Q. O  CD 3 TO  CD 3  3  .Q (0 K-  CO V  <  (D  N  I  >»  g" ^  c 1 oo  1  vP  O)  O  £ iS  s  ^ 0 0  QL  CO  _  91 O  TO cCO  O co"  CO  «= -o  'llill (0  c  CD _ C CO  o _l  Q.  CD  CD 3  _  CO  an  ISpNllit  0.  Py  CM  o  2  cu g  E  51  0  3 TO C CO TO  Chemical Analysis of Pyrite Samples  Table 4.2 Species  unit Huckleberry  Zimapan  Tizapa  Louv 1  Louv 2  Brunswick »iRy,.ci3nceotrate)j  wt% S S as S0 wt%  47.4 0.79  47.3 0.41  50.4 0.20  40.4 0.40  47.4 0.20  58.7 1.89  Fe Zn  wt% wt% wt%  46.7 45.2 0.01  46.9 42.5 0.70  50.2 45.4 0.53  40.0 39.4 0.004  56.8 38.2 2.60  Pb Cu Ni As Bi Si0 C0 Al 0 Na 0  wt% wt% wt% wt% wt% wt% wt% wt% wt%  0.01 0.02 <0.005 <0.02 0.024 0.50 0.14 0.09 0.04  1.39 0.02 <0.005 1.18 0.015 2.56 0.04 0.17  0.12 0.80 <0.005 0.19 0.031 0.30 0.01 0.11 0.07  0.04 0.04 <0.005 0.05 0.018 11.2 0.43 3.47 0.07  47.2 37.7 0.98 0.04 0.02 <0.005 0.06 0.016 8.70 0.10 6.50 0.07  8.00 0.32 0.006 0.28 0.030 0.95 0.74 2.12 0.01  CaO MgO MnO K20 C  wt%  0.45  0.45  wt% wt% wt%  0.05 0.13 0.02 0.06  0.50 0.01 0.10 0.02 0.26  0.05 0.15 0.02 0.09  0.55 0.71 0.13 0.02 0.09  0.57 0.31 0.08 0.02 0.08  0.20 0.11 0.02 0.01 0.9  <2 94.13  46 96.80  80 98.64  10 96.60  8 102.63  178 113.197  4  s2  2  3  2  3  2  Ag Sum Note:  wt% g/Tor  0.07  For all samples: Cd <0.005%, Cr <0.02%, Se<0.001 %, Te<0.002% Loss on ignition (LOI) not reported  52  c  m  o  i  d  in d  CO d to  cn  d  <£>  d  o T—  o  m o  o d  o  o d  to o d  CO d  CM  5  o _l  CM CO lO  CO CO  CD  CO  CO  o  d  T—  d  d  CM  lO iri  00 o d  d  T—  5 d  (0 Q_ (0 N  s  CO in  d  d  o d  d  o  o  CO d  CO CO rr  LO o d  m d  m d  LO  o d  d  T—  CO o d  d  CO  CO  o d  in d  CM CO d  CO d  T—  d  c  (0 c l  CO  £ N  in o d  CD IT) d  CO o d  CO CO d  CD  ja  CD  u 3  C CD  E  CO CO LO  CO  o  LU  53  SEM Micrographs of Unleached Pyrites from Polished Sections ( B a c k  Scatter v i e w )  Plate 4.1 Zimapan Pyrite  Plate 4.2 Brunswick Pyrite  Plate 4.5 Huckleberry Pyrite  Plate 4.6 Louvicourt-1 Pyrite  Legend: P y = p y r i t e ; S p = sphalerite; G a = galena; C p = c h a l c o p y r i t e ; A p = arsenopyrite; T e = tetrahydrite; G n = gangue  54  4.2  ELECTROCHEMISTRY OF THE UNLEACHED PYRITE SAMPLES  Stable open current potentials o f each pyrite sample were measured before beginning each voltamperometric test. Ocp values are presented i n Table 4.4. A correlation was found to exist between ocp values and their content o f mineral impurities. Unleached Huckleberry and Louvicourt-1 samples exhibited similar, relatively l o w ocp values, suggestive o f a relatively reactive sample compared to Louvicourt-2, Tizapa and Zimapan.  The latter 3 pyrite samples possessed more mineral impurities and were  characterized by higher ocp values, indicative o f a lower mineral reactivity. The l o w ocp of the Brunswick sample may be imparted by the high proportions o f sphalerite and galena i n that sample, which may have overwritten the ocp o f the pyrite i n the initial leaching cycles.  The evolution o f Brunswick sample reactivity with leaching time  (Section 4.5.5) supports this possibility.  Table 4.4 Open Current Potentials of Pyrite Samples (Volts, SSE) I'nlcachcd  4 weeks  10 weeks  20 weeks  -0.27  -0.22  -0.24  -0.22  -0.27  -0.22  -0.22  -0.20  Louvicourt-1  -0.27  -0.21  -0.20  -0.22  Louvicourt-2  -0.19  -0.21  -0.13  -0.22  Tizapa  -0.18  -0.20  -0.15  -0.18  Zimapan  -0.18  -0.18  -0.16  -0.20  Brunswick  -0.29  -0.26  -0.18  -0.13  Sample Huckleberry (water-leached) Huckleberry (column leached)  55  Figures 4.1 and 4.2 compare the voltamperometric response o f pyrites i n the unleached samples.  The arrows on Figure 4.1 indicate scan directions.  For the sake o f graph  clarity, scan direction arrows have not been added to the other voltamograms.  Each  sample had a distinct electrochemical response, indicative o f the reactivity differences between them.  Reactivity was measured by the potential at which oxidation o f the  sample reached a current o f 2.0 u A .  A l l anodic peaks showed a catalytic behaviour, where for a given potential, a lower current was generated during the forward scan (anodic current) than i n the reverse scan (cathodic current).  Although the cathodic response o f the leached mineral can reveal  interesting information on the properties o f precipitates covering mineral surfaces, this study focused on the anodic or oxidative behaviour o f pyrite to document the reactivity o f the pyrite surface after increasing periods o f oxidative leaching.  Figures 4.1 and 4.2 indicate that Huckleberry pyrite oxidized at the lowest potential and offered the least resistance as potential was increased. It was concluded to be the most reactive o f all unleached samples.  Following, in order o f decreasing reactivity, were  Louvicourt-1, Louvicourt-2 and Tizapa. The close up view o f the points o f initial current release (Figure 4.2) shows, for Brunswick, and slightly less pronounced for Zimapan, an initial current release prior to the typical curve o f pyrite oxidation. A n electrochemical study o f the oxidation o f galena and sphalerite in the same voltametric cell environment (Figures 4.3 and 4.4 respectively) suggested that the lower potential at which the oxidation o f the unleached sample appeared to be initiated was more likely attributable to the oxidation o f galena. The oxidation o f pyrite appeared to resume around 0.420 volts in  56  both the Brunswick and Zimapan curves as indicated by the sudden increase i n current. The relative reactivities o f the unleached Zimapan and Brunswick pyrites are not possible to establish because o f the presence o f galena.  Once pyrite oxidation was initialized however, Figure 4.1 shows that Brunswick pyrite was the most resistant to oxidation. Humidity cell studies o f fresh Brunswick tailings and mineralogical studies o f old tailings carried out by Noranda Technology Centre (1998f) found Brunswick pyrite to have a relatively low reactivity and to oxidize at a lower rate than what is commonly reported for pyrite.  The varying amount o f mineral impurities and distinct chemistry o f the pyrite samples, as well as the variable stoichiometric compositions o f the pyrites within one sample did not allow for a comparison o f pure FeS2 compounds.  The reactivity o f the pyrite i n each  sample could not be directly measured because o f these intrinsic differences impossible to eliminate from the sample. It was possible, however, through cyclic voltamperometry, to document the effects o f impurities on the reactivity o f pyrite as leaching progressed.  57  o o oo o o  <=> E u  _  13 M D. VI  a  o o oo  a ©  fi  73 O CO  a> «« fl  o  IX! O  o  CO  c  CU  13 j= oCZ5 TT cu sfl  o o  CM  o  CM  (vn) juaxino ON in  o a  o fl W Ph  a U o la <u a.  a •**  >  fl  13 O V S-  s  (yn) juejjno  4.3  EVOLUTION OF LEACHATE CHEMISTRY  The chemical analyses o f the leachate for all samples are presented i n Appendix I (Table 1-1). The evolution o f leachate composition is plotted as time-series graphs for each o f the following elements:  p H , conductivity, redox potential, sulfate, total iron, lead, zinc,  copper and arsenic (Figures 4.5 to 4.12). The artificial rain water solution (referred to as "Distilled Water") as well as the Huckleberry column solution used to leach a separate Huckleberry pyrite sample (referred to Hk-column) are also plotted as references.  The  solution-leached and water-leached Huckleberry samples are identified as Hk-1 and H k - w respectively.  After a rapid decline i n the first leaching cycles, p H stabilized for all pyrites (Figure 4.3). Redox values (Figure 4.4) remained in the 400-500 mvolt range ( S H E ) throughout the leaching cycles indicating that oxidizing conditions prevailed, as expected.  Conductivity, sulfate and dissolved iron were markedly higher for both Huckleberry pyrites (water and solution-leached) compared to the other pyrites (Figures 4.5, 4.6 and 4.7) indicating that Huckleberry pyrites were being oxidized at a considerably greater rate than the other pyrites throughout the leaching experiment.  Dissolved metal concentrations  varied considerably from one sample to another.  Leachate zinc concentrations (Figure 4.8) were directly proportional to the solid phase zinc content o f the unleached sample, occurring as sphalerite impurities. Samples with the highest sphalerite content such as Louvicourt-2, Tizapa, Zimapan and especially Brunswick all leached relatively high concentrations o f dissolved zinc. 60  Tizapa and  Zimapan leachates showed a decline in dissolved zinc concentration after 12 cycles (6 weeks) o f leaching whereas Brunswick and Louvicourt-2 zinc concentrations decreased much more slowly.  This indicates that sphalerite i n the Brunswick and Louvicourt-2  samples was largely exposed to the leaching solution whereas Tizapa and Zimapan sphalerite was partially locked inside pyrite grains.  Dissolved lead concentrations were also proportional to the initial solid phase lead (galena) content (Figure 4.9). The Brunswick pyrite leached out highest concentrations o f lead but only in the first 12 cycles o f leaching, after which aqueous lead concentrations reached steady state at approximately 10 mg/1, similar to Zimapan and Tizapa levels. S E M observations indicated that lead tended to precipitate as anglesite ( P b S 0 ) . 4  Copper and arsenic concentrations were relatively l o w in all leachates (Figures 4.10 and 4.11) reflecting the l o w chalcopyrite and arsenopyrite content o f the samples.  Tizapa's  aqueous copper concentration corresponds to its slightly higher chalcopyrite content observed i n the polished section o f the unleached sample.  10  +  . .'  *•* + * * +  +  +  +  +  ,* +  +  2  0 1  3  5  7  9  1 1  13  15  17  19  21  23  25  n o . of c y c l e s  Figure 4.5 Pyrite Leachate pH  61  27  29  31  33  35  37  39  —D—Hk-I •  % Pb:  Hk-w (0.007%)  — • — L o u 1 (0.04%) — • — L o u 2 (0.035%) -Q E  40  •  Tlz (0.12%)  — • — Z i m (1.39%) — 1 — B w k (8.0%) 1  3  5  7  9  11 13 15 17 19 21 23 25 27 29 31 33 35 37 39  no. of cycles Figure 4.11 Pyrite Leachate Lead Concentrations  63  30  25  £T 20 S  1 5  3  O 1C ram* 1  3  5  7  9  11 13 15 17 19 21 23 25 27 29 31 33 35 37 39  no. of cycles  Figure 4.12 Pyrite Leachate Copper Concentrations  30 25  1  3  5  7  9  11 13 15 17 19 21 23 25 27 29 31 33 35 37 39  no. of cycles  Figure 4.13 Pyrite Leachate Arsenic Concentration  64  4.4  EVOLUTION OF PRECIPITATE COATINGS DURING LEACHING  4.4.1 Huckleberry pyrite S E M observations showed differences i n precipitation products between the water and waste rock solution-leached Huckleberry samples as early as the 4  th  week o f leaching.  The water-leached pyrites conserved relatively sharp grain edges with very few visible coatings (Plate 4.7) or corrosion pits. The solution-leached pyrite surfaces, on the other hand, showed the presence o f amorphous precipitates after 4 weeks o f leaching, increasing in thickness and extent with advancing leaching cycles (Plate 4.8 and 4.9). Oxidation pits were also more extensive in the solution-leached pyrites (Plate 4.10).  Although precipitates were visible under S E M , infrared spectroscopy did not succeed i n identifying the secondary phases, as their amount was insufficient (less than 5%). Consequently, no data are available as to the exact composition o f these precipitates. A review o f the literature on surface oxidation products formed on pyrites suggests, however,  that  the  precipitates  observed  under  S E M may  be  amorphous  iron  oxyhydroxides similar to goethite (a-FeO(OH)) (Jambor, 1994; Bigham, 1994; K w o n g , 1993). Amorphous, hydrated iron sulfates can also be present on pyrite surfaces but are less common than iron oxyhydroxides in young mine wastes according to these authors. The thin, amorphous iron precipitates observed in this study w i l l therefore loosely be referred to as iron oxyhydroxides or F e O O H .  65  4.4.2  Louvicourt Pyrites  Mineralogically similar Huckleberry and Louvicourt-1 pyrite samples yielded similar precipitate coatings upon leaching. Plates 4.11, 4.12 and 4.13 show a precipitate cover similar i n form and abundance to Huckleberry solution-leached pyrites.  Very few  corrosion pits were visible after 20 weeks o f leaching.  The Louvicourt-2 pyrites were considerably different from Louvicourt-1 i n both the occurrence o f precipitates and the aspect o f oxidized pyrites. Pyrite surfaces remained relatively free o f precipitates until the 2 0  th  week, at which point visible amorphous iron  precipitates increased i n abundance. After 20 weeks o f leaching, however, the amount o f iron precipitates covering pyrite surfaces remained much lower than the Louvicourt-1 pyrites (Plates 4.14 and 4.15). Corrosion pits were abundant after 20 weeks o f leaching, contrary to Louvicourt-1 or Huckleberry pyrites (Plate 4.16). Sphalerite was observed i n Louvicourt-2 sample and, where both sphalerite and pyrite were present i n the same grain, sphalerite was extensively more riled and pitted than pyrite after 10 and 20 weeks o f leaching (Plate 4.17). Sphalerite was being preferentially oxidized over pyrite.  4.4.3  Tizapa Pyrite  Mineralogically similar Tizapa and Louvicourt-2 pyrites also showed similarities i n precipitate occurrence and aspect o f oxidized mineral phases.  The surface o f pyrite  grains remained relatively free o f precipitates throughout the 20 weeks o f leaching and, i n general, very few corrosion pits were observed (Plate 4.18).  Sphalerite was also  considerably more corroded than pyrite (Plate 4.19). In addition, sphalerite became more 66  extensively covered with precipitates than pyrite after 20 weeks o f leaching (Plate 4.20).  Galena i n the unleached Tizapa sample appeared as mineral impurities within the pyrite grains and exposed to the environment.  The exposed galena was readily available to  oxidize or dissolve. After 10 and 20 weeks o f leaching, no galena was detected but anglesite (PbSC^) was observed principally on altered galena, suggesting a direct oxidation (replacement) o f galena (Plate 4.21).  The precipitation o f lead sulfate as  anglesite explains the absence o f lead in the leachate.  4.4.4 Zimapan Pyrite The Zimapan pyrites showed a similar abundance o f amorphous iron precipitates to Louvicourt-1 pyrites.  After 20 weeks o f leaching, these precipitates  considerable portion o f most pyrite grains (Plate 4.22). observed as early as the 4  th  covered a  Oxidation pits were also  week o f leaching, attesting to the highly reactive character o f  Zimapan pyrite (Plate 4.23). Reactivity may be accentuated, in part, by the occurrence o f crystal lattice impurities o f A s known to act as an electron donor when present i n pyrite, thereby promoting pyrite oxidation.  N o galena was observed after the 4  th  week o f leaching, replaced by abundant euhedral  crystals o f anglesite occurring throughout the sample (Plate 4.24).  The distribution o f  anglesite suggests that nucleation formed from the leachate solution rather than by direct replacement o f galena phases, implying a previous dissolution or oxidation o f galena. The abundance o f anglesite appeared to remain constant with advancing leaching cycles.  67  4.4.5  Brunswick Pyrite  Relatively few iron hydroxide precipitates covered the surfaces o f pyrite grains after 20 weeks o f leaching (Plate 4.25). F e w pyrite corrosion pits were observed in the 4  th  week  o f leaching but their occurrence increased with advancing leaching cycles (Plate 4.26).  th Precipitates o f anglesite were observed in the 4  week o f leaching throughout the sample  along with galena (Plate 4.27). The occurrence o f anglesite suggests that, similar to the Zimapan sample, it was precipitated from solution, from previously oxidized galena. Galena was present i n the sample after 10 weeks, but absent after 20 weeks o f leaching. Sphalerite was not observed i n any o f the leached samples.  68  Huckleberry Pyrites  Plate 4.9 Solution-Leached, 20 weeks  Plate 4.10 Solution-Leached, 20 weeks  Plates 4.7 and 4.8 showing water-leached pyrites possess a smaller amount o f amorphous iron hydroxide precipitates o f than the solution-leached pyrites shown on Plates 4.9 and 4.10.  69  Louvicourt-1 Pyrite  Plate 4.12 20-week leached  Plate 4.11 4-week leached  Plate 4.13 20-week leached  Surface-covering  iron hydroxide precipitates  extensive with advancing leaching cycles.  70  on Louvicourt-1 pyrites  grew  Louvicourt-2 Pyrites  Plate 4.14 Pyrite surface at 10 weeks  Plate 4.15 Pyrite surface at 20 weeks  Plate 4.16 Pyrite at 10 weeks  Plate 4.17 20-week leached  Little difference was observed between pyrite surface precipitates after 10 and 20 weeks of leaching (Plates 4.14 and 4.15 respectively). Plate 4.16 shows corrosion pits on pyrite grain free o f visible mineral impurities. Plate 4.17 shows preferential sphalerite oxidation over pyrite.  71  Tizapa Pyrite  Plate 4.18 Pyrite surface at 20 weeks  Plate 4.19 Oxidized sphalerite, 20 weeks  Plate 4.20 Precipitate cover, 20 weeks  Plate 4.21 Oxidized galena, 20 weeks  After 20 weeks o f leaching, pyrite surfaces show very little oxidation and surface precipitates o f iron hydroxide (Plate 4.18). Plate 4.19 shows the preferential oxidation o f sphalerite over pyrite.  Plate 4.20 shows more extensive cover o f precipitates over  sphalerite and tetrahedrite than pyrite. Plate 4.21 shows the direct oxidation o f galena impurity characterized by surface precipitates o f anglesite (Ang).  72  Zimapan Pyrite  Plate 4.22 Pyrite surface at 20 weeks  Plate 4.23 Corrosion pits at 4 weeks  Plate 4.24 Anglesite precipitates, 20 weeks  Plate 4.22 shows the surface o f pyrite with extensive cover o f amorphous iron hydroxide precipitates.  Corrosion pits were abundant on pyrite grains after 4 weeks o f leaching  (Plate 4.23). Plate 4.24 shows anglesite (Ang) precipitates on the surface o f pyrite.  73  Brunswick Pyrite  Plate 4.27 Anglesite and galena, 4 weeks  Relatively small amounts o f amorphous iron hydroxide precipitates on pyrite surfaces after 20 weeks o f leaching (Plate 4.25). before 20 weeks o f leaching (Plate 4.26).  Corrosion pits were not commonly observed Anglesite (Ang) and galena (Ga) occurred  together on pyrite surfaces after 4 weeks o f leaching (Plate 4.27). observed i n subsequent cycles.  74  N o galena was  4.5  EVOLUTION OF PYRITE REACTIVITY  The evolution o f the reactivity o f pyrite with leaching time is demonstrated i n cyclic voltammograms  of  each  sample.  In  all  graphs,  each  curve  represents  the  voltamperometric response o f a particular pyrite sample at a given period (unleached and after 4, 10 and 20 weeks o f leaching). The navy blue curve represents the response o f the unleached sample and the purple, green and red curves represent the responses o f the same pyrite after 4, 10 and 20 weeks o f leaching respectively.  4.5.1 Huckleberry pyrites Figure 4.14 presents the voltamperometric responses o f Huckleberry pyrite leached i n water. The reactivity o f this pyrite decreased over the first 10 weeks, following which a slight gain i n reactivity was observed.  A net loss o f reactivity occurred over the 20  weeks o f leaching, characterized by a total positive displacement o f the initialization point o f 0.03 volts (measured at 2.0 u A ) . The solution-leached pyrite also showed a decreasing reactivity with leaching time, including a similar gain o f reactivity at the 20week mark (Figure 4.15). A net gain o f 0.07 volts occurred over the 20-week leaching period with respect to the unleached pyrite.  The  mineralogical observations,  leachate  chemistry  results  and  electrochemical  characterizations suggest that the Huckleberry pyrite is considerably reactive to oxidation  75  but that a passivation layer is likely formed on pyrite surfaces, effectively decreasing the 2  reactivity o f pyrite after only 4 weeks o f leaching. The passivating layer derived from the waste-rock column solution appeared to be more effectively decreasing pyrite reactivity than that derived from water leaching.  4.5.2  Louvicourt Pyrites  A s with sample mineralogy and evolution o f precipitate coatings, the Louvicourt-1 pyrite showed electrochemical responses after each leaching period, that were very similar to Huckleberry solution-leached pyrites (Figure 4.16). The loss o f reactivity was slightly lower i n the Louvicourt-1 sample than i n the Huckleberry solution-leached sample, with a net positive displacement o f only 0.055 volts compared to 0.07 volts for Huckleberry.  Contrary to Louvicourt-1, Louvicourt-2 pyrite became more reactive after 4 weeks o f leaching, with a negative displacement o f the point o f initial oxidation o f 0.05 volts (Figure 4.17). In the 10 and 2 0 week, however, pyrite reactivity decreased, resulting i n th  th  a net positive advancement o f 0.04 volts over the 20 weeks o f leaching with respect to the unleached pyrite.  The gain o f reactivity within the first 4 weeks o f leaching coincide  with the preferential oxidation and loss o f sphalerite over pyrite i n the sample observed under S E M and supported by the leachate chemistry data.  The subsequent passivation  could be attributable to the formation o f surficial, amorphous iron precipitates observed  The terms 'passivation' or 'passivated' in this study refer to a relative decrease in reactivity of the mineral measured as a positive displacement of the point of initialization of mineral oxidation, observed by cyclic voltamperometry. This term is not used to describe the mechanisms or the products responsible for the observed loss of reactivity. 2  76  in appreciable quantity i n the 20-week leached sample.  These precipitates could have  effectively passivated pyrite surfaces.  4.5.3  Tizapa pyrite  Tizapa pyrite reactivity evolved i n a similar way to that o f Louvicourt-2 pyrite, although at a smaller scale. Figure 4.18 shows an initial increase i n reactivity in the 4 leaching, followed by a reactivity decrease which remained until the 2 0  th  th  week o f week o f  leaching. A net positive displacement o f 0.05 volts occurred after 20 weeks o f leaching with respect to the unleached pyrites. The higher resistance to oxidation observed i n the 10 week compared to the 2 0 week, suggests that the precipitate coatings formed at an th  th  early stage (10 weeks) were different than those formed after 20 weeks o f leaching.  4.5.4 Zimapan pyrite The Zimapan voltammogram showed an increase i n the width o f the anodic peak with leaching time (Figure 4.19). W i t h advancing leaching cycles, the reverse scan o f the anodic peak released a consistently higher current than the forward scan, measured at a given potential. The forward scan showed a net advancement o f the initialization point o f pyrite oxidation o f 0.1 volts i n 20 weeks o f leaching.  The reverse scan showed an  increasing efficiency to oxidize the pyrite surfaces after initialization.  A possible  explanation for this behaviour is the observed increased abundance o f precipitates combined with the increased surface area created by corrosion pits as leaching progressed.  The thicker precipitate coating possibly offered increasing resistance to  oxidation, but once the potential became sufficiently high to break this barrier, more 77  current could be generated from the same grains because o f the increased surface area created by the corrosion pits. Although pit density remained constant with advancing leaching cycles, pits have a tendency to deepen inside the crystal rather than to enlarge at the surface o f the grain (Mustin, 1992). Consequently, the actual surface area o f a pitted grain could be increased without a notable increase i n the number or width o f the pits. The effect o f oxidation pits on the reactivity o f pyrite was not verified i n this study but merits further investigation.  The presence o f galena i n the unleached sample was observed i n Figure 4.2, Section 4.2. A close-up view o f the 4, 10 and 20-week anodic scan curves in Figure 4.20 shows that galena was no longer present in the sample after the 4  th  week o f leaching.  SEM  observations and leachate chemistry indicated that most o f the available galena was oxidized i n the first 4 weeks o f leaching where lead was reprecipitated as anglesite throughout the sample.  4.5.5 Brunswick The forward scan curves o f the Brunswick pyrite voltammograms showed an apparent increase i n reactivity with leaching time up to the 10  th  week o f leaching, followed by  relative decrease i n reactivity i n week 20 (Figure 4.21). The presence o f galena i n the unleached, 4-week and 10-week samples, however, concealed the exact point (or voltage) at which pyrite started to oxidize. It was therefore impossible to calculate the amount o f curve displacement. The apparent gain i n reactivity observed until week 10 was similar to the Louvicourt-2 response and also most likely attributable to the presence o f  78  sphalerite. Although sphalerite was not visible under S E M i n the leached samples, solid phase zinc concentration (corresponding to sphalerite) was the highest i n the Brunswick sample. Leachate chemistry indicated that large concentrations o f zinc were continually being leached form the Brunswick sample whereas iron was not. L i k e in the Zimapan and Louvicourt-2 pyrite samples,  sphalerite was most likely providing galvanic  protection to pyrite at least i n the early cycles.  With increasing leaching time, the  presence o f iron hydroxides and anglesite precipitates on pyrite surfaces may have reduced the availability o f sphalerite to oxidize in lieu o f pyrite. A close-up view o f Figure 4.22 indicated that the electrochemical signature o f galena was still discernible after 10 weeks o f leaching, although very diminished, and was absent form the 20-week voltamogram. Galena along with anglesite was observed under S E M in the 4-week sample, galena was not seen in the 10 and 20-week samples.  The reverse scan o f the Brunswick voltammogram showed a pattern similar to the Zimapan pyrites where the anodic peak widened with advancing leaching cycles. Oxidation pits were more abundant after 20 weeks o f leaching i n the Brunswick sample which, again, could possibly explain the increasing generation o f current on the reverse scan.  79  (vn) juajjnQ  4.5.6  Evolution of Pyrite Reactivity  Figures 4.23 to 4.25 show the voltametric responses o f the group o f pyrites after 4, 10 and 20 weeks o f leaching respectively.  The differences in reactivities between pyrites  were significantly decreased after 4 weeks o f leaching.  The opposing processes o f  increased passivation for the initially reactive pyrites (Huckleberry, Louvicourt-1) and loss o f galvanic protection for the initially less reactive pyrites (Louvicourt-2 and Tizapa and possibly Zimapan and Brunswick) appear to have effectively homogenized electrochemical responses o f the various pyrites.  Individual reactivity  the  differences  appeared again after 10 weeks o f leaching, developing further after 20 weeks as the characteristics  o f the pyrite surfaces were  likely dominated  by the presence o f  precipitation products and possibly oxidation pits.  Huck -water Huck -leach Louvicourt 1 Louvicourt 2 Tizapa Zimapan Brunswick  200  250  300  350  400  Potential (mV) Figure 4.23 All Pyrites, 4-week Leached 85  450  Table 4.5 shows the relative reactivity o f each pyrite at different leaching periods measured at the point o f 2.0 u A current release on the voltammograms (on close-up views not presented).  Samples exhibiting similar reactivities appear i n the same box.  Although the reactivity o f Huckleberry pyrites continually decreased with advancing leaching cycles, they remained the most reactive o f all pyrites after 20 weeks o f leaching. Similarly, Tizapa and Louvicourt-2 remained the least reactive.  The comparison with  Brunswick pyrite reactivity is difficult to establish because o f the galena peak effectively hiding the initialization point o f pyrite oxidation i n the unleached sample and after 4 and 10 weeks o f leaching.  This study suggested that the reactivity o f the pyrite samples studied depend greatly on the occurrence o f mineral impurities and, after some period o f leaching, on the composition o f the precipitate layer. The latter could be dictated by the various mineral phases being oxidized simultaneously with pyrite.  Table 4.5 Relative Reactivities of Pyrites  unleached  4 weeks  10 weeks  20 weeks  Huckleberry Louvicourt-1 Louvicourt-2 Tizapa  Huck. Water, Louvicourt-1, Huck. Column Louvicourt-2, Zimapan, Tizapa,  Huck. Water Huck. Column Louvicourt-1 Zimapan Tizapa Louvicourt-2  Huck. Water Huck. Column, Louvicourt-1 Zimapan Tizapa, Louvicourt-2, Brunswick  (Brunswick?)  (Brunswick?)  t  > o CU S-l ' in  cd <L>  t-i  o C  (Zimapan?) (Brunswick?)  87  5  5.1  RESULTS - C E M E N T E D PASTE BACKFILL STUDY  CHEMICAL ANALYSES OF TAILINGS, BINDER AND PASTE  Chemical analyses o f the tailings, paste and slag are presented i n Table 5.1. The paste mixture compositions were calculated from the chemical data o f Table 5.1 and are presented i n Table 5.2.  Analyses indicate that Tizapa tailings possess the highest  concentrations o f sulfide, iron, copper and zinc o f all other tailings.  They have the  second highest concentration o f lead and arsenic after Brunswick. Brunswick tailings have a composition similar to that o f Tizapa, with high concentrations o f sulfide, iron, zinc, lead and arsenic.  The maximum pyrite content is estimated at - 7 0 % for both  Brunswick and Tizapa tailings, assuming all the iron occurs as pyrite. This estimate is slightly higher than values reported by Brunswick M i n i n g Division Inc. (40 to 65% pyrite) (Noranda Technology Centre, 1998e). The discrepancy probably lies i n the 0-5% pyrrhotite content and the iron o f the sphalerite not being taken into account i n this summary calculation. The Tizapa estimate falls within the values reported by Penoles (Ybarra, 1998). Tizapa tailings also have the lowest concentration o f calcium whereas Brunswick has the highest.  F I M and Louvicourt tailings have a similar composition,  containing approximately one-third the concentration o f sulfide and half the iron than Brunswick or Tizapa, giving a maximum pyrite content o f 35% for each tailings. This estimate lies within the Noranda Inc. calculations o f 16 to 47% pyrite for Louvicourt. F I M and Louvicourt tailings also have l o w concentrations o f zinc, lead copper and arsenic compared to the other tailings.  A l l tailings have l o w concentrations 88  of  carbonates, a parameter associated with limestone (CaCCh) or dolomite (CaMg(C03)2), suggesting that all tailings have a low acid neutralization potential.  The chemical analyses presented in Table 5.2 indicate that the principal component o f O P C is calcium corresponding to the content o f tri- and di-calcium silicates ( C a S i 0 5 and 3  Ca2SiC>3 respectively) and tri-calcium aluminate ( C a s A L ^ ) , the principal binding agents of OPC.  The granulated blast furnace slag used i n the Louvicourt formulation has a  relatively l o w iron content and is therefore favourable for use against sulfate rich waters (Kosmatka et al., 1995). The slag is also rich i n calcium and silicon, carbonates and magnesium.  89  Table 5.1 Chemical Analyses of Tailings and Binders Used in Backfi I Mixtures  Water  %  Cement 1.3  S asSO, %  0.2  Stotal  s-  10.5 0.2  Tiz 37.4 0.4  34.3 3.3  8.2 16.1 0.11 0.03 0.001 <0.001 0.02 0.007 42.5 0.6 9.6 0.44  10.3 16.7 0.66 0.07 0.001 <0.001 0.005 <0.001 34.6 2.3 5.7 0.08  37.0 34.4 2.10 0.38 0.180 <0.005 0.31 O.001 11.2 2.3 2.7 0.07  31.0 30.5 1.08 1.08 0.100 0.007 0.37 0.13 21.7 1.6 1.1 0.14  2.8 6.7 0.20 0.02 1.3 89.5  8.8 2.0 1.44 0.06 2.4 85.3  0.7 0.7 0.09 0.05 0.4 92.9  14.3 4.2 0.12 0.11 0.2 110.97  %  1.7  Al 0 Na 0  %  4.5 0.63  CaO MgO MnO K 0 C Sum:  %  69.0 1.4 0.02 1.10 0.6 100.3  44.2 12.7 0.74 0.42 0.3 106.2  Fe Zn Pb Cu Ni As Cd Si0  co  2  3  2  3  2  1.1 1.2 0.007 0.003 0.057 <0.001 <0.005 <0.001 18.9  9.2 1.0  0.9 0.3 0.002 0.002 0.022 O.001 O.005 O.001 31.7 7.7 6.9 0.09  % % % % % % % % %  2  % % % %  2  %  For all solid samples: Bi <0.001%, Cr <0.02%, Se<0.001%, Te<0.002% L.O.I, not reported  Table 5.2 Final Composition of Paste Samples Element Stotal  S0 Sulfide Fe Zn Pb Cu Ni As Cd Si0 4  2  co  % % % % % % % % % % % %  3  Al 0 Na 0  %  CaO MgO MnO K 0 C Sum:  % % % %  2  3  2  2  %  %  Element  1.1 0.2  Louv 8.8 1.0 7.9 15.3 0.11 0.03 0.002 <0.001 0.02 <0.001 41.9 0.8 9.4 0.43  10.2 0.2 10.0 16.2 0.64 0.07 0.003 <0.001 0.005 0.005 34.1 2.3 5.6 0.10  Tiz 35.1 0.4 34.7 32.3 1.97 0.36 0.17 <0.005 0.29 O.001 11.7 2.3 2.8 0.10  4.9 6.8 0.22 0.04 1.3 90.2  10.6 2.0 1.40 0.09 2.3 85.7  5.0 0.7 0.09 0.12 0.4 93.3  32.6 3.2 29.5 29.0 1.03 1.03 0.10 0.007 0.35 0.12 21.6 1.6 1.3 0.16 17.0 4.0 0.12 0.16 0.2 110.44 90  41.8  S0  0.47 <0.1 <0.1 <0.05 <0.1 < 1.0  Fe Zn Pb Cu Ni As Cd Si  —  28.2 —  co  4  3  <0.3 64.8  Al Na  31.4 0.77 <0.1 8.7  Ca Mg Mn K C  —  : not analysed  5.2  A S P E C T OF C U R E D PASTE S A M P L E S  T h e F I M paste samples w e r e v e r y friable a n d h a d little c o h e s i o n after 14 days o f h u m i d curing.  S o m e F I M samples were s l i g h t l y damaged u p o n d e m o u l d i n g a n d p a r t i a l l y  d i s s o c i a t e d o r c r u m b l e d w h e n first b e i n g exposed to water o r the Fe2(S04)3 s o l u t i o n .  L o u v i c o u r t samples w e r e also r e l a t i v e l y soft a n d friable after 14 days o f h u m i d c u r i n g a l t h o u g h m o r e resistant than F I M samples.  N o d i s s o l u t i o n o r fracturing w a s observed  u p o n i m m e r s i o n o f the samples.  T h e c u r e d paste samples o f B r u n s w i c k a n d T i z a p a w e r e m u c h harder a n d appeared to have l o w e r p o r o s i t y that the other t w o mixtures.  T i z a p a paste w a s the densest a n d  hardest o f a l l m i x t u r e s . D u r i n g the c u r i n g p e r i o d , T i z a p a samples f o r m e d a t h i n o x i d i z e d c o a t i n g o n the paste surface i n s i d e the m o u l d ( F i g u r e 5.1). T h e coating penetrated less than 1 m m and w a s sanded d o w n to expose fresh surfaces to the l e a c h i n g solutions.  Figure 5.1 Superficial Oxidation of Freshly Cured Tizapa Paste Samples  91  5.3  WATER ABSORPTION BY THE PASTE  Measurements o f water absorption in the puck and cylinder samples after 5 and 20 weeks of leaching are presented in Tables 5.3 a and b respectively. With the exception o f the Louvicourt puck sample, the percentage o f water absorbed by the pucks and cylinder samples  remained  relatively unchanged  i n both  the  flooded  and  cycled  water  environments (Figures 5.2 and 5.3). Tizapa paste had an average water content o f 15.7 % after each leaching period, similarly Brunswick paste had a water content around 17.8 % , F I M 22.8% and Louvicourt 23.2 % water.  The paste with the lowest water content  (Tizapa) corresponding with the mixture having the highest proportion o f cement. Following this trend, Brunswick, F I M then Louvicourt had decreasing amounts o f cement and increasing water contents.  Equal water contents for samples o f similar composition but very different sizes and shapes (such as pucks and cylinders o f one particular mixture) suggest that the samples were completely saturated with water and with the ferric sulfate solution. Saturation o f the paste is likely achieved within 24 hours after immersion since the cycle-leached samples had similar water contents to the flooded-leached pastes.  Water absorption i n the ferric sulfate environment seemed to indicate an increased water content with time for all samples (Figure 5.4). Given that the samples became saturated with water shortly after immersion, the apparent increase in water content is instead a loss o f solid mass with leaching time due to dissolution o f the sample.  92  » pepAo - j a j B M  93 511.6  TIZAPA # 7 PUCK  CYLINDER  pepooy. - j e j B M 19.5 17.9  38.4 475.5  iri CO CM  CM  a  38.6 380.3  20.4 19.2  in  48.5 470.4  23.9  26.6  T—  BRWK #7 PUCK CYLINDER  24.0  32.6  396.2 301.5  CM  CYLINDER  25.5  orCO  30.9 390.6  23.4  295.9  397.4  29.4  oo  CYLINDER BRWK #8 PUCK CYLINDER  25.0  14.4  19.2 409.0 313.4  9.6  13.6  16.3 496.2 415.5  T  CYLINDER LOUV # 8 PUCK  PUCK  CYLINDER FIM #3  CN  22.9  318.9  413.5  CYLINDER LOUV#4 PUCK  25.8  16.0  21.5  PUCK  q CO CM  15.4  jeiBM  432.9  CO  19.3  17.4  18.7  <p CO CM  33.5  41.5  TIZAPA # 2 PUCK  404.2  43.9  CO  16.4  54.0 489.2  CYLINDER  CN  FIM #8  43.8  CYLINDER  pepoou -  17.5  >8 e>  36.6  383.5  17.6  33.3  40.5  O) CM CM  297.6  CM  23.3 388.1  CYLINDER  csi  BRWK # 5 PUCK  00  464.6  22.8  310.7  402.4  23.8 56.3  42.9  22.9  411.0 316.8  22.6  CYLINDER LOUV # 3 PUCK  25.6  o  CYLINDER BRWK # 10 PUCK  23.2  iri  40.2  00  52.4  23.4  33.1  in  280.2  PUCK  16.2 15.6  O)  FIM #7  423.6  44.3  CO  22.3  52.9 502.0  CYLINDER  c>  365.8  24.1  15.7  15.8  17.8  TIZAPA # 4 PUCK  17.8  38.5  467.2 384.1  46.8  CYLINDER  e> O  CYLINDER LOUV#2 PUCK  CYLINDER FIM # 1 1 PUCK 31.0  TIZAPA # 5 PUCK  17.2  23.0 23.2  CO CM  35.6  397.6  480.4  CYLINDER  BRWK #6 PUCK  286.7  30.8  40.0 373.2  CYLINDER  CM CO CM  18.2  (O CM CM  22.60  CN  419.5  23.5  28.7  CYLINDER BRWK # 1 PUCK  .—  s  42.3  309.7  CO ui  23.0  m'  CN  497.4  6.1  30.9  32.9 400.2  294.2  382.0  CO  CYLINDER LOUV # 5 PUCK  23.4  15.1  15.5  difference in water ab orption  3  22.6  22.8  409.3 315.8  9.8  12.7  PUCK  426.9  502.8  CYLINDER FIM # 10  44.5  52.7  %H20 average  O  CYLINDER LOUV#9 PUCK  22.5  15.5  14.6  12.0  PUCK  428.1  X a?  501.5  CM  FIM #4  O TIZAPA # 3 PUCK  a  15.9  5 ~  CYLINDER  Q ~  22.5  % H20 average 2  26.7  W dry (g) •a -g;  TIZAPA #6 PUCK  W wet (g) O CM X ^8 e> e> CO  CO 00  CM  o>  uoijnios e ( K ) S ) 2 9 d  oo  uounps e(frOS)39d  Tizapa  ni  ii  •  FiM Louv Bwk  *  AK  5  10  15  20  weeks leached  Figure 5.2 Water Absorption in Cyclic-Leaching (water) Environment  -Tizapa  o  u——  -Louv  3K—  5  -FiM  JK  10  -Bwk  20  15  weeks leached  Figure 5.3 Water Absorption in Flooded-Leaching (water) Environment  weeks leached  Figure 5.4 Water Absorption in Fe (S04)3 Solution 2  94  5.4  CHEMISTRY OF PASTE LEACHATES  Information on the effect o f the different chemical environments on paste was obtained by comparing the chemistry o f leachates o f the different leaching environments and the solid phase chemistry o f the unleached samples.  Results indicated that  leachate  chemistry was controlled more by the characteristics o f the leaching environments and the cement content o f the paste mixtures than by the mineralogy o f the different tailings.  5.4.1  Flooded and Cycled Water-Leached Environments  The chemistry o f the  flooded-environment and cycled-environment leachates  presented i n Appendix II (Tables I I - l and II-2 respectively).  are  Figures 5.5 to 5.32 present  graphs o f element concentrations with advancing leaching cycles for each water-leached environment. The solid phase concentration o f the element or ion appears i n the legend box o f each graph. Depletion rates were calculated for the elements showing the greatest change with time (calcium, potassium and magnesium) in order to verify the effect o f the different leachate volumes extracted from each leaching environment.  Flooded environment p H values remained relatively high (pH 8-9) throughout the 20 weeks o f leaching (Figure 5.5), whereas the cycled environment p H reached steady state at p H 6-7, closer to the p H o f the distilled water (Figure 5.6). Leachate conductivities were much higher in the early cycles o f the flooded environment and reached steady state at a slightly higher level that the leachate o f the cycled environment (Figures 5.7 and 5.8). Conductivity measurements are indicative o f a greater dissolution o f the sample i n the  95  flooded  environment  compared  to  the  cycled  environment.  Redox  potential  measurements are on average higher i n the cycled environments than i n the flooded environments (Figures 5.9 and 5.10). The higher redox levels in the cycled environment leachate are due to a higher dissolved oxygen concentration resulting from the exposure o f air between each cycle.  Trace concentrations o f dissolved iron are present i n F I M and Louvicourt leachates o f both the flooded and cycled environments in the early cycles o f leaching (Figures 5.11 and 5.12).  The aqueous iron concentrations appear to be unrelated to the iron content o f  the paste mixtures.  The presence o f aqueous iron either is from dissolution the  tetracalcium aluminoferrite phase o f the cement or from the dissolution o f pre-existing soluble iron salts originally contained in F I M and Louvicourt tailings.  Zinc concentrations in the flooded environment indicate that zinc is not mobile with the exception o f Brunswick in the first 3 cycles o f leaching where concentrations o f zinc (0.2-0.3 mg/1), slightly above detection limit (0.1 mg/1), are leached out o f the paste (Figure 5.13). Higher zinc mobility occurred in the cycled environment throughout the leaching cycles where zinc was mobilized from all pastes (Figure 5.14).  Cycled  environment zinc concentrations were close to detection limit for Louvicourt and F I M samples, both o f which contain the lowest solid phase zinc concentration. Brunswick and Tizapa samples showed higher concentrations o f zinc throughout the leaching cycles compared with the flooded environment and a general increase i n zinc concentration with leaching time.  Leachate zinc concentrations appear to be proportional to solid phase  concentrations o f zinc and possibly inversely proportional to the cement content o f the  96  samples.  Unlike zinc, lead was not very mobile i n either water-leached environment.  In the  flooded environment, Louvicourt, F I M and Brunswick leachate lead concentrations remained close to or below detection limit throughout the 20 weeks (Figure 5.15). N o lead was leached out o f the Tizapa samples.  L o w quantities o f lead were leached out  from all samples i n the early cycles o f leaching and out o f Louvicourt and Tizapa i n the last cycles (Figure 5.16).  A direct correlation cannot be made between solid phase lead  concentrations and leachate concentrations.  Sulfate concentrations followed a decreasing trend for all samples i n both the flooded and cycled water environments (Figures 5.17 and 5.18). p H values were too high to allow for significant sulfide oxidation hence the sulfate in the leachate is most likely derived from the initial sulfate content o f the tailings, most probably sulfate salts precipitated from the tailings water upon drying o f the tailings. Indeed, sulfate concentrations i n the leachate were proportional to the solid phase concentrations o f sulfate.  Calcium concentrations also showed a decreasing trend in both the flooded and cycled environments (Figures 5.19 and 5.20). Initial flooded leachate concentrations o f calcium were between 682 mg/1 ( F I M ) and 922 mg/1 (Brunswick), falling to 185 mg/1 and 590 mg/1 respectively after 20 weeks o f leaching. Louvicourt and Tizapa concentrations fall between those o f F I M and Brunswick. The difference i n the time series concentration o f calcium between both environments resulted from the different leachate sampling protocols. The calcium depletion curves are in fact similar for both environments (Figures 5.21 and 5.22).  The flooded 97  environment  showed  a  slightly  slower  depletion in the early cycles but quickly reached the values o f the cycled environment i n the later cycles.  Aqueous concentrations o f magnesium, potassium and silicon in both flooded and cycled environments are proportional to the solid phase concentration o f each sample for the respective element (Figures 5.23 to 5.32).  In the early cycles o f both water-leached environments, Louvicourt leachates had considerably higher concentrations o f magnesium than the other samples (Figures 5.23 and 5.24). Magnesium concentrations reached steady state shortly thereafter. Louvicourt concentrations remained slightly higher than the other samples.  A correlation can be  made between aqueous magnesium concentrations and the slag component o f the Louvicourt paste, the latter having the highest solid phase concentration o f magnesium o f all paste ingredients. It is most probable that the magnesium i n the leachate originated from the dissolution o f the incompletely cured slag component o f the Louvicourt sample. In fact, the granulated blast furnace slag used by Louvicourt requires up to 21 days o f curing before participating in cementicious reactions (Dallaire, 1997). The steady state conditions reached after the first cycles o f leaching were probably the consequence o f a more complete curing o f the binder, reducing the availability o f magnesium. Depletion profiles o f magnesium were similar for all samples in both water-leached environments, with the exception o f Tizapa (Figures 5.25 and 5.26). After 20 weeks o f leaching, the depletion o f magnesium from the Tizapa sample in the cycled environment was double that o f the flooded environment.  The continued depletion o f magnesium from Tizapa  may be related to the dissolution o f a magnesium mineral (i.e. dolomite: C a M g ( C 0 3 ) 2 ) ,  98  unaffected by the curing o f the cement phase.  In the case o f potassium, the most important source o f that element in the paste is the cement phase (refer to Table 5.2).  The highest leachate concentrations o f potassium  coincide with the Tizapa pastes, characterized by the highest proportion o f cement (Figure 5.27 and 5.28).  Potassium depletion rates are similar for both water-leached  environments for all samples, with the exception o f Louvicourt, which had a higher K depletion rate i n the flooded environment (Figures 5.29 and 5.30).  Silicon concentrations remained l o w in both water-leached environments (Figures 5.31 and 5.32). The general stability o f silicate minerals in the p H and redox conditions o f the leaching cells suggests that silicon is dissolving from the more soluble cement phases such as the tobermorite gel (Ca3Si2CV3H2C>), the principal binding agent o f cement. Continued decomposition o f this phase would have serious consequences on the stability o f the binder.  99  o o  C  N  O  t  O  H  d  r  O  O  o *  O  o o  O  o  O CM  o  O  r-  (S ) AiiAipnpuoo n  (I/6UJ) qd  (1/6) *os  o o  o o  O  CO  #  i  ?  s>  ?  <P  o o <D  o o  t  (l/Biu) BO  ( | / 6 l U ) BQ  o o CN  o  8  § 8 S 5 uoi}B|dap BO %  uouaidap eo %  o  CO  o  CO  o  tO  o  lO  o  o  CN  o  CN  0>  o  O  o  (0  o  CD  (|/6iu) 6|A|  o  CO  o  CO  o  o  m  co  csi  1 -  uojjeidap 6|/\| %  o  (|/6uj) is  5.4.2  Fe2(S04)3  Solution-Leached Environment  The chemical results o f the ferric sulfate leachate solutions are presented i n Appendix II (Table II-3). Time series concentration graphs, as well as calculated depletion rates for calcium, potassium and magnesium, are shown i n Figures 5.33 to 5.50.  The acidic p H (2.4 - 2.6) o f the ferric sulfate solution was almost instantly neutralized when coming in contact with the paste sample (Figure 5.33). In the first leaching cycles, ferric sulfate solutions were buffered to a p H o f 6-7 for all samples. After 40 cycles o f leaching (20 weeks), Louvicourt, Brunswick and Tizapa solutions were buffered to an average p H o f 3.5 whereas F I M leach solution was buffered to a slightly higher average p H o f 5. Conductivities o f the leachate solutions show a considerable drop i n the first four cycles o f leaching, stabilizing at a lower level (1500 uS) than that o f the ferric sulfate solution (2300 uS) (Figure 5.34).  This indicates that the ions are precipitating  from the solution upon contact with the samples. The ferric sulfate solution added to the samples bi-weekly has an average redox potential o f 620 milivolts S H E . Initial leachate redox values (after 3 to 4 day contact periods with the samples) are around 130 m V , rising to near ferric sulfate solution values in the final cycles (Figure 5.35). Louvicourt, Brunswick and Tizapa leachates had final redox values around 480 m V whereas F I M final redox conditions were around 330 m V . The evolution o f the redox values i n all leachates again indicates that the ferric sulfate solution is decreasingly altered or buffered by the sample as leaching progresses.  In the initial stages o f leaching, dissolved iron concentrations i n the leachate were much  108  lower than those o f the ferric sulfate leaching solution. Iron was precipitated out o f solution onto the surfaces o f the paste samples as a dark orange ferric hydroxide precipitate (Figure 5.36).  A s leaching progressed, leachate iron concentrations o f all  samples slowly increased, approaching the values o f the ferric sulfate solution (Figure 5.37). This indicates that the ferric hydroxide coating was forming much more slowly.  Unlike the water-leached environments, a considerable amount o f zinc was continuously leached out o f the Brunswick paste sample, averaging a concentration o f 15 mg/1 throughout the leaching cycles (Figure 5.38). Zinc was also leached out o f the Tizapa sample (average o f 4 mg/1), increasing i n the last five cycles to about 9 mg/1. Louvicourt and F I M samples released l o w concentrations o f zinc relative to the other samples (below 3 mg/1) but higher than i n the water leached environments. The samples with the highest solid phase zinc concentrations (Tizapa and Brunswick) yielded the highest leachate concentration o f zinc.  The mobility o f lead was slightly higher in the ferric sulfate environment than i n the water environments (Figure 5.39).  Tizapa and Brunswick leachates had the highest  concentration o f aqueous lead throughout the leaching cycles, both slowly increasing to an average o f 1.6 and 1.2 mg/1 respectively in the last 10 cycles. Lead concentrations i n the Louvicourt leachate were close to detection limit throughout the leaching cycles whereas F I M showed trace concentrations o f lead only in the final cycles. A s with zinc, lead concentration i n the leachate appears to be proportional to the solid phase concentration.  In the first cycles, sulfate concentrations in  the leachate were greater than the leaching 109  solution sulfate. After the third leaching cycle, leachate concentrations were essentially the same as the ferric sulfate solution. These concentrations were maintained throughout the remaining leaching cycles. This suggests that the sulfate ion o f the leaching solution is not participating i n cement-altering reactions within the pastes.  Calcium  was  leached from  all samples  throughout  the  test.  Relatively  high  concentrations were leached i n the early cycles, decreasing to steady state around the 10 leaching cycle (Figure 5.41).  th  The variation o f calcium concentrations between the  different samples is much greater i n the ferric sulfate environment than in water-leached environments. Louvicourt had the lowest aqueous calcium concentration as w e l l as the lowest solid phase concentration o f calcium.  F I M and Brunswick had generally the  highest  as well  leachate  concentration.  concentrations o f calcium  as the  highest  solid  phase  Leachate calcium concentrations are proportional to the solid phase  concentration prior to leaching. Ferric sulfate-leached depletion rates o f calcium (Figure 5.42) were slightly higher than in the water environments with the exception o f F I M for which the depletion rate was double that o f the water environments.  As  with the water-leached environments, aqueous  concentrations  o f magnesium,  potassium and silicon are also related to the solid phase concentration o f the element. Aqueous magnesium concentrations were highest in the Louvicourt leachate, which also possessed the highest solid phase concentration (Figure 5.43).  In the initial cycles,  leachate M g concentrations were similar to those o f the flooded environment. A s leaching progressed, however, the M g concentrations remained high i n the ferric sulfate solution. Tizapa, Brunswick and F I M leachate concentrations o f M g were also slightly  110  higher i n the ferric sulfate environment than in the flooded water environment. Depletion rates o f M g were, in fact, much higher i n the ferric sulfate environment (Figure 5.44) compared to the water-leached environment.  Potassium concentrations o f each leachate followed a trend similar to that o f the flooded environment (Figure 5.45). Leachate concentrations o f K in the first cycles were similar for both environments. However, the decrease i n concentrations occurred more suddenly and earlier i n the ferric sulfate environment, falling to near detection limit around the 5 leaching cycle.  th  Depletion rates o f K are similar i n all three environments, with the  exception o f Louvicourt for which the depletion rate is lower in the ferric sulfate than i n flooded water (Figure 5.46).  Leachate concentrations o f silicon were similar to those o f the flooded environments i n the initial cycles, the highest concentration being that o f the Louvicourt leachate (Figure 5.47).  Contrary to the water-leach environments, silicon concentrations in the ferric  sulfate environment showed a continued increase for all samples.  Given the l o w p H  conditions o f the leachate, especially i n the final cycles, the dissolved silicon i n this environment may be leached from silicate minerals as well as the tobermorite phase o f the binder.  Ill  Figure 5.33 Fe (S0 ) Cells - pH 2  4  3  4000 1 3  3000  -*  Tiz  "*•  Louv  ""•  Brw  -•  FIM  •- - -Solution •  1  4  7  10  13  16  19 22  25  ,  28  R  —  31  |  •  34  n o . of c y c l e s  Figure 5.34 Fe (S0 ) Cells - Conductivity 2  4  3  |  I  37 40  Figure 5.36 Ferric Hydroxide Coating on Paste Sample  1400  1  4  7  -*  Tiz (32.3% Fe)  -•  Louv (15.4%)  -•  Bwk (29.0%)  -*  FIM (16.2%) - "Solution  10 13 16 19 22 25 28 31 34 37 40 n o . of c y c l e s  Figure 5.37 Fe (S04>3 Cells - Dissolved Iron 2  30 -Tiz (2.0%Zn)  25 20  -Lou (0.1%)  15 -Bwk ( 1 . 0 % )  10 5  -FIM (0.6%)  0  V  1  4  I  I  I  I  I  I  I  I  I  I  1  7  10 13 16 19 22 25 28 31 34 37 40  1  n o . of c y c l e s  Figure 5.38 Fe (S0 ) Cells - Dissolved Zinc 2  4  3  113  ' 1  10 •Tiz ( 0 . 3 6 % P b ) - L o u (0.03%) •Bwk ( 1.0% ) "FIM (0.07%) 1  4  7  10  13  16  19  22  25  28  31  34  37  40  no. of cycles  Figure 5.39 Fe2(S0 )3 Cells - Dissolved Lead 4  (0.4%SO4) Louv (1.0%) •Bwk (3.2%) •FIM (0.2%) -  - -Solution  Figure 5.40 Fe2(S04)3 Cells - Sulfate Concentration  - * — Tiz (5.0% C a O ) - • — L o u v (4.9%) - • — B w k (17.0% ) ••—FIM (10.6%)  n o . of c y c l e s  Figure 5.41 Fe (S0 )3 Cells - Calcium Concentration 2  4  114  50  no. of c y c l e s  Figure 5.42 Fe2(S04)3 Cells - Calcium Depletion Rate  180  1  4  7  10 13  16 19 22 25 28 31  34 37 40  n o . of c y c l e s  Figure 5.43 Fe2(SC>4)3 Cells - Magnesium Concentration  10  1  4  7  10  13  16  19  22  25  28  31  34  37 40  no. of cycles  Figure 5.44 Fe (S0 )3 Cells - Magnesium Depletion Rate 2  4  115  180 -Tiz ( 0 . 1 2 % K 2 O )  150 s - 120  -Louv (0.04%)  "Si E  90  *  60  •Bwk (0.16%) -FIM (0.09%)  30  l t t l W M . H I JIJMMiJUlJWJ.—•••••••••  0 1  4  7  10  13  16  19 22  25  28  31  34  37 40  n o . of c y c l e s  Figure 5.45 Fe2(S04)3 Leach Cells - Potassium Concentration  Figure 5.46 Fe (S0 ) Leach Cells - Potassium Depletion Rate 2  4  3  Tiz ( 1 1 . 7 % S i 0 2 ) - • — L o u v (41.9%) - • — B w k (21.6%) - • — FIM (34.1%)  Figure 5.47 Fe (S0 )3 Leach Cells - Silicon Concentration 2  4  116  5.4.3  Summary of Observations from Leachate Chemistry  The chemistry o f the leachate together with the initial concentrations o f each element i n the paste provide the following information on processes occurring in the first 20 weeks o f leaching with water and artificial A R D solution:  Water-leached environments: •  The paste samples provided a high immediate buffering capacity to the leaching water.  The buffer was most likely coming from the dissolution o f the portlandite  phase o f the cement (Ca(OH)2). Higher p H i n the flooded environment suggests that portlandite dissolution was more intensive i n the flooded environment although calcium depletion rates were similar for both environments. The longer contact time between the leaching solution and the solid sample may have allowed for the precipitation o f a secondary calcium phase within the samples, thereby explaining the similarity between calcium depletion rates.  •  Conductivity values were indicative o f the extent o f sample dissolution.  Higher  conductivities occurred at steady-state conditions i n the flooded environments for all samples.  This indicates that more intensive dissolution occurred i n the flooded  environment throughout the leaching cycles because o f longer contact times between the samples and the leaching solution.  •  The combination o f high sulfate concentrations and l o w concentrations o f iron and zinc suggests that sulfate was not the product o f sulfide oxidation but rather was leached from a soluble sulfate phase  present i n the tailings (prior to cement 117  addition). Considering the high concentrations o f sulfate o f the tailings water (Table 2.2, Chapter 2), the sulfate in the samples likely originated from a sulfate salt precipitated when the tailings were dried, and redissolved when coming i n contact with the leaching solutions.  •  Comparison o f major ion chemistry (Ca, K , M g , and Si) i n the leachate and i n the solid phase o f the samples suggests that the binder was actively dissolving. Calcium was most likely leached from the highly soluble portlandite phase o f the cement (Ca(OH)2) (dissociation constant K e q = 6.3E22, Appelo and Postma, 1994; Pakhurst, 1995) whereas silica and potassium were likely dissolved from the tobermorite gel phase. Magnesium, i n relative abundance only in the Louvicourt leachate, was likely dissolved from the slag portion o f the binder. The decreasing loading o f major ions suggests the following: 1) a depletion o f the dissolving phases i n the outer layers o f the paste sample and/or 2) a continued curing o f the sample, binding the major ions into more stable, less soluble hydrated cement phases.  Ferric sulfate-leached environment: •  Leachate p H , redox potentials and aqueous iron concentrations evolved towards the values o f the ferric sulfate leaching solution with advancing leaching cycles.  This  suggests that the ferric sulfate solution was decreasingly modified or buffered by the paste with which it came i n contact.  •  The leaching solution sulfate did not appear to be consumed i n secondary, expansive mineral growth reactions in the samples over the 40 cycles o f leaching, as the sulfate  118  concentration o f the leaching solution and the leachates remained unchanged throughout the leaching period.  The relatively high aqueous zinc concentrations i n the Brunswick sample suggests either 1) active oxidation o f sphalerite that occurred in the outer layers o f the paste and/or 2) dissolution o f a soluble zinc phase (sulfate, oxide or hydroxide) present i n the tailings. The pyrite reactivity study indicated that sphalerite present in the Brunswick pyrite samples was readily oxidized, loading the leachate with dissolved zinc to a much greater extent than the other samples. The pyrite study supports the possibility o f active sphalerite oxidation. verifiable  from  the  leachate  data  Pyrite or pyrrhotite oxidation was not  as the  leaching solution iron  and  sulfate  concentrations masked any evidence o f oxidation o f these minerals.  The presence o f lead, albeit in low concentrations, suggest that some dissolution o f galena occurred in the Tizapa and Brunswick pastes, both o f which contained the largest concentration o f lead in the paste prior to leaching.  Major  ion chemistry (Ca, K , M g , and Si) compared with the  solid phase  concentrations o f the samples suggests that, as in the water-leached environment, binder phases were actively dissolving. Calcium was most likely leached from the portlandite phase (Ca(OH)2), with the O H - ions consumed to neutralize the acidic ferric sulfate solution. Higher depletion rates o f calcium, potassium and magnesium occurred i n the ferric sulfate environment (compared to the water) indicating that the A R D solution was more aggressive than water.  119  5.5  SOLID PHASE CHEMISTRY OF LEACHED PASTE  Tables 5.4 to 5.7 contain scanned images o f the leached paste samples showing concentric altered layers.  The sample  sections were  cut through either  pucks  (rectangular) or cylinder samples (pie-shaped). The above mentioned tables also contain comparative chemical analyses o f each altered layer o f the samples.  The detailed  analytical reports are presented i n Appendix III.  The chemistry o f each layer is presented  as the difference i n concentration (in  percentage) with respect to the core o f the particular sample i n the flooded environment. This method o f analysis was devised to show only the statistically significant differences in concentrations between the altered layers o f a particular backfill mixture, taking into consideration the 5% margin o f analytical error reported by the laboratory.  The  comparative studies are based on the premise that the core o f the flooded samples have undergone the least amount o f alteration, a hypothesis verified by various leaching studies o f cemented mixtures (DeCeukelaire, 1991; Revertegat et al, 1992; Kosmatka et al, 1995; Casanova et al, 1996).  5.5.1 Tizapa Water-leached environments Table 5.4 shows a net loss o f sulfate and calcium in the outer layers o f both waterleached environments compared to the core o f the flooded sample. The loss o f calcium is greater in the 4 m m outer layer o f the cycled sample than i n the 0.75 m m outer layer o f  120  the flooded layer. Calcium is most likely depleted from the highly soluble portlandite phase (Ca(OH)2) o f cement and possibly also from some de-calcification o f the tobermorite phase (Ca3Si20s-3H20), as reported i n other concrete leaching experiments (Revertegat et al, 1992; Carde and Francois, 1997). The less depleted concentration o f calcium i n the flooded paste may indicate the precipitation o f a secondary calcium phase in the outer layer.  Sulfide levels are similar in all layers suggesting that sulfide oxidation is not occurring in this layer. The sulfate leached out o f the sample therefore came from the dissolution o f a soluble sulfate phase already present i n the paste, a similar observation extracted from the leachate geochemistry data.  Ferric sulfate-leached environment Considerable differences i n solid phase chemistry were observed between the first and second alteration layers o f the sample exposed to the ferric sulfate solution. outermost  The  layer (layer 1) showed an enriched sulfate concentration but depleted  concentrations o f total sulfur and sulfide-sulfur, calcium, magnesium, zinc and lead compared to the underlying layer 2 and to core concentrations o f the flooded sample. These results indicate that oxidation o f sphalerite and dissolution o f galena was probably occurring i n this area.  N o information is available to verify the oxidation o f the iron  sulfides although leachate geochemistry data suggests that it is minimal.  Layer 2 was  depleted i n sulfate and further depleted i n calcium and magnesium compared to the layer 1 and the core o f the flooded sample.  Solid phase concentrations i n the 3  almost all reverted to the values o f the  rd  layer had  core sample i n the flooded environment, 121  with the exception o f calcium and magnesium which still remained slightly depleted.  A progressive depletion o f cement occurred from the centre o f the sample towards the edges.  The greater loss o f cement i n the outer layer exposed to the ferric sulfate leach  solution allowed sulfide mineral oxidation to take place.  5.5.2  Brunswick  Water-leached environments Table 5.5 shows that, like the Tizapa water-leached samples, the outer layers o f both water-leached Brunswick samples underwent a net loss o f sulfate and calcium, with a more severe depletion occurring in the cycled environment. The dissolved calcium and sulfate likely originated from cement phases (portlandite and tobermorite) and preexisting sulfate phases respectively, as described for the Tizapa samples.  In addition,  similar sulfide concentrations and sulfur to metals ratios in all water-leached layers suggest that no sulfide oxidation occurred i n any water-leached sample.  Ferric sulfate-leached environment Sulfate concentrations were significantly depleted only i n the 2 sulfate-leached paste.  n d  layer o f the ferric  Precipitation of a secondary sulfate phase from the leachate  solution most likely replaced the lost sulfate due to dissolution o f the solid phase, explaining the unchanged sulfate concentrations i n the outermost layer o f the paste. The depleted sulfate concentrations, unchanged sulfide concentrations and stable sulfur to metal ratios suggest that, like i n the Tizapa case, no oxidation occurred i n layers 2 and 3 122  o f the ferric sulfate-leached paste. Sulfide mineral oxidation was also unlikely i n layer 1 as indicated by the similar sulfide concentrations between layer 1 o f the ferric sulfate environment and the core o f the flooded paste.  A l s o similar to Tizapa, calcium was  progressively less depleted in underlying layers 2 and 3, reflecting the decreasing dissolution o f portlandite (and de-calcification o f tobermorite) with depth.  5.5.3  Louvicourt  Water leached environments N o alteration layers were evident upon observation o f the water-leached samples (Table 5.6).  A n outer layer o f 1 m m was, therefore, arbitrarily chosen and analysed for  verification. complete  The thin layers extracted did not provide enough sample to allow for  elemental  analyses.  Sulfur  species  analyses  showed  depleted  sulfate  concentrations in the outer layers o f both environments with respect to the core o f the flooded sample, resulting from the dissolution o f pre-existing sulfate phases.  The  absence o f significant changes i n sulfide concentrations and i n total sulfur to sulfide ratios suggests that no oxidation o f sulfide minerals occurred i n these layers.  Ferric sulfate-leached environment T w o distinct layers o f alteration were observed on this sample.  However, only the  second layer provided enough material to carry out elemental analyses.  Similar to the  Tizapa ferric sulfate-leached sample, layer 2 o f the Louvicourt sample showed depleted concentrations o f sulfate, calcium, magnesium, zinc, copper and lead. Some oxidation o f  123  sphalerite along with dissolution o f galena and possibly some soluble copper phases may have occurred i n this layer. The similar sulfide concentration with respect to core values does not, however, support the occurrence o f significant sulfide mineral oxidation.  5.5.4 Francisco I. Madero Water leached environments Table 5.7 shows a net loss o f sulfate in layer 1 o f both water-leached environments, as well as a net gain i n copper i n layer 1 o f the flooded sample.  The elevated copper  concentration suggests a mobilisation o f copper from an internal area to the edges o f the sample where it may have precipitated as a more stable phase. The migration o f elements may have been facilitated by the greater porosity and lower binder proportion o f the F I M backfill mixture compared to the others.  Ferric sulfate-leached environment Layer 1 was characterized by depleted concentrations o f sulfate, calcium and magnesium and a considerably enriched concentration o f iron.  A s with the other ferric sulfate-  leached samples, the depletion o f sulfate and major ions may be explained by the dissolution o f pre-existing sulfate phases and cement phases i n that layer, namely portlandite and perhaps tobermorite.  The enriched iron may be explained by a deeper  penetration o f surficial ferric hydroxide precipitates observed on all samples i n this environment because o f lower binder content o f the F I M paste mixture.  124  03  "8  00  00  2  o  il 13 SB C  c5  CN  o  •a I  >  III •XT  I 50  03  O o  5. ss  o o  £ £  5-.  03 •T3  as  "O  r-4  SO  ©"  o -9  CN  c  CM  <D  O k > O  'I  -2 J  w T3  o c  P CD  o  03 C  CO  u  p -a  oo  a £ cu  I  e T3  e  T3  CM  ^  -rt  o  J  c<~>  O  c  CN  Si Si o cn co ©  i  S3  oo  CL, -o CL UJ oo  CO _>>  o  03  C  03  W  o  -  I  £  i i  D  a a  UL.  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CL IT)  CN  a 00  O  o oo  I oo# 03 CU  TJ  CN  O  la  00 CD  CN  S-  s  fi,  oo  98  ii  £  00  CD  OJ  oo oo  l : x  a GO  CD  S? fe  CD  —  93 >.  03 fe  O  U  CD  fe  -C 0-  3JO0 UI0i} 90U3J3JJip UOIJBJJU30UO0 %  5.5.5  Summary of Observations from the Solid Phase Chemistry  Water-leached pastes: •  N o s i g n i f i c a n t o x i d a t i o n o f sulfide m i n e r a l s o c c u r r e d i n a n y layer o f a l l w a t e r - l e a c h e d samples.  Sulfate c o n t a i n e d i n the paste w a s r e a d i l y leached out o f the b a c k f i l l f r o m a  p r e - e x i s t i n g sulfate phase present i n the m i x t u r e . I n a l l but the F I M samples, sulfate d e p l e t i o n w a s m o r e severe i n the c y c l e d e n v i r o n m e n t than the f l o o d e d e n v i r o n m e n t .  •  C a l c i u m w a s r e a d i l y leached out f r o m the outermost layers o f a l l samples a n a l y s e d , i n both  water-leached  environments.  Depletion was more  e n v i r o n m e n t f o r b o t h T i z a p a a n d B r u n s w i c k samples. Louvicourt.  severe  i n the c y c l e d  N o data w e r e a v a i l a b l e f o r  T h e less depleted amount o f c a l c i u m i n the f l o o d e d paste m a y i n d i c a t e  the p r e c i p i t a t i o n o f a secondary c a l c i u m phase i n the outer layer i n that e n v i r o n m e n t . C a l c i u m w a s l i k e l y lost f r o m the d i s s o l u t i o n o f the cement phases o f portlandite (Ca(OH)2) a n d p o s s i b l y f r o m d e c a l c i f i c a t i o n o f tobermorite ( C a 3 S i 2 0 s - 3 H 2 0 ) . T h e latter w a s also reported b y Revertegat et al, (1992) a n d C a r d e a n d F r a n c o i s (1997) i n s i m i l a r experiments.  •  T h e m a x i m u m t h i c k n e s s o f the alteration layers i n b o t h water-leached  environments  w a s larger f o r paste m i x t u r e s c o n t a i n i n g the highest percentage o f p y r i t e a n d largest p r o p o r t i o n o f binder.  Indeed, the m a x i m u m thickness o f the alteration layer f o r  L o u v i c o u r t a n d F I M pastes w a s 1 m m i n 20 w e e k s , c o m p a r e d to 6 m m f o r T i z a p a a n d B r u n s w i c k mixtures.  129  Ferric sulfate-leached pastes: •  Similar or increased sulfate levels on the edges o f the samples suggest that one or both o f the following were occurring in the samples: 1) penetration and precipitation o f sulfate from the solution as the solution came in contact with the pore solution o f the paste and/or 2) active oxidation o f sulfide minerals (sphalerite and possibly some iron sulfides) producing sulfate and releasing metal ions.  Both cases involve the  advancement o f a ferric sulfate solution front into the paste.  •  Greater net losses o f calcium and magnesium occurred i n the ferric sulfate environment compared to the water-leached environments, indicating a more intensive dissolution o f portlandite and perhaps tobermorite (through decalcification). Portlandite, a hydroxide mineral, is considerably more soluble in an  acidic  environment than in the a near-neutral water-leached environment, the hydroxyl ion being consumed to buffer the acidic solution.  •  The alteration layers o f samples leached in ferric sulfate solution penetrated deeper than i n the  water  environments.  In addition, similar to the  water-leached  environments, mixtures with higher pyrite and high binder content had deeper penetrating alteration: 8 m m and 10 m m in 20 weeks for Tizapa and Brunswick respectively, compared to 2.75 m m and 1.8 m m for Louvicourt and for F I M pastes.  •  In general, as a result o f the more intensive dissolution o f the binder phases, the readily oxidized sulfide minerals such as sphalerite present in the outermost layers were most likely oxidized by the acidic (unbuffered) ferric sulfate solution. The  130  oxidation o f sulfide minerals did not, however, occur in all samples i n that environment. The net loss o f lead in the altered layers o f some samples indicated that galena was probably dissolved or oxidized from those areas.  5.6  BUFFERING CAPACITY AND ACID-BASE ACCOUNTING  Paste p H measurements and A B A results are presented in Tables 5.8 a to d.  Paste p H can give qualitative information on the immediate buffering capacity or acid production o f soil or rocks. In this study, alkaline paste p H suggested the presence o f a readily available source o f neutralizing minerals and little or no acidity. The paste p H o f unleached Louvicourt (8.4), Brunswick (10.3) and Tizapa (11.1) samples were high. These values corresponded to a binder content o f 4.5% for Louvicourt (0.9% cement and 3.6 % slag) and to cement contents o f 5% for Brunswick and 6.2% for Tizapa respectively. For comparison, the paste p H values o f these tailings ranged from 5.0 to 5.8. F I M paste sample had a paste p H o f 8.9, a smaller increase from the tailings p H o f 7.3.  These values reflect the smaller proportion o f cement (3%) o f the F I M paste  mixture. Paste p H values corresponded well to the cement content o f the paste mixtures.  5.6.1  Evolution of Acid Producing and Acid Neutralizing Potentials:  A s a preliminary step to acid-base accounting ( A B A ) , fizz ratings ranged from no fizz to moderate fizz for both the tailings and unleached paste mixtures o f Tizapa, Brunswick and Louvicourt.  F I M tailings and unleached paste produced a strong fizz.  Ratios o f  neutralizing potential (NP) to acid production potential ( A P ) or N P R o f all tailings, 131  unleached pastes and o f each altered layer o f the 20 week-leached pastes also figure i n Tables 5.8 a to d.  A l l tailings possessed N P R values below 0.5.  The addition o f cement provided some  neutralizing capacity to all the mixtures. However, N P R values all remained below 1, hence a strong potential for acid generation remained.  In all cases, the proportion o f  cement was not sufficient to change the classification o f the backfill to a potentially non acid-generating residue.  In all cases, as leaching progressed, A P values remained statistically similar (within the 5 % margin o f analytical error) while N P was being depleted. N P depletion was attributed to the dissolution o f portlandite, agreeing with the leachate and solid phase chemistry data. In addition, N P depletion was more severe in the cycle-leached environment than i n the flooded environment.  Resulting N P R values were lower for the outer layers o f  samples in the cycle-leached environments than i n the flooded environments.  Ferric sulfate or artificial A R D solution leaching depleted N P values to a greater extent than water-environments. 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'13 co  ©I  o CO  '« I©  CN  136  5.7  SCANNING ELECTRON MICROSCOPE (SEM) OBSERVATIONS  The sections shown i n Tables 5.4 to 5.7 discussed earlier, were observed under S E M to identify the mineralogical changes that occurred as leaching progressed.  The layer  identification system and thicknesses described in these tables was used to guide the S E M observations.  Energy dispersive x-ray analyses ( E D X ) were carried out on many  different mineral surfaces.  The fragility o f the surfaces did not allow the sections to be  polished flat since they would easily crumble. A l l E D X analyses reported i n this section are therefore qualitative and used i n conjunction with S E M micrographs to identify the general composition o f mineral phases. Plates 5.1 to 5.8 show the typical morphology and qualitative composition ( E D X scans) o f tobermorite, portlandite, gypsum and ettringite encountered i n this study.  5.7.1  Tizapa  The Tizapa paste contained the largest proportion o f cement (6.2%) making possible the observation o f many distinct phases o f hydrated cement.  Tizapa samples displayed the  greatest differences in cement mineralogy between the altered layers.  i) Flooded Water Environment Layer 1:  0 - 0 . 8 mm  A n extensive layer o f densely packed euhedral portlandite crystals covered the surface o f the sample (Plate 5.9). Portlandite was also abundant down to a depth o f 0.8 to 1 m m ,  137  along with a well hydrated but thin layer of tobermorite covering the tailings grains (Plates 5.10 and 5.11).  The local abundance o f euhedral portlandite in the outer layer  suggests that it was most likely secondary i n nature.  Below layer 1:  0.8-1.5 mm  A transition zone o f poorly developed tobermorite was encountered at this level, where tailings grains were not so well cemented compared to the overlying and underlying layers (Plate 5.12). Little portlandite was encountered i n this layer.  Core:  < 1.5 m m  The core o f the sample was characterised by well-developed and abundant tobermorite gel covering the tailings particles (Plate 5.13).  Acicular ettringite was also visible  covering partially hydrated tricalcium aluminate (C3A or Ca3Ali06) phases (Plate 5.14). Ettringite is normally one o f the first minerals to crystalize upon hydration o f the cement. Its purpose is to slow down the otherwise very rapid hydration o f C3A which causes "flash setting" o f the concrete (Taylor, 1997; Kosmatka et al, 1995). The occurrence o f ettringite on C3A indicates that it was a primary mineral rather than a product o f secondary, expansive mineral growth reactions within the sample.  In addition, the occurrence o f ettringite in this area indicated an abundance o f C a  2 +  and  SO4 " ions i n solution, as ettringite is formed according to the reaction (Vernet, 1994): 2  32 H 0 + C a A l 0 (C3A) 2  3  2  + 3 C a + 3 S 0 " -> (or dissolved gypsum) 2 +  6  2  4  C a ^ ^ M O H ^ ^ H z O (ettringite)  (5.1)  In l o w sulfate cement mixtures, the gypsum phase present i n the dry cement is eventually 138  exhausted (in the first hours o f hydration) and the ettringite proceeds to react with the excess C3A to produce monosulfoaluminate (Ca4Ai206(S04)45H20; a wavy, leaf-like morphology - not observed in this sample) according to:  2 Ca Al 0 (C A) 3  2  3  6  + Ca6Al2(S04)3(OH)i -26H 0 -> 3 C a 4 A l 0 ( S 0 4 ) - 1 5 H 0 + 17 H 0 (ettringite) (monosulfoaluminate) 2  2  2  6  2  2  (5.2)  Monosulfoaluminate forms away from the C3A surface, leaving the mineral surface free to continue to hydrate, consuming excess portlandite and providing some additional strength to the mixture (see reaction (3) i n Table 1.1, Chapter 1). In this case, however, sulfate continues to be present i n the pore water i n equilibrium with the primary ettringite, preventing further hydration o f C3A and consumption o f portlandite.  Both  factors impede the normal development o f strength.  ii) Cycled Water Environment Layer 1:  0 - 4 mm  The outer layer o f the cycle-leached sample was characterized by poorly developed tobermorite gel i n the exterior portion o f the layer (Plate 5.15), only partially covering tailings grains. Ettringite needles and abundant gypsum were also visible i n the upper portion o f this layer (Plate 5.16 and 5.17).  In the deeper portion (>2mm), the cement  phases were well developed and distributed. Plate 5.18 suggests that the cement phase was not very strong as many tailings grains were removed during the sample preparation process, leaving empty grain-shaped cavities.  139  Core: > 4 m m A s with the flooded core, this area o f the cycle-leached paste was characterised by abundant ettringite covering phases o f tobermorite and unhydrated C3A (Plate 5.19). Tobermorite cover was generally extensive and well developed (Plate 5.20).  iii) Ferric Sulfate Environment Layer 1:  0 - 2 mm  This layer o f the ferric sulfate-leached sample was generally porous with sparse hydrated cement phases (Plate 5.21). Most pyrite grains (as well as other tailings grains) showed only traces o f cement cover (Plate 5.22). Some isolated areas did have well-developed tobermorite (Plate 5.23).  Oolitic masses o f iron sulfate were exclusive to this layer  (Plates 5.24 and 5.25). This phase could either have been precipitated directly from the ferric sulfate solution i n contact with the pore water o f the sample or could have precipitated following sulfide mineral oxidation i n that area.  Layer 2:  2 - 6 mm  The hydrated cement cover was more uniform and better developed in this layer, offering better protective cover to all grains such as pyrite (Plate 5.26). Under S E M , the presence o f amorphous hydrated cement cover masked the presence o f amorphous iron phases such as hydrated iron oxide or iron oxyhydroxide. The rust colour o f this layer suggests, however, the presence o f these phases.  140  Layer 3:  6 - 8 mm  This was an apparent transition zone between layer 2 and the core o f the sample, where the mineralogy was similar to that o f the core o f the flooded sample.  5.7.2  Brunswick  i) Flooded Water-Leached Environment Layer 1:  0 - 3 mm  Similar to the Tizapa sample, portlandite was also present i n abundance i n this layer, along with fairly well developed tobermorite covering the tailings grains (Plates 5.27 and 5.28). The cement appeared to be less resistant than i n deeper layers as empty grain cavities were observed (Plate 5.29). The cavities were, however, less numerous than i n the ferric sulfate-leached sample.  Layer2:  3-6mm  This layer possessed well-developed and abundant tobermorite covering the tailings particles (Plate 5.30).  N o portlandite was visible.  A n acicular phase, most likely  ettringite, was also observed from this depth down to the core o f the sample.  141  ii) Cycled Water-Leached Environment Layer 1:  0 - 3 mm  Tobermorite appeared to be well developed, covering the tailings grains fairly well (Plate 5.31). The paste was porous but did not show any empty grain cavities (Plate 5.32). N o gypsum or acicular ettringite were observed at any depth in this sample.  Layer 2: 3 - 6 m m N o obvious mineralogical differences were observed under S E M between this layer and Layer 1.  iii) Ferric Sulfate-Leached Environment Layer 1:  0 - 1 mm  Macroscopically, this layer showed an outer rim o f iron oxyhydroxide precipitated from the leaching solution (refer to Figure 5.36). Under S E M , a 0.5 m m crust was visible with vertical fractures extending 1.5 to 2 m m down (Plate 5.34). M a n y grain cavities were visible suggesting that the strength o f the hydrated cement was relatively poor (Plates 5.34 and 5.35). Acicular minerals were also present i n this area but were too small to obtain a reliable E D X identification (Plate 5.36). The morphology o f the phase suggested ettringite.  142  Layer 2:  1 - 3 mm  The hydrated cement phases appeared to be more resistant at this depth, showing a smoother sample surface with less grain cavities (Plate 5.37).  Masses o f oolitic iron  sulfate phases were abundant in this layer, possibly a type o f jarosite (Plate 5.38). Similar to the Tizapa sample, this iron sulfate phase either was precipitated directly from the ferric sulfate solution in contact with the pore water o f the paste or was a product o f sulfide mineral oxidation in the area. Euhedral gypsum crystals were encountered inside sample pores (Plate 5.39). N o ettringite needles were identified i n this layer.  Layer 3:  3 - 1 0 mm  Macroscopically, a gradient in colour was visible from 3mm to 10 m m and a very pale line marked the 10mm depth. This fine layer was not noticeable under S E M and layer 3 appeared as a transition zone between layer 2 and the core o f the sample.  In the upper  and middle portion o f layer 3, cement binder strength appeared to be greater (Plate 5.40), the topographic lows were related to the porosity o f the paste rather than to empty grain cavities. Tailings grains were also better covered by cement (Plate 5.41). N o ettringite needles or gypsum was positively identified in this layer.  5.7.3  Louvicourt  The high slag proportion o f the Louvicourt paste gave the binder a different aspect from the O P C binder used in the other mixtures.  The hydrated slag was identified by its  morphology and occurrence, appearing as both masses o f subangular crystals (Plate 5.42)  143  and delicate, semi-dendritic crystals covering the tailings grains (Plate 5.43).  Both  morphologies are typical o f well-hydrated slag (Taylor, 1997). The composition o f the tobermorite developed from slag is slightly different from that from Portland cement, generally having a lower concentration o f C a but higher A l and M g . Plate 5.44 presents a typical E D X spectrograph o f slag tobermorite (from the flooded paste). Plate 5.45 shows the general aspect o f the paste.  i) Flooded Water-Leached Environment Layer 1:  0 - 1 mm  Contrary to Brunswick and Tizapa pastes, no portlandite was observed i n any area o f the flooded paste sample. Tobermorite was well developed, covering the tailings grain to a greater extent the previous two samples i n the same area. The tobermorite content was slightly lower i n some areas o f the outer layer (Plate 5.46) compared to deeper areas where tobermorite was more abundant and evenly distributed (Plate 5.47). N o ettringite, monosulfoaluminate or gypsum was identified i n any area o f the flooded paste.  ii) Cycled Water-Leached Environment Layer 1:  0 - 1 mm  A greater depletion o f tobermorite was observed down to 500um from the surface (Plate 5.48) compared to the flooded environment. Binder cover o f tailings grains i n this area was still greater than Tizapa and Brunswick at that depth (Plate 5.49). B e l o w 500 um, tobermorite abundance and distribution was similar to that o f the flooded paste (Plate  144  5.50).  Secondary gypsum growth occurred in the deeper portion o f the paste (Plate 5.51) which could have been a source o f internal stress within the sample. N o gypsum was found closer to the surface. N o portlandite, ettringite or monosulfo-aluminate was identified i n any area o f the cycled sample.  A vertical fracture traversing the height o f puck was observed under S E M (not visible to the naked eye) (Plate 5.52).  Dendritic and platy crystal forms o f Tobermorite were  observed lining the fracture walls (Plate 5.53) indicating that the fracture was formed during leaching rather than during later-stage sample preparation.  Secondary fracturing  reveals the occurrence o f internal stress after only 20 weeks o f leaching i n water.  iii) Ferric Sulfate-Leached Environment Layer 1:  0 - 0.75 mm  Tobermorite in the outermost layer was most depleted i n the ferric sulfate environment (Plate 5.54), however, not so much as i n samples containing O P C binder i n a similar environment. Its morphology was slightly altered compared to deeper areas, the masses were more rounded and amorphous (Plate 5.55), possibly resulting from an altered chemistry due to leaching o f some elements, as described by Taylor (1997). Tobermorite content and increased again at 1 m m (Plate 5.56), regaining its more typical morphology.  A partly separated crust was present on the surface o f the puck, visible under S E M (Plate 5.57).  N o obvious secondary minerals were observed on the walls o f the fracture  145  suggesting that fracturing may have occurred during sample preparation. The constant thickness o f the fracture indicates, however, that it developed along an existing plane o f weakness.  Layer2:  1-3 mm  L i k e i n Tizapa and Brunswick, some iron sulfate precipitates were observed i n this layer. The morphology o f these precipitates differed from the other samples. They occurred as larger botryoidal masses (20 um - Plate 5.58) rather than the loosely bound oolitic masses occurring i n Brunswick and Tizapa. E D X analysis suggests that these precipitates may be a form o f jarosite (Plate 5.59) or some other hydrated iron sulfate phase.  No  portlandite, gypsum or monosulfoaluminate phases were identified i n this sample at any depth.  L o n g , rectangular-shaped crystals (Plate 5.60) were analysed but the E D X  spectrogram was not sufficiently precise to provide a definite identification.  The  morphology was similar to that o f ettringite. Tobermorite gel was well developed (Plate 5.61) although slightly less abundant than deeper towards the core o f the sample.  B e l o w 3mm depth, tobermorite was abundant, well developed and w e l l distributed, similar to that encountered in the core o f the flooded sample.  N o gypsum or  monosulfoaluminate were observed although the rectangular ettringite-shaped crystals were still present.  5.7.4 Francisco I. Madero F I M samples contained the lowest proportion o f cement (3%) o f all paste mixtures.  146  Macroscopically, neither the  flooded nor the  cycled water-leached  environments  displayed visible alteration layers after 20 weeks o f leaching. S E M observations revealed few differences between superficial layers and the cores o f the samples.  i) Flooded Water-Leached Environment Layer 1:  0 - 1 mm  The outer layer o f the flooded paste showed a good distribution o f well-developed tobermorite (Plate 5.62), similar to Brunswick tobermorite distribution although F I M contains 60% o f the cement content. N o portlandite or ettringite were identified i n this layer or deeper.  Core:  > 1 mm  A gradual decline i n tobermorite content and degree o f development was visible between - 7 5 0 urn and 1mm, below which the tobermorite appeared as fibrous and scattered, incompletely covering the tailings grains (Plate 5.63). This morphology is suggested i n the literature to be typical o f decalcified tobermorite (Taylor, 1997). These phases were, unfortunately, too scattered and thin to provide reliable E D X analyses to measure the calcium to silicon ratio.  ii) Cycled Water-Leached Environment Layer 1:  0 - 1 mm  Tobermorite was moderately well developed but less abundant than in the flooded  147  sample.  The cement phases in the outer layer appeared as masses o f small individual  crystals (<lum size) gathered in bunches rather than intergrown (Plate 5.64).  This  morphology was encountered throughout the cycled sample and was distinct from the morphology encountered in any other mixture, where it appeared as amorphous masses. These cement phases could not be accurately identified, as the crystals were too small and not sufficiently abundant to obtain reliable E D X analyses. N o secondary minerals such as portlandite, ettringite or gypsum were observed at any depth i n this sample.  iii) Ferric Sulfate-Leached Environment Layer 1:  0 - 1 mm  Plates 5.65 and 5.66 show the porous but homogeneous aspect o f the outer edge o f the sample.  The small proportion o f cement i n the mixture appeared evenly distributed,  showing better tailings grain cover than Tizapa in a similar environment (Plate 5.67). Wavy, leaf-like minerals, most likely monosulfoaluminate, were discernible deeper, at 1.8 m m (Plates 5.68). Acicular ettringite could not be positively identified i n this area or anywhere i n the sample. The presence o f monosulfoaluminate over ettringite suggests that relatively l o w concentrations o f aqueous C a  2 +  and SO4 " ions were present i n the pore 2  solution i n the area - see reaction (5.2). This may be a consequence o f the l o w sulfate levels present i n the F I M tailings, the lowest o f all the samples. Cement cover o f tailings (pyrite) grains appeared to be slightly improved at 4 m m depth compared to the surface (Plate 5.69).  148  Tizapa - Flooded-Leached Paste Samples  Plate 5.9 Portlandite cover on sample surface  Plate 5.10 Bottom of layer 1  Plate 5.11 Bottom of layer 1  Plate 5.12 Poorly developped Tb, below layer 1  Plate 5.13 Core of sample  Plate 5.14 Primary ettringite on incompletely hydrated cement grain (Tea), core of sample  Pd = portlandite; Tb = tobermorite gel (amorphous); Py = pyrite; + = area o f E D X analysis 151  Tizapa - Cycle-leach Paste Samples  Plate 5.15 Poorly developped Tb, layer 1  Plate 5.16 Ettringite, upper layer 1  Plate 5.17 Secondary gypsum, upper layer 1  Plate 5.18 General paste aspect, layer 1  Plate 5.19 Ettringite in core of sample  Plate 5.20 Good tobermorite development at core of sample  Et = ettringite; Tea = tricalcium aluminate (unhydrated cement grain); G y = gypsum  152  Tizapa - F e ( S 0 ) Solution-leached Paste Samples 2  4  3  Plate 5.21 Porous layer 1  Plate 5.22 Depleted tobermorite, layer 1  Plate 5.23 Area of good tobermorite, layer 1  Plate 5.24 Layer 1, iron sulfate precipitate  Fe 0 JlFe  I  A l | ' J  1  I  0  2.0  T  —  4.0  1 ~T  6.0  Plate 5.25 EDX of iron sulfate precipitate  153  1  '  f B.O  1 10.C  Plate 5.26 Good tobermorite cover, layer 2  Brunswick - Flooded-leach Paste Samples  Plate 5.29 Less resistant tobermorite, layer 1 Plate 5.30 Abundant tobermorite, layer 2  154  Brunswick - Cycle-leach Paste Samples  Plate 5.31 Good tobermorite cover, layer 1  Plate 5.32 General aspect of paste, layer 1  Brunswick - F e ( S 0 ) Solution-leach Paste Samples 2  4  3  Plate 5.33 Paste surface  Plate 5.34 General aspect of paste, layer 1  Plate 5.35 Poor tobermorite cover, layer 1  Plate 5.36 Possible 2 ettringite, layer 1 ry  155  Brunswick - Fe (S04)3 Solution-leached Paste Samples 2  Plate 5.37 General aspect of paste, layer 2  Plate 5.38 Fe-sulfate precipitate, layer 2  Plate 5.39 Inside large pore, layer 2  Plate 5.40 General aspect of paste, layer  Plate 5.41 Layer 3  156  Louvicourt - Aspects o f Hydrated Slag-OPC binder  Plate 5.42 Massive tobermorite  Plate 5.43 Dendritic tobermorite  II hi  t  1  r "~  0.0  ——i—  2.0  s  c* —  • i- •  A •  4.0  Plate 5.44 EDX of tobermorite developed from slag  Plate 5.45 General aspect of paste mixture  157  i  i  i  6.0  8.0  10.0  Louvicourt - Flooded-leached Paste Samples  Plate 5.46 Less tobermorite in upper layer 1  158  Plate 5.47 More abundant tobermorite in lower layer 1  Louvicourt - Cycle-leached Paste Samples  Plate 5.48 Poorly developed Tb, upper layer 1  Plate 5.50 More developed Tb, lower layer 1  Plate 5.49 Upper layer 1  Plate 5.51 Secondary gypsum, lower layer 1  Plate 5.52 Tobermorite-Iined fracture in paste Plate 5.53 Close-up view of tobermorite  159  Louvicourt - Fe^SCVb Solution-leached Paste Samples  Plate 5.54 Depleted tobermorite, layer 1  Plate 5.55 Masses of tobermorite at upper layer 1  Plate 5.56 Increased Tb content, below layer 1 Plate 5.57 Fracture developed possibly along existing plane of weakness  160  Louvicourt -  Fe (S0 )3 2  4  Solution-leached Paste Samples  Plate 5.58 botryoidal mass of iron sulfate precipitate  e  Ca  >C—X-l—IT^  0.0  I  *  1  1  |  2.0  1—  ••  4.0  1  i '  Fe  L - A'  6.0  8.0  1 10.0  Plate 5.59 EDX of iron sulfate precipitate (type of jarosite?)  Plate 5.60 Possibly ettringite, layer 2  Plate 5.61 Well developed tobermorite, lower layer 2  161  Francisco I. Madero - Flooded-leached Paste Samples  Plate 5.62 Layer 1,floodedsample  Plate 5.63 Core offloodedsample  Francisco I. Madero - Cycle-leached Paste Samples  Plate 5.64 Layer 1, hydrated cement phase (may be Et, Pd, Tb and/or Gy)  162  Francisco I. Madero - F e ( S 0 ) Solution-leached Paste Samples 2  Plate 5.67 Layer 1  4  3  Plate 5.68 Monosulfoaluminate (?), below layer 1  Plate 5.69 Core of sample M s a = monosulfoaluminate  163  5.7.5  Summary of Observations from Scanning Electron Microscopy  Distinct secondary precipitates were observed between mixtures containing only O P C binder and the Louvicourt mixture containing 80% slag i n its binder.  In general, the  tobermorite developed from the slag-cement was observed to be more resistant to leaching i n all environments.  Flooded Water Environment: •  In O P C binder mixtures, Euhedral portlandite crystals were present i n relative abundance i n the outer edges o f flooded water-leached samples exclusively.  The  abundance o f portlandite appeared to be proportional to the cement content o f the sample. The longer contact time between the water and the sample in the flooded environment most likely allowed for the pore water to become supersaturated with calcium hydroxide and re-precipitate this phase. These observations agree with the solid phase and leachate chemistry data.  •  Tobermorite, the main binding agent o f cement, was present i n smaller amounts i n the outer edges o f the samples compared to the cores.  The amount o f depletion was  much less i n the slag-cement Louvicourt mixture. In all cases and all depths, the tobermorite that was present was evenly distributed and well developed.  •  Ettringite was easily discernible in the core o f samples that contained the largest amount o f O P C binder (Tizapa and Brunswick).  This suggests that dissolved C a  2 +  and SO4 " were abundant in the pore solution in the core o f these samples but were 2  164  leached out on the edges.  The concentration gradient would drive a gradual and  continued depletion o f these ions from deeper areas within the paste.  •  N o secondary, expansive minerals such as ettringite, gypsum or monosulfoaluminate were observed i n the slag-cement binder mixture.  This attests to the increased  resistance to water-leaching and sulfate attack o f this binder mixture.  Cycled water environment: •  N o portlandite was found in the outer edges o f any cycle-leached sample. Supersaturation o f portlandite was most likely inhibited i n the cycled environment as O H - ions were consumed to neutralize the leaching solution, and C a  2 +  was dissolved  and flushed out o f the system before equilibrium could be reached.  •  N o portlandite was observed in the slag-binder mixture as expected, as slag generally generates little portlandite upon hydration, another feature o f sulfate  resistance  (Taylor, 1997).  •  Tobermorite was depleted i n the outer edges o f the cycled water-leached pastes relative to the sample cores, the extent o f which appears to be greater than i n the flooded environments for both the O P C and slag-cement binder mixtures.  •  Presence or greater abundance o f grain-shaped voids in the outer edges o f the O P C binder pastes suggests a decreased cement binding strength in these areas relative to the core. The loss i n binding strength may be caused by the lower amount o f cement and/or by a degradation (i.e. decalcification) o f the cement phase in this area relative  165  to the core.  •  Although tobermorite was depleted in the outer layer o f the slag-cement mixture, no apparent loss o f binder strength or empty grain cavities were observed.  •  Secondary, euhedral gypsum crystals were observed i n the outer layers o f the Tizapa sample along with acicular ettringite.  Secondary gypsum was also observed i n the  Louvicourt paste below 4 m m depth. These minerals indicate that a source o f internal stress created by the pressure o f crystal growth is present within the backfill.  Ferric sulfate environment: •  Tobermorite was depleted i n the outer edges o f the ferric sulfate-leached pastes, to a similar extent than in the cycled environments, for mixtures o f both binder types. The level o f depletion was much lower in the Louvicourt sample than i n the others.  •  Oolitic or botryoidal. masses o f iron sulfate precipitates (possibly jarosite) were encountered in the 2  n d  layer o f alteration for all paste mixtures. This location may  indicate the penetration front o f the ferric sulfate leach solution at the interface with the high p H pore solution o f the samples.  •  Phases o f ettringite, monosulfoaluminate and/or gypsum were present at various 2"T"  depths i n the ferric sulfate-leached samples, indicating that dissolved C a ions were present in sufficient quantity in the pore solution.  2  and SO4 "  These ions were  therefore not completely depleted from the pore solution i n any area o f the ferric sulfate-leached samples. 166  5.8  COMPRESSIVE STRENGTH MEASUREMENTS  Unconfined compressive strength (ucs) measurements were carried out on unleached paste samples and on the 5, 10 and 20-week leached samples o f all 3 environments. The results shown on Figures 5.48 through 5.51 should be interpreted in a qualitative manner since only one sample was tested per environment, per leaching period.  Ucs requirements by each respective mine were achieved for all paste mixtures and surpassed for the Tizapa mixture (Table 5.9). appeared  to remain constant  environment.  as  Brunswick, Louvicourt and F I M ucs  leaching progresses,  regardless  o f the  leaching  A small decrease in ucs was observed in the Tizapa paste i n all  environments as leaching progressed. This loss may not be statistically significant.  Alteration o f the cement phases upon leaching was not only expected, from the literature review, but was also observed under S E M . The layers where alteration occurred were probably too thin after 20 weeks o f leaching to influence the compressive strength o f the entire sample.  Unconfined compressive strength o f any material is a function o f the surface area o f the sample subjected to compression according to the formula:  (5.3)  167  where F is the force applied to the sample by the compressor, i n kg, (multiplied by gravity) and A is the area o f the sample i n contact with the piston. Postulating that the loss o f cementing material seen under S E M resulted in a loss o f strength i n the oxidized layers, then the unaltered core o f the samples should possess higher strength than the unleached samples. Although no measurements were made to verify this, it is possible that the cores o f the leached samples gained strength over the leaching period because o f the longer curing time o f the samples i n an aqueous environment. This would suggest that uniaxial compressive strength may not constitute an appropriate measuring tool to characterize the strength o f weathered backfill samples.  168  o w  CO  o  OJ T3  o  >f  d  co o  CO  00  d  d  •^r CM co o d d  TJ  a>  X3  •sr  o o  o  t-  d  d  CN  o co c  CN  3  o  d  CO  o  o  co  CD  d CO  3 O O "> 3 O  co •o  d  oo d  d d  —I o  CD CO  ml  m co • CO o >. o N CNJ o CO  a.  TJ CD  TJ  co  N~ CN  I  a>  o o  csi  u_  c cu E cu cr cu  a:  TJ  0 o ro  JZ  c  CO CD CD  5  LO  T-  169  Flooded Cycled Fe2(S04)3 soln  0  5  10  20  Weeks of leaching  Figure 5.48 Tizapa Paste - UCS with Leaching Time  -Flooded -Cycled -Fe2(S04)3 solnl  5  10  20  Weeks of leaching  Figure 5.49 Brunswick Paste - UCS with Leaching Time  -Flooded -Cycled -Fe2(S04)3 soln  0  5  20  10 Weeks of leaching  Figure 5.50 Louvicourt Paste - UCS with Leaching Time  170  2  Weeks of leaching  Figure 5.51 Francisco I. Madero Paste - UCS with Leaching Time  171  6  DISCUSSION  6.1  PYRITE REACTIVITY STUDY  The mineral surface characterisation technique o f cyclic voltamperometry provided a measure o f the combined effects o f the different mineral characteristics on the reactivity of pyrite, such as chemistry, crystal morphology, stoichiometry and presence o f other sulfide phases associated with pyrite (termed mineral impurities).  This method o f  analysis established that, initially, Huckleberry pyrite was the most reactive o f all samples studied, most likely due to its relatively low concentration o f separate sulfide phases. Conversely, Louvicourt-2 and Tizapa pyrites were the least reactive because o f the considerable amount o f galena and sphalerite texturally associated with pyrite i n the samples.  This study suggested that the presence o f these mineral phases i n the pyritic  samples effectively decreased the reactivity o f pyrite to a greater extent than either stoichiometric proportions o f Fe:S, chemical impurities i n the crystal lattice, crystallinity or morphology o f pyrite crystals could have increased the reactivity o f pyrite.  The following sections discuss the effects o f the various mineral characteristics observed in this study, on the reactivity o f pyrites.  6.1.1  Higher  Effect of Precipitate Coatings on the Passivation of Pyrite: Huckleberry and Louvicourt-1 Samples dissolved iron  and  sulfate  concentrations  in the  leachate  supported  the  voltamperometric data suggesting that the unleached Huckleberry pyrite was the most 172  reactive o f all samples analyzed. The relatively high reactivity o f Huckleberry pyrites may, i n part, be caused by the greater amount o f fine size particles i n the sample. Grain size is not believed to be the only cause o f Huckleberry reactivity since the finer grain size Brunswick sample possessed the lowest reactivity after 20 weeks o f leaching. A s leaching progressed, however, both water and waste rock solution-leached Huckleberry pyrites were rapidly oxidized and became less reactive.  The amount o f iron oxyhydroxide precipitates on pyrite surfaces that were visible under S E M formed a very discontinuous layer even after 20 weeks o f leaching.  These  precipitates alone may not account for the decreasing reactivity exhibited i n the voltamperometric study. The visible precipitates may be however, local accumulations o f a more extensive F e O O H precipitates and/or polysulfide layers not visible under S E M . Mineral coatings o f F e O O H on oxidized pyrite were observed by many researchers. 1  Nicholson and others (1990) measured a continuous a-Fe203 (maghemite) precipitate layer on pyrite o f an average o f 0.6 urn thickness after 400 days o f leaching. In that study, the chemistry o f the precipitate layer was believed to be a dehydrated form o f yFeOOH  (lepidocrocite) transformed  by the  high vacuum o f the  auger  electron  spectroscopy ( A E S ) instrument, according to:  2 FeO(OH) -> F e 0 2  3  + H 0 2  (6.1)  The presence o f such thin oxidation rims on pyrite were also documented by Alpers and others (1994) and Jambor (1994). M a n y other researchers also documented the formation of iron oxide, iron oxyhydroxide, and/or sulfur, sulfoxianion or polysulfide layers on the  173  pyrite surface following electrode oxidation, all o f which effectively decreased  the  leachability (or reactivity) o f the underlying pyrite (Peters, 1984; Buckley et al., 1988; Ahmed, 1991; Z h u et al., 1992; Ahmed and Giziewicz, 1992). In addition, Moses and Herman (1991) proposed that accumulation o f Fe  on pyrite surfaces i n early stages o f  leaching effectively decreased the rate o f pyrite oxidation and that F e O O H deposits on the pyrite surfaces i n slowly-stirred, long-term leaching experiments were the cause o f decrease pyrite oxidation rates.  Although no specific information is available on the composition o f the oxidized layers observed by S E M on pyrite grains i n many samples, literature indicates that thin F e O O H , sulfur or polysulfide layers could be responsible for the increasingly efficient passivation of Huckleberry pyrites or any o f the other pyrites studied.  Louvicourt-1 and -2 samples came from the same mineral deposit but exhibited large differences in their chemical and stoichiometric composition as well as i n their oxidation behaviour.  The  unleached  Louvicourt-1  pyrite  sample  showed  a  similar  voltamperometric response to the Huckleberry sample and, likewise, a similar evolution of reactivity with increasing leaching-time.  The more abundant, amorphous  iron  precipitates covering the surfaces o f Louvicourt-1 pyrites did not appear however, to result i n a greater loss o f reactivity. It is probable that sulfur or polysulfide layers could have been more instrumental in passivating the pyrite surfaces than the apparently discontinuous iron oxyhydroxide precipitate coatings.  174  6.1.2  Effect of Stoichiometry on Pyrite Reactivity: Louvicourt-1 and Louvicourt-2 Samples  Pyrite stoichiometry was not found a determinant factor i n the reactivity o f pyrite compared to the presence o f sphalerite. The relative depletion o f iron in the Louvicourt-1 pyrite structure consisted o f an acceptor defect that should have theoretically protected the pyrite from oxidation (Shuey, 1975; K w o n g , 1993) compared to Louvicourt-2 pyrite having a S:Fe ratio o f exactly 2.0. Similarly, although arsenic contents were similar for both Louvicourt samples, S E M observations showed that Louvicourt-1 arsenic was mainly  contained  i n separate arsenopyrite  phases whereas  principally occurred as lattice impurities within pyrite.  Louvicourt-2  arsenic  The presence o f arsenic  impurities i n the crystal lattice o f pyrite creates a donor defect, which should theoretically promote  the  oxidation  of  Louvicourt-2 pyrite  (Shuey,  1975;  Kwong,  1993).,  Voltamperometric analyses o f the unleached Louvicourt pyrites indicated, however, that Louvicourt-2 pyrite was the less reactive one.  The presence o f arsenic may have  influenced the way i n which the Louvicourt-2 pyrite sample oxidized, as a greater density of oxidation pits were visible on Louvicourt-2 pyrites compared to Louvicourt-1 after 10 and 20 weeks o f leaching.  In summary, for all the pyrites studied, the effects o f iron to sulfur ratios and the presence o f arsenic impurities i n the pyrite crystal lattice were overcome by an opposing, more determinant factor on the reactivity o f pyrite. This factor is believed to be the presence o f sphalerite and galena impurities i n the pyrite samples, which offered galvanic protection to pyrite.  175  6.1.3  Effect of Mineral Impurities on Pyrite Reactivity  6.1.3.1 Galvanic Protection by Sphalerite: Louvicourt-2 and Tizapa Samples Tizapa and Louvicourt-2 pyrite samples possessed similar mineralogies and showed a similar evolution o f pyrite reactivity.  Both were characterized by an initial gain i n  reactivity after 4 weeks o f leaching, followed by a decreasing reactivity.  This study  suggests that during the leaching period, sphalerite grains in contact with pyrite acted as an anode and were preferentially oxidized over pyrite.  The cathodic reaction could  involve the reduction o f oxygen, as described by M u r r and Metha (1983):  A n o d i c reaction: sphalerite oxidation ZnS -> Z n  2 +  + S° + 2e-  (6.2)  Cathodic reaction: reduction o f water A 0  l  + 2H + 2 e - ^ H 0  (6.3)  +  2  2  K w o n g and Lawrence (1994) also proposed that the formation o f galvanic couples affected  the  reactivity o f pyrite to a greater extent than mineralogical  defects,  stoichiometric imbalances, grain size, crystallographic orientation or bacterial activity. Results o f both this and the K w o n g and Lawrence studies are supported by the research of Metha and M u r r (1983) who measured the contribution o f galvanic interaction on the reactivity o f pyrite in an acidic solution o f 1 M  H2SO4.  They proposed that galvanic  interactions were responsible for the cathodic protection o f a sulfide mineral with higher rest potential (pyrite) in contact with a mineral o f lower rest potential (sphalerite,  176  chalcopyrite) which acts as the anode and oxidized preferentially.  The gain o f reactivity i n both Louvicourt-2 and Tizapa samples coincided with higher concentrations  o f zinc i n both sample  leachates,  following  reaction 6.2.  SEM  observations showed that i n both samples, when sphalerite and pyrite occurred on the same grain, sphalerite phases were extensively corroded compared to pyrite phases. Pyrite generally possessed smooth surfaces and sharp grain edges i n the 4  th  and 10  th  weeks for Tizapa and Louvicourt-2 respectively.  The subsequent passivation o f pyrite coincided with the decline i n leachate zinc concentrations, occurring around the 5  th  week (10  th  cycle) for Tizapa and the 10  th  week  (20 cycle) for Louvicourt-2. Indeed, i n later leaching cycles S E M observations showed th  sphalerite grains becoming precipitate-covered and corroded, suggesting that sphalerite could also have been passivated and were no longer available to oxidize and offer galvanic protection to pyrite. Sphalerite passivation may be caused by a covering o f the surface by elemental sulfur shown in reaction 6.2. pitting were observed i n the 10  th  Increases i n the intensity o f pyrite  week for Tizapa and 2 0  th  week for Louvicourt-2,  suggesting a more intensive oxidation o f pyrite following the decreased oxidation o f sphalerite.  The increasing supply o f partly oxidized sulfur and dissolved iron to the  leachate could then have promoted the formation o f iron oxyhydroxide and/or sulfurpolysulfide coatings on all grains, effectively decreasing pyrite reactivity. A decrease i n the reactivity o f pyrite was indeed observed electrochemically in the 10  th  and 2 0  th  week  o f leaching i n the Tizapa and Louvicourt-2 samples respectively.  The  absence  of  sphalerite  in  the  Louvicourt-1 sample 177  probably  enabled  that pyrite surface to oxidize from the first leaching cycle onward, allowing for an earlier production o f a passivating layer. Voltamperometric studies indicated that Louvicourt-1 pyrite was more reactive prior to leaching but was passivated at an early stage.  6.1.3.2 Effect of the Presence of Galena: Zimapan and Brunswick Samples The presence o f galena was detected on the Zimapan and Brunswick voltamograms by the occurrence o f a characteristic peak around 0 m V . In that way, a concentration o f galena as l o w as 1.5 % was electrochemically detected in the Zimapan pyrite sample. A more intense peak was observed i n the Brunswick sample that contained a higher galena content (9%). The amplitude o f peaks attributed to galena appeared to be proportional to the galena content o f the sample.  The fact that galena is highly susceptible to oxidation is well documented in the literature (Metha and Murr, 1983; Nicholson, 1994; Alpers et al, 1994; Jambor, 1994; Blowes et al., 1995). In this study, galena was also found to oxidize rapidly: only traces o f galena were present in the Brunswick sample after 10 weeks o f leaching. Galena was replaced with anglesite (PbSC^) either by direct replacement (i.e. i n Tizapa where anglesite is found covering skeletal galena) or by precipitation from solution (i.e. in Brunswick and Zimapan where anglesite is dispersed throughout the sample).  The l o w solubility o f  anglesite (Table 6.1) explains why dissolved lead readily precipitated as lead sulfate i n the sample effectively scavenging the dissolved lead from the leachate solutions. Unlike lead, dissolved zinc does not tend to precipitate as a secondary zinc hydroxide or sulfate but tends to remain i n the leachate solution. In the case o f Brunswick, it is possible that  178  the leachate solution was oversaturated with respect to anglesite due to the large amount o f galena being oxidized.  Table 6.1 Dissociation Constants for Secondary Minerals of Zn and Pb Zinc Dissociation I .cad Secondary Minerals : Constant Secondary Minerals ! dog K.y 3.0 PbS0 ZnS0 7.5 Pb 0 S0 Zn (OH) S0 28.4 Pb 0 S0 Zn4(OH) S0 Zn(OH) Pb(OH) 11.5 '(Parkhust, 1995: PhreeqC database updated 1996) 4  4  2  2  6  4  3  2  4  4  3  2  6.1.4  4  4  2  Dissociation Constant (IogK)' -7.8 10.4 22.1 8.2  The Huckleberry Problem  The potentially acid generating Huckleberry pyrite was found the most reactive o f all pyrites tested. This classification may be due in part by the larger fine grain proportion o f the sample, increasing the surface area o f the sample and possibly increasing the oxidation rate o f pyrite ( M c K i b b e n and Barnes, 1986; K w o n g 1993).  This is not  necessarily the principal cause o f the Huckleberry's high reactivity, however, since the Brunswick sample, which had a higher proportion o f fines, exhibited the lowest reactivity (following the disappearance o f galena i n the sample). A n important factor remains that the reactivity o f Huckleberry pyrite declined from the 4  th  week onward.  Although  Huckleberry remained the most reactive pyrite sample throughout the 20-week leaching period, the waste rock solution-leached pyrite showed the greatest loss o f reactivity (0.07 volts) after Zimapan (0.1 volts).  The reactive nature o f these pyrite samples was  progressively overcome by the precipitate coatings, which are believed to have caused the reactivity decrease.  179  The slight gain i n reactivity observed in the 20  week o f leaching in the solution-leached  Huckleberry pyrites as well as in Tizapa and Louvicourt-2 samples, however, suggested that surface properties o f the pyrite (or its precipitate cover) continually evolved during leaching, affecting the reactivity o f pyrite. Evolution o f surface properties may include changes in precipitate chemistry or morphology, bacteriological alteration o f grain surfaces, or many other factors.  The electrochemically observed changes i n reactivity  could be paralleled with the constantly evolving chemistry o f mineral waste solids and leachates, making predictions difficult to establish with an acceptable degree o f certainty.  6.2  6.2.1  CHEMICAL STABILITY OF PASTE MIXTURES  Effect of Cementing Tailings Waste  Binder (Portland cement and slag-cement mixtures) does provide some buffering capacity to the tailings. The small proportion o f binder added to the reactive tailings studied (less than 6% i n this study) is, however, far from sufficient to neutralise all the potential acidity o f the tailings studied which had N P R values below 0.1. N P R values o f all the backfill mixtures studied remained well below 1 after cement addition, remaining potentially acid generating like the tailings that make up the backfill.  Paste p H measurements suggest that both binders ( O P C and slag-OPC) contain readily dissolved buffering minerals that provide efficient, early-stage buffering capacity to the backfill. A B A analyses as a function o f leaching time indicated that binder N P is readily consumed i n the early stages o f contact with water or A R D solution.  Depletion is  observed to be more severe i n A R D solution for all cases studied. A s a result, even i f 180  cementing material was added initially in sufficient quantity to raise the N P R o f the backfill to environmentally safe levels (i.e. N P R > 3), the N P may be depleted i n this environment at a very early stage, leaving the tailings unprotected.  In most cemented backfilling operations, the backfill is exposed to the environment almost immediately after placement. This study suggests that exposure o f the backfill to an aqeous leaching environment prior to being properly cured w i l l promote the loss o f binder material, which, in the cases studied, accounted for the majority o f backfill buffering capacity. This could potentially be accentuated where slag-cement binders are used, as slag is typically slower to react (hydrate and cure.) than O P C . The loss o f material was further accelerated in A R D solution. Fully cured backfill could, however, show a different response to leaching. Conversly, i f the backfill is not exposed to any leaching solution, for example i n a dry area o f the mine, the cement content should be conserved until the mixture is properly cured.  In the case o f Tizapa paste backfill which contained the highest proportion o f cement and the highest proportion o f sulfide minerals i n the tailings, surficial oxidation o f the samples began to occur within the 14-day curing period. This suggests that the sulfide portion o f the waste was not sufficiently covered or protected by binder to prevent its oxidation i n air. Paste containing such high proportions o f reactive sulfides would likely require complete submergence in water at an early stage to prevent oxidation o f the sulfide portion o f the tailings. Such a material may not be suitable for storage above ground, or underground above the water table, where it could be exposed to air and humidity for extended periods o f time.  181  6.2.2  Cemented Backfill Alteration with Leaching  In 20 weeks o f leaching, differences i n alteration were more influenced by the leaching environment and binder content ( O P C and slag-cement) o f the paste mixtures than by tailings composition. Binder content was found to be the most important factor affecting leachate water quality.  Calcium, potassium, silicon and magnesium loadings i n the  leachates were found to be directly proportional to the solid phase concentration o f these elements in the paste mixture, i n turn, proportional to the binder content o f the mixture.  6.2.2.1 Water-Leached Environments The leachate and solid phase geochemistry along with A B A analyses, indicated that calcium concentrations i n the leachate came principally from the dissolution o f the portlandite phase (Ca(OH) ) o f cement, where O H - ions were consumed to buffer the p H 2  of the leaching solution. This agrees with previous investigations on the effect o f water leaching o f cemented mixtures (Adenot and B u i l , 1992; Revertegat, 1992; Benzaazoua, 1997).  E D X microanalyses under S E M were not sufficiently precise (due to uneven  sample surfaces) to measure a quantifiable decalcification o f the tobermorite, as expected from the literature review.  The modified morphology o f the exposed tobermorite  observed i n altered regions, however, was similar to the morphological descriptions o f Ca-depleted tobermorite by Adenot and B u i l (1992), Carde and Francois (1997) and Taylor (1997).  They have found that i n situations o f severe leaching, tobermorite  generated from both O P C and slag-OPC binder lost its volume and took the shape o f dehydrated gel-like (silica-rich) filaments, loosing binding strength i n the process.  182  The Tizapa and Brunswick mixtures containing the highest percentage o f pyrite and largest proportion o f binder showed deeper penetrating alteration layers compared to the Louvicourt and F I M mixtures. The tobermorite developed from the slag binder i n the Louvicourt  paste was  considerably less  depleted  than  O P C tobermorite  i n all  environments and showed generally fewer expansive minerals after 20 weeks o f leaching. This indicates that the slag-cement binder o f Louvicourt offered increased resistance to water leaching as well as to sulfate attack compared to OPC-binder backfills.  Flooded vs Cyclic Water-Leached Environments Portlandite supersaturation was achieved i n the flooded environment for both the Tizapa and Brunswick samples, which contained the highest proportion o f binder. Masses o f euhedral portlandite lined the outer edge o f the samples and were observed to 1 m m and 3 m m depths respectively after 20 weeks o f leaching. N o portlandite was observed i n the Louvicourt or F I M pastes.  Penetration o f the leachate solution and the resulting concentration gradients o f calcium and hydroxyl ions between the sample pore solution and the leachate drove the dissolution o f portlandite and the diffusion o f these ions out o f the sample. The chemical conditions at the sample edge may have induced the precipitation o f secondary minerals at that location. The same can be suspected to have occurred for all soluble backfill components.  Supersaturation conditions must occur i n the leachate solution before the  secondary mineral may nucleate and grow. Longer contact times between the solution and the paste, such as i n the flooded environment simulated i n this study, facilitate the  183  achievement o f supersaturated conditions and precipitation o f secondary minerals such as gypsum or portlandite.  S E M observations showed, however, that no expansive ettringite or gypsum was present after 20 weeks o f leaching i n the flooded environment i n either Tizapa or Brunswick samples. Instead, secondary gypsum and ettringite were encountered i n the outer edges o f the cycled environment samples. It is clear then that i n the case o f cemented tailings backfill, the occurrence o f secondary minerals is a solubility-controlled process.  Table 6.2 presents the ranges o f solubility constants (Kps), associated solubilities i n water (in mg/1) and equilibrium constants (Keq) o f the secondary minerals encountered i n concrete mixture. L i m e , calcite and salt (halite) are also shown for comparison.  Mineral solubility can be evaluated using the solubility product constant (Ksp) i n instances where dissolution o f the mineral in pure water is not affected by water p H .  Given any dissolving mineral ' A B ' : x A B <-» y A + z B  (6.2)  K s p is evaluated from the activities o f species (activity o f a solid =1) according to: Ksp= [A] x[B] y  184  z  (6.3)  Table 6.2 Solubility Products and Dissociation Constants of Secondary Minerals Derived from Hydrated Portland Cement and Other Reference Minerals Mineral  Solubilit\ Product  Solubility  and Dissociation Reaction  (Ksp or molcs/1 at  (in water at 25"C) *  Portlandite: C a ( O H ) + H -> C a +  5.5 x 10"°  2  +2 H 0  2 +  ( 2 )  2  3 . 2 x 10"  depends on pHfor  5  Dissociation ( onslant (Kcqat25 'C) l  "'"" ' '6.3xl0 " r  6  2 2  pH<10.7 Ettringite: Ca Al (S0 )3(OH)i -26H 0 6  2  4  2  2  -> 6 C a + 2 A l + 3 S 0 " + 12 0 H " + 2 6 H 0 2 +  3 +  2  4  1.23x10^ (4) ! - l 12 P )  depends on pHfor  not available  pH<ll  0  2  Gypsum: C a S 0 - 2 H 0 4  -> C a  2 +  Calcite: -> C a  + S0 " + 2 H 0 2  4  2 +  w  2  9 x 10"°  ( 8 )  2  CaC0  2.5xl0"  (7)  5  > 2.8x10*  + C0 "  ( 8 )  2  3  4.5xl(T  > 3 mg/1  ( 8 )  9  2  ( 5 )  w  2 +  + 2OH  3 . 2 x 10"  3  3.1xl0"  9  4 mg/1  Lime: CaO + H 0 -> C a  l 0  2 1 0 0 mg/1 (i  (1 3  520 mg/1  w  6.3xl0  3 2  -  Halite (salt): N a C l  ^3.4x10'  K  -> N a + C l " +  " 360 000 mg/1  w  3 . 4 x 10  1  * Solubility calculated from Ksp values unless stated otherwise; (1) Brown and Lemay, 1977; (2)Duchesne and Reardon, 1995; (3) Myneni et. al., 1998; (4) Warren and Reardon, 1994; (5) Appelo and Postma, 1994; (6) Pakhurst, 1995 (PhreeqC database updated 1996); (7) Freeze and Cherry, 1979 (Ksp values calculated from solubilities); (8) Fetter, 1993  For minerals affected by p H , such as some oxides and hydroxides, mineral solubility is better represented by the equilibrium constant (Keq). K e q is determined from the law o f mass action, which states that the rate o f chemical reaction is proportional to the activity of the participating substances (including water):  G i v e n any two substances: w A + x B <-> y C + z D 185  (6.4)  K e q is evaluated from the activity o f species (activity o f solids and H 2 O = 1 ) according to: Keq=  rCPxrDf [ A f x [B]  (6.5) x  In minerals such as calcite, the relationship between K s p and K e q is explained by:  CaCO-3 (in water) <-> C a  2 +  + C0 "  K e q = K s p = [Ca ] x [ C 0 " ] = 3.2 x 10" 2+  (6.6)  2  3  2  9  3  (6.7)  However for portlandite, K s p and K e q differ because o f the dependence o f portlandite on the p H o f water. The simple dissolution o f portlandite follows:  Ca(OH)  Ca  2  2 +  + 2 OH'  (6.8)  A n d therefore: K s p = [Ca ] x [OH'] s 1.9 x 10" 2+  5  (6.9)  Whereas the solubility o f portlandite is expressed as:  Ca(OH)  2  +2H'<-> C a  2 +  + 2H 0 2  (6.10) i  Keq =  r C a l = 6.3 x 1 0 2+  22  (6.11)  [H ] +  Equation 6.11 shows that solution p H is inversly propotional to portlandite solubility, explaining why portlandite dissolves more readily i n a low p H A R D solution.  Portlandite has the same solubility range as gypsum but the above showed that  186  portlandite is much less stable in neutral or acidic p H conditions than gypsum or ettringite.  In these conditions, portlandite w i l l readily dissolve and saturate the pore  water with calcium. In the cases o f Tizapa and Brunswick backfill (highest proportions of cement), secondary portlandite was found in the outer sample edges o f the flooded environment only, whereas secondary gypsum was found i n the outer edges o f the samples i n the cycled environment only. One possible explanation for these occurrences is that the pore solution o f both water-leached environments becomes supersaturated with portlandite within 24 hours, hence precipitating this phase i n the outer edges o f the samples.  W i t h increasing leaching time, the secondary portlandite redissolves. i n  response to the declining concentration o f calcium i n the pore solution as other more stable mineral phases continue to dissolve and supply other elements (i.e. A l ) necessary for  ettringite or gypsum to become supersaturated, nucleate and precipitate out o f  solution.  This study suggests that in a flooded environment, where hydraulic conductivity o f the saturated material is l o w and the backfill mixtures contain relatively high proportions o f Portland cement, secondary expansive minerals may form. Consequently, in addition to the dissolution and loss o f binder components and eventually loss o f binder strength, expansive minerals w i l l create internal stresses accelerating the disintegration o f the backfill.  In cyclic leaching environments where the backfill is exposed to a leaching agent for short periods, the leachate contact time may not be sufficient to precipitate secondary, expansive ettringite and gypsum.  Instead, the leachate acts as an ion sink, driving  187  dissolution and depletion o f the binder components, especially the very unstable portlandite, and flushing them out before supersaturation and precipitation o f expansive mineral phases can occur.  Solid phase geochemistry showed that calcium and sulfate  were more effectively depleted on the edges o f the samples i n the cycled environments compared to flooded environment.  Cyclic leaching resulted in an accelerated loss o f  binder material and, in the case o f OPC-binder, a loss o f tobermorite integrity i n the periphery o f the samples as indicated by the more abundant grain cavities and greater depletion o f major ions in the outer layers o f the cycled samples.  The extent o f  tobermorite depletion was lower in the slag-cement binder mixture o f Louvicourt pastes.  S E M observations and solid phase geochemistry o f the altered layers o f all samples, indicate that backfill leaching is a surface phenomenon characterized by alteration layers widening inwards with increasing leaching time.  The data obtained i n this relatively  short term leaching study does not, however, provide information on the long-term rate o f advancement o f the leaching front, whether the leaching front would continue to progress or eventually cease its advancement. In all water-leached cases, the augmented porosity resulting from  binder dissolution,, combined with the precipitation o f expansive  secondary minerals suggests that advancement w i l l likely continue to proceed as porosity is augmented with time and the leachate can more easily penetrate the backfill.  Effect of low binder content The absence o f portlandite, gypsum or ettringite i n the F I M and, to some extent, i n the Louvicourt samples, is likely related to both the small proportion o f hydrated Portland cement available to be dissolved and to the  generally higher porosity o f these samples. 188  H i g h sample porosity w i l l facilitate contact between solute and solvent and the diffusion of dissolved elements out o f the sample making supersaturation more difficult to achieve. In addition, i f the initial concentration o f portlandite to be dissolved is small, pore solution supersaturation may not be attained.  This is especially true i n the case o f slag-  mixed binders where the hydration reaction and formation o f tobermorite consumes portlandite (Taylor, 1997; Kosmatka et al., 1995). In this case, supersaturation o f high calcium phases (such as ettringite) is more difficult to achieve without a readily available source o f dissolved calcium.  Gypsum supersaturation can occur, however, as was  observed i n the deeper areas o f the cyclic-leached pastes. internal pressure within the backfill.  This, i n time, w i l l create  Wide-spread secondary gypsum growth would be  responsible for internal breakdown and loss o f strength o f the backfill.  6.2.3  Ferric Sulfate or Artificial ARD Environments  Pore water i n concrete has a stable p H ~12, the stability p H o f portlandite. The artificial A R D solution p H o f 2.4-2.5 used i n this experiment is considerably more conducive to portlandite dissolution than water.  Decreased pore solution p H accelerates  the  dissolution o f portlandite whose hydroxide ions are consumed to buffer the solution p H . This was demonstrated i n these experiments by the leachate geochemistry data, where p H was buffered on contact with the pastes in the early cycles o f leaching. The A B A tests also attested to the high, early dissolution o f portlandite whereby N P depletion was considerably more severe and deeper penetrating in the ARD-leached pastes than i n the water-leached pastes for both O P C and slag-OPC binder mixtures. Alteration layers o f the ferric sulfate-leached pastes were almost double in thickness to those o f the water189  leached pastes. Indeed, ARD-leached pastes had a measurable lost o f mass in 20 weeks o f leaching, contrary to the water-leached pastes.  S E M observations o f the altered layers o f O P C binder backfills showed important losses of tobermorite volume as well as more pronounced changes i n tobermorite morphology (likely from the decalcification o f tobermorite) and apparent strength losses (increased grain cavities viewed with S E M ) in the altered layers compared to the water-leached samples.  Tobermorite loss was slightly lower i n the slag-OPC binder (Louvicourt)  pastes, but still higher than i n water-leached environments. The limited ucs tests carried out on all samples did not show, however, any strength loss with time. This is probably due to the relative thin r i m o f alteration compared to the diameter o f the entire sample.  A l l ARD-leached pastes samples were covered by surficial iron hydroxide precipitates. They also possessed similar amorphous, oolitic or botryoidal iron sulfate precipitates, possibly jarosite, i n the second alteration layer ( 2  nd  from edge).  It remains unclear,  however, whether these minerals were precipitated directly from the ferric sulfate solution or precipitated as a product o f the oxidation o f iron sulfide in these areas. Both cases imply the penetration o f an acidic ferric sulfate solution front into the paste. The sulfide minerals would have been oxidized i n the acidic environment whereas metal sulfates and hydroxides would have been precipitating on the pore solution side o f the front where p H and ionic strength could allow the precipitation o f these minerals. In any case, it is possible that continued growth o f both or either o f these precipitates (iron hydroxide and/or iron sulfate) eventually passivates the surface o f samples.  The  evolution o f leachate  p H , redox  potentials and aqueous iron concentrations 190  towards the values o f the leaching solution support the suggestion that the samples were passivating as leaching progressed. It is unclear whether the depressed reactivity o f the paste was a consequence o f precipitate coatings from the A R D - s o l u t i o n (ferric hydroxide precipitates visible on all ARD-leached pastes), or i f it was due to an exhaustion o f soluble buffering minerals i n the area actively leached. In the latter case, loading rates o f paste constituents to the leachate would be proportional to the infiltration rate o f the solution into the backfill.  Regardless o f the processes responsible for the apparent  passivation o f the paste, the implication is that paste backfill exposed to A R D w i l l eventually lose its ability to buffer the drainage, following which the acidic, metal-loaded mine drainage coming i n contact with the backfill would be free to flow unaltered around the backfill to the surrounding groundwater or surface water.  Another aspect o f the eventual passivation o f the paste is that the aggressive sulfate solution would eventually be prevented from penetrating the backfill and degrade the hydrated cement phases o f the mixture. In any case, the unchanging sulfate levels i n the leachates suggest that the sulfate i n the leaching solution may not actively participate i n cement altering reactions in the first 20 weeks o f leaching. A l l data suggest that i n 20 weeks o f leaching, external sulfate plays a minimal role i n cement-altering reactions when tailings already contain high concentrations o f sulfate.  The presence o f relatively high aqueous zinc concentrations in the Brunswick sample, combined with the results o f the pyrite reactivity study, support the possibility o f active sulfide mineral oxidation i n the outer layers o f the paste.  Absence o f dissolved iron i n  the leachate indicate one o f two possibilities: 1) no oxidation o f iron sulfide minerals  191  (pyrite, pyrrhotite), or more likely 2) oxidation or iron sulfides followed by precipitation o f iron sulfate within the backfill. The pyrite reactivity study suggests that pyrite may get passivated at an early stage. Verification o f the latter would require further investigation.  192  7  CONCLUSIONS  The following conclusions are made based on the results o f this study:  7.1  •  PYRITE REACTIVITY STUDY  Results o f the electrochemical study supported by chemical, mineralogical and leachate chemistry data, suggested that the presence o f non-pyrite sulfide mineral impurities, i n contact with pyrite was the most important parameter affecting pyrite reactivity i n initial leaching cycles.  In more advanced stages o f leaching,  electrochemical analyses together with S E M observations suggested that precipitate coatings, either visible F e O O H coatings or angstrom thick sulfur or polysulfide layers not visible under S E M , played a dominant role i n the reactivity o f pyrites, effectively decreasing pyrite reactivity as leaching progressed.  •  The presence o f sphalerite was believed to form a galvanic couple with pyrite until sphalerite itself became unreactive because o f precipitate coatings on its surface. Once the availability o f sphalerite became limited, pyrite became free to oxidize and develop a precipitate coating, which proceeded to decrease its reactivity.  •  The extent o f galvanic protection offered by sphalerite was proportional to the sphalerite content o f the sample. Lower concentrations o f sphalerite in Louvicourt-2 is believed to have offered either less efficient or shorter time o f galvanic protection compared to the Brunswick sample.  Higher leachate concentrations o f iron and  193  increased pitting o f the Louvicourt-2 pyrite suggested that this sample oxidized to a greater extent than the Brunswick pyrite.  Leachate concentrations unleached pyrite samples.  o f zinc were proportional to the sphalerite content o f The high solubility o f zinc sulfate and zinc hydroxide  explains why no zinc precipitates were found in any o f the leached samples.  The presence o f galena in pyrite samples was signalled by a current release at around 0 mvolts, lower than the potential at which pyrite oxidized i n that environment.  The  intensity o f the voltametric peak appeared to be proportional to the concentration o f galena i n the sample. Contrary to zinc, leachate lead concentrations remained l o w throughout the leaching period by precipitating out as anglesite (PbSC»4).  It is  possible that the precipitation o f secondary anglesite could have indirectly provided pyrite grains some resistance to oxidation by reducing the surface area o f pyrite available to oxidize.  After the 20-week leaching period, all pyrites had decreased reactivities with respect to its unleached reactivity regardless o f the presence o f mineral impurities. In many cases, however, a slight regain o f reactivity observed in the 2 0  th  week o f leaching  may indicate that surface precipitates were modified, no longer providing increasing levels o f passivation. This study does not provide information as to the resistance o f the passivating layers to changing conditions i n the leachate.  This factor commands  caution i n the long-term extrapolation o f the effectiveness o f passivation layers.  194  7.1.1  Huckleberry Pyrite  The potentially acid generating Huckleberry pyrite was found to be the most reactive o f all samples studied.  However, the reactivity o f Huckleberry pyrites began to decrease  within 4 weeks o f leaching, progressively overcoming its reactive nature. Reactivity loss was slightly greater i n the waste rock (low-grade ore)-leached sample, possibly because of the greater amount o f ions i n solution available to for precipitate coatings.  The  potentially reactive pyrite present i n the low grade ore for which were carried out 3-year kinetic tests may also have been reactive but surfaces o f the pyrites probably passivated from early cycles onward by the ionically charged leachate solution to which it was exposed.  Post leaching mineralogical analyses  should have been carried out to  characterize the surfaces o f these pyrites.  7.1.2  Application of this Study to Field Investigations  A s pyrite is never the only sulfide mineral i n a deposit or in mineral wastes, this study has shown that, at the scale o f the study, occurrence o f mineral impurities was the most determinant factor o f the reactivity o f pyrite, above crystal morphology or stoichiometric imbalances.  Optical and electronic microscopy i n conjunction with point  source  chemical analyses permitted to distinguish between solid solutions and separate mineral phases. The mode o f occurrence o f impurities present with pyrite dictates the ability o f that mineral phase to oxidize and provide galvanic protection. I f a sphalerite grain has a very small contact area with a pyrite grain or is separated by a micron size phase o f a different mineral, galvanic protection could be diminished or ineffective.  195  Some pyrite grains that were not i n contact with mineral impurities, for example i n the Tizapa and Zimapan samples, oxidized to a similar extent than pyrite grains i n the Huckleberry or Louvicourt-1 samples.  This indicates how the intrinsic mineralogical  heterogeneity o f a given geologic deposits or mineral waste pile make predicting A R D generation tentative at best. Although detailed stoichiometry and mineralogy can give indications as to the expected reactivity o f a mineral, these characteristics may not be the dominant factors affecting the water quality and/or overall reactivity o f a waste rock pile or tailings pond.  Megascopic features such as pile or pond construction, hydrology or  other site conditions also influence the possible generation o f A R D from a given mineral waste as observed by many researchers (Kwong, 1991; Ritchie, 1994; Schafer et al., 1994; Woyshner and St-Arnaud, 1994; Otwinowski, 1997). It is therefore imperative that representative mineralogical samples and megascopic characteristics, such as (but not limited to) multiple grain size fractions that could include sulfide mineral pockets, be included i n the make up o f rock samples used in kinetic tests in order to make more accurate A R D predictions.  At  a fundamental  level o f understanding  the reactivity o f pyrite, however,  the  electrochemical characterisation technique o f cyclic voltamperometry on carbon paste electrodes used i n this study was an effective tool to demonstrate how the various mineralogical characteristics work together to influence the overall reactivity o f pyrite. C y c l i c voltamperometry was found to effectively measure the ability o f pyrite to oxidize under the conditions o f the kinetic tests carried out i n this study. This type o f analyses can be revealing i n situations where kinetic test results are inconclusive, such as i n the  196  case o f Huckleberry. The high degree o f expertise involved i n interpreting the results o f an electrochemical study, however, can be an obstacle to the widespread use o f this technique as an A R D predictive tool.  7.2  CEMENTED PASTE BACKFILL LEACHING STUDY  The small proportion o f binder added to reactive tailings i n this study was insufficient to rise N P R ratios above 1 or change the status o f potentially acid generating tailings to non-acid generating backfill. The addition o f cement to reactive tailings does not necessarity constitute an effective prevention against generation o f A R D from reactive tailings.  Some backfill mixtures that contain a high concentration o f reactive pyritic tailings used as the principal aggregate (such as Tizapa backfill) may oxidize in air. This type o f backfill would likely require complete submergence i n water at an early stage to prevent oxidation o f the tailings. Such a material would not be suitable for storage above ground or underground above the water table, where it could be exposed to air and humidity for extended periods o f time.  •  The hydrated portland cement phases o f backfill readily dissolve i n water. The extent and rapidity o f depletion is generally more pronounced in a l o w p H solution (i.e. ARD).  The small addition o f neutralisation potential provided by the cement is  depleted at an early stage in the external layers o f the samples, leaving the tailings in these areas exposed to the environment. alteration were  i n general  more  In 20 weeks o f leaching, differences i n  pronounced 197  between  the  different  leaching  environments than between the different tailings mineralogy making up the backfill mixtures. Binder content (cement or cement/slag) o f the paste mixtures was found to be the most important factor affecting the release o f major ions to the leachate in all environments.  Backfill leaching is a surface phenomenon characterized by alteration layers widening towards the core o f the sample with increasing leaching time. The same would be suspected to occur i n underground mine workings filled with cemented backfill. The backfill would start to deteriorate at the contact with the host rock, with the depth o f penetration o f the alteration zone increasing with time.  The current short-term  leaching study does not provide, however, information on the long-term rate o f advancement o f the alteration zones, whether it would remain constant or eventually cease its advancement. It can be postulated from the experimental data, however, that in  water-leached  environments, the  augmented  porosity and  precipitation o f  expansive minerals w i l l encourage leachate penetration into the backfill.  Hence i n  these environments, the rate o f advancement o f the alteration zone would probably not slow down.  This study indicated that in ARD-leached environments, the backfill material become passivated with time. The passivation mechanism remains unclear, however, either due to an increasing thickness o f ARD-solution precipitates onto the backfill or due to the formation o f self-sealing iron precipitates resulting from sulfide oxidation i n the outer layers o f the backfill. In either case, the consequence is a decreased capacity o f the backfill to neutralize incoming A R D solution generated within the host rock. In  198  time, the acidic and metal-loaded mine drainage could flow unaltered around the backfill to the surrounding groundwater or surface water.  This study suggests that  cemented tailings may not constitute, i n the moderate to long-term, an effective buffering material to neutralize A R D generated within the mine or untreated A R D effluent.  Some oxidation o f sphalerite and galena occurred i n the A R D and cyclic leaching environment but not in the flooded water-leached environment.  Oxidation o f pyrite  and pyrrhotite may have occurred in the A R D or cyclic environments but would have reprecipitated within the backfill matrix.  The reactivity o f iron sulfides within  backfill would require specific investigation.  The apparent loss i n strength that was observed under S E M for some A R D and waterleached pastes was not manifested i n the ucs tests o f the leached samples.  These  results may not be statistically significant as only one sample per category was tested. Conversly, they could indicate that the thin alteration layer is not sufficiently significant to alter the strength o f the entire sample measured by the apparatus. The study methodology was not designed to measure the strength o f different areas within the sample.  Knowledge o f the peripheral strength would be very useful (i.e.  periphery o f backfilled stopes) as this area is exposed to blasting when mining next to the backfill.  Lower peripheral strength o f the backfill may cause higher than  expected dilution o f the ore. Ouellet et al. (1998) have pointed out that this was a particular problem at the Louvicourt mine with a previous backfill formulation.  199  This study showed a definite alteration o f the backfill when subjected to water or A R D leaching solution, resulting i n some apparent losses i n strength o f the backfill i n the altered layers.  Thicknesses o f the alteration zones, however, ranged from less  than 1mm to 10mm in 20 weeks, with the thicker alteration zones associated with the sulfide-rich tailings in A R D - l e a c h solutions. It is necessary to place these results into the context o f a mine setting where field rates o f alteration can be higher or lower, depending on the local groundwater regime and annual precipitation on the site. The possible impacts o f leaching paste backfill depend on the purpose for which the backfill is used. For example, a highly pyritic cemented backfill may not suit the purpose o f a buffering plug meant to neutralize pre-existing acidic drainage within the mine.  The same backfill could, however, effectively serve the purpose o f ground  support i f placed in an isolated stope and promptly submerged in a n o n - A R D environment or i f placed in a dry area.  This study highlights the importance o f carrying out leaching studies on cemented paste backfill mixtures in order to evaluate its stability under the conditions to which it w i l l be exposed.  200  8  8.1  RECOMMENDATIONS  PYRITE REACTIVITY STUDY  Electrochemical studies o f minerals should be considered where static test results are inconclusive as to the potential acid generation o f a mineral waste. These studies could be realized during kinetic tests, for one or a number o f sulfide minerals o f concern to a particular mine site. The test set-up could consist, for example, o f retrievable pockets o f the sulfide mineral o f concern placed into the column o f rock sample being kinetically tested or placed as an appendage, similar to the Huckleberry solution-leached set-up o f this study.  The evolution o f sulfide mineral reactivity could, i n this way, be verified  periodically along with the resistance o f passivation coatings.  The following recommendations  should be followed to facilitate interpretation and  enhance the understanding o f sulfide mineral leaching tests:  1. In order to facilitate the observation o f precipitate products and the evolution o f grain surfaces, the pyrite samples should be washed in an ultrasound bath to remove the small mineral particles electrostatically stuck to pyrite surfaces. This would not only yield particle-free pyrite surface but also provide a uniform grain size sample from which could be calculated oxidation rates per surface area o f pyrite.  2.  Further analyses o f the precipitate products should be carried out to define the composition o f the precipitates and test the resistance o f the passivation layer. This  201  would involve:  •  Detailed cathodic electrochemical analyses o f the leached grains to verify the relative resistance o f the grain coatings with advancing leaching cycles, increasing the cathodic current sweep to -1.5 volts.  •  Detailed quantitative mineral analyses o f the precipitation products or grain surfaces. This would require making resin-mounted polished sections o f leached samples and analyzing the grain edges by using either point source chemical analyses ( E D X , W D X or E A S ) or image analyses under S E M .  3.  Additional information should be obtained on the effect o f oxidation pits on the reactivity o f pyrite. C y c l i c voltamperometry could be performed on unleached and pitted grains o f the same sample, previously acid-washed to remove the precipitate coatings.  8.2  PASTE BACKFILL STUDY  The following recommendations could facilitate future backfill leaching studies and help to resolve issues pending from this study.  1. Longer-term leaching studies would be useful for the following:  •  To obtain information on the progression o f the leaching rate o f the backfill with advancing leaching times and determine i f the rates w i l l increase, decrease or w i l l remain the same as leaching progresses.  202  To obtain thicker leaching zones, which would facilitate separation o f the different layers and to provide a greater amount o f sample to carry out multiple analyses o f solid phase geochemistry, acid-base accounting or any other analyses. Bigger sample dimensions would also help i n this respect, keeping sample diameter = Vi height specification for the ucs test.  To provide the opportunity to observe under microscope and document the oxidation o f sulfide minerals.  Larger sample sizes would be necessary i n longer-term tests to keep a relatively unleached core for comparison purposes and for the determination o f the rate o f penetration o f the alteration zone, making sure that the alteration zone w i l l not fully penetrate the sample.  Control samples should be prepared and stored for the duration o f the tests i n a lime solution at p H 12. Both the leached samples and the control samples could then be chemically analyzed and physically tested to compare the effects o f leaching.  The following analyses should be carried out as they provide very useful information as to the geochemical changes occurring within the pastes:  evolution o f A B A  parameters in the altered layers compared to the control sample and to the core o f the flooded sample; elemental analyses o f the leachate, the solid phase o f the altered layers, o f the control sample and o f the core o f each leached sample. parameters should include the following: C a , A l , M g , K , S i , C 0  2 3  Analytical  \ Fe, Z n , C u , Pb,  trace elements (depending on the mineralogy o f the tailings, i.e. A s , H g , Sb, Cd) as  203  well as p H , conductivity and redox potentials o f the leachate solutions and the sulfide 2  speciation (S tai, SO4 to  2  S ") o f the solid phase samples.  2. Preparation o f the samples for analyses:  •  A polished section o f the altered paste could be prepared by using ultra low viscosity resin that would be deeply penetrating. Such samples would allow quantitative E D X analyses necessary to more easily identify the various hydrated binder phases and possibly the presence o f oxidation rims on sulfide minerals. These samples should be prepared i n addition to the flat unpolished sections that permit the observation o f 3dimentional crystal morphologies usefull for mineral identification.  3.  Carry out leaching analyses o f 3-month cured samples in parallel with 14-day cured samples (or earlier i f necessary) to verify the effects o f advanced curing on the leachability o f the hydrated cement in the backfill.  4.  Carry out multiple (minimum 3) leaching cells o f the same samples in order to perform multiple ucs analyses, for statistical significance. This could be organized as one large cell containing 3 samples, to ensure that all are subjected to similar conditions.  5. 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E . et ah, 1993: Electrochemistry of Pyrite, in: Hydrometallurgy; Fundamentals, Technology and Innovations, Proceedings o f the 4 International Symposium on Hydrometallurgy, Hiskey and Warren Editors, Salt Lake City, Utah, August 1993: (Chapter 5) 85-99. th  Warren C.J. and Reardon E.J.,1994: The Solubility of Ettringite at 25 deg. C, Cement and Concrete Research, v o l . 24, no. 8: 1515-1524. Williams D.J., 1997: Effectiveness of Co-disposing Coal Washery Wastes, in: Tailings and M i n e Waste '98 Conference Proceedings, Fort Collins Colorado, January 1998: 335-341. Woyshner M . R . and St-Arnaud L . , 1994: Hydrogeological Evaluation and Water Balance of a Thickened Tailings Deposit Near Timmins, ON, Canada, in: 3 International Conference on the Abatement o f A c i d i c Drainange, Pittsburgh, Pensylvania, A p r i l 1994: 198-207. r d  Ybarra, R . R . , 1998: Prediccion de la Generacion de Drenajes Acidos en una Presa de dales con Altos Contenidos de Pirita. M . S c . Thesis, Faculty o f Chemical Sciences, Universidad Autonoma de San Luis Potosi (Mexico).  212  Z h u X . , et al., 1992: Transpassive Oxidation of Pyrite, in: Proceedings o f the International Symposium on Electrochemistry i n Mineral and Metal Processing, Woods y Richardson Editors: 391-409.  i  213  APPENDIX I - ANALYTICAL RESULTS: PYRITE LEACHATE  214  Table 1-1 (cont'd)  Cycle  Pyrite Leach Solution Pb (mg/l) Hk-I  1 2 3 4  Pyrite - Leachate Chemistry  0.5 1.4  Hk-w Lou1 Lou2 4.9 4.3 3.8  6.8 7.4  87 92  2.8 8.3 7.7  0.1  5 6  92 71  0.1 0.3  7  46  0.1  8 9 10  0.4  11 12  0.3 <0.1 0.2 0.1  13 14  0.5  0.3  4.2 4.1  0.3  0.3  15  0.5  0.8  16 17 18  0.5 0.4 0.9  19 20  1.6 0.3  21 22  5.8  19 5.9 5.5  61 73  18 42  28 94  2.5  53 48 33 8.4  75 51  240 366 320 174  0.2 0.1  4.1 3.2  39 9.2 40 17  0.6  2.0  177 180 188  0.1 0.4 0.2 < 0.  3.0 2.1  31 10 42 12 36 9.0  68 59 43 52 36 32  225 273  3.0 3.3  2.2  5.4  8.0  49 37  1.3  5.2  6.0  27  0.2  0.1  2.0  4.3  2.3 1.7  7.2 6.6  17 13  21  0.8 0.4  6.8  2.3  2.0 1.4  10 11  0.1 < 0 . 0.1 < 0 . 0.3 1.2  0.5  3.3 3.4  1.8 1.8  0.5 0.2 1.7  3.7 2.8 2.4  1.0 0.8  0.6 0.2  3.1 0.6 2.7  2.3 2.6 2.6  23  0.2  0.5  24 25  0.1 0.2  26  0.2 0.4  0.9  14  9.5 <0.1 9.9 <0.1  0.7 1.7  0.5  1.2  6.1  13  6.5  0.2  0.1  1.0  6.0 6.3  13 12  7.0 8.4  0.3 < 0. 0.2 < 0.  0.8 1.2  1.5 0.5  5.9 6.0  13  9.4  7.6 6.5  8.3 7.4  0.1 0.2  0.6 0.5  10 11 14  2.6  0.7  7.8  5.1 0.1 5.4 <0.1  0.3  2.5  0.6  6.6  0.3  2.8  0.5  0.1  0.6  1.8  27  0.2  0.6  28  0.4  0.4  29  0.1  30 31 32  7.9 6.5  As (mg/l)  Hk-I Hk-w Lou1 Lou2 Tiz Zim Brw 1.6 0.4 0.2  1.4 1.1  6.1 5.1 5.0  Zn (mg/l) Zim Brw  0.3 2.3 0.5 0.1  1.0 0.7  0.4  Tiz  22 47  33 6.5 18 2.7 9 1.6 14 2.2  15 7 10 11  0.5 0.8  13  7.3 <0.1 < 0 .  0.5  6.8  11  6.9  0.2  0.6  0.2  7.2  5.7  4.2  0.2  0.5  1.3  0.2  5.5  6.4  4.1  0.3  0.3  0.5  5.3 1.2  1.8  0.5  3.9  2.0  4.4  0.4  0.2  0.5  1.1  1.3  0.3  6.3  5.9  3.7  0.2  0.3  0.8  4.6  1 1.5  0.2  0.3  1.4  0.3  6.7  8.1  0.1  0.5  1.7  0.4  7.1  7.9  0.4  0.6  0.1  0.5  0.2  0  0.5  36  0.3  0.3  0.6 <0.1  7.2  6.6  9.9  0.2  0  0.7  1.4  4.6  <1 <1 <1  <1 <1 <1  <1 <1 <1  1.4  4.1 4.2 3.4  <1 <1  <1 <1  <1 <1  1.3  <1 <1  <1 <1  <1 <1  1.5 1.4  3.6 3.4  <1 <1  <1  <1  <1  <1  4.6  <1  201  <1  <1  <1  <1  1.1  1.5  <1  1  248  <1  <1  <1  <1  <1  2.1  <1  3.8 1.3 0.7  213  <1  <1  <1  <1  <1  3  <1  246  <1  <1  <1  <1  1  2.8  <1  <1  <1  <1  <1  <1  3  <1  <1  <1  <1  <1  2.6  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  0.2 1.3 1.1  <1  <1  <1  <1  <1  1.7  <1  1  <1  <1  <1  <1  <1  2.1  <1  <1  <1  <1  <1  <1  2.3  <1  62  <1  <1  <1  <1  <1  2.7  <1  1 1.4  100  5 1.1  3.2 0.8 0.9 1.4 1.5  65  0.4  0.7  0.1  7.0  6.1  6.7  0.3  0.2  0.8  4.1  38  0.2  0.3  0.8  0.2  6.1  4.6  5.0  0.2  0.3  0.7  3.9 1.2  2  57  <1  <1  <1  <1  <1  2.5  39  0.6  0.5  0.5  0.2  5.6  3.3  5.6  0.6  0.2  0.4  3.3 1.2  1  27  <1  <1  <1  <1  <1  2.1  0.4 0.59 <0.1 <0.1  6.5  3.3  0.7  0.4  0.4  2.8  <1  <1  <1  <1  <1  1.8  1 0.8  <1 <1  0.2  215  <1 <1 <1  <1  1.6  3.4 4.1  37  40  <1  <1  4  1.6 1.6  <1 <1  187  4.4 1.1  7.2  <1  7.1 1.6 3.3  0.1  6.2  <1  <1 <1  5.7  0.4  1.9  <1  3.7 4.1  <1  6.7 7.2  <1 <1  <1  0.1  7.3  <1 <1  <1  0.8  5.4  <1 <1  <1  0.2  1.1  <1  <1  5.5  0.3  <1 <1  1.0 < 1 < 1  4.4 3.2  147  1.2  0.8  1.0  7.2 1.9 3.6  4 1.2 1.1  0.4  2.7  <1 <1  1.7  4.8 1.1 1.4  0.2  <1  <1 <1  <1  0.5  0.5  <1  <1 <1  <1  0.6  0.2  <1  <1 <1  <1  0.3  0.3  <1 <1  <1  0.1  35  <1 3.4  146  0.1  34  <1  <1 1.1 <1  131 151  0.3  0.3  <1  <1 <1  <1 <1 <1  2.8  0.3  <1  <1 <1  <1 <1 <1  0.1  33  <1  <1  0.3  <1 <1  <1 <1  149 145 135 133  <1 <1  <1  154  16 2.1 6.6 18 2.4 6.4 8.4 1.5 9.5 2.2  <1  <1 <1 <1 <1  <1 <1 <1  16 1.9 6.1 19 2.1 7.3 22 2.4  1.1  Zim Brw <1 <1 <1 <1  <1 <1 <1  <1  <1 <1  0.9 1.1  Tiz  <1 <1 <1  183 176  0.1 0.1  0.1  222 239  Hk-w Lou1 Lou2 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1  28 3.4 15 2.7 8.8 22 3.2 9.5  0.1  0.3  261 295  Hk-I  <1  Table 1-1 (cont'd) Cycle  Pyrite Leach Solution Fe total (mg/1) Hk-I  1  Pyrite Leachate Chemistry  2  388 848  679 764  3 4  738 475  818 427  80 80  85 97 69 108 44 66  5 6  345 419  448 364  70 47  7  340  286  39  8 9  10  273 302 246  336 280 304  49 48 50  11 12 13  247 146 80  33  14  128  173 122 156 287  15 16  256 152  203  17  129  18 19  17 23 41  75  46  66  38  58 64 57 52  35 43 47 45  35 21 26  68  0.3  53 45  0.3 0.2  45 40 34  0.2 0.2 0.2  30 19 24  27 15 19  0.2  1.1 0.9 0.7  0.3 0.8 0.9 0.4 0.4  0.8 0.5 0.9 0.4 0.4 0.4  51 40 36  33 37  20 21  181  35 34  39  160  35  30  29  17 14  0.1 0.2  125 139 142  180 161 163  46  28 31  28 31  14  0.3  0.1 0.5  13  0.3 0.5  0.6 0.4  156  47  33 33  14  112  24 21  13  0.4  22  106  123  28  14  26  12  23  75  139  46  17  28  11  0.5  24  91  89  26  12  12  0.7  25  60  99  28  12  9  1.1  0.3  2  2.1  20 21  26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  Cu (mg/l)  S04 (g/l) Hk-I  Hk-w Lou1 Lou2  Tiz  <0.1 <0.1 <0.1  0.2 <0.1 <0.1  0.1 0.4 2.0 <0.1 1.1 <0.1  0.2 0.2 <0.1 16 <0.1 <0.1 22 0.3 <0.1  <0.1  0.1  0.4  0.1  0.4  0.3 <0.1  0.2 0.3 0.1 0.1 0.0 0.1 0.2 0.1 0.0 0.0 0.1 0.2 0.1 0.2 0.1 0.1 0.2 0.1 0.2 0.2 0.0 0.0 0.2 0.2 0.2 0.1 0.1 <0.1 0.2 0.2 0.5 <0.1 <0.1  0.9  0.1  0.5 0.4  0.0 0.1  6.6 4.2  0.4 <0.1 0.3 <0.1  4.4 8.4  0.3 <0.1 0.3 <0.1  3.6 3.4  0.3 <0.1 0.3 <0.1  1.9 1.6 1.4  0.2 <0.1 0.2 0.2 0.2 <0.1  2.1 1.6  0.2 <0.1 0.2 0.2  Zim Brw Hk-I Hk-w Lou1 Lou2 Tiz Zim Brw 86 27 8.0 1.0 1.3 0.3 0.3 0.2 0.1 0.1 64 118 0.5 1.7 1.9 0.3 0.4 1.7 0.3 0.2 97 0.3 2.4 1.9 0.4 0.3 0.2 0.3 0.2 61 45 86 0.2 0.9 0.1 0.2 0.1  Hk-w Lou1 Lou2 Tiz  42 41  73  15  8  30  0.5  0.6 0.5 0.5  0.1 0.3 0.2 0.2 0.3 0.2 0.3 0.5 0.2 0.4 0.2 0.1  0.2 0.2 0.5 < 0.1 <0.1 0.1 0.1 0.1 < 0.1 <0.1 0.2 0.2 0.2 < 0.1 0.1 0.2 0.2 0.2 < 0.1 0.1 0.4 0.2 0.1 0.1 <0.1 0.1 0.1 0.0 0.1 < 0.1 0.12 0.1 0.1 0.0 0.2 < 0.1 <0.1 0.1 0.1 0.1 0.1 < 0.1 <0.1 0.2 0.2 0.3 0.2 < 0.1 <0.1 0.1 0.6 0.11 <0.1  0.1 0.1 0.4 <0.1 0.4 <0.1 0.2 <0.1 0.2 <0.1 0.3 <0.1 0.4 <0.1 0.3 <0.1 0.4 <0.1 0.4 <0.1 0.5 <0.1 0.4 <0.1 0.2 <0.1  <0.1 <0.1 0.3 <0.1 <0.1 <0.1 <0.1 <0.1 0.2 0.1  0.2  0.3  0.3  0.0  0.1 0.1  0.3  0.3  0.1  0.1 0.1 0.1  85  17  8  3  7.9  70  58  15  6  <0.1  16  47  67  20  8  3  31  47  69  19  9  12  3  73  20  11  15  3  46  103  68  98  55  71  87  15  11  17  2  66  58  82  21  12  19  3  82  0.2  0.2  60  67  16  10  15  3  76  0.2  0.2 <0.1 <0.1  60  88  21  13  16  3  89  0.2  0.2 <0.1  67  73  22  12  15  4  127  0.2  0.2 <0.1 <0.1  64  72  19  10  13  3  110  52.8  61.1  16  10  11  1.9  1.7 <0.1 <0.1 1.5 0.3 <0.1 1.9 0.2 <0.1 1.6 1.9  0.2 <0.1 0.3 <0.1  1.0  0.2 0.1  1.4  <0.1  0.2 0.00 0.15  0.2  0.1  1.6 <0.1 <0.1  0.14  0.18  0.2  0.2  1.4<0.1  0.2  0.00 0.13  0.2  0.1  0.8 <0.1  0.4  0.2 0.17 0.14  0.2  0.1  1.0<0.1  0.1  0.5 0.17  0.1 <0.1  0.9 <0.1  0.3  0.4  0.1  2.2 <0.1 <0.1  0.2 <0.1  0.00 <0.1 <0.1  Zim Brw  0.3  0.1 0.1  0.1  0.3  0.1  0.5 0.14 0.11  0.1  0.1  0.8 <0.1 0.5  0.2  0.3  0.1 0.1  0.12 0.13  0.2  0.1  0.9 0.1  53  0.6  0.2  0.5 0.13 <0.1  0.1  0.8  0.1 <0.1  57  0.3  0.5 0.00 0.28  0.1  0.1  1.1  0.1 1.5  0.4 0.13 0.15  0.2  0.1  0.9  0.1 2.4  0.3 <0.1 <0.1 0.1 0.1  0.4 0.12 0.16 <0.1 <0.1  1.3<0.1  4.2  0.1 <0.1 0.1 0.1  0.5 0.10 0.12 <0.1 <0.1  1.0 <0.1  6.4  0.2 0.2  0.1  0.1  0.1  0.1 0.1  0.1 0.2 0.1 0.1 0.1  0.1 0.2 0.1  0.2 <0.1 <0.1 0.1 0.1  216  0.3  0.00 <0.1 <0.1 <0.1  0.7 <0.1 4.3  0.00  1.1 <0.1 4.5  0.11  0.1 <0.1  0.11 <0.1  0.1 <0.1  1.0 <0.1  0.13 0.11 <0.1 <0.1  0.8 <0.1  <0.1 0.14  0.1 <0.1 <0.1 0.1  Pyrite Leachate Chemistry  Table 1-1 (cont'd) Cycle  Pyrite Leach Solution conductivity (uS) Hk-I  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32  33 34 35 36 37 38 39  40  1751 3900 2440 1744 2460 2430 1820 2420 1908 1945 1412 899 1119 1440 1123 1207 1167 1251 1365 1042 1124 757 950 818 944 674 526 231 588 629 653 748 754 701 712 654 697 691 575  Hk-w Lou Lou  1954 2900 2860 1909 2490 2360 1697 2400 1917 2160 1296 972 1111 1391 1088 1378 1190 1374 1301 1230 1379 883 1336 950 1087 695 829 798 684 679 720 837 797 736 701 580 685 635 619 453  508 576 424 344 367 405 322 385 402 381 299 195 227 358 320 296 329 418 417 335 435 269 416 262 301 142 196 212 223 189 244 241 222 166 231 152 180 190 158 128  548 676 573 444 439 450 391 470 430 361 277 207 256 396 328 701 281 275 292 245 244 159 182 173 160 110 136 120 154 124 130 157 146 128 141 110 119 118 99 91  Tiz  Zim  524 490 418 409 413 409 443 481 444 380 243 307 411 403 410 390 365 434 366 425 323 567 372 313 197 222 270 240 195 198 246 256 221 243 194 226 196 160 153  581 487 519 524 515 476 478 466 369 304 174 227 260 261 236 247 242 243 220 225 305 179 147 129 135 101 400 130 69 75 77 67 82 63 68 55 69 55 47  Hk-I Hk-w Lou1 Lou2 Tiz  947 647 609 529 541 511 538 573 668 196 665 687 629 473 487 478 473 423 445 410 434 503 485 429 517 603 683 655 666 654 874 731 720 783 830 739 771 802 721 795  3.1 2.4 2.4 2.6 2.4 2.5 2.5 2.3 2.6 2.5 2.6 3.0 2.6 2.7 2.9 2.6 2.7 2.7 2.7 2.7 2.8 2.9 2.8 2.9 2.9 3.3 3.2 3.9 3.3 3.2 3.2 3 3.1 3.1 3.1 3.1 3.1 3.1 3.3  3.8 2.7 2.4 2.6 2.4 2.4 2.6 2.4 2.5 2.5 2.6 2.9 2.6 2.7 2.8 2.6 2.6 2.6 2.6 2.5 2.6 2.8 2.7 2.8 2.8 3.0 2.8 2.9 3.1 3.1 3 2.9 2.9 3.1 3 3.1 3.1 3.1 3 3.2  4.3 3.6 3.4 3.5 3.5 3.4 3.6 3.4 3.4 3.4 3.4 3.8 3.4 3.4 3.3 3.4 3.3 3.3 3.3 3.2 3.3 3.4 3.3 3.4 3.5 4.1 3.5 3.7 3.7 3.9 3.7 3.6 3.7 4.1 3.7 3.8 3.8 3.7 3.9 4.0  217  4.5 3.9 3.6 3.8 3.8 3.8 3.9 3.6 3.7 4.4 3.7 5.6 3.5 3.7 3.6 3.6 3.6 3.7 3.6 3.6 3.7 3.8 3.7 3.8 4.0 3.6 3.8 4.1 4.2 4.2 4.2 3.8 4.0 4.5 4.0 4.1 4.3 4.1 4.3 4.2  4.1 3.5 3.3 3.4 3.3 3.4 3.3 3,2 3.2 3.2 3.2 3.7 3.2 3.2 3.1 3.1 3.1 3.2 3.2 3.1 3.1 3.2 3.2 3.2 3.3 3.6 3.3 3.5 3.6 3.6 3.7 3.5 3.5 3.8 3.6 3.5 3.6 3.6 3.7 3.7  Redox potential (mV) Zim Brw Hk-I Hk-w Lou1 Lou2 Tiz Zim Brwk 4.8 5.0 208 194 4.7 5.2 3.9 5.1 245 204 4.0 5.1 no data available 276 3.6 4.9 3.4 5.0 219 3.4 5.0 289 155 3.3 4.9 256 3.4 5.0 3.5 5.1 196 220 3.3 5.0 3.7 5.1 332 301 131 295 286 268 3.4 5.4 334 320 309 291 322 315 228 3.4 5.4 329 331 310 292 324 308 139 3.4 5.3 332 315 305 280 315 304 143 3.4 5.5 317 335 311 290 327 298 140 3.4 5.4 331 329 311 297 343 320 173 3.4 5.4 333 326 317 288 314 310 237 3.5 5.4 334 335 317 295 329 319 231 3.3 5.3 340 335 308 279 327 330 251 3.4 5.5 333 329 315 292 325 318 149 3.5 5.8 331 320 291 283 307 305 196 3.5 5.2 315 323 311 284 307 299 178 3.7 5.4 323 326 303 284 318 301 3.9 5.4 319 325 304 274 307 269 233 3.3 5.1 325 309 242 135 241 144 209 4.1 5.0 311 323 270 240 277 236 197 4.2 4.7 302 279 299 257 305 145 204 4.0 5.3 273 286 285 261 288 255 153 4.4 5.2 302 281 242 217 267 218 5.0 4.1 270 318 292 276 302 241 257 5.5 4.1 303 320 286 263 286 184 257 4.6 3.7 293 316 288 252 288 194 266 5.6 3.5 317 319 256 213 277 128 241 4.5 3.5 324 318 293 271 300 220 273 308 4.4 3.3 324 319 284 240 302 223 310 4.6 3.2 4.5 3.2 322 311 281 242 289 216 312 5.3 3.2 306 312 240 224 273 175 317 4.7 3.2 309 306 230 221 279 190 315  Table 1-1  Pyrite Leachate Chemistry  Days  Cycle con  Date  fuS)  0 3 6 10 14 17 20 23 27 30 34 38 41 44 55 62 65 69 72 76 80 83 86 90 94 97 104 107 111 111 114 118 121 125 128 132 136 141 143 146 Note:  Column Leach soln pH vol edo Fetot S04 Cu (ml) (mV) (mg/l)  07-Nov-97 1 233 6.7 228 10-Nov-97 2 14-Nov-97 3 340 5.5 245 18-Nov-97 4 221 7.8 213 21-Nov-97 5 288 6.8 212 24-Nov-97 6 277 6.2 225 27-Nov-97 7 258 6.4 205 01-Dec-97 8 238 6.2 202 04-Dec-97 9 240 6.2 203 08-Dec-97 10 247 6.3 217 12-Dec-97 11 235 6.4 206 15-Dec-97 12 188 6.6 202 18-Dec-97 13 163 6.6 202 29-Dec-97 14 157 6.6 195 05-Jan-98 15 295 5.9 205 08-Jan-98 16 287 6.1 210 17 181 6.3 194 12-Jan-98 15-Jan-98 18 219 6.1 204 19-Jan-98 19 169 6.4 198 23-Jan-98 20 170 5.7 200 26-Jan-98 21 151 6.2 200 22 154 6.4 205 29-Jan-98 02-Feb-98 23 138 6.1 197 24 163 5.7 181 06-Feb-98 09-Feb-98 25 163 5.7 181 12-Feb-98 26 147 5.7 185 16-Feb-98 27 143 6.0 184 19-Feb-98 28 143 5.9 171 23-Feb-98 29 130 6.4 226 26-Feb-98 30 163 6.3 205 02-Mar-98 31 152 6.1 191 05-Mar-98 32 166 6.0 190 09-Mar-98 33 160 5.9 204 12-Mar-98 34 170 5.8 214 16-Mar-98 35 165 6.2 192 20-Mar-98 36 166 5.8 168 37 168 5.8 190 25-Mar-98 38 169 6.1 187 27-Mar-98 30-Mar-98 39 165 6.2 187 02-Apr-98 40 166 6.2 185 empty spaces = no data available  0.31 0.10 0.42 0.10  90 209 231 122 137 174 161 231 159 157 180 165 165 217 148 167 193 157 203 206 123 171 164 161 160 204 198 142  0.04 0.01 0.00 0.23 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Zn  Ca Mg Na  K  (g'l) (mg/l (mg/l (mg/l (mg/l) (mg/l (mg/l)  <0.1 <0.1 <0.1 <0.1  31 30 26 23  3.6 1.7 1.5 1.3  26 24 22 20  9.5 9.0 8.5 7.4  0.12 <0.1 <0.1 0.08 <0.1 <0.1 0.09 <0.1 <0.1 <0.1 <0.1 0.25 <0.1 1.7 0.80 <0.1 <0.1 0.25 <0.1 <0.1 0.18 <0.1 <0.1 0.08 <0.1 <0.1 0.10 <0.1 <0.1 0.14 <0.1 <0.1 0.11 <0.1 <0.1 0.11 <0.1 <0.1 0.09 <0.1 <0.1 0.25 <0.1 <0.1 0.05 <0.1 <0.1 0.10 <0.1 <0.1 0.07 <0.1 <0.1 0.08 <0.1 <0.1 0.07 <0.1 <0.1 0.06 <0.1 <0.1 0.07 <0.1 <0.1 0.07 <0.1 <0.1 0.06 <0.1 <0.1 0.08 <0.1 <0.1 0.07 <0.1 <0.1 0.07 <0.1 <0.1 0.06 <0.1 <0.1 0.06 <0.1 <0.1 0.06 <0.1 <0.1 0.10 <0.1 <0.1 0.07 <0.1 <0.1 0.05 <0.1 <0.1 0.07 <0.1 <0.1 0.07 <0.1 <0.1  28 28 29 26 39 38 40 31 30 51 56 45 83 35 36 36 70 28 62 33 41 32 27 29 31 35 53 41 35 28 31 33 36 38 39  1.9 1.8 1.6 1.5 1.8 1.7 2.0 1.5 1.4 2.5 2.7 2.3 7.8 2.1 1.7 2.0 6.6 1.3 6.0 1.94 1.86 2.0 1.73 1.82 1.89 1.73 2.15 1.21 1.44 1.5 1.3 1.8 2.1 2.2 2.25  20 18 15 16 15 15 14 10 12 13 14 12 14 13 10 11 10 10 11 11 10 9.6 9.6 11 11 10 10 9.2 8.9 9.5 10 9.1 9.3 9 8.7  9.2 7.3 6.5 7.0 7.2 7.4 7.3 4.8 6.0 7.3 8.4 6.5 6.6 7.0 5.5 5.9 5.3 4.9 5.1 5.1 6.06 7.12 5.81 6.36 6.8 6.76 6.89 6.2 5.8 5.64 6.5 6.2 6.53 6.36 6.41  0.11 0.10 0.12 0.01  <0.1 <0.1 <0.1 <0.1  218  Distilled water pH cond Redox (uS)  5.4 5.6 2.1 5.3 1.6 5.5 5.6 19 5.6 1.8 5.5 4.6 5.6 2.6 5.5 10 5.6 16 5.5 28 5.4 21 5.4 26 5.5 27 5.5 17 5.5 8.3 5.4 7.3 5.6 9.2 5.4 9.4 5.4 10 5.5 4.5 5.7 8.7 5.5 6.1 5.3 4.6 5.5 6.4 5.5 6.4 5.6 9.3 5.6 9.5 5.5 18 5.6 39 5.3 6.9 5.3 5.8 5.6 5.4 5.5 8.8 5.6 48 5.4 4.1 5.3 16 5.2 3.4 5.6 14 5.6 14  (mV)  269 128 232 248 256 178 269 144 271 134 217 183 248 265 250 250 272 184 213 232 221 241 250 247 163 211 234 263 204 173  APPENDIX II - ANALYTICAL RESULTS: PASTE BACKFILL LEACHATE  219  Table 11-1  ay  Date  4 7 18 25 32 40 46 54 61 68 75 82 89 96 103 108 117 124 131 138  16-Dec-97 18-Dec-97 30-Dec-97 06-Jan-98 13-Jan-98 21-Jan-98 27-Jan-98 04-Feb-98 11-Feb-98 18-Feb-98 25-Feb-98 04-Mar-98 11-Mar-98 18-Mar-98 25-Mar-98 30-Mar-98 08-Apr-98 15-Apr-98 22-Apr-98 29-Apr-98  ay  Date  4 7 18 25 32 40 46 54 61 68 75 82 89 96 103 108 117 124 131 138  16-Dec-97 18-Dec-97 30-Dec-97 06-Jan-98 13-Jan-98 21-Jan-98 27-Jan-98 04-Feb-98 11-Feb-98 18-Feb-98 25-Feb-98 04-Mar-98 11-Mar-98 18-Mar-98 25-Mar-98 30-Mar-98 08-Apr-98 15-Apr-98 22-Apr-98 29-Apr-98  Paste Leachate Chemisty - Flooded Water Cells  ycl 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20  Distilled water data Paste leach water conductivity (uS) batch cond pH redox PH Louv Brw FIM Tiz Louv Brw FIM (uS) (mV) Tiz 1 28 5.5 499 3710 3670 3550 3100 11 7.4 7.7 7.8 17 5.6 422 3440 3450 3260 2720 11 6.6 6.7 6.7 2 3 33 5.6 429 3120 2940 2960 2520 11 7.8 7.2 8.3 4 6.1 5.6 509 2660 2540 2480 2230 10 8.5 8.1 8.5 5 18 5.5 469 2410 2310 1798 1920 11 8.5 8.7 8.9 6 5.8 5.58 442 1519 1668 1678 1475 11 7.8 7.5 8.4 7 6.3 5.61 488 1384 1403 1477 1225 11 8.3 8.1 8.3 16 5.61 466 1346 1496 1610 1185 11 8.5 9.2 9.2 8 9 3.3 5.4 459 1295 1398 1261 1091 11 9.4 9.2 9.4 5.5 5.4 555 1242 1111 1167 1031 11 9.4 9.1 9.5 10 11 39 5.6 443 1212 1082 1201 838 11 9.3 9.1 9.3 12 6.1 5.3 393 1256 1097 1218 811 11 9.0 8.9 9.5 7.8 5.3 475 1169 1005 1127 777 11 9.2 9.2 8.9 13 14 15 5.2 479 1101 982 1149 701 11 9.3 9.7 9.4 15 6.5 5.5 434 1080 952 1305 699 11 8.7 9.2 9.1 14 5.4 447 1067 937 1120 715 11 8.5 9.1 8.4 16 17 30.7 5.5 466 1113 1003 1221 737 9.8 8.3 8.8 8.5 18 65 5.6 452 1140 947 1228 732 10 7.9 8.4 8.5 19 47 5.6 469 1200 1026 1260 761 10 7.9 7.7 8.1 20 10.5 5.4 400 1033 890 1104 656 10 8.1 8.4 8.5  ycle 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20  Tiz <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  FPtotal(mg/l) Cu (mg/l) Louv Brw FIM Louv Brw FIM Tiz 0.74 <0.1 0.23 <0.1 <0.1 <0.1 <0.1 0.20 <0.1 0.5 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.20 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Tiz <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Zn (mg/l) Louv Brw <0.1 0.14 <0.1 0.17 <0.1 0.13 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  FIM <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Redox potential (mVSHEl  Tiz 286 311 303 280 257 290 261 276 304 280 288 257 213 234 236 233 260 252 243 228  Tiz 0.2 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Note: distilled water batch 1 put into cells Dec 12th '97, cycle 1 chemical data taken on Dec.16th '97  220  LOU  400 371 422 314 311 359 340 318 400 316 356 396 354 294 324 412 349 319 361 347  Brw 405 410 395 333 295 374 347 320 368 321 350 302 308 276 301 359 326 312 332 327  FIM 388 402 414 328 307 252 349 313 376 308 337 407 333 298 332 394 354 329 327 330  Pb (mg'l) Lou Brw FIM 0.1 0.1 <0.1 0.1 0.1 <0.1 0.1 0.1 <0.1 0.3 <0.1 0.2 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.2 <0.1 <0.1 <0.1 <0.1 0.1 0.2 <0.1 0.1 0.2 <0.1 0.1 0.2 <0.1 0.1 <0.1 <0.1 0.2 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Table l l - i (cont'd) ay  Date  ycle  4 7 18 25 32 40 46 54 61 68 75 82 89 96 103 108 117 124 131 138  16-Dec-97 18-Dec-97 30-Dec-97 06-Jan-98 13-Jan-98 21-Jan-98 27-Jan-98 04-Feb-98 11-Feb-98 18-Feb-98 25-Feb-98 04-Mar-98 11-Mar-98 18-Mar-98 25-Mar-98 30-Mar-98 08-Apr-98 15-Apr-98 22-Apr-98 29-Apr-98  ay  Date  ycle  4 7 18 25 32 40 46 54 61 68 75 82 89 96 103 108 117 124 131 138  16-Dec-97 18-Dec-97 30-Dec-97 06-Jan-98 13-Jan-98 21-Jan-98 27-Jan-98 04-Feb-98 11-Feb-98 18-Feb-98 25-Feb-98 04-Mar-98 11-Mar-98 18-Mar-98 25-Mar-98 30-Mar-98 08-Apr-98 15-Apr-98 22-Apr-98 29-Apr-98  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20  Paste Leachate Chemisty - Flooded Water Cells  S0 (mg/l) Ca (m g/i) Tiz Louv Brw FIM Tiz Louv Brw FIM 1.3 1.3 1.4 1.1 840 845 922 682 2.2 2.3 2.2 1.5 883 1.9 2.1 1.9 1.7 784 945 857 852 2.1 1.7 1.7 1.5 782 776 773 774 1.5 1.6 1.6 1.3 725 860 755 664 1.5 1.5 1.5 1.2 700 760 751 556 1.3 1.3 1.5 0.9 610 622 646 452 1.2 1.2 1.4 0.8 626 617 699 437 1.2 1.2 1.5 0.7 595 591 697 392 1.1 1.1 1.4 0.6 572 538 692 347 1.1 1.0 1.4 0.6 547 526 675 301 0.9 1.0 1.4 0.4 554 600 880 310 1.0 0.9 1.4 0.5 546 540 870 308 0.9 0.8 1.4 0.4 504 510 760 278 1.0 0.9 1.4 0.5 515 530 855 294 0.9 0.8 1.3 0.5 417 460 810 280 492 390 650 190 1.0 0.9 0.7 1.3 0.4 467 330 623 209 0.9 0.8 1.3 0.4 472 365 610 198 0.6 0.7 1.3 0.4 312 320 590 185 4  Mg (mg/l) K (mg/l) Tiz Louv Brw FiM Tiz Lou Brw FIM 3.6 47 11 4.8 127 51 107 32 4.0 127 19 8.3 158 85 136 53 80 14 5.0 146 74 117 45 2.8 2.7 59 10 4.5 138 60 99 41 4.7 41 7.0 2.8 117 45 66 34 5.5 30 6.4 2.0 78 36 44 28 1.3 22 5.0 4.5 50 27 32 22 1.2 16 4.4 1.9 41 25 25 20 1.4 13 4.1 1.9 29 21 17 19 1.2 10 3.6 1.7 20 17 12 16 1.2 11 3.6 1.5 16 15 10 15 12 3.2 1.0 10 12 6.0 10 0.9 1.0 11 3.2 1.0 8.3 9.2 6.2 10 1.0 9.0 2.7 0.8 7.2 8.2 5.1 9.3 0.9 13 3.2 1.5 5.7 7.9 4.8 10 0.7 12 3.0 1.7 4.8 7.1 4.7 10 1.3 8.4 2.4 1.3 4.1 5.0 5.3 9.0 0.7 12 2.4 1.8 4.6 6.5 5.0 11 0.6 14 2.3 1.9 4.4 6.0 5.1 10 0.8 12 2.0 1.6 5.5 5.2 4.3 9.1  Al (mg/l) Si (mg/1) Louv Brvv FIM Tiz Tiz Louv Brw FIM Tiz 2.2 9.7 1.9 2.4 < 0.30 < 0.30 < 0.30 < 0.30 <1.0 3.3 9.8 2.0 3.4 < 0.30 < 0.30 < 0.30 < 0.30 <1.0 2.9 9.7 2.2 3.2 < 0.30 < 0.30 < 0.30 < 0.30 <1.0 2.0 9.7 1.6 2.7 < 0.30 < 0.30 < 0.30 < 0.30 <1.0 <1.0 10 2.2 3.2 2.9 <1.0 3.0 9.8 1.9 3.0 <1.0 3.0 8.6 1.5 2.5 2.2 9.9 1.5 2.0 <0.3 <0.3 <0.3 <0.3 <1.0 2.4 10 1.2 2.6 <0.3 <0.3 < 0.3 <0.3 <1.0 2.2 9.7 1.0 2.1 <0.3 <0.3 <0.3 <0.3 <1.0 1.9 9.8 1.4 2.6 <0.3 <0.3 <0.3 <0.3 <1.0 1.5 6.8 0.5 1.6 0.32 < 0.3 <0.3 <0.3 <1.0 1.6 6.1 0.5 1.3 0.32 < 0.3 <0.3 <0.3 <1.0 1.7 6.7 0.5 1.5 0.34 < 0.3 <0.3 <0.3 <1.0 1.7 6.6 0.8 1.6 0.33 < 0.3 <0.3 <0.3 <1.0 1.5 6.6 0.9 2.0 0.63 < 0.3 <0.3 <0.3 <1.0 <0.3 <0.3 <0.3 <0.3 <1.0 0.8 1.6 8.9 2.0 4.1 <0.3 <0.3 <0.3 <0.3 <1.0 1.3 9.0 2.0 4.1 <0.3 <0.3 <0.3 <0.3 <1.0 2.5 8.6 1.7 3.6 <0.3 <0.3 <0.3 <0.3 <1.0  As (mg/l) IIP** Louv Brw FIM <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0  Note: distilled water batch 1 put into cells Dec 12th '97, cycle 1 chemical data taken on Dec. 16th '97  221  Table 11-2 Day 1 6 12 20 28 34 41 48 50 57 64 71 78 84 90 99 107 113 120 138 Day 1 6 12 20 28 34 41 48 50 57 64 71 78 84 90 99 107 113 120 138  Date 12-Dec-97 18-Dec-97 30-Dec-97 07-Jan-98 15-Jan-98 21-Jan-98 28-Jan-98 04-Feb-98 11-Febr98 18-Feb-98 25-Feb-98 04-Mar-98 11-Mar-98 18-Mar-98 24-Mar-98 30-Mar-98 08-Apr-98 16-Apr-98 22-Apr-98 29-Apr-98  Paste Leachate Chemisty - Cycled Water Cells Distilled water data Cycle bate cond PH cdox (mV) Tiz (uS) 1 1 27.5 5.5 499 1880 2 2 28.8 5.5 359 1000 3 3 26.8 5.6 485 1038 4 4 6.09 5.6 509 885 5 5 17.7 5.5 469 1192 6 5.83 5.58 442 692 6 7 7 6.3 5.61 488 935 4.5 5.52 778 8 8 4.0 5.32 498 795 9 9 779 10 10 17.9 5.42 11 11 34.9 5.64 432 803 12 12 6.57 5.34 404 810 13 13 14.9 5.2 479 684 14 14 2.0 5.2 464 758 15 15 8.8 5.55 481 721 15 5.59 490 745 16 16 17 17 55 5.62 411 820 18 18 48 5.6 470 863 19 19 4.62 5.3 389 785 767 20 20 18 5.59  Paste leach water conductivity (uS) pH Louv Brw FIM Tiz Louv Brw FIM 1660 1902 1565 10 8.2 7.6 8.4 1090 1330 900 9.3 7.5 7.6 7.4 1054 1268 892 7.3 6.6 6.5 6.3 846 1060 724 8.1 7.3 6.3 6.3 1104 1322 906 7.8 6.6 6.4 7.3 639 842 495 6.6 6.3 6.1 6.7 775 1025 633 6.8 6.9 6.5 6.8 609 860 491 6.4 6.8 6.5 6.7 635 838 504 6.7 7.1 6.8 7.1 601 813 476 7.6 7.7 7.3 7.7 647 928 528 7.4 7.5 7.2 7.4 596 818 509 7.2 7.5 7.1 7.5 459 669 400 7.0 7.4 7.1 7.6 540 758 459 6.7 7.6 7.1 7.3 496 708 432 6.7 6.8 6.6 6.8 520 724 745 6.9 7.5 7.0 7.5 625 847 693 6.9 7.3 7.0 7.5 623 801 676 6.9 7.2 6.9 7.7 567 765 773 6.9 7.3 7.5 7.6 592 749 628 7.3 7.4 7.4 7.8  Cu (mg/l) Fc total (mg/l) Tiz Louv Brw FIM Tiz Louv Brw FIM Tiz 12-Dec-97 1 <0.1 0.74 <0.1 0.23 <0.1 <0.1 <0.1 <0.1 <0.1 18-Dec-97 2 <0.1 0.5 <0.1 1.78 <0.1 <0.1 <0.1 <0.1 <0.1 30-Dec-97 3 <0.1 0.3 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 4 <0.1 0.7 <0.1 0.2 <0.1 <0.1 <0.1 <0.1 <0.1 07-Jan-98 15-Jan-98 5 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 21-Jan-98 6 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 7 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 28-Jan-98 04-Feb-98 8 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.1 0.1 <0.1 0.1 11-Feb-98 9 <0.1 <0.1 <0.1 <0.1 <0.1 0.1 0.1 <0.1 0.1 18-Feb-98 10 <0.1 <0.1 <0.1 <0.1 <0.1 0.1 <0.1 <0.1 0.1 25-Feb-98 11 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.12 12 <0.1 <0.1 <0.1 <0.1 <0.1 04-Mar-98 11-Mar-98 13 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.11 18-Mar-98 14 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.14 24-Mar-98 15 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.11 30-Mar-98 16 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.3 08-Apr-98 17 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 1.2 16-Apr-98 18 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.5 22-Apr-98 19 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.14 29-Apr-98 20 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.3 Note: distilled water put into cells for 24 hours prior to leachate analysis Date  Cycle  222  Zn (mg/l) Louv Brw <0.1 0.1 <0.1 0.3 <0.1 0.5 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.2 <0.1 0.3 0.1 0.4 0.1 0.2 0.1 0.6 <0.1 <0.1 <0.1 0.3 <0.1 0.3 <0.1 0.3 <0.1 0.5 <0.1 2.2 <0.1 1.4 <0.1 0.5 <0.1 1.3  FIM <0.1 0.1 0.2 <0.1 <0.1 <0.1 <0.1 0.1 0.1 0.1 0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Redox potential (mV SHE) Tiz Louv Brwk FIM 300 358 368 361 367 405 417 411 399 420 421 423 356 367 371 356 300 362 364 334 321 323 338 329 385 390 383 390 413 409 412 412 369 374 385 385 375 378 388 400 352 356 347 361 345 328 365 352 379 373 390 392 381 405 396 412 445 419 434 433 418 393 398 394 374 393 396 390 390 395 378 357 348 358 357 360 332 329 338 338  Tiz 0.2 0.2 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.7 0.3  Pb (mg/l) Louv Brw 0.1 0.1 0.1 0.1 0.1 0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 0.6 <0.1 0.42 <0.1  FIM <0.1 <0.1 <0.1 0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1 <0.1  Table 11-2 (cont'd) Day  Date  Cycle  1 6 12 20 28 34 41 48 50 57 64 71 78 84 90 99 107 113 120 138  12-Dec-97 18-Dec-97 30-Dec-97 07-Jan-98 15-Jan-98 21-Jan-98 28-Jan-98 04-Feb-98 11-Feb-98 18-Feb-98 25-Feb-98 04-Mar-98 11-Mar-98 18-Mar-98 24-Mar-98 30-Mar-98 08-Apr-98 16-Apr-98 22-Apr-98 29-Apr-98  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20  Day  Date  Cycle  12-Dec-97 18-Dec-97 30-Dec-97 07-Jan-98 15-Jan-98 21-Jan-98 28-Jan-98 04-Feb-98 11-Feb-98 18-Feb-98 25-Feb-98 04-Mar-98 11-Mar-98 18-Mar-98 24-Mar-98 30-Mar-98 08-Apr-98 16-Apr-98 22-Apr-98 29-Apr-98  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20  1 6 12 20 28 34 41 48 50 57 64 71 78 84 90 99 107 113 120 138  Paste Leachate Chemistry - Cycled Water Cells SO, (mg/l) Ca (mg/l) Tiz Louv Brw FJM Tiz Louv Brw FIM Tiz 1.3 1.3 1.4 1.1 840 845 922 682 3.6 0.9 0.8 1.0 0.7 475 543 621 530 3.0 0.6 0.7 0.9 0.8 318 333 431 332 2.6 0.5 0.5 0.8 0.4 232 225 370 204 2.0 0.8 0.7 1.1 0.6 340 261 481 276 3.0 0.4 0.3 0.5 0.2 167 214 238 111 4.5 0.4 0.3 0.6 0.3 208 200 265 123 4.8 0.4 0.3 0.4 0.2 210 148 257 120 2.2 0.4 0.3 0.5 0.2 230 178 296 125 2.4 0.3 0.3 0.4 0.2 225 150 250 116 2.3 0.4 0.3 0.6 0.2 255 174 329 129 2.3 0.4 0.3 0.5 0.2 340 170 353 120 2.1 0.4 0.3 0.5 0.2 340 164 325 160 2.8 0.4 0.3 0.5 0.2 335 178 320 158 2.3 0.4 0.2 0.5 0.2 285 170 310 160 3.0 0.5 0.3 0.5 0.3 354 190 350 270 2.8 271 190 260 210 2.4 0.5 0.3 0.5 0.3 0.6 0.4 260 152 232 158 2.7 0.4 0.3 0.4 0.3 233 124 195 168 1.8 0.6 0.3 0.4 0.2 249 150 233 142 2.0  Mg (mg/l) K (mg/l) Louv Brw FIM Tiz Louv Brw FIM 47 11 4.8 127 51 107 32 35 6.3 4.0 113 36 86 25 30 7.8 3.8 80 27 59 20 24 5.6 2.6 55 19 32 16 48 6.0 6.8 44 19 25 17 29 4.0 2.0 17 8.2 9.0 8.5 22 4.7 2.0 13 6.7 6.7 8.6 8.8 3.3 1.6 10 5.9 5.5 8.4 10 3.9 1.7 8.6 6.1 4.9 8.9 8.3 3.4 1.6 5.9 1.9 3.5 7.9 10 4.1 1.8 6.3 5.9 5.3 9.1 1.1 3.8 1.0 4.1 3.7 2.2 5.7 9.0 2.9 1.9 3.5 3.0 1.9 5.0 9.0 2.2 1.5 3.3 3.2 1.8 5.1 8.7 2.5 1.3 3.0 3.2 1.9 5.1 12 3.6 2.1 3.0 3.6 2.0 8.6 3.1 1.7 3.0 5.0 7.4 13 6.4 12 2.0 1.8 4.4 4.8 4.7 8.1 9.2 2.3 1.4 1.7 2.4 1.4 5.2 11 3.2 1.3 2.4 3.3 2.4 5.0  As (mg/l) Si (mg/l) Al (mg/l) Tiz Louv Brw FIM Tiz Louv Brw FIM Tiz Louv Brw 2.2 9.7 1.9 2.4 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 2.0 7.1 2.0 2.9 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 1.5 5.0 1.8 2.8 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 <0.5 4.5 <0.5 1.4 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 <0.5 6.1 1.3 2.8 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 <0.5 4.4 <0.5 1.1 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 <0.5 4.2 <0.5 1.7 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 0.8 5.4 1.2 1.9 <0.3 0.5 <0.3 <0.3 <1.0 <1.0 <1.0 1.0 3.7 1.4 1.7 <0.3 0.3 <0.3 <0.3 <1.0 <1.0 <1.0 1.4 4.0 1.4 2.0 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 2.5 5.7 2.1 4.2 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 1.4 3.8 1.0 1.6 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 1.2 3.2 0.9 1.3 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 1.4 4.0 0.9 1.6 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 1.6 4.1 1.0 1.9 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 2.0 5.1 1.4 3.0 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 2.3 5.2 1.2 1.4 <0.3 <0.3 <0.3 < 0.3 <1.0 <1.0 <1.0 5.3 9.7 4.8 6.0 <0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0 2.0 4.8 1.7 2.9 <0.3 <0.3 < 0.3 <0.3 <1.0 <1.0 <1.0 3.6 7.3 3.0 4.2 < 0.3 <0.3 <0.3 <0.3 <1.0 <1.0 <1.0  223  FIM <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0 <1.0  Table 11-3  Paste Leachate Chemisty - Fe (S0 ) Solution Cells  Day  Paste leach solution Fe (S0 ) Solution Cycle bate cond pH OR Feto SO, conductivity (uS)  2  2  0.5  Date  4  4  3  Redox potential (mV SHE)  3  no.  (uS)  (mV (mg/l) (g/l)  12-Dec-97  1  1  2735  2.9 587  920 1 5  Tiz  Louv  Brw  FIM  pH Tiz Lou Brw FIM 6.3  6.6  6.2  6.1  6.0 181  6.7  6.2  6.0 228 239 199 252 5.9 234 223 226 327 6.1 278 304 243 182 5.6 290 374 258 360  6  17-Dec-97  2  2  2360  2.3 596  29-Dec-97  2610 1836  826 1 5 3240 1890 3020 1800 754 1 4 1855 1605 1623 1432  05-Jan-98  3 4  2.6 623  25  3 4  2.6 624  780  1 4  1472  1362  1339  1315  6.3  6.3  6.3 6.2  28 32  08-Jan-98 12-Jan-98  5 6 7  1967 1977  740 734 1 5  1338 1391  1311 1328  1268 1363  1350 1337  5.9 5.7  5.7 5.6  5.9 5.9  1 6 1407 1335 1258 1314 1331 1298 1275 1488 1418  1350 1307  5.2  1298 1351  5.3 5.2  5.3 5.2 5.1  5.5 5.7 5.6  4.1  3.9  1453 1349  1232  4.0  3.6  5.0 4.4  1512 1448 1487 1434  1399 1316 1304  4.5 4.4  4.3 4.0 3.8  5.3 4.9 4.1  3.9  4.2 4.4 4.1  35  15-Jan-98  1929  2.6 625 2.6 620 2.4 669  39 43  19-Jan-98 23-Jan-98 26-Jan-98  8 9 10  8 9  1942 2420  2.6 629 2.6 623  10  2320  2.5 626  2370 1979  2.6 626 2.5 624 2.5 623  46 49  29-Jan-98  11  11  53 57 60  02-Feb-98 06-Feb-98 09-Feb-98  12 13 14  12 13 14  63 67  12-Feb-98  15  16-Feb-98 19-Feb-98 23-Feb-98  16 17  15 16 17  1911 1894 2460  18  18 19 20 21 22  2510 1896  70 74 77  2700 2400  689  726 1 4 1318 772 1 7 1320 692 1 6 1496 723 1 3 1448  723 1 5 1424 687 1 6 1474 638 1 3 1502  2.5 630 2.5 621  1517 1552 679 1 .4 1447 1397 1439 548 1 .3 1439 1420 1390 665 1 .4 1494 1490 1504  2230  2.6 630  611  1 .3 1556  1460  2410 2330  2.6 624 2.7 612 2.5 628  609 645 483 695  1 .4 1426 1 .4  1380 1400 1375 1354  1 .3 1325 1 .5 1462  1320 1431  2.5 625 2.7 617  1219  3.9 4.0  1272 1320  4.1 3.9  1503 1328  4.0  1225 1232 1169  81 84  26-Feb-98 02-Mar-98 05-Mar-98  88  09-Mar-98  19 20 21 22  91  23 24  23  2550  2.5 628 2.5 621  95 99  12-Mar-98 16-Mar-98  24  2500  2.6 619  20-Mar-98  25  25  2.7 622  103  24-Mar-98  26  109  30-Mar-98  27  26 27  2460 2480 2540  2.5 623  1 .5 1536 564 1 .5 1407 1355 1391 1143 604 1 .5 1466 1377 1395 1199 630 1 .4 1396 1364 1375 1143  112  02-Apr-98  28  641  06-Apr-98  29  1951 2360  2.5 627  116  28 29  118  08-Apr-98  30  30  1934  2.5 616  123  13-Apr-98  31 32  2.5 615  16-Apr-98  31 32  2370  126  1901  2.6 624  721  130  20-Apr-98  33  33  2390  725 1 .6  133 137  23-Apr-98 27-Apr-98  34  34  35  35  2580 1937  2.5 621 2.5 617  140  30-Apr-98  36  36  2530  144 0 4 - M a y - 9 8  37  37  2570  38  2410  4.3 4.0  3.9 4.0 3.8 4.0  3.9 4.1 3.8 4.1  4.5 3.9 4.4  3.7  3.7 3.8  3.9 4.0  3.8  3.9  4.1  4.1  4.4  3.9  3.9  4.1  3.7  3.6  3.8  3.8 3.5  3.8  4.0  3.5  3.6  3.9 3.7  3.7  4.0  3.6  3.8  1320  3.9  3.7  4.0  3.6 3.8  3.4  3.6 3.7  2.5 617  745 1 .5 1405 1399 1417 1275 727 1 .5 1564 1539 1596 1310 910 1 .4 1584 1565 1576 1374  2.5 614  550 470  2.5 621  2.5 627  2.5 615  11-May-98  39  39  2430  154 14-May-98  40  40  2470  2.5 624  151  38  598  3.9 3.9 4.0  3.9 4.1  2.6 625 2.5 622  147 0 7 - M a y - 9 8  550 1 .5 1374  1361 1457 1233 1353 1387 1280 1503 1555 1290  1 .4 1458 1390 520 1 .4 1523 1479 630 1 .5 1465 1438  1420  1489  1461  1478  1535 1429  1188 1272 1216 1302  1 .6 1454  1434  1418  Lou Br.v FIM  6.4  18  5 6 7  Tiz  6.5 6.2  154 155 169  5.1 412 409 298 5.8 341 329 278 5.7 302 299 239 5.1 433 470 317  449 265 235 331  5.2 428 508 372 306 5.7 400 419 365 5.5 405 447 381 5.1 469 495 451 5.4 433 459 420  289 289 370 294  5.6 435 457 412 311 5.2 443 439 430 344 5.5 436 432 408 289 5.2 435 428 421 322 5.6 440 431 409 302 5.0 463 477 450 389 5.6 434 419 394 287 5.3 466 484 442 399 5.1 439 452 437 345 5.5 433 444 419 291 5.8 426 427 404 255 5.6 430 439 405 254 5.3 450 457 444 304 5.6 450 451 398 281 5.1 473 488 460 312 5.6 428 451 408 277 5.6 4.9 432 4 4 5 411  333  1583  1280  3.7  3.5  3.7  4.2 469 512 466 395 5.2 443 489 448 330 4.9 462 506 482 342 5.3 4 5 5 494 470 306  1 .4 1608 1543 1636 840 1 .7 1615 1587 1596 725 1 .4 1493 1466 1479  1411  3.6  3.4  1320  3.7  3.5  3.5 3.7  4.6 459 493 467 357 5.1 454 481 444 307  1305  3.5  3.4  3.6  4.7 465 479 455 369  1 .4 1565 1542  Notes: Empty s p a c e s denote unavailable results O R P : Oxidation-reduction (redox) potential (SHE)  224  3.6  3.5 3.4  3.6  Table 11-3 (cont'd) Day 0.5 6 18 25 28 32 35 39 43 46 49 53 57 60 63 67 70 74 77 81 84 88 91 95 99 103 109 112 116 118 123 126 130 133 137 140 144 147 151 154  Date  Cycle  12-Dec-97 17-Dec-97 29-Dec-97 05-Jan-98 08-Jan-98 12-Jan-98 15-Jan-98 19-Jan-98 23-Jan-98 26-Jan-98 29-Jan-98 02-Feb-98 06-Feb-98 09-Feb-98 12-Feb-98 16-Feb-98 19-Feb-98 23-Feb-98 26-Feb-98 02-Mar-98 05-Mar-98 09-Mar-98 12-Mar-98 16-Mar-98 20-Mar-98 24-Mar-98 30-Mar-98 02-Apr-98 06-Apr-98 08-Apr-98 13-Apr-98 16-Apr-98 20-Apr-98 23-Apr-98 27-Apr-98 30-Apr-98 04-May-98 07-May-98 11-May-98 14-May-98  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  Paste Leachate Chemisty - F e ( S 0 ) Solution Cells 2  Cu (mg/l) Fetotal(mg/l) Tiz Louv SWmBMi Tiz Lou Brw FllWiil 260 400 328 348 <0.1 0.1 <0.1 <0.1 94 280 154 169 <0.1 0.1 <0.1 <0.1 73 75 74 138 <0.1 0.1 <0.1 <0.1 88 216 190 230 <0.1 0.2 <0.1 <0.1 290 329 290 155 0.6 0.5 0.4 0.4 224 290 279 249 0.6 0.5 0.3 0.4 183 201 201 134 0.6 0.7 0.3 0.4 248 368 304 261 0.6 0.3 0.4 0.4 278 397 323 271 0.3 0.4 0.3 0.3 340 417 366 371 0.5 0.4 0.4 0.5 388 438 388 335 0.4 0.9 0.3 0.4 346 399 357 267 0.4 0.4 0.5 0.6 333 451 357 321 <0.1 0.1 0.2 <0.1 438 443 379 339 0.2 0.1 0.1 <0.1 417 459 378 361 0.1 0.3 0.1 <0.1 0.1 0.3 0.1 406 421 336 424 492 359 343 0.1 0.3 0.1 <0.1 343 436 345 304 <0.1 0.2 0.1 <0.1 402 479 355 356 <0.1 0.2 0.1 <0.1 329 387 290 236 0.2 0.2 0.2 0.2 490 411 270 260 0.1 0.1 0.2 0.2 438 407 230 221 0.1 0.1 0.2 0.2 518 510 393 250 0.2 0.1 0.3 0.4 502 384 310 305 0.1 0.3 0.2 0.2 476 370 295 260 <0.1 0.1 0.2 0.2 348 420 254 166 <0.1 0.2 0.1 0.3 416 417 360 180 <0.1 0.2 0.2 0.4 448 497 350 215 <0.1 0.2 0.2 0.3 433 370 325 312 <0.1 0.1 0.2 0.2 549 502 390 310 <0.1 0.3 0.2 0.1 552 594 605 423 0.3 0.1 0.3 0.5 530 494 900 370 0.3 0.2 0.7 0.4 530 935 800 395 0.2 0.3 0.7 0.4 630 1180 640 530 0.0 0.3 0.4 0.3 570 605 554 580 0.2 0.2 0.4 0.2 700 690 498 520 0.2 0.3 0.3 0.1 720 730 610 474 0.1 0.3 0.3 <0.1 490 485 470 330 <0.1 0.1 <0.1 <0.1 555 480 433 380 <0.1 0.1 <0.1 <0.1 472 430 460 393 <0.1 0.1 <0.1 <0.1  225  4  3  Zn (mg/l) 1.3 0.6 0.4 0.9 3.3 3.1 5.2 4.2 3.6 4.3 4.1 4.1 3.2 3.6 3.3 3.8 3.2 3.2 3.2 3.6 4.0 5.0 3.8 4.7 4.7 4.7 4.5 4.7  6.9 6.2 6.7 5.9 6.6 7.1 9.5 7.7 12 8.9  0.6 0.4 3.8 0.5 1.4 1.2 1.8 1.7 1.3 1.5 1.8 1.5 1.4 1.2 1.5 1.6 1.4 1.5 1.4 1.3 1.5 1.6 1.3 1.3 1.3 1.7 1.8 1.8 1.6 1.0 1.4 1.3 3.0 2.5 1.8 1.6 2.1 1.4 1.7 1.4  11 8.6 4.7 9.1 17 18 23 18 18 19 18 17 19 17 15 16 14 15 13 14 12 9.7 15 12  16 11 16 22 14 14 14 11 13 13 13 12  Pb(mg/I) Tiz LOU Brw FIM,, 1.3 0.5 0.1 <0.1 <0.1 1.0 <0.1 <0.1 0.1 <0.1 0.8 <0.1 <0.1 0.2 <0.1 1.1 <0.1 <0.1 <0.1 <0.1 1.3 <0.1 0.2 <0.1 <0.1 1.6 <0.1 0.2 0.2 <0.1 2.1 0.6 0.2 0.3 <0.1 1.8 0.4 0.2 0.2 <0.1 2.0 0.3 <0.1 0.3 <0.1 2.1 0.8 <0.1 0.9 <0.1 2.1 0.9 <0.1 1.4 <0.1 1.7 0.8 <0.1 0.7 <0.1 1.8 0.6 <0.1 1.2 <0.1 2.0 1.0 <0.1 1.6 <0.1 1.7 0.9 <0.1 1.4 <0.1 0.8 0.1 1.0 1.8 1.0 0.1 1.4 <0.1 1.5 0.8 <0.1 1.1 <0.1 1.8 1.1 <0.1 1.2 <0.1 1.7 0.9 <0.1 0.9 0.1 1.6 0.3 1.3 1.0 0.3 0.6 1.5 0.3 1.7 1.2 0.1 1.0 1.0 0.2 0.9 0.3 1.1 0.2 1.6 0.2 2.8 1.8 <0.1 <0.1 <0.1 2.9 2.3 <0.1 <0.1 <0.1 2.3 1.4 0.2 1.5 0.2 2.7 1.9 0.5 3.3 0.2 2.2 1.3 0.6 1.2 0.2 2.3 1.9 0.7 2.1 0.2 2.1 1.4 0.3 1.5 0.2 2.3 2.1 0.3 1.4 0.2 2.4 2.0 0.3 1.5 0.2 3.0 1.4 1.0 <0.1 2.2 1.2 <0.1 <0.1 <0.1 3.1 1.3 <0.1 <0.1 <0.1  Table 11-3 (cont'd) Days  Date  Paste Leachate Chemisty - Fe (S0 ) Solution Cells 2  SO, (mg/l)  Cycle Tiz  0.5 6  12-Dec-97 17-Dec-97  1 2  1.9 2.1  LOUV Brw F i M  1.9 2.5  2.0 2.1  1.7 2.0  586 756  Louv Brw F I M . 498 691  590 742  3  Mg (mg/l)  Ca (mg/l) Tiz  4  525 736  Tiz  K (mg/l)  Louv Brw FIMj 63  8.0 142 39.3  104  8.0  108  10 177 55.3  122  36  51  28 20  18  29-Dec-97  3  1.8  2.0  1.5  1.6  835  773  770  631  8.0  115  8.3  25  05-Jan-98  4  1.6  1.7  1.5  1.5  761  581  800  546  8.0  61  6.4  28  08-Jan-98  5  1.6  1.6  1.6  1.6  562  390  477  583  27  7.1  8.8  45  8.9  7.5  120  32  46  18.2  20  6.3  5.8  14 9.2  11  32  12-Jan-98  6  1.8  1.6  1.6  1.5  611  355  561  440  9.7  48  11  6.3  6.3  4.3  4.4  35  15-Jan-98  7  1.6  1.5  1.5  1.4  543  381  557  433  14  72  16  8.5  4.4  4.0  3.6  8.0  39  19-Jan-98  8  1.6  1.6  1.5  1.4  494  401  484  411  9.3  57  9.7  7.0  2.6  3.7  2.1  6.2  43  23-Jan-98  9  1.7  1.5  1.6  1.7  559  350  522  522  9.2  47  9.6  8.0  2.4  3.0  2.1  7.0  46  26-Jan-98  10  1.5  1.4  1.5  1.6  446  241  435  416  8.3  35  8.3  5.5  1.4  1.9  1.5  4.8  49  29-Jan-98  11  1.6  1.4  1.7  1.4  419  211  443  394  8.7  39  9.8  6.4  1.5  2.3  1.8  4.3  53  02-Feb-98  12  1.7  1.7  1.8  1.5  472  275  461  466  9.7  50  11  7.6  1.3  2.4  1.5  3.9  57  06-Feb-98  13  1.5  1.5  1.5  1.4  420  277  529  429  9.0  42  11  6.5  1.4  2.4  1.9  3.7  60  09-Feb-98  14  1.4  1.3  1.4  1.4  353  209  444  385  9.4  34  9.7  6.4  1.0  1.9  1.4  3.8  63  12-Feb-98  15  1.4  1.5  1.3  1.5  344  244  427  395  9.9  43  9.9  6.4  1.0  2.6  1.5  3.7  67  16-Feb-98  16  1.6  1.3  1.4  393  271  442  12  51  11  1.0  1.9  1.4  70  19-Feb-98  17  1.4  1.3  1.2  1.4  307  237  381  367  9.3  45  8.6  6.2  0.9  2.1  1.5  3.6  74  23-Feb-98  18  1.2  1.4  1.3  1.3  330  266  443  372  11  56  11  6.7  0.9  2.1  1.5  3.0  77  26-Feb-98  19  1.3  1.3  1.2  1.5  305  229  337  372  10  49  8.4  6.4  1.0  2.2  1.8  3.4  1.6  1.6  1.3  356  220  407  343  12  45  9.0  5.9  81  02-Mar-98  20  1.5  84  05-Mar-98  21  1.4  1.5  1.2  285  180  280  10  37  7.0  1.2  2.1  2.6  3.0  22  1.6  1.6  1.2  377  222  294  15  69  8.6  1.2  2.4  3.1  3.0  1.8  1.4  244  113  340  8.2  25  9.0  1.0  1.8  2.4  1.0  328  213  320  13  48  8.8  1.0  2.1  2.2  2.0  357  218  14  56  0.9  1.8  1.0  2.0  320  16  79  0.9  2.5  1.0  2.0  254  13  62  0.8  2.0  1.0  2.0  180  11  42  0.8  1.7  1.0  2.0  0.8  2.0  1.4  2.9  2.5  1.7  2.4  3.8 3.5  88 91  09-Mar-98 12-Mar-98  23  1.5  95  16-Mar-98  24  1.5  1.3  1.4  99  20-Mar-98  25  1.6  1.2  1.4  103 109 112 116 118  24-Mar-98 30-Mar-98 02-Apr-98 06-Apr-98 08-Apr-98  26 27 28 29 30  1.7 1.6 1.6 1.7 1.5  1.8  447  1.6  379  1.8  297  1.6 1.4 1.8  1.8 1.4  1.6 1.7  364 225  169 107  420 230  58  410  32  285  14 7.3  9.0 7.0  123  13-Apr-98  31  1.8  1.7  2.0  1.8  355  180  440  460  13  63  11  7.8  1.6  2.2  2.4  126  16-Apr-98  32  1.6  1.3  3.0  1.6  279  134  560  420  10  40  15  9.6  1.7  2.2  3.8  3.2  130  20-Apr-98  33  1.5  2.8  2.7  1.6  320  270  480  380  12  93  15  8.6  1.2  2.7  2.8  2.9  133  23-Apr-98  34  1.5  2.7  1.7  1.5  236  207  290  280  7.9  70  7.9  6.1  1.0  2.1  1.5  2.0  137  27-Apr-98  35  1.6  1.6  1.7  1.7  284  152  327  333  11  50  9.5  7.0  0.8  1.2  1.3  2.1  140  30-Apr-98  36  1.8  1.7  1.5  1.5  280  133  255  294  10  49  6.6  5.8  0.5  1.2  1.0  1.6  144  04-May-98  37  2.0  1.9  1.5  1.6  340  154  320  348  14  64  8.6  6.7  0.6  1.1  0.9  1.7  147  07-May-98  38  1.6  1.3  1.6  1.4  314  140  321  443  8.0  45  11  9.0  0.5  1.6  1.2  2.0  151  11-May-98  39  1.7  1.5  1.7  1.6  466  183  355  400  17  65  13  9.0  0.6  2.0  0.9  1.6  1.6  1.4  1.6  1.5  296  132  290  350  10  43  11  9.0  0.5  1.7  0.9  1.8  154  14-May-98  40  226  Table 11-3 (cont'd) Days  Date  Paste Leachate Chemisty - F e ( S 0 ) Solution Cells 2  Cycle  S i (mg/l) hz  0.5 6 18 25 28 32 35 39 43 46 49 53 57 60; . 63 67 70 74 77 81 84 88 91 95 99 103 109 112 116 118 123 126 130 133 137 140 144 147 151 154 ;  1  12-Dec-97 17-Dec-97 29-Dec-97 05-Jan-98 08-Jan-98 12-Jan-98 15-Jan-98 19-Jan-98 23-Jan-98 26-Jan-98 29-Jan-98 02-Feb-98 06-Feb-98 09-Feb-98 12-Feb-98 16-Feb-98 19-Feb-98 23-Feb-98 26-Feb-98 02-Mar-98 05-Mar-98 09-Mar-98 12-Mar-98 16-Mar-98 20-Mar-98 24-Mar-98 30-Mar-98 02-Apr-98 06-Apr-98 08-Apr-98 13-Apr-98 16-Apr-98 20-Apr-98 23-Apr-98 27-Apr-98 30-Apr-98 04-May-98 07-May-98 11-May-98 14-May-98  1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40  7.0 6.6 5.4 7.0 9.6 9.7 15 12 12 12 13 13 12 13 13 14 13 12 14 13 13 13 11 12 12 12 13 12 15 17 17 19 17 15 15 16 18 15 19 15  1  ouv Brw  12 14 14 12 17 17 23 23 21 19 20 19 20 18 19 19 20 20 20 17 16 17 15 12 13 16 18 17  19 20 36 33 18 18 20 15 16 14  9.8 8.9 12 8.9 10 11 10 10 12 12 11 10 10 10 10 9.7 8.6 8.5 8.3 10  15 29 22 16 14 13 13 12 11 11  FIM 6.2 6.9 6.2 5.9 9.4 8.0 9.3 8.0 8.6 7.7 8.3 9.6 8.6 8.4 8.1 8.6 8.3 10 8.0  15 16 15 12 13 12 13 13 10 11  Tiz  A l (mg'l) Louv brw  <0.3 <0.3 <0.3 <0.3 <0.5 <0.5 1.2 0.6 <0.5 0.6 0.8 0.7 0.7 1.4 1.0 1.4 1.0 1.1 <0.3 <0.3 1.8 1.2 2.3 1-7 1.1 0.5 0.7 1.7 <0.3 <0.3 1.1 1.6 1.3 2.3 1.4 2.0 1.8 1.8 2.1 2.0  227  <0.3 <0.3 <0.3 <0.3 <0.5 <0.5 1.82 <0.5 <0.5 1.63 <0.5 1.1 1.5 2.3 2.0 2.5 1.7 1.7 1.6 <0.3 1.6 0.9 2.1 1.3 0.8 0.7 0.9 1.4 <0.3 <0.3 1.1 1.3 2.1 1.4 2.1 2.4 2.0 2.0 1.5 1.7  <0.3 <0.3 <0.3 <0.3 <0.5 <0.5 <0.5 <0.5 <0.5 <0.5 <0.5 <0.5 0.9 0.9 0.8 0.9 1.0 0.8 0.8 <0.3 0.5 0.7 1.2 0.7 1.0 <0.3 <0.3 1.0 <0.3 <0.3 1.0 2.1 1.5 2.1 1.4 1.3 1.1 1.4 0.9 1.3  4  3  As (mg/l)  FIM <0.3 <0.3 <0.3 <0.3 2.7 0.7 <0.5 <0.5 <0.5 <0.5 <0.5 <0.5 <0.3 <0.3 <0.3 0.5 0.6 <0.3 <0.3 1.0 <0.3 1.0 1.0 <0.3 <0.3 <0.3 <0.3 <0.3 <0.3 <0.3 <0.3 0.3 1.6 1.0 0.9 0.5 0.4 <0.3 0.7  •Tiz-, Louv Brw  <1 <1 <1  <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1  <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1  FIM <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1  <1 <1 <1  <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1 <1  <1 <1  <1 <1  1  1.2 <1 1.0 2.0 1.4 <1 1.3 2.0  <1  <1 <1  <1  <1  <1  <1  1 <1 <1 <1 <1  <1  1 <1 <1 <1  <1  <1  <1 <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1  <1 <1  <1  <1  <1  <1  <1 <1  <1  APPENDIX I I I - ANALYTICAL RESULTS: PASTE BACKFILL SOLID PHASE  228  to CD o> NT  CCh  o o o o o o o o o o o o o o o o o o o o o o o o o o o o o o o • »o o o •o o o o - o o o o o o o o o o o o o •• •o o • o o i n o o o m i n o o o i n ••o o •o o o o •o o o o o o • •o •  CL -  onodon TH r-i tH  OOOCNOLOOOOOOOOOOOOOOOOOOO Hriri HHHHHHHdHHdririHririOOriO  rHH H  z O NHOIflrtrllflriHHHHHlflHHHOrtririHOnHlflHHHriH •OH • O • OOO O iH O * H OOOOOOO  o  Q LU  CC Q_l <  Q  .  .  . .  OOO  oioioioioioioioii  o  O  •  o cc  < z  l i  < cc  o< <  CD yj cc UJ  sis  £ w  < >  COLLLUo U-OCtS 0 |  •o  O  .  O  O  OOOOOOO  rr  o  Z  o  C0 OJ  CO CO LU O  X  •  O  Z  -2~-  C L U COLLI  •  •  •  •  •  •  •  •  i • 9! •  ; 01 01 01 01  01  •  •  •  I  I  oi oi oi w oi « *g I l l l l O O O O>  I  r^HHHHHHHHHHHHHHHHHSHHHHHaa U UO i  CL _l  <  o  fc —I  QCC  < z <  O CO  yjCCtogN ZLLI t no  <5  m>>  O  ^? O rH u o H to  O  H  ^? a? oj  OJ OJ  n-CO  < zco 3  inLnininLninLnMLfltALfltAiAinifliAtnLniAiAiAUtiAuinm^onrori  2Q LLJO  03  X CD  *3 C0 "U  L  (D OJ CM CO CO  rajo o> ja * J2-i§ 9 o0  6  CO O)  <  3 0 • U 00  < m  | ° u J O  o  t- ^- c x  O  e J3 >g  CM 00  — m -c "Z.  1  a  I  CMCCiO-  CO  LU I<  O  LL I-  cc LU  o  H o a  cc m LL  43  O  CO CO CM  | CO cc  DC  LU >  z 3 O * CC 'p"<  o a 43 -H  •o a o ii a 43  . M U 0  Pi n o a u  01 EH  a b> H i  z  z o  < cc < a.  6  iOrr  I0S  E  o  143 Q  LU  cc CL LU  -1 0.  u g •rl 0 •0 44 a to t> 9 •H ) 0 0 'H a o -a s ca to i U  S ft  2 « H U  NUMBER SAMPLES  CC  CD _ sz CC  SA  C0  5°  co  ^ i—  CJSUJ SQ LLJO-  5"  116  m m rH rH  in in CN 00 CN CN  

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