Ubc 2011 Spring Drozdiak Jeffrey

A Pilot-Scale Examination of a Novel High Pressure Grinding Roll / Stirred Mill Comminution Circuit for Hard-Rock Mining Applications

by

Jeffrey Adam Drozdiak

B.A.Sc., The University of British Columbia, 2006

A THESIS SUBMITTED IN PARTIAL FULFILLMENT OF

THE REQUIREMENTS FOR THE DEGREE OF

MASTER OF APPLIED SCIENCE

in

The Faculty of Graduate Studies

(Mining Engineering)

THE UNIVERSITY OF BRITISH COLUMBIA

(Vancouver)

April 2011

© Jeffrey Adam Drozdiak, 2011

ii

Abstract

The mining industry will be faced with new challenges as the need to develop lower grade ore

deposits expands to meet the rising demand for raw resources. Low-grade deposits require a

substantially increased tonnage to achieve adequate metal production and have caused the

consumption of energy in mining practices such as comminution to rise dramatically. If

improvements could be made in the processes employed for metal extraction, the mining

industry could remain sustainable for future generations. This research focused on the

development of a novel comminution circuit design to addresses these issues. The circuit

design incorporated two, known energy efficient technologies, the High Pressure Grinding Roll

(HPGR) and the horizontal high-speed stirred mill, and examined the technical feasibility of a

circuit operating without the need for a tumbling mill.

The main objectives of this research were to setup pilot-scale research equipment and develop

the design criteria necessary to operate an HPGR / stirred mill circuit. Testing consisted of

using a copper-nickel sulphide ore from Teck Limited’s Mesaba deposit to evaluate a circuit

comprised of two stages of HPGR comminution followed by stirred mill grinding. To evaluate

the potential energy benefits of this novel circuit arrangement, energy consumption related to

comminution was calculated for the circuit using power draw readings off the main motor and

the throughput recorded during testing. To provide a basis for comparison, the energy

requirements for two conventional circuits, a cone crusher / ball mill and an HPGR / ball mill,

were determined through HPGR pilot-scale testing, Bond grindability testing and JK SimMet®

flowsheet simulation.

Results from this research showed that operating the first-stage HPGR in open circuit and the

second stage in closed circuit with a 710µm screen, resulted in a circuit energy requirement of

14.85kWh/t, a reduction of 9.2 and 16.7% over the HPGR / ball mill and cone crusher / ball mill

circuits, respectively. To assist in future HPGR / stirred mill studies, a refined testing procedure

was developed with a reduced sample commitment and the ability to perform an energy

comparison with a Semi-Autogenous Grinding (SAG) mill / ball mill circuit.

iii

Preface

The work presented in the following thesis was performed solely by the author. Some of the

information provided in Chapter 4 was published in the following paper:

Drozdiak, J., Nadolski, S., Bamber, A., Klein, B., & Wilson, S. (2010). A Comparison of the

energy requirements of an HPGR / stirred mill circuit and conventional grinding circuits

for the comminution of mesaba ore. 42nd Annual Meeting of the Canadian Mineral

Processors, Ottawa, ON, Canada.

The co-authors made contributions regarding the structure and layout of the paper and all

experimental testing was carried out by the principle author.

My research committee, consisting of Dr. Bern Klein, Dr. Andrew Bamber, Dr. Marek Pawlik,

Josh Rubenstein, and Mike Larson, provided input into design of the test program and provided

advice on the research project.

Stefan Nadolski, of Koeppern Machinery Australia, provided assistance during HPGR pilot-scale

testwork and provided a summary for HPGR operating data.

iv

Table of Contents

Abstract ......................................................................................................................................... ii

Preface ......................................................................................................................................... iii

Table of Contents ......................................................................................................................... iv

List of Tables ................................................................................................................................ vi

List of Figures ............................................................................................................................. viii

Acknowledgements ..................................................................................................................... xii

1 Introduction ............................................................................................................................ 1

2 Literature Review ................................................................................................................... 4

2.1 High Pressure Grinding Rolls ......................................................................................... 5

2.1.1 Background ............................................................................................................. 5 2.1.2 Technology Overview .............................................................................................. 6 2.1.3 HPGR Operating Parameters .................................................................................. 8 2.1.4 Energy Efficient Comminution ............................................................................... 10 2.1.5 HPGR Flowsheet Considerations .......................................................................... 13 2.1.6 Advantages and Disadvantages ............................................................................ 16

2.2 Stirred Media Mills ........................................................................................................ 21

2.2.1 Background ........................................................................................................... 21 2.2.2 Vertical Stirred Mill Technology ............................................................................. 22 2.2.3 Horizontal Stirred Mill Technology ......................................................................... 24 2.2.4 Horizontal Stirred Mill Operating Parameters ........................................................ 27 2.2.5 Energy Efficiency for Stirred Media Mills ............................................................... 29 2.2.6 Horizontal Stirred Mill Flowsheet Options ............................................................. 32 2.2.7 Process Benefits of Horizontal Stirred Mills........................................................... 35

2.3 HPGR / Stirred Mill Circuit ............................................................................................ 37

2.4 Literature Summary ...................................................................................................... 43

3 Experimental Procedure ...................................................................................................... 44

3.1 Definition of Comminution Circuits ............................................................................... 45

3.1.1 Cone Crusher / Ball Mill Circuit ............................................................................. 45 3.1.2 HPGR / Ball Mill Circuit ......................................................................................... 46 3.1.3 HPGR / Stirred Mill Circuit ..................................................................................... 46

3.2 Sample Description ...................................................................................................... 49

3.3 Equipment .................................................................................................................... 51

3.3.1 High Pressure Grinding Roll .................................................................................. 51 3.3.2 Horizontal Stirred Mill ............................................................................................ 53 3.3.3 Vibrating Screen .................................................................................................... 55 3.3.4 Bond Test Ball Mill ................................................................................................. 55

v

4 Testing and Simulation Results ........................................................................................... 57

4.1 Cone Crusher / Ball Mill Circuit Results ........................................................................ 58

4.1.1 Flowsheet Simulation ............................................................................................ 58 4.1.2 Specific Energy Calculations ................................................................................. 60

4.2 HPGR / Ball Mill Circuit ................................................................................................. 62

4.2.1 HPGR Pilot-Scale Testing ..................................................................................... 62 4.2.2 Flowsheet Simulation ............................................................................................ 71 4.2.3 Specific Energy Calculations ................................................................................. 74

4.3 HPGR / Stirred Mill Circuit ............................................................................................ 76

4.3.1 The Stirred Mill Circuit ........................................................................................... 76 4.3.2 The HPGR Circuit .................................................................................................. 84 4.3.3 Circuit Energy Summary ....................................................................................... 91

5 Discussion of Results .......................................................................................................... 92

5.1 Assessing Operating Parameters for Pilot-Scale Testing ............................................. 93

5.1.1 HPGR Operating Parameters ................................................................................ 93 5.1.2 Stirred Mill Operating Parameters ......................................................................... 94

5.2 Comparison of Comminution Circuits ........................................................................... 97

5.3 Preliminary HPGR / Stirred Mill Circuit Flowsheet ...................................................... 102

5.4 Refined Procedure for Future Testing ........................................................................ 104

6 Conclusions and Recommendations ................................................................................. 107

References ................................................................................................................................ 111

Appendix A – JK SimMet® Data ................................................................................................ 122

Appendix B – Bond Work Index Data ....................................................................................... 154

Appendix C – HPGR Data ........................................................................................................ 171

Appendix D – Stirred Mill Data .................................................................................................. 220

vi

List of Tables

Table 2-1 Summary of Energy Consumption for Comminution .................................................. 11

Table 2-2 Summary of Grinding Media Wear Rates ................................................................... 26

Table 2-3 Normalized Effect of Decreasing Ball Size ................................................................. 28

Table 2-4 Summary of Ball Mill Size Over the Years .................................................................. 30

Table 2-5 Summary of Power Density for Grinding Mills ............................................................ 31

Table 3-1 HPGR Machine Specifics ........................................................................................... 51

Table 4-1 Equipment Selection for Cone Crusher / Ball Mill Circuit ............................................ 58

Table 4-2 Bond Work Indices for Cone Crusher / Ball Mill Circuit .............................................. 60

Table 4-3 Feed Conditions for Pressing Force Tests ................................................................. 62

Table 4-4 Results for Cycle Four of Closed Circuit Testing ........................................................ 70

Table 4-5 Equipment Selection for HPGR / Ball Mill Circuit ........................................................ 71

Table 4-6 Summary Bond Ball Mill Work Indices for Cone Crusher and HPGR Product ............ 74

Table 4-7 Test Conditions for the 355µm Signature Plot ............................................................ 77

Table 4-8 Test Conditions for the 710µm Signature Plot ............................................................ 78

Table 4-9 Summary of Mill Operating Conditions for 1.2mm Testing ......................................... 79

Table 4-10 Revised Test Conditions for 710µm Signature Plots ................................................ 82

Table 4-11 Summary of HPGR Results for First Stage Open and Closed Circuit Testing .......... 86

Table 4-12 Comparison of Wet and Dry Screening .................................................................... 91

Table 4-13 Summary of HPGR / Stirred Mill Energy Requirements ........................................... 91

Table 5-1 Statistics Summary of Circuit Energy Values ............................................................. 99

Table C-1 HPGR Pilot-Scale Test Key ..................................................................................... 172

Table C-2 HPGR Operating Data Phase One .......................................................................... 173

Table C-3 HPGR Operating Data Phase One (continued) ....................................................... 174



Table C-6 HPGR Operating Data Phase Two .......................................................................... 177

Table C-7 HPGR Operating Data Phase Two (continued) ....................................................... 178



Table C-10 HPGR Feed Size Distributions Phase One ............................................................ 181

Table C-11 HPGR Feed Size Distributions Phase One (continued) ......................................... 182

vii

Table C-12 T1A01 Product Size Distributions .......................................................................... 183











Table C-23 HPGR Feed Size Distributions Phase Two ............................................................ 205

Table C-24 HPGR Feed Size Distributions Phase Two (continued) ......................................... 206

Table C-25 HPGR Feed Size Distributions Phase Two (continued) ......................................... 207


Table C-27 T2B01 Product Size Distributions .......................................................................... 210





viii

List of Figures

Figure 2-1 Diagram of HPGR Comminution ................................................................................. 6

Figure 2-2 Flowsheet for Cerro Verde ........................................................................................ 15

Figure 2-3 Flowsheet Comparison for Peňasquito ...................................................................... 16

Figure 2-4 Standard Photographic Evidence of Micro-cracking .................................................. 18

Figure 2-5 Example of Gravity-Induced Vertical Stirred Mill Technology (Vertimill®) .................. 23

Figure 2-6 Examples of Fluidized Vertical Stirred Mill Technologies .......................................... 24

Figure 2-7 IsaMillTM Layout ......................................................................................................... 25

Figure 2-8 IsaMillTM Grinding Mechanism ................................................................................... 25

Figure 2-9 Flowsheet for Mount Isa Mines .................................................................................. 33

Figure 2-10 Flowsheet for Regrind Circuit at Kumtor Mine ......................................................... 34

Figure 2-11 Original Flowsheet for McArthur River ..................................................................... 35

Figure 2-12 Modified McArthur River Flowsheet with Coarse IsaMillTM Grinding ........................ 35

Figure 2-13 HPGR Flowsheet for Fines Production in the Cement Industry .............................. 38

Figure 2-14 HPGR Flowsheet the Sukhoy Gold Plant ................................................................ 39

Figure 2-15 Example of an HPGR / Stirred Mill Circuit ............................................................... 40

Figure 2-16 A Proposed HPGR / IsaMillTM Circuit ....................................................................... 40

Figure 2-17 Anglo Platinum’s HPGR Test Circuit ....................................................................... 41

Figure 2-18 Anglo Platinum’s HPGR / Stirred Mill Testing Flowsheets ...................................... 42

Figure 3-1 Cone Crusher / Ball Mill Flowsheet ........................................................................... 45

Figure 3-2 HPGR / Ball Mill Flowsheet ....................................................................................... 46

Figure 3-3 HPGR / Stirred Mill Flowsheet (Open Circuit) ........................................................... 47

Figure 3-4 HPGR / Stirred Mill Flowsheet (Closed Circuit) ......................................................... 47

Figure 3-5 Geographic Location of Mesaba ................................................................................ 49

Figure 3-6 Mesaba Feed Size Distribution .................................................................................. 50

Figure 3-7 Pilot-Scale HPGR Installation .................................................................................... 52

Figure 3-8 M20 Stirred Mill Installation ....................................................................................... 53

Figure 3-9 M20 Stirred Mill with Mixing Tanks ............................................................................ 54

Figure 3-10 ZS40 SWECO® Vibrating Screen ............................................................................ 55

Figure 3-11 Bond Test Ball Mill ................................................................................................... 56

Figure 4-1 Cone Crusher / Ball Mill JK SimMet® Flowsheet ....................................................... 59

Figure 4-2 Comparison of Specific Pressing Force and Product Size ........................................ 63

ix

Figure 4-3 Comparison of Specific Pressing Force and Specific Throughput ............................ 63

Figure 4-4 Comparison of Specific Pressing Force and Specific Energy Consumption ............. 64

Figure 4-5 Comparison of Moisture Content and Product Size .................................................. 65

Figure 4-6 Comparison of Moisture Content and Specific Throughput ....................................... 65

Figure 4-7 Comparison of Moisture Content and Specific Energy Consumption ........................ 66

Figure 4-8 Comparison of Roller Speed and Product Size ......................................................... 67

Figure 4-9 Comparison of Roller Speed and Specific Throughput ............................................. 67

Figure 4-10 Comparison of Roller Speed and Specific Energy Consumption ............................ 68

Figure 4-11 Product Size for Closed Circuit Testing ................................................................... 69

Figure 4-12 Specific Throughput for Closed Circuit Testing ....................................................... 69

Figure 4-13 Specific Energy Consumption for Closed Circuit Testing ........................................ 70

Figure 4-14 HPGR / Ball Mill JK SimMet® Flowsheet ................................................................. 73

Figure 4-15 Summary of Bond Work Indices .............................................................................. 74

Figure 4-16 Signature Plot for Top Size Testing of 355µm ......................................................... 77

Figure 4-17 Signature Plot for Top Size Testing of 710µm ......................................................... 78

Figure 4-18 Stirred Mill Dynamic Classifier Pegs ........................................................................ 79

Figure 4-19 Summary of Mill Parameters for 1.2mm Test #1 ..................................................... 80



Figure 4-22 Replacement of Grinding Disc with Spacer ............................................................. 82

Figure 4-23 710µm Signature Plot Results with Revised Operating Conditions ......................... 83

Figure 4-24 Malvern and Screening Comparison for T1 ............................................................. 84

Figure 4-25 Malvern and Screening Comparison for T2 ............................................................. 84

Figure 4-26 Particle Size Distributions for Option A .................................................................... 87

Figure 4-27 Particle Size Distributions for Option B .................................................................... 87

Figure 4-28 Product Size for Second Stage Closed Circuit Testing ........................................... 89

Figure 4-29 Specific Throughput for Second Stage Closed Circuit Testing ................................ 89

Figure 4-30 Specific Energy Consumption for Second Stage Closed Circuit Testing ................. 90

Figure 5-1 Summary Layout for Cone Crusher / Ball Mill Circuit ................................................ 97

Figure 5-2 Summary Layout of HPGR / Ball Mill Circuit ............................................................. 97

Figure 5-3 Summary Layout of HPGR / Stirred Mill Circuit ......................................................... 98

Figure 5-4 Summary of Specific Energy Consumption for Each Circuit ..................................... 99

Figure 5-5 Product Size Distributions for Each Comminution Circuit ........................................ 101

x

Figure 5-6 Preliminary Layout for an HPGR / Stirred Mill Circuit .............................................. 103

Figure 5-7 Scoping Level Testing Procedure for HPGR / Stirred Mill Evaluation ..................... 106

Figure B-1 Bond Work Index Data Crusher Product (150µm) .................................................. 155

Figure B-2 Bond Work Index Data Crusher Product (150µm) (continued) ............................... 156

Figure B-3 Bond Work Index Data 3N/mm2 HPGR Product (150µm) ....................................... 157

Figure B-4 Bond Work Index Data 3N/mm2 HPGR Product (150µm) (continued) .................... 158





Figure B-9 Bond Work Index Data Crusher Product (106µm) .................................................. 163

Figure B-10 Bond Work Index Data Crusher Product (106µm) (continued) ............................. 164

Figure B-11 Bond Work Index Data 3N/mm2 HPGR Product (106µm) ..................................... 165

Figure B-12 Bond Work Index Data 3N/mm2 HPGR Product (106µm) (continued) .................. 166





Figure C-1 T1A01 Particle Size Distributions ............................................................................ 184









Figure C-10 T1A10 Particle Size Distributions .......................................................................... 202



Figure C-13 T2B01 Particle Size Distributions .......................................................................... 211



xi



Figure D-1 355µm Top Size Test Signature Plot Data .............................................................. 221

Figure D-2 355µm Top Size Test Particle Size Distributions .................................................... 222

Figure D-3 710µm Top Size Test Signature Plot Data .............................................................. 223

Figure D-4 710µm Top Size Test Particle Size Distributions .................................................... 224

Figure D-5 T2C02 Signature Plot Data (T1) ............................................................................. 225

Figure D-6 T2C02 Particle Size Distributions (T1) .................................................................... 226

Figure D-7 T2C03 Signature Plot Data (T2) ............................................................................. 227

Figure D-8 T2C03 Particle Size Distributions (T2) .................................................................... 228

xii

Acknowledgements

I would like to thank Xstrata Technology and the National Science and Engineering Research

Council (NSERC) for their generous financial support for my research. Special thanks are given

to Josh Rubenstein and Mike Larson for their indispensable advice and technical knowledge,

which has allowed me to better understand the fundamentals of stirred milling.

Special thanks are also given to Dr. Andrew Bamber and BC Mining Research for their

continued support of my research and their role in introducing me to the field of high-pressure

grinding. The ability to complete a Masters degree while running a “business unit” was a

challenge, but the skills I have taken out of it will help me immensely as I move forward.

I would like to thank my research committee and especially my faculty advisor Dr. Bern Klein for

helping me through the process and providing invaluable advice and guidance along the way. I

would also like to thank Pius Lo and UBC for providing me the facilities to complete my

research.

I would like to acknowledge Teck Ltd. and especially Steve Wilson for supporting my research

and allowing me access to samples. I hope the results of this thesis will help in making Mesaba

an economically viable mine in the future.

Special thanks go out to Stefan Nadolski, of Koeppern Machinery Australia, for his assistance

with this research and for his support as both a friend and a colleague. The help in refurbishing

the M20, provided by both himself and Darcy Houlahan, allowed this research to happen and for

that I am truly grateful.

Most of all, I would like to thank my family for their continuing support of my endeavours and for

allowing me to be the best that I can be.

1

1 Introduction

The mining industry will be faced with new challenges in the years ahead. The exponentially-

increasing global population has resulted in an increased demand for raw resources. With the

known rich, coarse-grained deposits depleted, attention has turned to development of low-grade

deposits requiring increased tonnages to achieve adequate metal production. This increased

tonnage has resulted in an increased energy demand associated with metal extraction.

Coupled with this, society is becoming increasingly conscious of their footprint on the

environment, and serious attempts have begun, to reduce carbon emissions and increase

energy efficiency (Norgate and Haque, 2010). To adapt to this changing landscape, the mining

industry must begin to accept and adapt new, more energy-efficient technologies and begin to

focus on developing flowsheets capable of addressing the above issues.

Comminution, the process of crushing and grinding ore to liberate valuable minerals, is the most

energy-intensive part of the processing flowsheet, and accounts for upwards of 75-80% of the

overall energy consumption of the processing plant (Abouzeid and Fuerstenau, 2009; Tromans,

2008). In addition, the main unit operations employed in this process, tumbling mills, are as low

as 1% efficient (Fuerstenau and Abouzeid, 2002).

Currently, the main comminution circuits employed in the mining industry to process hard-rock,

low-grade deposits include some form of tumbling mill. This equipment utilizes steel balls (ball

mills), competent ore (Autogenous Grinding (AG) mills) or the combination of the two (Semi-

Autogenous Grinding (SAG) mills) to fracture rock using the breakage mechanisms of impact

and abrasion. The rotation of these large, cylindrical mills requires a considerable amount of

energy. Although their established circuit design and ability to process high tonnages is a huge

benefit, the increased energy demand and inability to efficiently grind to liberation sizes below

45µm could slowly decrease their role in flowsheet designs of the future.

In the past 20 years, new, more energy-efficient technologies have been developed and

adapted for hard-rock mining comminution. High Pressure Grinding Rolls (HPGR), an

innovative technology adapted from the cement and briquetting industries, have begun to be

considered for more base metal projects now that roll surfaces have been developed to treat

hard, abrasive ores (Dunne, 2006). Operating with two counter-rotating rolls, HPGRs create a

compressive bed of particles between the rolls, utilizing the process of inter-particle breakage.

2

This form of breakage results in improved comminution performance with a decreased demand

on energy (Klymowsky et al., 2006). Additionally, unlike tumbling mills, which require steel balls

to act as an energy transfer medium, HPGRs transfer energy directly from the rolls to the bed of

material, resulting in an increase in energy efficiency (Fuerstenau and Kapur, 1995). Another

technology, known as a horizontal stirred mill or IsaMillTM, was adapted from the pharmaceutical

industry in the early 1990s to help effectively process fine-grained ore bodies (Johnson et al.,

1998). The IsaMillTM consists of a cylindrical tube with a centrally-rotating shaft, mounted with

evenly-spaced grinding discs. Loaded with small ceramic grinding media (2-6mm) and operated

at high speeds, the equipment utilizes high-intensity attrition breakage to reduce particles in

size. The rotation of a central shaft, as opposed to the entire grinding chamber (tumbling mills),

results in decreased energy requirements; while the combination of small grinding media and

increased media velocity, has been shown to improve the energy efficiency of grinding in

particle sizes below an f80 of 150µm (Burford and Clark, 2007).

The goal of this research was to examine the possibility of incorporating the above-mentioned

energy-efficient equipment into a single flowsheet and eliminating the need for a tumbling mill.

The biggest obstacle surrounding this research was that the proposed circuit would be operating

both pieces of equipment outside of their normal operating range. As HPGRs began being

adapted to the hard-rock mining sector, they found the most functionality in a tertiary crushing

role, preparing feed for the ball mill (Morley, 2006a). Therefore, the process envelope for an

HPGR operating in hard-rock circuits typically has feed sizes of up to 70mm, and products

normally no finer than 4mm (Gruendken et al., 2010). At the same time, horizontal stirred mill

technologies such as the IsaMillTM have begun to be well-established in ultrafine grinding as a

regrind mill, providing a more energy-efficient alternative for processing rougher concentrates

with an f80 no larger than 100µm (Gao and Holmes, 2007). To design the proposed circuit, a

suitable transfer size needed to be established, to utilize both pieces of equipment effectively. A

review of the literature found that a suitable circuit layout would comprise of two stages of

HPGR, followed by stirred milling (Daniel, 2007b). The literature provided very little operating

data for this circuit layout and therefore, optimization of operating parameters was fundamental

in making this circuit technically feasible.

3

The following were the proposed research objectives for the work summarized in this thesis:

• The organization and setup of appropriate pilot-scale research equipment, capable of

testing the proposed circuit. Included in this objective was a complete refurbishing of a

Netzsch M20 stirred mill, complete with an appropriate mixing system capable of

handling the coarse particle sizes tested.

• Determination of a suitable transfer size between the second-stage HPGR and the

stirred mill.

• Examination of possible circuit layouts for the two stages of HPGR comminution. This

included assessing changes in operating parameters and their effects on comminution

performance.

• Determination of the potential specific energy requirements necessary to operate the

proposed circuit.

• Comparison of the determined energy requirements with two conventional circuits

currently being used in the industry, a cone crusher / ball mill circuit and an HPGR / ball

mill circuit. This analysis included a combination of testing and simulation to gain an

appropriate baseline for comparison.

• Development of a preliminary HPGR / stirred mill circuit flowsheet.

• Formulation of a practical testing procedure, which could be applied for future

examination of the proposed circuit with other ore types.

4

2 Literature Review

The following chapter will review the current literature related to high pressure grinding rolls,

stirred media mills, and the combination of the two. This review will include a summary of the

fundamentals, an explanation for energy efficiency, the current flowsheet design for the

equipment, and the potential advantages and disadvantages of each technology.

5

2.1 High Pressure Grinding Rolls

2.1.1 Background

The technology of high pressure grinding and its adaptation to comminution was first realized by

Professor Klaus Schoenert in the late 1970s, with fundamental work on fracture physics in

comminution (Schoenert, 1979). His results concluded that increased energy efficiency was

possible with the use of compressive beds and the effects of inter-particle breakage. Using a

similar machine design to those used for the briquetting of coal, he adapted this concept into the

technology now known as High Pressure Grinding Rolls, or HPGRs for short (Schoenert, 1988).

HPGRs initially found considerable success being used to grind soft clinker material in the

cement industry. Converting closed circuit ball mills to semi-finish grinding circuits and

incorporating HPGRs for feed preparation, the industry was able to reduce energy consumption

by 15-30% (Patzelt, 1992). After its establishment in the cement industry, advocates of the

technology began to look for new applications. In 1990, the Argyle diamond mine of Australia

adopted the HPGR technology to help process the increased ore hardness of their deeper,

competent, un-weathered lamproite (Lane et al., 2009). Due to its ability to select pressures

strong enough to break the host rock, but weak enough to leave the valuable stones intact, the

applications of the HPGR in the diamond industry have become well-established (Anguelov et

al., 2008). Since then, HPGRs have also found a niche application crushing iron ore for pellet

feed preparation (Pyke et al., 2006).

The first attempt to apply the HPGR technology in hard-rock, base metal mining, occurred at the

Cyprus Sierrita mining complex near Green Valley, AZ in July 1995 (Thompsen et al., 1996).

The plant trials conducted on the copper-molybdenum ore resulted in important findings in the

areas of circuit design and wear rates for continuous operation. Although this application of

HPGRs was not successful in becoming a lasting operation, the trials presented a success in

further understanding the challenges that lay ahead, before the HPGR technology could

successfully be adapted to hard-rock mining (Morley, 2008).

Over the next decade, continuous improvements in roller wear, especially in the area of studded

lining, allowed HPGRs to become a more viable option for base metal mining. The

improvements of roller wear life, inter alia, allowed the successful installation of the HPGR

technology at the Cerro Verde copper mine in 2007, becoming the first large-scale HPGR

6

installation in hard-rock mining (Vanderbeek et al., 2006). Since then, installations at

Boddington (gold), PT Freeport Indonesia (copper, gold) and Amplats Potgietersrust (platinum)

have incorporated the HPGR technology into their flowsheets (Rosario and Hall, 2010).

2.1.2 Technology Overview

The technology of HPGRs is comprised of two counter-rotating rolls mounted on a sturdy frame

(refer to Figure 2-1). One roll is fixed, while the other is allowed to “float” and move horizontally.

Unlike the rolls crusher typically found in the coal industry, a force is exerted on the floating

roller by a hydraulic oil cylinder system, exposing material to pressures as high as 300N/mm2

(Daniel, 2002; Schoenert, 1988). Material is choke-fed between the two rolls, creating a

compressive bed of particles and reducing the material in size through inter-particle breakage.

To cushion the constant shock on the rollers, a set of nitrogen accumulators is installed behind

the floating roller, providing a smooth operating gap. To maintain high pressures between the

rolls, removable cheek plates are installed in the transverse direction. Due to this arrangement,

a pressure profile is created along the rolls, resulting in a finer crush at the centre and a coarser

crush along the edge.

Figure 2-1 Diagram of HPGR Comminution (Daniel and Morrell, 2004; Napier-Munn et al., 1996)

Due to the high compressive forces exerted on the material, the product tends to agglomerate,

producing what is commonly referred to as “flake.” The competency of this flake varies

depending on material type and although publications refer to flake competency testing (Morley,

2008); no quantitative method has been developed. Due to the presence of flake in the product,

7

some form of de-agglomerator, such as an impact crusher or scrubber, may need to be

considered when designing HPGR circuits (Schoenert, 1988).

Another feature unique to HPGR comminution is the concept of the floating operating gap.

Unlike other crushing equipment, there is no closed side setting when operating an HPGR.

Since the roller is floating, the operating gap varies during operation and the actual distance

between the rolls is dependent on the material and the frictional forces between itself and the

roll surface. Due to this unique operating gap, it is difficult to produce accurate throughput

models for HPGRs without performing pilot-scale testing (Klymowsky et al., 2006).

Normally, HPGRs are sized using the width and diameter of the rolls. Different aspect ratios

between the diameter and width are used, depending on the supplier. The main supplier of

HPGRs, Krupp-Polysius, uses a high aspect ratio, which is more expensive, but produces

longer wear life, since a smaller roller width allows less contact with the material. The other two

suppliers, KHD and Koeppern, utilize a smaller aspect ratio, which increases roller width,

creating a relatively finer product, due to a decreased edge effect and an increased pressure

near the centre of the roll (Morley, 2008). The roller size also dictates the allowable top size of

the feed. Since HPGRs promote compressive bed breakage, too large a top size will result in

single particle breakage, thus eliminating the benefits of HPGR comminution. According to

Polysius’ website, the largest HPGR has a throughput upwards of 3,000tph, with a feed top size

of 75mm (Polysius, 2011).

Currently, the only reliable way to scale-up, and properly size and select an HPGR, is to perform

pilot-scale testing. Unfortunately, this requires a large quantity of material and usually poses a

challenge to greenfield operations, since normally; only expensive drill core is available. The

lack of a reliable small-scale suite of tests that can be used for accurate size and selection is

one of the biggest hindrances preventing HPGRs from being explored in more projects (Daniel,

2002; Morrell et al., 1997).

HPGRs are mechanically very reliable, with typical availability as high as 98% (Morley, 2008).

The main reason for downtime remains the wear-life of different parts of the equipment. The

main wear components of HPGRs are typically the parts in direct contact with the material.

These parts include the feed chute, the cheek plates, and the roller surface (Dunne, 2006). The

roller surface is the most problematic, and a majority of downtime is associated with roller

8

change-out. Currently, the two main profiles used for roller surface are a smooth tyre and

studded lining.

The smooth surface roll profile was initially made of a Ni-hard, wear-resistant steel (Oberheuser,

1996), but over time, newer surfaces, such as Koeppern’s Hexadur® wear lining, have been

developed. Hexadur® consists of a hard, abrasive-resistant material set into a matrix of softer

material. This softer material wears quicker and creates a grooved profile that promotes the

formation of an autogenous layer. In the cement industry, Hexadur® has been found to last for

30,000 hours of continuous operation, but currently the only example of Hexadur® lining being

used for processing hard ore is at the Bendigo gold mine in Australia and no wear rates are

available (Pyke et al., 2006). Hexadur® can operate at pressing forces up to 6N/mm2 for

industrial applications (Morley, 2008).

The more applicable roll liner for hard-rock applications is the studded lining. Consisting of

tungsten carbide studs mounted on to a tyre, studded lining improves wear life through

formation of an autogenous layer developed between the studs. This autogenous layer reduces

contact between the roller surface and the abrasive material, and allows a wear life of between

4,000 and 8,000 hours (Klymowsky et al., 2006). The literature has suggested that not only

does studded lining promote longer wear life, it also improves throughput, although at the price

of increased energy consumption and higher grinding forces (Lim and Weller, 1999). Studded

lining is limited to a specific pressing force of 1-4.5N/mm2, after which the pressing force will

cause damage to the metal studs (Morley, 2008).

2.1.3 HPGR Operating Parameters

The two main operating parameters available when running an HPGR are the rotational speed

of the rolls and the horizontal pressing force applied by the floating roll. These two parameters

provide flexible control to an HPGR operator. A change in roller speed allows the operator to

adjust the throughput of the machine, up to a certain point. Increasing the roller speed will

subsequently increase throughput, decreasing operating gap and specific throughput, and

resulting in a narrower product size distribution (Lim et al., 1997). Koeppern recommends that a

nominal roller speed of 19.1 RPM or the equivalent of 1 roll diameter per second, be employed

for optimal results. By selecting the appropriate pressing force to be exerted by the floating

roller, operators have the ability to vary product size and control the p80 for downstream

processing. A decrease in pressing force will result in a coarser product, while an increase will

9

result in a finer product. Unfortunately, past a certain point, an increase in force produces little

size reduction, and results in an increase in wasted energy attributed to heat (Djordjevic, 2010;

Schoenert, 1988). To ensure effective operation of an HPGR, too high a pressing force should

not be used, as this would prevent the floating roller from floating. In this situation, an unknown

amount of force is being applied directly to the frame, considerably reducing the effectiveness

and efficiency of the machine. Therefore, determining the optimal pressing force is critical, since

this value varies depending on the properties of the ore.

To properly scale-up and size HPGR machines, a number of parameters have been developed

over the years. The following is a list of the main terms used in HPGR sizing and selection.

Process Specific Throughput Constant (m-dot)

The specific throughput constant, or m-dot, of the HPGR provides a normalized value allowing

for throughput comparison of different-sized roller presses. The m-dot is a function of roller

width, roller diameter, roller speed, and press throughput. The m-dot corresponds to the

throughput (t/h) of an HPGR fitted with roller dimensions of 1m width, 1m diameter and rotating

at 1m/s. The m-dot is material specific and used to size the HPGR roller dimensions for a given

press throughput (Klymowsky et al., 2006).

m-dot = W___ D * L * v

(1)

Where: m-dot (ts/hm3) = Specific throughput

W (t/h) = Press throughput

D (m) = Roller diameter

L (m) = Roller width

v (m/s) = Roller peripheral speed

Specific Pressing Force (Fsp)

The specific pressing force refers to the amount of force being applied by the floating roller to

material located between the rolls, as a function of roller width and roller diameter. The specific

pressing force is independent of roller size, allowing for comparison of process performance for

HPGRs that vary in size (Klymowsky et al., 2006).

10

Fsp = F__ D * L

(2)

Where: Fsp (kN/mm2) = Specific pressing force

F (kN) = Total pressing force exerted

Net Specific Energy Consumption (Esp)

The net specific energy consumption is the amount of energy transferred to the material running

through the HPGR. This value can be calculated by recording the average power consumption

of the machine before and after a test, and subtracting it from the power being consumed during

a test. The net power consumption is then divided by the throughput (t/h) of the machine during

stable operating conditions. Net specific energy consumption is used for motor sizing and was

the value used for comminution energy requirements in HPGR testing.

Esp = Pt - PiW

(3)

Where: Esp (kWh/t) = Net specific energy consumption

Pt (kW) = Total main motor power draw

Pi (kW) = Idle main motor power draw

2.1.4 Energy Efficient Comminution

The process of comminution has been well-documented as a very energy-intensive process.

Tromans (2008), using energy figures collected by national energy departments, summarized

the actual figures related to this demand on energy. Tromans found that upwards of 39% of the

overall energy consumption for mining activities is spent in the processing plant, and of that

figure, 75% can be attributed to the process of comminution. Table 2-1 summarizes the

author’s findings for the energy consumption of mining between 2001 and 2002, for the USA,

Canada, Australia, and South Africa.

11

Table 2-1 Summary of Energy Consumption for Comminution (Tromans, 2008)

Country Annual Energy Consumption

For Mining (PJ*) Percentage of Total National Energy Usage Consumed by Comminution

United States 1.343 x 103 ~0.39%

Canada 7.89 x 103 ~1.86%

Australia 260.9 ~1.48%

South Africa 309.5 ~1.8%

* 1 PJ = 1 x 1015 J

Overall, the demand for energy in mining is considerable, and the ability to reduce this

consumption by even a fraction would be very beneficial. To improve mining economics of the

future, focus should be on improving the energy efficiency of circuit flowsheets. It follows from

the above discussion, that comminution represents an area for potentially larger energy savings.

The fracture of rock in comminution occurs when compressive forces cause pre-existing flaws in

the rock to experience tensile stresses normal to the crack length (Hu et al., 2001; Tromans and

Meech, 2002). Failure results once an increased propagation of these cracks produces new

surface area, and a release of strain energy at the crack tip (Rumpf, 1973).

The Merriam-Webster online dictionary defines “efficiency” as “the ratio of the useful energy

delivered by a dynamic system to the energy supplied by it” (Merriam-Webster Dictionary,

2011). For comminution, and tumbling mills in particular, this could be defined as the energy

required for the breakage and size reduction of rock, over the mechanical energy delivered to

the system by rotating the mill. If this definition is used, efficiencies for comminution in the

range of 0.1-2% have been well-documented in the literature (Fuerstenau and Abouzeid, 2002;

Tromans and Meech, 2002; Tromans and Meech, 2004).

To provide a more meaningful number for the efficiency of comminution, other definitions have

been put forward by researchers. Some have argued that the output energy used in efficiency

calculations is not reflective of the process itself. Fuerstenau and Kapur (1995) argued that,

traditionally, only the surface energy required in the generation of new surfaces has been used,

ignoring the strain energy required for the growth of crack length. With the inclusion of strain

energy, the authors suggested that the baseline for comminution efficiency should be

determined by the energy resulting from single-particle fracture experiments. This argument

12

stems from claims that the energy utilization of comminution for single particle breakage is the

most efficient (Schoenert, 1979). If this definition were used for comminution efficiency, then

ball milling has efficiencies lying mostly within the range of 7 and 12%, with some as high as

17%, while HPGR comminution can be as high as 45% (Fuerstenau et al., 1996; Gutsche and

Fuerstenau, 1999).

Tromans (2008) hypothesized that comminution efficiency must have some maximum ideal

limiting efficiency. Using a theoretical analysis of fracture mechanics, the author found a value

somewhere between 5 and 10%, depending on Poisson’s ratio. Tromans summarized that

using this limiting efficiency, relative efficiencies ratios for comminution are between 3 and 26%,

depending on the material.

Whittles et al. (2006) found that in terms of fracture mechanics, the most efficient forms of

comminution are slow compression of single particles, followed by compression of a bed of

particles. Although slow compression of a single particle, which minimizes energy loss due to

heat and noise, is the most efficient form of comminution, this is not practical for large-scale

applications. Instead, Schoenert (1988) determined that, for continuous operations, the

application of a compressive force to a bed of particles was the most effective process. This

concept allowed for the development of the HPGR technology, improving upon the inefficiencies

inherent in ball milling. Although one of the main mechanisms in ball mill comminution is

through particle bed breakage, the low probability of particle collisions makes ball milling

inefficient (Gutsche and Fuerstenau, 1999). Ball milling involves a considerable waste of

energy in the lifting and dropping of steel balls, resulting in imperfect collisions which may or

may not actually produce enough impact force to result in particle breakage. Coupled with this,

a large quantity of energy is put into the wasted generation of heat, and the adverse wear of

liners and steel balls (Fuerstenau and Kapur, 1995).

Compared to ball mills, the breakage in HPGR comminution is a much more direct consequence

of the process. No medium is required in the transfer of energy to the material. The

compression exerted by the rolls is transferred directly to the material, resulting in an improved

utilization of energy (Fuerstenau and Kapur, 1995). Breakage results due to very high stresses

generated at the contact points between particles in the confined compressive bed. Because of

this inter-particle interaction, the pressure is amplified within the bed of particles, resulting in

pressures high enough to exceed the Uniaxial Compressive Strength (UCS) of the rock. This

13

process results in improved energy efficiency over traditional tumbling mills (Fuerstenau et al.,

1991; Fuerstenau et al., 1996). Although wasted heat is generated in HPGR processing,

Djordjevic (2010) found this to be unavoidable in rock fragmentation, due to the activation of

material friction and shear stresses acting along the fractured surface.

Regardless of the definition used for comminution efficiency, HPGR comminution has been well-

documented as improving energy utilization in comparison to conventional grinding practices;

however HPGR circuit configurations require an increased reliance on materials handling

equipment and the overall energy benefits associated with HPGR comminution are reduced.

Even with the increased energy demand of auxiliary equipment, several publications have

documented the overall energy benefit of HPGR circuits over conventional hard-rock circuits

such as the SAG Ball Mill Comminution Circuit (SABC). Oestreicher and Spollen (2006)

conducted a comparative study between an SABC circuit and an HPGR / ball mill circuit using a

combination of testing results, operational data, and simulation. The authors found an overall

reduction in energy of 19.6%. Rosario and Hall (2010) presented a study comparing two case

studies examining SABC and HPGR / ball mill circuits. The authors concluded that if only

specific energy consumption of the comminution process was compared, then Case A had a

reduction in energy of 25.1% and Case B had a reduction in energy of 30.2%; however if the

overall circuit was examined, including all auxiliary equipment, then the energy reduction

dropped by only 11.7% and 18.4%, respectively. Finally, Anguelov et al. (2008) summarized a

number of trade-off studies performed by Wardrop Engineering for various mining projects. An

average energy savings of 25% was determined when incorporating the HPGR into hard-rock

flowsheets.

2.1.5 HPGR Flowsheet Considerations

Several flowsheets options have been developed to capitalize on the energy benefits associated

with the HPGR technology summarized in Section 2.1.4. One of the most important aspects to

consider when designing an HPGR circuit is that as pressure (energy) is increased in an HPGR,

there is a limiting factor where the generation of fines begins to clog pores, preventing further

breakage. At this threshold, an increased amount of energy is wasted in the form of heat

(Djordjevic, 2010). Therefore, when designing an HPGR circuit, final preparation of product

must incorporate a grinding mill to produce the adequate amount of fines for further processing.

14

When HPGRs were first introduced into the cement industry, their role in the flowsheet was to

perform much of the work that was traditionally being done by a ball or tube mill (Aydogan et al.,

2006). In most circumstances, the introduction of an HPGR before the ball mill, acting as a

“booster,” helped improve throughput and reduce specific energy consumption (Fuerstenau et

al., 1991). As HPGRs began to be adapted to hard-rock mining, the best use for the technology

was found to be in replacing a cone crusher (tertiary crushing role) in a standard three-stage

crushing circuit. This arrangement typically requires a feed top size of between 50 and 70mm,

and produces a product size between 4 and 6mm. When operating an HPGR for this duty, the

industry developed some guidelines after the Cyprus Sierrita plant trials of the mid 1990s, the

most important of which was the careful preparation of feed for the circuit (Morley, 2006b). The

presence of tramp metal in the feed can be detrimental to the roller lining, and a metal detector

should be placed on the conveyor prior to the HPGR feed chute. Also, the feed top size should

be no larger than the operating gap. This prevents the possibility of single particle breakage or

the damaging of metal studs. To achieve this, the secondary crusher should be placed in

closed circuit with a screen prior to the HPGR.

Possible circuit configurations for HPGRs include:

• Open Circuit

•

– A possibility if employing two-stage grinding or designing a grinding

circuit that could handle the coarse fraction. This configuration is the simplest, since no

recycle feed is required, and is ideal for products with hard competent flake. The

drawback of this circuit could be increased energy demand (Gruendken et al., 2010).

Closed Circuit with Dry Screening

•

– A possibility if flake competency is weak and

efficient screening can be achieved. This configuration is ideal over wet screening,

since it prevents unwanted water from entering the circuit. The main drawback of this

configuration is excessive dust generation (Morley, 2006a).

Closed Circuit with Wet Screening

•

– The most likely circuit when dealing with more

competent flake, or with less efficient screening (Morley, 2006a). Unfortunately,

increasing moisture content in the feed leads to lower throughput and higher energy

consumption (Fuerstenau and Abouzeid, 2007). Therefore proper circuit design must

compensate for these limitations.

Closed Circuit with Product Splitter – This involves taking a split of the product,

preferably the edge product, and recycling the coarser material without the need for

classification. This configuration is ideal since no screening is required; however, since

15

HPGR product will still contain a fraction of coarse particles when entering the ball mill,

the situation could arise where the ball mill is unable to effectively break these particles

and a critical build-up could ensue (Gruendken et al., 2010).

HPGRs have an advantage over conventional crushers when operating in closed-circuit

operations. HPGRs can still produce an effective crush of oversize material, since unlike

crushers, there is no closed side setting, limiting the probability of breakage (Gruendken et al.,

2010).

Morley and Daniel (2009) examined what future HPGR flowsheets should entail. The authors

believe that the next generation of HPGR flowsheets should attempt to eliminate the need for

auxiliary equipment, and allow the secondary crusher or the HPGR to operate in some form of

open circuit. This would greatly decrease the capital cost and complexity of HPGR circuits.

Although several ideas are put forward, each suggestion is met with substantial operating

problems, making these flowsheet concepts unrealistic at the present time.

The role of an HPGR as a tertiary crusher has thus placed it in direct competition with the SAG

Ball Mill Comminution (SABC) circuit currently being used as the standard for ball mill feed

preparation (Gruendken et al., 2010). Although an SABC circuit is currently the industry

standard, an HPGR circuit could provide increased operating benefits for hard-rock, high-

tonnage operations. The advantages that an HPGR circuit has over an SABC circuit (refer to

Figure 2-2) were first observed with the HPGR installation at Cerro Verde in Peru (Vanderbeek

et al., 2006).

Figure 2-2 Flowsheet for Cerro Verde (Rosario et al., 2011; Vanderbeek et al., 2006)

16

Another application currently being implemented at Gold Corps’ Peňasquito operation in

Mexico, utilizes an HPGR to treat pebble crusher product in their SABC circuit (refer to Figure 2-

3). The purpose of this arrangement is to improve the overall throughput of the comminution

circuit, by further reducing the size of the pebble crusher product. The authors estimated that a

30% increase in total capacity can be achieved, with only an additional 4.8kWh/t of energy

expenditure (Dixon et al., 2010).

Figure 2-3 Flowsheet Comparison for Peňasquito (Dixon et al., 2010)

2.1.6 Advantages and Disadvantages

When applying HPGRs to a process flowsheet, there are several benefits that make this

technology more appealing than the conventional SAG / ball mill and three-stage crushing

circuits. A review of the literature has suggested that proper design of HPGR circuits can lead

to the following advantages:

• Decreased Operating Costs – Due to the reduction in energy consumption provided by

HPGR comminution, lower energy costs will lead to lower overall operating costs. The

elimination of steel grinding media can also lead to cost savings. Morley (2008) stated

that, although roll tyre replacement is still required, the cost is approximately the same

as SAG mill liners. Vanderbeek et al. (2006) estimated that a reduction in operating

costs of $0.368/t was achievable by implementing an HPGR based circuit in place of an

SABC circuit.

17

• Inter-Particle Breakage

•

– The unique compressive bed and increased production of

fines created in HPGR comminution provides an advantage over conventional crushers

(cone) for preparation of ball mill feed. This benefit could lead to an increase in overall

circuit throughput, providing an improvement in ball mill circuits (Danilkewich and

Hunter, 2006).

Lower Sensitivity to Ore Variability

•

– Unlike SAG mill operations, HPGRs can handle a

large change in ore hardness, with little effect on throughput and comminution

performance. Due to the ability to adjust the specific pressing force, an operator can

directly account for increased ore hardness (Rosario, 2010).

Small Machine Footprint

•

– With the size of an HPGR being considerably smaller than

that of the SAG mills currently in operation, a reduced machine footprint can increase

available space in the mill, increasing flexibility for operations (Danilkewich and Hunter,

2006).

Short Equipment Lead Time

The above advantages provided by HPGRs are hard to dispute, but the other main advantages

resulting from the production of micro-cracks within the product, have yet to be fully proven.

Due to the high stresses created within the compression bed, some of the product does not

fracture, but does contain several micro-cracks. This can weaken the material and potentially

improve downstream processes, such as flotation and leaching.

– HPGRs have a relatively short lead time in comparison to

SAG mills, with differences between 6 and 14 months (Morley, 2008).

After a review of the literature, a standard photograph (Figure 2-4) was presented when the

topic of micro-cracking was discussed. No background on experimental procedure could be

found. Daniel (2007a) provided the first detailed account that HPGR product exhibits micro-

cracking. In Daniel’s doctoral dissertation, photographic existence of micro-cracking was

provided for a number of different ore types, using a Scanning Electron Microscope (SEM).

18

Figure 2-4 Standard Photographic Evidence of Micro-cracking (Morley, 2008)

Several publications have reported that micro-cracks result in a reduction in the Bond ball mill

work index of 10-25%, in comparison to conventional crusher product (Daniel, 2007a;

Danilkewich and Hunter, 2006; Muranda, 2009; Norgate and Weller, 1994; Rule et al., 2008).

Some have argued this is due to the higher presence of fines in HPGR product, providing the

illusion that HPGR product is weaker than conventional crusher product. To prove this concept

wrong, Rule et al. (2008) performed regular Bond work index tests, as well as a modified Bond

work index test, where fines were removed from the HPGR product to create a comparable size

distribution to the crusher product. A reduction of 12% was obtained for the unaltered HPGR

product, while a reduction of 7% resulted from the modified product.

According to the literature, increasing the specific pressing force results in an increased

reduction in the Bond work index. Norgate and Weller (1994) performed Bond work index

testing on zinc and gold ores at specific pressing forces, ranging between 1 and 12N/mm2. The

reduction was greatest between 4 and 8N/mm2, with little difference resulting from higher

pressures. The authors concluded from the results that the reduction in Bond work index should

be used as additional criteria when selecting an optimum specific pressing force. In addition to

size reduction and specific throughput, specific energy consumption of both HPGR, and ball mill

grinding, should be evaluated.

Tavares (2005) performed a comparative study between the HPGR and conventional crushing

equipment (roll crusher and hammer mill), to evaluate the reduction of impact energy necessary

to break particles. Copper and gold ores were tested at narrow size fractions using an impact

load cell. Particle weakening occurred in HPGR product, but diminished with decreasing

particle size. Tavares concluded that, below 1.5mm, particle weakening became insignificant

19

between the different pieces of equipment. Although this could be considered proof that particle

weakening does not occur at finer size fractions, the impact load cell may not be the ideal way

to test the strength of small particles. Daniel (2007a) found that extensive micro-cracking is still

present in fractions finer than 850µm.

To demonstrate that particle weakening leads to a reduction in the energy demand in ball mill

grinding, HPGR and conventional crusher (cone) product could be processed through a pilot-

scale ball mill. Under continuous operation, energy requirements for grinding to a specified

product size could be compared for each feed preparation method. So far, no literature was

found that incorporates this approach.

Unlike the presence of micro-cracking and the reduction of ball mill energy, the benefits micro-

cracking provides to downstream processing are less clearly understood. Micro-cracks are

believed to create a more porous material, leading to better leaching characteristics for gold and

silver ores. In a report by Golden Queen Mining, for their Soledad Mountain Project in Southern

California, bottle roll and column leach tests were performed to compare the HPGR with a

Vertical Shaft Impact (VSI) crusher. Higher recoveries and shorter leach times were achieved

with HPGR product, and subsequently, the choice of HPGR comminution was selected for their

heap leach operation (Klingmann, 2005). In the case of flotation, no conclusive evidence has

been put forth to confirm that HPGR product improves flotation kinetics and increases metal

recovery (Palm et al., 2010).

Hosten and Ozbay (1998) speculated that compressive bed breakage leads to material

fracturing along grain boundaries, resulting in liberation of mineral grains at coarse size

fractions. Daniel (2007a) attempted to answer the question of preferential liberation created by

HPGR comminution. The author’s work found no conclusive proof of this phenomenon.

Although HPGRs provide a veritable array of benefits, there are a few drawbacks to this

technology. The following are the main disadvantages associated with HPGR technology:

• Circuit Complexity – Due to the stipulations placed on the feed requirements for HPGRs,

circuit design typically includes a secondary crusher in closed circuit with a screen, a

metal detector to prevent tramp metal from damaging the rolls, and a screening circuit to

handle HPGR product prior to feeding the ball mill. These specifications require an

20

increased amount of materials handling equipment and increased capital costs

associated with the circuit (Morley, 2008).

• Increased Capital Cost

•

– Due to the present limitation on bearing size, and thus roller

size, an HPGR machine is limited to a maximum throughput of 2,500-3,000tph. For

larger operations, multiple machines are required to perform the same duty as one SAG

mill. Coupled with the increased auxiliary equipment mentioned above, HPGR circuits

typically have higher capital costs. Vanderbeek et al. (2006) estimated that Cerro Verde

capital costs for the complete HPGR comminution circuit (primary crushing through ball

milling) were ~29% higher than a complete SABC circuit. Although higher in capital, the

decrease in overall operating costs produced approximately 1.5% higher internal rate of

return for the project.

No Standard Energy and Throughput Model

•

– Although not a disadvantage of the

technology, currently, there are no standard small-scale tests available to accurately

predict energy and throughput for a given ore. Until small-scale tests are available that

require a small quantity of material, HPGRs will not be considered in early stage circuit

designs.

Industry Acceptance

•

– Due to the mining industry’s reluctance to embrace new

technology, HPGRs are not considered as a primary option in circuit design. Although

examples of HPGRs operating in hard-rock, high-tonnage operations are beginning to

materialize, until this technology is successfully introduced in a Canadian mining

operation, the status quo will remain in effect on this.

Inability to Process Clayish Ore

•

– HPGRs are unable to process sticky clayish ores, due

to slippage on the rolls, reduced throughput, and production of unreasonably large and

competent flakes. Although this is the current situation, work has been done on a circuit

design to handle such ores (Rosario et al., 2011).

Poor Performance with Increasing Moisture Content

Although there are limitations to the HPGR technology, proper circuit design and continuing

research and development should lead to the mitigation of the associated risks.

– When high moisture content is

present in the feed, poor performance in terms of throughput and wear rates can be

experienced. When processing wet material, the inability to produce a continuous

autogenous layer on the roller surface can drastically decrease roller life (Fuerstenau

and Abouzeid, 2007; Morley, 2008).

21

2.2 Stirred Media Mills

2.2.1 Background

Stirred mill technology, or the concept of a centrally rotated shaft to agitate media, was first

developed during the 1950s in Japan. This technology utilized a vertical orientation, and was

used for grinding in the minerals industry. In 1979, the Metso grinding division (then known as

Koppers) acquired the license, with the intention of applying it to base metal mining operations.

Unfortunately, the design was not suitable for the rigors of hard and abrasive ore, and

subsequent work was done to improve grinding efficiency, minimize maintenance and

downtime, and improve wear rates. These developments led to the creation of Svedala’s

Vertimill®, a tower mill with the ability to process material as coarse as 6mm, and produce

product as fine as 20µm (Allen, 2009).

In the early 1990s, Mt. Isa Mines Limited (now a part of Xstrata) was investigating technologies

to economically process two of their fine-grained lead/zinc deposits: Mount Isa in Queensland

and McArthur River in the Northern Territory. The McArthur River deposit was originally

discovered in 1955, but no company had been able to economically process the fine-grained

deposit with the grinding technology available. In the case of Mount Isa, the gradual decrease

in metallurgical performance in the mid 1980s, due to finer liberation size, resulted in recoveries

dropping to 50% by the early 1990s (Anderson and Burford, 2006; Burford and Niva, 2008).

High media costs, impractical energy requirements, and poor flotation performance from steel

media contamination, led to the decision that the available grinding technologies, ball mills and

tower mills, were unsuitable for ultrafine grinding to sub-10µm, and new technology was needed

to address this challenge. After researching other industries that require ultrafine grinding, the

process team at Mount Isa settled on the horizontal stirred mill technology, manufactured by

Netzsch of Germany. The technology was being used to process high value manufactured

products, such as printer inks, pharmaceuticals, paint pigments, and chocolate. These

applications required small mills, run in batch operation, using high cost, sanitary grinding

media. To adapt the mill to the metals industry, work was done to increase the mill capacity,

allow for continuous operation, and apply low-cost grinding media (Pease, 2007). This

development work resulted in the creation of the M3,000 IsaMillTM, leading to its installation at

Mount Isa in 1994 and McArthur River in 1995 (Burford and Clark, 2007).

22

Since the mid 1990s, the development of inert ceramic grinding media and increased mill size to

the M10,000 have allowed the IsaMillTM to move away from ultrafine grinding (<10µm) and

establish itself as a regrind mill, producing product with a p80 between 20 and 40µm. The

ability to further liberate rougher concentrate, without contaminating mineral surfaces, has led to

successful installations at Kumtor (gold), Western Limb (platinum tailings), Prominent Hill

(copper/gold), and Potgietersrust (platinum) (Anderson and Burford, 2006; Burford and Clark,

2007; Curry et al., 2005).

2.2.2 Vertical Stirred Mill Technology

There are two main orientations for stirred mill technology: vertical and horizontal. Each

orientation has its own advantages and disadvantages, but both use attrition as the main

breakage mechanism for size reduction.

Vertical stirred mill technologies can be classified into two sub-categories, gravity-induced and

fluidized, depending on how the grinding media is circulated within the mill and the speed at

which the shaft operates. In the case of gravity-induced circulation, a centrally-mounted,

double-helical screw is suspended into the cylindrical grinding chamber, and rotated at a low

speed in the range of 100 RPM (Sinnott et al., 2006). The chamber is filled with grinding media,

typically steel for coarse applications, and ceramic for fine applications. As the screw rotates,

media is drawn up the centre of the mill, and eventually cascades off the edge of the screw,

creating a gravity-induced, downward flow of media along the mill perimeter (Allen, 2009;

Sinnott et al., 2006). Material, fed as slurry, enters at the top of the mill and circulates down

along the perimeter, being drawn back upwards with the aid of the rotating screw (Cleary et al.,

2006). This action creates continuous contact with grinding media, initiating size reduction

through attrition. As material is ground finer, it overflows the mill and is sent to a classifier,

where coarse material is re-circulated back to the mill, and finer material is sent on as final

product. Figure 2-5 shows an example of particle flow in the Vertimill®. Examples of gravity-

induced stirred mills include the tower mill and the Vertimill®. Lichter and Davey (2006) stated

that tower mills are more efficient at a coarser feed size; however most installations operate in

regrind circuits at fine particles sizes (Allen, 2009).

23

Figure 2-5 Example of Gravity-Induced Vertical Stirred Mill Technology (Vertimill®) (Metso, 2010)

For fluidized stirred mill technology, a centrally-rotating shaft is suspended within a cylindrical

grinding chamber, but unlike gravity-induced circulation, the shaft is equipped with either pins or

grinding discs and operates at a high speed in the range of 250 RPM (Sinnott et al., 2006). The

rotating shaft agitates the grinding media, creating a fluidized bed. Slurry is fed in at the bottom

(Deswik) or top (Stirred Media Detritor) of the mill, and passes through the fluidized bed of

media, resulting in high intensity media – particle interactions. Product then passes through the

media retention screens and overflows the mill as product. Typically, these mills operate with

ceramic or sand grinding media, and are best suited for ultrafine grinding applications (Metso,

2010; Rule et al., 2008). Examples of fluidized vertical stirred mill technologies include Imerys’

Stirred Media Detritor (SMD) (pin configuration) and the Deswik mill (disc configuration). Both

of these technologies are shown in Figure 2-6.

Vertical stirred mill technology is a more efficient comminution technology compared with

conventional tumbling mills, and Metso has claimed that a 30-50% reduction in energy can be

achieved, depending on how fine a grind is required (Metso, 2010). Since a ball mill relies on

rotation of the entire grinding chamber to create slurry – media interactions, a greater

expenditure of energy is required. Size reduction is usually achieved by attrition and impact, but

impact is not as effective, due to the probability that media will collide with other media or the

mill lining, resulting in wasted energy. Allen (2009) proposed that the most efficient zone in the

24

ball mill is what is referred to as the kidney, in reference to its shape. In this zone, media and

particles are in constant contact with each other, resulting in an increased rate of attrition. A

similar zone of intense attrition is consistent throughout the stirred mill chamber, resulting in

improved grinding efficiency. Although this restricts the top size fed to the mill because this

zone reduces impact breakage, limiting fracture of coarse particles.

Figure 2-6 Examples of Fluidized Vertical Stirred Mill Technologies Stirred Media Detritor (Left) and Deswik Mill (Right) (Capstick, 2010; Metso, 2010)

2.2.3 Horizontal Stirred Mill Technology

In the metal mining industry, the main example of a horizontal stirred mill technology is the

IsaMillTM. This technology is comprised of a centrally-rotating shaft, enclosed by a fixed

cylindrical grinding chamber (refer to Figure 2-7). The shaft is installed with 7-8 evenly-spaced

polyurethane grinding discs, and operates at very high speeds, between 1,200 and 2,000 RPM

(Larson et al., 2008). The rotation of the grinding discs creates tip speeds of 19-23m/s, while a

tower mill and the SMD operate at 3m/s and 8m/s, respectively (Anderson and Burford, 2006;

Parry, 2006). Material is fed as slurry at one end of the mill and passes through the fluidized

media zone, where high-intensity attrition reduces the particles in size (Arburo and Smith,

2009). High-intensity attrition allows the IsaMillTM to process fine particles at a high throughput.

Attached to the end of the rotating shaft is a dynamic classifier that utilizes centrifugal forces to

retain grinding media and coarse particles, while allowing fine particles to exit the mill. A

25

diagram detailing this process is shown in Figure 2-8. Currently, the largest IsaMillTM available

is the M10,000, which is equipped with a 3MW motor.

Figure 2-7 IsaMillTM Layout (Burford and Clark, 2007)

Figure 2-8 IsaMillTM Grinding Mechanism (Burford and Clark, 2007)

Originally, IsaMillTM operations used close-proximity grinding media, including slag, ore gravel,

and sand. This grinding media was beneficial because it was cheap and in constant supply.

Unfortunately, this type of media is not hard enough to produce efficient grinding, and suffers

26

from high wear rates. Kwade and Schwedes (2002) stated that the stress intensity exerted by

grinding media, adheres to the following relationship:

SI = d3 * ρ * v2 (4)

Where: SI (N*m) = Stress intensity per media particle

d (m) = Media diameter

ρ (kg/m3) = Media density

v (m/s) = Media velocity

Since an IsaMillTM already operates at high speeds, to improve the effectiveness of grinding

media, an increase in diameter or density is required. The original grinding media had a low

Specific Gravity (SG) (2.4) and small diameter (<1mm), leading to milling inefficiencies and

limitation of feed size. With the introduction of ceramic grinding media, exhibiting higher SG

(3.7) and larger diameter (3.5mm), the IsaMillTM can operate at a coarser feed size (<150µm)

while providing a lower media wear rate (Burford and Niva, 2008). Table 2-2 shows typical wear

rates for different grinding media, including MT1, a ceramic grinding media manufactured by

Magotteaux International.

Table 2-2 Summary of Grinding Media Wear Rates (Curry and Clermont, 2005)

Media Type Consumption Rate (g/kWh)

Relative Consumption

MT1 (-4 +3 mm) 15 1.0

Alumina 1 (-4 +3 mm) 128 8.5

Alumina 2 (-4 +3 mm) 295 19.7

Australian River Pebble (-4 +3 mm) 200 13.3

Australian Silica Sand (-6 +3 mm) 781 52.1

Ni Slag (-4 +1 mm) 1305 87.0

The energy requirements for a full-scale IsaMillTM can be determined using a laboratory mill.

Gao et al. (1999) determined that a 1:1 energy scale-up exists between a lab-scale M4 mill and

industrial-scale M4,000 mill. This ratio is attributed to the grinding mechanism shown in Figure

2-8, which prevents short-circuiting in the mill, allowing for uniform grinding. Curry et al. (2005)

reported that results obtained in an M4 can accurately scale-up to an M10,000.

27

2.2.4 Horizontal Stirred Mill Operating Parameters

When running a horizontal stirred mill, several operating parameters pertain to performance in

the mill. Some parameters are related to the feed conditions entering the mill (feed density and

volumetric flow rate), while others relate to the operating conditions of the mill (mill speed,

media volume, and media size). The following section will discuss each of these parameters

and their effect on mill performance.

Feed Density

The feed density is related to the density of the solids component and the percentage by weight

occupied in the slurry. Several publications reported that operating at low percent solids, below

30-40%, results in lower energy efficiency (Gao et al., 1999; Lichter and Davey, 2006); however

this limit may be material-dependent, depending on SG of the ore (Larson et al., 2008). Larson

et al. (2008) suggested that operating at 50% solids results in optimal energy efficiency and

greater than 65% may result in poor efficiency due to viscosity issues. Feed density is a main

parameter used to control retention time in the mill. An increase in solids content allows the mill

to operate at the same throughput, but decreases the amount of energy transmitted to the

material.

Volumetric Flow Rate

The volumetric flow rate refers to the amount of slurry passing through the mill in a given time

interval. Larson et al. (2008) found that the effect of flow rate has little influence on energy

efficiency and only affects residence time in the mill. This residence time will affect the size

reduction of the product, but the energy usage will adhere to the same curve.

Mill Speed

The mill speed refers to the rotational speed of the agitator. Typical values for operation vary

depending on mill size (disc diameter), but result in tip speeds of 19-23m/s. Larson et al.

(2008), using a lab-scale M4 mill, determined that mill speed has very little effect on energy

efficiency, and a linear relationship exists between speed and mill power draw. Parry (2006)

suggested that varying mill speed can control the stress intensity exerted by the grinding media

and could be used to process soft and hard minerals.

28

Media Volume

The media volume refers to the percentage of bulk media occupying the grinding chamber when

the shaft and disc volume are removed. The generally accepted operating range for media

volume is between 60 and 80%. When operating below this range, insufficient media is

available for grinding and the possibility of unbroken solids can lead to clogging of the mill

(Larson et al., 2008). The adjustment of media volume is one of the options available to

operators to prevent over-grinding when a decrease in throughput is experienced. Termed “turn

down,” operators can decrease the media volume in the mill to reduce the energy input into the

material and operate at a lower throughput with the same grind (Curry et al., 2005).

Media Size

The media size, measured by the diameter, is the most critical parameter available to optimize

energy efficiency. The selection of media size is crucial, since coarse media is required to

break the larger particles, but must be small enough to provide efficient grinding for finer

particles. Jankovic (2003) found a difference in energy efficiency of 40% when poorly selected

media size was tested. The impact of media size is shown in Table 2-3, where a decrease in

media size leads to an increase in media surface area, resulting in an increase in media –

particle collisions.

Table 2-3 Normalized Effect of Decreasing Ball Size (Lichter and Davey, 2006)

Ball Size (mm)

Surface Area (m2/t)

Number of Balls (per tonne)

Number of Balls Normalized

20 83.3 66,3125 1

15 111.1 157,190 2.4

10 166.7 530,516 8

5 333.3 4,144,132 62

3 555.6 19,648,758 296

2 833.3 66,314,560 1,000

Mankosa et al. (1986) suggested a selection ratio for media size to mean (80% passing) particle

size of 20:1 for fine grinding. Using this ratio, a feed f80 of 300µm would require a media top

size of 6mm. This ratio provides a balance between an increased probability of media – particle

29

interaction and the capability for media to catch and break particles. Yue and Klein (2006)

confirmed these assumptions using a geometric analysis and suggested that this ratio allowed

for the capture of 4 to 5 particles within bead voids. So far no literature was found related to

media selection for coarse stirred milling and this ratio may be lower for coarser applications.

2.2.5 Energy Efficiency for Stirred Media Mills

As discussed in Section 2.1.4, operation of conventional tumbling mills requires a substantial

amount of energy to rotate large cylindrical mills filled with steel media and slurry. This rotating

action creates the lift for steel balls to tumble, thereby reducing coarse particles in size through

impact breakage, while providing the motion necessary to grind particles between steel balls for

attrition breakage. The combination of these mechanisms allows ball mills to be applicable to a

wide range of sizes from an f80 of 4mm down to a p80 of 45µm, below which these

mechanisms lose an increasing amount of energy to ball – ball and ball – liner collisions. These

two mechanisms of breakages are effective; however the low probability of particle – ball

collisions leads to low energy efficiency (Fuerstenau and Abouzeid, 2002). Section 2.1.4

referred to the kidney zone in a ball mill. In this zone, constant ball – particle collisions results in

high fines generation through attrition. Unfortunately, this zone is not consistent throughout the

mill, because an open volume is still required for media to tumble and to create impact

breakage. Due to these shortcomings, over the years, compensation for low efficiencies has

resulted in the installation of larger ball mills, increasing the power requirements for grinding

(refer to Table 2-4). Stirred mill technology has evolved over the years to help improve upon

this increased energy requirement. The energy benefits associated with increased media –

particle interactions, resembling the kidney zone in a ball mill, accompanied with lower power

draw necessary to rotate a central shaft, has led stirred milling to become a viable option for

regrind mills (Lichter and Davey, 2006).

30

Table 2-4 Summary of Ball Mill Size Over the Years (Daniel, 2007a; Lynch and Rowland, 2005)

Year Diameter

(m) Length

(m) Power (kW)

1909 1.2 2.1 11

1912 1.9 2.3 41

1927 2.4 2.4 168

1940 3.1 2.8 447

1963 3.9 5.5 1,491

1970 5.6 6.4 3,169

1990 6.1 9.3 5,593

1997 7.3 10.5 11,440

Kwade and Schwedes (2002) stated that the stress intensity exerted by media is proportional to

the velocity squared. The speed of media in tumbling mills is limited by the speed at which

centrifuging of mill contents begins and effective breakage from cascading media ceases,

referred to as critical speed (Kapur et al., 1992). This limits the size of the grinding media in a

ball mill because a smaller media size cannot subject particles to the high-stress intensities

required for breakage (Wang and Forssberg, 2007).

Breakage characteristics in stirred mills are dependent on the stress intensity exerted on the

particle and the number of stress events experienced by feed particles and their resulting

daughter fragments (Kwade and Schwedes, 2002). For grinding of coarse particles, the stress

frequency is high because larger particles have a higher probability of making contact with

media. In addition, coarse particles also exhibit a higher degree of flaws, resulting in lower

required stress intensity for breakage (Wang and Forssberg, 2007). If an appropriate-sized

media were selected, then it should be possible to efficiently grind coarse particles in a stirred

media mill.

As particles are reduced in size, they require a higher number of collisions and increased stress

intensity to cause further breakage. To achieve this, stirred mill technology, such as the

IsaMillTM operates at high impeller speeds and small media size, resulting in increased energy

intensity and a higher frequency of collisions. Table 2-5 compares the power density in an

IsaMillTM to other grinding mills. The power density in a stirred mill is considerably higher than in

31

a tower mill. If this extra energy is utilized effectively through the optimization of operating

parameters, then increased energy efficiency should be achievable.

Table 2-5 Summary of Power Density for Grinding Mills (Burford and Niva, 2008)

Mill Type Installed Power

(kW) Mill Volume

(m3) Power Density

(kW/m3)

Autogenous Mill 6,400 353 18

Ball Mill 2,600 126 21

Regrind Mill 740 39 19

Tower Mill 1,000 12 42

IsaMillTM – M10,000 3,000 10 300

Several publications found that for fine size ranges up to 150µm, the mechanisms mentioned

above result in stirred mill technology being a more energy-efficient option over ball mills,

independent of orientation (Allen, 2009; Anderson and Burford, 2006; Anyimadu et al., 2007;

Lichter and Davey, 2006; Pease, 2007). As the feed increases in particle size, the effectiveness

of the ball mill becomes more apparent. Shi et al. (2009) conducted a study to determine

whether stirred mills could achieve higher energy efficiency than ball mills at coarse particle

sizes. The first test examined the energy comparison between a vertical stirred mill and a Bond

ball mil for processing material with a feed top size of 3.35mm down to a p80 of 75µm. The

stirred mill achieved energy reductions of 25%, 37%, and 27% for the three ore types tested.

These results are similar to other researchers’ suggestions that vertical stirred mills are more

energy-efficient than ball mills for primary grinding applications (Lichter and Davey, 2006). Allen

(2009) noted that although Vertimills® have been applied to regrind applications, they are best

suited for primary grinding.

The second test by Shi et al. (2009) investigated the energy requirements of an M4 IsaMillTM

and a batch ball mill (300mm x 300mm) for processing a sample with a feed top size of 1mm

down to varying product sizes. The ball mill was able to achieve slightly lower energy

requirements for coarser grinds (p80 > 40µm), but became less efficient with finer grinds.

Although the ball mill was more efficient at coarser particle sizes, it is evident that operating

conditions for testing were not optimized. Testing conditions consisted of operating at 30%

solids, using 3.5mm ceramic grinding media. Section 2.2.4 mentioned that media size plays an

32

important role in the effectiveness of breakage at coarse sizes. The recommended media size

ratio of 20:1 was not used, which may explain the ineffectiveness of grinding at coarse sizes.

More energy was required by the smaller media to break the coarse particles, resulting in a

lower efficiency. Based on the 20:1 ratio, a media top size of 6-8mm should have been used.

Other publications have examined the energy efficiency between tower mills and the IsaMillTM

and found that the IsaMillTM operates more effectively at finer sizes while a tower mill becomes

more efficient at coarser sizes (Burford and Niva, 2008; Harbort et al., 2010). Parry (2006)

performed a comparative study between tower mills installed at the Red Dog mine in Alaska and

lab-scale stirred mills. Using an IsaMillTM and a Stirred Media Detritor, Parry found a 50%

reduction in energy for grinding from an f80 of 29µm to a p80 of 22µm. Although the tower mill

may function more effectively at coarse sizes, its orientation poses considerable problems in

scale-up. When designing vertical stirred mills, the motor selection is dependent upon the

required torque necessary to rotate the media from rest. With larger units, this torque begins to

dominate the mechanical design, requiring a substantially larger motor and support frame. Due

to the orientation of the stirrer in a horizontal configuration, scaling up does not result in this

problem and the design of larger mills is more feasible (Pease, 2007).

Overall, there are benefits and limitations to both stirred mill orientations, but the use of stirred

mill technology has the potential to improve the energy efficiency of grinding and with proper

flowsheet design, could be a viable option over conventional ball milling.

2.2.6 Horizontal Stirred Mill Flowsheet Options

Originally the IsaMillTM was developed for ultrafine grinding applications, to efficiently process

finely-disseminated ores. Typical flowsheets place the IsaMillTM in the regrind circuit, accepting

rougher concentrate product from the regrind ball mill, at an f80 of 25µm, and further liberating

minerals by grinding to below a p80 of 10µm (Gao et al., 2002). An example of this is shown in

Figure 2-9 where the IsaMillTM is installed at Mount Isa Mines to treat the lead and zinc

concentrate for further liberation.

33

Figure 2-9 Flowsheet for Mount Isa Mines (Gao et al., 2002)

After establishing itself as an energy-efficient alternative for ultrafine grinding to sub 10µm, the

IsaMillTM is now being considered for coarser applications, accepting feed sizes up to an f80 of

150µm. Combining the increased grinding effectiveness of harder, higher SG ceramic media

with an overall increase in mill size, design of regrind circuits are beginning to incorporate

IsaMiIlsTM as the mill accepting feed directly from rougher flotation (Burford and Clark, 2007).

An IsaMillTM installed at the Kumtor gold mine in Kyrgyzstan was originally designed to process

product from a regrind ball mill, but during ball mill maintenance, the M10,000 was used as its

replacement (refer to Figure 2-10). During this time, the IsaMillTM was required to accept an f80

of 130-150µm and produce a product p80 of 60-65µm. Although there was insufficient time to

optimize media top size and other operating conditions, the circuit was able to operate

effectively. Had these conditions been optimized, plant operators believed that a reduction in

power draw and a decrease in product size could have been achieved as well (Anderson and

Burford, 2006).

34

Figure 2-10 Flowsheet for Regrind Circuit at Kumtor Mine (Burford and Clark, 2007)

The Phu Kham copper/gold mine in Laos, currently in development, will utilize an M10,000

IsaMillTM for regrinding flotation concentrate. The circuit will operate at a throughput of 168tph,

processing an f80 of 106µm and producing a product p80 of 38µm (Burford and Clark, 2007).

IsaMillTM technology is beginning to be applied to coarse grinding applications. Pease (2007)

presented a summary of an IsaMillTM at McArthur River operating as a secondary grinding mill,

treating SAG mill product. Originally the IsaMillTM was installed to grind 50µm concentrate down

to a p80 of 7µm, while a SAG mill was used to prepare feed for flotation (p80 of 45µm). To

increase tonnage a tower mill was installed to process a portion of the SAG underflow stream

(refer to Figure 2-11). Lab and pilot-scale tests were carried out on SAG underflow product to

determine the potential grinding limits for coarser IsaMillTM applications. With the plus 1mm

fraction screened out, the IsaMillTM achieved a finer product for the same energy input as the

installed tower mill. Unfortunately a build-up of steel scats and coarse particles provided

problems for continuous operation. The IsaMillTM was subsequently tested using minus 1mm

cyclone overflow with a reduced f80 of 300µm. Plans are now underway to increase throughput

at McArthur River by installing two M10,000 mills to treat SAG screened product (f80 of 300µm)

and grind to a flotation feed p80 of 45µm (refer to Figure 2-12).

35

Figure 2-11 Original Flowsheet for McArthur River (Pease, 2007)

Figure 2-12 Modified McArthur River Flowsheet with Coarse IsaMillTM Grinding (Pease, 2007)

2.2.7 Process Benefits of Horizontal Stirred Mills

There are several process benefits available for horizontal stirred mills which can help improve

the economics of the circuit. The following are the main advantages offered:

• Small Machine Footprint

•

– Although not a process benefit, the compact size and energy

intensity offered by stirred mills allows for a small footprint, resulting in an increase in

available space, in case plant expansion and an increased throughput is desired.

Inert Grinding Media – The use of ceramic grinding media eliminates the potential for

steel contamination of mineral surfaces, a problem that may hinder flotation kinetics.

Several publications have noted, especially in the precious metals sector, that the use of

36

inert ceramic grinding media can help improve flotation response for fines flotation (Rule

et al., 2008). The use of steel media can lead to deposits of iron hydroxide on the

surface of sulphide minerals, resulting in poor flotation selectivity and increased reagent

dosage. The use of ceramic grinding media has eliminated this problem and can allow

for cleaner liberated mineral surfaces at finer sizes, improving the economic benefits of

fine-grained mineral deposits (Arburo and Smith, 2009; Pease et al., 2006).

• Internal Classification

– The dynamic classifier installed at the end of the mill agitator

shaft, allows the mill to operate in open circuit, producing a sharper product size

distribution. This configuration leads to the elimination of a recycle stream, reducing

maintenance costs, and increasing throughput capacity. The uniform grinding action

experienced throughout the mill, leads to a reduction of over-grinding and the prevention

of ultrafines, which may prove problematic in the flotation circuit (Burford and Clark,

2007).

37

2.3 HPGR / Stirred Mill Circuit

To take advantage of the two energy efficient technologies summarized in Sections 2.1.4 and

2.2.5, an appropriate circuit flowsheet must be developed. Section 2.1.5 stated that an HPGR is

limited in the size reduction achievable from one pass through the rolls. To create a product

fine enough to process through a stirred mill, a second stage of HPGR comminution is

necessary. Several publications have documented the effects of processing material through

multiple passes of an HPGR. Norgate and Weller (1994) performed tests on a gold ore to

determine whether operating several stages of HPGR at a lower specific pressing force could

be more energy-efficient than single-stage operation at a higher specific pressing force. The

results showed that with subsequent passes through the mill, the specific energy consumption

trended downward. By pass four, specific energy consumption began to flatten out. The largest

decrease, at 31%, was experienced between pass one and two. In terms of size reduction, a

positive linear relationship was found between the total specific energy consumed for all passes

and the reduction ratio. It should be noted that this test was also repeated with a de-

agglomeration step performed between each pass, resulting in no significant change in results.

The data presented by Norgate and Weller (1994) shows that the operation of multiple HPGRs

in series can lead to an increased size reduction, producing the feed range acceptable for

stirred milling.

Daniel (2007b) conducted tests on a pilot-scale HPGR to assess whether several HPGRs in

series could produce a similar grind size to a ball mill, at a lower specific energy consumption.

The first two passes through the HPGR produced the highest size reduction ratios and

subsequent passes became less efficient. Daniel concluded that two passes through the mill is

the limit for efficient crushing of hard ores.

Rule et al. (2008) conducted testwork on a Labwal HPGR to determine the effect of passing

material through two open-circuit stages of HPGR comminution, while varying the specific

pressing force in the second stage. Good size reduction was achieved between the first and

second pass, but no significant difference in size reduction resulted from increasing the specific

pressing force in stage two. To optimize the specific energy consumption for operation of a two

stage HPGR circuit, an increase in specific pressing force in stage two is unnecessary.

Fuerstenau et al. (1999) summarized work showing the benefits that an HPGR / ball mill circuit

can achieve by optimizing the ball size. The characteristics of HPGR product, including fines

38

content and micro-cracking, allows for the ability to improve the efficiency in the ball mill by

reducing ball size. The authors predicted that optimization may be reached when 50-60% of the

total energy load is handled by the HPGR.

Pease (2007) found that the IsaMillTM can process particles sizes as coarse as 1mm. If the

main attributes of HPGR product, including the high proportion of fines and the presence of

micro-cracking, are also considered, a successful transition between the two pieces of

equipment should be feasible. To prevent coarse particles from entering the stirred mill and

causing critical build-up, either a screen or air classifier, is required. Since HPGR product tends

to form a cake, depending on the competency, either an impact style de-agglomerator or wet

screening is required. HPGRs operating in closed circuit to achieve a fine product cut size can

be found in the cement industry. An example of this circuit design is shown in Figure 2-13. An

impact style de-agglomerator is employed to break up flake and an air classifier is used to

prepare a fine product for ball milling (Aydogan et al., 2006).

Figure 2-13 HPGR Flowsheet for Fines Production in the Cement Industry (Aydogan et al., 2006)

The Sukhoy gold plant in Russia, shown in Figure 2-14, operates an HPGR in closed circuit with

a 1.4mm screen and utilizes a scrubber for product de-agglomeration. The main challenge

experienced in operating this circuit, was proper control of the moisture content in the HPGR

feed (Gruendken et al., 2010).

39

Figure 2-14 HPGR Flowsheet the Sukhoy Gold Plant (Gruendken et al., 2010)

Wang et al. (1998) and Wang et al. (2006) performed testing on an HPGR / stirred mill

combination using lab-scale equipment. Calcium carbonate and limestone wet filter cake

(<150µm) were subjected to multiple passes through a lab-scale HPGR and processed through

a horizontal stirred mill or attrition mill. Increasing the number of passes through the HPGR

resulted in improved throughput and size reduction in the stirred mill. The authors concluded

that the breakage mechanisms in HPGR comminution can lead to a subsequent increase in the

breakage rates in stirred milling.

Valery and Jankovic (2002) proposed the first concept of a combination HPGR / stirred mill

circuit in a study examining the need for a reduction in the energy requirements of comminution.

Simulating results for a more energy efficient circuit, a high-intensity blasting, two stage HPGR /

Vertimill® circuit (refer to Figure 2-15) was compared to a conventional blasting, SAG / ball mill

circuit. The simulation results predicted an energy savings of 45%, but no actual testwork was

conducted.

40

Figure 2-15 Example of an HPGR / Stirred Mill Circuit (Valery and Jankovic, 2002)

Figure 2-16 A Proposed HPGR / IsaMillTM Circuit (Pease, 2007)

Pease (2007) presented the concept of an HPGR / IsaMillTM circuit in his discussion of coarse

stirred milling at McArthur River (refer to Figure 2-16). No testing was carried out but Pease

predicted that this circuit could be an example of comminution flowsheet design of the future.

41

Ayers et al. (2008) described the first operation of an HPGR / IsaMillTM circuit using pilot-scale

equipment. The authors documented Anglo Platinum’s research into applying the IsaMillTM to

coarser feed applications. A test rig was set up, incorporating two 5tph HPGRs in series with

screens (refer to Figure 2-17).

Figure 2-17 Anglo Platinum’s HPGR Test Circuit (Ayers et al., 2008)

A continuously operating circuit was establishing using the coarse HPGR in closed circuit with a

dry screen, followed by wet screening of the undersize, at a cut size of 850µm (refer to Figure 2-

18). The screen product was fed to an M250 IsaMillTM operating with 3.5mm MT1 ceramic

grinding media. With an f80 of 300µm and a product p80 of 45µm, the IsaMillTM circuit achieved

1.3tph, with a specific energy consumption of 75kWh/t and a total circuit energy consumption of

80kWh/t.

Another circuit tested in this research project incorporated a ball mill before the IsaMillTM (refer

to Figure 2-18) and resulted in improved IsaMillTM circuit performance, but at a higher total

circuit specific energy consumption of 85kWh/t. In both cases the optimal media size was not

used and improved circuit energy requirements could have been achieved with an increased

media size.

42

Figure 2-18 Anglo Platinum’s HPGR / Stirred Mill Testing Flowsheets (Ayers et al., 2008)

The authors concludes with examples of future circuit arrangements to be tested by Anglo

Platinum, but as of the writing of this thesis, no published literature summarizing this work can

be found in the public domain.

43

2.4 Literature Summary

The HPGR can achieve improved energy efficiency over a SAG mill due to the application of

inter-particle breakage, and the ability to transfer input energy directly to the material via the

grinding rolls (Section 2.1.4). A stirred mill can achieve improved energy efficiency over a ball

mill, at fine particle sizes, due to reduced energy requirements associated with utilizing a

centrally-rotating shaft and the ability to grind efficiently with high speeds and small grinding

media (Section 2.2.5). The incorporation of these two energy-efficient comminution devices

could result in an overall reduction in the specific energy requirements for comminution. HPGR

product, with a high percentage of fines and micro-cracks, could successfully be transferred to a

stirred mill circuit with two successive passes through the HPGR. Since very few operating

examples were found (Section 2.3), pilot-scale testing would be required to successfully

determine appropriate design criteria.

Overall, combining an HPGR and a stirred mill to produce a more energy-efficient circuit has the

potential for a wide variety of processing and operating advantages. With the benefits of lower

operating costs related to reduced energy consumption and operation of an open-circuit

grinding configuration, coupled with the flexibility available to grind efficiently to finer particle

sizes with inert grinding media, this combination has the potential to be the future for energy-

efficient comminution.

44

3 Experimental Procedure

This chapter describes the methodology and the equipment used to address the objectives of

this research. The main objective was to examine the technical feasibility of combining an

HPGR and a stirred mill into a novel flowsheet. To achieve this objective, a pilot-scale testing

program was carried out on a copper-nickel sulphide ore from Teck Limited’s Mesaba deposit in

Minnesota.

Evaluation of the potential energy benefits for the proposed circuit design required a basis for

comparison. Lab and simulation work was carried out on two alternate comminution circuits, a

cone crusher / ball mill circuit and an HPGR / ball mill circuit. Results from this study were used

to draw conclusions on which of the three circuits required the lowest specific energy

consumption for comminution.

45

3.1 Definition of Comminution Circuits

The three circuits examined for the energy comparison study were: a cone crusher / ball mill

circuit, an HPGR / ball mill circuit and the novel HPGR / stirred mill circuit. The feed size to

each circuit was fixed at an f80 of 21mm, and a product p80 of 75µm was chosen as a suitable

feed size for flotation. The circuits were evaluated solely on the power consumed per tonne of

material in order to achieve an equivalently sized product from an equivalently sized feed.

Energy requirements of material handling equipment such as conveyors, pumps and screens

were not taken into account.

The approach in all cases was to determine an appropriate set of design criteria for each

flowsheet and to calculate the specific work index for each stage of comminution based on the

work index determined and the transfer sizes selected.

3.1.1 Cone Crusher / Ball Mill Circuit

The first circuit examined was a cone crusher / ball mill circuit, typically found in a three-stage

crushing flowsheet. This circuit was the industry standard for hard-rock comminution prior to the

establishment of SAG mill technology. The circuit comprised of a cone crusher in closed circuit

with a screen followed by a ball mill in closed circuit with a cyclone. The flowsheet of this circuit

is shown in Figure 3-1. Data for the circuit was generated from a combination of Bond

grindability testing and simulation using JK SimMet® software.

Figure 3-1 Cone Crusher / Ball Mill Flowsheet

Feed

Productp80 = 75µm

Cone Crusher

4mm Screen

Ball Mill

Cyclone

f80 = 21mm

46

3.1.2 HPGR / Ball Mill Circuit

The second circuit examined was an HPGR / ball mill circuit. This circuit mimics the standard

HPGR comminution flowsheet currently being used in the hard-rock mining sector (refer to

Section 2.1.5). The circuit comprised of a high pressure grinding roll in closed circuit with a

screen followed by a ball mill in closed circuit with a cyclone (refer to Figure 3-2). Data for this

circuit was generated using a combination of HPGR pilot-scale testing, Bond grindability testing

and simulation using JK SimMet® software.

. Figure 3-2 HPGR / Ball Mill Flowsheet

For HPGR pilot-scale evaluation, tests were carried out to assess the influence of different

process parameters on comminution performance. These tests included the variation of specific

pressing force, roller speed and feed moisture content, as well as closed-circuit testing with a

4mm screen. Data from this study was entered into JK SimMet® to model fit an appropriate

HPGR model for Mesaba ore. The T10H and HPGR power coefficient model parameters were

fitted using the procedure outlined by Daniel and Morrell (2004). After calibration of the HPGR

model, simulation was carried out for the HPGR / ball mill circuit.

3.1.3 HPGR / Stirred Mill Circuit

The final circuit tested comprised of two stages of HPGR followed by a horizontal stirred mill.

Since no operating examples were found in the literature, pilot-scale testing on both pieces of

equipment was performed to determine appropriate transfer sizes between each stage of

Feed

f80 = 21mm

Productp80 = 75µm

HPGR

4mm Screen

Ball Mill

Cyclone

47

comminution. From these tests, an appropriate circuit design and layout was generated for

comparison to the above mentioned circuits.

The appropriate transfer size between each stage of HPGR comminution was evaluated using

two separate flowsheet options. Option A examined the first-stage HPGR in open circuit (Figure

3-3), while Option B examined the first-stage HPGR in closed circuit with a 4mm screen (Figure

3-4). Each option was evaluated to assess how a change in transfer size between HPGRs

affected specific throughput, specific energy consumption and size reduction.

Figure 3-3 HPGR / Stirred Mill Flowsheet (Open Circuit)

Figure 3-4 HPGR / Stirred Mill Flowsheet (Closed Circuit)

For determination of transfer size between the second-stage HPGR and the stirred mill, three

different top sizes (355µm, 710µm and 1.2mm) were evaluated. These particle sizes tested the

limits for stirred mill grinding and an evaluation was based on the specific energy requirements

Feed

Product

f80 = 21mm

p80 = 75µm

HPGR

Fine Screen

Stirred Mill

HPGR

Feed

Product

Fine Screen4mm Screen

HPGRHPGR

f80 = 21mm

p80 = 75µmStirred Mill

48

for comminution and whether the stirred mill could operate and grind effectively. Once a

transfer size of 710µm was selected, pilot-scale testing was performed to generate data for the

circuit.

49

3.2 Sample Description

The Mesaba copper-nickel deposit is located in the Mesabi Range of the Duluth intrusive

complex located in North-eastern Minnesota (refer to Figure 3-5). This complex is comprised of

mafic volcanics (tholeiitic basalt) with layered intrusions of primarily a gabbro-troctolite

composite (Minnesota Geological Survey, 2010). Mineralogy of the Mesaba deposit comprises

mainly of massive and disseminated sulphides with the main minerals of interest being

chalcopyrite (copper), cubanite (copper) and pentlandite (nickel). The inferred resource stands

at 700Mt, with a grade of 0.46% Cu and 0.12% Ni (Infomine, 2001).

Figure 3-5 Geographic Location of Mesaba (Mayhew et al., 2009)

The samples used for this study were originally excavated as part of a bulk sample taken in

2001 after Teck Limited’s acquisition of the property. The majority of the sample was used to

create a bulk flotation concentrate for hydrometallurgical testing, while the remaining material

was kept stored on site. As part of hydrometallurgical testing, the head grade of the bulk

sample was determined to be 0.32% Cu and 0.12% Ni (Teck, 2010).

In December 2009, approximately 5 tonnes of sample, at nominally 100% minus 100mm, was

shipped to UBC. The material was screened and crushed in a laboratory jaw crusher to 100%

minus 32mm and homogenized and split into sixteen 45 gallon drums using a rotary sample

splitter. A representative sample was taken for size distribution, bulk density and moisture

50

content determination. A moisture content of 1% and a bulk density of 2.16t/m3 were

established for the ore. The particle size distribution of the sample is shown in Figure 3-6.

Figure 3-6 Mesaba Feed Size Distribution

51

3.3 Equipment

The following section describes the main pieces of test equipment used and the methodology

used for calculating specific energy consumption.

3.3.1 High Pressure Grinding Roll

HPGR testing was conducting using a pilot-scale unit manufactured by Koeppern. The pilot unit

is custom made for obtaining design information for sizing and selection of industrial-scale units.

Table 3-1 summarizes the technical data provided by Koeppern for the machine.

Table 3-1 HPGR Machine Specifics

Roller Diameter 750 mm

Roller Width 220 mm

Press Drive Dual Output Shaft Gear Reducer

Feed System Gravity

Wear Surface Hexadur® WTII

Installed Power 200 kW

Maximum Pressing Force 1800 kN

Maximum Specific Pressing Force 8.5 N/mm2

Variable Speed Drive up to 40 RPM (1.55 m/s)

Experimental data was recorded every 200ms through the programmable logic controller (PLC)

data logger and downloaded to a laptop. The computer system measures: time, roller gap (left

and right), pressing force (left and right) and power draw. A picture of the HPGR pilot unit is

shown in Figure 3-7.

.

.

52

Figure 3-7 Pilot-Scale HPGR Installation

A pilot test with the HPGR comprises the crushing of one 45 gallon drum of material (~375kg).

The material is loaded into a feed hopper with the use of an overhead crane and drum tipper.

Once the machine conditions are stabilized, the slide gate of the feed hopper is opened and the

test begins. The material flows with the aid of gravity through the HPGR rollers and drops on to

the product conveyor located below the rolls. Using the equations presented in Section 2.1.3,

specific throughput and specific energy consumption are then determined for the test.

Since the HPGR does not grind uniformly across the roller width, a splitter gate is installed on

the end of the product conveyor to separate the product into centre, edge and waste streams.

The centre portion is finer than the edge portion and during testing a particle size distribution is

performed on each to accurately predict size distributions for full-scale operations. For square

rollers, where roll diameter is equal to roll width, the proportion of centre and edge product is

observed to be approximately 85% centre and 15% edge for industrial units. All of the HPGR

product size distributions presented in Chapter 4 account for this through scaling of centre and

edge size distributions at a ratio of 85:15. Material collected during unstable operation, initial

response and material run-out periods, was designated as waste material and only material

which has been crushed during stable press operation was collected for analysis.

53

3.3.2 Horizontal Stirred Mill

Stirred mill testing was carried out using Netzsch’s M20 horizontal stirred mill. The mill has a

capacity of 20 litres and installed with an 18.6kW motor. The mill was upgraded to include a

new mechanical seal, updated grinding disc configuration and the replacement of a hand-crank,

variable-speed pulley system with a fixed pulley system and Variable Frequency Drive (VFD).

The installation of the VFD allowed for direct readings of mill power and mill speed. To monitor

the mill, sensors were installed for feed pressure, and both feed and product temperature. A

PLC interface and data logger was also installed to control the mill settings and record all

important mill parameters during testing. A picture of the upgraded mill is shown in Figure 3-8.

The mill configuration, including grinding disc design, was based on recommendations from

Xstrata Technology and allowed for the ability to scale-up results to what would be expected for

industrial units.

Figure 3-8 M20 Stirred Mill Installation

A Watson-Marlow and Bredel SPX 25 hose pump and corresponding VFD were used to feed

the mill. The pump has a capacity of 25L/min and was designed to handle viscous slurries. The

installation of a VFD for the 1.5kW pump motor allowed for accurate monitoring and control of

mill flow rate.

54

The mixing system was comprised of two 180L-capacity mix tanks with corresponding 250W

variable speed agitators and it was designed to mix slurries at upwards of 60% solids with a

particle top size as coarse as 1.2mm. The piping system for the circuit was setup so that each

mix tank could easily be switched from product to feed with little hesitation. The final setup is

shown in Figure 3-9.

Figure 3-9 M20 Stirred Mill with Mixing Tanks

For testing of the stirred mill energy requirements, a graph of specific energy consumption and

p80 grind size was generated. This graph, known as a signature plot, is the common method

used in industry for accurate sizing of full-scale IsaMillsTM and has a scale-up ratio of 1:1 (Gao

et al., 1999). The procedure entails running the material through the mill a select number of

times and recording the energy requirements and product size after each pass. The passes are

carried out consecutively in order to observe the energy consumption as the size of the product

decreases. The results provide a series of points plotted on a log – log graph that shows the

relationship between energy input and product size (p80).

55

Particle sizing for this work was done using a Malvern Mastersizer 2000. This laser sizing

equipment utilizes the principle that grains of different sizes diffract light at different angles; a

decrease in size produces an increase in diffracted angle. This equipment has become the

standard for analyzing size ranges unrealistic for screening (Larson et al., 2008).

3.3.3 Vibrating Screen

All screening work carried out for HPGR closed-circuit testing was performed using a SWECO®

Vibro-Energy® Separator. This vibrating screen, model ZS40, is equipped with a 373W motor

and a counterweight system to produce both vertical and horizontal vibrating motion. The

screener is equipped with 1m diameter wire mesh screens. A picture of the equipment is shown

in Figure 3-10.

Figure 3-10 ZS40 SWECO® Vibrating Screen

3.3.4 Bond Test Ball Mill

Energy requirements for ball mill grinding were determined using Bond ball mill work indices for

cone crusher and HPGR product. Representative samples were screened at minus 3.35mm

and processed through a standard Bond ball mill measuring 305mm in length and 305mm in

56

diameter, with a 285 ball charge weighing 20,125 g (refer to Figure 3-11). Testing was carried

out using the standard Bond Ball Mill Grindability Test procedure developed by Bond (1961).

For the crushing work index, insufficiently sized material was available to perform impact

testing; therefore a traditional approach was taken and the Bond work index was used. The

resulting indices were then used with the Bond equation to calculate specific energy

consumption for both crushing and grinding.

Figure 3-11 Bond Test Ball Mill

57

4 Testing and Simulation Results

The following chapter summarizes the results obtained for each circuit described in Section 3.1.

Simulation and lab results from each circuit are presented and the specific energy consumption

for comminution is calculated.

58

4.1 Cone Crusher / Ball Mill Circuit Results

4.1.1 Flowsheet Simulation

The specific energy consumption of comminution for the circuit was determined with a flowsheet

developed using JK SimMet® software. The circuit was designed for 250tph capacity and

equipment was sized based on a product p80 of 75µm. Table 4-1 summarizes the equipment

sized for the circuit.

Table 4-1 Equipment Selection for Cone Crusher / Ball Mill Circuit

Closed Side Setting (mm) Cone Crusher

2.8 Re-circulating Load (%) ~30%

Aperture Size (mm) Product Screen

4

Diameter (m) Ball Mill

5 Length (m) 10

Critical Speed (%) 70% Media Charge (%) 40%

Media Top Size (mm) 35 Re-circulating Load (%) ~250%

Quantity Hydrocyclones

6 Cyclone Diameter (mm) 420

Inlet Diameter (mm) 175 Vortex Finder Diameter (mm) 150 Apex (Spigot) Diameter (mm) 113

Length (mm) 500 Cone Angle (degree) 20

A flowsheet was simulated to determine the appropriate transfer size between the cone crusher

and the ball mill. The flowsheet generated in JK SimMet® is shown in Figure 4-1. Simulation

results determined that the appropriate transfer size between the crushing circuit and the ball

mill circuit would be 80% passing 2.12mm. For a summary of JK SimMet® results refer to

Appendix A.

59

Figure 4-1 Cone Crusher / Ball Mill JK SimMet® Flowsheet

60

4.1.2 Specific Energy Calculations

To calculate the overall specific energy consumption for the cone crusher and ball mill, work

indices were determined for the material. In the case of the cone crusher, no material was

available for the size requirements, 50-75mm, necessary to perform impact testing (Tavares

and Carvalho, 2007); therefore a traditional approach was taken and the Bond ball mill work

index was used. Locked-cycle testing was performed using two sieve sizes (106µm and

150µm), to allow for the comparison of different product sizes. The results for the work indices

of the circuit are shown in Table 4-2. For a complete breakdown of results refer to Appendix B.

Table 4-2 Bond Work Indices for Cone Crusher / Ball Mill Circuit

Ball Mill Work Index (kWh/t)

150 µm 15.9 106 µm 16.5

Using the Bond work indices and the transfer size determined in Section 4.1.1, the theoretical

energy requirements were calculated for the proposed circuit using the Bond equation (Bond,

1961).

𝐖 = 𝟏𝟎 ∗ 𝐁𝐖𝐢 ∗ � 𝟏�𝐩𝟖𝟎

− 𝟏√𝐟𝟖𝟎

� (5)

Where: W (kWh/t) = Specific Energy Consumption

BWi (kWh/t) = Bond Work Index

p80 (µm) = 80% Passing Product Size

f80 (µm) = 80% Passing Feed Size

Calculation of the cone crusher energy requirements used the Bond work index at a sieve size

of 150µm. This coarser screen size provided a lower estimate for the energy requirements of a

cone crusher and provides a best-case scenario for the crushing circuit. Calculation of the ball

mill energy requirements used the Bond work index at a sieve size of 106µm. The final product

for the circuit was set at a p80 of 75µm and the Bond work index, at a sieve size of 106µm,

better reflects the grinding energy requirements to grind to this finer particle size.

61

Cone Crusher Circuit

𝑊 = 10 ∗ 15.9 𝑘𝑊ℎ 𝑡⁄ ∗ � 1�2,120µ𝑚

− 1�21,000µ𝑚

�

Specific Energy Consumption for Crushing = 2.36kWh/t

Ball Mill Circuit

𝑊 = 10 ∗ 16.5 𝑘𝑊ℎ 𝑡⁄ ∗ � 1�75µ𝑚

− 1�2,120µ𝑚

�

Specific Energy Consumption for Grinding = 15.47kWh/t

TOTAL ENERGY CONSUMPTION = 17.83kWh/t

62

4.2 HPGR / Ball Mill Circuit

This section summarizes the testing and simulation work carried out to determine the specific

energy consumption required for an HPGR / ball mill circuit.

4.2.1 HPGR Pilot-Scale Testing

Section 2.1.2 stated that the only reliable method for assessing a material’s response to HPGR

comminution, and determining scalable operating parameters, is to perform pilot-scale testing.

The first parameter evaluated was the appropriate specific pressing force that achieved the

optimum balance of specific throughput, specific energy consumption and size reduction. Once

identified, testing was done using this pressing force to assess the effect of different operating

conditions on HPGR performance. A complete summary of HPGR operating data, including

particle size distributions, can be found in Appendix C.

Identifying Specific Pressing Force

Four initial tests were done to determine the effect of specific pressing force on the material.

Pressures of 2N/mm2, 3N/mm2, 4N/mm2 and 5N/mm2 were chosen and comparisons were

made with respect to product size, net specific energy consumption and m-dot. The feed

conditions for each test are shown in Table 4-3. All tests were performed at a roller speed of

0.75m/s.

Table 4-3 Feed Conditions for Pressing Force Tests

Moisture Content 2.5% Bulk Density 2.16 t/m3

f80 21.32 mm f50 13.96 mm

The comparison of product particle size at different specific pressing forces is shown in Figure

4.2. As the pressing force increased, both the p50 and p80 decreased, although the effect on

p80 was more pronounced than the effect on p50. This result is due to an increased force being

exerted on the particles as they flow through the rolls. An increased force would promote

increased breakage and the effect would be more pronounced on larger sized particles, hence

the steeper trend for p80.

63

Figure 4-2 Comparison of Specific Pressing Force and Product Size

The comparison of specific throughput (m-dot) at different specific pressing forces is shown in

Figure 4-3. As the pressing force increased, the specific throughput decreased. This trend is

due to the gap between the rollers decreasing slightly with increasing pressing forces, resulting

in the reduction of throughput in the machine.

Figure 4-3 Comparison of Specific Pressing Force and Specific Throughput

64

The comparison of specific energy consumption at different specific pressing forces is

summarized in Figure 4-4. As the pressing force increased, the energy consumption also

increased. This is typical of the process because more energy is being transmitted into the

material at higher pressures.

Figure 4-4 Comparison of Specific Pressing Force and Specific Energy Consumption

A specific pressing force of 4N/mm2 was selected for the remainder of pilot-scale testing. The

results indicated that a pressing force of 4N/mm2 provided a fine balance between energy

consumption and size reduction without a significant change in specific throughput.

Variation of Moisture Content

Section 2.1.6 described that moisture can have an adverse effect on HPGR comminution.

Testing was performed to assess the effect of moisture on size reduction, specific throughput

and energy consumption. Both drier (1%) and wetter (5%) tests were performed at the selected

specific pressing force (4N/mm2) with a roller speed of 0.75m/s. The effect on size reduction is

shown in Figure 4-5. An increase in moisture tends to increase the product size, although not to

a significant degree. This trend may result from the presence of moisture causing slippage on

the rolls and a slight decrease in the effectiveness of the compression bed.

65

Figure 4-5 Comparison of Moisture Content and Product Size

The comparison of specific throughput (m-dot) at different feed moisture contents is shown in

Figure 4-6. An increase in moisture content had a drastic effect on the specific throughput,

dropping it to 276.58ts/hm3. This may result from moisture acting as a lubricant, decreasing the

frictional forces between the material and the roll surface, and decreasing the operating gap.

Figure 4-6 Comparison of Moisture Content and Specific Throughput

66

The comparison of specific energy consumption at different feed moisture contents is shown in

Figure 4-7. An increase in moisture content caused an increase in energy consumption. This

may be caused by the wet material requiring more power to draw it through the rolls, coupled

with the decreased operating gap, reducing flow through the rolls.

Figure 4-7 Comparison of Moisture Content and Specific Energy Consumption

The results indicate that increased moisture content had a negative effect on HPGR

comminution performance and drier material produced better results. Unfortunately in

operations, dry material requires a complex dust suppression system, and therefore, a

compromise on moisture content must be made. The remainder of pilot-scale tests were

conducted at 2.5% moisture as a compromise.

Varying Roller Speed

The effect of roller speed on comminution performance was assessed using two tests at higher

(0.9m/s) and lower (0.6m/s) speeds. The effect on size reduction is shown in Figure 4-8. The

graph shows that roller speed has very little effect on size reduction.

67

Figure 4-8 Comparison of Roller Speed and Product Size

The comparison of roller speed with specific throughput is shown in Figure 4-9. An increase in

roller speed caused an increase in throughput. This should be expected, since a faster rotation

of the rolls causes an increase in the amount of material that can be processed in a given time

interval.

Figure 4-9 Comparison of Roller Speed and Specific Throughput

68

The effect on specific energy consumption at different roller speeds is shown in Figure 4-10.

The change in roller speed had no effect on the specific energy consumption.

Figure 4-10 Comparison of Roller Speed and Specific Energy Consumption

Closed Circuit Testing

To test the effect of closed-circuit operation, locked-cycle testing was conducted using a 4mm

screen. Material was processed through the HPGR at 4N/mm2 and the product screened at

4mm using a SWECO® 1m vibrating screen. Using the product size distributions from testing,

the percentage of minus 4mm was calculated (at 90% screening efficiency) and then combined

with fresh feed and re-run through the HPGR. This process was repeated two more times to

simulate closed-circuit operation. Results were generated to determine size reduction, specific

throughput and specific energy consumption for each cycle. The resulting product size for each

cycle is shown in Figure 4-11. The chart shows that the introduction of a re-circulating load

decreased the product size and began to stabilize by cycle four.

69

Figure 4-11 Product Size for Closed Circuit Testing

The effect on specific throughput for closed-circuit testing is shown in Figure 4-12. Closed-

circuit operation had little effect on specific throughput. The variation between each cycle can

probably be attributed to testing error.

Figure 4-12 Specific Throughput for Closed Circuit Testing

70

The results for specific energy consumption are displayed in Figure 4-13. The closed-circuit

testing had little to no effect on specific energy consumption.

Figure 4-13 Specific Energy Consumption for Closed Circuit Testing

Results from the last cycle of testing are summarized in Table 4-4. These results will be used

for energy calculations, as well as the experimental data required for model fitting with JK

SimMet®.

Table 4-4 Results for Cycle Four of Closed Circuit Testing

F80 21.77 mm

F50 13.38 mm

p80 6.61 mm

p50 1.91 mm

Percentage Passing -4 mm 67.4%

Net Specific Energy Consumption 1.45 kWh/t

(-4 mm) Net Specific Energy Consumption 2.15 kWh/t

Specific Throughput 304 ts/hm3

71

4.2.2 Flowsheet Simulation

Using the results from the last cycle of HPGR closed-circuit testing, model fitting of an HPGR

circuit was performed using JK SimMet®. The T10H and HPGR power coefficient model

parameters were fitted using the model fit tool in JK SimMet®. This tool uses an iterative

function to fit experimental data to simulated data by adjusting model parameters until a

correlation can be achieved. The T10h and HPGR power coefficient parameters relate to the

breakage mechanisms in the compression zone of the HPGR and the product size for closed-

circuit testing was used as the experimental data. The procedure used for calibrating the HPGR

model was outlined by Daniel and Morrell (2004). The resulting model fit was able to simulate a

product size distribution similar to the one generated experimentally.

Once an HPGR model was calibrated for use with Mesaba ore, a flowsheet was designed for

250tph capacity with a product p80 of 75µm. The equipment sized for the flowsheet is

summarized in Table 4-5.

Table 4-5 Equipment Selection for HPGR / Ball Mill Circuit

Roller Diameter (mm) High Pressure Grinding Roll

1,200 Roller Width (mm) 1,000

Re-circulating Load (%) ~45%

Aperture Size (mm) Product Screen

4

Diameter (m) Ball Mill

5 Length (m) 9.1

Critical Speed (%) 70% Media Charge (%) 40%

Media Top Size (mm) 27.5 Re-circulating Load (%) ~250%

Quantity Hydrocyclones

7 Cyclone Diameter (mm) 350

Inlet Diameter (mm) 175 Vortex Finder Diameter (mm) 150 Apex (Spigot) Diameter (mm) 113

Length (mm) 450 Cone Angle (degrees) 20

72

A flowsheet was simulated to determine the appropriate transfer size between the HPGR and

the ball mill. The flowsheet generated in JK SimMet® is shown in Figure 4-14. Simulation

results determined a transfer size of 80% passing 1,6mm between the HPGR circuit and the ball

mill circuit. For a summary of JK SimMet® results, refer to Appendix A.

73

Figure 4-14 HPGR / Ball Mill JK SimMet® Flowsheet

74

4.2.3 Specific Energy Calculations

Section 2.1.6 described that one of the benefits of HPGR comminution is particle weakening

and the subsequent reduction in energy requirements for ball mill grinding. To confirm this

advantage, Bond ball mill work indices were determined for HPGR product at different specific

pressing forces. Samples were taken from HPGR centre product and screened at 3.35mm with

no additional crushing. As with Bond work indices for cone crusher product, two separate

screens sizes (106µm and 150µm) were tested to allow for the comparison of different product

sizes. The results, including cone crusher product for comparison, are summarized in Table 4-6

and shown in Figure 4-15. For a complete breakdown of results, refer to Appendix B.

Table 4-6 Summary Bond Ball Mill Work Indices for Cone Crusher and HPGR Product

Bond Ball Mill Work Indices (kWh/t)

Locked Cycle Screen Size

(µm)

Cone Crusher Product

HPGR Product 3 N/mm2 4 N/mm2 5 N/mm2

150 15.9 14.5 14.5 13.3 106 16.5 15.8 15.7 15.7

Figure 4-15 Summary of Bond Work Indices

A reduction in Bond work index was achieved between cone crusher and HPGR product;

however the reduction went from 8.8% to 4.8% with a decrease in screen size. Results also

showed a reduction in Bond work index with increasing specific pressing force, although this

75

effect is reduced with a decrease in screen size. The reduction in screen size may have caused

a decrease in the effectiveness of product micro-cracking and a relatively higher amount of

energy was required to produce the finer product.

Using the results in Table 4-6, coupled with HPGR tests data, specific energy consumption can

be calculated for the circuit.

HPGR Circuit

Results from closed-circuit HPGR testing (Section 4.2.1) produced an HPGR Specific Energy

Consumption equal to 2.15kWh/t

Ball Mill Circuit

Using the Bond work index resulting from a specific pressing force of 4N/mm2 and a screen size

of 106µm, coupled with the transfer size determined in Section 4.2.2, the following energy

calculation can be made:

𝑊 = 10 ∗ 15.7 𝑘𝑊ℎ 𝑡⁄ ∗ � 1�75µ𝑚

− 1�1,600µ𝑚

�

Specific Energy Consumption for Grinding = 14.2kWh/t

TOTAL ENERGY CONSUMPTION = 16.35kWh/t

76

4.3 HPGR / Stirred Mill Circuit

The HPGR / stirred mill circuit required considerably more pilot-scale testing than the previous

two circuits, since very few operating examples could be found in the literature. The results of

pilot-scale testing determined the appropriate transfer size between each step of comminution

and provided the corresponding specific energy consumption for circuit energy summation.

4.3.1 The Stirred Mill Circuit

Section 2.2.6 described that the typical f80 for an IsaMillTM ranges from 30-150µm, usually

operating with 3.5mm or finer ceramic grinding media. Only a few examples were found of the

IsaMillTM grinding a coarser feed size; however these examples did not employ a large enough

media size (Ayers et al., 2008; Shi et al., 2009). Design of an HPGR / stirred mill flowsheet

requires selection of an effective feed top size for stirred mill grinding. Using properly sized

grinding media, transfer sizes were tested to achieve a balance between energy consumption

and practicality. Once an appropriate transfer size was selected, additional pilot-scale testing

was performed with improved operating conditions in order to assess the specific energy

requirements for grinding to a p80 of 75µm.

Three feed top sizes were chosen to cover a range from fine to coarse and included 355µm,

710µm and 1.2mm. Each feed size was run through an IsaMillTM to produce a signature plot to

compare the energy versus size reduction relationships.

Traditionally signature plots are performed using an M4 IsaMillTM, due to a small sample

requirement of only 15kg. The results from these tests have been found to accurately scale-up

1:1 to industrial sized units (Curry et al., 2005). Tests are typically conducted with a media top

size no larger than 3.5mm, but when running coarse grinding tests, a larger media size is

required. Unfortunately size constraints in the mill chamber prevent an M4 from operating

effectively with a larger media size. Therefore, when running tests at coarse particle sizes

above a p80 of 200µm, signature plots can no longer be run with an M4 unit and testing

requires a larger mill to operate effectively with media size above 3.5mm. For analysis of the

355µm sample, an M4 could still be used, but for the 710µm and 1.2mm samples, an M20 was

employed.

Testing of the 355µm sample was performed using a graded charge (57.5% 3mm, 30% 2mm

and 12.5% 3.5mm) of MT1 ceramic grinding media manufactured by Magotteaux International.

77

The operating conditions for the test followed the standards laid out in Section 2.2.4 and are

summarized in Table 4-7. An f80 of 204µm was determined for the feed. This corresponds to a

media to mean (80% passing) particle size ratio of 17.5:1. Had the 20:1 ratio suggested by

Mankosa et al. (1986) been applied, the media top size would have been 4mm. For a complete

breakdown of test results, refer to Appendix D.

Table 4-7 Test Conditions for the 355µm Signature Plot

f80 204 µm Feed Wt. 15 kg

Solids Content 43% Flow Rate 2.5 L/min

Media Volume 70% Mill Speed 1,215 RPM

A graph of the resulting signature plot is shown in Figure 4-16. The graph shows that grinding

to a product p80 of 75µm required 14.17kWh/t.

Figure 4-16 Signature Plot for Top Size Testing of 355µm

Testing of the 710µm sample was performed using a composite of two types of media: 70% by

volume 8mm diameter ceramic media manufactured by Rojan Advanced Ceramics and 30% by

volume 3mm diameter MT1 ceramic media manufactured by Magotteaux International. In

retrospect, this mixture may not have been ideal for testing. The media top size appears to

78

have been oversized for the application, as well as a combination of two different types of

ceramic media, with different sizes and SGs, may not have been as efficient as a graded charge

comprising of a single type of media. The test conditions used for the signature plot are shown

in Table 4-8. An f80 of 321µm corresponds to a media to particle size ratio of 25:1. Had the

20:1 ratio suggested by Mankosa et al. (1986) been applied, the media top size would have

been 6.4mm. For a further breakdown of results, refer to Appendix D.

Table 4-8 Test Conditions for the 710µm Signature Plot

f80 321.4 µm Feed Wt. 75 kg

Solids Content 40% Flow Rate 12.15 L/min

Media Volume 70% Mill Speed 1,169 RPM

The resulting signature plot is shown in Figure 4-17. The graph shows that the first pass

produced a very fine product (p80 of 38µm) and that extrapolation was necessary for calculating

the specific energy consumption to grind to a p80 of 75µm. With extrapolation, an estimated

specific energy consumption of 14.72kWh/t was required.

Figure 4-17 Signature Plot for Top Size Testing of 710µm

y = 3839.1x-1.289

R² = 0.9916

y = 10469x-1.232

R² = 0.9715

10.0

100.0

1000.0

10.0 100.0

Net S

Peci

fic E

nerg

y Co

nsum

ptio

n (k

Wh/

t)

Particle Size (µm)

p80 p98

79

For testing the 1.2mm sample, the same ratio of 8mm Rojan Ceramic media and 3mm

Magotteaux media was used. This testing was expected to approach the limits of feasibility for

coarse grinding in a stirred mill, and as a result, numerous tests were conducted at a variety of

test conditions. Since coarse particles may have trouble passing through the dynamic classifier

and exiting the mill, a test was conducted to assess whether removal of dynamic classifying

pegs (refer to Figure 4-18) helped improve coarse grinding. Table 4-9 summarizes the

operating conditions for each test.

Figure 4-18 Stirred Mill Dynamic Classifier Pegs Six Classifying Pegs (Left) and Three Classifying Pegs (Right)

Table 4-9 Summary of Mill Operating Conditions for 1.2mm Testing

Test No. f80 (µm)

Solids Content

(%)

Media Volume

(%)

Mill Speed (RPM)

Mill Flow Rate (L/min)

Number of Dynamic

Classifier Pegs 1 500 40 70 1,200 14.6 6 2 500 40 70 1,050 12.1 6 3 500 40 70 1,050 11 3

The first test was unable to grind coarse particle sizes with the operating conditions selected.

The first pass through the mill recorded a power draw 56% higher than the allowable power

rating for the motor. The VFD safety featured shut down the mill at the 75 second mark to

prevent damage to the motor. The stirred mill was attempting to grind material from a feed top

size of 1.2mm to a product p80 of 50µm in one pass, requiring more power than was available.

A summary of key mill parameters is shown in Figure 4-19.

80

Figure 4-19 Summary of Mill Parameters for 1.2mm Test #1

The second test was also unable to successfully grind coarse particle sizes, with similar results

to test one. The VFD safety feature shut down the mill at the 2 minute mark because the mill

was drawing 26% more power than the motor could handle. A summary of key mill parameters

is shown in Figure 4-20.


0

200

400

600

800

1000

1200

1400

0 25 50 75 100

Time (s)

Mill

Sha

ft Sp

eed

(rpm

)

0

10

20

30

40

50

60

Mot

or P

ower

(kW

) , F

lowr

ate (

l/m) ,

Pro

duct

Te

mp

(deg

rees

C)

Shaft Speed (shaft sensor) Product Temp Flowrate (pump spd) Motor Power

0

200

400

600

800

1000

1200

1400

0 25 50 75 100 125 150Time (s)

Mill

Sha

ft Sp

eed

(rpm

)

0

10

20

30

40

50

60

Mot

or P

ower

(kW

) , F

lowr

ate (

l/m) ,

Pro

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Te

mp

(deg

rees

C)


81

The third test was carried out to assess whether removal of the dynamic classifier pegs helped

improve coarse grinding. The mill recorded a dramatic drop in power draw during the first pass;

however an increase in feed pressure at the 380 second mark resulted in the mill shutting down.

The main mill parameters plotted in Figure 4-21 show that during testing, the motor power

slowly increased and the shaft speed slowly decreased. This is consistent with a hypothesis

that, due to a decrease in energy input, critical sized particles were being retained in the mill;

slowly building up until the mill chamber was full, triggering an increase in pressure at the feed

inlet.


For the range of operating conditions tested, grinding a feed size of 1.2mm was not possible;

however if a higher rated motor was installed, the possibility of operating at these conditions

may have been successful.

With the results presented above, a transfer size of 710µm was selected as the most feasible

option for an HPGR / stirred mill circuit. Although a transfer size of 355µm produced the lowest

energy requirement for grinding to a p80 of 75µm, the savings of 0.5kWh/t was not sufficient to

overcome the increased energy requirements for operating an HPGR in closed circuit with a

355µm screen.

0

200

400

600

800

1000

1200

1400

0 25 50 75 100 125 150 175 200 225 250 275 300 325 350 375 400 425Time (s)

Mill

Sha

ft Sp

eed

(rpm

)

0

10

20

30

40

50

60

Mot

or P

ower

(kW

) , F

lowr

ate (

l/m) ,

Pro

duct

Te

mp

(deg

rees

C)


82

After designating 710µm as the transfer size between the second-stage HPGR and the stirred

mill, two additional pilot-scale tests were performed to attempt to reduce the specific energy

requirements for the circuit. Operating conditions were chosen to target a specific energy input

of 7-9kWh/t per pass through the mill. This would create a coarse first pass for the signature

plot, eliminating the problem of extrapolation experienced with the original test. The first

grinding disc on the mill shaft was replaced with a spacer (refer to Figure 4-22) to reduce the

energy input of the mill. A summary of the revised operating conditions for testing are shown in

Table 4-10. For a complete summary of results, refer to Appendix D.

Figure 4-22 Replacement of Grinding Disc with Spacer

Table 4-10 Revised Test Conditions for 710µm Signature Plots

f80 340 µm Feed Wt. 100 kg

Percent Solids 57% Flow Rate 20.4 L/min Mill Speed 1,169 RPM

Improvements were also made in the selection of grinding media for the tests. A graded charge

(40% 4-6mm, 30% 2-4mm, 20% 2.8-3.2mm and 10% 2-2.2mm) of Cenotec ceramic grinding

media was used at a media charge volume of 70%. The ratio of media top size (6mm) to f80

(340µm) was adjusted, representing a ratio of 17.6:1. The results for both tests are shown in

Figure 4-23. The graph shows that substantially reduced energy consumption was achieved

with properly sized media and revised operating conditions.

83

Figure 4-23 710µm Signature Plot Results with Revised Operating Conditions

Testing showed an average specific energy consumption of 9.73kWh/t was required to grind to

a p80 of 75µm, an improvement of 34% over the original test. This value was selected to be

used for energy calculations of the HPGR / stirred mill circuit.

The size measurements used to generate the signature plots in Figure 4-23 were performed

using a Malvern Mastersizer 2000. For a comparison, pass one product was also sized using

screens. Since Malvern sizing is based on volume, while screening is based on weight, results

will not be identical. All other testwork performed for this flowsheet relied on size results from

screening; therefore a comparison should be made. Malvern and screening comparisons for T1

and T2 are shown in Figure 4-24 and Figure 4-25, respectively. The screening results indicated

a finer product than the Malvern results. These results show that the signature plots generated

using Malvern sizing, can be considered a conservative estimate for energy consumption, since

size results may have been finer using screens. Unfortunately screening is impractical below

38µm (the product size after pass two), so Malvern sizing was used for all stirred mill products in

order to remain consistent.

84

Figure 4-24 Malvern and Screening Comparison for T1

Figure 4-25 Malvern and Screening Comparison for T2

4.3.2 The HPGR Circuit

Pilot-scale testing was conducted to produce suitable data for the HPGR section of the HPGR /

stirred mill circuit. Since size reduction is limited with one stage of HPGR comminution (refer to

Section 2.3), design of an HPGR / stirred mill circuit required at least two consecutive stages of

85

HPGR comminution to produce a particle size acceptable for stirred milling. With the transfer

size between the second-stage HPGR and the stirred mill determined in Section 4.3.1, work

was done to determine the appropriate transfer size between each stage of HPGR crushing.

Two options were examined to find the appropriate circuit layout. In Option A, the first-stage

HPGR was placed in closed circuit with a 4mm screen, while in Option B, the first stage

remained open circuit and the second stage accepted product directly from stage one.

For Option A, product from closed-circuit testing in Section 4.2.1 was processed again through

the HPGR at the same roller speed (0.75m/s) and specific pressing force (4N/mm2). The use of

the same specific pressing force for second-stage HPGR crushing stems from work performed

by Rule et al. (2008), in which they found that no difference was observed when changing the

specific pressing force in the second stage of two-stage HPGR crushing. For Option B, fresh

feed was processed through two consecutive stages of HPGR comminution using the same

roller speed and specific pressing force as Option A. The results for both options are

summarized in Table 4-11. The size distributions for Options A and B are presented in Figure

4-26 and Figure 4-27, respectively. For a complete breakdown of results, refer to Appendix C.

86

Table 4-11 Summary of HPGR Results for First Stage Open and Closed Circuit Testing

OPTION A OPTION B

HPGR Stage 1

f80 21.77 mm 21.54 mm f50 13.38 mm 13.7 mm

HPGR p80 6.61 mm 7.68 mm HPGR p50 1.91 mm 1.88 mm Circuit p80 1.86 mm 7.68 mm Circuit p50 489 µm 1.88 mm

Circuit Reduction Ratio 11.7 2.8 Net Specific Energy Consumption 1.45 kWh/t 1.54 kWh/t

Percentage Passing 4 mm 67.4% (-4 mm) Net Specific Energy Consumption 2.15 kWh/t

Specific Throughput 304 ts/hm3 307 ts/hm3

HPGR Stage 2

f80 1.86 mm 7.68 mm f50 489 µm 1.88 mm

HPGR p80 1.12 mm 2.79 mm HPGR p50 222 µm 462 µm Circuit p80 332 µm 339 µm Circuit p50 124 µm 142 µm

Circuit Reduction Ratio 5.6 22.6 Net Specific Energy Consumption 1.2 kWh/t 1.23 kWh/t

Percentage Passing 0.71 mm 71.3% 56.5% (-0.71 mm) Net Specific Energy Consumption 1.68 kWh/t 2.18 kWh/t

Specific Throughput 236 ts/hm3 311 ts/hm3 TOTAL SPECIFIC ENERGY CONSUMPTION 3.83 kWh/t 3.72 kWh/t

87

Figure 4-26 Particle Size Distributions for Option A

Figure 4-27 Particle Size Distributions for Option B

88

Operating the first stage of HPGR crushing in open circuit required less energy compared with

operating in closed circuit with a screen. If looked at strictly from an energy perspective, very

little difference is gained choosing one circuit over the other, but if design and operating factors

are considered, the choice of operating the first stage in open circuit becomes the better option.

The ability to operate the circuit without a screen allows for the elimination of extra auxiliary

equipment such as screens and conveyors, while the absence of an additional stage of wet

screening would help to reduce the adverse effects that increased moisture content would have

on HPGR performance. Although the increased re-circulating load resulting in the second stage

would require an increase in tonnage and machine size, this would be countered by the

decreased machine size required for stage one. Overall, the reduced complexity offered by

open circuit configuration led to selecting this configuration for further testing.

Once the open circuit configuration was selected for stage one, additional pilot-scale testing was

performed to evaluate how comminution performance would be affected by operating the

second sage in closed circuit with a 710µm screen. Testing was conducted in a similar manner

to Section 4.2.1. Product from Option B was screened at 710µm and a calculated split of

oversize was mixed with fresh product from stage one and processed through the HPGR. This

procedure was repeated two more times in order to simulate closed-circuit operation. The

resulting product size for each cycle is shown in Figure 4-28. The product size increased

slightly with the introduction of a re-circulating load. This is in contrast to the results in Section

4.2.1, where the introduction of a re-circulating load caused a decrease in product size. This

increase may have been the result of a finer re-circulating load reducing the breakage within the

compressive bed. For a complete breakdown of results, refer to Appendix C.

89

Figure 4-28 Product Size for Second Stage Closed Circuit Testing

The results for the effect of closed-circuit operation on specific throughput are displayed in

Figure 4-29. The introduction of a re-circulating load had no effect on specific throughput.

Figure 4-29 Specific Throughput for Second Stage Closed Circuit Testing

90

The results for the effect of closed-circuit operation on specific energy consumption are

summarized in Figure 4-30. As with specific throughput, the introduction of a re-circulating load

had no effect on specific energy consumption.

Figure 4-30 Specific Energy Consumption for Second Stage Closed Circuit Testing

To achieve efficient screening at 710µm for an industrial operation, the practice of wet screening

is necessary. Section 2.1.6 showed that the introduction of moisture to an HPGR circuit leads

to adverse effects on throughput and energy consumption. The effect of moisture on second-

stage HPGR crushing was tested using product from the final closed-circuit cycle. The sample

was wet screened over a 710µm screen to determine the potential moisture content for oversize

in a closed-circuit operation. The saturated oversize, with a measured moisture content of

10.5%, was then used to run an additional closed-circuit cycle. A summary of the results are

presented in Table 4-12 To allow for a direct comparison of the effects of wet screening, the

results from cycle four (dry) are presented as well. For a complete breakdown of results, refer

to Appendix C.

As expected, the results show an adverse effect on throughput and energy consumption,

although the product size became considerably finer. The data generated for the wet screening

cycle represents the worst-case scenario, and thus will be used for the energy calculations for

the circuit.

91

Table 4-12 Comparison of Wet and Dry Screening

Dry Cycle Wet Cycle

Feed Moisture Content 2.4% 5.8% F80 5.69 mm 6.41 mm F50 1.79 mm 1.95 mm p80 3.16 mm 2.88 mm p50 718 µm 523 µm

Percentage Passing -710µm 49.8% 54.8% Net Specific Energy Consumption 1.45 kWh/t 1.96 kWh/t

(-710µm) Net Specific Energy Consumption 2.91 kWh/t 3.58 kWh/t Specific Throughput 304 ts/hm3 232 ts/hm3

4.3.3 Circuit Energy Summary

The results obtained from pilot-scale testing and the resulting specific energy consumption for

the circuit are summarized in Table 4-13. For a comparison, results from dry and wet screening

for second-stage HPGR are included. Results show that the implementation of wet screening

would result in a 4.7% increase in specific energy consumption for the circuit.

Table 4-13 Summary of HPGR / Stirred Mill Energy Requirements

Comminution Circuit

Section Reference

Specific Energy Consumption with

Dry Screening (kWh/t)

Specific Energy Consumption with

Wet Screening (kWh/t)

First Stage HPGR 4.32 1.54 1.54 Second Stage HPGR 4.32 2.91 3.58

Stirred Mill 4.31 9.73 9.73

Total Specific Energy Consumption 14.18 14.85

92

5 Discussion of Results

The following chapter describes research outcomes based on results presented in Chapter 4.

With these results, conclusions are reached on the choice of operating parameters for pilot-

scale testing and what changes could be made for future testing. This chapter will also discuss

which circuit required the lowest specific energy consumption for comminution; while preliminary

design work is presented of a potential flowsheet for an HPGR / stirred mill circuit. The chapter

concludes with a refined testing procedure for future HPGR / stirred mill studies.

93

5.1 Assessing Operating Parameters for Pilot-Scale Testing

Operating parameters used to conduct HPGR and stirred mill testing were chosen based on a

review of the literature and recommendations made from industry. This section will review the

main parameters used for testing and comment on future variations to be studied.

5.1.1 HPGR Operating Parameters

The main operating parameters identified in Section 4.2.1 as having the most influence on

comminution performance were specific pressing force and feed moisture content.

Specific Pressing Force

The specific pressing force selected for testing of Mesaba ore was 4N/mm2. This force provided

a balance of size reduction and specific energy consumption without drastically changing the

specific throughput. This force also approaches the limits for safe operation with studded lining.

Morley (2008) indicated that the safe operating range for studding lining is 1-4.5N/mm2 and

anything higher risks damaging the metal studs. Therefore selection of 4N/mm2 can be safely

operated in industrial units and provides good size reduction at low specific energy

consumption.

The choice to keep specific pressing force constant for second-stage HGPR crushing was

based on results presented by Rule et al. (2008) and summarized in Section 2.3, in which

varying specific pressing force had little effect on size reduction for second-stage HPGR

crushing in a Labwal. Further testing could be done to confirm this conclusion with pilot-scale

results.

Feed Moisture Content

The feed moisture content used during testing was 2.5% by weight. The moisture content was

selected taking into account results obtained in Section 4.2.1. Lower moisture content resulted

in improved HPGR comminution performance; however dust suppression must be accounted

for. The optimum results occurred at 1%, when specific throughput was high, specific energy

consumption was low and product size was finer. With increasing moisture content, a decrease

in comminution performance was observed. Unfortunately operating at 1% moisture is

unfeasible due to excessive dust generation. In both pilot-scale testing and industrial

94

operations, dust suppression is an issue and higher moisture content mitigates these risks. The

selection of 2.5% provided a good balance in comminution performance and a substantial

reduction in dust generation. For industrial applications, moisture content is rarely as low as 1%

and values above 2-3% are more realistic.

Moisture content testing for second-stage HPGR crushing (Section 4.3.2) showed a

considerable increase in specific energy consumption when addition of saturated oversize was

used for closed-circuit testing. The effect of wet screening resulted in an increase in moisture

content to 5.8%. The data generated for this test provide an indication of what is expected

when an HPGR operates in closed circuit with wet screening. Further testing could be done to

examine a full range of moisture contents for second-stage HPGR crushing.

5.1.2 Stirred Mill Operating Parameters

The main operating parameters discussed in Section 2.2.4 for stirred milling were feed density,

flow rate, mill speed and media size. Section 4.3.1 found that mill geometry also has an effect

on grinding.

Feed Density

The average solids content used for 710µm stirred mill testing was ~57% by weight and the SG

of the ore was measured to be 3.0. This resulted in a solids content by volume of 30.6% and a

slurry SG of 1.61. This solids content was chosen to reduce grinding, by decreasing the amount

of energy transmitted per pass through the mill. Normally a solids content of 50% is

recommended to achieve optimum results. Larson et al. (2008) claims that operating with

higher solids content reduces efficiency due to increased viscosity. For testing Mesaba ore,

viscosity issues were negligible because power draw dropped with subsequent passes through

the mill. High viscosity would have caused power draw to remain constant or increase with finer

size. The sharp product size distribution produced during testing resulted in a lower ultrafines

content (<15µm), which allowed solids content to remain high without clogging the mill. Further

testing could be performed to determine whether a decrease in solids content has an improved

effect on grinding efficiency.

95

Mill Flow Rate

The flow rate chosen for feeding slurry to the mill was 20.4L/min. This flow rate was chosen to

reduce the grinding residence time in the mill. An increased flow rate resulted in a coarser

product, a desirable characteristic when generating results for grinding to a p80 of 75µm. Mill

flow rate does not have an effect on grinding efficiency because feed and mill properties are not

changing. The only change is the speed at which particles travel through the mill. The only

circumstances where mill flow rate would affect stirred mill operations are at very high and very

low flows. If too high a flow rate was selected, material may have a hard time exiting the mill,

resulting in clogging the mill. If too low a flow rate was selected, solids would have time to settle

in the feed line, resulting in blockage. The flow rate chosen for testing exhibited neither

situation and no further testing is required to assess different flow rates.

Mill Speed

The speed chosen for testing was set at 1,200 RPM with 1,169 RPM achieved, the maximum

attainable speed for the mill with the currently installed motor and drive system. This

corresponds to a tip speed of ~11m/s. Larson et al. (2008) found that mill speed has little

influence on grinding efficiency at high speeds. The authors claim that mill speed affects the

power draw of the mill and the energy transmitted to the material but does not affect the grinding

efficiency. Further testing could be done to validate this claim for coarse stirred milling by

running tests at higher and lower speeds.

Media Size

The size of grinding media used in testing was a graded charge with a top size of 6mm. The

recommended ratio for fine grinding is 20:1 media size to mean particle size (80% passing).

The f80 for 710µm testing was 340µm, corresponding to a ratio of 17.6:1. Had a 20:1 ratio been

used, the media would have had a top size of 6.4mm. Since only one media size was used,

further testing could be done to confirm an optimum bead size. Different media top sizes could

be tested, ranging from 3-8mm, improving the media selection guidelines for coarse stirred

milling.

96

Mill Geometry

The mill geometry was varied during stirred mill testing by removing the first grinding disc and

reducing the number of classifying pegs. The former resulted in a decrease in grinding energy,

while the latter resulted in a drop in grinding efficiency. Further work could be done to examine

the spacing of grinding discs and the effect on grinding efficiency.

Feed Top Size

The maximum feed top size tested for coarse stirred milling was 1.2mm. The operating

conditions chosen for these tests did not result in effective grinding. Further work could be done

to evaluate this top size with a larger-sized motor. With the use of a larger motor size,

increased power draw could be achieved and some of the operating conditions selected could

be fully explored. Further testing could also be done on an intermediate top size of 1mm.

97

5.2 Comparison of Comminution Circuits

From the conclusions reached in Chapter 2, it was predicted that the incorporation of both an

HPGR and a stirred mill into a novel circuit design, could lead to a reduction in the specific

energy requirements for comminution. In Chapter 4, lab and simulation work was done to

investigate whether these conclusions were valid. The resulting transfer sizes, operating work

indices and circuit layout are summarized in Figures 5-1 to 5-3.

Figure 5-1 Summary Layout for Cone Crusher / Ball Mill Circuit

Figure 5-2 Summary Layout of HPGR / Ball Mill Circuit

Feed

Product

f80 = 21mm

WI = 15.47 kWh/t

t80 = 2.12mmp80 = 75µm

WI = 2.36 kWh/t

Feedf80 = 21mm

WI = 2.15 kWh/t

Product

WI = 14.2 kWh/t

t80 = 1.6mmp80 = 75µm

98

Figure 5-3 Summary Layout of HPGR / Stirred Mill Circuit

Using the results presented above, a bar graph was generated to summarize the specific

energy consumption for each stage of comminution (refer to Figure 5-4). The graph shows that

the HPGR / stirred mill circuit required the lowest specific energy consumption and achieved a

reduction of 9.2% and 16.7% over the cone crusher / ball mill and HPGR / ball mill circuits,

respectively.

The results presented in this thesis were obtained from pilot-scale testing on a single test for

each operating variable. Since pilot-scale testing required a significant quantity of material per

test, 350kg for HPGR and 100kg for stirred mill, the reproducibility and standard deviation could

not be determined for each changing variable; however some repeatability testing was

performed on pilot-scale HPGR testing using 5 homogenized drums. Results showed that

specific energy consumption had a standard deviation of 0.0167 and specific throughput a

standard deviation of 11.43. For stirred mill testing, since only two signature plots were

generated at similar conditions, the standard deviation could not be calculated and instead the

median of 0.23 was considered. The energy figures associated with Bond grindability testing

were found to have a standard deviation of 0.0548, when comparing the three results of HPGR

product at a screen size of 106µm. With these results, testing errors were calculated for each

circuit at a 95% confidence interval. Table 5-1 summarizes the statistics related to each circuit

energy result.

Feed

Productt80 = 0.34mm

WI = 3.58 kWh/t

f80 = 21mm t80 = 6.4mm

p80 = 75µm

WI = 9.73 kWh/t

WI = 1.54 kWh/t

99

Figure 5-4 Summary of Specific Energy Consumption for Each Circuit

Table 5-1 Statistics Summary of Circuit Energy Values

The error values displayed in Table 5-1 show that the HPGR / stirred mill circuit contained the

most potential for a variation in reported results. With the inclusion of testing error, the HPGR /

stirred mill circuit still required the lowest specific energy consumption for comminution. The

testing error presented here can be considered only an approximation because the results are

based on only a few tests and the accuracy of JK SimMet® modelling is not accounted for.

Although JK SimMet® is a sophisticated tool for process simulation; the outputs are still dictated

by a mathematical model and are not a result of actual testing.

Sample Set

Mean (kWh/t)

Standard Deiation

Standard Deviation of the

Mean

95% Confidence

IntervalUpper Limit

Lower Limit

Cone Crusher Specific Energy Value 3 2.36 0.0548 0.0316 0.0620 2.42 2.30 Ball Mill Specific Energy Value 3 15.47 0.0548 0.0316 0.0620 15.53 15.41

HPGR Energy Value 5 2.15 0.0167 0.00747 0.0146 2.16 2.14 Ball Mill Energy Value 3 14.2 0.0548 0.0316 0.0620 14.26 14.14

Stage 1 HPGR Energy Value 5 1.54 0.0167 0.00747 0.0146 1.55 1.53 Stage 2 HPGR Energy Value 5 3.58 0.0167 0.00747 0.0146 3.59 3.57 Stirred Mill Energy Value* 2 9.73 0.23 0.163 0.319 10.05 9.41

*Median used intead of standard deviation17.83 +/- 0.0916.35 +/- 0.0614.85 +/- 0.32

Cone Crusher / Ball Mill Circuit HPGR / Ball Mill Circuit HPGR / Stirred Mill Circuit

Total Specific Energy Consumption With 95% Confidence Interval

100

The energy values determined in this study did not take into account any auxiliary equipment for

the circuits. Each circuit would require additional energy requirements for feeders, conveyors,

screens, pumps and cyclones. Additional energy requirements for the cone crusher / ball mill

circuit would result from screens and conveyors for the crushing circuit and pumps and cyclones

for the ball mill circuit. For the HPGR / ball mill circuit, increased energy requirements would

result from screens and conveyors in the HPGR circuit and pumps and cyclones in the ball mill

circuit. The energy requirements for the HGPR / stirred mill circuit would increase with a feed

conveyor for first-stage HPGR crushing, screens and conveyors for second-stage HPGR

crushing and pumps for the stirred mill circuit. The extra energy required for the increased

quantity of conveyors would be counteracted by the reduction in energy related to an open-

circuit grinding configuration. The energy requirements for a de-agglomerator were not

necessary for Mesaba ore due to a low flake competency; however this energy requirement

would need to be considered for an ore that produced more competent flake. Overall, the

increased energy requirements for all three circuits when incorporating auxiliary equipment

should not affect the results significantly.

The product size distribution generated from each circuit is shown in Figure 5-5. The cone

crusher / ball mill circuit and HPGR / ball mill circuit were created from JK SimMet® simulation of

the cyclone overflow product; while the HPGR / stirred mill circuit was generated from product

obtained from pass one of 710µm testing. The graph shows that the product from each circuit

resembles the same trend. The stirred mill product is shown to be a bit coarser; however the

product produced in pass one had a p80 of 83.8µm, so only the shape of the curve should be

compared. Although these size distributions resemble a similar trend, the size distributions of

different mineral phases could be considerably different. A softer material, possibly the valuable

sulphide minerals, could be much finer than is shown here.

101

Figure 5-5 Product Size Distributions for Each Comminution Circuit

102

5.3 Preliminary HPGR / Stirred Mill Circuit Flowsheet

A preliminary design of an HPGR / stirred mill circuit is based on pilot-scale testing summarized

in Section 4.3, coupled with HPGR flowsheet requirements presented in Section 2.1.5. The

current status of HPGR flowsheet design incorporates a closed-circuit crusher prior to HPGR

processing. This step prevents oversized material from entering the HPGR circuit, ensuring the

prevention of single particle breakage and damage of the metal studs. A metal detector must

be installed on a conveyor prior to entering the HPGR circuit to protect the HPGR roller lining

from tramp metal.

For configuration of the HPGR circuit, Section 4.3.2 concluded that two stages of HPGR

comminution were required to produce a product fine enough to grind efficiently in a stirred mill.

Section 4.3.2 also concluded that the feasible layout for these two stages would place the first

stage in open circuit and the second stage in closed circuit with a screen. This layout would

provide the advantage of reducing circuit complexity, while similarly reducing the water addition

to the circuit. To achieve efficient screening for the second stage, inclusion of both a de-

agglomeration step and wet screening could be incorporated to handle competent flake.

To ensure optimum feed density for efficient stirred milling, undersize from the HPGR circuit

would be fed to a mixing tank, where water would be added to control pulp density. Section

2.2.4 stated that ideal operating conditions for stirred milling require a solids content of 40-50%.

This could be achieved with a simple process control loop installed between the water addition

tank and the IsaMillTM feed tank. For operation of the IsaMillTM circuit, the simplicity available

with the dynamic classifier and open-circuit configuration would limit any need for a recycle

stream. A preliminary design of the circuit is shown in Figure 5-6. The design of the secondary

crushing circuit was similar to the configuration installed at Cerro Verde (Section 2.1.5). A cone

crusher is placed in a reversed closed-circuit arrangement, which would reduce throughput and

improve crushing efficiency for the cone crusher by screening out fine particles from the feed.

103

Figure 5-6 Preliminary Layout for an HPGR / Stirred Mill Circuit

Coarse Ore Stockpile

Cone Crusher

Metal Detector

HPGR

Coarse Screen

HPGR

Impact De-agglomerater

Fine Screen

IsaMill Feed Tank

IsaMill

Product

Water Addition

104

5.4 Refined Procedure for Future Testing

Chapter 4 showed that a considerable amount of testing was conducted to arrive at conclusions

on which circuit required the lowest specific energy consumption for comminution. This large

volume of testing required a considerable amount of material, a quantity not often available for

greenfield operations. Reducing sample requirements would allow greenfield operations to

conduct a preliminary estimation of the energy benefits possible for an HPGR / stirred mill

configuration. To this end, a refined testing program was developed with a reduced sample

commitment of four 45 gallon drums or 1,200kg.

The refined procedure would require a feed particle size of minus 150mm. This would provide

the size specifications necessary to perform JK Drop Weight testing, an impact test performed

to provide a, b and ta parameters. These parameters measure the impact energy for breakage

(Napier-Munn et al., 1996) and supply the data necessary to simulate crusher and SAG mill

circuits in JK SimMet®. This inclusion allows for the estimation of an SABC circuit for

comminution energy comparison. The revised procedure also retains testing of crusher and

HPGR product for Bond work index determination, since these values are necessary for

predicting the energy requirements for grinding.

As for HPGR testing, estimation of energy benefits would only require four specific pressing

force tests. The results from these tests would provide valuable performance data for sizing an

HPGR, beneficial to a project even if adoption of an HPGR / stirred mill circuit is not pursued

further. With these results, an optimal specific pressing force would be selected and the product

from that test would be re-run through the HPGR and standard test data recorded to determine

energy requirements for second-stage HPGR crushing. The remaining product from the other

three tests would be re-run through the HPGR, solely for production of feed for stirred mill

testing. An option is available to use this material to assess the effect of moisture on second-

stage HPGR crushing. All second-stage products would then be screened at 710µm.

Undersize from screening would be blended and two 100kg splits would be run through an M20

stirred mill, producing two signature plots for energy calculations.

Upon completion of testing, results would be used to perform JK SimMet® simulation for three

comparative circuits: a cone crusher / cone crusher / ball mill circuit, a cone crusher / HPGR /

ball mill circuit and an SABC circuit. Each circuit would be simulated for a feed top size of

105

125mm and a product p80 of 75µm. A summary of energy requirements for these circuits would

then be compared to a circuit similar in design to the layout developed in Section 5.3.

The resulting study would provide a preliminary estimation of the potential energy benefits

associated with exploring an HPGR / stirred mill option. Overall, this procedure could be looked

at as a scoping level study for whether further HPGR / stirred mill testing is desirable. Summary

of a testing flowsheet is shown in Figure 5-7.

106

Figure 5-7 Scoping Level Testing Procedure for HPGR / Stirred Mill Evaluation

1,200kg@ 100% -150mm

Screen32mm

Jaw Crusher

100kgJK Drop Weight

Testing

Drum #1

Blend Sample

Drum #2

Drum #3

Drum #4

Particle Size Distribution

Bulk Density

HPGR Pressing Force #1

HPGR Pressing Force #4OPTIMAL



Stage Crushing10kg

Bond Work Index Testing

Crusher Product

@ -3.35mm

TestedProduct

TestedProduct

TestedProduct

TestedProduct

Centre Particle Size Distribution

Edge Particle Size Distribution

Flake Density

Bond Work Index Testing

HPGR Product

10kgScreened @ 3.35mm

HPGR HPGR HPGR HPGR

710µmScreen

M20 Stirred Mill Testing

Signature Plot Generation

TestedProduct

Centre Particle Size Distribution

Edge Particle Size Distribution

Flake Density

107

6 Conclusions and Recommendations

The research summarized in this thesis focused on the first steps necessary in the development

of a novel HPGR / stirred mill circuit. A review of the literature found that, separately, HPGRs

and stirred mills offered higher energy efficiency over existing tumbling mills and incorporating

them into a novel circuit layout could lead to eliminating the need for a tumbling mill. This could

result in an improvement in comminution efficiency for hard-rock mining applications. The

literature review found that very little operating examples of such a circuit were available, thus

pilot-scale testing was used to assess its technical feasibility. Section 2.3 concluded that a

layout would need to consist of two stages of HPGR comminution to achieve an acceptable feed

size for coarse stirred milling.

The first step necessary in designing the circuit was to develop appropriate design criteria and

feasible transfer sizes between each stage of comminution. The following were the critical

design considerations for the circuit:

• Specific Pressing Force

•

– The specific pressing force for Mesaba ore was determined to

be 4N/mm2. This force provided a balance between size reduction and specific energy

consumption with no significant drop in specific throughput.

HPGR Moisture Content

•

– The moisture content for testing was 2.5%. This value

provided a balance between dust suppression and comminution performance.

HPGR Circuit Configuration

•

– Two stages of HPGR comminution were selected to

produce adequate feed size for stirred milling. Two configurations were evaluated,

operating the first-stage HPGR in open circuit or in closed circuit with a 4mm screen.

Open circuit configuration was determined to be the better option for the circuit.

Stirred Mill Geometry

•

– The first grinding disc for the stirred mill was removed to produce

a coarser grind per pass through the mill. This resulted in accurate energy figures for

coarse grind sizes.

Grinding Media Size – The grinding media top size used for stirred mill testing was 6mm

ceramic. The selection of this top size provided effective grinding for coarse stirred

milling.

108

• Stirred Mill Flow Rate

•

– The stirred mill flow rate of 20.4L/min was selected to help

ensure a short residence time in the mill. This helped improved prediction of grinding

energy for coarser products.

Mill Feed Solids Content

•

– The solids content for the feed was tested at 57%. This

produced lower specific energy consumption per pass through the mill, helping achieve

an accurate signature plot with evenly spaced data points.

Mill Speed

•

– The mill speed was set at 1,200 RPM and achieved an actual recorded

value of 1,169 RPM. This corresponded to a disc tip speed of 11m/s.

HPGR / Stirred Mill Transfer Size

Energy data was recorded during testing, and specific energy consumption was summarized for

the proposed HPGR / stirred mill circuit. To evaluate whether this circuit provided the energy

benefits suggested in the literature, a combination of testing and JK SimMet® simulation was

conducted on cone crusher / ball mill and HPGR / ball mill circuits. The results provided data

used to calculate specific energy requirements for comparison.

– Three separate transfer sizes were evaluated,

355µm, 710µm and 1.2mm. Results showed that 710µm was the most suitable option

for the operating conditions tested.

Results showed that an HPGR / stirred mill circuit required a specific energy consumption of

14.85kWh/t to reduce a feed f80 of 21mm down to a product p80 of 75µm. Upon comparison to

the conventional circuits, this corresponded to 9.2 (HPGR / ball mill circuit) and 16.7% (cone

crusher / ball mill circuit) reduction in the required energy for comminution. With results from the

study, a proposed circuit flowsheet was generated and a revised testing procedure

recommended, establishing a basis for moving forward. The revised testing procedure reduced

the testing sample requirements to 1,200kg. Results from this procedure would provide data to

compare the specific energy requirements for the following circuits:

• Cone crusher / cone crusher / ball mill circuit

• Cone crusher / HPGR / ball mill circuit

• SAG / ball mill circuit

• HPGR / stirred mill circuit

Overall, this research provided a valuable addition to the body of knowledge for potential

comminution circuit designs of the future. Chapter 1 stated that if the mining industry was to

109

remain sustainable for future generations, a shift was needed to implement more energy

efficient practices. Establishing the technical feasibility of an HPGR / stirred mill circuit is a first

step in the ascension towards adopting a more effective use of energy for comminution. With

the establishment that combining HPGRs and stirred mills together can reduce specific energy

consumption for comminution, the foundation has been created to allow future engineers to

adopt this concept to a larger sample set.

Future work is recommended for further evaluation of an HPGR / stirred mill circuit:

1. Application of the refined testing procedure with different ore types. The results

presented here represent only one ore deposit and it is recommended that more ore

types be tested to increase the database on the energy benefits of HPGR / stirred mill

comminution. Evaluation of different base metal and precious metal deposits will help

determine whether further energy benefits are achievable. The inclusion of a SAG mill /

ball mill component will help determine whether the energy benefits achieved in this

study are similar when an SABC circuit is included.

2. Examination of the total specific energy requirements for the circuit. Future studies

should include analysis of materials handling equipment to determine total specific

energy requirements for each circuit. This would help finalize whether the energy

benefits determined in this study can translate to energy improvements in a complete

comminution circuit.

3. Evaluation of different grinding media sizes for coarse stirred milling. Further work

should be done in establishing media selection criteria for coarse stirred milling. A

media size ranging between 3-8mm should be tested on a 710µm feed top size and

signature plot comparisons should be performed to determine which size produced the

highest grinding efficiency.

4. Further evaluation of feed top size testing for coarse stirred milling. The coarsest

particle size tested in the stirred mill was 1.2mm. The operating conditions used for this

study resulted in a power draw higher than the 18.6kW motor could handle. Using a

larger motor size may help successfully grind 1.2mm material. Further work should be

done on assessing whether sizes between 0.71 and 1.2mm can be effectively processed

through a stirred mill.

110

5. Testing the variation in operating conditions on coarse stirred milling. Other operating

conditions, including increasing mill speed, reducing feed solids content and changing

mill geometry, should be examined to confirm their affects on grinding efficiency.

6. Further design work on the second-stage HPGR screening circuit. The screening circuit

prior to the stirred mill should be explored further, including testing the technical

feasibility of screening a fine product while still achieving high comminution performance

for a continuously-operating HPGR.

7. Establishing a continuously operating pilot-scale circuit. This would help improve the

identification of potential bottlenecks in the circuit and provide valuable data for scaling

up to a full-scale circuit.

111

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press - stirred ball milling; feinzerkleinerung von kalkstein in gutbettwalzenmuehlen und

ruehrwerkskugelmuehlen. Aufbereitungs-Technik, 39(6), 267-278.

Wang, Y., Forssberg, E., & Lin, P. (2006). Hybrid comminution with high-pressure roller and

stirred bead milling. Particle and Particle Systems Characterization, 22(5), 336-344.

121

Whittles, D. N., Kingman, S., Lowndes, I., & Jackson, K. (2006). Laboratory and numerical

investigation into the characteristics of rock fragmentation. Minerals Engineering, 19(14),

1418-1429.

Yue, J., & Klein, B. (2006). Effects of bead size on ultrafine grinding in a stirred bead mill. In S.

Komar Kawatra (Ed.), Advances in comminution (pp. 87-98). Littleton, CO, USA: Society for

Mining, Metallurgy and Exploration Inc.

122

A. Appendix A – JK SimMet® Data

154

B. Appendix B – Bond Work Index Data

155

Figure B-1 Bond Work Index Data Crusher Product (150µm)

BOND MILL GRINDABILITY TEST REPORT

Test: 150µm Test #2 Date: September 21, 2009Sample: Mesaba HPGR Product (3n/mm^2) Project: MASc Research Thesis

TEST CONDITIONS

Cycle Oversize Wt. Product Wt. Feed Undersize Net Product Product per Rev. Required Rev. Predicted Recirculating (grams) (grams) (grams) (grams) (grams/rev.) (rev.) Rev Load

1 838 563 354 209 2.09 1002 1,024 377 142 235 1.89 124 124 2713 1,022 379 95 284 1.76 161 161 2704 999 402 96 306 1.77 173 173 2495 994 407 102 305 1.80 169 169 2446 1,007 394 103 291 1.77 165 165 2557 994 408 100 308 1.81 170 170 244

SIZE ANALYSIS TEST RESULTS

Sieve Size % PassingTyler mesh µm Feed Product

6 3360 99.5 Material Charge Wt.-700 mL(g) = 1,40110 1700 77.5 Test Screen (µm) = 15014 1180 68.1 Undersize in Feed (%)= 25.320 841 59.1 Circulating Load (%) = 24428 595 51.6 Gbp (ave.) = 1.7935 417 44.5 Product P80 (µm) = 125.648 297 37.0 Feed F80 (µm) = 1,85465 210 31.4 W (kWh/ton) = 13.2

100 150 25.3 94.2 W (kWh/tonne) = 14.5150 106 20.0 68.4200 75 15.1 39.1270 53 9.2 21.2325 45 3.9 4.9400 38 3.1 0.4

156

Figure B-2 Bond Work Index Data Crusher Product (150µm) (continued)

SIZE ANALYSISSize Feed Product

Tyler µm Weight Individual Passing Weight Individual Passingmesh (g) (%) (%) (g) (%) (%)

6 3360 5.8 0.4 99.6 0.0 100.010 1700 405.9 27.9 71.7 0.0 100.014 1180 146.4 10.1 61.6 0.0 100.020 841 128.5 8.8 52.8 0.0 100.028 595 101.3 7.0 45.8 0.0 100.035 417 93.0 6.4 39.4 0.0 100.048 297 95.2 6.5 32.9 0.0 100.065 210 74.9 5.2 27.7 0.0 100.0100 150 76.7 5.3 22.4 13.2 1.8 98.2150 106 68.1 4.7 17.8 194.5 25.9 72.3200 75 67.1 4.6 13.1 213.0 28.4 43.9270 53 70.7 4.9 8.3 131.9 17.6 26.3325 45 69.3 4.8 3.5 135.9 18.1 8.2400 38 7.3 0.5 3.0 35.6 4.7 3.5-400 43.8 3.0 0.0 25.9 3.5 0.0Total 1454.0 100.0 750.0 100.0

F 80 P80S1 (µm) 3360 S1 (µm) 150P1 (%) 99.6 P1 (%) 98.2S2 (µm) 1700 S2 (µm) 106P2 (%) 71.7 P2 (%) 72.3

m 0.48 m 0.88b 0.68 b 0.17

F80 2133 P80 118.9

157

Figure B-3 Bond Work Index Data 3N/mm2 HPGR Product (150µm)



TEST CONDITIONS


1 838 563 354 209 2.09 1002 1,024 377 142 235 1.89 124 124 2713 1,022 379 95 284 1.76 161 161 2704 999 402 96 306 1.77 173 173 2495 994 407 102 305 1.80 169 169 2446 1,007 394 103 291 1.77 165 165 2557 994 408 100 308 1.81 170 170 244




100 150 25.3 94.2 W (kWh/tonne) = 14.5150 106 20.0 68.4200 75 15.1 39.1270 53 9.2 21.2325 45 3.9 4.9400 38 3.1 0.4

158

Figure B-4 Bond Work Index Data 3N/mm2 HPGR Product (150µm) (continued)



6 3360 6.6 0.5 99.5 0.0 100.010 1700 309.3 22.1 77.5 0.0 100.014 1180 131.4 9.4 68.1 0.0 100.020 841 126.4 9.0 59.1 0.0 100.028 595 104.6 7.5 51.6 0.0 100.035 417 99.1 7.1 44.5 0.0 100.048 297 105.8 7.6 37.0 0.0 100.065 210 78.2 5.6 31.4 0.0 100.0100 150 85.5 6.1 25.3 45.8 5.8 94.2150 106 74.3 5.3 20.0 205.8 25.9 68.4200 75 68.5 4.9 15.1 233.5 29.3 39.1270 53 82.9 5.9 9.2 141.9 17.8 21.2325 45 73.5 5.2 3.9 130.1 16.3 4.9400 38 10.8 0.8 3.1 35.6 4.5 0.4-400 44.1 3.1 0.0 3.3 0.4 0.0Total 1401.0 100.0 796.0 100.0


m 0.37 m 0.92b 1.62 b -0.07

F80 1854 P80 125.6

159




TEST CONDITIONS


1 862 644 456 187 1.87 1002 1,078 427 195 232 1.86 125 125 2523 1,095 410 129 281 1.73 162 162 2674 1,068 437 124 313 1.78 176 176 2445 1,075 430 132 297 1.77 168 168 2506 1,071 434 130 304 1.80 169 169 2477 1,079 426 132 294 1.77 166 166 253




100 150 30.3 96.9 W (kWh/tonne) = 14.5150 106 23.9 68.1200 75 15.9 37.5270 53 9.1 15.6325 45 2.3 4.2400 38 0.3 0.9

160




6 3360 8.8 0.6 99.4 0.0 100.010 1700 327.7 21.8 77.6 0.0 100.014 1180 134.1 8.9 68.7 0.0 100.020 841 118.9 7.9 60.8 0.0 100.028 595 96.1 6.4 54.4 0.0 100.035 417 85.7 5.7 48.8 0.0 100.048 297 98.1 6.5 42.2 0.0 100.065 210 78.1 5.2 37.0 0.0 100.0100 150 101.9 6.8 30.3 24.2 3.1 96.9150 106 96.2 6.4 23.9 224.2 28.8 68.1200 75 120.2 8.0 15.9 238.3 30.6 37.5270 53 102.4 6.8 9.1 170.4 21.9 15.6325 45 102.0 6.8 2.3 88.7 11.4 4.2400 38 30.2 2.0 0.3 25.7 3.3 0.9-400 4.6 0.3 0.0 6.9 0.9 0.0Total 1505.0 100.0 778.4 100.0


m 0.36 m 1.02b 1.65 b -0.52

F80 1849 P80 124.2

161




TEST CONDITIONS


1 727 662 403 259 2.59 1002 992 398 192 206 2.60 79 79 2493 1,050 339 115 224 2.07 108 108 3104 1,010 380 98 282 1.96 144 144 2665 1,003 386 110 276 1.88 147 147 2606 979 411 112 299 1.97 152 152 2387 987 403 119 284 2.01 141 141 2458 997 393 117 276 1.99 139 139 254




100 150 29.0 97.6 W (kWh/tonne) = 13.3150 106 23.1 72.9200 75 16.9 42.3270 53 9.7 21.3325 45 2.4 6.6400 38 0.4 1.9

162




6 3360 2.9 0.2 99.8 0.0 100.010 1700 232.4 16.7 83.1 0.0 100.014 1180 118.5 8.5 74.5 0.0 100.020 841 121.4 8.7 65.8 0.0 100.028 595 107.9 7.8 58.0 0.0 100.035 417 105.0 7.6 50.5 0.0 100.048 297 112.1 8.1 42.4 0.0 100.065 210 84.4 6.1 36.3 0.0 100.0100 150 101.4 7.3 29.0 18.2 2.4 97.6150 106 82.4 5.9 23.1 184.4 24.6 72.9200 75 85.7 6.2 16.9 229.5 30.7 42.3270 53 101.3 7.3 9.7 157.2 21.0 21.3325 45 100.7 7.2 2.4 109.6 14.6 6.6400 38 28.1 2.0 0.4 35.7 4.8 1.9-400 5.3 0.4 0.0 13.9 1.9 0.0Total 1389.5 100.0 748.5 100.0


m 0.30 m 0.84b 2.19 b 0.37

F80 1497 P80 118.4

163

Figure B-9 Bond Work Index Data Crusher Product (106µm)


Test: 106µm Test #1 Date: January, 2011Sample: Mesaba Crusher Product Project: MASc Research Thesis

TEST CONDITIONS


1 1,108 315 247 69 0.69 1002 904 519 55 465 0.91 512 512 1743 968 455 90 365 1.05 349 349 2134 982 441 79 362 1.15 313 313 2235 1,055 368 76 291 1.02 286 286 2876 981 442 64 378 1.13 336 336 2227 999 424 77 347 1.18 293 293 2368 1,026 397 73 324 1.15 282 282 2589 1,026 397 69 328 1.12 294 294 259



6 3360 99.4 Material Charge Wt.-700 mL(g) = 1,422.910 1700 69.0 Test Screen (µm) = 10614 1180 59.9 Undersize in Feed (%)= 17.320 841 50.0 Circulating Load (%) = 25928 595 42.4 Gbp (ave.) = 1.1535 417 36.6 Product P80 (µm) = 8048 297 30.7 Feed F80 (µm) = 2,24165 210 26.4 W (kWh/ton) = 15.0

100 150 21.7 W (kWh/tonne) = 16.5150 106 17.3 100.0200 75 76.3270 53 39.8325 45 30.3400 38 22.7

164

Figure B-10 Bond Work Index Data Crusher Product (106µm) (continued)



6 3360 2.7 0.6 99.4 0.0 100.010 1700 136.9 30.4 69.0 0.0 100.014 1180 40.9 9.1 59.9 0.0 100.020 841 44.7 9.9 50.0 0.0 100.028 595 34.0 7.6 42.4 0.0 100.035 417 25.9 5.8 36.6 0.0 100.048 297 26.9 6.0 30.7 0.0 100.065 210 19.4 4.3 26.4 0.0 100.0100 150 20.8 4.6 21.7 0.0 100.0150 106 20.1 4.5 17.3 0.0 0.0 100.0200 75 93.6 23.7 76.3270 53 143.8 36.5 39.8325 45 37.3 9.5 30.3400 38 30.0 7.6 22.7Pan 77.7 17.3 0.0 89.6 22.7 0.0Total 450.0 100.0 394.3 100.0


m 0.54 m 0.78b 0.25 b 0.96

F80 2241 P80 80

165



Test: 106µm Test #2 Date: January, 2011Sample: Mesaba HPGR Product (3n/mm^2) Project: MASc Research Thesis

TEST CONDITIONS


1 1,029 396 359 37 0.37 1002 519 906 100 806 0.96 841 841 573 985 439 228 211 1.13 186 186 2244 1,031 394 111 283 1.08 262 262 2625 996 428 99 329 1.16 284 284 2336 981 443 108 335 1.30 258 258 2217 1,025 400 112 288 1.27 227 227 2568 1,026 399 101 298 1.23 242 242 257



6 3360 99.3 Material Charge Wt.-700 mL(g) = 1,424.410 1700 79.0 Test Screen (µm) = 10614 1180 71.4 Undersize in Feed (%)= 25.220 841 61.9 Circulating Load (%) = 25728 595 54.8 Gbp (ave.) = 1.2735 417 48.6 Product P80 (µm) = 81.548 297 41.9 Feed F80 (µm) = 1,76565 210 36.9 W (kWh/ton) = 14.4

100 150 31.2 W (kWh/tonne) = 15.9150 106 25.2 100.0200 75 74.6270 53 40.2325 45 30.4400 38 21.5

166




6 3360 3.5 0.7 99.3 0.0 100.010 1700 106.3 20.3 79.0 0.0 100.014 1180 39.5 7.6 71.4 0.0 100.020 841 49.9 9.6 61.9 0.0 100.028 595 37.2 7.1 54.8 0.0 100.035 417 32.0 6.1 48.6 0.0 100.048 297 35.0 6.7 41.9 0.0 100.065 210 26.5 5.1 36.9 0.0 100.0100 150 29.6 5.7 31.2 0.0 100.0150 106 31.4 6.0 25.2 0.0 0.0 100.0200 75 101.4 25.4 74.6270 53 137.7 34.5 40.2325 45 38.9 9.7 30.4400 38 35.8 9.0 21.5Pan 131.6 25.2 0.0 85.8 21.5 0.0Total 522.5 100.0 399.6 100.0


m 0.34 m 0.85b 1.87 b 0.66

F80 1765 P80 81 .5

167




TEST CONDITIONS


1 976 505 413 92 0.92 1002 993 489 141 348 1.13 307 307 2033 1,043 439 136 303 1.20 253 253 2384 1,073 408 122 286 1.14 252 252 2635 1,053 428 114 314 1.15 272 272 2466 1,020 461 119 342 1.30 263 263 2217 1,052 430 129 301 1.33 227 227 2458 1,082 399 120 279 1.22 228 228 2719 1,072 409 111 298 1.17 255 255 262

10 1,023 458 114 344 1.30 265 265 22311 1,064 417 128 289 1.27 227 227 255




100 150 36.0 W (kWh/tonne) = 15.8150 106 27.9 100.0200 75 75.0270 53 35.5325 45 19.9400 38 11.2

168




6 3360 1.9 0.4 99.6 0.0 100.010 1700 106.1 20.6 79.0 0.0 100.014 1180 33.8 6.6 72.4 0.0 100.020 841 41.1 8.0 64.4 0.0 100.028 595 33.3 6.5 58.0 0.0 100.035 417 28.4 5.5 52.4 0.0 100.048 297 30.3 5.9 46.5 0.0 100.065 210 25.5 5.0 41.6 0.0 100.0100 150 28.9 5.6 36.0 0.0 100.0150 106 41.4 8.0 27.9 0.0 0.0 100.0200 75 102.7 25.0 75.0270 53 162.6 39.6 35.5325 45 63.9 15.5 19.9400 38 35.9 8.7 11.2Pan 143.6 27.9 0.0 45.9 11.2 0.0Total 514.3 100.0 411.0 100.0


m 0.34 m 0.83b 1.84 b 0.73

F80 1764 P80 81 .1

169




TEST CONDITIONS


1 928 504 398 106 1.06 1002 971 461 140 321 1.26 255 255 2103 1,049 383 128 254 1.14 223 223 2744 1,021 411 106 305 1.15 265 265 2485 1,012 420 114 305 1.19 256 256 2416 1,005 427 117 310 1.26 246 246 2357 1,017 415 119 296 1.29 230 230 2458 1,034 398 115 283 1.24 228 228 260




100 150 34.2 W (kWh/tonne) = 15.7150 106 27.8 100.0200 75 76.9270 53 39.4325 45 27.2400 38 19.8

170




6 3360 3.2 0.4 99.6 0.0 100.010 1700 146.1 19.4 80.2 0.0 100.014 1180 50.6 6.7 73.5 0.0 100.020 841 64.3 8.5 64.9 0.0 100.028 595 51.6 6.9 58.1 0.0 100.035 417 45.7 6.1 52.0 0.0 100.048 297 48.6 6.5 45.5 0.0 100.065 210 39.8 5.3 40.3 0.0 100.0100 150 45.2 6.0 34.2 0.0 100.0150 106 48.6 6.5 27.8 0.0 0.0 100.0200 75 90.9 23.1 76.9270 53 147.5 37.5 39.4325 45 48.1 12.2 27.2400 38 29.1 7.4 19.8Pan 209.3 27.8 0.0 77.7 19.8 0.0Total 753.0 100.0 393.3 100.0


m 0.24 m 0.76b 2.61 b 1.06

F80 1682 P80 79.0

171

C. Appendix C – HPGR Data

172

Table C-1 HPGR Pilot-Scale Test Key

Test ID Test Description PH

ASE

ON

E T1A01 5N/mm Specific Pressing Force Testing T1A02 4N/mm Specific Pressing Force Testing / Closed Circuit Testing Cycle 1 T1A03 3N/mm Specific Pressing Force Testing T1A04 2N/mm Specific Pressing Force Testing T1A05 0.6m/s Roller Speed Testing T1A06 0.9m/s Roller Speed Testing T1A07 1% Moisture Content Testing T1A08 5% Moisture Content Testing T1A09 Closed Circuit Testing with 4mm Screen Cycle 2 T1A10 Closed Circuit Testing with 4mm Screen Cycle 3 T1A11 Closed Circuit Testing with 4mm Screen Cycle 4

PHA

SE T

WO

T2A01 Generating Product for Second Stage Closed Circuit Testing T2B01 Second Stage Dry Closed Circuit Testing Cycle 1 T2B02 Second Stage Dry Closed Circuit Testing Cycle 2 T2B03 Second Stage Dry Closed Circuit Testing Cycle 3 T2B04 Second Stage Dry Closed Circuit Testing Cycle 4 T2B05 Second Stage Wet Closed Circuit Testing Cycle 5

173

Table C-2 HPGR Operating Data Phase One

Roller Diameter (D) [m] 0.750Roller Width (W) [m] 0.220

Symbol Unitν [m/s] 0.75 0.75 0.75 0.75 0.60 0.90n [rpm] 19.10 19.10 19.10 19.10 15.30 22.90

Static Gap X0 [mm] 9.0 9.0 9.0 9.0 9.0 9.0Hydraulic Pressure P [bar] 103 82 62 41 82 82

Pressing Force F [kN] 828.4 659.5 498.6 329.7 659.5 659.5Specific Pressing Force FSP [kN/m2] 5 4 3 2 4 4

Test Time t [s] 19.00 18.41 20.99 18.41 15.00 15.80Average Actual Speed: ωAV [m/s] 0.75 0.75 0.75 0.77 0.61 0.90

Standard Deviation σω 0.05 0.06 0.05 0.25 0.17 0.06Actual Roller gap (average) XgAV [mm] 21.46 22.53 24.10 25.39 23.68 23.18

Standard Deviation σX 0.68 0.59 1.13 0.93 0.74 0.68Actual Hydraulic Pressure (average) PAV [bar] 100.7 80.2 61.1 42.5 81.2 80.6

Standard Deviation 0.87 1.79 0.96 1.66 0.81 1.21Actual Pressing Force (average) FAV [kN] 810 645 491 342 653 648

Actual Specific Pressure (average) FSPAV [kN/m2] 4.92 3.92 2.98 2.08 3.97 3.94Idle Power Draw Pi [kW] 8.12 8.04 9.18 9.36 6.88 12.20

Power Draw P [kW] 77.94 64.77 54.52 43.58 53.44 79.75Total Specific Energy Consumption ESP [kWh/t] 2.17 1.77 1.42 1.09 1.75 1.73Net Specific Energy Consumption ESP net [kWh/t] 1.94 1.55 1.18 0.85 1.53 1.46

Press throughput W [t/h] 35.98 36.55 38.36 40.11 30.48 46.17Specific Throughput Constant m dot [ts/hm3] 292.20 295.69 310.97 315.37 304.22 311.50

Press Constant

Data Description Test Number:

Proc

ess

Set

Poin

ts

Speed

Proc

ess

Dat

a

T1A01 T1A02 T1A06T1A05T1A04T1A03

174

Table C-3 HPGR Operating Data Phase One (continued)

Symbol Unitν [m/s] 0.75 0.75 0.75 0.75 0.75 0.75n [rpm] 19.10 19.10 19.10 19.10 19.10 19.10

Static Gap X0 [mm] 9.0 9.0 9.0 9.0 9.0 9.0Hydraulic Pressure P [bar] 82 82 82 82 82 82

Pressing Force F [kN] 659.5 659.5 659.5 659.5 659.5 659.5Specific Pressing Force FSP [kN/m2] 4 4 4 4 4 4

Test Time t [s] 21.39 20.60 19.21 20.60 18.80 18.59Average Actual Speed: ωAV [m/s] 0.75 0.76 0.75 0.76 0.76 0.77

Standard Deviation σω 0.08 0.18 0.06 0.17 0.14 0.21Actual Roller gap (average) XgAV [mm] 23.43 20.59 22.07 22.47 22.13 17.56

Standard Deviation σX 1.56 0.90 0.50 0.60 0.57 0.32Actual Hydraulic Pressure (average) PAV [bar] 80.4 80.8 80.5 79.9 80.1 80.8

Standard Deviation 0.61 0.84 0.81 1.46 0.87 0.43Actual Pressing Force (average) FAV [kN] 647 650 647 642 644 650

Actual Specific Pressure (average) FSPAV [kN/m2] 3.93 3.95 3.93 3.90 3.92 3.95Idle Power Draw Pi [kW] 9.11 8.56 9.16 10.25 10.00 9.63

Power Draw P [kW] 64.04 73.87 62.17 64.70 64.94 45.49Total Specific Energy Consumption ESP [kWh/t] 1.70 2.13 1.71 1.75 1.72 1.53Net Specific Energy Consumption ESP net [kWh/t] 1.46 1.89 1.46 1.47 1.45 1.20

Press throughput W [t/h] 37.63 34.60 36.31 37.01 37.83 29.77Specific Throughput Constant m dot [ts/hm3] 304.77 276.58 293.66 296.99 304.00 235.81

T1A11 T1B01Pr

oces

s Se

t Po

ints

Speed

Data Description Pr

oces

s D

ata

Test Number: T1A07 T1A08 T1A09 T1A10

175


Symbol UnitAverage Flake Density ρF [t/m3] 2.59 2.57 2.55 2.51 2.55 2.53

Standard Deviation 0.06 0.04 0.06 0.02 0.04 0.05Flake Thickness Average XF [mm] 25.2 24.7 25.6 26.6 24.7 25.2

Standard Deviation 1.6 2.5 2.6 3.8 1.3 2.2Feed Moisture [%] 2.5% 2.5% 2.5% 2.5% 2.5% 2.5%

Feed Bulk Density [t/m3] 2.16 2.16 2.16 2.16 2.16 2.16Particle Size DistributionFeed: 80% Passing Size F80 [mm] 21.33 21.33 21.33 21.33 21.33 21.33Feed: 50% Passing Size F80 [mm] 13.96 13.96 13.96 13.96 13.96 13.96

Centre: 80% Passing Size P80 [mm] 6.85 7.10 7.43 8.88 7.46 6.91Centre: 50% Passing Size P50 [mm] 1.60 1.74 1.67 2.29 1.84 1.70Edge: 80% Passing Size P80 [mm] 8.83 10.93 12.12 12.87 9.00 11.27Edge: 50% Passing Size P50 [mm] 2.18 3.30 4.59 5.30 2.35 3.70

Scale-up: 80% Passing Size P80 [mm] 7.15 7.67 8.13 9.48 7.69 7.56Scale-up: 50% Passing Size P50 [mm] 1.69 1.97 2.11 2.74 1.92 2.00

Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 2.98 2.78 2.62 2.25 2.77 2.82Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 8.28 7.07 6.62 5.09 7.28 6.98

Mass BalanceTotal Feed Material MF [kg] 383 366 375 370 372 321Total Centre Product MC [kg] 151.3 149.8 166.4 144.5 146.1 129.2

Centre Product % of Centre & Edge Material MCE% [%] 79.7% 80.1% 74.4% 70.5% 58.9% 69.5%Total Edge Product ME [kg] 38.6 37.1 57.3 60.6 101.9 56.8

Edge Product % of Centre & Edge Material MEF% [%] 20.3% 19.9% 25.6% 29.5% 41.1% 30.5%Edge Product % of Centre Product MEC% [%] 25.5% 24.8% 34.4% 41.9% 69.7% 44.0%

Total Waste Product MW [kg] 179.2 168.1 138.5 157.3 113.2 124.3Waste Product % of Total Feed MWF% [%] 46.8% 45.9% 36.9% 42.5% 30.4% 38.7%

Total Recovered Product MP [kg] 369 355 362 362 361 310Mass Reconciliation (+ "gain; - "loss") MPF% [%] 3.6% 3.0% 3.4% 2.1% 2.9% 3.3%

T1A05Description Test Number: T1A01 T1A02 T1A03 T1A04 T1A06

176


Symbol UnitAverage Flake Density ρF [t/m3] 2.54 2.48 2.56 2.56 2.56 2.30

Standard Deviation 0.12 0.08 0.04 0.06 0.03 0.03Flake Thickness Average XF [mm] 27.5 22.9 24.4 23.8 24.2 18.0

Standard Deviation 1.2 3.3 2.0 2.8 2.6 2.0Feed Moisture [%] 1.0% 5.0% 2.5% 2.5% 2.5% 2.2%

Feed Bulk Density [t/m3] 2.16 2.16 2.22 2.20 2.16Particle Size DistributionFeed: 80% Passing Size F80 [mm] 21.33 21.33 19.02 19.73 21.77 1.86Feed: 50% Passing Size F80 [mm] 13.96 13.96 11.23 10.58 13.38 0.49

Centre: 80% Passing Size P80 [mm] 6.12 7.37 5.68 6.43 6.04 1.09Centre: 50% Passing Size P50 [mm] 1.14 1.68 1.42 1.72 1.72 0.22Edge: 80% Passing Size P80 [mm] 11.46 9.90 9.67 10.15 9.74 1.28Edge: 50% Passing Size P50 [mm] 4.20 3.31 3.51 3.71 3.38 0.25

Scale-up: 80% Passing Size P80 [mm] 6.92 7.75 6.28 6.99 6.60 1.12Scale-up: 50% Passing Size P50 [mm] 1.60 1.92 1.73 2.02 1.97 0.22

Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 3.08 2.75 3.03 2.82 3.30 1.66Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 8.73 7.25 6.48 5.24 6.80 2.21

Mass BalanceTotal Feed Material MF [kg] 373 393 364 362 362 249Total Centre Product MC [kg] 151.6 145 147.2 155.2 141 101.3

Centre Product % of Centre & Edge Material MCE% [%] 67.8% 73.2% 76.0% 73.3% 71.4% 65.9%Total Edge Product ME [kg] 72.0 53.0 46.5 56.6 56.6 52.4

Edge Product % of Centre & Edge Material MEF% [%] 32.2% 26.8% 24.0% 26.7% 28.6% 34.1%Edge Product % of Centre Product MEC% [%] 47.5% 36.6% 31.6% 36.5% 40.1% 51.7%

Total Waste Product MW [kg] 138.5 179.6 159.9 139.3 154.5 89.0Waste Product % of Total Feed MWF% [%] 37.1% 45.7% 43.9% 38.5% 42.7% 35.7%

Total Recovered Product MP [kg] 362 378 354 351 352 243Mass Reconciliation (+ "gain; - "loss") MPF% [%] 2.9% 3.9% 2.9% 3.0% 2.7% 2.5%

T1A10 T1A11 T1B01T1A07 T1A08 T1A09Test Number:Description

177

Table C-6 HPGR Operating Data Phase Two

Roller Diameter (D) [m] 0.750Roller Width (W) [m] 0.220

Symbol Unitν [m/s] 0.75 0.75 0.75 0.75 0.75n [rpm] 19.10 19.10 19.10 19.10 19.10

Static Gap X0 [mm] 9.0 9.0 9.0 9.0 9.0Hydraulic Pressure P [bar] 82 82 82 82 82

Pressing Force F [kN] 660.0 660.0 660.0 660.0 660.0Specific Pressing Force FSP [kN/m2] 4 4 4 4 4

Test Time t [s] 21.00 21.21 20.40 21.80 19.20Average Actual Speed: ωAV [m/s] 0.76 0.74 0.77 0.75 0.76

Standard Deviation σω 0.14 0.08 0.24 0.06 0.16Actual Roller gap (average) XgAV [mm] 22.07 22.25 21.93 22.12 22.40

Standard Deviation σX 2.14 1.45 1.13 0.54 0.41Actual Hydraulic Pressure (average) PAV [bar] 80.8 80.7 81.0 81.1 80.6

Standard Deviation 0.83 0.78 0.66 1.15 1.09Actual Pressing Force (average) FAV [kN] 650 649 651 652 648

Actual Specific Pressure (average) FSPAV [kN/m2] 3.95 3.95 3.96 3.96 3.94Idle Power Draw Pi [kW] 8.86 8.11 8.33 8.31 8.65

Power Draw P [kW] 65.48 67.40 67.51 67.50 56.61Total Specific Energy Consumption ESP [kWh/t] 1.78 1.74 1.79 1.74 1.46Net Specific Energy Consumption ESP net [kWh/t] 1.54 1.53 1.57 1.53 1.23

Press throughput W [t/h] 36.84 38.75 37.78 38.74 38.89Specific Throughput Constant m dot [ts/hm3] 295.62 319.80 298.21 314.14 310.86

Proc

ess

Set

Poin

ts

Speed

Proc

ess

Dat

a

Test Number: T2A01 T2A02 T2A03 T2A04 T2B01

Press Constant

Data Description

178

Table C-7 HPGR Operating Data Phase Two (continued)

Symbol Unitν [m/s] 0.75 0.75 0.75 0.75n [rpm] 19.10 19.10 19.10 19.10

Static Gap X0 [mm] 9.0 9.0 9.0 9.0Hydraulic Pressure P [bar] 82 82 82 82

Pressing Force F [kN] 660.0 660.0 660.0 660.0Specific Pressing Force FSP [kN/m2] 4 4 4 4

Test Time t [s] 20.20 20.00 20.40 24.20Average Actual Speed: ωAV [m/s] 0.73 0.73 0.74 0.77

Standard Deviation σω 0.07 0.09 0.14 0.21Actual Roller gap (average) XgAV [mm] 23.48 22.93 22.62 17.26

Standard Deviation σX 0.30 0.37 0.19 1.21Actual Hydraulic Pressure (average) PAV [bar] 81.0 80.7 80.8 80.5

Standard Deviation 0.85 1.27 0.91 2.38Actual Pressing Force (average) FAV [kN] 651 649 650 648

Actual Specific Pressure (average) FSPAV [kN/m2] 3.96 3.95 3.95 3.94Idle Power Draw Pi [kW] 8.21 8.09 7.55 10.56

Power Draw P [kW] 52.89 52.31 52.16 67.60Total Specific Energy Consumption ESP [kWh/t] 1.41 1.41 1.42 2.32Net Specific Energy Consumption ESP net [kWh/t] 1.19 1.20 1.21 1.96

Press throughput W [t/h] 37.52 36.98 36.79 29.17Specific Throughput Constant m dot [ts/hm3] 311.62 309.82 300.26 231.68

Proc

ess

Set

Poin

ts

Speed

Proc

ess

Dat

aT2B02 T2B03 T2B04 T2B05Data Description Test Number:

179


Symbol UnitAverage Flake Density ρF [t/m3] 2.41

Standard Deviation 0.03Flake Thickness Average XF [mm] 24.6

Standard Deviation 2.1Feed Moisture [%] 2.5% 2.5% 2.5% 2.5% 2.4%

Feed Bulk Density [t/m3] 2.16 2.16 2.16 2.16 2.16Particle Size DistributionFeed: 80% Passing Size F80 [mm] 21.54 21.54 21.54 21.54 8.10Feed: 50% Passing Size F80 [mm] 13.70 13.70 13.70 13.70 2.27

Centre: 80% Passing Size P80 [mm] 2.54Centre: 50% Passing Size P50 [mm] 0.42Edge: 80% Passing Size P80 [mm] 4.97Edge: 50% Passing Size P50 [mm] 0.89

Scale-up: 80% Passing Size P80 [mm] 2.79Scale-up: 50% Passing Size P50 [mm] 0.46

Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 2.90Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 4.91

Mass BalanceTotal Feed Material MF [kg] 311 364 366 371 338Total Centre Product MC [kg] 214.9 228.3 214.1 234.6 132.6

Centre Product % of Centre & Edge Material MCE% [%] 63.9%Total Edge Product ME [kg] 74.8

Edge Product % of Centre & Edge Material MEF% [%] 36.1%Edge Product % of Centre Product MEC% [%] 56.4%

Total Waste Product MW [kg] 94.6 131.3 151.6 135.5 128.5Waste Product % of Total Feed MWF% [%] 30.4% 36.1% 41.4% 36.5% 38.0%

Total Recovered Product MP [kg] 310 360 366 370 336Mass Reconciliation (+ "gain; - "loss") MPF% [%] 0.5% 1.2% 0.1% 0.2% 0.6%

T2A04 T2B01Description Test Number: T2A01 T2A02 T2A03

180


Symbol UnitAverage Flake Density ρF [t/m3] 2.56 2.58 2.56 2.57

Standard Deviation 0.05 0.04 0.03 0.02Flake Thickness Average XF [mm] 21.8 23.5 22.7 19.3

Standard Deviation 3.4 2.2 1.9 1.4Feed Moisture [%] 2.4% 2.4% 2.4% 5.8%

Feed Bulk Density [t/m3] 2.25 2.25 2.25 2.25Particle Size DistributionFeed: 80% Passing Size F80 [mm] 6.61 6.16 5.69 6.41Feed: 50% Passing Size F80 [mm] 2.27 2.06 1.79 1.95

Centre: 80% Passing Size P80 [mm] 3.04 2.71 2.92 2.79Centre: 50% Passing Size P50 [mm] 0.64 0.59 0.65 0.49Edge: 80% Passing Size P80 [mm] 5.36 5.26 4.43 3.39Edge: 50% Passing Size P50 [mm] 1.47 1.43 1.19 0.71

Scale-up: 80% Passing Size P80 [mm] 3.42 3.09 3.15 2.88Scale-up: 50% Passing Size P50 [mm] 0.74 0.71 0.73 0.52

Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 1.93 1.99 1.81 2.23Reduction Ratio F80/P80 (Centre (0.85) Edge(0.15)) 3.09 2.88 2.45 3.72

Mass BalanceTotal Feed Material MF [kg] 338 337 338 334Total Centre Product MC [kg] 173.5 157.5 159 141.2

Centre Product % of Centre & Edge Material MCE% [%] 82.4% 76.6% 76.3% 72.0%Total Edge Product ME [kg] 37.0 48.0 49.5 54.9

Edge Product % of Centre & Edge Material MEF% [%] 17.6% 23.4% 23.7% 28.0%Edge Product % of Centre Product MEC% [%] 21.3% 30.5% 31.1% 38.9%

Total Waste Product MW [kg] 121.0 125.5 121.5 131.6Waste Product % of Total Feed MWF% [%] 35.8% 37.2% 35.9% 39.4%

Total Recovered Product MP [kg] 332 331 330 328Mass Reconciliation (+ "gain; - "loss") MPF% [%] 1.9% 1.8% 2.4% 1.9%

T2B03 T2B04 T2B05T2B02Description Test Number:

181

Table C-10 HPGR Feed Size Distributions Phase One

32 0 0.00% 100.00% 0 0.00% 100.00%25 880.1 7.21% 92.79% 667 5.97% 94.03%19 2551.9 28.12% 71.88% 1575.2 20.05% 79.95%16 1851.1 43.28% 56.72% 1228.1 31.04% 68.96%

12.5 1407.6 54.81% 45.19% 1523.9 44.67% 55.33%8 1520.7 67.27% 32.73% 2118.7 63.62% 36.38%

5.6 724.8 73.21% 26.79% 1092.2 73.39% 26.61%4 531.9 77.57% 22.43% 832.4 80.83% 19.17%

2.8 454.8 81.29% 18.71% 462.5 84.97% 15.03%2 309.1 83.82% 16.18% 202.1 86.78% 13.22%

1.4 280 86.12% 13.88% 151.5 88.13% 11.87%1 188.4 87.66% 12.34% 106.7 89.09% 10.91%

0.71 137 88.78% 11.22% 120.9 90.17% 9.83%0.5 158.5 90.08% 9.92% 111.3 91.16% 8.84%

0.355 148 91.30% 8.70% 106.4 92.11% 7.89%0.25 156.1 92.57% 7.43% 109.6 93.10% 6.90%0.18 144.4 93.76% 6.24% 123.7 94.20% 5.80%0.125 147.4 94.96% 5.04% 153.6 95.58% 4.42%0.09 179.8 96.44% 3.56% 140.2 96.83% 3.17%0.063 174 97.86% 2.14% 147 98.14% 1.86%0.045 103.4 98.71% 1.29% 107.4 99.10% 0.90%Pan 157.4 100.00% 0.00% 100.1 100.00% 0.00%Total 12206.4 11180.5

Percent Passing (%)

Size (mm)

Weight (g)

Percent Accumulated (%)

Percent Passing (%)

Weight (g)


Mesaba Feed Sample T1A09 Feed

182

Table C-11 HPGR Feed Size Distributions Phase One (continued)

32 0 0.00% 100.00% 49.9 0.46% 99.54%25 609.5 4.99% 95.01% 1049.2 10.21% 89.79%19 2089.1 22.09% 77.91% 1959.6 28.42% 71.58%16 1220.9 32.09% 67.91% 1433 41.73% 58.27%

12.5 1286.7 42.62% 57.38% 1190.9 52.79% 47.21%8 2110.2 59.90% 40.10% 1594 67.60% 32.40%

5.6 1297.1 70.52% 29.48% 850.3 75.50% 24.50%4 1090.6 79.45% 20.55% 729.1 82.28% 17.72%

2.8 649.3 84.77% 15.23% 455 86.50% 13.50%2 256.2 86.86% 13.14% 196.5 88.33% 11.67%

1.4 161.8 88.19% 11.81% 140.8 89.64% 10.36%1 113.5 89.12% 10.88% 88.7 90.46% 9.54%

0.71 106.4 89.99% 10.01% 95.3 91.35% 8.65%0.5 122.2 90.99% 9.01% 98.8 92.26% 7.74%

0.355 110.9 91.90% 8.10% 88 93.08% 6.92%0.25 122.5 92.90% 7.10% 91.3 93.93% 6.07%0.18 134.7 94.00% 6.00% 92.7 94.79% 5.21%0.125 142.2 95.17% 4.83% 101.4 95.73% 4.27%0.09 160.3 96.48% 3.52% 95.8 96.62% 3.38%0.063 177.8 97.93% 2.07% 167 98.17% 1.83%0.045 116.9 98.89% 1.11% 94.1 99.05% 0.95%Pan 135.4 100.00% 0.00% 102.4 100.00% 0.00%Total 12214.2 10763.8

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)


T1A10 Feed T1A11 FeedPercent Passing

(%)

183

Table C-12 T1A01 Product Size Distributions

32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 52.5 0.54% 99.46% 62.9 0.72% 99.28% 0.57% 99.43%16 87.1 1.44% 98.56% 186 2.86% 97.14% 1.65% 98.35%

12.5 363.5 5.19% 94.81% 598.8 9.74% 90.26% 5.87% 94.13%8 1072.7 16.26% 83.74% 1093.7 22.31% 77.69% 17.17% 82.83%

5.6 757.8 24.08% 75.92% 789.3 31.38% 68.62% 25.17% 74.83%4 752.5 31.84% 68.16% 650.1 38.85% 61.15% 32.89% 67.11%

2.8 717.9 39.25% 60.75% 585.1 45.58% 54.42% 40.20% 59.80%2 601 45.45% 54.55% 498 51.30% 48.70% 46.33% 53.67%

1.4 653.7 52.20% 47.80% 474.1 56.75% 43.25% 52.88% 47.12%1 411.6 56.44% 43.56% 292.3 60.11% 39.89% 56.99% 43.01%

0.71 342.2 59.97% 40.03% 346.4 64.09% 35.91% 60.59% 39.41%0.5 439.2 64.51% 35.49% 333.4 67.92% 32.08% 65.02% 34.98%

0.355 388 68.51% 31.49% 310.6 71.49% 28.51% 68.96% 31.04%0.25 406.7 72.71% 27.29% 321.8 75.18% 24.82% 73.08% 26.92%0.18 387.5 76.70% 23.30% 311.5 78.76% 21.24% 77.01% 22.99%0.125 412.6 80.96% 19.04% 340 82.67% 17.33% 81.22% 18.78%0.09 395.4 85.04% 14.96% 379.8 87.04% 12.96% 85.34% 14.66%0.063 412 89.29% 10.71% 346.7 91.02% 8.98% 89.55% 10.45%0.045 219.5 91.56% 8.44% 201.5 93.34% 6.66% 91.82% 8.18%Pan 818.2 100.00% 0.00% 579.9 100.00% 0.00% 100.00% 0.00%Total 11417.7 8701.9

T1A01 85:15 Scaled ProductPercent

Accumulated (%)Percent Passing

(%)

T1A01 Centre Product T1A01 Edge ProductSize (mm)

Weight (g)


Percent Passing (%)

Weight (g)


Percent Passing (%)

184

Figure C-1 T1A01 Particle Size Distributions

185


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 23.1 0.23% 99.77% 0 0.00% 100.00% 0.20% 99.80%19 30 0.54% 99.46% 105.4 1.20% 98.80% 0.64% 99.36%16 156.2 2.11% 97.89% 449.7 6.33% 93.67% 2.75% 97.25%

12.5 385.9 6.01% 93.99% 755.4 14.94% 85.06% 7.35% 92.65%8 1096.8 17.09% 82.91% 1273.5 29.47% 70.53% 18.95% 81.05%

5.6 766 24.83% 75.17% 783.2 38.40% 61.60% 26.86% 73.14%4 758.5 32.49% 67.51% 670 46.04% 53.96% 34.52% 65.48%

2.8 785.3 40.42% 59.58% 599.2 52.87% 47.13% 42.29% 57.71%2 656.4 47.05% 52.95% 441.3 57.90% 42.10% 48.68% 51.32%

1.4 680.8 53.93% 46.07% 460.3 63.15% 36.85% 55.31% 44.69%1 510.9 59.09% 40.91% 286.1 66.42% 33.58% 60.19% 39.81%

0.71 375.1 62.88% 37.12% 241.9 69.17% 30.83% 63.83% 36.17%0.5 415.4 67.08% 32.92% 300.3 72.60% 27.40% 67.91% 32.09%

0.355 376.4 70.88% 29.12% 276.7 75.75% 24.25% 71.61% 28.39%0.25 386.6 74.78% 25.22% 280.5 78.95% 21.05% 75.41% 24.59%0.18 368.7 78.51% 21.49% 351.5 82.96% 17.04% 79.18% 20.82%0.125 406.1 82.61% 17.39% 270 86.04% 13.96% 83.13% 16.87%0.09 371.9 86.37% 13.63% 331.6 89.82% 10.18% 86.89% 13.11%0.063 401.4 90.42% 9.58% 301.4 93.26% 6.74% 90.85% 9.15%0.045 219.3 92.64% 7.36% 169.7 95.19% 4.81% 93.02% 6.98%Pan 728.8 100.00% 0.00% 421.4 100.00% 0.00% 100.00% 0.00%Total 9899.6 8769.1

Percent Passing (%)


Percent Passing (%)

T1A02 Centre Product T1A02 Edge Product T1A02 85:15 Scaled ProductSize (mm)

Weight (g)


Percent Passing (%)

Weight (g)


186


187


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 29.6 0.26% 99.74% 22.7 0.18% 99.82% 0.25% 99.75%19 117.7 1.31% 98.69% 376.2 3.08% 96.92% 1.58% 98.42%16 229.1 3.36% 96.64% 688.4 8.40% 91.60% 4.11% 95.89%

12.5 510.3 7.91% 92.09% 1308.4 18.51% 81.49% 9.50% 90.50%8 1159.5 18.24% 81.76% 2280.8 36.13% 63.87% 20.93% 79.07%

5.6 834.4 25.68% 74.32% 1162.5 45.11% 54.89% 28.60% 71.40%4 779.1 32.63% 67.37% 1006.2 52.89% 47.11% 35.67% 64.33%

2.8 832.1 40.05% 59.95% 860.5 59.53% 40.47% 42.97% 57.03%2 694.3 46.24% 53.76% 635.8 64.45% 35.55% 48.97% 51.03%

1.4 772.6 53.13% 46.87% 627.5 69.30% 30.70% 55.55% 44.45%1 619.7 58.65% 41.35% 470.1 72.93% 27.07% 60.79% 39.21%

0.71 472.6 62.87% 37.13% 336.9 75.53% 24.47% 64.77% 35.23%0.5 505.4 67.37% 32.63% 349 78.23% 21.77% 69.00% 31.00%

0.355 459.9 71.47% 28.53% 331.9 80.79% 19.21% 72.87% 27.13%0.25 483.3 75.78% 24.22% 337.2 83.40% 16.60% 76.92% 23.08%0.18 510 80.33% 19.67% 327.8 85.93% 14.07% 81.17% 18.83%0.125 470 84.52% 15.48% 363.1 88.73% 11.27% 85.15% 14.85%0.09 454.8 88.57% 11.43% 334.1 91.32% 8.68% 88.98% 11.02%0.063 374.6 91.91% 8.09% 360.5 94.10% 5.90% 92.24% 7.76%0.045 229.1 93.96% 6.04% 193 95.59% 4.41% 94.20% 5.80%Pan 677.9 100.00% 0.00% 570.6 100.00% 0.00% 100.00% 0.00%Total 11216 12943.2

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)


T1A03 Centre Product T1A03 Edge Product T1A03 85:15 Scaled Product

188


189


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 173.7 1.93% 98.07% 307.6 2.70% 97.30% 2.05% 97.95%16 275.2 5.00% 95.00% 868.7 10.34% 89.66% 5.80% 94.20%

12.5 519.1 10.78% 89.22% 1228.2 21.13% 78.87% 12.33% 87.67%8 1029.2 22.24% 77.76% 1968.7 38.43% 61.57% 24.67% 75.33%

5.6 790 31.04% 68.96% 1150.8 48.54% 51.46% 33.66% 66.34%4 678.9 38.60% 61.40% 882.4 56.29% 43.71% 41.25% 58.75%

2.8 670.9 46.07% 53.93% 758.5 62.95% 37.05% 48.60% 51.40%2 551.2 52.20% 47.80% 582.6 68.07% 31.93% 54.58% 45.42%

1.4 580 58.66% 41.34% 488.9 72.37% 27.63% 60.72% 39.28%1 423.1 63.37% 36.63% 334.2 75.31% 24.69% 65.16% 34.84%

0.71 382.9 67.64% 32.36% 291.3 77.87% 22.13% 69.17% 30.83%0.5 364.8 71.70% 28.30% 311.8 80.61% 19.39% 73.03% 26.97%

0.355 323.1 75.30% 24.70% 281.1 83.08% 16.92% 76.46% 23.54%0.25 331.7 78.99% 21.01% 285.5 85.58% 14.42% 79.98% 20.02%0.18 305.4 82.39% 17.61% 281.2 88.05% 11.95% 83.24% 16.76%0.125 302 85.75% 14.25% 275.8 90.48% 9.52% 86.46% 13.54%0.09 335.5 89.49% 10.51% 260.9 92.77% 7.23% 89.98% 10.02%0.063 323.2 93.09% 6.91% 310.4 95.50% 4.50% 93.45% 6.55%0.045 182.1 95.12% 4.88% 164.1 96.94% 3.06% 95.39% 4.61%Pan 438.6 100.00% 0.00% 348.3 100.00% 0.00% 100.00% 0.00%Total 8980.6 11381

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



190


191


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 31.6 0.37% 99.63% 141.5 1.23% 98.77% 0.50% 99.50%16 224.3 2.97% 97.03% 274.3 3.62% 96.38% 3.07% 96.93%

12.5 376.6 7.34% 92.66% 687.7 9.60% 90.40% 7.68% 92.32%8 926.9 18.09% 81.91% 1538.1 22.97% 77.03% 18.82% 81.18%

5.6 733.1 26.59% 73.41% 993.5 31.62% 68.38% 27.34% 72.66%4 675 34.42% 65.58% 874.7 39.22% 60.78% 35.14% 64.86%

2.8 689 42.41% 57.59% 859.8 46.70% 53.30% 43.06% 56.94%2 506.4 48.29% 51.71% 677.4 52.59% 47.41% 48.93% 51.07%

1.4 554.3 54.72% 45.28% 722.7 58.88% 41.12% 55.34% 44.66%1 378.7 59.11% 40.89% 498 63.21% 36.79% 59.72% 40.28%

0.71 282.6 62.39% 37.61% 391 66.61% 33.39% 63.02% 36.98%0.5 378.9 66.78% 33.22% 470.5 70.70% 29.30% 67.37% 32.63%

0.355 339.6 70.72% 29.28% 414.3 74.30% 25.70% 71.26% 28.74%0.25 357.2 74.86% 25.14% 436 78.10% 21.90% 75.35% 24.65%0.18 364.3 79.09% 20.91% 418.9 81.74% 18.26% 79.49% 20.51%0.125 321.2 82.81% 17.19% 385.6 85.09% 14.91% 83.16% 16.84%0.09 342.8 86.79% 13.21% 469.6 89.18% 10.82% 87.15% 12.85%0.063 364.4 91.02% 8.98% 368.6 92.38% 7.62% 91.22% 8.78%0.045 191 93.23% 6.77% 211.3 94.22% 5.78% 93.38% 6.62%Pan 583.5 100.00% 0.00% 664.5 100.00% 0.00% 100.00% 0.00%Total 8621.4 11498

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



192


193


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 0 0.00% 100.00% 165.1 1.78% 98.22% 0.27% 99.73%16 107.3 1.34% 98.66% 469.5 6.85% 93.15% 2.16% 97.84%

12.5 178.2 3.56% 96.44% 816.4 15.67% 84.33% 5.38% 94.62%8 691.5 12.18% 87.82% 1469.3 31.54% 68.46% 15.08% 84.92%

5.6 628.8 20.01% 79.99% 815.7 40.35% 59.65% 23.06% 76.94%4 695.7 28.69% 71.31% 726.5 48.20% 51.80% 31.61% 68.39%

2.8 661 36.92% 63.08% 668.1 55.41% 44.59% 39.70% 60.30%2 516.5 43.36% 56.64% 491 60.71% 39.29% 45.96% 54.04%

1.4 608.2 50.94% 49.06% 515.8 66.29% 33.71% 53.24% 46.76%1 473.8 56.85% 43.15% 388.9 70.49% 29.51% 58.89% 41.11%

0.71 380 61.58% 38.42% 311.4 73.85% 26.15% 63.42% 36.58%0.5 378.4 66.30% 33.70% 284.8 76.92% 23.08% 67.89% 32.11%

0.355 299.5 70.03% 29.97% 262.9 79.76% 20.24% 71.49% 28.51%0.25 336.3 74.22% 25.78% 270.1 82.68% 17.32% 75.49% 24.51%0.18 326.2 78.29% 21.71% 259.4 85.48% 14.52% 79.37% 20.63%0.125 306.7 82.11% 17.89% 305 88.78% 11.22% 83.11% 16.89%0.09 335 86.29% 13.71% 266.3 91.65% 8.35% 87.09% 12.91%0.063 330.6 90.41% 9.59% 288.4 94.77% 5.23% 91.06% 8.94%0.045 226.8 93.23% 6.77% 162.6 96.52% 3.48% 93.73% 6.27%Pan 542.8 100.00% 0.00% 321.9 100.00% 0.00% 100.00% 0.00%Total 8023.3 9259.1

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



194


195


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 31.4 0.34% 99.66% 165.1 1.24% 98.76% 0.48% 99.52%16 169 2.20% 97.80% 774.9 7.04% 92.96% 2.93% 97.07%

12.5 279.5 5.27% 94.73% 1195.4 15.99% 84.01% 6.88% 93.12%8 855.6 14.66% 85.34% 2305.7 33.26% 66.74% 17.45% 82.55%

5.6 621.4 21.48% 78.52% 1286.6 42.90% 57.10% 24.69% 75.31%4 600.9 28.07% 71.93% 1086 51.03% 48.97% 31.52% 68.48%

2.8 588.5 34.53% 65.47% 965 58.26% 41.74% 38.09% 61.91%2 540.8 40.47% 59.53% 735.1 63.76% 36.24% 43.96% 56.04%

1.4 591.4 46.96% 53.04% 710.9 69.09% 30.91% 50.28% 49.72%1 426.8 51.64% 48.36% 512.7 72.93% 27.07% 54.84% 45.16%

0.71 366.6 55.67% 44.33% 426 76.12% 23.88% 58.73% 41.27%0.5 485.2 60.99% 39.01% 388 79.02% 20.98% 63.70% 36.30%

0.355 445.1 65.88% 34.12% 356.9 81.70% 18.30% 68.25% 31.75%0.25 440.1 70.71% 29.29% 356.7 84.37% 15.63% 72.76% 27.24%0.18 415.7 75.27% 24.73% 323.7 86.79% 13.21% 77.00% 23.00%0.125 426.3 79.95% 20.05% 335.1 89.30% 10.70% 81.35% 18.65%0.09 386.1 84.19% 15.81% 294.4 91.51% 8.49% 85.28% 14.72%0.063 392 88.49% 11.51% 352.8 94.15% 5.85% 89.34% 10.66%0.045 240.5 91.13% 8.87% 184.5 95.53% 4.47% 91.79% 8.21%Pan 808.4 100.00% 0.00% 596.6 100.00% 0.00% 100.00% 0.00%Total 9111.3 13352.1

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



196


197


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 29.6 0.23% 99.77% 0.03% 99.97%19 14.5 0.16% 99.84% 66.9 0.74% 99.26% 0.25% 99.75%16 111.8 1.38% 98.62% 273.7 2.84% 97.16% 1.60% 98.40%

12.5 386 5.59% 94.41% 1024.7 10.71% 89.29% 6.36% 93.64%8 1110.5 17.72% 82.28% 2094.7 26.78% 73.22% 19.08% 80.92%

5.6 800.5 26.46% 73.54% 1300.8 36.77% 63.23% 28.01% 71.99%4 717.8 34.30% 65.70% 1147.2 45.57% 54.43% 35.99% 64.01%

2.8 640.4 41.30% 58.70% 1000.7 53.26% 46.74% 43.09% 56.91%2 510.6 46.87% 53.13% 732 58.87% 41.13% 48.67% 51.33%

1.4 540.5 52.78% 47.22% 699.8 64.25% 35.75% 54.50% 45.50%1 445.4 57.64% 42.36% 516 68.21% 31.79% 59.22% 40.78%

0.71 368.3 61.66% 38.34% 403.3 71.30% 28.70% 63.11% 36.89%0.5 347.6 65.46% 34.54% 397.3 74.35% 25.65% 66.79% 33.21%

0.355 328.4 69.04% 30.96% 379.6 77.26% 22.74% 70.28% 29.72%0.25 364.1 73.02% 26.98% 386.7 80.23% 19.77% 74.10% 25.90%0.18 383.3 77.21% 22.79% 405 83.34% 16.66% 78.13% 21.87%0.125 404 81.62% 18.38% 416.7 86.54% 13.46% 82.36% 17.64%0.09 404.7 86.04% 13.96% 399.8 89.61% 10.39% 86.57% 13.43%0.063 354.9 89.91% 10.09% 377.5 92.51% 7.49% 90.30% 9.70%0.045 197.5 92.07% 7.93% 208.9 94.11% 5.89% 92.38% 7.62%Pan 726.1 100.00% 0.00% 767.4 100.00% 0.00% 100.00% 0.00%Total 9156.9 13028.3

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



198


199


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 17.3 0.13% 99.87% 81.3 0.71% 99.29% 0.22% 99.78%16 62.9 0.62% 99.38% 278.8 3.15% 96.85% 1.00% 99.00%

12.5 359.7 3.39% 96.61% 818.1 10.31% 89.69% 4.43% 95.57%8 1170.1 12.40% 87.60% 1763.3 25.73% 74.27% 14.40% 85.60%

5.6 1021.7 20.27% 79.73% 1191.8 36.16% 63.84% 22.65% 77.35%4 1068.7 28.50% 71.50% 1153.8 46.25% 53.75% 31.16% 68.84%

2.8 1046.1 36.56% 63.44% 1041.4 55.37% 44.63% 39.38% 60.62%2 893.6 43.44% 56.56% 711.9 61.59% 38.41% 46.16% 53.84%

1.4 886.6 50.27% 49.73% 653 67.31% 32.69% 52.82% 47.18%1 731.2 55.90% 44.10% 492.4 71.61% 28.39% 58.26% 41.74%

0.71 607.9 60.58% 39.42% 393.3 75.05% 24.95% 62.75% 37.25%0.5 561.8 64.91% 35.09% 350.7 78.12% 21.88% 66.89% 33.11%

0.355 492.8 68.71% 31.29% 263.8 80.43% 19.57% 70.46% 29.54%0.25 555.8 72.99% 27.01% 286.7 82.94% 17.06% 74.48% 25.52%0.18 557.6 77.28% 22.72% 308.5 85.64% 14.36% 78.53% 21.47%0.125 587.3 81.80% 18.20% 273.6 88.03% 11.97% 82.74% 17.26%0.09 573.9 86.22% 13.78% 341.4 91.02% 8.98% 86.94% 13.06%0.063 489.3 89.99% 10.01% 337.2 93.97% 6.03% 90.59% 9.41%0.045 255.3 91.96% 8.04% 177.3 95.52% 4.48% 92.49% 7.51%Pan 1044 100.00% 0.00% 512.2 100.00% 0.00% 100.00% 0.00%Total 12983.6 11430.5

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



200


201


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 37.6 0.37% 99.63% 131.7 0.96% 99.04% 0.46% 99.54%16 179.8 2.12% 97.88% 464.5 4.33% 95.67% 2.45% 97.55%

12.5 276.5 4.82% 95.18% 931.7 11.09% 88.91% 5.76% 94.24%8 999.5 14.57% 85.43% 2346.7 28.13% 71.87% 16.60% 83.40%

5.6 848.6 22.85% 77.15% 1383.2 38.17% 61.83% 25.15% 74.85%4 932.2 31.95% 68.05% 1343.8 47.92% 52.08% 34.34% 65.66%

2.8 882.6 40.56% 59.44% 1169.8 56.41% 43.59% 42.94% 57.06%2 661.3 47.01% 52.99% 783.1 62.10% 37.90% 49.27% 50.73%

1.4 659 53.44% 46.56% 741.6 67.48% 32.52% 55.55% 44.45%1 507.4 58.39% 41.61% 517.9 71.24% 28.76% 60.32% 39.68%

0.71 412 62.41% 37.59% 396.7 74.12% 25.88% 64.17% 35.83%0.5 387.6 66.19% 33.81% 376.8 76.86% 23.14% 67.79% 32.21%

0.355 359.9 69.70% 30.30% 377.3 79.60% 20.40% 71.19% 28.81%0.25 387.1 73.48% 26.52% 372.5 82.30% 17.70% 74.80% 25.20%0.18 403.4 77.42% 22.58% 372 85.00% 15.00% 78.55% 21.45%0.125 505.8 82.35% 17.65% 436.4 88.17% 11.83% 83.22% 16.78%0.09 372.3 85.98% 14.02% 378 90.91% 9.09% 86.72% 13.28%0.063 497.5 90.84% 9.16% 364 93.56% 6.44% 91.25% 8.75%0.045 172.8 92.52% 7.48% 203.5 95.03% 4.97% 92.90% 7.10%Pan 766.3 100.00% 0.00% 684.3 100.00% 0.00% 100.00% 0.00%Total 10249.2 13775.5

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



202


203


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 18.2 0.20% 99.80% 134.5 0.97% 99.03% 0.31% 99.69%16 122.2 1.53% 98.47% 429.8 4.08% 95.92% 1.91% 98.09%

12.5 236.5 4.11% 95.89% 935.8 10.84% 89.16% 5.12% 94.88%8 806.3 12.92% 87.08% 2065.5 25.77% 74.23% 14.85% 85.15%

5.6 793 21.58% 78.42% 1406.3 35.93% 64.07% 23.73% 76.27%4 852.3 30.88% 69.12% 1330 45.55% 54.45% 33.08% 66.92%

2.8 809.2 39.72% 60.28% 1199.4 54.21% 45.79% 41.89% 58.11%2 677.1 47.11% 52.89% 804.1 60.03% 39.97% 49.05% 50.95%

1.4 576.5 53.40% 46.60% 809.4 65.88% 34.12% 55.27% 44.73%1 464.8 58.48% 41.52% 569.5 69.99% 30.01% 60.20% 39.80%

0.71 405.5 62.90% 37.10% 486.2 73.50% 26.50% 64.49% 35.51%0.5 349.4 66.72% 33.28% 421.1 76.55% 23.45% 68.19% 31.81%

0.355 319 70.20% 29.80% 383.7 79.32% 20.68% 71.57% 28.43%0.25 338.8 73.90% 26.10% 399.5 82.21% 17.79% 75.15% 24.85%0.18 339.9 77.61% 22.39% 413.5 85.20% 14.80% 78.75% 21.25%0.125 397.6 81.95% 18.05% 447.3 88.43% 11.57% 82.92% 17.08%0.09 371.7 86.01% 13.99% 355.5 91.00% 9.00% 86.76% 13.24%0.063 363.4 89.98% 10.02% 384.8 93.78% 6.22% 90.55% 9.45%0.045 202.1 92.18% 7.82% 226.6 95.42% 4.58% 92.67% 7.33%Pan 716 100.00% 0.00% 634 100.00% 0.00% 100.00% 0.00%Total 9159.5 13836.5

T1A11 Centre Product T1A11 Edge Product T1A11 85:15 Scaled ProductSize (mm)

Weight (g)


Percent Passing (%)

Weight (g)


Percent Passing (%)


Percent Passing (%)

204


205

Table C-23 HPGR Feed Size Distributions Phase Two

32 614.028 0.78% 99.22% 0 0.00% 100.00%25 5883.32 8.25% 91.75% 0 0.00% 100.00%19 16040.4 28.61% 71.39% 43 0.35% 99.65%16 10617.9 42.09% 57.91% 285.9 2.68% 97.32%

12.5 9488.95 54.13% 45.87% 525.3 6.95% 93.05%8 9306.87 65.94% 34.06% 1640.9 20.30% 79.70%

5.6 4775.57 72.01% 27.99% 1210.7 30.16% 69.84%4 3572.83 76.54% 23.46% 1026.4 38.51% 61.49%

2.8 2847.8 80.16% 19.84% 918.8 45.99% 54.01%2 2061.37 82.77% 17.23% 743.9 52.04% 47.96%

1.4 1622.67 84.83% 15.17% 684.4 57.61% 42.39%1 1434.94 86.65% 13.35% 510.9 61.77% 38.23%

0.71 1013.4 87.94% 12.06% 455.7 65.47% 34.53%0.5 1137.21 89.38% 10.62% 432.7 69.00% 31.00%

0.355 1021.79 90.68% 9.32% 460.9 72.75% 27.25%0.25 1038.84 92.00% 8.00% 459.6 76.49% 23.51%0.18 955.794 93.21% 6.79% 673.7 81.97% 18.03%0.125 1048.5 94.54% 5.46% 394.7 85.18% 14.82%0.09 742.453 95.49% 4.51% 373.2 88.22% 11.78%0.063 819.648 96.53% 3.47% 385.8 91.36% 8.64%0.045 539.052 97.21% 2.79% 221.9 93.16% 6.84%Pan 2196.93 100.00% 0.00% 840.2 100.00% 0.00%Total 78780.3 12288.6

T2A01 Feed T2B01 FeedSize (mm)

Weight (g)


Percent Passing (%)

Weight (g)


Percent Passing (%)

206

Table C-24 HPGR Feed Size Distributions Phase Two (continued)

32 0 0.00% 100.00% 0 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00%19 20.7 0.18% 99.82% 20.7 0.20% 99.80%16 79.5 0.88% 99.12% 107.9 1.26% 98.74%

12.5 331.1 3.77% 96.23% 308.5 4.29% 95.71%8 1184.5 14.12% 85.88% 855.4 12.70% 87.30%

5.6 1161.7 24.28% 75.72% 968.6 22.21% 77.79%4 1102.3 33.91% 66.09% 1004.2 32.08% 67.92%

2.8 1167.7 44.12% 55.88% 1003.1 41.93% 58.07%2 1007.9 52.92% 47.08% 892.4 50.70% 49.30%

1.4 1030.1 61.93% 38.07% 868 59.23% 40.77%1 719.8 68.22% 31.78% 687.8 65.98% 34.02%

0.71 755.8 74.82% 25.18% 826.6 74.10% 25.90%0.5 450.1 78.76% 21.24% 479.3 78.81% 21.19%

0.355 258.2 81.01% 18.99% 236.3 81.13% 18.87%0.25 269.7 83.37% 16.63% 237.6 83.47% 16.53%0.18 254.8 85.60% 14.40% 220.8 85.64% 14.36%0.125 286.6 88.10% 11.90% 251.7 88.11% 11.89%0.09 308.1 90.80% 9.20% 241.6 90.48% 9.52%0.063 250.3 92.98% 7.02% 247.1 92.91% 7.09%0.045 154 94.33% 5.67% 124.7 94.14% 5.86%Pan 648.7 100.00% 0.00% 596.9 100.00% 0.00%Total 11441.6 10179.2


Percent Passing (%)

T2B02 Feed T2B03 FeedSize (mm)

Weight (g)


Percent Passing (%)

Weight (g)

207

Table C-25 HPGR Feed Size Distributions Phase Two (continued)

32 0 0.00% 100.00% 0 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00%19 16.5 0.15% 99.85% 53.6 0.48% 99.52%16 90.6 0.96% 99.04% 124.2 1.60% 98.40%

12.5 284.9 3.52% 96.48% 377.1 4.99% 95.01%8 951.8 12.07% 87.93% 989.5 13.88% 86.12%

5.6 915.2 20.30% 79.70% 1027.5 23.11% 76.89%4 909.4 28.47% 71.53% 996.4 32.07% 67.93%

2.8 1059.2 37.98% 62.02% 1034.6 41.36% 58.64%2 957.8 46.59% 53.41% 887.8 49.34% 50.66%

1.4 1060.6 56.12% 43.88% 912.3 57.54% 42.46%1 841.8 63.68% 36.32% 635.3 63.25% 36.75%

0.71 972.4 72.42% 27.58% 724.2 69.75% 30.25%0.5 574.7 77.58% 22.42% 520.6 74.43% 25.57%

0.355 267.1 79.98% 20.02% 353.8 77.61% 22.39%0.25 281.9 82.52% 17.48% 337 80.64% 19.36%0.18 289 85.11% 14.89% 342.4 83.72% 16.28%0.125 266.3 87.51% 12.49% 351 86.87% 13.13%0.09 271.4 89.94% 10.06% 329.5 89.83% 10.17%0.063 261.8 92.30% 7.70% 305.2 92.57% 7.43%0.045 173.9 93.86% 6.14% 127.1 93.72% 6.28%Pan 683.5 100.00% 0.00% 699.2 100.00% 0.00%Total 11129.8 11128.3

Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)


T2B04 Feed T2B05 Feed

208


32 0 0.00% 100.00% 0 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00%19 52.5 0.54% 99.46% 43 0.35% 99.65%16 87.1 1.44% 98.56% 285.9 2.68% 97.32%

12.5 363.5 5.19% 94.81% 525.3 6.95% 93.05%8 1072.7 16.26% 83.74% 1640.9 20.30% 79.70%

5.6 757.8 24.08% 75.92% 1210.7 30.16% 69.84%4 752.5 31.84% 68.16% 1026.4 38.51% 61.49%

2.8 717.9 39.25% 60.75% 918.8 45.99% 54.01%2 601 45.45% 54.55% 743.9 52.04% 47.96%

1.4 653.7 52.20% 47.80% 684.4 57.61% 42.39%1 411.6 56.44% 43.56% 510.9 61.77% 38.23%

0.71 342.2 59.97% 40.03% 455.7 65.47% 34.53%0.5 439.2 64.51% 35.49% 432.7 69.00% 31.00%

0.355 388 68.51% 31.49% 460.9 72.75% 27.25%0.25 406.7 72.71% 27.29% 459.6 76.49% 23.51%0.18 387.5 76.70% 23.30% 673.7 81.97% 18.03%0.125 412.6 80.96% 19.04% 394.7 85.18% 14.82%0.09 395.4 85.04% 14.96% 373.2 88.22% 11.78%0.063 412 89.29% 10.71% 385.8 91.36% 8.64%0.045 219.5 91.56% 8.44% 221.9 93.16% 6.84%Pan 818.2 100.00% 0.00% 840.2 100.00% 0.00%Total 11417.7 12288.6

Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)


T2A01 Centre + Edge Product T2A01 Centre + Edge + Waste

209


210

Table C-27 T2B01 Product Size Distributions

32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%16 10.1 0.10% 99.90% 15.3 0.22% 99.78% 0.11% 99.89%

12.5 16.2 0.25% 99.75% 147.9 2.35% 97.65% 0.56% 99.44%8 346.2 3.55% 96.45% 605.6 11.05% 88.95% 4.67% 95.33%

5.6 386.6 7.23% 92.77% 435.8 17.32% 82.68% 8.74% 91.26%4 535.4 12.32% 87.68% 477.4 24.18% 75.82% 14.10% 85.90%

2.8 605.4 18.08% 81.92% 447.5 30.61% 69.39% 19.96% 80.04%2 632.4 24.10% 75.90% 402.6 36.40% 63.60% 25.95% 74.05%

1.4 585.3 29.68% 70.32% 492.4 43.48% 56.52% 31.75% 68.25%1 548.3 34.90% 65.10% 311.3 47.95% 52.05% 36.85% 63.15%

0.71 725.2 41.80% 58.20% 364.5 53.19% 46.81% 43.51% 56.49%0.5 559.1 47.12% 52.88% 321.1 57.81% 42.19% 48.72% 51.28%

0.355 531.5 52.18% 47.82% 275.1 61.76% 38.24% 53.62% 46.38%0.25 626.9 58.15% 41.85% 335.4 66.58% 33.42% 59.41% 40.59%0.18 650.3 64.34% 35.66% 344.5 71.53% 28.47% 65.42% 34.58%0.125 756.1 71.53% 28.47% 396 77.23% 22.77% 72.39% 27.61%0.09 600.4 77.25% 22.75% 284.4 81.31% 18.69% 77.86% 22.14%0.063 541.7 82.41% 17.59% 308.3 85.75% 14.25% 82.91% 17.09%0.045 379.7 86.02% 13.98% 208.7 88.75% 11.25% 86.43% 13.57%Pan 1468.6 100.00% 0.00% 782.9 100.00% 0.00% 100.00% 0.00%Total 10505.4 6956.7

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)


T2B01 Centre Product T2B01 Edge Product T2B01 85:15 Scaled Product

211

Figure C-13 T2B01 Particle Size Distributions

212


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%16 15.6 0.16% 99.84% 52.7 0.59% 99.41% 0.22% 99.78%

12.5 71.7 0.89% 99.11% 169.6 2.47% 97.53% 1.12% 98.88%8 333.5 4.27% 95.73% 733.7 10.62% 89.38% 5.22% 94.78%

5.6 472.5 9.07% 90.93% 729.8 18.72% 81.28% 10.52% 89.48%4 568.2 14.84% 85.16% 769.6 27.27% 72.73% 16.70% 83.30%

2.8 637.1 21.30% 78.70% 768.5 35.80% 64.20% 23.48% 76.52%2 650.8 27.91% 72.09% 674.5 43.29% 56.71% 30.22% 69.78%

1.4 699.9 35.02% 64.98% 678.8 50.83% 49.17% 37.39% 62.61%1 572.8 40.83% 59.17% 593.5 57.42% 42.58% 43.32% 56.68%

0.71 731.7 48.26% 51.74% 592.6 64.00% 36.00% 50.62% 49.38%0.5 545.5 53.80% 46.20% 424.4 68.72% 31.28% 56.03% 43.97%

0.355 506.8 58.94% 41.06% 309.6 72.16% 27.84% 60.92% 39.08%0.25 573.4 64.76% 35.24% 329.8 75.82% 24.18% 66.42% 33.58%0.18 555 70.40% 29.60% 300.9 79.16% 20.84% 71.71% 28.29%0.125 635.2 76.84% 23.16% 350.2 83.05% 16.95% 77.78% 22.22%0.09 470.6 81.62% 18.38% 334.8 86.77% 13.23% 82.39% 17.61%0.063 421.1 85.90% 14.10% 302.4 90.13% 9.87% 86.53% 13.47%0.045 317.4 89.12% 10.88% 171.9 92.04% 7.96% 89.56% 10.44%Pan 1071.9 100.00% 0.00% 717.2 100.00% 0.00% 100.00% 0.00%Total 9850.7 9004.5

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



213


214


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%16 0 0.00% 100.00% 27.8 0.24% 99.76% 0.04% 99.96%

12.5 11.9 0.13% 99.87% 185.6 1.85% 98.15% 0.39% 99.61%8 303.2 3.34% 96.66% 955.8 10.16% 89.84% 4.36% 95.64%

5.6 350.5 7.04% 92.96% 919.4 18.14% 81.86% 8.71% 91.29%4 525.2 12.60% 87.40% 1009.9 26.92% 73.08% 14.75% 85.25%

2.8 627.2 19.24% 80.76% 969.7 35.34% 64.66% 21.66% 78.34%2 630.8 25.92% 74.08% 858.3 42.80% 57.20% 28.45% 71.55%

1.4 652.7 32.83% 67.17% 872.5 50.37% 49.63% 35.46% 64.54%1 567.6 38.84% 61.16% 722.3 56.65% 43.35% 41.51% 58.49%

0.71 742.5 46.69% 53.31% 774.6 63.38% 36.62% 49.20% 50.80%0.5 537.7 52.39% 47.61% 588.1 68.49% 31.51% 54.80% 45.20%

0.355 476.8 57.43% 42.57% 423.1 72.16% 27.84% 59.64% 40.36%0.25 553.9 63.29% 36.71% 430.8 75.90% 24.10% 65.19% 34.81%0.18 482.2 68.40% 31.60% 393.9 79.33% 20.67% 70.04% 29.96%0.125 558.2 74.31% 25.69% 514.3 83.79% 16.21% 75.73% 24.27%0.09 467.2 79.25% 20.75% 332.6 86.68% 13.32% 80.37% 19.63%0.063 480.6 84.34% 15.66% 407.1 90.22% 9.78% 85.22% 14.78%0.045 260.3 87.09% 12.91% 236.5 92.27% 7.73% 87.87% 12.13%Pan 1219.4 100.00% 0.00% 889.5 100.00% 0.00% 100.00% 0.00%Total 9447.9 11511.8

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



215


216


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%16 0 0.00% 100.00% 29.8 0.25% 99.75% 0.04% 99.96%

12.5 50.9 0.50% 99.50% 159.2 1.59% 98.41% 0.66% 99.34%8 278.4 3.21% 96.79% 762.9 8.01% 91.99% 3.93% 96.07%

5.6 467.6 7.78% 92.22% 785.3 14.62% 85.38% 8.80% 91.20%4 589 13.53% 86.47% 876.4 21.99% 78.01% 14.80% 85.20%

2.8 736.1 20.71% 79.29% 1021.1 30.58% 69.42% 22.19% 77.81%2 776.1 28.29% 71.71% 936.3 38.46% 61.54% 29.81% 70.19%

1.4 716.8 35.28% 64.72% 963.1 46.56% 53.44% 36.98% 63.02%1 602.3 41.16% 58.84% 795.7 53.26% 46.74% 42.98% 57.02%

0.71 741.3 48.40% 51.60% 856.9 60.46% 39.54% 50.21% 49.79%0.5 584 54.10% 45.90% 678.8 66.18% 33.82% 55.91% 44.09%

0.355 517.7 59.15% 40.85% 477.2 70.19% 29.81% 60.81% 39.19%0.25 552.5 64.55% 35.45% 456.4 74.03% 25.97% 65.97% 34.03%0.18 503 69.46% 30.54% 422.5 77.59% 22.41% 70.68% 29.32%0.125 532.8 74.66% 25.34% 428.6 81.19% 18.81% 75.64% 24.36%0.09 351.4 78.09% 21.91% 336.6 84.02% 15.98% 78.98% 21.02%0.063 463.3 82.61% 17.39% 388.7 87.29% 12.71% 83.31% 16.69%0.045 311.4 85.65% 14.35% 251.5 89.41% 10.59% 86.21% 13.79%Pan 1470.3 100.00% 0.00% 1258.8 100.00% 0.00% 100.00% 0.00%Total 10244.9 11885.8

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



217


218


32 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%25 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%19 0 0.00% 100.00% 0 0.00% 100.00% 0.00% 100.00%16 0 0.00% 100.00% 6 0.04% 99.96% 0.01% 99.99%

12.5 35.8 0.28% 99.72% 23.7 0.21% 99.79% 0.27% 99.73%8 391.2 3.32% 96.68% 497.9 3.81% 96.19% 3.40% 96.60%

5.6 607.4 8.05% 91.95% 858.3 10.02% 89.98% 8.34% 91.66%4 673.5 13.28% 86.72% 863.5 16.26% 83.74% 13.73% 86.27%

2.8 851.5 19.91% 80.09% 1024.6 23.67% 76.33% 20.47% 79.53%2 790.4 26.05% 73.95% 977.6 30.74% 69.26% 26.76% 73.24%

1.4 901.4 33.06% 66.94% 983.3 37.85% 62.15% 33.78% 66.22%1 684.1 38.38% 61.62% 825 43.82% 56.18% 39.20% 60.80%

0.71 770.4 44.38% 55.62% 844.7 49.92% 50.08% 45.21% 54.79%0.5 689 49.74% 50.26% 755.4 55.38% 44.62% 50.58% 49.42%

0.355 581.6 54.26% 45.74% 625.3 59.91% 40.09% 55.11% 44.89%0.25 629.3 59.15% 40.85% 700.5 64.97% 35.03% 60.03% 39.97%0.18 523.6 63.23% 36.77% 660 69.74% 30.26% 64.20% 35.80%0.125 489.1 67.03% 32.97% 640.2 74.37% 25.63% 68.13% 31.87%0.09 308.5 69.43% 30.57% 504.1 78.02% 21.98% 70.72% 29.28%0.063 405.6 72.58% 27.42% 537.2 81.90% 18.10% 73.98% 26.02%0.045 165.9 73.87% 26.13% 283 83.95% 16.05% 75.38% 24.62%Pan 3359.2 100.00% 0.00% 2220.3 100.00% 0.00% 100.00% 0.00%Total 12857.5 13830.6

Percent Passing (%)


Percent Passing (%)

Size (mm)

Weight (g)


Percent Passing (%)

Weight (g)



219


220

D. Appendix D – Stirred Mill Data

221

Figure D-1 355µm Top Size Test Signature Plot Data

Project MASc Research Thesis DateDuty HPGR/Stirred Mill Circuit Location UBC CMPOre Mesaba Ore (Cu/Ni Ore) IsaMill Type M4Test Name 355µm Top Size Test Media Spec. MT1 3.5mm graded charge

Solids SG (t/m3): 3.00 Media Vol (L): 2.5Pass # N (rpm) NLP (kW)Q (sec/L) Pump % kg/L Temp C E (Wh) Time (h)

1 1214 0.68 20 100 1.44 27 279 0.1182 1215 0.68 24 100 1.36 37 420 0.1403 1212 0.68 24 100 1.37 47 407 0.1384 1212 0.68 24 100 1.38 49 396 0.1415 1209 0.68 27 100 1.35 52 396 0.1586 >>> Averaged 1.38 <<<7 Calculated 1.40

Pass # Gross kW Net kW Q (m3/h) % Solids M (t/h) E (kWh/t) Cumul. E p80 p98 CSIFeed 204.0 337.0 1.7

1 2.37 1.69 0.183 54.5% 0.137 12.3 12.3 74.9 175.6 2.32 3.00 2.32 0.149 43.1% 0.089 26.2 38.5 36.9 88.1 2.43 2.95 2.27 0.149 43.1% 0.089 25.6 64.1 17.7 46.3 2.64 2.81 2.13 0.149 42.5% 0.087 24.4 88.5 14.5 36.1 2.55 2.51 1.83 0.134 42.5% 0.079 23.2 111.7 10.3 26.0 2.5

Target p80 Size 75 kWh/t @ Target: 14.17

28-Oct-09

Test Data

Calculated Data

Comments

IsaMill Grinding Test Report

Signature Plot

y = 1518.5x-1.083

R² = 0.972

y = 5000.8x-1.135

R² = 0.9744

1.0

10.0

100.0

1000.0

1.0 10.0 100.0 1000.0

Spec

ific

Ener

gy (k

Wh/

t)

Size (microns)p80 P98 Power (p80) Power (P98)

222

Figure D-2 355µm Top Size Test Particle Size Distributions

% Passing 98 95 90 80 70 60 50 40 30 20 10 5Pass 5 26 18 14 10 8 7 5 4 3 2 1 1Pass 4 36 28 21 14 11 9 7 5 4 2 1 1Pass 3 46 34 26 18 13 10 8 6 4 2 1 1Pass 2 88 68 52 37 29 22 16 12 8 5 2 1Pass 1 176 127 101 75 59 46 36 27 19 10 4 2Feed 337 310 286 204 151 124 88 67 49 32 14 5

Sizings (µm)Sizing Data

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% P

assi

ng

Sizing (µm)


FeedPass 1Pass 2Pass 3Pass 4Pass 5

223

Figure D-3 710µm Top Size Test Signature Plot Data

Project MASc Research Thesis DateDuty HPGR/Stirred Mill Circuit Location UBC CMPOre Mesaba Ore (Cu/Ni Ore) IsaMill Type M20Test Name 710µm Top Size Test Media Spec. 70% 8mm Rojan and 30% 3mm MT1


1 1170 6.72 5 100 1.33 41.55 2928 0.1492 1169 6.72 5 100 1.29 50.44 2621 0.1493 1169 6.72 5 100 1.28 51.97 2608 0.1544 1169 6.72 5 100 1.27 55.46 2426 0.14956 >>> Averaged 1.29 <<<7 Calculated 1.41

Pass # Gross kW Net kW Q (m3/h) % Solids M (t/h) E (kWh/t) Cumul. E p80 p98 CSIFeed 332.0 664.0 2.0

1 19.70 12.98 0.729 39.9% 0.375 34.6 34.6 38.0 98.0 2.62 17.57 10.85 0.727 36.1% 0.339 32.0 66.6 24.3 67.2 2.83 16.91 10.19 0.729 36.1% 0.340 30.0 96.6 16.6 42.1 2.54 16.24 9.52 0.727 36.0% 0.338 28.1 124.7 14.6 37.1 2.55

Target p80 Size 75 kWh/t @ Target: 14.72

30-Oct-09

Test Data

Calculated Data

Comments


Signature Plot

y = 3839.1x-1.289

R² = 0.9916y = 10469x-1.232

R² = 0.9715

10.0

100.0

1000.0

10.0 100.0

Spec

ific

Ener

gy (k

Wh/

t)

Size (microns)p80 P98 Power (p80) Power (P98)

224

Figure D-4 710µm Top Size Test Particle Size Distributions

% Passing 98 95 90 80 70 60 50 40 30 20 10 5Pass 4 37.1 28.5 21.6 14.6 11.2 8.7 6.6 4.8 3.3 2.1 1.2 0.7Pass 3 42.1 31.9 24.7 16.7 12.6 9.7 7.2 5.1 3.5 2.1 1.2 0.7Pass 2 67.2 47.1 34.9 24.3 17.6 13.1 9.8 6.8 4.4 2.6 1.3 0.7Pass 1 98.0 74.0 56.3 38.0 28.5 21.4 15.4 11.0 7.1 3.9 1.7 0.9Feed 664 595 485 332 239 174 124 95 63 36 18 9


0

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% P

assi

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Sizing (µm)


Feed

Pass 1

Pass 2

Pass 3

Pass 4

225

Figure D-5 T2C02 Signature Plot Data (T1)

Project Name MASc Research Thesis Date(s) TestedDuty HPGR/Stirred Mill Circuit Date IssuedOre Mesaba Ore (Cu/Ni Ore) Location UBCTest Number T2C02 IsaMill Type M20Contact Person Jeff Drozdiak Media Spec. Cenotec 6mm graded charge)


1 1170 6.39 3 100 1.59 26.84 1088 0.0662 1170 6.39 3 100 1.59 37.65 1004 0.0613 1169 6.39 3 100 1.59 44.16 935 0.0594 1169 6.39 3 100 1.59 49.64 875 0.0585 1169 6.39 3 100 1.59 54.33 852 0.0586 1170 6.39 3 100 1.59 57.65 819 0.0577 1168 6.39 3 100 1.59 60.43 780 0.055

>>> Avg. 1.59 <<<Calc. 1.61

Pass # Gross kW Net kW Q (m3/h) % Solids M (t/h) E (kWh/t) Cumul. E P80 P98 CSIFeed 338.6 639.2 1.9

1 16.53 10.14 1.224 56.9% 1.108 9.1 9.1 83.8 262.9 3.12 16.50 10.10 1.229 56.9% 1.112 9.1 18.2 42.0 106.6 2.53 15.96 9.56 1.224 56.9% 1.108 8.6 26.9 31.5 74.8 2.44 15.21 8.82 1.224 56.9% 1.108 8.0 34.8 25.5 59.5 2.35 14.67 8.27 1.224 56.9% 1.108 7.5 42.3 21.6 49.6 2.36 14.39 7.99 1.224 56.9% 1.108 7.2 49.5 19.4 44.1 2.37 14.18 7.79 1.224 56.9% 1.108 7.0 56.5 17.0 39.0 2.3

Target P80 Size (if applic. 75 kWh/t @ Target: 10.0 Media Consumption (g/kWh):

Test Data

Calculated Data

Comments

IsaMill Grinding Test Report08-Nov-1009-Dec-10

Signature Plot

y = 1512x-1.164

R² = 0.9966

y = 1813.9x-0.963

R² = 0.9896

1.0

10.0

100.0

1000.0

1.0 10.0 100.0 1000.0

Spec

ific E

nerg

y (kW

h/t)

Size (microns)P80 P98 Power (P80) Power (P98)

226

Figure D-6 T2C02 Particle Size Distributions (T1)

% Passing 98 95 90 80 70 60 50 40 30 20 10 5

Pass 7 39.0 30.7 24.0 17.0 12.9 10.0 7.6 5.5 3.8 2.3 1.2 0.8

Pass 6 44.1 34.9 27.3 19.4 14.8 11.2 8.7 6.3 4.3 2.6 1.4 0.8

Pass 5 49.6 38.9 30.4 21.6 16.5 12.7 9.6 7.0 4.7 2.9 1.4 0.9

Pass 4 59.5 46.4 36.0 25.5 19.2 14.8 11.2 8.1 5.4 3.2 1.6 0.9

Pass 3 74.8 58.2 44.9 31.5 23.7 18.1 13.6 9.9 6.6 3.8 1.8 1.0

Pass 2 106.6 80.7 61.2 42.0 31.0 23.3 17.4 12.4 8.2 4.7 2.1 1.1

Pass 1 262.9 175.4 125.9 83.8 60.6 44.4 32.0 22.4 14.5 8.1 3.1 1.5

Feed 639.2 532.1 451.2 338.6 253.5 192.1 142.4 97.4 63.7 37.4 18.7 9.4


0

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Sizing (µm)


FeedPass 1Pass 2Pass 3Pass 4Pass 5Pass 6Pass 7

227

Figure D-7 T2C03 Signature Plot Data (T2)

Project Name MASc Research Thesis Date(s) Tested 17-Jan-11Duty Description HPGR/Stirred Mill Circuit Date Issued 20-Jan-11Ore /Conc. Type Mesaba Ore (Cu/Ni Ore) Location UBCTest Number T2C03 IsaMill Type M20Contact Person Jeff Drozdiak Media Spec. Cenotec 6mm Graded Charge


1 1169 6.61 3 100 1.62 21.1 866 0.0632 1168.5 6.61 3 100 1.62 27.92 749 0.0583 1169 6.61 3 100 1.62 33.37 746 0.0584 1168.5 6.61 3 100 1.62 38.34 723 0.0565 1169 6.61 3 100 1.62 42.9 700 0.0566 1168 6.61 3 100 1.62 46.2 670 0.0547 1168 6.61 3 100 1.62 49.6 637 0.053

>>> Avg. 1.62 <<<Calc. 1.62

Pass # Gross kW Net kW Q (m3/h) % Solids M (t/h) E (kWh/t) Cumul. E P80 P98 CSIFeed 340.2 647.7 1.9

1 13.74 7.13 1.224 57.1% 1.133 6.3 6.3 121.7 374.2 3.12 12.96 6.35 1.226 57.1% 1.135 5.6 11.9 54.5 140.5 2.63 12.91 6.31 1.220 57.1% 1.130 5.6 17.5 38.0 88.1 2.34 12.83 6.22 1.224 57.1% 1.133 5.5 23.0 31.9 72.4 2.35 12.54 5.94 1.220 57.1% 1.130 5.3 28.2 26.8 60.6 2.36 12.37 5.77 1.220 57.1% 1.130 5.1 33.3 24.1 53.3 2.27 12.13 5.52 1.220 57.1% 1.130 4.9 38.2 20.9 45.4 2.2

Target P80 Size (if applic. 75 kWh/t @ Target: 9.5 Media Consumption (g/kWh):

Test Data

Calculated Data

Comments


Signature Plot

y = 864.93x-1.044

R² = 0.9838

y = 969.22x-0.867

R² = 0.978

1.0

10.0

100.0

1000.0

1.0 10.0 100.0 1000.0

Spec

ific E

nerg

y (kW

h/t)

Size (microns)P80 P98 Power (P80) Power (P98)

228

Figure D-8 T2C03 Particle Size Distributions (T2)

% Passing 98 95 90 80 70 60 50 40 30 20 10 5

Pass 7 45.4 36.4 28.9 21.0 16.1 12.5 9.6 7.1 4.8 2.9 1.5 0.9

Pass 6 53.3 42.3 33.4 24.1 18.5 14.4 11.0 8.0 5.4 3.2 1.6 0.9

Pass 5 60.6 47.7 37.3 26.8 20.5 15.8 12.1 8.9 5.9 3.5 1.7 1.0

Pass 4 72.4 57.2 44.8 31.9 24.3 18.8 14.3 10.4 7.0 4.0 1.9 1.0

Pass 3 88.1 69.3 53.8 38.0 28.7 22.0 16.7 12.1 8.2 4.7 2.1 1.1

Pass 2 140.5 103.3 78.8 54.5 40.4 30.4 22.7 16.3 10.8 6.1 2.5 1.3

Pass 1 374.2 278.8 200.1 121.7 87.1 61.9 43.4 29.5 18.9 10.5 3.9 1.8

Feed 647.7 554.2 458.9 340.2 255.9 193.3 144.3 99.0 65.4 36.3 18.2 9.1


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Sizing (µm)


FeedPass 1Pass 2Pass 3Pass 4Pass 5Pass 6Pass 7

Ubc 2011 Spring Drozdiak Jeffrey

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