Kitaly Venance D. - Blast Design Optimization to Improve Material Fragmentation-complete report

THE UNIVERSITY OF DODOMA

School of Mines and Petroleum Engineering

DEPARTMENT OF MINING AND MINERAL PROCESSING

B. Sc. MINING ENGINEERING

FINAL YEAR PROJECT

ON

BLAST DESIGN OPTIMIZATION TO IMPROVE MATERIAL

FRAGMENTATION

Course Name/Code: MINING PROJECT 2 (MN 499)

Student’s Name: KITALY VENANCE D.

Registration No: T/UDOM/2009/08472

Signature: ………………

Submission Date: 7th June, 2013

Supervised by: Eng. H. KAMANDO and

Mr. LUPYANA, Samwel D.

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page i

STATEMENT OF DECLARATION

This project report is specifically intended for the partial fulfilment of the bachelor’s degree

in Mining Engineering as per requirement by the Department of Mining and Mineral

Processing Engineering, College of Earth Sciences – The University of Dodoma – Tanzania.

I hereby declare that this is my own original work and has never been submitted before to

any other institution for the award of degree or diploma. Wherever other people’s work

and/or ideas have been used in this work, it has been cited and referenced appropriately.

© KITALY Venance D. (2013)

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page ii

ACKNOWLEDGEMENTS

I would like to show my sincere gratitude to the Almighty God for giving me life and health

for which I have been able to conduct this study.

I also want to thank my supervisors, Mr. Lupyana Samwel D. and Eng. H. Kamando for their

help and guidance throughout the time I have been doing this study. I also thank them for

their inspiring attitude towards me which has always given me more belief in my capabilities

and potentials.

I would also like to thank Williamson Diamonds Limited (WDL) for providing me with the

site and facilities for doing my study. The information they provided were very helpful

throughout this study. Specifically, I would like to thank Eng. Sixtus Massota and Mr.

Timothy Kabondo (site geologist) for their cooperation in providing me with the necessary

information especially historical data which have been used in this study.

I won’t forget to thank my fellow colleagues, B Sc. Mining Engineering (4th

year) for their

significant help and support in this study. Their comments and suggestions have helped me a

lot in accomplishing this work.

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page iii

ABSTRACT

This project report is aimed to explain the study done at Williamson Diamonds Limited

(WDL). The main objective of this study was to optimize the blast design for improving

material fragmentation.

To do this study, data were collected from WDL during February-2013. The data collected

included raw drilling and blasting data and their associated fragmentation and historical data

concerning drilling and blasting operations for the year 2012. These data were then analysed

and the result were discussed so as to come up with optimal values for spacing, burden

powder factor and hole depth.

Direct measurement of spacing, burden, hole diameter and hole depth were done in field by

using tape measure and the results were recorded in a field note book. Explosives, charging

techniques, delay set up, tie up, initiation system and fragmentation were observed directly

from the field and digital camera were used to take some illustrative photos. From explosives

utilization and charging techniques, powder factor was estimated by dividing the explosive

used (kg) by the volume of material blasted (m3). Historical data were obtained from

company records by asking some questions regarding the data of interest from responsible

site supervisors.

Tables and charts (by Microsoft office- excel) were used to represent the quantitative data

prior to analysis. The data were then analysed by performing trend analysis on the historical

data. Graphs for cost against fragmentation were plotted to obtain the optimal fragmentation

(at minimum total cost). Design parameters (spacing, burden, powder factor) were also

plotted against fragmentation to obtain the optimal values by using excel package.

The main findings for this study were predicted optimal parameters. Fragmentation was

predicted to be 95% (-300mm), spacing 2m, burden 1.9m and powder factor 1.3kg/m3. The

hole diameter was maintained at 102mm.

It was concluded that, the predicted values were optimal and had to be applied in blocks of

the pit where fragmentation were bad. The hole depth suitable for these parameters was

concluded to be 5.7m or more.

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page iv

It was recommended to do a study on viability of installing a crushing system in the

treatment plant in order to accommodate less fragmented material and reduce explosive costs

as well as increasing recovery of diamonds from the ore. For further study, it was

recommended that these parameters be applied and see their results so that the real optimal

parameters can be obtained.

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page v

LIST OF ABBREVIATIONS

ANFO – Ammonium Nitrate + Fuel Oil mixture (explosive)

BOUMA – fine and course sequences of bouma facies

BVK – Brecciated Volcanic Kimberlite (kimberlitic dyke)

CL – Clay

Dh – hole diameter

GB – Granite Breccias

GN – 100% granite

NE – North East

NW – North West

RVK – Re-worked Volcanic Kimberlite

SE – South East

SW – South West

WDL – Williamson Diamonds Limited

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page vi

TABLE OF CONTENTS

STATEMENT OF DECLARATION ..................................................................................... i

ACKNOWLEDGEMENTS .................................................................................................. ii

ABSTRACT ........................................................................................................................ iii

LIST OF ABBREVIATIONS ................................................................................................v

TABLE OF CONTENTS .................................................................................................... vi

CHAPTER 1 .........................................................................................................................1

INTRODUCTION .................................................................................................................1

1.1: Overview of Drilling and Blasting Operations .............................................................1

1.2: Overview of Williamson Diamonds Limited (WDL) ...................................................2

1.2.1 Location ................................................................................................................2

1.2.2: Drilling and blasting at WDL ...............................................................................3

1.3: Problem Statement ......................................................................................................4

1.4: Aims and objectives ....................................................................................................5

1.4.1: Aims .....................................................................................................................5

1.4.2: Objectives ............................................................................................................5

1.4.2.1: Main Objective ..............................................................................................5

1.4.2.2: Specific Objectives ........................................................................................5

1.5: Project scope ...............................................................................................................5

CHAPTER 2 .........................................................................................................................6

LITERATURE REVIEW ......................................................................................................6

2.1: Background.................................................................................................................6

2.2: Design parameters .......................................................................................................7

2.2.1: Hole Diameter ......................................................................................................8

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page vii

2.2.2: Burden .................................................................................................................8

2.2.3: Spacing ................................................................................................................9

2.2.4: Sub-grade ........................................................................................................... 10

2.2.5: Stemming............................................................................................................ 10

2.2.6: Bench height/Hole depth .................................................................................... 11

2.2.7: Timing/Delay setup ............................................................................................ 11

2.2.8: Powder factor..................................................................................................... 14

2.3: Costs/benefits associated ........................................................................................... 15

CHAPTER 3 ....................................................................................................................... 19

METHODOLOGIES ........................................................................................................... 19

CHAPTER 4 ....................................................................................................................... 22

DATA COLLECTION AND ANALYSIS ........................................................................... 22

4.1: Collection of Data ..................................................................................................... 22

4.2: Analysis of the Data .................................................................................................. 22

CHAPTER 5 ....................................................................................................................... 30

RESULTS AND DISCUSSIONS ........................................................................................ 30

5.1: Results ...................................................................................................................... 30

5.2: Discussions ............................................................................................................... 30

CHAPTER 6 ....................................................................................................................... 34

CONCLUSIONS ................................................................................................................. 34

CHAPTER 7 ....................................................................................................................... 35

RECOMMENDATIONS ..................................................................................................... 35

REFERENCES.................................................................................................................... 36

APPENDICES .................................................................................................................... 40

Appendix 1: Field raw data .............................................................................................. 40

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page viii

Appendix 2: Extracted data from WDL archive (historical data) for various purposes ...... 58

Appendix 3: Some photos showing fragmentation at different blasts (Dh = 102mm)......... 63

LIST OF TABLES

Table 2.2: Selected Factors for First-Approximation Surface Blast Designs ..........................7

Table 2.2.3: Spacing formulae under different conditions (all parameters in feet (ft))............9

Table 2.3: Comparison of Drilling and Blasting Costs for Various Mining Methods ........... 16

Table 4.2: Extracted data for plotting blasting parameters models ....................................... 27

Table 9.1(a): Blasting raw data as at 22nd

February, 2013(Dh = 102mm) ............................. 40

Table 9.1(b): Explosive utilization for the 22-Feb-13 blast (NW Block C) .......................... 54

Table 9.1(c): Explosive utilization for the 27-Feb-13 blast (SW Block C) ........................... 56

Table 9.2(a): The whole data range analysed for costs (Dh = 102mm) ................................. 58

Table 9.2(b): The whole data range analysed for specific parameters (Dh = 102mm) ........... 60

LIST OF FIGURES

Fig. 1.1: Tanzania map showing location of Mwadui ............................................................3

Fig. 2.2: Typical initiation patterns for surface blasting showing initiation by rows (Be and

Se are effective burden and spacing, respectively) ............................................................... 12

Fig. 2.2.8: Cost – powder factor relationship ....................................................................... 15

Fig. 2.3: Optimal fragmentation on cost basis ...................................................................... 17

Fig. 4.2(a): WDL pit major blocks showing rock types and characteristics (grade and

density) ............................................................................................................................... 23

Fig. 4.2(b): Drilled holes deviations from the designed depths ............................................ 24

Fig. 4.2(c): Fragmentation optimization models .................................................................. 27

Fig. 4.2(d): Spacing optimization model ............................................................................. 28

Fig. 4.2(e): Burden optimization model ............................................................................... 29

Fig. 4.2(f): Powder factor optimization model ..................................................................... 29

Fig. 5.2: Very poor fragmentation at NW Block C due to improper charging (16- Feb-12) .. 33

Fig. 9.1(a): Depicted timed, tied up pattern for the blast at NW Block C (on 22-Feb-13) ..... 54

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page ix

Fig. 9.1(b): Blasting sequence showing direction of the muck-pile (V-cut) at NW Block C on

22-Feb-13 ............................................................................................................................ 55

Fig. 9.1(c): Good rock fragmentation with spacing and burden of 2.5m x 2.5m as at 22-Feb-

13 at NW Block C (CL-RVK) ............................................................................................. 55

Fig. 9.1(d): A photo showing the pattern for blasting at SW Block C on 27-Feb-13 ............ 56

Fig. 9.1(e): Poor rock fragmentation with spacing and burden of 2.5 x 2.5m as at 27-Feb-13

at SW Block C (NCL-RVK and GB) ................................................................................... 57

Fig. 9.3(a): Poor fragmentation with burden and spacing 3m x 3m as at 27-Jan-2012 at NW

Block C (CL-RVK) ............................................................................................................. 63

Fig. 9.3(b): Poor rock Fragmentation with burden and spacing 3m x 3.5m as at 6-Feb-2012 at

SW Block C (GB/BVK) ...................................................................................................... 64

Fig. 9.3(c): Good rock Fragmentation with Burden and Spacing of 2.5m x 2.5m.as at 16-Feb-

2012 at NW Block C (CL-RVK) ......................................................................................... 64

© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 1

CHAPTER 1

INTRODUCTION

1.1: Overview of Drilling and Blasting Operations

Back in early 16th Century, people used various methods to break the rock. One of the

common methods used was fire setting where the rock would be heated up to very high

temperatures, then quenched with a stream of cold water which resulted into thermal shock

that broke the rock.

It was until 1627 when the first explosive in history (black powder) was used in Hungary for

rock breakage. Since then, explosives have been one of the cheapest and most efficient

means of rock breakage in hard rock mining (Society of Explosives and Engineers Inc.,

2010).

To break the rock, explosives need to be well confined so that the energy released can’t

escape to the atmosphere and get lost. For primary and sometimes secondary blasts,

confinement is obtained through drilling.

Drilling is the process of making holes on the rock for various purposes especially blasting.

Drilling is done by either of the following mechanisms;

Down The Hole (DTH)

Rotation

Percussion

Percussion-Turn (P-Turn)

Blasting normally follows after drilling. The drilled holes are loaded/charged with explosives

and tied up then initiated electrically or non-electrically. Proper designed and organised blast

provides good fragmentation of the blasted material.

Material fragmentation is one of the most important aspects to be considered by any mining

engineer. Degree of fragmentation to the large extent affects the productivity of mining

operations especially in surface mines (Ramulu, 2012).


The use of blasting as a method to break the rock requires proper selection of explosives and

blasting devices, the careful design of borehole patterns, loading characteristics, and delay

blasting sequence, and the control of ground vibration, air-blast, and fly-rock. Efficient blast

designs produce the desired particle size distributions and placement of muck-piles for ease

of rock removal and handling (Hartman, 1996).

1.2: Overview of Williamson Diamonds Limited (WDL)

Williamson Diamond Limited (WDL) is an open pit mine and is the largest operating

diamond mine in the world, found in Tanzania. During the time of this study, mining

activities in WDL were divided into two distinct operations which were

In-pit mining; this involved mining within the kimberlite pipe through open pit

mining and

Alluvial/Gravel mining; this involved mining of alluvial/placer diamonds which

have been eroded from their area of origin and get deposited somewhere else. It was

done by removing the upper dark soil to uncover the diamond hosting soil which was

then treated to recover the diamonds in it.

1.2.1 Location

WDL is located at Mwadui area in Kishapu District in Shinyanga region. Mwadui is located

at a few kilometres from Shinyanga – Mwanza road, and just several kilometres from a small

town known as Maganzo. The climate of Mwadui is dry and semi arid having two seasons,

namely, wet season (Nov to Apr) and dry season (May to Oct). Average temperature varies

from 17 degrees centigrade to 33 degrees centigrade.


Fig. 1.1: Tanzania map showing location of Mwadui (WDL Archive)

1.2.2: Drilling and blasting at WDL

Drilling at WDL was normally done by using DTH drill rigs. The diameter used was

normally small diameter (102mm) due to the capacity of the available drill rigs. WDL

benches were 5m high and 450 in slope as per design. This limited the hole depths in most of

the cases to be 5m, though sometime more than 5m holes were also drilled. Spacing and

burden used at WDL were 2.5m and 2.5m respectively in a staggered drilling pattern.

Charging operations were done by both mechanised and manual methods. Two types of main

charge/explosives were normally used. Emulsion was mostly used in wet holes and its

charging was mechanised. On the other hand, ANFO was used in dry holes and its charging

was manual. Pentolite boosters were normally used for making primers for initiation of the

main charge. Stemming length used at WDL was 2m as per design and no sub-drill was done.

As for stemming material drill chips were normally used.


Timing was normally delayed, with inter-hole delay 42ms and inter-row delay also 42ms.

Tying up pattern used was normally V-cut, but firing parallel to a single free face was also

common.

Blasting was initiated by combined electrical and non-electrical initiation system. This was

done by means of electrical detonator and a detonating cord. The detonating cord connected

the entire shot to electrical cables through an electric detonator. The blasting machine was

then used to send an electrical signal to the shot through electrical cables.

Drilling and blasting operations at WDL mine were expected to produce good fragmentation

of at least 80% pass (-300mm) through the plant scalp bin grizzle. However, attaining this

fine fragmentation had been quite a challenge.

1.3: Problem Statement

Poor fragmentation had been a problem in many mining companies in Tanzania particularly

at Williamson Diamonds Limited. This had been one of the sources of under-performance in

terms of material handling as well as ore treatment operations.

The WDL treatment plant didn’t have any crushing system. This led WDL to use blasting as

the only means of reducing size of materials to the size which can be handled by milling

plant. The required size is -300mm.

Blasting parameters used however, were producing large amount of boulders (of up to 1.5m)

that have been leading to frequent plant stoppages and delays, which also affects loading and

hauling operations.

From the company’s treatment plant statistics for the month of July 2012 for instance ,

41.87% of the total available hours was wasted as delays from various reasons (including

plant maintenance); and 10.65% of these delays was related to the oversize occasions! This

in turn was found to contribute 18.44% of the queuing time for dump truck DT 41.

All these were burden to the company as they increased operational costs but also handling

boulders (both from in-pit and plant scalp bin) to the tailing dump was quite a costing task.

These costs could be reduced if a careful study was done on fragmentation improvement.


1.4: Aims and objectives

1.4.1: Aims

To study the current drilling and blasting parameters and the associated fragmentation

of the blasted material.

To observe if whether all drilling, charging, timing and initiation operations are done

as indicated in the current designs or not.

To determine if whether the poor fragmentation is caused by unsuitable design

parameters or something else.

1.4.2: Objectives

1.4.2.1: Main Objective

To determine the optimal drilling and blasting parameters that will give good

fragmentation of the blasted material.

1.4.2.2: Specific Objectives

To estimate the optimal spacing, burden and hole depth that can ensure good

fragmentation; by considering rock properties (density, ore and/or waste).

To determine the optimal powder factor considering explosive costs as well as

fragmentation required.

To ensure proper drilling of holes with required depth and at exact locations (as

indicated in the drilling pattern) in each blast.

1.5: Project scope

This project concentrated on important parameters that were required for achieving good

fragmentation. Such parameters included hole diameter, burden, spacing, hole depth and

powder factor. Other design parameters don’t affect fragmentation directly, and was not dealt

with in detail in this study.


CHAPTER 2

LITERATURE REVIEW

2.1: Background

Material fragmentation has been one of the main focuses of engineers in mining industry

especially in blasting operations. This is of great importance because to the large extent, the

successiveness of all other operations in the production chain depends mainly on the

effectiveness of blasting (fragmentation attained).

In the field of blasting technology however, the researchers are confronted with the problem

of developing adequately accurate quantity indexes for determining the rock fragment size

distribution in mass blasting. The difficulties are to the greatest part caused by the fact that

the rock is neither homogeneous nor isotropic, the structural properties in the rock mass may,

even when the rock type is the same, change from one site to another (Strelec, 2011).

Despite these challenges, many researchers have done a great job and came up with some

solutions on how to design a sound blasts, and to the large extent, their work has been

providing satisfying results including reasonable fragmentation in the blasted material. Some

formulae and software have been developed, which can predict and give the base

approximations towards establishment of suitable drilling and blasting parameters.

Some of the very important parameters to be addressed in drilling and blasting design

include;

Hole diameter

Hole depth/Bench height

Hole spacing

Burden

Stemming and /or Charge length

Sub-drill/sub-grade length and

Timing/delay setup

Powder factor


2.2: Design parameters

There are no specific design principles or procedures in drilling and blasting that can work

perfectly in each site/occasion. An empirical approach is therefore taken in blast design. This

approach is necessary due to the many factors that cannot be controlled, such as geology and

explosive loading conditions.

According to Ash (1963), Pugliese (1973), Van Ormer (1973), Hagen (1981), Dick et al.

(1983), and many others; among other parameters, borehole diameter and burden are perhaps

the most important factors used in design. Fragmentation and size distribution are a function

of burden and hole diameter.

Ash (1963) has provided simple empirical formulas to compute burden, spacing, sub-grade,

and stem lengths using “K factors,” as shown in table 2.2 below;

Table 2.2: Selected Factors for First-Approximation Surface Blast Designs

Parameter Information

Burden B = KBD

Where;

D is hole diameter

Using ANFO:

KB = 22 for rock density < 2.7 g/cm3

= 30 for rock density > 2.7 g/cm3

Using slurry, dynamite or other high explosive:

= 27 for rock density < 2.7 g/cm3

= 35 for rock density > 2.7 g/cm3

Spacing S= KSB KS = 1 to 2, depending on initiation

Sub-grade J = KJB KJ = 0.2 to 0.5 (average 0.3)

Stemming T = KTB KT = 0.5 to 1.3 (average 0.7)

Source: Ash, 1963.


2.2.1: Hole Diameter

Hole diameter is usually set by the drill rig capacity, which is matched to the range of hole

depths anticipated for the job. It is therefore convenient to select hole diameter (based on

your drilling equipment) and from which, the burden can be established. In this study, small

hole diameter (102mm) is selected due to the capacity of the available drilling rigs.

2.2.2: Burden

Burden values should be selected based on geology and explosive energy output. Excessive

burden resists penetration by explosion gases to effectively fracture and displace the rock and

part of the energy may become seismic intensifying blast vibrations. This phenomenon is

most evident in pre splitting blasts, where there is total confinement and vibration levels can

be up to five times those of bench blasting.

Small burden on the other hand lets the gases escape and expand with high speed towards the

free face, pushing the fragmented rock and projecting it uncontrollably, provoking an

increase in overpressure of the air, noise and fly-rock.

Numerous formulas have been suggested to calculate the burden, which take into account one

or more of the parameters (like hole diameter and bench height); however, their values all fall

in the range of 20 to 40 D, depending fundamentally upon the properties of the rock mass

(Rajpot, 2009).

For example Ash (1963) suggested the following formula for burden;

Burden B = KBD

Where;

KB is a constant ranging from 22 to 40

Austin Powder Company (2002) provides the formula for burden based on densities of rock

and explosive and the diameter of the coupled explosive column as shown below;

B = De * [2* (de/dr) + 1.5]


Where;

De is the diameter of the coupled explosive (in)

de is density of explosive (g/cm3)

dr is density of rock (g/cm3) and

B is burden (ft)

2.2.3: Spacing

Spacing is calculated as a function of burden, delay timing between blast-holes and initiation

sequence. Very small spacing causes excessive crushing between charges and superficial

crater breakage, large blocks in front of the blast-holes and toe problems. Excessive spacing

between blast-holes causes inadequate fracturing between charges, along with toe problems

and an irregular face (Jimeno, 1995).

Austin Powder Company (2002) provides formulae for spacing based on the bench height

and burden ratio (H/B) as shown in table 2.2.3 below;

Table 2.2.3: Spacing formulae under different conditions (all parameters in feet

(ft))

Timing Formula

If H/B ≥ 4 If H/B < 4

Instantaneous S =0.2B S = [H + 2B]/3

Delayed S =1.4B S = [H + 7B]/8

Source: Austin Powder Company (2002)

Again; Ash (1963) suggests that spacing can be given by the formula;

S = KsB where Ks is a constant ranging from 1 to 2 (depends on initiation)


2.2.4: Sub-grade

If the sub-drilling is small, then the rock will not be completely sheared off at floor level,

which will result in toe appearance and a considerable increase in loading costs. However, if

sub-drilling is excessive, the following will occur:

An increase in drilling and blasting costs.

An increase in vibration level.

Excessive fragmentation in the top part of the underlying bench, causing drilling

problems of the same and affecting slope stability in the end zones of the open pit.

Increase in risk of cut-offs and over-break, as the vertical component of rock

displacement is accentuated (Rajpot, 2009).

For vertical blast-holes when a bench is massive, the sub-drilling distance suggested by Ash

(1968), Gustafsson (1973), Jimeno et al. (1995) should be approximately equal to 30% of the

burden. Hustrulid (1999), on the other hand proposes that the drilled distance of the hole to

the toe elevation (the sub-drilling distance) should be equal to 8 diameters.

2.2.5: Stemming

Stemming is one of the important parameters that affect directly the effectiveness of blast

including material fragmentation. Like other parameters, stemming has to be designed and

optimised and this should be done very carefully. It is a common practice to take stemming

with no much care especially in terms of selection of stemming material, where drilling chips

has been readily used; taking consideration of their simple availability near the blast-hole.

However, sometimes this practise may affect much the effectiveness of blast (specifically

material fragmentation). Konya (1990) and Jimeno et al. (1995) suggests an optimum

borehole diameter to stemming material particle diameter ratio of about 17:1.

In terms of stemming length, Jimeno et al. (1995) proposes that optimum lengths of

stemming should increase as the quality and competence of the rock decrease, varying

between 20D and 60D, where D is the diameter of the borehole. Whenever possible, a

stemming length of more than 25D should be maintained in order to avoid problems of air-

blast, fly-rock, cut-offs, and over-breaks. Another study was done by Ash (1968) who


concluded that the amount of stemming or collar should be used as a direct function of the

burden as follows;

T = KTB

Where;

B is burden and

KT is a constant (ranges from 0.5 to 1.3)

However both approaches give the acceptable results and any of them can be used in this

work.

2.2.6: Bench height/Hole depth

Bench height is usually determined by the working specifications of loading equipment. The

bench height limits the size of the charge diameter and the burden.

When the bench height to burden ratio is large, it is easy to displace and deform rock,

especially at the bench centre. The optimum ratio (Hb/B) is larger than 3. If Hb/B = 1, the

fragments will be large, with over-break/back-break around holes and toe problems. With

Hb/B = 2, these problems are attenuated and are completely eliminated when Hb/B >3 (Ash,

1968).

When Hb is small, any variation in the burden B or spacing S has a great influence on the

blasting results. When Hb increases, with B kept constant, spacing can increase to maximum

value without affecting fragmentation. However, if the bench height is very large, there can

be problems of blast-hole deviation, which will not only affect rock fragmentation but will

also increase risk of generating strong vibrations, fly-rock, and over-break because the

drilling pattern and subsequently the explosives consumption will not remain constant in the

different levels of the blast-hole (Rajpot, 2009).

2.2.7: Timing/Delay setup

Delay blasting techniques are employed to improve fragmentation, control of rock

movement, over-break, and to reduce ground vibrations. The delay patterns used in design


will determine the sequence of hole or deck initiations, thereby, dictate the overall direction

of blasted rock movement and resulting fragmentation.

Timing between detonating charges depends on the spacing/burden (S/B) ratio, and it is

normally recorded in milliseconds (ms). This determines the muck-pile displacement height

and the distance from the bench.

Depending on initiation sequence, an effective burden Be and effective spacing Se result as

shown in Fig.2.2. The effective spacing is the distance between holes in a row defined by

adjacent time delays (e.g., delays by rows). Effective burden is the distance in the direction

of resultant rock mass movement. The V and echelon (diagonal) patterns are used when rock

placement is restricted. Designs using two free faces usually provide improved fragmentation

and throw control over those using a single face.

Fig. 2.2: Typical initiation patterns for surface blasting showing initiation by rows (Be and

Se are effective burden and spacing, respectively) (Hartman, 1996).


Timing however must be optimised. If the inter-hole delay is too short, the movement of row

burdens is restricted and fragmentation is poor. Again, if inter-hole delays are too long, cut-

offs of surface delays may occur. The minimum time for design is controlled by the stress

wave travel time (= 2Be) in order for radial cracking to begin to develop, contributing to the

detachment of the rock mass in the vicinity of the hole. This detachment forms an internal

free face (or relief) to which successive detonations will interact with the reflection of stress

waves. The minimum timing is, therefore;

t = 2Be/Cp *103 Where;

t is stress wave travel time in ms

Be is effective burden or distance from the hole to the free face in feet

and

Cp is velocity of sound for the rock in fps (Hartman, 1996).

The maximum timing is that at which the burden is fully detached and accelerating as gas

pressures build.

Hagen (1977) noted the time to burden movement ranges from 5 to 50 ms, and suggests an

optimum range of timing for design between 1.5 to 2.5 ms/ft of Be. Timing studies have been

performed to investigate resulting fragmentation and muck pile shapes.

Reduced-scale research using a variation in delay ratios suggests improved fragmentation for

timing between 11 to 17 ms/ft of Be (Stagg, 1987), while Bergmann et al. (1974)

demonstrated improved fragmentation for S/B ratios of two at timing ratios of 1 ms/ft of Be

or greater. Production-scale, multiple-row blasting has resulted in recommended timing to

improve fragmentation. Andrews (1981) suggests delays of 1 to 5 ms/ft within rows and 2 to

15 ms/ft of Be between rows (or on the echelon). Anderson et al. (1982) measured flyrock

velocity, or gas venting, through the collar stemming to establish a 3.4 ms/ft of hole spacing

and 8.4 ms/ ft of Be recommendation for optimum breakage and forward movement.


Similar work in which muck pile profiles were mapped indicates that optimum forward

throw and muck pile height reduction occur for delay ratios of 4.2 ms/ft of Se and 10 ms/ft of

Be, while forward throw is minimized, resulting in high muck piles, with ratios of 1.5 to 2

ms/ft of Se and 5 to 6 ms/ft of Be (Winzer, 1981).

Hagen (1977) has shown for single row production shooting and S/B of 1.2 to 1.6 that timing

ratios greater than 1.2 ms/ft of Be are ideal. Hagen recommended 1.2 ms/ft of Be for

multiple-row production blasting in hard rock, while using high powder factors and short

stem lengths. A 2.4 ms/ft of Be was recommended for soft rock with long stem lengths and

low powder factors. To control ground vibrations, Kopp (1987) recommended that a 1.3

ms/ft of S and 1.2 to 4.3 ms/ft of Be be used.

The timing ratios cited are found to vary over a wide range. A great deal of research on the

effects of initiation timing cannot be compared due to the lack of similar variables such as

geology, scale, and explosive type.

Winzer et al. (1983) recognized the need to qualify delay ratios, in a general way, based on

existing fracture density. Competent dense rock requires lower delay ratios to achieve fine

fragmentation, while weak fractured rock fragments best with higher delay ratios.

2.2.8: Powder factor

Powder factor refers to the volume or tonnage of material broken by a given weight of

explosives. It is expressed in (kg/m3) or (kg/tonne). It can serve a variety of purposes, such as

an indicator of how hard the rock is, or the cost of the explosives needed, or even as a guide

to planning a shot.

Very small powder factor will lead to poor fragmentation and subsequent increase in

operation costs. On the other hand, very large powder factor will lead to over-fragmentation

and again increase in overall operating costs. As other parameters therefore, powder factor

has to be optimised in order to realise cost effectiveness of the entire mining operation

(Assakkaf, 2003).

The relationship between powder factor and the overall total cost is shown in figure 2.2.8


Fig. 2.2.8: Cost – powder factor relationship (Mishra, 2009)

2.3: Costs/benefits associated

Drilling and blasting are unit operations required for development and production. The

components of costs for drilling and blasting include labour, direct costs of operating

equipment, and supplies. In surface mines, the basis of cost is computed per ton (tonne) of

ore produced or per cubic yard (cubic meter) of material broken for removal. The costs are

directly related to powder factor and depend on geology, type of explosives, and the size of

the blast-holes and excavating equipment.

For equipment, such as a drill rig, the costs can be summarized with the following

relationship:

Drill cost ($/ft) = [(Cown + Cop)BL(hr) +BC]/BL(ft)

Where; Cown is cost per hour to own the rig,

Cop is cost per hour to operate the rig

BL is bit life in hours or in feet and

BC is bit cost (Hartman, 1996).


The cost to own includes taxes, interest, insurance, amortization, and depreciation. The cost

to operate includes labour, fuel, and parts and supplies, such as tires and drill steel. Labour

costs, also applied to blasting costs, include base salary plus benefits. Benefits, which range

from 30 to 40% of base salary, include insurance, health care, pension, and vacation.

Incentive pay, as a percentage of base salary, is often provided when productivity increases

over a predetermined average. Productivity is measured as feet (meters) drilled for the drill

crew or loaded and shot per employee-shift for the blasting crew. Blasting costs comprise

explosives, boosters and primers, initiation systems, and other expendables. Labour costs

include the hours spent by the blasting crew to handle and transport explosives, load holes,

detonate the shot, and take inventory and prepare paperwork. The cost of bulk loading and

storage equipment is also included (Hartman, 1996).

Blasting costs are directly related to powder factor and the cost per pound of the main

explosive charge. Labour costs can represent 5 to 40% of the total blasting costs, while the

cost of expendable blasting accessories such as primers and initiators is generally less than

20% of total costs. A comparison of drilling and blasting costs for various mining methods is

shown in Table 2.3.

Table 2.3: Comparison of Drilling and Blasting Costs for Various Mining

Methods

Extraction Methods Cost ($/yd3)

Metal mines and quarries 0.08-2.00

Construction 0.03-0.35

Tunnelling 0.80-2.00

Long-hole stoping 0.50-1.20

Vertical crater retreat mining 0.45-0.90

Cut and fill stoping 0.60-1.00

Shrinkage stoping 0.50-0.75

Conversion factor: $1 /yd3 = $1.31 /m3.

Source: Aimone, 1979.


However, in a company what matters the most is the total cost for the whole operation

(mineral production). Drilling and blasting is just the part of the entire operation, and in cost

optimization process, the focus shouldn’t only be on drilling and blasting.

The total costs for an operation are significantly reduced by the use of smaller drill hole

patterns (spacing and burden). The costs for drilling and explosive do indeed increase in that

way, but the costs of loading, transportation, crushing and grinding of the mineral are

significantly smaller (Strelec, 2011).

This however must be optimized in order to realize the costs reduction in the operation (see

figure 2.3 below). The optimal fragmentation (Xo) is obtained at the minimum total cost.

Fig. 2.3: Optimal fragmentation on cost basis (Strelec, 2011)

From fig. 2.3 it can be seen that, if the cost of drilling and blasting is set to minimum, we get

coarser fragmentation, which then significantly increases the cost of loading, transport, post-

fragmentation and ore treatment.


To the large extent, owners of open mines and quarries are, out of ignorance, well disposed

towards savings on drilling and blasting. These savings, however, will often disappear with

increased costs of loading, transport, and subsequent reduction in the size of oversize blocks

by hydraulic hammer and the increased costs of ore treatment (Strelec, 2011).

Optimization of fragmentation is therefore very important, and as for this study this will be

done basing on drilling and blasting costs against loading and hauling costs. Ore treatment

costs caused by poor fragmentation (including ore loss and plant delays/stoppages) has been

very difficult to quantify in this case and will just be considered as the added advantage out

of improved blasting results (good fragmentation).


CHAPTER 3

METHODOLOGIES

The methods for data collection and analysis in this study were as shown below;

Direct measurements in field;

Spacing

Burden

Hole diameter

Hole depth

Charge length and/or Stemming length

Equipment used;

A tape measure

Field note book and pen

The data obtained were quantitative.

Direct observation in field;

Explosive utilization and charging techniques ( for estimation of powder

factor)

Delay set up

Tie up and initiation system

Fragmentation

Equipment used;

A digital camera (to take some photos for illustration)

Field note book and pen

Scientific calculator

Both quantitative and qualitative data were obtained.


Questionnaires on;

Explosives types

Explosives cost

Explosive properties (e.g. density, detonation velocity etc.)

Historical data for drilling and blasting parameters and their respective results

(in terms of fragmentation)

Material handling costs in each blast design that have been employed

Questions asked included;

What type of explosive and explosive accessories has been used

for blasting over past few years to present?

How much do these explosives cost?

What are densities and detonation velocity of these explosives?

May I get the historical data for blasting and associated

fragmentation over a past few years?

What has been the cost for material handling over past few years?

Most of the data obtained here were quantitative historical data for

prediction of optimal parameters

Logical presentation of data using various statistical tools like;

Tables and

Charts

Tools used included;

Microsoft office (word and/or excel)

Scientific calculator


Trend analysis on;

Design data Vs fragmentation

Costs Vs fragmentation

Tools;

Microsoft office (excel)

Graphs/models for each design parameter Vs fragmentation were

generated here.

Prediction of most probable;

Optimal design parameters and

Optimal costs from the models generated

Tools;

Microsoft office (excel)


CHAPTER 4

DATA COLLECTION AND ANALYSIS

4.1: Collection of Data

Data were collected from the site (Williamson Diamonds Limited) by using various methods

as indicated in chapter 4. Some of the data that were collected were historical data

concerning drilling and blasting and the associated fragmentation and costs for the year 2012.

Some raw data on drilling and blasting operations for the period of two weeks during the

month of February 2013 were also collected. Some of the important data collected were

represented in the appendices of this work.

4.2: Analysis of the Data

Most of the data used in this project were from just a part of WDL pit. The pit was divided

into 16 major blocks but just 7 blocks were in regular operation. Figure 4.2(a) shows the

WDL pit major blocks and their properties (yellow shaded blocks were the ones in regular

operation).


Fig. 4.2(a): WDL pit major blocks showing rock types and characteristics (grade and

density)

Drilling and blasting data (table 9.1(a), appendix 1) were analysed for drilling accuracy as

follows;

Drill-holes’ drilled depths were compared with the designed hole depth, and the deviations

obtained were plotted as shown in figure 4.2(b).


Fig. 4.2(b): Drilled holes deviations from the designed depths

The actual powder factor (PFa) for this blast (22-Feb-13) was obtained as;

PFa = (P100 + ANFO)/VT

Where; P100 is kgs of emulsion recorded by the loading truck (= 7895kg)

ANFO is kgs of ANFO taken from the explosives magazine (= 4470kg); table 9.1(b)

VT is the volume of material blasted (= 10937.5m3)

Thus; PFa = (7895+4470)/10937.5

PFa = 1.13kg/m3.................................................................................(4.2.1)

The designed/calculated powder factor (PFc) for this blast was obtained as;


PFc = (nP100*LCP100*ρLP100 + nANFO* LCANFO* ρLANFO)/VT

Where; nP100 & nANFO is number of holes charged with emulsion and ANFO respectively

LCP100 & LCANFO is charge length for emulsion and ANFO respectively and

ρLP100 & ρLANFO is loading density for emulsion and ANFO respectively

nP100 = 150, nANFO = 100, LCP100 = LCANFO = 5m, ρLP100 = 9.8kg/m, ρLANFO = 6.54kg/m (see

table 9.1(b) in appendix 1).

Thus; PFc = (150*5*9.8+100*5*6.54)/10937.5

PFc = 0.97kg/m3....................................................................................(4.2.2)

The same procedure were used for 27-Feb-13 data (table 9.1(c) in appendix 1) and provided;

PFa = 1.17kg/m3....................................................................................(4.2.3)

PFc = 1.09kg/m3....................................................................................(4.2.4)

The whole historical data range was used to estimate the relative costs per tonne for drilling

& explosives and material handling as follows;

Drilling cost (DC) = meters drilled (M) * Cost of drilling one meter (Cm)

Cm was 4.87 US$/m

Thus; DC = 4.87*M

Explosive cost (EC) = cost of main charge (MC) + cost of explosive accessories (AC)

But MC = (Amount (kg) of emulsion used (P100) * cost of one kg of emulsion (CP100)) +

(Amount (kg) of ANFO used * cost of one kg of ANFO (CANFO))

CP100 was 1.246 US$/kg

CANFO was 0.956 US$/kg

AC was 0.15MC (15% of cost of main charge)


Thus; EC = (1.246*P100 + 0.956*ANFO) + 0.15MC

Drilling and explosive cost (DEC) = DC + EC

Now, DEC per tonne (DECpt) can be obtained from;

DECpt = DEC/T

Where; T is tonnage of muck-pile.

NB: (computations were done for each blast; see table 9.2(a) in appendix 2)

Material handling cost per tonne (HCpt) for each blast was estimated by the company by

considering;

Fuel price

Labour’s hourly rate

Owning & operating cost for equipment and

The working condition (see table 9.2(a) in appendix 2).

The total cost per tonne (TCpt) = DECpt + HCpt (table 9.2(a))

To obtain the optimal fragmentation, the data in table 9.2(a) were used to plot the graphs for

associated costs (including total cost) on the same X-Y axes (fig. 4.2(c)); and at the minimum

total cost, the optimal fragmentation was obtained.


Fig. 4.2(c): Fragmentation optimization models

The historical data range were also analysed for optimal spacing, burden and powder factor.

The previous spacing and burden and their associated powder factor and fragmentation were

used in this case. For values of spacing and burden that were used several times, their

average powder factor and fragmentation were taken. The data in table 9.2(b) (in appendix 2)

were thus reduced to few data (see table 4.2(b)) which were used to plot the optimization

models.

Table 4.2: Extracted data for plotting blasting parameters models

Burden (m) Spacing (m) Powder Factor(kg/m3) % (-300mm)

3 3.5 0.31 60%

3 3 0.47 68%

2.5 2.5 0.82 80%

2 2.5 0.82 85%

2 2 1.45 95%


To generate the models, scatter plot for each variable versus fragmentation was plotted on the

X-Y axes. Trend lines (with forecast option) that fitted best to the data were added on scatter

points so that the optimal parameters could be predicted.

Optimal spacing was obtained by reading the value of spacing corresponding with the

optimal fragmentation value through the trend line model shown in figure 4.2(d).

Fig. 4.2(d): Spacing optimization model

The same procedure was done for burden and powder factor for prediction of the optimal

parameters as shown in figures 4.2(e) and 4.2(f).


Fig. 4.2(e): Burden optimization model

Fig. 4.2(f): Powder factor optimization model


CHAPTER 5

RESULTS AND DISCUSSIONS

5.1: Results

The following results were obtained after the analysis on the data collected;

Average deviation of holes drilled (on 22-Feb-13) was -0.01m (table 9.1(a))

The actual powder factor for this blast was 1.13kg/m3 (equation 4.2.1).

The designed/calculated powder factor was 0.97kg/m3 (equation 4.2.2).

For the 27-Feb-13 blast;

The actual powder factor was 1.17kg/m3 (equation 4.2.3)

The designed/calculated powder factor was 1.09kg/m3 (equation 4.2.4)

The optimal fragmentation was found to be 95% (-30mm) (from figure 4.2(b)).

This in turn gave the following results;

Spacing = 2m (figure 4.2(c))

Burden = 1.9m (figure 4.2(d))

Powder factor = 1.3kg/m3 (figure 4.2(e)).

5.2: Discussions

In both blasts performed, actual powder factor seemed to be higher than the

designed/calculated one. In terms of cost this meant more cost was incurred but it could have

positive effects in terms of fragmentation.

In most cases overcharging though can cause a number of problems, but in terms of

fragmentation of material it would normally have a positive effect. For this reason, the

fragmentation problem in parts of WDL pit could have not been caused by explosives usage

unless improper explosive (for example using ANFO in wet holes) was used.


The problem of overcharging here could be caused by bad charging practices. From direct

observation on charging operations, some holes were often loaded with explosive until they

flooded. This could be due to the fact that no stamping stick was used to ensure that

explosives were loaded only to the required height, leaving stemming length specified in the

blast design.

The optimization results were obtained from analysis on the data which to the large extent

reflects the current situation at WDL. As stressed out before (in the problem statement), the

mine doesn’t have a crushing system thus the only way of comminution is through blasting.

It is therefore very important that blasting results yields very well fragmented material so as

to enhance the smoothness of the following operations after blasting.

The predicted optimal fragmentation of 95% (-300mm) seems too high to achieve but it is

important that it is achieved, though the drilling and explosives costs would be considerably

high but as a result it would save a lot of cost in the after blasting operations.

By just comparing the material handling costs with the explosive costs, this fragmentation

was justified (it yields the minimum total cost), but the benefits of achieving such

fragmentation could be much higher than it can be thought. This is because for some reasons

including time limit and privacy of the company itself, some data could not be found, which

could put more weight to the obtained results.

Some of the data that couldn’t be collected were the costs for ore treatment including plant

stoppages (caused by oversize problems), and the amount of money lost due to reduced

recovery caused by throwing away the oversize kimberlitic material into waste dump. These

cost aspects would normally rise up with poor fragmentation.

However, if the stated fragmentation (95% -300mm) could be met, the oversize material

would be much minimized and so would be handling and ore treatment costs. Recovery

would be increased and so would be the revenue resulting from diamonds sales.

To achieve the optimal fragmentation of 95% (-300mm) the optimal spacing of 2m would be

required. The optimal burden would be 1.9m. The powder factor in this case would be about

1.3kg/m3.


Again these values (for spacing and burden) seems too small and the increased explosive

consumption can be reflected by the high powder factor obtained (1.3kg/m3). Various

literatures suggests values for burden and spacing which are relatively high than these values.

For WDL average rock density of 2.02kg/t and the hole diameter of 102mm used, the burden

should be at least 2.2m (Ash, 1963), 2.0m (Rajpot, 2009) or 3.3m (Austin Powder Company,

2002). The spacing should be 2.9m (Austin Powder Company, 2002) or 2.7m (Ash, 1963);

(see sections 2.2.2 and 2.2.3 for formulae). All these values are higher than the predicted

values. The powder factor should be less than 1 (Mishra, 2009); (see section 2.2.8).

But again the WDL case is a bit different; comminution is totally dependent upon blasting

which is not the case for the values given in literatures. It can therefore (until there is

introduction of crushing system in ore treatment), prove worth to apply these values for

production.

There are number of factors which could make it difficult to meet this fragmentation even if

the predicted optimal values of 2m spacing, 1.9m burden and 1.3kg/m3 powder factor were to

be used. Some of the factors could be;

Short drill holes; It is known from various literatures that, the drill-hole depth to

burden ratio (Hb/B) should be greater than 3 (section 2.2.6). For the burden of 1.9m

therefore, the hole depth should be at least 5.7m. Drilling holes with depth less than

5.7m could result to poor fragmentation.

Over/under charge the drill-holes; Charging operations at WDL were not done very

careful and over charging and undercharging normally occurred interchangeably.

They didn’t use the tamping stick to ensure that the required stemming length was

maintained and as a result, some holes were charged too little while others were

charged too much. This is known to cause problems like air-blasts, ground vibrations

and uneven fragmentation (section 2.2.5).

Improper use of explosive; The design might be perfect but if the explosive used is

not proper, fragmentation could be very poor and very high costs might result. WDL

pit consisted of wet and dry areas, which resulted to dry and wet holes. Wet holes

were to be charged by emulsion and dry holes were to be charged by ANFO.


Charging wet holes by ANFO (by any mistake) would only mean that the shot

wouldn’t fire; while on the other hand, charging dry holes by emulsion meant

increase in explosive cost (emulsion price per kg is higher than that of ANFO). This

actually happened on 16th February, 2012 at NW Block C (in blast which was done

with spacing and burden of 2.5m). See figure 5.2 below.

Fig. 5.2: Very poor fragmentation at NW Block C due to improper charging (16- Feb-12)


CHAPTER 6

CONCLUSIONS

The following were concluded for this study;

Optimal fragmentation was 95% (-300mm).

Optimal spacing was 2m.

Optimal burden was 1.9m.

Optimal powder factor was 1.3kg/m3.

For improved fragmentation the hole depth should be at least 5.7m.

Poor fragmentation was just for some parts of the pit (at a time, NW Block A, B and

D; SW Block C and D and sometimes Block B) so the predicted optimal parameters

should be applied there.

Other blocks which were not in operation but had similar rock properties like the

mentioned should also be dealt with by the predicted parameters if they were to be in

operation.


CHAPTER 7

RECOMMENDATIONS

Following the study done the following recommendations were made;

The study should be done to see the viability of installing a crushing system to the

treatment plant to improve recovery and reduce the drilling and blasting costs.

Before charging, holes need to be evaluated to observe any wet holes and charge

them with proper explosive (emulsion) to avoid failures in blasting.

All dry holes must be identified and be charged with ANFO to reduce the powder

factor (explosives cost) that could result if they are charged with emulsion.

If further study is to be done on this problem, it should include all cost aspects, and

there should be implementation of the parameters obtained in order to come to the

real optimal values.


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APPENDICES

Appendix 1: Field raw data

Table 9.1(a): Blasting raw data as at 22nd

February, 2013(Dh = 102mm)

Hole Depth Deviation

Spacing Deviation

Burden Deviation Desig

n Drilled Desig

n Drille

d Desig

n Drille

d

DH1 6 6.01 0.01

2.5 2.5 0 F

DH2 6 6 0

2.5 2.51 0.01

DH3 6 5.9 -0.1 I

2.5 2.5 0

DH4 6.3 6.1 -0.2

2.5 2.5 0 R

DH5 6.8 6.7 -0.1

2.5 2.5 0

DH6 7 6.95 -0.05 S

2.5 2.5 0

DH7 7 7 0

2.5 2.52 0.02 T

DH8 7 7 0

2.5 2.5 0

DH9 7 7.03 0.03

2.5 2.49 -0.01 R

DH10 6.7 6.71 0.01

2.5 2.48 -0.02

DH11 6.5 6.48 -0.02 O

2.5 2.5 0

DH12 6.4 6.39 -0.01

2.5 2.5 0 W

DH13 6.2 6.15 -0.05

2.5 2.51 0.01 2.5 2.5 0

DH14 6 6 0

2.5 2.49 -0.01 S

DH15 6 6 0

2.5 2.5 0 E


Table 9.1(a) cont….

DH16 6.5 6.49 -0.01

2.5 2.5 0 C

DH17 6.4 6.41 0.01

2.5 2.48 -0.02 O

DH18 6.9 6.88 -0.02

2.5 2.51 0.01 N

DH19 7 7 0

2.5 2.5 0 D

DH20 7 6.99 -0.01

2.5 2.5 0

DH21 7 7 0

2.5 2.5 0

DH22 6.9 6.89 -0.01 R

2.5 2.49 -0.01

DH23 6.7 6.7 0

2.5 2.49 -0.01 O

DH24 6.5 6.43 -0.07

2.5 2.5 0

DH25 6.4 6.41 0.01 W

2.5 2.5 0

DH26 6.3 6.3 0

2.5 2.51 0.01 2.5 2.5 0

DH27 6 5.89 -0.11

2.5 2.5 0 T

DH28 6 6 0

2.5 2.5 0

DH29 6 6.04 0.04 H

2.5 2.5 0

DH30 6.6 6.57 -0.03

2.5 2.49 -0.01 I

DH31 6.8 6.79 -0.01

2.5 2.505 0.005

DH32 7 6.98 -0.02 R

2.5 2.5 0

DH33 7 7 0

2.5 2.5 0 D

DH34 7 7 0



2.5 2.495 -0.01

DH35 7 7 0

2.5 2.5 0 R

DH36 7.1 7 -0.1

2.5 2.5 0

DH37 7 7 0 O

2.5 2.5 0

DH38 6.9 6.95 0.05

2.5 2.49 -0.01 W

DH39 6.8 6.77 -0.03

2.5 2.49 -0.01 2.5 2.5 0

DH40 6 6 0

2.5 2.5 0 F

DH41 6 6 0

2.5 2.51 0.01 O

DH42 6.4 6.38 -0.02

2.5 2.5 0 U

DH43 7 7 0

2.5 2.5 0 R

DH44 7 7 0

2.5 2.5 0 T

DH45 7 7 0

2.5 2.51 0.01 H

DH46 7 7.01 0.01

2.5 2.5 0

DH47 7 7 0

2.5 2.5 0

DH48 7 7.01 0.01 R

2.5 2.52 0.02

DH49 7 7.03 0.03

2.5 2.47 -0.03 O

DH50 7 7 0

2.5 2.5 0

DH51 7 7 0 W

2.5 2.49 -0.01

DH52 7 7 0

2.5 2.5 0 2.5 2.5 0



DH53 6.5 6.39 -0.11

2.5 2.5 0 F

DH54 6.5 6.49 -0.01

2.5 2.5 0

DH55 6.5 6.51 0.01 I

2.5 2.5 0

DH56 7 7.01 0.01

2.5 2.5 0 F

DH57 7 7 0

2.5 2.5 0

DH58 7 6.99 -0.01 T

2.5 2.5 0

DH59 7 6.98 -0.02

2.5 2.5 0 H

DH60 7 7.02 0.02

2.5 2.5 0

DH61 7.2 7.25 0.05

2.5 2.5 0 R

DH62 7.3 7.38 0.08

2.5 2.5 0

DH63 7.2 7.15 -0.05 O

2.5 2.51 0.01

DH64 7.1 7.09 -0.01

2.5 2.5 0 W

DH65 7 7 0

2.5 2.49 -0.01 2.5 2.49 -0.01

DH66 6.5 6.5 0

2.5 2.5 0 S

DH67 7 6.96 -0.04

2.5 2.48 -0.02

DH68 7 7 0 I

2.5 2.51 0.01

DH69 7 7.01 0.01

2.5 2.5 0 X

DH70 7 7 0

2.5 2.5 0

DH71 7 7 0 T



2.5 2.5 0

DH72 7 6.99 -0.01

2.5 2.5 0 H

DH73 7.4 7.38 -0.02

2.5 2.5 0

DH74 7.7 7.71 0.01

2.5 2.51 0.01 R

DH75 7.6 7.62 0.02

2.5 2.5 0

DH76 7.4 7.38 -0.02 O

2.5 2.5 0

DH77 7 7 0

2.5 2.5 0 W

DH78 7 6.99 -0.01

2.5 2.49 -0.01 2.5 2.5 0

DH79 7 7 0

2.5 2.5 0 S

DH80 7 7 0

2.5 2.5 0 E

DH81 7 7 0

2.5 2.5 0 V

DH82 7 7 0

2.5 2.5 0 E

DH83 7 6.97 -0.03

2.5 2.5 0 N

DH84 7 6.99 -0.01

2.5 2.5 0 T

DH85 7.2 7.18 -0.02

2.5 2.5 0 H

DH86 7.5 7.51 0.01

2.5 2.48 -0.02

DH87 7.8 7.78 -0.02

2.5 2.52 0.02 R

DH88 7.5 7.49 -0.01

2.5 2.5 0

DH89 7.5 7.48 -0.02 O

2.5 2.5 0



DH90 7.3 7.29 -0.01

2.5 2.5 0 W

DH91 7.1 7.08 -0.02

2.5 2.5 0 2.5 2.5 0

DH92 7 7 0

2.5 2.5 0 E

DH93 7 7 0

2.5 2.51 0.01 I

DH94 7 7.01 0.01

2.5 2.49 -0.01 G

DH95 7 7 0

2.5 2.5 0 H

DH96 7.3 7.28 -0.02

2.5 2.5 0 T

DH97 7.5 7.46 -0.04

2.5 2.5 0 H

DH98 7.7 7.69 -0.01

2.5 2.5 0

DH99 7.6 7.59 -0.01

2.5 2.5 0

DH100 7.6 7.6 0 R

2.5 2.5 0

DH101 7.5 7.46 -0.04

2.5 2.51 0.01 O

DH102 7.4 7.42 0.02

2.5 2.49 -0.01

DH103 7.2 7.18 -0.02 W

2.5 2.5 0

DH104 7 7 0

2.5 2.5 0 2.5 2.5 0

DH105 7 7 0

2.5 2.5 0 N

DH106 7 7.01 0.01

2.5 2.5 0

DH107 7.1 7 -0.1 I

2.5 2.5 0

DH108 7.2 7.21 0.01



2.5 2.5 0 N

DH109 7.3 7.29 -0.01

2.5 2.5 0

DH110 7.5 7.45 -0.05 T

2.5 2.5 0

DH111 7.6 7.59 -0.01

2.5 2.5 0 H

DH112 7.8 7.81 0.01

2.5 2.51 0.01

DH113 7.8 7.78 -0.02

2.5 2.49 -0.01 R

DH114 7.5 7.48 -0.02

2.5 2.48 -0.02

DH115 7.6 7.57 -0.03 O

2.5 2.49 -0.01

DH116 7.3 7.35 0.05

2.5 2.5 0 W

DH117 7 7 0

2.5 2.5 0 2.5 2.5 0

DH118 7 6.92 -0.08

2.5 2.51 0.01 T

DH119 7.2 7.18 -0.02

2.5 2.5 0

DH120 7.3 7.29 -0.01 E

2.5 2.5 0

DH121 7.4 7.37 -0.03

2.5 2.5 0 N

DH122 7.4 7.42 0.02

2.5 2.5 0

DH123 7.9 7.85 -0.05 T

2.5 2.5 0

DH124 7.8 7.83 0.03

2.5 2.5 0 H

DH125 7.7 7.69 -0.01

2.5 2.5 0

DH126 7.6 7.54 -0.06

2.5 2.5 0 R



DH127 7.8 7.77 -0.03

2.5 2.5 0

DH128 7.7 7.7 0 O

2.5 2.51 0.01

DH129 7.2 7.21 0.01

2.5 2.49 -0.01 W

DH130 7 7 0

2.5 2.48 -0.02 2.5 2.46 -0.04

DH131 7.5 7.48 -0.02

2.5 2.5 0 E

DH132 7.4 7.39 -0.01

2.5 2.51 0.01 L

DH133 7.5 7.48 -0.02

2.5 2.51 0.01 E

DH134 7.5 7.51 0.01

2.5 2.5 0 V

DH135 7.6 7.59 -0.01

2.5 2.5 0 E

DH136 7.8 7.84 0.04

2.5 2.5 0 N

DH137 7.7 7.72 0.02

2.5 2.5 0 T

DH138 7.8 7.79 -0.01

2.5 2.51 0.01 H

DH139 7.5 7.49 -0.01

2.5 2.5 0

DH140 7.8 7.8 0

2.5 2.5 0 R

DH141 7.6 7.6 0

2.5 2.5 0 O

DH142 7.5 7.6 0.1

2.5 2.5 0 W

DH143 7.4 7.34 -0.06

2.5 2.5 0 2.5 2.56 0.04

DH144 7.5 7.45 -0.05

2.5 2.52 0.02 T

DH145 7.5 7.5 0



2.5 2.5 0 W

DH146 7.6 7.5 -0.01

2.5 2.5 0 E

DH147 7.8 7.79 -0.01

2.5 2.5 0 L

DH148 7.8 7.78 -0.02

2.5 2.51 0.01 F

DH149 7.5 7.47 -0.03

2.5 2.5 0 T

DH150 7.5 7.53 0.03

2.5 2.5 0 H

DH151 7.5 7.52 0.02

2.5 2.49 -0.01

DH152 7.4 7.38 -0.02

2.5 2.5 0 R

DH153 7.3 7.24 -0.06

2.5 2.51 0.01

DH154 7.3 7.3 0 O

2.5 2.5 0

DH155 7.3 7.22 -0.08

2.5 2.5 0 W

DH156 7.6 7.58 -0.02

2.5 2.5 0 2.5 2.5 0

DH157 7.5 7.5 0

2.5 2.5 0 T

DH158 7.5 7.45 -0.05

2.5 2.5 0 H

DH159 7.6 7.58 -0.02

2.5 2.5 0 I

DH160 7.7 7.7 0

2.5 2.5 0 R

DH161 7.8 7.78 -0.02

2.5 2.5 0 T

DH162 7.6 7.5 -0.1

2.5 2.5 0 E

DH163 7.4 7.5 0.1

2.5 2.5 0 E



DH164 7.5 7.5 0

2.5 2.51 0.01 N

DH165 7.5 7.51 0.01

2.5 2.49 -0.01 T

DH166 7 7 0

2.5 2.5 0 H

DH167 7 6.99 -0.01

2.5 2.5 0 R

DH168 7 6.97 -0.03 O

2.5 2.5 0 W

DH169 7.2 7.19 -0.01

2.5 2.5 0 2.5 2.55 0.005

DH170 7.7 7.68 -0.02

2.5 2.5 0 F

DH171 7.8 7.82 0.02

2.5 2.5 0 O

DH172 7.9 7.9 0

2.5 2.5 0 U

DH173 7.8 7.79 -0.01

2.5 2.51 0.01 R

DH174 7.6 7.62 0.02

2.5 2.49 -0.01 T

DH175 7.4 7.42 0.02

2.5 2.5 0 E

DH176 7.4 7.35 -0.05

2.5 2.5 0 E

DH177 7.4 7.4 0

2.5 2.5 0 N

DH178 7.4 7.41 0.01

2.5 2.5 0 T

DH179 7 7.01 0.01

2.5 2.5 0 H

DH180 7 6.96 -0.04

2.5 2.5 0 R

DH181 7 7 0 O

2.5 2.5 0 W

DH182 7.1 7 -0.1



2.5 2.5 0 2.5 2.5 0

DH183 7.4 7.37 -0.03

2.5 2.48 -0.02 F

DH184 7.6 7.6 0

2.5 2.5 0 I

DH185 7.4 7.38 -0.02

2.5 2.5 0 F

DH186 7.6 7.55 -0.05

2.5 2.52 0.02 T

DH187 7.5 7.55 0.05

2.5 2.5 0 E

DH188 7.2 7.2 0

2.5 2.5 0 E

DH189 7.1 7.12 0.02

2.5 2.5 0 N

DH190 7.2 7.14 -0.06

2.5 2.5 0 T

DH191 7 7 0

2.5 2.5 0 H

DH192 7 7.01 0.01

2.5 2.5 0 R

DH193 6.7 6.6 -0.1

2.5 2.5 0 O

DH194 6.5 6.53 0.03

2.5 2.51 0.01 W

DH195 6.8 6.79 -0.01

2.5 2.48 -0.02 2.5 2.5 0

DH196 7.4 7.37 -0.03

2.5 2.5 0 S

DH197 7 7 0

2.5 2.5 0 I

DH198 7.4 7.3 -0.1

2.5 2.51 0.01 X

DH199 7.4 7.45 0.05

2.5 2.5 0 T

DH200 7.2 7.2 0

2.5 2.5 0 E



DH201 7 7.1 0.01

2.5 2.5 0 E

DH202 7 7 0

2.5 2.5 0 N

DH203 6.9 6.8 -0.1

2.5 2.5 0 T

DH204 7 7 0

2.5 2.5 0 H

DH205 6.7 6.69 -0.01

2.5 2.5 0 R

DH206 6.5 6.53 0.03 O

2.5 2.5 0 W

DH207 6.6 6.6 0

2.5 2.49 -0.01 2.5 2.5 0

DH208 7.3 7.31 0.01

2.5 2.5 0 S

DH209 6.9 7 0.1

2.5 2.5 0 E

DH210 7.1 7 -0.1

2.5 2.5 0 V

DH211 7.4 7.36 -0.04

2.5 2.5 0 E

DH212 7.1 7 -0.01

2.5 2.5 0 N

DH213 6.9 7 0.01

2.5 2.5 0 T

DH214 6.6 6.59 -0.01

2.5 2.49 -0.01 E

DH215 6.7 6.73 0.03

2.5 2.5 0 E

DH216 6.8 6.79 -0.01

2.5 2.5 0 N

DH217 6.6 6.6 0

2.5 2.5 0 T

DH218 6.4 6.43 0.03

2.5 2.51 0.01 H

DH219 6.2 6.21 0.01



2.5 2.5 0 ROW

DH220 6 6 0

2.5 2.5 0 2.5 2.495 -0.01

DH221 7 7 0

2.5 2.5 0 E

DH222 6.8 6.76 -0.04

2.5 2.5 0 I

DH223 7 7 0

2.5 2.5 0 G

DH224 6.7 6.76 0.06

2.5 2.5 0 H

DH225 6.8 6.77 -0.03

2.5 2.5 0 T

DH226 6.5 6.5 0

2.5 2.51 0.01 E

DH227 6.4 6.5 0.1

2.5 2.49 -0.01 E

DH228 6.5 6.45 -0.05

2.5 2.5 0 N

DH229 6.5 6.5 0

2.5 2.5 0 T

DH230 6 6 0

2.5 2.5 0 H

DH231 6 6.1 0.1

2.5 2.5 0 ROW

DH232 5.7 5.69 -0.01

2.5 2.5 0 2.5 2.51 0.01

DH233 7 7 0

2.5 2.5 0 N

DH234 6.8 6.8 0 I

2.5 2.5 0 N

DH235 6.7 6.7 0 E

2.5 2.5 0 T

DH236 6.6 6.57 -0.03 E

2.5 2.5 0 E

DH237 6.3 6.31 0.01 N

2.5 2.5 0 T



DH238 6.3 6.32 0.02 H

2.5 2.495 -0.01

DH239 6.2 6.17 -0.03

2.5 2.48 -0.02 R

DH240 6.2 6.2 0

2.5 2.52 0.02

DH241 6.2 6.21 0.01 O

2.5 2.51 0.01

DH242 6 6 0

2.5 2.5 0 W

DH243 6 6.01 0.01

2.5 2.5 0 2.5 2.5 0

DH244 6 6 0

2.5 2.5 0 T

DH245 6.1 6 -0.1 W

2.5 2.5 0 E

DH246 6.1 6.12 0.02 N

2.5 2.5 0 T

DH247 6 6 0 I

2.5 2.5 0 E

DH248 5.8 5.81 0.01 T

2.5 2.5 0 H

DH249 5.8 5.8 0

2.5 2.5 0 R

DH250 5.7 5.69 -0.01

2.5 2.5 0 O

DH251 5.5 5.5 0

2.5 2.5 0 W

DH252 5.3 5.25 -0.05

Total 1765 1762.7 -2.28

Average 7.00 6.99 -0.01


More data for 22-Feb-13 blast

Table 9.1(b): Explosive utilization for the 22-Feb-13 blast (NW Block C)

Explosive Actual

amount (kg)

Loading

density (kg/m)

Holes

charged

Average charge

length (m)

Emulsion (P100) 7895 9.80 150 5

ANFO 4470 6.54 100 5

Patterns drilled, tie up and blasting sequence

Fig. 9.1(a): Depicted timed, tied up pattern for the blast at NW Block C (on 22-Feb-13)


Fig. 9.1(b): Blasting sequence showing direction of the muck-pile (V-cut) at NW Block C on

22-Feb-13

Fig. 9.1(c): Good rock fragmentation with spacing and burden of 2.5m x 2.5m as at 22-Feb-

13 at NW Block C (CL-RVK)


Data for 27-Feb-13 Blast

Table 9.1(c): Explosive utilization for the 27-Feb-13 blast (SW Block C)

Explosive Actual amount

(kg)

Loading

density (kg/m)

Average charge

length (m)

Holes charged

Emulsion 14260 9.08 7.2 193

ANFO 5470 6.54 7.2 100

The drilling pattern for this blast was as shown in figure 9.1(d). The number of holes was 293

and the average depth was 9.2m.

Fig. 9.1(d): A photo showing the pattern for blasting at SW Block C on 27-Feb-13


Fig. 9.1(e): Poor rock fragmentation with spacing and burden of 2.5 x 2.5m as at 27-Feb-13

at SW Block C (NCL-RVK and GB)


Appendix 2: Extracted data from WDL archive (historical data) for various purposes

Table 9.2(a): The whole data range analysed for costs (Dh = 102mm)

BLASTING DATE

ESTIMATED % (-

300mm)

EXPLOSIVE(kg) BLASTING

COST(US$/t)

MATERIAL HANDLING

COST (US$/t)

TOTAL COST

(US$/t) EMULSION ANFO

27-Jan-12 70% 600 1000 0.41 4.99 5.40

02-Feb-12 95% 11550 0 1.55 1.08 2.63

06-Feb-12 60% 300 2375 0.37 6.80 7.17

07-Feb-12 85% 3272 2250 0.88 2.10 2.98

13-Feb-12 85% 5000 1850 0.96 1.99 2.95

16-Feb-12 80% 12350 0 1.05 2.36 3.41

01-Mar-12 82% 7331 0 1.00 2.01 3.01

02-Mar-12 65% 11646 1250 0.64 4.99 5.63

05-Mar-12 80% 7063 775 0.84 2.38 3.22

10-Mar-12 85% 9660 0 0.88 2.14 3.02

13-Mar-12 85% 4010 1425 0.80 2.48 3.28

19-Mar-12 80% 11329 3800 0.86 3.14 4.00

21-Mar-12 88% 5413 750 0.82 1.58 2.40

24-Mar-12 85% 6385 1250 0.80 2.17 2.97

27-Mar-12 85% 6621 250 0.91 1.23 2.14

02-Apr-12 75% 10415 5050 0.84 3.85 4.69

04-Apr-12 85% 5593 0 0.82 2.47 3.29

06-Apr-12 80% 5550 0 0.82 3.04 3.86

11-Apr-12 85% 13645 1000 0.88 2.79 3.67

17-Apr-12 75% 5000 3375 0.75 4.68 5.43

19-Apr-12 75% 8015 2040 0.85 3.78 4.63

24-Apr-12 85% 12220 0 0.83 2.58 3.41

30-Apr-12 75% 7185 1125 0.79 4.32 5.11

04-May-12 85% 13313 1750 0.84 2.18 3.02

08-May-12 80% 13114 625 0.83 2.45 3.28

14-May-12 85% 13360 3500 0.82 1.99 2.81

16-May-12 80% 4100 3725 0.80 2.24 3.04

24-May-12 85% 16096 8975 0.82 2.03 2.85

05-Jun-12 80% 15910 7100 0.86 2.57 3.43

12-Jun-12 70% 8010 16900 0.82 3.98 4.80

18-Jun-12 85% 3205 5850 0.85 2.18 3.03

23-Jun-12 90% 0 6750 0.77 1.11 1.88

26-Jun-12 75% 240 3500 0.80 3.89 4.69



28-Jun-12 80% 3000 3500 0.83 3.22 4.05

02-Jul-12 80% 5716 8750 0.84 3.58 4.42

04-Jul-12 80% 990 750 0.83 3.88 4.71

08-Jul-12 85% 8597 7475 0.89 5.02 5.91

06-Jul-12 70% 2229 3975 0.64 3.37 4.01

14-Jul-12 75% 620 2050 0.84 4.26 5.10

18-Jul-12 85% 8500 900 0.91 2.78 3.69

19-Jul-12 70% 2520 11475 0.75 4.64 5.39

21-Jul-12 85% 1500 3250 0.83 2.88 3.71

25-Jul-12 85% 5320 3000 0.84 2.97 3.81

27-Jul-12 70% 4300 1250 0.93 4.50 5.43

01-Aug-12 85% 10309 500 0.93 2.75 3.68

02-Aug-12 75% 15047 2250 0.93 3.42 4.35

07-Aug-12 80% 11034 3500 0.92 2.87 3.79

10-Aug-12 85% 6095 0 0.94 2.38 3.32

16-Aug-12 80% 4315 3825 0.89 3.28 4.17

20-Aug-12 70% 8464 3625 0.87 4.27 5.14

25-Aug-12 75% 9500 2750 0.93 4.18 5.11

28-Aug-12 75% 5190 11000 0.83 3.96 4.79

31-Aug-12 80% 13000 375 0.98 3.08 4.06

03-Sep-12 80% 6400 0 0.96 2.98 3.94

07-Sep-12 70% 11659 500 0.96 4.13 5.09

10-Sep-12 80% 4350 2500 0.91 2.99 3.90

14-Sep-12 80% 14900 4600 0.70 2.12 2.82

20-Sep-12 85% 8260 0 0.96 2.09 3.05

21-Sep-12 70% 14335 625 0.74 3.89 4.63

22-Sep-12 85% 8200 0 0.96 2.00 2.96

25-Sep-12 80% 17856 0 1.14 1.89 3.03

28-Sep-12 80% 19530 0 0.92 2.09 3.01

01-Oct-12 85% 11640 300 0.94 1.98 2.92

03-Oct-12 80% 0 6625 0.76 2.56 3.32

09-Oct-12 80% 27375 8127 1.00 1.56 2.56

12-Oct-12 85% 15159 2000 1.01 1.14 2.15

15-Oct-12 85% 8100 2500 0.91 1.38 2.29

19-Oct-12 85% 2620 2000 0.76 2.47 3.23

19-Oct-12 85% 3970 1500 0.89 2.26 3.15

22-Oct-12 80% 8000 3500 0.84 2.56 3.40

24-Oct-12 80% 1000 900 0.68 3.64 4.32



26-Oct-12 85% 6446 4375 0.78 2.63 3.41

02-Nov-12 80% 14006 0 0.97 2.74 3.71

03-Nov-12 75% 5644 3100 0.86 3.36 4.22

07-Nov-12 70% 6582 0 0.96 3.82 4.78

10-Nov-12 80% 6500 1000 0.90 2.28 3.18

12-Nov-12 80% 10008 4058 0.92 2.30 3.22

16-Nov-12 80% 18666 1000 0.95 2.01 2.96

19-Nov-12 80% 3240 0 1.13 1.86 2.99

22-Nov-12 75% 4860 1125 0.93 2.72 3.65

26-Nov-12 85% 10590 1850 0.94 1.45 2.39

30-Nov-12 80% 7900 1075 0.95 2.18 3.13

03-Dec-12 75% 13230 750 0.92 2.98 3.90

07-Dec-12 80% 7960 1275 0.90 2.14 3.04

10-Dec-12 85% 5878 0 0.97 1.99 2.96

12-Dec-12 80% 2200 0 1.08 1.78 2.86

14-Dec-12 82% 12450 0 0.91 1.15 2.06

17-Dec-12 75% 3580 0 0.79 3.81 4.60

22-Dec-12 84% 15787 2386 1.00 1.27 2.27

27-Dec-12 80% 1405 6250 0.81 2.26 3.07

Table 9.2(b): The whole data range analysed for specific parameters (Dh =

102mm)

BLASTING

DATE

AREA AV.

DEPTH

(m)

BUR

DEN

(m)

SPAC

ING

(m)

ROCK TYPE INSITU

GROUND

EST. %(-

300mm)

PF

(kg/m3)

27-Jan-12 NW-C 3.5 3 3 CL-RVK MASSIVE-

SOFT_ROCK

70% 0.35

02-Feb-12 NW-C 5.3 2 2 CL-RVK SOFT_ROCK 95% 1.45

06-Feb-12 SW-C 5.3 3 3.5 NCL-RVK MASSIVE-

SOFT_ROCK

60% 0.31

07-Feb-12 NW-C 7.0 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.86

13-Feb-12 NE-

D/W.C

3.9 2 2.5 NCL-RVK SOFT_ROCK 85% 0.82

16-Feb-12 NW-C 7.0 2.5 2.5 CL-RVK SOFT_ROCK 80% 1.01

01-Mar-12 NW-C 3.5 2.5 2.5 CL-RVK SOFT_ROCK 82% 0.95

02-Mar-12 SW-C 5.1 3 3 NCL-RVK SOFT_ROCK 65% 0.59

05-Mar-12 NW-C

& D

3.8 2.5 2.5 CL-RVK SOFT_ROCK 80% 0.73


13-Mar-12 NW-

C&D

3.8 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.71


Table 9.2(b) cont....

19-Mar-12 SW-C 7.0 2.5 2.5 CL-RVK SOFT_ROCK 80% 0.80




02-Apr-12 NE-

D/W.C

7.0 2.5 2.5 NCL-RVK MASSIVE,SOF

T_ROCK

75% 0.77

04-Apr-12 NW-C 2.7 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.69



17-Apr-12 SW-C 4.7 2.5 2.5 NCL-RVK SOFT-

HARD_ROCK

75% 0.66

19-Apr-12 NW-C 5.1 2.5 2.5 CL-RVK SOFT_ROCK-

HARD

75% 0.76


30-Apr-12 SW-C 5.5 2.5 2.5 NCL-RVK SOFT-

HARD_ROCK

75% 0.67

04-May-12 NW-C 5.5 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.73

08-May-12 SW-C 5.8 2.5 2.5 NCL-

RVK/BOUMA

SOFT_ROCK 80% 0.73

14-May-12 NW-C 5.8 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.73

16-May-12 SW-B 4.0 2.5 2.5 NCL-RVK SOFT_ROCK 80% 0.74

24-May-12 NW-C 5.6 2.5 2.5 CL-RVK 85% 0.74

05-Jun-12 NW-C 5.4 2.5 2.5 CL-RVK SOFT_ROCK 80% 0.78

12-Jun-12 SW-B 5.2 2.5 2.5 BOUMA SOFT_ROCK 70% 0.72


23-Jun-12 NW-C 6 2.5 2.5 CL-RVK SOFT_ROCK 90% 0.78

26-Jun-12 SW-C 3.9 2.5 2.5 NCL-RVK SOFT_ROCK 75% 0.83


02-Jul-12 SW-C 8.7 2.5 2.5 NCL-RVK SOFT_ROCK 80% 0.82

04-Jul-12 SW-B 2.4 2.5 2.5 CL-RVK SOFT_ROCK 80% 0.77

08-Jul-12 NW-C 4.4 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.88

06-Jul-12 SW-C 10 2.5 2.5 BVK HARD-

SOFT_ROCK

70% 0.71

14-Jul-12 SW-C 2.9 2.5 2.5 NCL-RVK SOFT_ROCK 75% 0.86


19-Jul-12 SW-C 10.1 2.5 2.5 BVK HARD-

SOFT_ROCK

70% 0.87



27-Jul-12 NE-

D/W.C

7.4 2.5 2.5 NCL-RVK SOFT-

HARD_ROCK

70% 0.87

01-Aug-12 NW-C 3.9 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.85



02-Aug-12 NW-

A,B

6.5 2.5 2.5 CL-RVK SOFT TO

HARD ROCK

75% 0.86

07-Aug-12 SW-C 7.7 2.5 2.5 NCL-RVK SOFT ROCK 80% 0.87

10-Aug-12 NW-C 3.8 2.5 2.5 CL-RVK SOFT ROCK 85% 0.85

16-Aug-12 NW-

C,D

4.6 2.5 2.5 CL-RVK SOFT ROCK 80% 0.87

20-Aug-12 NW-B 7.2 2.5 2.5 CL-RVK HARD ROCK 70% 0.81

25-Aug-12 NW-B 7.1 2.5 2.5 CL-RVK HARD-SOFT

ROCK

75% 0.87

28-Aug-12 SW-

C,D

6 2.5 2.5 NCL-RVK SOFT TO

HARD ROCK

75% 0.82

31-Aug-12 SW-C 7.3 2.5 2.5 NCL-RVK SOFT ROCK 80% 0.91

03-Sep-12 NW-

A,B

4.4 2.5 2.5 CL-RVK SOFT ROCK 80% 0.87

07-Sep-12 NW-B 8.0 2.5 2.5 CL-RVK HARDTO

SOFT ROCK

70% 0.88

10-Sep-12 NW-B 4.2 2.5 2.5 CL-RVK SOFT ROCK 80% 0.88

14-Sep-12 NW-D 4.3 2.5 2.5 BVK SOFT ROCK 80% 0.74

20-Sep-12 NW-D 7.2 2.5 2.5 CL-RVK SOFT ROCK 85% 0.88

21-Sep-12 SW-C 5.6 2.5 2.5 BVK SOFT TO

HARD ROCK

70% 0.76




01-Oct-12 NW-B 8.6 2.5 2.5 CL-RVK SOFT ROCK 85% 0.85

03-Oct-12 SW-C 4.7 2.5 2.5 NCL-RVK SOFT ROCK 80% 0.77

09-Oct-12 SW-C 9.8 2.5 2.5 NCL-RVK SOFT ROCK 80% 0.98



19-Oct-12 NW-C 3.3 2.5 2.5 CL-RVK SOFT ROCK 85% 0.65

19-Oct-12 NW-

C,D

2.1 2.5 2.5 CL-RVK SOFT ROCK 85% 0.83


24-Oct-12 NW-

C,D

4.3 2.5 2.5 CL-RVK SOFT ROCK 80% 0.54


02-Nov-12 NW-B 8.7 2.5 2.5 CL-RVK SOFT ROCK 80% 0.89

03-Nov-12 SW-C 6.7 2.5 2.5 NCL-RVK SOFT ROCK 75% 0.80

07-Nov-12 NW-

A,B

5.0 2.5 2.5 NCL-RVK SOFT ROCK 70% 0.87







22-Nov-12 NW-

A,D

3.4 2.5 2.5 CL-

RVK,BVK?

SOFT ROCK 75% 0.86

26-Nov-12 NW-C 5.2 2.5 2.5 CL-RVK SOFT ROCK 85% 0.88


03-Dec-12 NW-

D,A

5.0 2.5 2.5 NCL-RVK SOFT ROCK 75% 0.84

07-Dec-12 SW-C 4.9 2.5 2.5 NCL-RVK SOFT ROCK 80% 0.82

10-Dec-12 NW-C 9.0 2.5 2.5 CL-RVK SOFT ROCK 85% 0.89

12-Dec-12 NW-C 9.0 2.5 2.5 CL-RVK SOFT ROCK 80% 1.04

14-Dec-12 NW-

B,C

4.8 2.5 2.5 CL-RVK SOFT ROCK 82% 0.81

17-Dec-12 SW-C 3.5 2.5 2.5 NCL-

RVK/BVK

SOFT TO

HARD ROCK

75% 0.68

22-Dec-12 NW-B 5.4 2.5 2.5 CL-RVK SOFT ROCK 84% 0.97

27-Dec-12 SW-C 4.8 2.5 2.5 NCL-

RVK/BOUMA

SOFT ROCK 80% 0.83

Appendix 3: Some photos showing fragmentation at different blasts (Dh = 102mm).

Fig. 9.3(a): Poor fragmentation with burden and spacing 3m x 3m as at 27-Jan-2012 at NW

Block C (CL-RVK)


Fig. 9.3(b): Poor rock Fragmentation with burden and spacing 3m x 3.5m as at 6-Feb-2012

at SW Block C (GB/BVK)

Fig. 9.3(c): Good rock Fragmentation with Burden and Spacing of 2.5m x 2.5m.as at 16-

Feb-2012 at NW Block C (CL-RVK)

Documents

Kitaly Venance D. - Blast Design Optimization to Improve Material Fragmentation-complete report