this project was done in Williamson diamond mine

1. ASSESSMENT OF DRILLING AND BLASTING PARAMETERS TO
MINIMISE TOES PROBLEM
Case study at Williamson diamond mine
A Final Year Project Submitted in Partial Fulfilment of the
Requirements for the Degree of Bachelor of Science (Mining Engineering)
of the University of Dodoma
University of Dodoma
July, 2016

2. i
CERTIFICATION
The Undersigned certify that they have read and hereby recommend for examination by
the University of Dodoma, a final year project entitled: Assessment of drilling and
blasting parameters to minimise toes problems, in partial fulfilment of the
requirements for the Degree of Bachelor of Science (Mining engineering) of the
University of Dodoma
Supervisor…………………………………….
Eng P. R. Mwaria.
Date........................................................................

3. ii
DECLARATION
I, Marcelli Joseph C, declare that this final year project is my own original work and
that it has not been presented to any other University for a similar or any other degree
award.
Signature…………………………………………………………

4. iii
ACKNOWLEDGMENT
Firstly I would like to thank GOD for giving me strength and health toward the
compilation of this project within a given limited period of time.
I extend my deeply gratitude to my parents who were always providing me a great
support mentally but mostly financially in the way along to completion of this project.
Also I want to express my gratitude to my supervisor, Eng. Peter R Mwaria for his help
and guidance throughout the time that i have been doing this project.
Great appreciations goes to the department of mining at Williamson diamond limited to
give me golden opportunity to conduct my project at their company and with their
pleasure they provided with me everything necessary for my smooth complication of
this project.
Finally special thanks to my fellow classmates BSc. Mining Engineering for their
support in this project. Their comments and advice help me a lot towards compilation of
this project.

5. iv
ABBREVIATIONS
ABS- Absolute bulk strength
ANFO - Ammonium nitrate fuel oil
ANFO-Ammonium nitrate fuel oil
B - Burden
BSR- Burden Stiffness Ratio
BVK- Brecciate Volcanic Kimberlite
J- Sub drill
LD- Loading Density
mm - Millimetre
m -Meter
RC-Relative Confinement
RVK- Reworked Volcanic Kimberlite
S- Spacing
VED- Vertical Energy Distribution

6. v
ABSTRACT
Toes refers to the humps present on the floor after blasting have been done. This project
is done at Williamson diamond mine and it have the aim of analysing and suggesting the
possible means to minimise the toes formation so as to obtain the toe less blasting and
minimise drilling and blasting cost.
The datas were collected at Williamson diamond mine site located at Mwadui,
Shinyanga Region. Datas that were collected are burden, spacing, stemming height,
borehole depth and borehole diameter. Data analysis was done by computing vertical
energy distribution, relative confinement and burden stiffness ratio so as to find out the
cause of the toes and possible means of elimination.
After data analysis it was observed that the cause of the toes is due to poor design of
drilling and blasting parameter and poor consideration of relative confinement, burden
stiffness ratio and vertical energy distribution, then proposed parameter are burden
3.10m, spacing 3.56m, sub-drill 1.0m, borehole depth 10.31m stemming height 1.86m
and diameter of borehole 102mm.
It is concluded that unavoidable factors contributing to the formation of toes be non-
existent then a drill block with pre-calculated values of drilling and blasting parameters
and with holes being marked properly and drilled up to the requisite depths, the face
being cleared off all the muck from the previous blasts and without any existing toe can
produce a perfectly toe less blasting, it is recommended that regular review of drilling
and blasting parameters is compulsory and quality assurances and quality control of
holes has to be done before blasting

7. vi
CONTENTS
CERTIFICATION ..............................................................................................................i
DECLARATION .............................................................................................................. ii
ACKNOWLEDGMENT.................................................................................................. iii
ABBREVIATIONS ..........................................................................................................iv
ABSTRACT.......................................................................................................................v
CHAPTER ONE ................................................................................................................1
INTRODUCTION .............................................................................................................1
1.1 Background of the problem......................................................................................1
1.2 Williamson diamond mine (Mwadui mine) .............................................................2
1.2.1 Location of the study area..................................................................................2
1.2.2 Regional geology...............................................................................................3
1.3 PROBLEM STATEMENT ......................................................................................3
1.3.1 OBJECTIVES .......................................................................................................4
1.3.2 MAIN OBJECTIVE ..........................................................................................4
1.3.3 SPECIFIC OBJECTIVES..................................................................................4
1.3.4 JUSTIFICATION OF THE STUDY .................................................................4
1.3.5 PROJECT SCOPE.............................................................................................4
1.3.6 IMPORTANCE OF THE STUDY ....................................................................5
1.3.7 PROBLEM QUESTIONS .................................................................................5
CHAPTER TWO ...............................................................................................................6
LITERATURE REVIEW...................................................................................................6
2.1 Drilling and blasting.................................................................................................6
2.2 Rock drilling.............................................................................................................6

8. vii
2.2.1 Drilling pattern ......................................................................................................7
2.3 Rock Blasting ...........................................................................................................7
2.3.1 Explosives..........................................................................................................7
2.3.2 Types of explosive.............................................................................................8
2.3.3 Properties of explosives.....................................................................................9
2.4 Drilling and blasting design parameters...................................................................9
2.4.1 Drilling and blasting design basic principles...................................................10
2.4.2 Hole diameter (D) ............................................................................................11
2.4.3 Bench height (Hb) ............................................................................................12
2.4.4 Spacing (S).......................................................................................................13
2.4.5 Burden (B) .......................................................................................................13
2.4.6 Stemming (T)...................................................................................................14
2.4.7 Charge length (C).............................................................................................15
2.4.8 Sub-drilling (J).................................................................................................15
2.4.9 Blast hole inclination...........................................................................................17
2.5 Energy distribution.................................................................................................18
CHAPTER THREE..........................................................................................................20
METHODOLOGY...........................................................................................................20
3.0 Introduction ............................................................................................................20
3.1 Primary data ...........................................................................................................20
3.2 Secondary data .......................................................................................................20
CHAPTER FOUR............................................................................................................21
DATA ANALYSIS AND RESULT DISCUSSION .......................................................21
4.1 Data collection........................................................................................................21

9. viii
4.2 Data analysis...........................................................................................................22
4.3 Analysis of causes toe formation............................................................................24
4.3.1 Vertical energy distribution.................................................................................24
4.3.2 Burden stiffness ratio.......................................................................................25
4.3.3 Relative confinement ratio ..................................................................................26
4.3.3 Sub drilling..........................................................................................................28
4.4 Toe minimisation analysis......................................................................................28
4.5 Results discussion...................................................................................................32
CONCLUSION................................................................................................................33
RECOMMENDATION ...................................................................................................34
References........................................................................................................................35
APPENDICES .................................................................................................................37
Appendix 1: Drilling and blasting data ........................................................................37
Appendix 2: Rock density data ....................................................................................42
Appendix 3: Explosive properties data.........................................................................42
Appendix 4: Historical data..........................................................................................43
List of figures
Figure 1; Mwadui diamond mine.......................................................................................2
Figure 2; Tanzania map shows Mwadui diamond mine (Tanzania mining, 2016)............3
Figure 3; Drill rig (Asia infrastructure limited, 2016) .......................................................6
Figure 4: Drilling geometry .............................................................................................11
Figure 5; graph of hole inclination against sub drilling...................................................16
Figure 6: Poor energy distribution results into toes formations.......................................19
Figure 7: Graph number 1 shows borehole deviations.....................................................22

10. ix
Figure 8: Graph number 2 shows Burden deviation ........................................................23
Figure 9; Graph number 3 shows stemming deviation ....................................................23
Figure 10: Holes with no or very small stemming height................................................24
Figure 11: Graph number 4 shows VED against average toes height..............................25
Figure 12: Graph number 5 BSR against average toe height...........................................26
Figure 13: Graph number 6 RC against average toes height ...........................................27
Figure 14: Graph number 7 of actual against designed BSR, WED and CR..................27
Figure 15: Toes at WDL ready for secondary blasting....................................................28
Figure 16: Graph number 8 shows borehole depth and borehole diameter .....................29
Figure 17; Graph number 9 shows borehole depth against stemming height..................30
Figure 18: Graph number 10 shows burden and spacing.................................................31
List if tables
Table 1: Burden stiffness ratio.........................................................................................12
Table 2: Burden table.......................................................................................................14
Table 3: Sub drill table.....................................................................................................16
Table 4: Drilling and blasting data...................................................................................21

11. 1
CHAPTER ONE
INTRODUCTION
1.1Background of the problem
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 (engineers,
2010).
Drilling is the process of making holes on the rock for various purposes especially
blasting, the drilled holes are loaded or charged with explosives and tied up then
initiated electrically or non-electrically. Proper designed and organized blast provides
good fragmentation of the blasted material. (Ramulu, 2012)
Blasting is an essential part of the mining cycle. In virtually all forms of mining, rock is
broken by drilling and blasting the rock. Blasting technology is the process of fracturing
material by the use of a calculated amount of explosive so that a predetermined volume
of material is broken. Good blast design and execution are essential to successful mining
operations. (Rock blasting, 2015)
Toe refers to humps present on the floor after blasting have been done, it was observed
that the toe formation has always been a drawback in the opencast mines. There are
certain factors that result in toe formation like the burden and spacing, size of drill block,
condition of drill holes and condition of face before blasting; charging of blast holes and
the type of initiation are the factors that can be avoided. But the strata variation,
fractured strata and watery holes are unavoidable. So it should be tried to achieve a drill
block where the unavoidable factors are non-existent. It is marked with crest, burden,
spacing. They were of the view that blast holes must be charged as per proper charging
pattern with appropriate percentage of booster, base and column and holes by charging
from bottom initiation leads to toe-less blasting. (Reddy, 1999). Due to the problem of

12. 2
toe occurrence Williamson diamond mine minimize these toes by using dozer and
secondary blasting method.
1.2 Williamson diamond mine (Mwadui mine)
Williamson diamond mine is the one of the huge and oldest currently operating mining
in Tanzania, the mine owned by Petra Diamond(75%) and Government of Tanzania
(25%), the company own the diamond open pit mine at Mwadui, and the mineral present
in the pit which are diamond gemstone, The mine has been operational for 70 years. It
still possesses significant diamond resources yet to be mined. The current mine plan for
Williamson is for 18 years. The potential life of the mine is more than 50 years.
Figure 1; Mwadui diamond mine
1.2.1 Location of the study area
Williamson diamond mine pit is located in Mwadui area found at Kishapu district in
Shinyanga region located thirty kilometer from the Shinyanga town and three kilometer
from Mwanza- Dar es Salaam tarmac road. The mine covers about 146 hecteres,
currently the mine have developed up to 90m deep from the surface and its estimate to
go up to 350m deep from the surface.

13. 3
Figure 2; Tanzania map shows Mwadui diamond mine (Tanzania mining, 2016)
1.2.2 Regional geology
The large portions of central and North Western Tanzania are covered by what has been
termed the Tanzanian craton. The craton is covered by Archean rocks, known as the
Dodoman System in central Tanzania, the Nyanzian System and Kavirondian System in
the northern part of the country. The Archean rocks have been intruded by kimberlites of
various ages, some of them were diamond bearing such as the famous (Williamson
kimberlite) and a number of the kimberlites in the Mabuki area and some in the Tabora
region.
1.3 PROBLEM STATEMENT
Explosive rock breakage or blasting is the mostly used method to break rocks in mining
and quarries, rock blasting becomes the cheapest method of breaking rocks than
mechanical methods, this method is used where there is no other means of breaking
rocks such as rock breakers, blasting is also used as means of increasing production in
the mining sector (Nicholson, 2005). Poor design or implementation of the designed
drilling and blasting parameter leads to the extra cost to the company by giving out poor

14. 4
shape of the ground after blasting due to the presence of toes this problem is mainly
observed in Williamson diamond mine (Mwadui). The present toes gives cost to the
company since it require the use of dozer or secondary blasting to remove them during
material handling, floor control and preparation for next blasting, also they result into
low volume of material obtained compared to the calculated volume by surveyors which
also lead to the poor estimation of the production.
1.3.1 OBJECTIVES
1.3.2 MAIN OBJECTIVE
To determine the drilling and blasting parameters which will minimize toes occurred
after blasting and irregular pit floor.
1.3.3 SPECIFIC OBJECTIVES
I) To investigate drilling and blasting parameters which are diameter of the
borehole, Hole depth or Bench height, inclination of the blast holes, Spacing,
Burden, Sub-drill or sub-grade length and burden stiffness ratio that results
into presence of toes and irregular pit floor after blasting.
II) To investigate the rock properties and type of explosive used.
III) To analyse the formation of toe and irregular pit floor
1.3.4 JUSTIFICATION OF THE STUDY
The aim of this project is to investigate the cause of toes occurred after blasting that
result into irregular pit floor or benches, this study will give out the cause and the proper
solution of the problem that can be applied to minimize the problem for all mine with
similar geological characteristics.
1.3.5 PROJECT SCOPE
This project concentrated on important parameters that are required for minimizing toes
and obtain regular pit floor. Such parameters included hole diameter, burden, spacing,

15. 5
hole depth, sub drill, bench height to burden ratio. Other geological factors such as rock
fracture, strata variation, fractured strata and watery holes are not included in this
project.
1.3.6 IMPORTANCE OF THE STUDY
The main importance of this project is to discover the main cause of toe during blasting
and indicate the best troubleshooting parameters which could be used to minimize this
problem and obtain irregular floor, the conclusion drawn from this work should be
employed to minimize toe in any mine with similar geological characteristics to that of
Mwadui mine, By doing so the cost will be reduced and minimize dozer operations and
secondary blast.
1.3.7 PROBLEM QUESTIONS
I. What are the causes of toes after blasting?
II. How to minimize toes obtained so as to reduce the use of dozer during floor
control?
III. What are the best drilling and blasting parameters to be used to minimize this
problem?

16. 6
CHAPTER TWO
LITERATURE REVIEW
2.1Drilling and blasting
Drilling and blasting are the major unit operations in opencast mining. In spite of the
best efforts to introduce mechanization in the opencast mines, blasting continue to
dominate the production. Therefore to cut down the cost of production optimal
fragmentation from properly designed blasting pattern has to be achieved. (Parida, 2007)
Rock fragmentation refers to the process of breaking the huge rock into small particles
or fragments which are simple to carry and easy to crush this can be done by using two
methods that is by using explosives or mechanical methods, rock fragmentation involve
two main activities that is drilling and blasting.
2.2 Rock drilling
Drilling is the use of drill rig to create holes for exploration or for loading with
explosives for blasting or ventilation purpose. Drill rig is the machine applies rotation,
percussion (hammering), or a combination of both to make holes.
Figure 3; Drill rig (Asia infrastructure limited, 2016)

17. 7
2.2.1 Drilling pattern
Drilling patterns vary greatly and depend upon the blast hole diameter, explosives
properties, rock properties, the degree of fragmentation and the displacement required
and the height of the face. Borehole patterns are drilled square (S/B = 1) or
rectangular (S/B greater than 1) on centre or offset (staggered).
Types of firing pattern is of mainly five types:
 Rectangular pattern
 Triangular pattern
 V pattern
 Zigzag pattern
 Staggered pattern
(Singh, 2005)
2.3 Rock Blasting
Rock blasting is the controlled use of explosives and other methods such as gas pressure
blasting pyrotechnics or plasma processes, to excavate, break down or remove rock. It is
practiced most often in mining, quarrying and civil engineering such as dam or road
construction. Except in mining, the result of rock blasting is often known as a rock cut.
Rock blasting currently utilizes many different varieties of explosives with different
compositions and performance properties. Higher velocity explosives are used for
relatively hard rock in order to shatter and break the rock, while low velocity explosives
are used in soft rocks to generate more gas pressure and a greater heaving effect. (Rock
blasting, 2016)
2.3.1 Explosives
An explosive, or blasting agent, is a compound or a mixture of compounds, which, when
initiated by heat, impact, friction, or shock, is capable of undergoing a rapid

18. 8
decomposition, releasing tremendous amounts of heat and gas. The decomposition is a
self-propagating, exothermic reaction called an explosion. The stable end products are
gases that are compressed, under elevated temperature, to very high pressure. (Hartman,
HowaldScott G. Britton,Donald W. Gentry,W. Joseph Schlitt,Michael, 1996)
2.3.2 Types of explosive
Many commercial or industrial explosives are classified as High explosive because they
contain critical amounts of military explosives or nitro-glycerine, and usually they are
cap sensitive. Others, such as dry blasting agents, are not classified as high explosive,
and require boosters or primers of high explosive for initiation. (Akhavan, 2005)
2.3.2.1 Ammonium Nitrate and Fuel Oil
It is a mixture of dry porous prilled ammonium nitrate and fuel oil, at the ratio of
94.3/5.7. The performance of this explosive depend on sensitivity prill properties. It does
not detonate ideally and its performance properties depend upon charge diameter and
confinement. For dry hole condition it is excellent since its density is less than that of
water, and also it should be initiated as soon as it is loaded.
2.3.2.2 Emulsions
This is a two-liquid phase containing microscopic droplets of aqueous nitrates of salts
mainly ammonium nitrate distribute widely in fuel oil, wax, or paraffin using an
emulsifying agent. The watering-oil structure depends on entrapped air or microspheres
for sensitivity, thereby eliminating the need for expensive explosive compounds.
Densities range from 1.15 to 1.45 this make emulsions to have excellent water resistant
properties, and they remain stable at low temperatures.

19. 9
2.3.3 Properties of explosives
The following are properties of explosive or explosive selection criteria;
2.3.3.1 Water Resistance
This explain the ability of an explosive to withstand exposure to water for long periods
of time without loss of strength or ability to detonate defines the water resistance. The
presence of moisture in amounts greater than 5% dissolves chemical components in dry
blasting agents and alters the composition of gases produced, contributing to the
formation of noxious fumes and lower energy output. (Adhikari G.R. and Venkatesh
H.S, 1999)
2.3.3.2 Density of explosive
The density of an explosive is defined as the weight per unit volume or the specific
gravity. Commercial explosives range in density from 0.5 to 1.7. Explosives with a
density less than 1 will float in water. Therefore, in water filled holes, an explosive with
a density greater than 1 is required. Density is most useful in determining the loading
density (LD) or the weight of explosives one can load per unit length of borehole (in
kilogram per meter). Note that knowledge of loading density is required for blast-design
calculations, and is calculated in English units as
L.D= 0.3405 βD2
Where; β is density
D is explosive column diameter in inches
2.4 Drilling and blasting design parameters
Preliminary blast design parameters are based on rock mass-explosive-geometry
combinations, which are later adjusted on the basis of field feedback using that design.
The primary requisites for any blasting round are that it ensures optimum results for
existing operating conditions, possesses adequate flexibility, and is relatively simple to

20. 10
employ. It is important that the relative arrangement of blast holes within a round be
properly balanced to take advantage of the energy released by the explosives and the
specific properties of the materials being blasted. There are also environmental and
operational factors peculiar to each mine that will limit the choice of blasting patterns.
The design of any blasting plan depends on the two types of variables; uncontrollable
variables or factors such as geology, rock characteristics, regulations or specifications as
well as the distance to the nearest structures, and controllable variables or factors. The
blast design must provide adequate fragmentation, to ensure that loading, haulage, and
subsequent disposal or processing is accomplished at the lowest cost. Further to the cost,
the design of any blast must encompass the fundamental concepts of an ideal blast
design and have the flexibility to be modified when necessary to account for local
geologic conditions. (Biran, 1994)
2.4.1 Drilling and blasting design basic principles
In designing a blast, three principles should be kept in mine
 Explosive force functions best when the rock being blasted has a free face toward
which it can break
 There must be an adequate void or open space into which the broken rock can
move and expand(swell)
 To properly utilize the energy available, the explosive product should be well-
confined within the rock.
If a blast is lacking in one or more of these three principles, the results will generally be
less than desired.
Some years ago, the late Richard Ash gathered data from a large number of blasts and
develop empirical formulas from that data to show the average relationship between hole
diameter, burden, spacing hole length, sun-drilling and stemming height. These
relationships were later published in 1972 by the Bureau of Mines in information
circular. Some of the very important parameters to be addressed in drilling and blasting
design include;

21. 11
 Hole diameter (D)
 Hole depth or Bench height(Hb)
 Spacing (S)
 Burden (B)
 Stemming (T)
 Charge length (L)
 Sub-drill or sub-grade length (J)
(Bendel, 1999)
Figure 4: Drilling geometry
2.4.2 Hole diameter (D)
The hole diameter is selected such that in combination with appropriate positioning of
the holes, will give proper fragmentation suitable for loading, transportation equipment
and crusher used. Additional factor that should be considered in the determination of the
hole diameter are Bench height. Hole diameter varies from 35 in small benches up to
440 mm in large benches. Langefors and Kihlstrom suggested that the diameter be kept
between 0.5 to 1.25 percent of the bench height.

22. 12
Diameter of the blast hole also play an important role in controlling toe formation in the
sense that the diameter of the hole should match with the geometry of the blast design
that is bench height, burden and spacing.
The rule of thumb for calculating the hole diameter is given by the following formula
Blast hole diameter in mm ≤ 15 x Bench height (BH) in metres
2.4.3 Bench height (Hb)
Usually the working specifications of loading equipment determine the height of the
bench. The bench height limits the size of the charge diameter and the burden. (Ash,
1968), states that 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 or back break around holes
and toe problems. With Hb/ B = 2, these problems are attenuated and are completely
eliminated when Hb/ B >3. The condition Hb / B >3, is usually found in quarries and coal
strip mining operations. In metal mining the bench height is conditioned by the reach of
the loading machine and the dilution of the mineral as well. When Hbis 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. 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).
Table 1: Burden stiffness ratio
Burden stiffness ratio = Hb /B 2 to 3.5 good fragmentation

23. 13
2.4.4 Spacing (S)
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, C. L., Jimeno, E. L. and
Francisco, J. A. C, 1995)
Rule of thumb formula for calculating spacing is given by the following formula
Spacing (S) = 1.15 x B (This gives an equilateral pattern)
Spacing (S) = 1 to 2 times the burden. (Nobel, 2010)
The value of the spacing to burden ratio (S: B) which has been commonly used in
different formulas lies between 1 and 2. From the production scale test with the spherical
charges breaking to crater geometry, many workers suggested that the spacing be kept
about 1.3 times the burden. When this ratio increases more than 2, unexpected results
were found. (Mishra, 2009). Design parameters Borehole patterns are drilled square (S/B
= 1) or rectangular (S/B = 1) on centre or offset (staggered) (S>B).
2.4.5 Burden (B)
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.
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
> 3.5 very good fragmentation

24. 14
properties of the rock mass (Rajpot, 2009). For the purpose of toe minimization at the
bench the average burden can be calculated from the following formula
Average burden =
𝑐𝑟𝑒𝑠𝑡 𝑏𝑒𝑛𝑐ℎ+𝑡𝑜𝑒 𝑏𝑒𝑛𝑐ℎ
2
Also the rule of thumb to calculate burden can be expressed as follows
Burden (B) = (25 to 40) x (D)
Table 2: burden table
Burden B = KBD Using ANFO
K B = 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
2.4.6 Stemming (T)
The primary function of the stemming is to confine the gas produced by the explosive
until they have adequate time to fracture and move the ground. A suitable stemming
column of suitable length and consistency enhances fracture and displacement by gas
energy. The enough energy at the bottom of the borehole will minimize the effect of toe
occurrence. When the burden has a high frequency of natural crack and planes of
weakness relatively long stemming column can be used. When the rock is hare and
massive the stemming should be shortest which will prevent excessive noise, air blast
and back brake. (Mishra, 2009). The rule of thumb used to compute the stemming length
and stemming material dimensions respectively

25. 15
Stemming (T) ≥ 20 x D or (0.7 - 1.2) x B
Stemming material size = D/10 to D/20 (Angular material with minimum fines)
Increase the multiplier if drill cuttings are used for stemming holes are wet, Decrease the
multiplier if stone chips are for stemming and/ or holes are dry. (Bendel, 1999)
2.4.7 Charge length (C)
Charge length this refers to the length of the explosive from bottom of the hole up the
point where stemming are filled to the bore hole, this length is always the different
between the sum of bench height and sub drilling subtracting the stemming height.This
is the explosive column in a blast hole and should be at least 20D in order to utilize fully
the explosion-generated strain in the rock. The rule of thumb of charge length is given
by the following formula.
Charge length (C) ≥ 20 D
2.4.8 Sub-drilling (J)
The amount of hole that is drilled below the intended floor of the excavation. Except in
those situations where the rock is horizontal bedding planes, the detonating charge will
usually leave a crater at the bottom of the hole rather than shearing the rock on the
horizontal plane. 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.
 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.
(Sethi N.N. & Dey N.C.A, 2004)

26. 16
If the toe formation will not avoid it may increase the operating costs for loading,
hauling equipment. The optimum effective sub drilling depends on density of the rock,
effective burden, type of explosive, blasthole diameter and inclination, the structural
formation, location of initiators in the charge.
The breakage in the bottom of blast hole is produced in the shape of inverted cones,
whose angle to the horizontal depend upon structure of the rock mass and on the residual
stresses. Normally they vary between 10 to 30 degrees
Table 3: Sub drill table
Different rock formation J/B
Open bedding plane at toe
Horizontal stratification 0
Easy toe. Soft rock 0.1-0.2
Normal. Medium hard rock 0.3
Difficult toe. Hard rock 0.4-0.5
Figure 5; graph of hole inclination against sub drilling

27. 17
The value of sub-drilling that produces the intersection of the cone shaped surface at
bench level is usually around J=0.3B because it has been shown that;
S=1 to 1.4B also
J= tan α ×(
S
2
)
With α taking on the indicating values.
The normal ratio of J/B for the bench blast are shown on the table.
In order to reduce sub-drilling the use of explosives which gives high concentration of
energy per unit of length in the bottom part of the charge and the drilling inclined blast
holes is recommended.
In horizontal bedding plane coal mining operations, in order to eliminate the crushing
effect of the ends of the charges, sub-drilling takes on a negative value as the bottom of
the blast hole is backfilled to a length of approximately 4D.
2.4.9 Blast hole inclination
The benefits of inclined drilling are better fragmentation, displacement and swelling of
the muck pile, less sub-drilling and better use of the explosive energy, lower vibration
levels and less risk of toe appearance.
The disadvantages of inclined holes are the following:
 Increased drilling length and deviation when drilling long blast hole.
 More wear on the bits, drill steel and stabilizers.
 Less mechanical availability of the drilling rig.
 Poor flushing of drill cuttings due to friction forces, requiring an increase in air
flow.
There are few management factors which are disadvantageous with the inclined holes
and are as follows:
 Difficulty in positioning of the drills.

28. 18
 Necessity of close supervision which creates work lapses.
 Lower drill feed, which means that in hard rock the penetration rate is limited in
direct proportion to the angle of inclination of the mast.
 Less productivity with rope shovels due to lower height of the muck pile.
 Problems in charging the explosive, especially in blast holes with water. (Pal,
U.K. and Ghosh, N, 2002)
In recent year attention has been given by open pit operators to the drilling of blast holes
up to 20 degree vertical. The benefits from inclined charges are Reduction of collar and
toe region less sub drilling requirement Uniformity of burden throughout the length of
blast hole Drilling of next bench is easier. Air blast and fire rock may occur more easily
due to smaller volume of material surrounding the collar inclined hole are successively
used in Europe where high benches and smaller diameter holes in medium to higher
strength rock exist. In case the face is high the use of vertical blast holes produce a
considerable variation in burden between the top and bottom face which is the basic
cause in the formation of toe. Angle greater than 25 degree are less used because of
difficulty in maintaining blast hole alignment excessive bit wear and difficulty in
charging blast holes. The blast hole length L increases with inclination. To calculate L,
the following equation is used:
Where, β in degrees represents the angle with respect to the vertical. (Mishra, 2009)
2.5 Energy distribution
Energy confinement play a huge role in toe formation, this explain the energy
distribution on the bore hole when energy is very poor on the bottom of the bore hole
toes will occur and when the energy is properly distributed toes will not occur.

29. 19
Figure 6:Poor energy distribution results into toes formations
On figure 6 shows toes occurrence due to poor energy distribution inside the boreholes,
also the borehole with sub drills have high energy distribution than area borehole
without sub drills.
Energy distribution is computed as relative confinement and vertical distribution, when
relative confinement is less than 1.4 then it means poor energy distribution and result
into toes that is irregular floor, fly rocks and stemming ejection, when relative
confinement is greater than 1.4 it means good energy distribution.
Energy distribution also is determined by considering vertical energy distribution that is
charge length divide by bench height this should be greater than 80% for good
fragmentation and toeless basting.
Relative confinement =
(𝑠𝑡𝑒𝑚𝑚𝑖𝑛𝑔 𝑙𝑒𝑛𝑔𝑡ℎ ×210000)+(𝑐ℎ𝑎𝑟𝑔𝑒 𝑑𝑖𝑎𝑚𝑒𝑡𝑒𝑟 ×600)
𝑐ℎ𝑎𝑟𝑔𝑒 𝑑𝑖𝑎𝑚𝑒𝑡𝑒𝑟 ×𝑐ℎ𝑎𝑟𝑔𝑒 𝑒𝑛𝑒𝑟𝑔𝑦 𝐴𝐵𝑆
Where; ABS explosive = AWS explosive× DE
DE is diameter of borehole, ABS is absolute bulk strength and AWS is absolute weight
strength.
Also to avoid poor energy distribution the inclined bore hole are the most effectively
method of minimize the energy loss so as to minimize toe formation.
Boreholes
Toes

30. 20
CHAPTER THREE
METHODOLOGY
3.0 Introduction
In this project the methods used to collect data can be categorized into two parts which
are direct observation from the field and direct measurements from the field also
computation was done on the field, by using this methods both qualitative data and
quantitative data are obtained
3.1 Primary data
To measure the following drilling and blasting parameters and quantitative data are
obtained;
 Diameter of the borehole
 Spacing
 Burden
 Bench height
 Subgrade length
 Inclination of the blast hole
To obtain the designed data from drilling and blasting department at Williamson
diamond limited and literature review.
3.2 Secondary data
Qualitative data are those data which include direct observation from the field and some
simple questions to the blasting department, the data include.
 To investigate the toe resulted after blasting
 To interview workers about type of rock and its geological structure
 To interview workers about type of explosive and its density
Finally is to analyses the data obtain so as to achieve the objective of the project.

31. 21
CHAPTER FOUR
DATA ANALYSIS AND RESULT DISCUSSION
4.1 Data collection
In this project data were collected at Williamson diamond mine at block D north east
and block A south east, the method used in data collection are indicated on chapter four,
and data obtained are the actual drilling and blasting parameters, rock and explosive
properties and historical data of 2016 from January to February, drilling and blasting
data recorded at the pit on 31st
march, 2016 are indicated on appendix one.
The following table shows the drilling and blasting data, that is historical data and actual
measured data from the field.
Table 4: Drilling and blasting data
Date Burden
(m)
Spacing
(m)
Borehole
depth (m)
Stemming
(m)
Material type Average toe
height(m)
5/1/2016 3.0 3.0 3.5 1.0 RVK 0.3
7/1/2016 2.2 2.5 5.1 1.5 RVK 0.6
19/1/2016 2.0 2.0 5.3 1.5 RVK 0.7
22/1/2016 2.5 2.5 10.0 2.5 BVK 0.1
29/1/2016 3.5 4.0 6.0 1.5 MUDSTONE 0
4/2/2016 2.5 2.5 10.1 2.5 BVK 0.3
9/2/2016 3.5 4.0 6.0 1.5 MUDSTONE 0
16/1/2016 3.0 3.0 3.9 1.0 BVK
0.8
25/2/2016 2.5 2.5 9.8 1.5 RVK
0.1
3/3/2016 2.0 2.0 4.9 1.5 BVK
0.5
23/3/2016 2.5 2.5 3.8 1.0 RVK
0.6
31/3/2016 3.5 4.0 11.2 2.5 SHALE
0.2

32. 22
4.2 Data analysis
Drilling and blasting data measured from the field were analysed to check the accuracy
of the designed data against actual data by obtaining the deviations. The deviation was
computed from the following formula
Deviation= actual – designed
On referring to appendix one, hole depth deviation can be Cleary shown on the graph
number 1 below.
Figure 7: Graph number 1 shows borehole deviations
The graph number 1 shows slightly deviation of actual depth from the designed depth of
maximum +1.3m on only one point and minimum of -2.4m on only one point the
average deviation of the bore holes depth is - 0.01687 which is equal to -1.687%.
The deviation of actual burden from the designed burden was obtain to check the
accuracy of the actual burden as shown on the graph number 1 by referring appendix
one.
-3
-2.5
-2
-1.5
-1
-0.5
0
0.5
1
1.5
B1
B5
B9
B13
B17
B21
B25
B29
B33
B37
B41
B45
B49
B53
B57
B61
B65
B69
B73
B77
B81
B85
B89
B93
B97
B101
B105
B109
B113
B117
B121
B125
Deviationfromrequredborehole
depth(m)
Borehole number
Graph shows dole depth deviation

33. 23
Figure 8: Graph number 2 shows Burden deviation
The graph number 2 shows slightly deviation of actual burden from the designed burden
of maximum +0.4m on only two points and minimum of -0.3m on only three points the
average deviation of the bore holes burden is 0.517428%. The deviation of spacing to
burden ratio is computed to be 1.13 while the designed value is 1.15 which deviate from
actual value by -0.019.
The deviation of actual stemming from the designed stemming was obtain to check the
accuracy of the actual stemming as shown on the graph number 3 by referring appendix
two.
Figure 9; Graph number 3 shows stemming deviation
-0.4
-0.3
-0.2
-0.1
0
0.1
0.2
0.3
0.4
0.5
B1
B6
B11
B16
B21
B26
B31
B36
B41
B46
B51
B56
B61
B66
B71
B76
B81
B86
B91
B96
B101
B106
B111
B116
B121
B126
Deviatonfromrequred
burden(m)
Hole number
Burden deviation
-3
-2.5
-2
-1.5
-1
-0.5
0
0.5
1
B1
B6
B11
B16
B21
B26
B31
B36
B41
B46
B51
B56
B61
B66
B71
B76
B81
B86
B91
B96
B101
B106
B111
B116
B121
B126
Deviationfromrequred
stemming(m)
Hole number
Stemming deviation

34. 24
The graph number 3 shows deviation of actual stemming height from the designed
stemming, the deviation is occurred mostly on the negative side due to the poor accuracy
in charging process as shown in some holes there are no stemming height at all, the
percentage deviation is 0.9132%.
Figure 10: Holes with no or very small stemming height
4.3 Analysis of causes toe formation
After data collection and checking for the deviation the causes of the formation of toes
after blasting was determined by considering the main three factors that contributing to
the formation of toes based on the literature review, those factors include sub drilling,
burden stiffness ratio, vertical energy distribution and relative confinement.
4.3.1 Vertical energy distribution
The required vertical energy distribution at Williamson diamond mine is lying within
80% to 90% when referring to literature review the recommended vertical energy
distribution must be greater than 80% so as energy to be distributed properly and
minimize the possibility of obtaining toes, the data obtained at the site are analysed as
follows to find the vertical energy distribution.

35. 25
Figure 11: Graph number 4 shows VED against average toes height
From the graph number 4 it shows that VED is less than the required by WDL that is
80% which is 74%, the lower the VED results into toes occurrence especially in hard
material such as RVK and BVK, expect in soft rocks that is mudstone there is no
occurrence of toes based on VED since the energy is distributed properly in this kind of
rocks due to the softness of the rock.
4.3.2 Burden stiffness ratio
The appropriate burden stiffness ratio at Williamson diamond mine is 3.0 which provide
the best size of fragmentation and less toes occurrence, the collected data was analysed
on the graph number 5 to show how toes occurred with respect to the burden stiffness
ratio.
RVK RVK RVK BVK
MUDS
TONE
BVK
MUDS
TONE
BVK RVK BVK RVK SHALE
VED 71.428670.588271.6981 75 75 75.2475 75 74.35984.693969.387873.684277.6786
average toes height 0.3 0.6 0.7 0.1 0 0.3 0 0.8 0.1 0.5 0.6 0.2
0
0.1
0.2
0.3
0.4
0.5
0.6
0.7
0.8
0.9
0
10
20
30
40
50
60
70
80
90
Averagetoeheight(m)
VED(%)
Material type
VED against average toe height

36. 26
Figure 12: Graph number 5 BSR against average toe height
From the graph number 5 it is observed that when the BSR is less than 3.0 which is 2.4
the toes formation become very high this full fill the literature review, and when the
BSR is greater than 3 the average toes height become very low this conclude that the
BSR is inversely proportional to the toes occurrence.
4.3.3 Relative confinement ratio
Confinement ratio competed by considering the diameter of the borehole that is 102mm
and explosive material absolute bulk strength (ABS) obtained at Williamson diamond
mine, the ABS is computed by considering the density of explosive from the following
formula
ABS explosive = AWS explosive× DE
Where DE is the density of explosive, the AWS of ANFO is 880cal/cm3
and density of
ANFO is 0.86g/cm3
therefore ABS of ANFO is3168j/cm3
.Graph number 14 shows
computed CR from the measured stemming height
RVK RVK RVK BVK
MUDS
TONE
BVK
MUDS
TONE
BVK RVK BVK RVK SHALE
BSR 1.16667 2.04 2.65 4 1.5 4.04 1.5 1.3 3.92 2.45 1.52 2.8
Average toes height 0.3 0.6 0.7 0.1 0 0.3 0 0.8 0.1 0.5 0.6 0.2
0
0.1
0.2
0.3
0.4
0.5
0.6
0.7
0.8
0.9
0
0.5
1
1.5
2
2.5
3
3.5
4
4.5
Averagetoeheight
BSR
Materal type
BSR against Average toes height

37. 27
Figure 13: Graph number 6 RC against average toes height
Graph number 6 shows that when the RC is less than 1.4 which is 1.2 the occurrence of
toes increase and when the CR is greater than 1.4 the toes occurrence is very low, that’s
for hard rocks and in soft rock the relative confinement is very is suitable enough since
the density of rock is very low therefore the average toes height is 0.
The average actual VED, BSR, and RC is compared with the designed parameters as
shown on graph number 7.
Figure 14: Graph number 7 of actual against designed BSR, WED and CR
RVK RVK RVK BVK
MUDS
TONE
BVK
MUDS
TONE
BVK RVK BVK RVK SHALE
RC 0.839 1.164 1.164 1.814 1.164 1.814 1.164 0.839 1.164 1.164 0.839 1.814
average toe height 0.3 0.6 0.7 0.1 0 0.3 0 0.8 0.1 0.5 0.6 0.2
0
0.5
1
1.5
2
Material type
RC against average toes height
VED BSR RC
ACTUAL 0.7447 2.41 1.245
DESIGNED 0.8 3 1.4
0
0.5
1
1.5
2
2.5
3
3.5
Actual agaist designed VED, BSR and RC
ACTUAL DESIGNED

38. 28
4.3.3 Sub drilling
Sub drill is one of the most importance factor which control the formation of toes but in
Williamson diamond mine there is no consideration of sub drill, due to the minimisation
of charging and drilling cost.
Figure 15: Toes at WDL ready for secondary blasting
4.4 Toe minimisation analysis
The process of toes minimisation require analysis of the VED,RC,BSR, and sub drill to
obtain the best drilling and blasting parameters which will minimize the problem, this
process involving the designed parameters obtained at the drilling and blasting
department at Williamson diamond mine. The VED should be 80%, RC is 1.4 and BSR
is 3.0
From the formulas of BSR, CR and VED this parameters are fixed constant to obtained
measured ground designed parameters as follows. The CR is considered first
1.4 =
(𝑠𝑡𝑒𝑚𝑚𝑖𝑛𝑔 ℎ𝑒𝑖𝑔ℎ𝑡 ×210000)+(𝑑𝑖𝑎𝑚𝑒𝑡𝑒𝑟 𝑜𝑓 𝑏𝑜𝑟𝑒ℎ𝑜𝑙𝑒 ×600)
𝑑𝑖𝑎𝑚𝑒𝑡𝑒𝑟 𝑜𝑓 𝑏𝑜𝑟𝑒ℎ𝑜𝑙𝑒 ×3168
Making diameter of borehole the subject the equation become
Diameter of borehole = 54.75 × stemming length………………………….i
Then the second equation is the VED equation as follows

39. 29
80% =
𝑏𝑜𝑟𝑒ℎ𝑜𝑙𝑒 𝑑𝑒𝑝𝑡ℎ−𝑠𝑡𝑒𝑚𝑚𝑖𝑛𝑔 ℎ𝑒𝑖𝑔ℎ𝑡
𝑏𝑜𝑟𝑒ℎ𝑜𝑙𝑒 𝑑𝑒𝑝𝑡ℎ
×100%
On making stemming height the subject the following equation is obtained
Stemming height = 0.2 × borehole depth…………………………ii
On solving equation (i) and (ii) graphically it gives the following parameters by using
borehole diameter of 102mm
Figure 16: Graph number 8 shows borehole depth and borehole diameter
Graph number 8 gives the best point at which toe less blasting can be obtained this point
gives the borehole depth that is equal to 9.31m, diameter of the borehole equal to
102mm. stemming height can be shown from graph number 9, the stemming height for
toeless blasting is observed to be 1.86m which is approximate equal to 1.9m.
8.6
8.8
9
9.2
9.4
9.6
9.8
10
10.2
94 96 98 100 102 104 106 108 110 112
Boreholedepth(m)
borehole diameter (mm)
Graph of borehole depth and borehole diameter

40. 30
Figure 17; Graph number 9 shows borehole depth against stemming height
Then finally consider the BSR equation as follows
3 =
𝑏𝑜𝑟𝑒ℎ𝑜𝑙𝑒 𝑑𝑒𝑝𝑡ℎ
𝑏𝑢𝑟𝑑𝑒𝑛
On making the burden the subject the equation become
Burden =
𝑏𝑜𝑟𝑒ℎ𝑜𝑙𝑒 𝑑𝑒𝑝𝑡ℎ
3
…………………………iii
Therefore burden and spacing can be obtained from the graph number10 after obtaining
the value of the best borehole diameter from graph number 8.
1.7
1.75
1.8
1.85
1.9
1.95
2
2.05
94 96 98 100 102 104 106 108 110 112
Stemmingheight(m)
Borehole diameter (mm)
Graph of stemming height agaist borehole diameter

41. 31
Figure 18: Graph number 10 shows burden and spacing
After reading the value of borehole depth then finally is to obtain the new burden and
spacing with produce toe less blasting, from graph number 10the value of burden is
3.10m and value of spacing is 3.56m. The borehole diameter should remain the same so
as to minimize cost of buying new drill bit which are more expensive.
4.4.1 Sub drilling
Since Williamson diamond mine does not consider sub drilling then it must be designed
from rule of thumb by considering the rock properties according to literature sub drill is
not applied only when the rock consist of bedding plan but rock at the site does not have
bedding plan and they are medium to hard rock with density less than 2.7g/cm3
and
explosive used is ANFO, from graph number 10 burden is computed to be 3.10m by
using sub drilling rule of thumb (J= 0.3B) then the required sub drilling to be applied
should be 0.93m approximately to 1.0m.
0
0.5
1
1.5
2
2.5
3
3.5
4
4.5
96 98 100 102 104 106 108 110 112
Borehole diameter (mm)
Ghaph shows borehole deph, burden and spacing
burden spacing

42. 32
4.5 Results discussion
Toe problem at Williamson diamond mine is observed due to the poor drilling and
blasting practice and not implementation of the designed parameters because when
considering graph number 1 shows the deviation of the bore hole deviation, graph 2 and
3 which shows stemming and burden deviation respectively, this shows that there is a
very small deviation from the designed data to the actual data and the deviation cannot
result into toes problem.
When considering graph number 4, 5 and 6 which shows how average toes height varies
with VED, BSR and RC respectively its observed that the designed parameters varies
inversely proportional to the toes formation that is when one parameter increase the toe
formation decrease except on the mudstone where no toes observed due to the softness
of the rock. Graph number 7 shows how actual VED, BSR and CR does not meet the
designed parameters this also is the one of the major cause of toes during blasting
The absence of sub-drill result into toes formation to the areas where the VED, RC, BSR
is suitable to the required level therefore to solve this problem sub drill should be
introduced so as to obtained toe less blasting, by using rule of thumb the sub drill
computed is obtain to be 1.0m.
The fixed VED, RC, and BSR is used to determine the new drilling and blasting
parameters which will result into toeless blasting, this can be shown from graph number
8 ,9 and 10. Finally the borehole depth is increased by the sub drill 1.0m to be 10.31m
and this is the best borehole depth which will produce toeless blasting and good
fragmentation.

43. 33
CONCLUSION
Toe present after blasting is due to the poor achievements of the designed VED, BSR
and CR is it shown that BSR is 2.41 while the required is 3.0, VED is 74.47% while the
required is 80% and RC is 1.245 while the required is 1.4m. This deviation is caused by
poor designed of the drilling and blasting parameters. Therefore the best parameters
which will result into toeless blasting should be burden 3.10m, spacing is 3.56m, sub-
drill is 1.0m, borehole depth is 9.31m when sub drill is added it become 10.31m ,
stemming height is 1.86m and diameter of borehole is 102mm.
Should the unavoidable factors contributing to the formation of toes bein non- existent
then a drill block with pre-calculated values of burden, spacing, stemming height,
borehole diameter, borehole depth and with holes being marked properly and drilled up
to the requisite depths, the face being cleared off all the muck from the previous blasts
and without any existing toe can produce a perfectly toe less blasting using suggested
pattern of charging and initiation system.

44. 34
RECOMMENDATION
The following are recommended are made:
 Regular review of the drilling and blasting parameters is recommended so as to
identify the accuracy of the designed parameters on the ground.
 There should be and introduction of sub drill during drilling and blasting so as
to eliminate toes problem.
 Bottom hole initiation system should be considered so as to produce toeless
blasting due to the fact that is minimise premature release of gases through
stemming column.
 More research is recommended based on elimination of toes problem so as to
minimise operation cost.
 Blasting supervisors should conduct surveys during charging so as to ensure all
holes are charged correct and actual stemming height is equal to the designed
made.
 Drilling supervisor should ensure the depth of holes are correct drilled and there
should be no deviation from designed depth so as not to affect charging process.
 Quality assurances and quality control of holes has to be done before blasting

45. 35
References
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48. 37
APPENDICES
Appendix 1: Drilling and blasting data
Designed drilling and blasting data measured at Williamson diamond mine
Parameter Value
Date of blasting 31/3/2016
Block name S-E A
Required elevation 1170m
Bit diameter 102mm
burden 3.5m
Spacing 4.0m
Material type Shalestone
Inclination of borehole 90°
Stemming 2.5m
VED 80%
BSR 3.0
RC 1.4
Sub drilling 0m
Average borehole depth 11.2m
Explosive type ANFO
Actual drilling and blasting data
Borehole
number
Actual
burden
(m)
Actual
spacing
(m)
Actual
stemming
(m)
Actual
hole
depth
(m)
Burden
deviation
(m)
Spacing
deviation
(m)
Stemming
deviation
(m)
Hole
depth
deviation
(m)
B1 3.5 3.8 0.9 10.9 0 -0.2 -1.6 -0.3
B2 3.7 4 2 11.3 0.2 0 -0.5 0.1
B3 3.5 3.9 2.6 10.7 0 -0.1 0.1 -0.5
B4 3.5 4.1 0.6 11.4 0 0.1 -1.9 0.2
B5 3.5 4.3 0.9 10 0 0.3 -1.6 -1.2
B6 3.6 4 1.2 11.3 0.1 0 -1.3 0.1
B7 3.8 4.1 1.9 10.8 0.3 0.1 -0.6 -0.4
B8 3.5 4 1.7 10.4 0 0 -0.8 -0.8

49. 38
B9 3.5 4 1.5 10.9 0 0 -1 -0.3
B10 3.4 4.3 2 10.7 -0.1 0.3 -0.5 -0.5
B11 3.6 4.2 2.1 11.2 0.1 0.2 -0.4 0
B12 3.5 4.5 0 11.3 0 0.5 -2.5 0.1
B13 3.5 4.6 2.3 11.5 0 0.6 -0.2 0.3
B14 3.5 4.2 2.1 10 0 0.2 -0.4 -1.2
B15 3.5 4.1 2.3 9.8 0 0.1 -0.2 -1.4
B16 3.5 4.1 2.5 11.3 0 0.1 0 0.1
B17 3.2 4 2.6 11.9 -0.3 0 0.1 0.7
B18 3.5 4 2.4 9.9 0 0 -0.1 -1.3
B19 3.6 4 2.6 9.7 0.1 0 0.1 -1.5
B20 3.5 4.1 2.5 12 0 0.1 0 0.8
B21 3.5 4 2.3 11.8 0 0 -0.2 0.6
B22 3.5 4.2 2 11.4 0 0.2 -0.5 0.2
B23 3.5 4.2 2.6 9.7 0 0.2 0.1 -1.5
B24 3.5 4.1 2.1 11.2 0 0.1 -0.4 0
B25 3.5 4 2.3 11.3 0 0 -0.2 0.1
B26 3.5 4.3 2.6 11.5 0 0.3 0.1 0.3
B27 3.6 4.1 2.5 11.6 0.1 0.1 0 0.4
B28 3.5 4 2.1 11 0 0 -0.4 -0.2
B29 3.5 3.8 2.3 10.7 0 -0.2 -0.2 -0.5
B30 3.7 3.9 2.1 11.8 0.2 -0.1 -0.4 0.6
B31 3.5 3.6 2.5 9.9 0 -0.4 0 -1.3
B32 3.5 3.9 2.4 10.9 0 -0.1 -0.1 -0.3
B33 3.5 4.3 2.6 11.1 0 0.3 0.1 -0.1
B34 3.5 3.9 2.1 9.8 0 -0.1 -0.4 -1.4
B35 3.4 4 2.4 12 -0.1 0 -0.1 0.8

50. 39
B36 3.5 3.9 2.5 11.9 0 -0.1 0 0.7
B37 3.5 3.9 2.6 11.2 0 -0.1 0.1 0
B38 3.5 4 1.6 11.2 0 0 -0.9 0
B39 3.5 4 0 9.7 0 0 -2.5 -1.5
B40 3.5 3.8 0.5 10.9 0 -0.2 -2 -0.3
B41 3.6 4 0.4 11.3 0.1 0 -2.1 0.1
B42 3.9 4 0.6 11.7 0.4 0 -1.9 0.5
B43 3.6 4.1 2 11.4 0.1 0.1 -0.5 0.2
B44 3.5 4 0.3 10 0 0 -2.2 -1.2
B45 3.6 4 0.9 11.3 0.1 0 -1.6 0.1
B46 3.6 4.1 2.7 10.8 0.1 0.1 0.2 -0.4
B47 3.6 4.2 2 10.4 0.1 0.2 -0.5 -0.8
B48 3.5 4 2.3 10.9 0 0 -0.2 -0.3
B49 3.4 4 2.1 11.7 -0.1 0 -0.4 0.5
B50 3.5 4 0.5 11.2 0 0 -2 0
B51 3.4 3.9 1.2 11.3 -0.1 -0.1 -1.3 0.1
B52 3.4 4 1.4 11.5 -0.1 0 -1.1 0.3
B53 3.5 4 2 10 0 0 -0.5 -1.2
B54 3.4 4 0.6 9.8 -0.1 0 -1.9 -1.4
B55 3.4 4.1 1.5 11.3 -0.1 0.1 -1 0.1
B56 3.8 4.1 2.6 11.9 0.3 0.1 0.1 0.7
B57 3.5 4 0 10.9 0 0 -2.5 -0.3
B58 3.6 4 2.9 9.7 0.1 0 0.4 -1.5
B59 3.5 4 2.7 12 0 0 0.2 0.8
B60 3.5 4 1.3 11.8 0 0 -1.2 0.6
B61 3.5 4 1.8 11.4 0 0 -0.7 0.2
B62 3.5 3.9 1.7 9.7 0 -0.1 -0.8 -1.5

51. 40
B63 3.5 3.8 1.9 12.2 0 -0.2 -0.6 1
B64 3.5 3.6 0 11.3 0 -0.4 -2.5 0.1
B65 3.6 3.9 1.2 11.5 0.1 -0.1 -1.3 0.3
B66 3.5 3.8 1 11.6 0 -0.2 -1.5 0.4
B67 3.5 3.6 1.3 11 0 -0.4 -1.2 -0.2
B68 3.4 3.7 1.9 10.7 -0.1 -0.3 -0.6 -0.5
B69 3.5 3.7 2 11.8 0 -0.3 -0.5 0.6
B70 3.2 3.8 0 9.9 -0.3 -0.2 -2.5 -1.3
B71 3.5 3.9 2.3 10.9 0 -0.1 -0.2 -0.3
B72 3.5 3.8 2.4 11.1 0 -0.2 -0.1 -0.1
B73 3.5 3.6 2.1 9.8 0 -0.4 -0.4 -1.4
B74 3.5 3.7 2.5 12 0 -0.3 0 0.8
B75 3.2 3.8 2.6 11.9 -0.3 -0.2 0.1 0.7
B76 3.5 3.9 2 11.2 0 -0.1 -0.5 0
B77 3.5 3.8 1 11.1 0 -0.2 -1.5 -0.1
B78 3.5 3.9 1.6 9.7 0 -0.1 -0.9 -1.5
B79 3.5 4.1 1.5 11.7 0 0.1 -1 0.5
B80 3.5 4.5 0 11.2 0 0.5 -2.5 0
B81 3.5 3.8 0.2 11.3 0 -0.2 -2.3 0.1
B82 3.5 3.7 0.6 11.5 0 -0.3 -1.9 0.3
B83 3.5 3.5 0.5 10 0 -0.5 -2 -1.2
B84 3.4 4.1 0.7 9.8 -0.1 0.1 -1.8 -1.4
B85 3.6 4.1 2 11.3 0.1 0.1 -0.5 0.1
B86 3.6 4.1 1.6 11.9 0.1 0.1 -0.9 0.7
B87 3.4 4.2 1.8 10.9 -0.1 0.2 -0.7 -0.3
B88 3.6 4 0.8 9.7 0.1 0 -1.7 -1.5
B89 3.5 3.9 0 12 0 -0.1 -2.5 0.8

52. 41
B90 3.6 4.1 2 11.8 0.1 0.1 -0.5 0.6
B91 3.5 3.8 0.9 11.4 0 -0.2 -1.6 0.2
B92 3.4 4.1 2.7 9.7 -0.1 0.1 0.2 -1.5
B93 3.6 4 2.8 12.2 0.1 0 0.3 1
B94 3.8 4 2.3 11.3 0.3 0 -0.2 0.1
B95 3.9 4 2.1 11.5 0.4 0 -0.4 0.3
B96 3.5 4 0 11.6 0 0 -2.5 0.4
B97 3.6 3.9 2.6 11 0.1 -0.1 0.1 -0.2
B98 3.4 3.9 2.1 10.7 -0.1 -0.1 -0.4 -0.5
B99 3.5 3.9 2.5 11.8 0 -0.1 0 0.6
B100 3.5 4 1.3 9.9 0 0 -1.2 -1.3
B101 3.5 4 1.2 10.9 0 0 -1.3 -0.3
B102 3.5 4 0.6 11.1 0 0 -1.9 -0.1
B103 3.5 4 0.4 9.8 0 0 -2.1 -1.4
B104 3.5 4 0.5 12 0 0 -2 0.8
B105 3.5 4 2.5 11.9 0 0 0 0.7
B106 3.5 3.8 0.3 11.2 0 -0.2 -2.2 0
B107 3.5 3.9 2.6 11.1 0 -0.1 0.1 -0.1
B108 3.6 3.7 1.4 9.7 0.1 -0.3 -1.1 -1.5
B109 3.8 3.9 1.8 11.7 0.3 -0.1 -0.7 0.5
B110 3.5 3.8 1.7 11.2 0 -0.2 -0.8 0
B111 3.5 3.9 0.7 11.3 0 -0.1 -1.8 0.1
B112 3.5 4.2 0.9 11.5 0 0.2 -1.6 0.3
B113 3.5 4 2.3 10 0 0 -0.2 -1.2
B114 3.5 4 2.1 9.8 0 0 -0.4 -1.4
B115 3.5 3.9 2.6 11.3 0 -0.1 0.1 0.1
B116 3.5 3.8 0.1 11.9 0 -0.2 -2.4 0.7

53. 42
B117 3.5 3.9 0.7 10.9 0 -0.1 -1.8 -0.3
B118 3.5 3.9 0 9.7 0 -0.1 -2.5 -1.5
B119 3.5 3.7 2.6 8.8 0 -0.3 0.1 -2.4
B120 3.5 3.9 2.8 11.8 0 -0.1 0.3 0.6
B121 3.5 3.8 0 11.4 0 -0.2 -2.5 0.2
B122 3.5 4.3 2.4 9.7 0 0.3 -0.1 -1.5
B123 3.6 4.1 0.3 12.2 0.1 0.1 -2.2 1
B124 3.5 4.3 2.7 11.3 0 0.3 0.2 0.1
B125 3.6 3.7 0.1 12.5 0.1 -0.3 -2.4 1.3
B126 3.5 3.8 0.4 11.6 0 -0.2 -2.1 0.4
B127 3.5 4.2 0 11 0 0.2 -2.5 -0.2
Average 3.518110 3.977165 1.596850 11.01417 0.018110 -0.02283 -0.903149 -0.18582
Appendix 2: Rock density data
Material type Density (g/cm3
) content
RVK 2.03 40% granite
BVK 2.48 70% granite
Mudstone 1.92 Very fine grained laminated
kimberlite
Shale 2.05 Clay and silt sized sediments
Bouma 1.97 Intermediate between RVK
and shale
Appendix 3: Explosive properties data
Explosive type RBS Density (g/cm3
)
ANFO 100 0.86

54. 43
Appendix 4: Historical data
Date Designed
Burden
(m)
Designed
Spacing
(m)
Borehole
depth
(m)
Rock type Stemming
(m)
Average
toe
height
(m)
Block Number
of holes
5/1/2016 3 3
3.5 RVK
1 0.3 N-W
C 198
7/1/2016 2.5 2.5
5.1 RVK
1.5 0.4 N-W
C 205
19/1/2016 2 2
5.3 RVK
1.5 0.3 S-W
C 138
22/1/2016 2.5 2.5
10 BVK
2.5 0.7 N-W
C 312
29/1/2016 3.5 4
6 MUDSTONE
1.5 0 S-W
C 131
4/2/2016 2.5 2.5
10.1 BVK
2.5 0.7 S-E
A 156
9/2/2016 3.5 4
6 MUDSTONE
1.5 0 SW-
C 231
16/2/2016 3 3
3.9 BVK
1 0.5 SW-
C 318
25/2/2016 2.5 2.5
9.8 RVK
1.5 0.4 NW-
B 437
3/3/2016 2.0 2.0 4.9 BVK 1.5 0.5 NW-
C,D 281
23/3/2016 2.5 2.5 3.8 RVK 1 0.6 NE-D 101
31/3/2016 3.5 4.0 11.2 SHALE 2.5 0.2 SE-A 127